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EPA
United States
Environmental Protection
Agency
Effluent Guidelines Division
WH-552
Washington DC 20460
EPA 440/1-82/061-b
May 1982
Water and Waste Management
Development
Document for
Effluent Limitations
Guidelines and
Standards for the
Proposed
EPA REGION VII IRC
077400
Ore Mining and Dressing
Point Source Category
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AUG 2 6 1982
DEVELOPMENT DOCUMENT
FOR PROPOSED
EFFLUENT LIMITATIONS GUIDELINES AND
NEW SOURCE PERFORMANCE STANDARDS
FOR THE
ORE MINING AND DRESSING
POINT SOURCE CATEGORY
Anne E. Gorsuch
Admi ni strator
Frederic A. Eidsness, Jr.
Assistant Administrator for Water
Stephen Schatzow
Di rector
Water Regulations and Standards
Jeffery Denit
Acting Director, Effluent Guidelines Division
B. Matthew Jarrett
Project Officer
May 1982
Effluent Guidelines Division
Office of Water
U.S. Environmental Protection Agency
Washington, D.C. 20460
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TABLE OF CONTENTS
Section Page
I EXECUTIVE SUMMARY 1
Best Available Technology Economically
Achievable (BAT) , 3
New Source Performance Standards.... .. 5
BCT Effluent Limitations........ 6
II INTRODUCTION.. 7
PURPOSE 7
LEGAL AUTHORITY 7
III INDUSTRY PROFILE, 17
ORE BENEFICIATION PROCESSES 17
ORE MINING METHODS.... .... ... 26
INDUSTRY PRACTICE 34
IV INDUSTRY SUBCATEGORIZATON Ill
FACTORS INFLUENCING SELECTION OF SUBCATEGORIES. . Ill
SUBCATEGORIZATION 115
COMPLEXES. 117
V SAMPLING AND ANALYSIS METHODS 119
SITE SELECTION.... 119
SAMPLE COLLECTION, PRESERVATION, AND
TRANSPORTATION. 125
SAMPLE ANALYSIS 131
VI WASTE CHARACTERIZATION 155
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TABLE OF CONTENTS (Continued)
Section Page
SAMPLING PROGRAM RESULTS 155
REAGENT USE IN FLOATION MILLS 161
SPECIAL PROBLEM AREAS 164
VII SELECTION OF POLLUTANT PARAMETERS 195
DATA BASE 196
SELECTED TOXIC PARAMETERS 196
EXCLUSION OF TOXIC POLLUTANTS THROUGHOUT
THE ENTIRE CATEGORY 196
EXCLUSION OF TOXIC POLLUTANTS BY SUBDIVISION
AND MILL PROCESS 201
CONVENTIONAL POLLUTANT PARAMETERS 202
NON-CONVENTIONAL POLLUTANT PARAMETERS 202
CONVENTIONAL AND NON-CONVENTIONAL PARAMETERS
SELECTED 203
SURROGATE/INDICATOR RELATIONSHIPS 203
VIII CONTROL AND TREATMENT TECHNOLOGY 215
IN-PROCESS CONTROL TECHNOLOGY 215
END-OF-PIPE TREATMENT TECHNIQUES 226
PILOT AND BENCH-SCALE TREATMENT STUDIES 255
HISTORICAL DATA SUMMARY 276
CONTROL AND TREATMENT PRACTICES 296
IX COST, ENERGY, AND NON-WATER QUALITY ASPECTS 401
DEVELOPMENT OF COST DATA BASE 401
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TABLE OF CONTENTS (Continued)
Section Page
CAPITAL COST 401
ANNUAL COST 403
TREATMENT PROCESS COSTS 404
MODULAR TREATMENT COSTS FOR THE ORE MINING
AND DRESSING INDUSTRY 414
NON-WATER QUALITY ISSUES 414
X BEST AVAILABLE TECHNOLOGY ECONOMICALLY
AVAILABLE 497
SUMMARY OF BEST AVAILABLE TECHNOLOGY 498
GENERAL PROVISIONS 500
BAT OPTIONS CONSIDERED FOR TOXICS REDUCTION 505
SELECTION AND DECISION CRITERIA 509
ADDITIONAL PARAGRAPH 8 EXCLUSIONS 520
XI BEST CONVENTIONAL POLLUTANT CONTROL TECHNOLOGY.. 523
XII NEW SOURCE PERFORMANCE STANDARDS (NSPS) 525
GENERAL PROVISIONS 525
SPS OPTIONS CONSIDERED 526
NSPS SELECTION AND DECISION CRITERIA 526
XIII PRETREATMENT STANDARDS 529
XIV ACKNOWLEDGEMENTS 531
XV REFERENCES 533
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TABLE OF CONTENTS (Continued)
Sect|on JLiiLl
SECTION III 533
SECTION V 534
SECTION VI.... 535
SECTION VII 535
SECTION VIII 536
SECTION IX 542
SECTION X 543
XVI GLOSSARY 545
APPENDIX A. , 564
APPENDIX B 608
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LIST OF TABLES
Number
III-l
III-2
III-3
III-4
III-5
III-6
III-7
III-8
III-9
111-10
III-ll
111-12
111-13
111-14
111-15
II 1-16
111-17
T T - 1 8
1 1 1 - 1 9
111-20
111-21
1
PROFILE OF IRON MINES
PROFILE OF IRON MILLS
PROFILE OF COPPER MINES
PROFILE OF COPPER MILLS
PROFILE OF LEAD/ZINC MINES
PROFILE OF LEAD/ZINC MILLS
PROFILE OF MISCELLANEOUS LEAD/ZINC MINES
PROFILE OF MISCELLANEOUS LEAD/ZINC MILLS
PROFILE OF GOLD MINES
PROFILE OF GOLD MILLS
PROFILE OF MISCELLANEOUS GOLD AND SILVER MINES...
PROFILE OF MISCELLANEOUS GOLD AND SILVER MILLS...
PROFILE OF SILVER MINES
PROFILE OF SILVER MILLS
PROFILE OF MOLYBDENUM MINES
PROFILE OF MOLYBDENUM MILLS
PROFILE OF ALUMINUM ORE MINES
PROFILE OF TUNGSTEN MINES (PRODUCTION GREATER
THAN 5000 MT ORE/YEAR)
PROFILE OF TUNGSTEN MINES (PRODUCTION LESS
THAN 5000 MT ORE/YEAR)
PROFILE OF TUNGSTEN MILLS (PRODUCTION LESS
THAN 5000 MT/YEAR)
PROFILE OF TUNGSTEN MILLS (PRODUCTION GREATER
THAN 5000 MT/YEAR)
3age
55
58
61
67
71
76
79
80
81
82
84
86
87
88
89
90
91
92
93
95
97
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LIST OF TABLES (Continued)
Number Page
111-22 PROFILE OF MERCURY MINES 98
111-23 PROFILE OF MERCURY MILLS 99
111-24 PROFILE OF URANIUM MINES 100
111-25 PROFILE OF URANIUM MILLS 102
111-26 PROFILE OF URANIUM (IN-SITU LEACH) MINES 104
111-27 PROFILE OF ANTIMONY SUBCATEGORY 106
111-28 PROFILE OF TITANIUM MINES 107
111-29 PROFILE OF TITANIUM DREDGE MILLS 108
111-30 PROFILE OF NICKEL SUBCATEGORY 109
111-31 PROFILE OF VANADIUM SUBCATEGORY 110
IV-1 PROPOSED SUBCATEGORIZATION FOR BAT - ORE
MINING AND DRESSING 118
V-l TOXIC ORGANICS 142
V-2 TOXIC METALS, CYANIDE AND ASBESTOS 147
V-3 POLLUTANTS ANALYZED AND ANALYSIS TECHNIQUES/
LABORATORI ES 148
V-4 LIMITS OF DETECTION FOR POLLUTANTS ANALYZED 149
V-5 LIMITS OF DETECTION FOR POLLUTANTS ANALYZED
FOR COST - SITE VISITS BY RADIAN CORPORATION 150
V-6 COMPARISON OF SPLIT SAMPLE ANALYSES* FOR
CYANIDE BY TWO DIFFERENT LABORATORIES USING
THE BELACK DISTILLATION/PYRIDINE-PYROZOLANE
METHOD 151
V-7 ANALTYICAL QUALITY CONTROL PERFORMANCE OF
COMMERCIAL LABORATORY PERFORMING CYANIDE*
ANALYSES BY EPA APPROVED BELACK DISTILLATION
METHOD 152
viii
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LIST OF TABLES (Continued)
Number Page
V-8 SUMMARY OF CYANIDE ANALYSIS DATA FOR SAMPLES
OF ORE MINING AND PROCESSING WASTEWATERS 153
VI-1 DATA SUMMARY ORE MINING DATA ALL SUBCATEGORIES... 166
VI-2 DATA SUMMARY ORE MINING DATA SUBCATEGORY IRON
SUBDIVISION MINE MILL PROCESS MINE DRAINAGE 172
VI-3 DATA SUMMARY ORE MINING DATA SUBCATEGORY IRON
SUBDIVISON MILL MILL PROCESS PHYSICAL AND/OR
CHEMICAL 173
VI-4 DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MINE MILL PROCESS
MINE DRAINAGE 174
VI-5 DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MILL MILL PROCESS
CYAN ID AT I ON 175
VI-6 DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MILL MILL PROCESS
FLOTATION (FROTH) 176
VI-7 DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MINE/MILL MILL
PROCESS HEAP/VAT/DUMP LEACHING 177
VI-8 DATA SUMMARY ORE MINING DATA SUBCATEGORY
ALUMINUM SUBDIVISION MINE MILL PROCESS
MINE DRAINAGE 178
VI-9 DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM
MOLYBDENUM SUBDIVISION MINE/MILL MILL
PROCESS GRAVITY SEPARATION 179
VI-10 DATA SUMMARY ORE MINING DATA SUBCATEGORY
TUNGSTEN SUBDIVISION MILL 180
VI-11 DATA SUMMARY ORE MINING DATA SUBCATEGORY
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LIST OF TABLES (Continued)
Number Page
MERCURY SUBDIVISION MILL MILL PROCESS
FLOTATION (FROTH) 181
VI-12 DATA SUMMARY ORE MINING DATA SUBCATEGORY
URANIUM SUBDIVISION MINE MILL PROCESS MINE
DRAINAGE 182
VI-13 DATA SUMMARY ORE MINING DATA SUBCATEGORY
URANIUM SUBDIVISION MILL MILL PROCESS AND
LOCATIONS 183
VI-14 DATA SUMMARY ORE MINING DATA SUBCATEGORY
TITANIUM SUBDIVISION MINE MILL PROCESS
MINE DRAINAGE 184
VI-15 DATA SUMMARY ORE MINING DATA SUBCATEGORY
TITANIUM SUBDIVISION MILLS WITH DREDGE
MINING MILL PROCESS PHYSICAL AND/OR
CHEMICAL 185
VI-16 DATA SUMMARY ORE MINING DATA SUBCATEGORY
VANADIUM SUBDIVISION MINE MILL PROCESS
NO MILL PROCESS 186
VI-17 DATA SUMMARY ORE MINING DATA SUBCATEGORY
VANADIUM SUBDIVISION MILL MILL PROCESS
FLOTATION (FROTH) 187
VI-18 SUMMARY OF REAGENT USE IN ORE FLOATION MILLS 188
VII-1 DATA SUMMARY ORE MINING DATA ALL SUBCATEGORIES... 206
VII-2 POLLUTANTS CONSIDERED FOR REGULATION 212
VII-3 PRIORITY METALS EXCLUSION BY SUBCATEGORY,
SUBDIVISION, MILL PROCESS 213
VII-4 TUBING LEACHING ANALYSIS RESULTS 214
VIII-1 ALTERNATIVES TO SODIUM CYANIDE FOR FLOATION
CONTROL 306
DESTRUCTION BY OZONATION AT MILL 6102 307
VIII-3 RESULTS OF LABORATORY TESTS AT MILL 6102
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LIST OF TABLES (Continued)
Number Pag^e
DEMONSTRATING EFFECTS OF RESIDENCE TIME, pH,
AND SODIUM HYPOCHLORITE CONCENTRATIONS ON
CYANIDE DESTRUCTION WITH SODIUM HYPOCHLORITE 308
VIII-4 EFFECTIVENESS OF WASTEWATER-TREATMENT
ALTERNATIVES FOR REMOVAL OF CHRYSOTILE AT
PILOT PLANTS 309
VIII-5 EFFECTIVENESS OF WASTEWATER-TREATMENT
ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS
AT ASBESTOS-CEMENT PROCESSING PLANT.. 310
VIII-6 EFFECTIVENESS OF WASTEWATER-TREATMENT
ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS
AT ASBESTOS, QUEBEC, ASBESTOS MINE 311
VIII-7 EFFECTIVENESS OF WASTEWATER-TREATMENT
ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS
AT BAIE VERTE, NEWFOUNDLAND ASBESTOS MINE 312
VIII-8 COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS
FOR TOTAL FIBERS AND CHRYSOTILE AT SEVERAL
FACILITIES SURVEYED 313
VIII-9 EFFLUENT QUALITY ATTAINED BY USE OF BARIUM
SALTS FOR REMOVAL OF RADIUM FROM WASTEWATER
AT VARIOUS URANIUM MINE AND MILL FACILITIES .. 316
VIII-10 RESULTS OF MINE WATER TREATMENT BY LIME
ADDITION AT COPPER MINE 2120 317
VIII-11 RESULTS OF COMBINED MINE AND MILL WASTEWATER
TREATMENT BY LIME ADDITION AT COPPER MINE/
MILL 2120 318
VIII-12 RESULTS OF COMBINED MINE WATER + BARREN LEACH
SOLUTIONS TREATMENT BY LIME ADDITION AT COPPER
MINE/MILL 2120 319
VIII-13 RESULTS OF COMBINED MINE WATER + BARREN LEACH
SOLUTIONS MILL TAILINGS TREATMENT BY LIME
ADDITION AT COPPER MINE/MILL 2120 320
VIII-14 CHARACTERISTICS OF RAW MINE DRAINAGE TREATED
DURING PILOT-SCALE EXPERIMENTS IN NEW
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LIST OF TABLES (Continued)
Number Page
BRUNSWICK, CANADA 321
VIII-15 EFFLUENT QUALITY ATTAINED DURING PILOT-SCALE
MINE-WATER TREATMENT STUDY IN NEW BRUNSWICK,
CANADA 322
VIII-16 RESULTS OF MINE-WATER TREATMENT BY LIME
ADDITION AT GOLD MINE 4102 323
VIII-17 RESULTS OF LABORATORY-SCALE MINE-WATER
TREATMENT STUDY AT LEAD/ZINC MINE 3113 324
VIII-18 EPA-SPONSORED WASTEWATER TREATABILITY STUDIES
CONDUCTED BY CALSPAN AT VARIOUS SITES IN ORE
MINING AND DRESSING INDUSTRY 325
VIII-19 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
ZINC MINE/MILL 3121 326
VIII-20 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
ZINC MINE/MILL 3121 DURING PERIOD OF MARCH
19-29, 1979 327
VIII-21 OBSERVED VARIATION WITH TIME OF pH AND COPPER
AND ZINC CONCENTRATIONS OF TAILING-POND DECANT
AT LEAD/ZINC MINE/MILL 3121 328
VIII-22 SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED
WITH PILOT-SCALE UNIT TREATMENT PROCESS AT
MINE/MILL 3121 DURING AUGUST STUDY 329
VIII-23 SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED
WITH PILOT-SCALE UNIT TREATMENT PROCESS AT
MINE/MILL 3121 DURING MARCH STUDY 330
VIII-24 RESULTS OF OZONATION FOR DESTRUCTION OF
CYANIDE 331
VIII-25 EFFLUENT FROM LEAD/ZINC MINE/MILL/SMELTER/
REFINERY 3107 PHYSICAL/CHEMICAL-TREATMENT
PLANT 332
VIII-26 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
XI
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LIST OF TABLES (Continued)
Number P a g e
TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
ZINC MINE/MILL/SMELTER/REFINERY 3107 333
VIII-27 SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED
WITH PILOT-SCALE UNIT TREATMENT PROCESSES AT
MINE/MILL/SMELTER/REFINERY 3107 334
VIII-28 CHARACTER OF MINE DRAINAGE FROM LEAD/ZINC
MINE 3113 335
VIII-29 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
ZINC MINE 3113 336
VIII-30 SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
AT MINE 3113 337
VIII-31 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT ALUMINUM
MINE 5102 338
VIII-32 SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
AT MINE 5102 339
VIII-33 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT MILL
9402 (ACID LEACH MILL WASTEWATER) 340
VIII-34 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT MILL
9402 (ACID LEACH MILL WASTEWATER) 341
VIII-35 SUMMARY OF A WASTEWATER TREATABILITY RESULTS
USING A LIME ADDITION/BARIUM CHLORIDE
ADDITION/SETTLE PILOT-SCALE TREATMENT SCHEME 342
VIII-36 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT MILL
9401 343
VIII-37 SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED
WITH PILOT-SCALE TREATMENT SYSTEM AT MILL
9401 344
VIII-38 RESULTS OF BENCH-SCALE AC ID/ALKALINE MILL
XI
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LIST OF TABLES (Continued)
Number Pa^ge
WASTEWATER NEUTRALIZATION 345
VIII-39 CHARACTERIZATION OF INFLUENT TO WASTEWATER
TREATMENT PLANT AT COPPER MINE/MILL/SMELTER/
REFINERY 2122 (SEPTEMBER 5-7 1979) 346
VIII-40 CHARACTERIZATION OF EFFLUENT FROM WASTEWATER
TREATMENT PLANT (INFLUENT TO PILOT-SCALE
TREATMENT PLANT) AT COPPER MINE/MILL/SMELTER/
REFINERY 2122 (SEPTEMBER 5-10, 1979) 347
VIII-41 SUMMARY OF TREATED EFFLUENT QUALITY FROM
PILOT-SCALE TREATMENT PROCESSES AT COPPER
MINE/MILL/SMELTER/REFINERY 2122
(SEPTEMBER 5-10, 1979) ., 348
VIII-42 CHARACTERIZATION OF EFFLUENT FROM WASTEWATER
TREATMENT PLANT (INFLUENT TO PILOT-SCALE
TREATMENT PLANT) AT COPPER MINE/MILL/SMELTER/
REFINERY 2121 (SEPTEMBER 18-19, 1979) 350
VIII-43 SUMMARY OF TREATED EFFLUENT QUALITY FROM
PILOT-SCALE TREATMENT PROCESSES AT COPPER
MINE/MILL/SMELTER/REFINERY 2121
(SEPTEMBER 18-19, 1979) 351
VIII-44 HISTORICAL DATA SUMMARY FOR IRON ORE MINE/
MILL 1808 (FINAL DISCHARGE) 352
VIII-45 HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL
2121 (FINAL TAILING-POND DISCHARGE: TREATMENT
OF MINE PLUS MILL WATER) 353
VIII-46 HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL
2120 (TAILING-POND OVERFLOW: TREATMENT OF
UNDERGROUND MINE, MILL, AND LEACH-CIRCUIT
WASTEWATER STREAMS) 354
VIII-47 HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL
2120 (TREATMENT SYSTEM-BARREL POND-EFFLUENT:
TREATMENT OF MILL WATER AND OPEN-PIT MINE
WATER) 355
VIII-48 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3105 (TREATED MINE EFFLUENT). 356
xi v
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LIST OF TABLES (Continued)
Number Page
VIII-49 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE
3130 (UNTREATED MINEWATER) 357
VIII-50 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE
3130 (TREATED EFFLUENT) 358
VIII-51 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3101 (TAILING-POND DECANT TO POLISHING
PONDS) 359
VIII-52 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3101 (FINAL DISCHARGE FROM POLISHING
POND) 360
VIII-53 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3102 (FINAL DISCHARGE) 361
VIII-54 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3103 (TAILING-POND EFFLUENT TO SECOND
SETTLING POND) 362
VIII-55 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3103 (EFFLUENT FROM SECOND SETTLING
POND) 363
VIII-56 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
MILL 3104 (TAILING-LAGOON OVERFLOW)
(EFFLUENT) 364
VIII-57 HISTORICAL DATA SUMMARY FOR LEAD/ZINC MILL
3110 (TAILING-LAGOON OVERFLOW) 365
VIII-58 HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE
6103 (TREATED EFFLUENT) 366
VIII-59 HISTORICAL DATA SUMMARY FOR MOLYBDENUM MILL
6101 (TREATED EFFLUENT) 367
VIII-60 HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/
MILL 6102 (CLEAR POND BLEED STREAM-INFLUENT
TO TREATMENT SYSTEM) 368
VIII-61 HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/
MILL 6102 (FINAL DISCHARGE FROM RETENTION
x v
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LIST OF TABLES (Continued)
Number Page
POND-TREATED EFFLUENT) 369
VIII-62 HISTORICAL DATA SUMMARY FOR BAUXITE MINE 5102
JANUARY 1979 - DECEMBER 1980, DISCHARGE 008 370
VIII-63 HISTORICAL DATA SUMMARY FOR BAUXITE MINE 5102,
FEBRUARY 1979 - DECEMBER 1980, DISCHARGE 009 371
VIII-64 HISTORICAL DATA SUMMARY FOR BAUXITE MINE 5102,
FEBRUARY 1979 - DECEMBER 1980, DISCHARGE 010 372
VIII-65 HISTORICAL DATA SUMMARY FOR BAUXITE MINE 5101,
JUNE 1978 - DECEMBER 1980, DISCHARGE 001 373
VIII-66 HISTORICAL DATA SUMAMRY FOR BAUXITE MINE 5101,
JANUARY 1978 - DECEMBER 1980, DISCHARGE 007 374
VIII-67 HISTORICAL DATA SUMMARY FOR BAUXITE MINE 5101,
JANUARY - SEPTEMBER 1980, DISCHARGE 009 375
VIII-68 HISTORICAL DATA SUMMARY FOR TUNGSTEN MINE
6104 (TREATED EFFLUENT) 376
VIII-69 HISTORICAL DATA SUMMARY FOR URANIUM MINE
7708 (TREATED MINE WATER) 377
VIII-70 HISTORICAL DATA SUMMARY FOR NICKEL MINE/MILL
6106 (TREATED EFFLUENT) 378
VIII-71 HISTORICAL DATA SUMMARY FOR VANADIUM MINE 6107,
JULY 1978 - DECEMBER 1980, DISCHARGE 005 379
VIII-72 HISTORICAL DATA SUMMARY FOR TITANIUM MINE/MILL
9906, OCTOBER 1975 - DECEMBER 1979 380
VIII-73 CHARACTERIZATION OF RAW WASTEWATER (TAILING-
POND EFFLUENT AS INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT COPPER MILL 2122
DURING PERIOD OF 6-14 SEPTEMBER 1978 381
VIII-74 CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
TO PILOT-SCALE TREATMENT TRAILER) AT BASE
AND PRECIOUS METALS MILL 2122 DURING PERIOD
8-19 JANUARY 1979 382
XVT
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LIST OF TABLES (Continued)
Number Pa g e
VIII-75 SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
PERFORMED AT MILL 2122 DURING PERIOD OF
6-14 SEPTEMBER 1978 383
VIII-76 PERFORMANCE OF A DUAL MEDIA FILTER WITH TIME-
FILTRATION OF TAILING POND DECANT AT MILL
2122 384
VIII-77 RESULTS OF CHLORINATION BUCKET TESTS FOR
DESTRUCTION PHENOL AND CYANIDE IN MILL 2122
TAILING POND DECANT 385
VIII-78 RESULTS OF PILOT-SCALE OZONATION FOR DESTRUCTION
OF PHENOL AND CYANIDE IN MILL 2122 TAILING POND
DECANT 386
VIII-79 EFFLUENT QUALITY ATTAINED AT SEVERAL PLACER
MINING OPERATIONS EMPLOYING SETTLING-POND
TECHNOLOGY 387
IX-1 COST COMPARISONS GENERATED ACCORDING TO
TREATMENT PROCESS AND ORE CATEGORY 416
IX-2 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (IRON ORE) 417
IX-3 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (COPPER ORE) 426
IX-4 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (LEAD-ZINC ORE) 433
IX-5 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (GOLD-SILVER) 441
IX-6 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (ALUMINUM ORE) 445
IX-7 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
xvi
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LIST OF TABLES (Continued)
Number Pajje
ORE MINED (FERROALLOY ORE) 446
IX-8 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (MERCURY ORE) 454
IX-9 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (TITANIUM ORE) 455
IX-10 COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
ORE MINED (URANIUM ORE) 456
IX-11 SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID
LEACHATE TEST 461
XVI t 7
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LIST OF FIGURES
Number Page
VIII-1 POTABLE WATER TREATMENT FOR ASBESTOS REMOVAL
AT LAKEWOOD PLANT, DULUTH, MINNESOTA 388
VIII-2 EXPERIMENTAL MINE-DRAINAGE TREATMENT SYSTEM
FOR UNIVERSITY OF DENVER STUDY 389
VIII-3 CALSPAN MOBILE ENVIRONMENTAL TREATMENT PLANT
CONFIGURATIONS EMPLOYED AT BASE AND PRECIOUS
METAL MINE AND MILL OPERATIONS 390
VIII-4 MOBILE PILOT TREATMENT SYSTEM CONFIGURATION
EMPLOYED AT URANIUM MILL 9402 391
VIII-5 PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED
AT URANIUM MILL 9401 392
VIII-6 MODE OF OPERATION OF SEDIMENTATION TANK DURING
PILOT-SCALE TREATABILITY STUDY AT MINE/MILL
9402 393
VIII-7 FRONTIER TECHNICAL ASSOCIATES MOBILE TREATMENT
PLANT CONFIGURATION EMPLOYED AT MINE/MILL/
SMELTER/REFINERIES #2121 and #2122 394
VIII-8 DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT
AT COPPER MINE/MILL 2120 395
VIII-9 DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT
AT COPPER MINE/MILL 2120 396
VIII-10 SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT
FACILITIES AT LEAD/ZINC MINE/MILL 3103 397
VIII-11 PLOTS OF SELECTED PARAMETERS VERSUS TIME AT
NICKEL MINE/MILL 6106 (1972-1974) 398
VIII-12 PLOT OF TSS CONCENTRATIONS VERSUS COPPER
CONCENTRATIONS IN TAILING-POND DECANT AT
MINE/MILL/SMELTER/REFINERY 2122 399
IX-1 SECONDARY SETTLING POND/LAGOON - TYPICAL LAYOUT.. 466
IX-2 ORE MINING WASTEWATER TREATMENT SECONDARY
SETTLING POND/LAGOON COST CURVES 467
IX-3 ORE MINE WASTEWATER TREATMENT SETTLING PONDS -
LINING COST CURVES 468
xi x
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LIST OF FIGURES (Continued)
Number Page
IX-4 FLOCCULANT (POLYELECTROLYTE) PREPARATION AND
FEED-FLOW SCHEMATIC 469
IX-5 ORE MINE WASTEWATER TREATMENT FLOCCULANT
(POLYELECTROLYTE) PREPARATION & FEED SYSTEM
COST CURVES 470
IX-6 OZONE GENERATION AND FEED FLOW SCHEMATIC 471
IX-7 ORE MINE WASTEWATER TREATMENT OZONE GENERATION
& FEED SYSTEM COST CURVES 472
IX-8 ALKALINE-CHLORINATION FLOW SCHEMATIC 473
IX-9 ORE MINE WASTEWATER TREATMENT ALKALINE
CHLORINATION COST CURVES.. 474
IX-10 ION EXCHANGE FLOW SCHEMATIC 475
IX-11 ORE MINE WASTEWATER TREATMENT ION EXCHANGE
COST CURVES 476
IX-12 GRANULAR MEDIA FILTRATION PROCESS FLOW
SCHEMATIC 477
IX-13 ORE MINE WASTEWATER TREATMENT GRANULAR MEDIA
FILTRATION PROCESS COST CURVES 478
IX-14 pH ADJUSTMENT FLOW SCHEMATIC 479
IX-15 ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT
CAPITAL COST CURVES 480
IX-16 ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT
ANNUAL COST CURVES 481
IX-17 WASTEWATER RECYCLE FLOW SCHEMATIC 482
IX-18 ORE MINE WASTEWATER TREATMENT RECYCLING
CAPITAL COST CURVES 483
IX-19 ORE MINE WASTEWATER TREATMENT RECYCLING
ANNUAL COST CURVES 484
xx
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LIST OF FIGURES (Continued)
Number Page
IX-20 ACTIVATED CARBON ADSORPTION FLOW SCHEMATIC 485
IX-21 ACTIVATED CARBON ADSORPTION CAPITAL COST
CURVE FOR PHENOL REDUCTION IN BPT EFFLUENT
DISCHARGED FROM BASE AND PRECIOUS METAL ORE
MILLS 486
IX-22 ACTIVATED CARBON ADSORPTION ANNUAL COST CURVE
FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
FROM BASE AND PRECIOUS METAL ORE MILLS 487
IX-23 CHEMICAL OXIDATION - HYDROGEN PEROXIDE FLOW
SCHEMATIC 488
IX-24 HYDROGEN PEROXIDE TREATMENT CAPITAL COST CURVE
FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
FROM BASE & PRECIOUS METAL ORE MILLS 489
IX-25 HYDROGEN PEROXIDE TREATMENT ANNUAL COST CURVE
FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
FROM BASE & PRECIOUS METAL ORE MILLS 490
IX-26 CHEMICAL OXIDATION-CHLORINE DIOXIDE FLOW
SCHEMATIC 491
IX-27 CHLORINE DIOXIDE CAPITAL COST CURVE FOR PHENOL
REDUCTION IN BPT EFFLUENT DISCHARGED FROM
BASE AND PRECIOUS METAL ORE MILLS 492
IX-28 CHLORINE DIOXIDE ANNUAL COST CURVE FOR PHENOL
REDUCTION IN BPT EFFLUENT DISCHARGED FROM
BASE AND PRECIOUS METAL ORE MILLS 493
IX-29 CHEMICAL OXIDATION-POTASSIUM PERMANGANATE
FLOW SCHEMATIC 494
IX-30 POTASSIUM PERMANGANATE TREATMENT CAPITAL COST
CURVE FOR PHENOL REDUCTION IN BPT EFFLUENT
DISCHARGED FROM BASE AND PRECIOUS METAL ORE
MILLS 495
IX-31 POTASSIUM PERMANGANATE TREATMENT ANNUAL COST
CURVE FOR PHENOL REDUCTION FROM BASE AND
PRECIOUS METAL ORE MILLS 496
xxi
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SECTION I
EXECUTIVE SUMMARY
This development document presents the technical data base
developed by the EPA to support effluent limitations guidelines
for the Ore Mining and Dressing Point Source Category. The Clean
Water Act of 1977 sets forth various levels of technology to
achieve these limitations. They are defined as best available
technology economically achievable (BAT), best conventional
pollutant control technology (BCT), and best available demon-
strated technology (BADT). Effluent limitations guidelines based
on the application of BAT and BCT are to be achieved by 1 July
1984. New source performance standards (NSPS) based on BADT are
to be achieved by new facilities. These effluent limitations
guidelines and standards are required by Sections 301, 304, 306,
307, and 501 of the Clean Water Act of 1977 (P.L. 95-217). They
augment the interim final regulations based on BPT, which were
first proposed on 6 November 1975. After extensive judicial
review, the final BPT regulations were published on 11 July 1978
and sustained by the 10th Circuit Court of Appeals on 10 December
1979.
Although the Clean Water Act of 1977 established the primary
legal framework for proposal of these limitations, EPA has also
been guided by a series of legally-binding judicial actions.
These include a series of settlement agreements, etc. into which
EPA entered with the National Resources Defense Council (NRDC)
and other environmental groups. The latest of these is NRDC v_._
Train, 8 ERC 2120 (D.D.C. 1976), modified, 12 ERC 1833 (D.D.C.
1979), aff'd and remd'd, EOF v_._ Costle, 14 ERC 2161 (D.D.C.
1980). The settlement agreement outlines a strategy for
regulation of toxic pollutant discharges according to a schedule
running through 1981 for 65 designated pollutant classes in 21
major industries, one of which is Ore Mining and Dressing. For
the purpose of regulation, the list of 65 pollutant classes
evolved into a list of 129 specific pollutants called "priority
pollutants" because of their importance of controlling discharges
of these toxic compounds. The priority pollutants serve as basis
for EPA's development of effluent limitations based on BAT and
BADT.
At present there are over 500 known major active ore mines (total
operations may number as many as 1000) and over 150 active ore
milling operations in the United States. Approximately
two-thirds of these mines and mills are existing point source
dischargers. The remainder do not discharge any process water.
There are no known existing indirect dischargers and no new
source indirect dischargers are anticipated. (Indirect
dischargers are those facilities which discharge to a publicly
owned treatment works.) Consequently, pretreatment standards,
which control the level of pollutants which may be discharged
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from an industrial plant to a publicly owned treatment works, are
not being proposed.
To recognize inherent differences in the industrial category, EPA
established subcategories within the larger category. The BPT
regulation for the ore mining and milling industry was divided
into 7 major subcategories based upon metal ore and 21
subdivisions based upon whether the facility was a mine or mill
and then further based upon the process employed at the mill.
The BPT subcategorization is retained under BAT with one
modification. The Ferroalloy ores subcategory which included
tungsten and molybdenum ore mines and mills has been split apart.
Molybdenum ore mines and mills are moved to the subcategory that
already includes copper, lead, zinc, gold, silver, and platinum
ore mines and mills. This new subcategory is renamed as the
copper, lead, zinc, gold, silver, platinum, and molybdenum ores
subcategory. Tungsten ore mines and mills are placed in a new
and separate subcategory. Three new subcategories have been
added since the time the court sustained the final BPT rule.
These are to apply to BAT, BCT, and NSPS. Each of the three new
subcategories consists of a single facility.
An extensive sampling and analysis effort was undertaken in 1977
and extends to the present. As part of this effort, 20
facilities were visited under screening and 14 facilities under
verification sampling, six facilities were visited for solid
waste and wastewater sampling, 12 treatability studies were
performed at nine sites, and data collected by EPA Regions VI,
VII, VIII, and X were reviewed to identify available treatment
technologies and to determine effluent levels that could be
achieved by these technologies. Six facilities were visited to
collect cost information as well as wastewater samples. Four
separate studies were performed by EPA's Industrial Environmental
Research Laboratory in Cincinnati on the treatability of
antimony, treatment alternatives for uranium mills, alternative
flotation reagents to replace cyanide compounds, and the
precision and accuracy of the analytical method for cyanide. The
data base also includes the BPT record, National Pollutant
Discharge Elimination System (NPDES) monitoring records, and data
submitted by the industry.
Three studies have been performed to determine the cost of
implementation of the candidate technologies. The first exercise
determined the cost of technologies based on model (typical)
facilities. The second costs the technologies in 1976 dollars
based on actual data from approximately 90 mines and mills which
had replied to an economic survey. These costs were verified in
a third study since the industry is so economically sensitive.
The costs presented in this document have been adjusted to 1979
dollars with appropriate inflation factors.
Executive Order 12291 (46 FR 13193-13198) requires that EPA and
other agencies perform Regulatory Impact Analyses of major regu-
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lations. The three conditions that determine whether a regula-
tion is classified as major are:
1. An annual effect on the economy of $100 million or more;
2. A major increase in costs or prices for consumers, individual
industries, federal, state, or local government agencies, or
geographic regions; or
3. Significant adverse effects on competition, employment,
investement productivity, innovation, or on the ability of United
States based enterprises to compete with foreign based
enterprises in domestic or export markets.
Under Executive Order 12291, EPA must judge whether a regulation
is "major" and therefore subject to the requirement of a
Regulatory Impact Analysis. This regulation is not major and
does not require a Regulatory Impact Analysis because the annual
effect on the economy is less than $100 million, it will not
cause a mjaor increase in costs, or significant adverse effects
on the industry.
This regulation was submitted to the Office of Management and
Budget for review as required by Executive Order 12251. Any
comments from OMB and EPA's responses to those comments are
available for public inspection at the EPA Public Information
Reference Unit, Room 2922 (EPA Library), Environmental Protection
Agency, 401 M Street, S.W., Washington, B.C.
BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE (BAT)
The presence or absence of the 129 toxic pollutants and several
conventional and nonconventional pollutants has been determined
as a result of the sampling and analysis program. One hundred
twenty-four of the 129 toxic pollutants have been excluded in all
subcategories based upon criteria contained in the Settlement
Agreement cited previously: (1) they were not detected, (2) they
were present at levels not treatable by known technologies, or
(3) they were effectively controlled by technologies upon which
other effluent limitations are based. The five remaining toxics
were excluded in some individual subcategories. Where specific
toxic pollutants are to be controlled with effluent limitations,
i.e., they were not excluded from the entire category or
individual subcategories, effluent limitations for those
pollutants are proposed. A number of end-of-pipe treatment
alternatives were considered for BAT, but were reduced to three
alternatives: (1) secondary settling; (2)
flocculation/coagulation; and (3) granular media filtration. The
remaining alternatives were eliminated because of high costs and
because some technologies were not applicable to an industrial
discharge characterized by extremely high flows and comparatively
low concentrations of pollutants in treated effluents. The three
options considered for controlling toxic metals were "add on" to
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BPT facilities which consist of lime precipitation and settling.
Of the three alternatives, no statistically significant
differences were discerned among the effluents from these
technologies.
Of these alternatives, secondary settling would require the least
expenditure. A statistical analysis of plant data from
facilities using secondary settling was used to derive achievable
levels which are more stringent than BPT. However, based on the
following considerations, the Agency has determined that
nationally applicable regulations based on secondary settling are
not warranted. First, in each subcategory, at least 95 percent
of the relevant pollutants are removed by BPT. Those pollutants
remaining are generally sulfi'de and oxide compounds in the form
of ore and gangue. Second, the Agency's environmental assessment
concluded that for the industrial category, the only
environmentally significant pollutants after stream flow dilution
are cadmium and arsenic and there is no appreciable reduction of
these between BPT and the derived levels. Finally, the BPT
limitations in this industry are generally more stringent than
BAT limitations being considered in other industries.
The BPT regulation provides for relief from effluent limitations,
including zero discharge, during periods of precipitation. The
basis of the precipitation exemption is that treatment facilities
must be designed, constructed, and maintained to include the
volume of water that would result from a 10-year, 24-hour
precipitation event. The same storm provision is retained in
BAT.
Where BPT is zero discharge for a subcategory, BAT is also zero
discharge. In subcategories where toxic pollutants were found,
BAT effluent limitations are proposed at BPT levels. As stated
previously all but five toxic pollutants are excluded from
regulation. These five are cadmium, copper, lead, mercury and
zinc.
BAT effluent limitations are not being established for asbestos.
BPT and BCT effluent limitations for TSS will effectively control
the discharge of asbestos. Available data demonstrated that, as
TSS levels are reduced in wastewater from mines and mills,
asbestos levels are reduced concomitantly, although the reduction
can not be quantified precisely. However, when TSS is reduced to
less than or equal to 30 mg/1 the data indicate that asbestos is
reduced to levels near observed background concentrations
(roughly 10s fibers per liter).
Uranium mills are excluded from BAT because pollutants from the
subcategory are from a single source and are uniquely related to
that source.
Cyanide is not regulated under BAT. A special study of the
precision and accuracy of the method was performed as part of the
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BAT review. Specific technology for the destruction of cyanide
was considered, but is not necessary because in-process controls
and retention of wastewater in tailing ponds reduce cyanide to
less than 0.4 mg/1 based on the precision and accuracy of the
analytical method for wastewater discharges from ore mines and
mills.
Gold placer mines are not regulated under the proposed BAT and
the subpart is reserved. Almost all gold placer mines are
located in remote areas of Alaska, No economic analysis has been
performed on these placer mines because no data are available
despite requests to industry for information. The Preamble to
the_ Proposed Regulation requests specific effluent data and cost
and cash flow data from palcer mines in order that an economic
impact assessment can be made before the regulation is proposed,
SOURCE PERFORMANCE STANDARDS NSPS
New facilities have an opportunity to implement the best and most
efficient ore mining and milling processes and wastewater
treatment technologies. Accordingly, Congress directed EPA to
consider the best demonstrated process changes and end-of-pipe
treatment technologies capable of reducing pollution to the
maximum extent feasible through a standard of performance which
includes, "where practicable, a standard permitting zero dis-
charge of pollutants".
NSPS for uranium mills is proposed as zero discharge. Zero
discharge is well demonstrated at existing uranium mills ( 1 8 of
19 do not discharge). New uranium mills in arid areas can
achieve zero discharge as cheaply as they can install treatment
to meet BPT limitations. Arid areas are those in which the
volume of water evaporated exceeds the volume resulting from
precipitation. The Agency knows of no mills actually planned for
humid areas.
NSPS for froth-flotation mills is proposed as zero discharge.
Zero discharge, based on total impoundment and recycle, or
evaporation, or a combination of these technologies, is demon-
strated as practicable at 46 of the 90 existing mills using froth
flotation for which EPA has data. Zero discharge (based on
recycle) was rejected as BAT because of the cost of retrofitting
the process in some existing mills and the potential changes
required in some existing mill processes where two or more
concentrates are recovered from the raw ore. The Agency's data
indicates that new source froth flotation mills can achieve zero
discharge as cheaply as they can install treatment to meet BPT.
The storm provision would depend on whether a facility is subject
to zero discharge. For NSPS requiring zero discharge, the
excursion applies only when a 10-year, 24-hour or greater storm
occurs and for NSPS allowing discharge subject to limitations,
the excursion will continue to be tied to the design criteria.
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BCT EFFLUENT LIMITATIONS
BCT was not intended to act as another limitation; rather, it
replaces BPT for control of the conventional pollutants: total
suspended solids (TSS), pH, biochemical oxygen demand (BOD), oil
and grease (O&G), and fecal coliform. Fecal coliform, BOD, and
O&G are not found in significant concentrations in this industry.
TSS and pH are central to control and treatment of the toxic
metals and are limited under BPT.
Since BCT is equivalent to BPT, no cost is implied to meet BCT
effluent limitations. BCT would pass any cost test since there
is no cost to implement BCT. The Agency is currently developing
a new BCT cost methodology. It is possible, though unlikely,
that a treatment technology more stringent than BPT will provide
additional removal of conventional pollutants and pass the new
cost test, when developed. In that event, the Agency will
propose new BCT limitations.
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In general, ore mines and mills are located in rural areas, far
from a POTW.
The 1977 Amendments added Section 301(b)(2)(E) to the Act
establishing "best conventional pollutant control technology"
(BCT) for discharges of conventional! pollutants from existing
industrial point sources. Conventional pollutants are those
defined in Section 304(a)(4) [biological oxygen demanding
pollutants (BODS), total suspended solids (TSS), fecal coliform,
and pH], and any additional pollutants defined by the
Administrator as "conventional" (oil and grease, 44 FR 44501,
July 30, 1979] .
BCT is not an additional limitation but replaces BPT for the
control of conventional pollutants. In addition to other factors
specified in section 304(b)(4)(B), the Act requires that BCT
limitations be assessed in light of a two part
"cost-reasonableness" test. American Paper Institute v. EPA, 660
F.2d 954 (4th Cir. 1981). The first test compares the cost for
private industry to reduce its conventional pollutants with the
costs to publicly owned treatment works for similar levels of
reduction in their discharge of these pollutants. The second
test examines the cost-effectiveness of additional industrial
treatment beyond BPT. EPA must find that limitations are
"reasonable" under both tests before establishing them as BCT.
In no case may BCT be less stringent than BPT.
EPA publisyed its methodology for carrying out the BCT analysis
on August 29, 1979 (44 FR 50732). In the case mentioned above,
the Court of Appeals ordered EPA to correct data errors
underlying EPA's calculation of the first test, and to apply the
second cost test. (EPA had argued that a second cost test was
not required.)
While EPA has not yet proposed or promulgated a revised BCT
methodology in response to the American Paper Institute v. EPA
decision mentioned earlier, EPA is proposing BCT limitations for
this category. These limits would be identical to those for BPT.
As BPT is the minimal level of control required by law, no
possible application of the BCT cost tests could result in BCT
limitations lower than those proposed today. Accordingly, there
is no need to wait until EPA revises the BCT methodology before
proposing BCT limitations.
Prior EPA Regulations
On 6 November 1975, EPA published interim final regulations
establishing BPT requirements for existing sources in the ore
mining and dressing industry (see 40 FR 51722). These regula-
tions became effective upon publication. However, concurrent
with their publications, EPA solicited public comments with a
view to possible revisions. On the same date, EPA also published
proposed BAT, NSPS, and pretreatment standards for this industry
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SECTION II
INTRODUCTION
PURPOSE
This study determined the presence and concentrations of the 129
toxic or "priority" pollutants in the ore mining and dressing
point source category for possible regulation. This development
document presents the technical data base compiled by EPA with
regard to these pollutants and their treatability for regulation
under the Clean Water Act. The concentrations of conventional
and nonconventional pollutants were also examined for the estab-
lishment of effluent limitations guidelines. Treatment technolo-
gies were also assessed for designation as the best available
demonstrated technology (BADT) upon which new source performance
standards (NSPS) are based. This document outlines the technol-
ogy options considered and the rationale for selecting each
technology level. These technology levels are the basis for the
proposed regulation effluent limitations.
LEGAL AUTHORITY
The regulations are proposed under authority of Sections 301,
304, 306, 307, 308, and 501 of the Clean Water Act (the Federal
Water Pollution Control Act Amendments of 1972, 33 USC 1251 et
seq., as amended by the Clean Water Act of 1977, P.L. 95-217)
(the "Act"). These regulations are also proposed in response to
the Settlement Agreement in Natural Resources Defense Council,
Inc., v. Train, 8 ERC 2120 (D.D.C. 1976), modified, 12 ERC 1833
(D.D.C. 1979).
The Clean Water Act
The Federal Water Pollution Control Act Amendments of 1972
established a comprehensive program to "restore and maintain the
chemical, physical, and biological integrity of the Nation's
waters," Section 101(a). By 1 July 1977, existing industrial
dischargers were required to achieve "effluent limitations
requiring the application of the best practicable control
technology currently available" (BPT), Section 301(b)(1)(A). By
1 July 1983, these dischargers were required to achieve "effluent
limitations requiring the application of the best available
technology economically achievable . . . which will result in
reasonable further progress toward the national goal of
eliminating the discharge of all pollutants" (BAT), Section
301(b)(2)(A). New industrial direct dischargers were required to
comply with Section 306 new source performance standards (NSPS),
based on best available demonstrated technology. The
requirements for direct dischargers were to be incorporated into
National Pollutant Discharge Elimination System (NPDES) permits
issued under Section 402 of the Act. Althouqh Section 402(a)(l)
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of the 1972 Act authorized the setting of requirements for direct
dischargers on a case-by-case basis, Congress intended that for
the most part, control requirements would be based on regulations
promulgated by the Administrator of EPA. Section 304(b) of the
Act required the Administrator to promulgate regulations
providing guidelines for effluent limitations setting forth the
degree of effluent reduction attainable through the application
of BPT and BAT. Moreover, Sections 304(c) and 306 of the Act
required promulgation of regulations for NSPS. In addition to
these regulations for designated industry categories, Section
307(a) of the Act required the Administrator to promulgate
effluent standards applicable to all dischargers of toxic
pollutants. Finally, Section 501(a) of the Act authorized the
Administrator to prescribe any additional regulations "necessary
to carry out his functions" under the Act. EPA was unable to
promulgate many of these regulations by the dates contained in
the Act. In 1976, EPA was sued by several environmental groups,
and in settlement of this lawsuit EPA and the plaintiffs executed
a Settlement Agreement which was approved by the Court. This
Agreement required EPA to develop a program and adhere to a
schedule for promulgating BAT effluent limitations guidelines,
and new source performance standards covering 65 classes of toxic
pollutants (subsequently defined by the Agency as 129 specific
"priority pollutants") for 21 major industries. See Natural
Resources Defense Council, Inc. v. Train, 8 ERC 2120 (D.D.C.
1976), modified, 12 ERC 1833 (D.D.C. 1979).
On 27 December 1977, the President signed into law the Clean
Water Act of 1977 ("the Act"). Although this law makes several
important changes in the Federal Water Pollution Control Program,
its most significant feature is its incorporation of several
basic elements of the Settlement Agreement program for toxic
pollution control. Sections 301(b)(2)(A) and 301(b)(2)(C) of the
Act now require the achievement, by 1 July 1984, of the effluent
limitations requiring application of BAT for toxic pollutants,
including the 65 priority pollutants and classes of pollutants
that Congress declared toxic under Section 307(a) of the Act.
Likewise, EPA's programs for new source performance standards are
now aimed principally at toxic pollutant controls. Moreover, to
strengthen the toxics control program, Section 304(e) of the Act
authorizes the Administrator to prescribe "best management
practices" (BMPs) to control the release of toxic and hazardous
pollutants from plant, site runoff; spillage or leaks; sludge or
waste disposal; and drainage from raw material storage associated
with, or ancillary to, the manufacturing or treatment process.
The proposed regulations provide effluent limitations guidelines
for BAT and establish NSPS on the basis of the authority granted
in Sections 301, 304, 306, 307, and 501 of the Clean Water Act.
Pretreatment Standards (PSES and PSNS) are not proposed for the
ore mining and dressing category since no known indirect
dischargers exist nor are any known to be in th'e planning stage.
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(see 40 FR 51738). Comments were also solicited on these
proposals.
On 24 May 1976, as a result of the public comments received, EPA
suspended certain portions of the interim final BPT regulations
and solicited additional comments (see 41 FR 21191). EPA
promulgated revised, final BPT regulations for the ore mining and
dressing industry on 11 July 1978, (see 43 FR 29711, 40 CFR Part
440). On 8 February 1979, EPA published a clarification of the
regulations as they apply to storm runoff (see 44 FR 7953). On 1
March 1979, the Agency amended the final regulations by deleting
the requirements for cyanide applicable to froth flotation mills
in the base and precious metals subcategory (see 44 FR 11546).
On 10 December 1979, the United States Court of Appeals for the
Tenth Circuit upheld the BPT regulations, rejecting challenges
brought by five industrial petitioners, Kennecott Copper Corp. y_._
EPA, 612 F.2d 1232 (10th Cir. 1979). The Agency withdrew the
proposed BAT, NSPS, and pretreatment standards on 19 March 1981
(see 46 FR 17567).
Industry Overview
The ore mining and dressing industry is both large and diverse.
It includes the ores of 23 separate metals and is segregated by
the U.S. Bureau of the Census Standard Industrial Classification
(SIC) into nine major codes: SIC 1011, Iron Ore; SIC 1021,
Copper Ores; SIC 1031, Lead and Zinc Ores; SIC 1041, Gold Ores;
SIC 1044, Silver Ores; SIC 1051, Aluminum Ore; SIC 1061,
Ferroalloy Ores including Tungsten, Nickel, and Molybdenum; SIC
1092 Mercury Ores; SIC 1094, Uranium, Radium, and Vanadium Ores;
and SIC 1099, Metal Ores, Not Elsewhere Classified including
Titanium and Antimony. Over 500 active mining and over 150
milling operations are located in the United States. Many are in
remote areas. The industry includes facilities that mine ores to
produce metallic products and all ore dressing and beneficiating
operations at mills operated either in conjunction with a mine
operation or at a separate location.
Summary of_ Methodology
From 1973 through 1976, EPA emphasized the achievement of
limitations based on application of best practicable technology
(BPT) by 1 July 1977. In general, this technology level
represented the average of the best existing performances of
well-known technologies for control of familiar pollutants
associated with the industry. In this industry, many metal
pollutants that Congress subsequently designated as toxic were
also regulated under BPT.
This rulemaking ensures the achievement, by 1 July 1984, of
limitations based on application of the best available technology
economically achievable (BAT). In general, this technology level
10
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represents the oest economically achievable performance in any
industry category or subcategory. Moreover, as a result of the
Clean Water Act of 1977, the emphasis of EPA's program has
shifted from control of "classical" pollutants to the control of
toxic substances.
EPA's implementation of the Act is described in this section and
succeeding sections of this document. Initially, because in many-
cases no public or private agency had done so, EPA, its
laboratories, and consultants had to develop analytical method?
for toxic pollutant detection and measurement. EPA then gathered
technical and economic data about the industry. A number of
steps were involved in arriving at the proposed limitations.
First, EPA studied the ore mining and dressing industry to
determine whether differences in raw materials; final products;
manufacturing processes; equipment. age, and size of plants;
water usage; wastewater constituents; or other factors required
the development of separate effluent limitations and standards
for different subcategori.es and segments of the industry. This
study included identifying raw waste and treated effluent
characteristics, including: the sources and volume of water
used, the processes employed, and the sources of pollutants and
wastewater in the plant and the constituents of wastewater,
including toxic pollutants. EPA then identified the constituents
of wastewaters that should be considered for effluent limitations
guidelines and standards of performance.
Next, EPA identified several distinct control and treatment
technologies, including both in-plant and end-of-process
technologies, that are in use or capable of being used in the ore
mining and dressing industry. The Agency compiled and analyzed
historical and newly generated data on the effluent quality
resulting from the application of these technologies. The long-
term performance, operational limitations, and reliability of
each treatment and control technology were also identified. In
addition, EPA considered the non-water quality environmental
impacts of these technologies, including impacts on air quality,
solid waste generation, water availability, and energy
requirements.
The Agency then estimated the costs of each control and treatment
technology from unit cost curves developed by standard
engineering analyses as applied to ore mining and dressing
wastewater characteristics. EPA derived unit process costs from
representative plant characteristics (production and flow)
applied to each treatment process (i.e., secondary settling, pH
adjustment and settling, granular-media filtration, etc.). These
unit process costs were added to yield total cost at each
treatment level. After confirming the reasonableness of this
methodology by comparing EPA cost estimates with treatment system
costs supplied by the industry, the Agency evaluated the economic
impacts of these costs.
11
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After considering these factors, EPA identified various control
and treatment technologies as BAT and NSPS. The proposed
regulation, however, does not require the installation of any
particular technology or limit the choices of technologies that
may be used in specific situations. Rather, it requires
achievement of effluent limitations that represent the proper
design, construction, and operation of these or equivalent
technologies.
The effluent limitations for ore mining and dressing BAT, BCT,
and NSPS are expressed in concentrations (e.g., milligrams of
pollutant per liter of wastewater) rather than loading per
unit(s) of production (e.g., kg of pollutant per metric ton of
product) because correlating units of production and wastewater
discharged by mines and mills was not possible for this category.
The reasons are:
1. The quantity of mine water discharged varies considerably
from mine to mine and is influenced by topography, climate,
geology (affecting infiltration rates) and the continuous nature
of water infiltration regardless of production rates. Mine water
may be generated and required to be treated and discharged even
if production is reduced or terminated.
2. Consistent water use and loss relationships for ore mills
could not be derived from facility to facility within a
subcategory because of wide variations in application of specific
processes. The subtle differences in ore mineralogy and process
development may require the use of differing amounts of water and
process reagents but do not necessarily require different
wastewater treatment technology(ies).
The Agency is not proposing pretreatment standards because it
does not know of any existing facilities that discharge to POTWs
or any that are planned.
Data Gathering Efforts
Data gathering for the ore mining and dressing industry included
an extensive collection of information:
1. Screening and verification sampling and analysis
programs
2. Enginereing cost site visits
3. Supporting data from EPA regional offices
4. Treatability studies
5. Industry self-monitoring sampling
6. BPT data base
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7. Placer study
8. Titanium sand dredges study
9. Uranium study
10. Solid waste study
EPA began an extensive data collection effort during 1974 and
1975 to develop BPT effluent standards. These data included
results from campling programs conducted by the Agency at mines
and mills and an assimilation of historical data supplied by the
industry, the Bureau of Mines, and other sources. This informa-
tion characterized waste-waters from ore mining and milling opera-
tions according to what were then considered key parameters-total
suspended solids, pH, lead, zinc, copper, and otb«r ^eta.ls,
However, little information on other environmental parameters,
such as other toxic metals and orgariics, was available from
industry or government sources. To establish the levels of thase
pollutants, the Agency instituted a second sampling and analysis
program to specifically address these toxic substances, Including
129 specific toxic pollutants for which regulation was mandated
by the Clean Water Act.
EPA began the second sampling and analysis program (screening and
verification sampling) in 1977 to establish the quantities of
toxic, conventional, and nonconyentional pollutants in ore mine
drainage and mill processing effluents. EPA visited 20 and 14
facilities respectively for screening and verification sampling.
EPA selected at least one facility in each major BPT subcategory.
The sites selected were representative of the operations and
wastewater characteristics present in particular subcategories.
These facilities were visited from April through November 1977.
To determine these sites, the agency reviewed the BPT data base
and industry as a whole, with consideration to:
1. Those using reagents or reagent constituents on
the toxic pollutants list;
2. Those using effective treatment for BPT regulated
pollutants;
3. Those for which historical data were available as
a means of verifying results obtained during
screening; and
4. Those suspected of producing wastewater streams
that contain pollutants not traditionally monitored.
After reviewing screen sampling analytical results, EPA selected
14 sites for verification sampling visits. Because most of the
organic toxic pollutants were either not detected or detected
13
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only at low concentrations in the screen samples, the Agency
emphasized verification sampling for total phenolics (4AAP),
total cyanide, asbestos (chrysotile), and toxic metals.
EPA revisited six of the facilities to collect additional data on
concentrations of total phenolics (4AAP), total cyanide, asbestos
(chrysotile), and to confirm earlier measurements of these
parameters.
After completing verification sampling, EPA conducted sampling of
two additional sites. At one molybdenum mill operation, a
complete screen sampling effort was performed to determine the
presence of toxic pollutants and to collect data on the
performance of a newly installed treatment system. The second
facility, a uranium mine/mill, was sampled to collect data on a
facility removing radium 226.by ion exchange. Samples collected
at this facility were not analyzed for organic toxic pollutants.
The Agency conducted a separate sampling effort to evaluate
treatment technologies at Alaskan placer gold mines. This study
was undertaken because gold placer mining was reserved under BPT
rulemaking and because little data were previously available on
the performance of existing treatment systems.
Industrial self-sampling was conducted at three facilities
visited during screen sampling to supplement and expand the data
for these facilities. The programs lasted from two to twelve
weeks. EPA selected two operations because they had been identi-
fied during the BPT study as two of the best treatment facili-
ties; the third because additional data on long-term variations
in waste stream charactersitics at these sites were needed to
supplement the historical discharge monitoring data, to reflect
any recent changes or improvements in the treatment technology
used, and to confirm that variations in raw wastewater levels did
not affect concentrations in treated effluents.
The Agency's regional surveillance and analysis groups performed
additional sampling at 14 facilities: nine in Colorado, Idaho,
Wyoming, and Montana; one in Arkansas; and four in Missouri.
Discharge monitoring reports were collected from EPA regional
offices for many of the ore producing facilities with treatment
systems. These data were used in evaluating the variations in
flow and wastewater characteristics associated with mine drainage
and mill wastewater.
The Agency took samples during the cost-site visits, although the
primary reason for the visits was to collect data that would
assist the Agency in developing unit process cost curves and that
would verify the cost assumption made. However, since many of
the sites had been sampled previously, the new sampling data
obtained served as additional verification of waste characteriza-
tion data.
14
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EPA conducted 13 treatability studies to characterize performance
of alternative treatment technologies on ore mine and mill
wastewaters. Secondary settling, flocculation, granular media
filtration, ozonation, alkaline chlorination and hydrogen
peroxide treatment were all examined in bench- and pilot-scale
studies. The data obtained from these studies were compared with
data obtained on the performance of these systems in actual
operations on pilot and full scale. In addition, the data were
used to determine the range of variability that might be expected
for these technologies, especially during periods of steady
running.
EPA obtained the data for its economic analysis primarily from a
survey conducted under Section 308 of the Clean Water Act. The
Agency sent questionnaires to 138 companies engaged in mining and
milling of metal ores. The data collected included production
levels, employment, revenue, operating costs, working capital,
ore grade, and other relevant information. The economic survey
data were supplemented by data from government publications,
trade journals, and visits to several mine/mills.
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16
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SECTION III
INDUSTRY PROFILE
ORE BENEFICIATION PROCESSES
As mined, most ores contain the valuable metals (values)
disseminated in a matrix of less valuable rock (gangue). The
purpose of ore beneficiation is the separation of the metal
bearing minerals from the gangue to yield a product which is
higher in metal content. To accomplish this, the ore must
generally be crushed and/or ground small enough so that each
particle contains mostly the mineral to be recovered or mostly
gangue. The separation of the particles on the basis of some
difference between the ore mineral and the gangue can then yield
a concentrate high in metal value, as well as waste rock
(tailings) containing very little metal. The separation is never
perfect, and the degree of success which is attained is generally
described by two numbers: (1) percent recovery and (2) grade of
the product (usually a concentrate). Widely varying results are
obtained in beneficiating different ores; recoveries may range
from 60 percent or less to greater than 95 percent. Similarly,
concentrates may contain less than 60 percent or more than 95
percent of the primary ore mineral. In general, for a given ore
and process, concentrate grade and recovery are inversely
related. Higher recovery is achieved only by including more
gangue, thereby yielding a lower grade concentrate. The process
must be optimized, trading off recovery against the value (and
marketability) of the concentrate produced. Depending on end
use, a particular grade of concentrate is desired, and specific
gangue components are limited as undesirable impurities.
Many properties are used as the basis for separating valuable
minerals from gangue, including: specific gravity, conductivity,
magnetic permeability, affinity for certain chemicals, solubil-
ity, and the tendency to form chemical complexes. Processes for
effecting the separation may be generally considered as:
1. gravity concentration
2. magnetic separation
3. electrostatic separation
4= ilutation
5. leaching
Amalgamation and cyanidation are variants of the leaching pro-
cess. Solvent extraction and ion exchange are widely applied
techniques for concentrating metals from leaching solutions and
for separating them from dissolved contaminants.
17
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These processes are discussed in general terms in the paragraph,-
which follow. This discussion is not meant to be all inclusive;
rather, it is to discuss the primary processes in current use in
the ore mining and milling industry. Details of some processes
used in typical mining and milling operations have been discussed
including presentation of process flowcharts, in Appendix A,
Industry Processes.
G_r_ay i.ty Concentration Processes
Gravity concentration processes exploit differences in density to
separate valuable ore minerals from gangue. Several techniques
(jigging, tabling, spirals, sink/float separation, etc.) are used
to achieve the separation. Each is effective over a somewhat
limited range of particle sizes, the upper bound of which is set
by the size of the apparatus and the need to transport ore within
it, and the lower bound by the point at which viscous forces
predominate over gravity and render the separation ineffective.
Selection of a particular gravity based process for a given ore
will be strongly influenced by the size to which the ore must be
crushed or ground to separate values from gangue as well as by
the density difference and other factors.
Most gravity techniques depend on viscous forces to suspend and
transport gangue away from the heavier valuable mineral. Since
the drag forces on a particle depend on its area, and its weight
depends on its volume, particle size as well as density will have
a strong influence on the movement of a particle in a gravity
separator. Smaller particles of ore mineral may be carried with
the gangue despite their higher density, or larger particles of
gangue may be included in the gravity concentrate. Efficient
separation, therefore, requires a process feed with uniform
particle sizes. A variety of classifiers (spiral and rake
classifiers, screens, and cyclones) are used to assure a
reasonably uniform feed. At some mills, a number of sized
fractions of ore are processed in different gravity separation
units.
Viscous forces on the particles set a lower particle size limit
for effective gravity separation by any technique. For very
small particles, even slight turbulence may suspend the particle
for long periods of time, regardless of density. Such
suspensions (slimes), cannot be recovered by gravity techniques
and may cause very low recoveries in gravity processing of
friable ores, such as scheelite (calcium tungstate, CaWOO.
Jigs
Jigs of many different designs are used to achieve gravity
separation of relatively coarse ore usually between 0.5 mm (0.02
inch) and 25 mm (1 inch) in diameter. In general, ore is fed as
a thick slurry to a chamber in which agitation is provided by a
pulsating plunger or other such mechanism. The feed separates
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;.Vito layers
drawn off
mineral
coarser
by density within the jig, the lighter gangue being
at the top, with the water overflow and the denser
drawn off at a screen on the bottom. Often a bed of
ore or iron shot is used to aid the separation; the dense
ore mineral migrates down through the bed under
the agitation within the jig. Several jigs are
series to achieve both acceptable recovery and
grade. Jigs are employed in the processing of
gold, and ferroalloys.
Tables
the influence of
often used in
high concentrate
ores of iron,
Shaking tables of a wide variety of designs have found widespread
use as an effective means of achieving gravity separation of
003 inch) to 2.5 mm (0.1 inch)
n
finer ore particles 0.08 mm
diameters. Fundamentally, they are tables over which flow ore
particles suspended in water, A series of ridges or riffles
perpendicular to the water flow traps heavy particles while
lighter ones are suspended and flow over the obstacles with the
water stream. The heavy particles move along the ridges to the
edge of the table and are collected as concentrate (heads), while
the light material which follows the water flow is generally a
waste stream (tails). Between these streams may be some material
(middlings) which nas been partially diverted by the riffles,
These are often collected separately and returned to the table
feed. Reprocessing of either heads or tails, or both, and
multiple stage tabling are common. Tables may be used to
separate 'minerals of minor density differences, but uniformity of
feed becomes extremely important in such cases. Tables are
employed
titanium,
Spirals
Humphreys
provide an
volumes
inch) in
in the
monazitc
conduit
of ore is
in che processing
and zirconium.
of ores of gold, ferroalloys
spiral separators, a relatively recent development,
efficient means of gravity separation for large
of material between 0.1 mm and 2 mm (.004 inch to .08
diameter. They have been widely applied, particularly,
processing of heavy sands for ilmenite (FeTiOS.) and
(a rare earth phosphate). Spirals consist of a helical
(usually of five? turns) about a vertical axis. A slurry
fed to the conduit at the top and flows down
the spiral
under gravity. The heavy minerals concentrate along the inner
edge of the spiral from which they may be withdrawn through a
series of ports. Wash water; may also be added through ports
along the inner edge to improve the separation efficiency. A
single spiral may typically be used to process 0.5 to 2.4 metric
tons (O.SS to 2.64 short tons) of ore per hout; in large plants,
as many as several hundred spirals may be run in parallel.
Spirals are used for processing ores of ferroalloys, titanium,
and zirconium.
Sink/Float Separation
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Sink/float separators differ from most gravity methods in that
bouyancy forces are used to separate the various minerals on the
basis of density. The separation is achieved by feeding the ore
to a tank containing a medium of higher density than the gangue
and less than the valuable ore minerals. As a result, the gangue
floats and overflows the separation chamber, and the denser
values sink and are drawn off at the bottom by a bucket elevator
or similar contrivance. Because the separation takes place in a
relatively still basin and turbulence is minimized, effective
separation may be achieved with a more heterogeneous feed than
for most gravity separation techniques. Viscosity does, however,
place a lower bound on separable particle size because small par-
ticles settle very slowly, limiting the rate at which ore may be
fed. Further, very fine particles must be excluded since they
mix with the separation medium, altering its density and
viscosity.
Media commonly used for sink/float separation in the ore milling
industry are suspensions of very fine ferrosilicon or galena
(PbS) particles. Ferrosilicon particles may be used to achieve
medium specific gravities as high as 3.5 and are used in heavy-
medium separation. Galena, used in the "Huntirigton-Heberlein"
process, allows the achievement of somewhat higher densities.
The particles are maintained in suspension by a modest amount of
agitation in the separator and are recovered for reuse by washing
both values and gangue after separation.
Sink/float separation techniques are employed for processing ores
of iron.
Magnetic Separation
Magnetic separation is widely applied in the ore milling
industry, both for the extraction of values from ore and for the
separation of different valuable minerals recovered from complex
ores. Extensive use of magnetic separation is made in the pro-
cessing of ores of iron, columbium, and tungsten. The separation
is based on differences in magnetic permeability (which, although
small, is measurable for almost all materials). This method is
effective in handling materials not normally considered magnetic.
The basic process involves the transport of ore through a region
of high magnetic field gradient. The most magnetically permeable
particles are attracted to a moving surface by a large electro-
magnet. The particles are carried out of the main stream of ore
by the moving surface and as it leaves the high field region the
particles drop off into a hopper or onto a conveyor leading to
further processing.
For large scale applications (particularly in the iron ore
industry) large, rotating drums surrounding the magnet are used.
Although dry separators are used for rough separations, drum
separators are most often run wet on the slurry produced in
20
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griiiding mills, Where smaller amounts of material are handled,
wet and crossed-belt separators are frequently employed.
Magnetic separations is used in the beneficiation of ores of iron,
ferroalloys, titanium, and zirconium are discussed in Appendix A.
Electrostatic Separation
Electrostatic separation is used to separate minerals on the
basis of their conductivity. It is an inherently dry process
using very high voltages (typically 20,000 to 40,000 volts). In
a typical implementation, ore is charged to 20,000 to 40,000
volts, and the charged particles are dropped onto a conductive
rotating drum. The conductive particles discharge very rapidly,
are thrown off, and collected. The nonconductive particles keep
their charge and adhere by electrostatic attraction to be removed
from the drum separately. Specific instances in which electro-
static separation has been used for processing ores of ferro-
alloys, titanium, and zirconium, are discussed in Appendix A.
Flotation Processes
Basically, flotation is a process where particles of one mineral
or group of minerals are made by addition of chemicals to adhere
preferentially to air bubbles. When air is forced through a
slurry of mixed minerals, the rising bubbles carry with them the
particles of the mineral(s) to be separated from the matrix. If
a foaming agent is added which prevents the bubbles from bursting
when they reach the surface, a layer of mineral laden foam is
built up at the surface of the flotation cell which may be
removed to recover the mineral. Requirements for the success of
the operation are that particle size be small, that reagents be
compatible with the mineral, and that water conditions in the
cell not interfere with attachment of reagents to the mineral or
to air bubbles.
Flotation concentration has become a mainstay of the ore milling
industry. Because it is adaptable to very fine particle sizes
(less than 0.001 cm), it allows high rates of recovery from
slimes, which are inevitably generated in crushing and grinding
and which are not generally amenable to physical processing. As
a physio-chemical surface phenomenon, it can often be made highly
specific, allowing production of high grade concentrates from
relatively low grade ore (e.g., over 95 percent MoS2_ concentrate
from 0.3 percent ore). Its specificity also allows separation of
different ore minerals (e.g., CuS, PbS, and ZnS) and operation
with minimum reagent consumption since reagent interaction is
typically only with the particular materials to be floated or
depressed.
Details of the flotation process (exact dosage of reagents,
fineness of grinds, number of regrinds, cleaner flotation steps,
etc.) differ at each operation where it is practiced and may
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often vary with time at a given mill. A complex system of rea-
gents is generally used, including four basic types of compounds:
collectors, frothers, activators, and depressants. Collectors
serve to attach ore particles to air bubbles formed in the flota-
tion cell. Frothers stablilize the bubbles to create a foam
which may be effectively recovered from the water surface.
Activators enhance the attachment of specific kinds of particles
to the air bubbles, and depressants prevent it. Frequently,
activators are used to allow flotation of ore which has been
depressed in an earlier stage of the process. In almost all
cases, use of each reagent in the mill is low (generally, less
than 0.5 kg per ton of ore processed), and the bulk of the
reagent adheres to tailings or concentrates.
Sulfide minerals are readily recovered by flotation using similar
reagents in small doses; although reagent requirements vary
throughout the class. Sulfide flotation is most often carried
out at alkaline pH. Collectors are most often alkaline xanthates
having two to five carbon atoms, for example, sodium ethyl
xanthate (NaS2COC2_H5^ , Frothers are generally organics with a
soluble group and a nonwettable hydrocarbon. Pine oil hydroxyl
(C6Hr20H), for example, is widely used to allow separate recovery
of metal values from mixed sulfide ores. Sodium cyanide is
widely used as a pyrite depressant. Activators useful in sulfide
ore flotation may include cuprous sulfide and sodium sulfide.
Sulfide minerals of copper, lead, zinc, molybdenum, silver,
nickel, and cobalt are commonly recovered by flotation.
Many minerals in addition to sulfides may be, and often are,
recovered by flotation. Oxidized ores of iron, copper, manga-
nese, the rare earths, tungsten, titanium, columbium and tanta-
lum, for example, may be processed in this way. Flotation of
these ores involves a very different group of reagents from
sulfide flotation and has, in some cases, required substantially
larger dosages. These flotation processes may be more sensitive
to feed water conditions than sulfide floats. They are less
frequently run with recycled water or untreated water. Collector
reagents used include fatty acids (such as oleic acid or soap
skimmings), fuel oil, various amines, and compounds such as
copper sulfate, acid dichromate, and sulfur dioxide as
conditioners.
Flotation is also used to process ores of iron, copper, lead and
zinc, gold, silver, ferroalloys, mercury, and titanium.
Leaching
Ores can be beneficiated by dissolving away either gangue or
values in aqueous acids or bases, liquid metals, or other
specific solutions. This process is called leaching.
Leaching solutions are categorized as strong, general' solvents
(e.g., acids) and weaker, specific solvents (e.g., cyanide). The
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acids dissolve any metals present, which often include gangue
constituents (e.g., calcium from limestone). They are convenient
to use since the ore does not have to be very finely ground, and
separation of the tailings from the value bearing (pregnant)
leach is then not difficult.
Specific solvents attack only one (or, at most, a few) ore
constituent(s). Ore must be finely ground to expose the values.
Heat, agitation, and pressure are often used to speed the actions
of the leach, and it is difficult to separate the solids (often
in the form of slimes) from the pregnant leach.
Countercurrent leaching, preneutralization of lime in the gangue,
leaching in the grinding process, and other combinations of
processes are often seen in the industry. The values contained
in the pregnant leach solution are recovered by one of several
methods, including precipitation (e.g., of metal hydroxides from
acid leach by raising pH), electrowinning (a form of
electroplating), and cementation. Ion exchange and solvent.
extraction are often used to concentrate values before recovery.
Ores can be exposed to leach in a variety of ways. In vat
leaching, the process is carried out in a container (vat), often
equipped with facilities for agitation, heating, aeration, and
pressurization (e.g., Pachuca tanks). In situ leaching is
employed in shattered or broken ore bodies on the surface, or in
old underground workings. Leach solution is applied either by
plumbing or percolation through overburden. The leach is allowed
to seep slowly to the lower levels of the ore body or mine where
it is pumped from collection sumps to a metal recovery or precip-
itation facility. In situ leaching is most economical when the
ore body is surrounded by an impervious matrix which minimizes
loss of leach solution. However, when water suffices as a leach
solution and is plentiful, in situ leaching is economical, even
in previous strata.
Low-grade ore, oxidized ore, or tailings can be treated above
ground by heap or dump leaching. Dump leaching is usually
employed for leaching of low-grade ore. Most leach dumps are
deposited on existing topography. The dump site is often
selected to take advantage of impermeable surfaces and to utilize
the natural slope of ridges and valleys for the collection of
pregnant leach solution. Heap leaching is employed to leach ores
of higher grade or value. Heap leaching is generally done on a
specially prepared impervious surface (asphalt, plastic sheeting,
or clay) that is furrowed to form drains and launders (collecting
troughs). This configuration is employed to minimize loss of
pregnant leach solution. The leach solution is typically applied
by spraying, and the launder effluent is treated to recover metal
values.
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Gold (cyanide leach) and uranium and copper (sulfuric ado leach)
are recovered by leaching processes. Leaching is also used in
processing ores of ferroalloys, radium, and vanadium.
Amalgamation
AmaJgamation is the process by which mercury is alloyed,
generally to gold or silver, to produce an amalgam. This process
is applicable to free milling of precious metal ores; that is,
those in which the gold is free, relatively coarse, and has clean
surfaces. Lode or placer gold and silver that is partly or com-
pletely filmed with iron oxides, greases, tellurium, or sulfide
minerals cannot be effectively amalgamated. Hence, prior to
amalgamation, auriferous ore is typically washed and ground to
remove any films on the precious metal particles. Although the
amalgamation process was used extensively for the extraction of
gold and silver from pulverized ores, it has largely been
superseded by the cyanidation process due to environmental
considerations.
A more complete description of amalgamation practices for the
recovery of gold values can be found in Appendix A.
Cyanidation
With occasional exceptions, lode gold and silver ores now are
processed by cyanidation. Cyanidation is a process for the
extraction of gold and/or silver from finely crushed ores, con-
centrates, tailings, and low grade mine run rock by means of
potassium or sodium cyanide used in dilute, weakly alkaline
solutions. The gold is dissolved by the solution according to
the reaction:
4Au + SNaCN + 2H20 + 02 - 4NaAu(CN)2 + 4NaOH
and subsequently adsorbed onto activated carbon (carbon-in-pulp
process) or precipitated with metallic zinc according to the
reaction (Reference 1):
2NaAu(CN)2_ + 4NaCN + 2Zn + 2E20
2Na2Zn(CN)4 + 2Au + H2 + 2NaOH
The gold particles are recovered by filtering, and the filtrate
is returned to the leaching operation.
The carbon-in-pulp process was developed to provide economic
recovery of gold from low grade ores or slimes. In this process,
gold which has been solubilized with cyanide is brought into
contact with activated coconut charcoal in a series of tanks.
The ore pulp and enriched carbon are air lifted and discharged
onto small vibrating screens between tanks, where the carbon is
separated and moved to the next adsorption tank. Gold-enriched
carbon from the last adsorption tank is leached with hot caustic
24
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cyanide solution to desorb the gold. This hot, high grade solu-
tion containing the leached gold is then sent to electrolytic
cells where the gold and silver are deposited onto stainless
steel wool cathodes.
Pretreatment of ores containing only finely divided gold and
silver usually includes multistage crushing, fine grinding, and
classification of the ore pulp into sand and slime fractions.
The sand fraction is then leached in vats with dilute, well-
aerated cyanide solution. After the slime fraction has thick-
ened, it is treated by agitation leaching in mechanically or air
agitated tanks, and the pregnant solution is separated from the
slime residue by thickening and/or filtration. Alternatively,
the entire finely ground ore pulp may be leached by agitation
leaching and the pregnant solution recovered by thickening and
filtration.
When this process is employed, the pulp is also washed by
countercurrent decantation (CCD) to maximize the efficiency of
gold recovery. A CCD circuit consists of a number of thickeners
connected in series. Direction of the overflow through the
thickeners is countercurrent to the direction of the underflow
(pulp). Wash solution used in the CCD circuit is subsequently
dosed with cyanide and used in the agitation leaching process.
Pulp moving through the CCD circuit is discharged from the last
thickener to tailings disposal.
In all of these leaching processes, gold or silver is recovered
from the pregnant leach solutions; however, different types of
gold/silver ore require modification of the basic flow scheme.
Efficient low-cost dissolution and recovery of the gold and
silver are possible only by careful process control of the unit
operations involved.
A more complete description of cyanidation practices for the
recovery of gold values can be found in Appendix A.
Ion Exchange and Solvent Extraction
These processes are used on pregnant leach solutions to
concentrate values and to separate them from impurities. Ion
exchange and solvent extraction are based on the same principle:
polar organic molecules tend to exchange a mobile ion in their
greater charge or a smaller ionic radius. For example, let R be
the remainder of a polar molecule (in the case of a solvent) or
of a polymer (for a resin), and let X be the mobile ion. Then,
the exchange reaction for a uranyltrisulfate complex is:
4RX + (U02(S04)3 > R4 U02 (S04)3 + 4X
This reaction proceeds from left to right in the loading process.
Typical resins adsorb about 10 percent of their mass in uranium
and increase by about 10 percent in density. In a concentrated
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solution of the mobile ion (for example, in N-hydrochloric acid),
the reaction can be reversed, and the uranium values are eluted
(in this example, as hydrouranyl trisulfuric acid). In general,
the affinity of cation-exchange resins for a metal lie cation
increases with increasing valence:
Cr ]|j 0 Mg || > Na|
and, because of decreasing ionic radius, with atomic number:
92U > 42MO > 23V
The separation of hexavalent 92U cations by ion exchange or
solvent extraction should prove to be easier than that of any
other naturally occurring element.
Uranium,, vanadium, and molybdenum (the latter being a common ore
constituent) usually appear in aqueous solutions as oxidized ions
iuranyl, vanadyl, or molybdate radicals). Uranium and vanadium
are additionally complexed with anionic radicals to form
trisulfates or tricarbonates in the leach. Since the complexes
react anionically, the affinity of exchange resins and solvents
is not simply related to fundamental properties of the heavy
metal (U, V, or Mo) as is the case in cationic-exchange
reactions. Secondary properties of the pregnant solutions
influence the adsorption of heavy metals. For example, seven
times more vanadium than uranium is adsorbed on one resin at pH
9; at pH 11, the ratio is reversed with 33 times more uranium
than vanadium being captured. Variations in affinity, multiple
columns, and leaching time with respect to breakthrough (the time
when the interface between loaded and regenerated resin arrives
at the end of the column) are used to make an ion exchange
process specific for the desired product.
In solvent extraction, the type and concentration of a polar
solvent in a nonpolar diluent (e.g., kerosene) affect separation
of the desired product. Solvent handling ease permits the con-
struction of multistage, concurrent and countercurrent, solvent
extraction concentrators which are useful even when each stage
effects only partial separation of a value from an interferent.
Unfortunately, the solvents are easily polluted by slime so
complete liquid/solid separation is necessary. Ion exchange and
solvent extraction circuits can be combined to take advantage of
the slime resistance of resin-in-pulp ion exchange and the
separatory efficiency of solvent extraction (Eluex process).
Ion excnange and solvent extraction methods are applied in the
processing of ores of uranium/radium/vanadium.
ORE MINING METHODS
Metal-ore mining is conducted by a variety of surface, under-
ground, and in situ procedures. The terminology used to describe
26
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L.hese procedures :ias been defined in a United States Bureau of
Hines (USBM) mining dictionary (Reference 2).
Surface mining includes quarrying, open-pit, open-cut, open-cast,
stripping, placering, and dredging operations. The USBM
Dictionary definitions provide no clear distinctions between
ouarrying, open-pit, open-cut, open-cast, and strip mining. A
preference can be discerned for using the word "quarrying" in
connection with surface mining of stone, although it often is
used in connection with surface mining of ail construction
materials. Strip mining appears to be the preferred term for
surface mining by successive parallel cuts that are filled, in
turn, with overburden. Red-bed copper in Oklahoma and bauxite in
Arkansas are mined in this way.
fV,
The terms "open-pit" and "'open-cut" identify surface mines other
than quarries, strip mines, and placers, but are often applied to
these types also. The term open-pit is used more specifically
for surface mines in relatively thick ore bodies characterized by
permanent disposal of wastes and terraced or benched slopes.
Most of the crude ores of. copper and iron come from surface mines
of this type.
Placer mining, which employs a variety of equipment and tech-
niques including dredging, is used in mining and concentration of
alluvial gravels and elevated beach sands. All the illmenite
production in Florida and New Jersey is obtained by the dredging
of elevated beach sand deposits. Although hydraulic placer
r.iin'ing of gold in California was enjoined by court decree in the
1880's, placering for gold, by other methods, continues in
California and Alaska.
Underground mining is conducted through adits or shafts by a
variety of methods that include room-and-pillar, block caving,
timbered stopes. open stopes, shrinkage stopes, sublevel stopes
and others (Reference 3). Underground mining usually is inde-
pendent of surface mining, but sometimes preceeds or follows it.
Waste removal is proportionately much less in underground than in
surface mining, but still requires surface waste disposal areas.
Underground mines supply substantially all the lead and zinc
mined domestically.
In situ mining procedures include the leaching of uranium and
copper.
Considerations given to the choice of mining method and brief
descriptions of the methods typically employed are given below.
Surface Mining Operations
Whether an ore body will be mined by surface or underground
methods will be determined by the economics of the operation. In
general, surface or open-pit mining is more economical than
-------
underground mining especially when the ore body is large and the
depth of overburden is not excessive.
Some predominant advantages inherent in the open-pit method are
as follows:
a) The open-pit method is quite flexible in that it often
allows for large increases or decreases in production on
short notice without rapid deterioration of the
workings.
b) The method is relatively safe. Loose material can be
seen and removed or avoided. Crews can be readily
observed at work by supervisors.
c) Selective mining is usually possible without difficulty.
Grade control can be easily accomplished by leaving lean
sections temporarily unmined or by mining for waste.
d) The total cost of open-pit mining, per ton recovered, is
usually only a fraction of the cost of underground
mining. Further, the cost spread between the two
methods is growing wider as larger-scale methods are
applied to open pits.
Drilling
Drilling is the basic part of the breaking operation in open-pit
mining; considerable effort has therefore been expended to
develop equipment for drilling holes at the lowest possible cost.
There are several types of drills which can be used. These
include churn drills, percussion drills, rotary drills and jet-
piercing drills. All are designed with one objective in mind: to
produce a hole of the required diameter, depth, and direction in
rock for later insertion of explosives.
Blasting
Basically, explosives are comprised of chemicals which, when
combined, contain all the requirements for complete combustion
without external oxygen supply. Early explosives consisted
chiefly of nitroglycerine, carbonaceous material and an oxidizing
agent. These mixtures were packaged into cartridges for
convenience in handling and loading into holes. Many explosives
are still manufactured and packaged to the basic formulas.
In recent years, it has been discovered that fertilizer-grade
ammonium
detonated
spread to
use this
blasting .
nitrate mixed with about six percent fuel oil could be
by a high explosive primer. This new application has
the point where virtually all open-pit mining companies
mixture (called ANFO) for some or all of the primary
28
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Typicaly, iuultiplu charges are placed in two rows of drill holes
and fired simultaneously to break free the upper part of the face
and prevent "back break" beyond the line of holes. In
horizontally bedded deposits, the face is often held right on the
line of holes. Fifty feet is almost standard for the height of a
bench in hard ground, but 30-foot and 40-foot benches are common.
The height depends on the reach of the shovel and the character
of the ore.
Stripping
Material overlying an ore body may consist of earth, sand,
gravel, rock or even water. Removal of this material generally
falls under the heading of stripping. Normally, stripping of
rock will be considered a mining operation. Generally, stripping
will be accomplished with heavy earth-moving equipment such as
large shovels, mounted on caterpillar treads and driven by
electricity or diesel power, and bulldozers. In the past,
railroad cars were used to haul the stripped material to the dump
area. Now, however, they have been replaced by large trucks
except for situations where long hauls are required.
Loading
After the ore has been broken down, it is transferred to the mill
for treatment. In small pits various kinds of small loaders,
such as scrapers or tractor loaders, are sometimes used, but in
most places the loading is done with a power shovel. Tractors
equipped with dozer blades are used for pushing ore over banks so
that it can be reached by the shovel, but their most effective
use is in cleaning up after the shovel, pushing loose ore back
against the toe of the pile where it can be readily picked up on
the next cut.
Underground Mining Operations
Historically, the mining method most often used has been some
form of open stope. Generally, to reach the ore a shaft is sunk
near the ore body. Horizontal passages are cut from the shaft at
various depths to the ore. The ore is then removed, hoisted to
the surface, crushed, concentrated and refined. Waste rock or
classified mill tailings may be returned to the mine as fill for
the mined-out areas or may be directed to a disposal basin
(tailings area).
Caving systems of mining ore have been developed as economical
approaches to mining extensive low-grade ore bodies.
The Shaft
The shaft is the surface opening to the mine which provides a
means of entry to or exit from the mine for men and materials,
and for the removal of ore or waste from underground to the
29
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surface. It may be vertical or inclined. (A passageway or
opening driven horizontally into the side of a hill generally for
the purpose of exploring or otherwise opening a mineral deposit
is called an adit. Strictly speaking, an adit is open to the
atmosphere at one end).
With the advent of modern day mining equipment which has greatly
increased the speed of shaft sinking it is -presently more
economical to sink deep hoisting shafts, and vertical shafts are
preferred to inclines. In the U.S., mines have ranged to a depth
of 2286m (7500 feet). Although it is unusual for a single shaft
to be deeper than 1219 meters (4000 feet), one shaft has been
sunk to a depth of 2286 meters (7500 feet) in the Coeur d'Alene
Mining District of Northern Idaho and another has exceeded 2620
meters (8600 feet) in South Dakota.
In the United Jtates, what are known as "square shafts," which
have two skip compartments and one or two large cage compart-
ments, are now the most popular, because they allow the use of
large cages, on which mine timbers can be taken into the mine on
trucks without rehandling. These shafts have the additional
advantage of getting the crew into the mine and out again in a
relatively short time. Shafts sunk from underground levels are
called winzes. Winzes are established to permit mining at deeper
depths.
Shaft conveyances include buckets, skips, cages or skip-cage
combinations. The first two are for hoisting rock or ore and
they vary in load capacity from one to eighteen tons. They
travel at approximately 610-914 meters (2000-3000 feet) per
minute. Cages are used for men and materials and can transport
as many as 85 men per load at slower speeds. Safety devices
exist to prevent shaft conveyances from failing, should cables
fail.
There are generally two types of hoists in use. The Koepe or
friction-drive hoist, in common use in Europe since 1875, was
first introduced to North America approximately two decades ago.
Many are now in service. In this type of hoisting operation,
ropes (cables) pass over a drum with counter-balancing weights or
loads on either side. These are raised or lowered via friction
between the ropes and the drum treads on which they rest. The
ropes pass over the drum only once. The arc of contact between
rope and drum is normally 180 degrees. On the conventional drum-
winder hoist the rope is wound onto the drum and, as such, loads
are raised or lowered by a simple winding or unwinding operation.
Levels
Levels are horizontal passes in a mine. They are generally
driven from the shaft at vertical intervals of 100-200 feet.
That part of the level driven from the shaft to the ore body is
known as the crosscut, and that part which continues along the
30
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ore body is known as a drift. Crosscuts and drifts vary in size
from about 2' x 7' to about 9' x 16' depending on the size of the
haulage equipment in use. A raise is an opening made in the back
(roof) of a level to reach the level above.
Scopes
A stope is an excavation where the ore is drilled, blasted and
removed by gravity through chutes to ore cars on the haulage
level below. Stopes require timbered openings (manways) to
provide access for men and materials. Normally, raises connect a
stope to the level above and are used for ventilation, for con-
venience in getting men and materials into the stope, and for
admitting backfill.
Stope Mining Methods
Today more than half the metallic ore produced from underground
methods is mined by open stopes with rooms and pillars.
Nearly all of the lead, zinc, gold, and silver mined from
underground in the U.S. is mined by this method as well as much
of the uranium and some copper and iron. The three commonly used
stoping methods are cut-and-fill stoping, square-set stoping, and
shrinkage stoping. The stoping method used normally depends on
the stability of the walls and roof as ore removal progresses.
The cut-and-fill stope is used in wider irregular ore bodies
where the walls require support to minimize dilution (i.e. waste
from walls falling into the broken ore). In its simplest form
this mining system consists of blasting down a horizontal cut
across the vein for a length of 15 meters (50 feet) or more,
removing the broken ore and filling the opening thus made with
waste (or mill tailings) until it is high enough to attack the
back again. Chutes and manways are raised prior to each addition
of fill. Waste material (fill) is dumped into the stope through
a waste-pass raise to the surface until it is level with the
chutes and manways. Flooring (wood or concrete) is placed over
the fill before the next ore cut is drilled and blasted.
Scrapers or diesel endloaders are used to remove ore to the
chutes and to level the waste backfill in the stope.
The square-set stope is used in an ore body where the walls anc
ore require support during ore removal. After each blast,
square-set timbers are erected and made solid by blocking to the
walls and back. Square-sets alone will not support large blocks
of ground, and therefore their primary function is to serve1 as a
working platform for the miners and as a protection from falling
ground. Consequently, square-sets have become recognized as a
system of timbering rather than a system of mining. In good
practice, not more than two sets high are allowed to stand open.
On the top floor the ore is drilled and blasted, and is allowed
to fal] to the floor below, where it is shoveled into chutes.
31
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The chutes and manways are raised and backfill is placed in the
stope as in the cut-and-fill method.
The shrinkage stope is used chiefly in narrow regular ore bodies
where the walls and ore require little support. After eacn
blast, sufficient ore is pulled from the chutes to make room for
the miners to drill and blast the next section. As the stope
progresses upwards the manways are raised slightly above the
level of the broken ore. When the stope reaches the level above,
it will be full of broken ore. On removal of the ore, the stopes
may be filled with waste material.
Undercut Block Caving Mining Method
This system of mining is applicable to large thick deposits of
weak ore which are undercut by a gridwork of drifts and cross-
cuts. The small pillars thus blocked out are reduced in size
until they cave, and the whole mass is allowed to settle and
crush, A variation of this used in some places relies on induced
caving, blasting being used to start the ore movement.
Generally, 91 meters (300 feet) is an economical height for
caving and ore is mined in panels in a retreating system. In
each panel the ore is undercut on a sublevel, the width of the
unsupported section depending on the strength of the ore.
In thick ore an elaborate system of branch raises carries the
broken ore to the main level, caving being regulated by the
amount of ore drawn off through finger raises immediately under
the undercutting level.
In thinner ore bodies, and in places where such an elaborate
system of branch raises is not justified, various expedients are
used instead of the branch raises. Scrapers in transfer drifts,
pulling the ore from finger raises to main chutes, are one
successful approach, and shaking conveyors for the same purpose
are another.
Undercut caving has been one of the most successful and
revolutionary of the new mining systems, and by its reduction in
cost has changed tremendous quantities of what would otherwise be
waste rock into profitable ore.,
Sublevel Caving Mining Method
The sublevel caving method is somewhat similar to block caving
except that it is adaptable to smaller more irregular deposits
and to softer, stickier ore. The ore is mined downward from a
series of subleveis, using fan blasts to break the ore. Since
only the subleveis must be kept open, the method is applicable to
heavy ground. The capping must cave easily but should not break
fine in comparison to the ore or excessive dilution may result.
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Placer Mining Operations
Placer deposits consist of alluvial gravels or beach sands
containing valuable heavy minerals. In Alaska and parts of the
Northwestern U.S., placer deposits are mined for gold (minor
amounts of platinum, tin and tungsten may also be recovered).
Two basic mining methods are employed. The most widely used
method is the use of heavy earth moving equipment such as bull-
dozers, front-end loaders and backhoes to push or carry the pay
gravels to a sluicebox. Generally, either a backhoe is used to
load the gravels into the sluicebox or the sluicebox is situated
such that the gravel can be pushed directly into its head-end by
a bulldozer. The same earth moving equipment is used to strip
overburden, when required.
At a few sites in Alaska, bucket-line dredges are still used to
mine gold from placers. Prior to dredging, the frozen gravels
are thawed by circulating water through 3.8 cm (1.5 inches) pipes
contained in drill holes spaced on 4.9 meter (16 foot) centers
and drilled to bedrock. Thawed gravels are dredged with a chain
of buckets which dump their contents into a hopper on the dredge.
Titanium minerals contained in sand deposits in New Jersey and
ancient .beach placers in Florida are also mined by dredging
methods. In these operations, a pond is constructed above the
ore body, and a dredge is floated on the pond. The dredges
currently used are normally equipped with suction head cutters to
mine the mineral sands.
Solution Mining
In situ or solution mining techniques are used in some parts of
Arizona, Nevada and New Mexico to recover copper and in Texas to
recover uranium. In situ mining involves leaching the desired
metal from mineralized ground in place. During in situ leaching,
the ore body must be penetrated and permeated by the leaching
solution, which must flow through the mineralized zone and then
be recovered for processing at the surface. An impermeable
underlying bed, such as shale or mudstone, is desirable to pre-
vent downward flow below the ore zone. Usually, in the solution
mining of copper, abandoned underground ore bodies previously
mined by block caving methods are leached. Although, in at least
one case, an ore body on the surface of a mountain was leached
after shattering the rock by blasting. In underground workings,
leach solution (dilute sulfuric acid or acid ferric sulfate) is
delivered by sprays, or other means, to the upper areas of the
mine and allowed to seep slowly to the lower levels from which
the solution is pumped to a precipitation plant at the surface.
Solution mining of uranium generally involves the leaching of
previously unmined, low-grade ore bodies. Injection and pro-
duction wells are drilled through the mineralized zone, the
drilling density depending on the nature of the body. The ore
33
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body may be fractured to improve permeability and ieachab.i. 1 ity.
The leaching solution is generally a dilute acid or carbonate.
An oxidant, such as sodium chlorate, may be added to improve
leaching, and a flocculant may improve flow.
INDUSTRY PRACTICE
The processes discussed above are used variously throughout the
ore mining and milling category. A profile of the mining and
milling practices used in each subcategory follows.
Iron Ore_ Subcategory
General
American icon ore shipments increased from 1968 to 1973. In
1973, the United States shipped 92,296,400 metric tons
(101,738,320 short tons) of iron ore. Shipments declined to a
level of 76,897,300 metric tons (84,763,893 short tons) in 1975
and leveled off in }976 to 77,957,500 metric tons (85,932,552
short tons) (Reference 4). Iron ore shipments decreased to
54,915,000 metric tons (60,539,000 short tons) in 1977, Ship-
ments increased to 84,538,000 metric tons (93,191,000 short tons)
in 1978 and to 87,597,000 metric tons (96,564,000 short tons) in
1979 (Reference 5). The general trend in the iron ore industry
is to produce increasing amounts of pellets and less "run of
mine" quantities (coarse, fines, and sinter). Total pellet
production in 1976 was 68,853,800 metric tons (75,897,543 short
tons), or 38.3~ percent of all iron ore shipped, whereas only 70
percent of all iron ore shipped in 1973 was in the form of
pellets.
Based on production figures, 54 percent of the U.S. iron ore
industry uses milling operacions which result in no discharge, 31
percent discharge to surface waters, and the discharge practices
for 15 percent are unknown. For palletizing operations alone, 56
percent of total production is represented by operations prac-
ticing no discharge of process wastewater, 35 percent discharge
to surface waters, and the discharge practices of 8 percent are
unknown. A summary presented in Tables III-l and II1-2 shows
production data, processes and wastewater technology employed,
and discharge methods and volumes.
Unlike the milling segment, the mining segment of the iron ore
industry does discharge, either directly to the environment or
into the mill water circuit, either as the primary source of pro-
cess water or as makeup water. Water can cause a variety of
problems if allowed to collect in mine workings. Therefore,
water is collected and p'jmped out of the mine.
rp
The primary discharge water treatment used in mining and mining/
milling operations is removal of. suspended solids by settling, A
34
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single facility u~-.es alum and a long-chain polymer as
flocculation aids for fine~grained suspended solids.
In 1978, one facility (Mill 1113) which formerly discharged was
expected to achieve no discharge of process wastewater. This
facility accounted for approximately 13 percent of the total U.S.
production of iron ore pelleto, In 1978, approximately 69 per-
cent of iron ore pellet production will come from zero-discharge
facilities, and zero-discharge for the Mesabi Range subcategory
is required for BPT.
Recent Trends
A new technology to obtain an acceptable iron ore product has
been developed recently and is currently being used at Mill 1120.
If successful, it could result in a shift from the current trend
of mining magnetic taconite ores to the mining of fine-grained
hematite ores. Due to the fine-grained nature cf these ores (85
percent is less than 25 micrometers (0.001 inch)}, very fine
grinding is necessary to liberate the desired mineral. Conven-
tional flotation techniques used for co -rse-grained hematite ores
have proven unsuccessful because of tl-e slimes developed by the
fine-grinding process.
A selective flocculation technique has been developed that
reduces the slimes which are so detrimental. In this process,
the iron minerals are flocculated selectively from a starch
product while the siliceous slimes are dispersed using sodium
silicate at the proper pH. After desliming, cationic flotation
is used, incorporating an amine for final upgrading. A simpli-
fied flow sequence for liberation of the fine-grained hematites
is illustrated in Figure A-4 of Appendix A. Careful control of
water hardness is necessary for the process to function properly.
Recycled water is lime-treated to create a water-softening
reaction.
Copper7 Lead, Zinc, Gold, Silver, Platinum, and Molybdenum
This subcategory includes many types of ore metals which are
milled by similar processes and which have similar wastewaters.
Copper Mining
Based on the profile of copper mines shown in Table III-3, there
are presently 33 operations engaged in the mining of copper.
This listing includes those operations whose status is active,
exploratory, or under development. The vast majority of the
mines listed in the tables are located in Arizona, while the
others are located in seven other states. In addition to the
mines listed in Table III-3, a recent MSHA (Mine Safety and
Health Administration) tabulation indicates that 22 smaller
operations with an average employment of about 10 people are
35
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presently engaged in copper mining. However, production, process
and water use information for these small mines is not available.
The tabulation below provides a production cross section of the
major copper mining states in 1976 (Reference 6):
State
Arizona
Utah
New Mexico
Montana
Nevada
Michigan
Tennessee
Idaho
Production
1000 Metric Tons 1000 Short Tons
157,339
26,817
22,690
15,220
7,092
3,448
1 ,845
128
173,472
29,567
25,016
16,781
7,820
3,801
2,034
141
The total domestic copper mine production from 1968 to 1979 is
shown below (1968-1973 production - Reference 7, 1974-1976
production - Reference 6, 1977-1979 production - Reference 5):
Year
1968
1969
1970
1971
1972
1973
1974
1975
1976
1977
1978
1979
Production
1000 Metric Tons 1000 Short Tons
154,239
202,943
233,760
220,089
242,016
263,088
266,153
238,544
257,349
235,844
239,247
264,790
170,054
223,752
257,729
242,656
266,831
290,000
293,443
263,003
283,736
259,973
263,724
291,881
As shown in Table III-3, 19 operations employ surface mining
methods, while 10 operations mine underground. Four operations
use both methods. The U.S. Bureau of Mines reports that 84
percent of the copper and 90 percent of the copper ore was
produced from open-pit mines in 1976.
Water handling practices at most mines result in the use of mine
water as makeup for leaching or milling circuits. However, mine
water is discharged to surface waters (direct discharge) at as a
result of "dormant" operational status. Mine water discharge
practices at seven operations are unknown because of exploratory
or development status.
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Many western copper mines use leaching operations. Leaching
operations currently employ sulfuric acid (5 to 1C percent) or
iron sulfate to dissolve copper from the oxide or mixed oxide-
sulfide ores in dumps, heaps, vats or in situ. The copper is
subsequently recovered from solution in a highly pure form via
precipitation, electrolytic deposition (electrowinning), or
solvent extraction-electrowinning. Production of cement and
electrowon copper contributes a significant quantity (17.6
percent in 1976) of the recoverable copper produced through
ni Hi rig.
Since leaching circuits require makeup water, total recycle of
leach circuit water is common practice. Therefore?, the BPT
effluent limitations guidelines for mines and mills which employ
dump, heap, in situ, or vat leach processes for the extraction of
copper from ores or ore waste was no discharge of process waste-
water. A clarification of the limitations was also promulgated
(44 FR 7953) which addressed the intent of the limitation, the
applicability of relief from effluent limitations, and defined
areas of coverage.
Copper Milling
Nearly all copper mines are associated with mills, as seen in the
copper mill profile (Table 111-4}. Froth flotation, a process
designed for the extraction of copper minerals from sulfide ores,
is the predominant ore beneficiation technique. The ore is
crushed and ground to a suitable mesh size and is sent through
flotation cells. Copper sulfide concentrate is lifted in the
froth from the crushed material and collected, thickened and
filtered. The final concentrate from the mill may contain 15 to
30 percent copper.
Many metal byproducts are claimed, from the copper concentrates
produced at mills/ however, most byproduct values are realized at
smelters and refineries. A major byproduct associated with the
copper mills is molybdenum concentrate, and molybdenum containing
byproducts were reported from 14 of the mills in 1971.
The final concentrate from the mill is sent to the smelter for
production of blister copper (98 percent Cu). The refinery pro-
duces pure copper (99.88 to 99.9 percent Cu) from the blister
copper, which retains impurities such as gold, silver, antimony,
lead, arsenic, molybdenum, selenium, tellurium, and iron. These
impurities are removed at the refinery.
Milling wastewater handling practices differ throughout the
industry, but most operations recycle mill water due to a nega-
tive water balance (net evaporation). Only two milling opera-
tions, both in the East., recycle a minor portion or none of the
milling wastewater. Four ore beneficiation operations practice
discharge to surface waters. One operation combines mine, leach
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and mill wastewater and reports a discharge. At least one
milling operation is reported to discharge intermittently.
Lead/Zinc
Lead and zinc are often found in the same ore and as such, are
discussed together throughout this document. The domestic lead/
zinc mining industry presently consists of 29 mining operations
(which may have more than one mine) and 23 concentrators located
in 12 states. Three of these mines and four of the concentrators
have begun production since 1974. This increased production
capability has been offset somewhat by the closing of seven mines
and six concentrators during the same time period.
U.S. mine production of lead in 1975 dropped six percent to
565,000 metric tons (621,500 short tons) from a record high of
603,500 metric tons (663,900 short tons) achieved in 1974
(Reference 6). Lead production has continued to decline. Pro-
duction in 1977 totaled 537,499 metric tons (592,500 short tons)
(Reference 5). Production of lead continued to decline and in
1979 production totaled 525,569 metric tons (579,300 short tons)
(Reference 5). Missouri remained the leading producer, with 89.8
percent of the nation's total lead production, followed by Idaho,
Colorado, and the other states. The seven leading mines, all in
Missouri, contributed 79 percent of the total U.S. mine produc-
tion, and the 12 leading mines produced 91 percent of the total
(Reference 8). Although lead production declined between 1974
and 1979, domestic lead prices- continued to rise from a price of
47.4 cents per kilogram (21.5 cents per pound) in 1975 to $1.16
per kilogram (52.6 cents per pound) in 1979. The current price
(15 September 1981) is 93 cents per kilogram (42 cents per pound)
(Reference 9).
Domestic mine production of zinc decreased from 454,500 metric
tons (499,900 short tons) in 1974 to 426,700 metric tons (469,400
short tons) in 1975. It then recovered to about 440,000 metric
tons (484,500 short tons) in 1976 (Reference 5). Zinc production
decreased to 407,900 metric tons (449,600 short tons) in 1977 and
further to 302,700 metric tons (333,700 short tons) in 1978 and
267,300 metric tons (294,600 short tons) in 1979 (Reference 5).
Tennessee was the leading zinc producing state in 1979 with 32
percent of total production and was followed by Missouri, 23 per-
cent; New Jersey, 12 percent; and Idaho, 11 percent (Reference
5). Tennessee led the nation in production of zinc for 15 con-
secutive years prior to 1973 and regained that status in 1975,
1977, 1978, and 1979 due to the opening of two new concentra-
tions.
In contrast to climbing lead prices, zinc prices have followed a
downward trend over the past decade in terms of real dollars.
Following inflationary price increases during the period 1972 to
1975, zinc prices declined from the average 1975 price of 85.91
cents per kilogram (38.96 cents per pound) to 81.61 cents per
38
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kilogram (37.01 cents per pound) in 1976 to 75.85 cents per kilo-
gram (34,4 cent per pound) in 1977 and then increased to the
current (15 September 1981) price of $1.09 per kilogram (49.25
cents per pound) (Reference 9),
The domestic mine production of lead/zinc continues to come
chiefly from ores mined primarily for their lead and zinc con-
tent. Less than 1 percent of the total lead/zinc production is
derived as byproduct or coproduct of ores mined for copper, gold,
silver, or fluorspar (Reference 10). The complex ores of the
Rocky Mountain area are particularly dependent on economic
extraction of the mineral's aggregate metal value, and the metal
of highest value is variable. Byproduct recovery at the smelter
or refinery from the processing of lead and zinc concentrates
also yields a significant portion of the domestic production of
antimony, bismuth, tellurium, and cadmium.
Lead and zinc ores are produced almost exclusively from under-
ground mines. Several mines began as open-pit operations and
have developed into underground mines. Conversely, a number of
underground mines have surfaced while following an ore body,
resulting in small, open-pit operations. At present, only one
open-pit, mine is in operation, and it is actually the intersec-
tion of an underground lead/zinc ore body with an adjacent open-
pit copper mine.
A general description of the lead/zinc mining industry is con-
tained in Tables III-5 to III-8 and includes processes, products,
location, age, wastewater treatment technology, and discharge
method and volume (see also References 11 and 12). As previously
indicated, many of the mines and mills shown as lead/zinc also
mine or mill for other metal values that are interspersed within
the lead/zinc ore matrix. These metal values are usually copper,
gold, or silver. However, the mines and mills shown as lead/zinc
are characterized based on their primary products, lead and zinc.
Gold
The four leading U.S. gold produc
total production during 1975.
production came from « 5 mines or
which were operated primarily for
Thirty-six percent of total pror"
duct of other mining -for exampi
primary metals); tht remainder-
placer operations.
ers accounted for 73 percent of
Approximately 95 percent of all
mine/mill operations, 10 of
recovery of gold (Reference 8).
uction was recovered as a bypro-
, where copper or lead/zinc are
was recovered at gold lode and
Gold prices have riser significantly in the past year. The price
of gold on the open market reached a previous high of nearly $200
per 31.1 gram (1 troy ounce) during 1974; in January 1980 the
price was over $80^.00 (Refe-
ounce). It is currently (15 Mar-
troy ounce) (Ref ere.-'ce 9),
ence 13) per 31.1 gram (1 troy
ft 1982) $315 per 31.1 gram (1
.1 expected response to the high
-------
price of gold is the increased gold prospecting arid development
activities reported in gold mining areas of the country
(Reference 14). In addition, plans for significant investment in
new production facilities or renovation of inactive facilities
have been announced by a number of mining companies during the
past 3 to 4 years. During 1975 and 1976, four to six new
cyanidation operations began full scale operation. However,
despite these prospects, domestic production has continued to
decline during recent years. Reported domestic production for
1975 was 32.7 metric tons (1,052,000 troy ounces), a decline of
27 percent from reported production of 45.1 metric tons
(1,450,000 troy ounces) in 1972.
The steady decline in domestic production of gold is due to
several factors: (1) inflation and shortages of equipment, mate-
rial, and labor have limited new mine developments; (2) in most
instances, lower grade ores are being mined, but mine and mill
limitations have generally allowed little expansion of tonnage
handled; (3) diminished copper production due to low copper
prices in 1977 to 1978 led to a decline in byproduct production
of gold; and (4) depletion of ore at two major producing gold
lode operations has resulted in the suspension of all production
at one during October 1977, while the second is scheduled for
permanent closure. These two mine/mill operations accounted for
18 percent of reported domestic primary gold production in 1974.
A summary description of gold mine/mill operations is presented
in Tables 111-9 to 111-12. As indicated, most operations employ
the cyanidation process for recovery of gold (see description
under Ore Beneficiation Processes) and Appendix A for specific
applications). This is especially true of lode mining operations
which have recently become active and are located predominantly
in Nevada. At these sites, heap leaching or agitation leaching
processes have been the methods of choice. In addition, the
preferred process for recovery of the gold from solution has been
the recently developed carbon-in-pulp process (see Appendix A).
The simplicity of its operation and the low capital and operating
costs have made this process economically superior to the
conventional zinc-precipitation process. This factor and the
current high selling price of gold have served to make
development and mining of some small or low-grade gold ore bodies
economically feasible.
Spent leach solutions are recycled at cyanidation leaching
operations. This practice has generally been implemented for
conservation of both reagent and process water.
Of the lode mill operations operating for the primary recovery of
gold, two report discharge of wastewater. One of these is
building facilities which will provide the equivalent of zero
discharge of mill wastewater. This mill uses the cyanidation
process and under the BPT regulation, zero discharge of process
wastewater is required.
40
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The second lode mill operation which discharges wastewater
recovers gold by amalgamation and lead/zinc by flotation. The
alkaline wastewater which results is settled in a multiple pond
system prior to final discharge.
The exact number of lode gold mines which have a discharge and
are not directly associated with a mill is not known. Treatment
of wastewater at placer mining operations is often not practiced.
At the large dredging operations and at the smaller hydraulic and
mechanical excavation operations, settling ponds have been
provided.
Silver
Domestic mine production of silver in 1975 totaled 1,085.4 metric
tons (34.9 million troy ounces). The largest percentage of this
production continued to be a byproduct of base-metal mineral
mining. During 1974, only three of the 25 leading producers
mined ore primarily for its silver content. These three were the
first, second, and eighth leading producers and accounted for
approximately 30 percent of total domestic primary production.
Production at a fourth major silver mine was curtailed during
1973 due to depletion of known ore reserves. Two recently
developed silver mine/mill operations began operation during 1976
and are expected to become major producers.
The selling price of silver, like that of gold, reached an all-
time high during 1979. During January 1980, the selling price
reached $48.00 (Reference 13) per 31.1 grams (1 troy ounce). The
current price of silver (15 March 1982) is $7.20 per 31.1 grams
(1 troy ounce) (Reference 9).
A summary description of silver mine/mill operations is presented
in Tables 111-13 and 111-14. More than 300 U.S. mines supply ore
from which silver is recovered. However, as previously stated,
most of this ore is exploited primarily for its copper, lead, or
zinc content. Byproduct silver is typically recovered from base-
metal flotation concentrates during smelting and refining
processes. The large operations which are exploiting ore pri-
marily for its silver content also recover the sulfide minerals
by flotation. Only 3 small fraction of a percent of total silver
production is recovered from placer mining or by amalgamation.
Approximately 1 percent is recovered by the cyanidation process.
Wastewater treatment practices at major silver mine/mill
operations typically employ a pond for collection and retention
of the bulk flotation-circuit tailings. In some instances,
multiple pond systems are employed to optimize control of
suspended solids. At. one of the four flotation mills operating
during 1977, clarified decant was recycled from the tailing pond
to the mill. Partial recycle is practiced at a second operation.
Two mills situated in a river valley in Idaho presently discharge
to a common tailing pond system.
41
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A major silver producing operation which opened in 1976 is
employing a vat leach (cyanidation), zinc-precipitation circuit.
Wastewater generated in this process is impounded and recycled
for reuse within the mill.
At one operation, mine drainage is settled in a multiple pond
system prior to final discharge and at a second operation, mine-
water is directly discharged without treatment. Minewater at all
other silver mining operations (for which information is avail-
able) is either used as makeup in the mill or is discharged into
the mill tailing pond treatment, system.
Platinum
The platinum-group mining and milling industry includes those
operations which are involved in the mining and/or milling of ore
for the primary or byproduct recovery of platinum, palladium,
iridium, osmium, rhodium, and ruthenium. Domestic production of
these platinum-group metals results as a oyproduct of copper
refining. Until recently, production from a single placer opera-
tion in Alaska accounted for all the U.S. production from mining
primary ores. In 1976, this facility, located in the Goodnews
Bay District of Alaska, ceased operation. Total "mine" produc-
tion of platinum-group metals in 1978 was 258.2 kilograms (8,303
troy ounces), which was recovered entirely as a byproduct of
copper refining. Of this, approximately 3?.1 kilograms (1,258
troy ounces) represented platinum metal itself. Total secondary
recovery of platinum-group metals (from scrap) was 7,998.6 kilo-
grams (257,191 troy ounces) in 1978.
Platinum-group metals are recovered as secondary metals in many
places within the United States. Minor amounts have been
recovered from gold placers in California, Oregon, Washington,
Montana, Idaho, and Alaska, but significant amounts as primary
deposits have been produced only from the Goodnews Bay area in
Alaska.
In 1976, the single platinum mining operation employed techniques
similar to those used for recovery of gold from placer deposits,
i.e., bucketline dredging. The coarse, gravelly ore was
screened, jigged and tabled. Chromite and the magnetite were
removed by magnetic separation techniques. After drying, air
separation techniques were applied, and a 9U percent concentrate
was obtained. Water used for the initial processing was dis-
charged from the dredge into a settling pond and subsequently
discharged from the pond after passing through coarse tailings.
Removal of total suspended solids to below 30 mg/1 was
accomplished.
In the United States, the major part of platinum production has
always been recovered as a byproduct of copper refining in
Maryland, New Jersey, Texas, Utah and Washington. Byproduct
platinum-group metals are sometimes refined oy electrolysis and
42
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chemical means to yield recoveries of over 99 percent. The price
of platinum as quoted recently (15 September 1981) on the spot
market was $475 per 31.1 grams (1 troy ounce) (Reference 9).
Molybdenum
Production of molybdenum has been, generally, increasing over the
past 30 years as illustrated below (References 5, 7, 15, 16, 17):
Production
Year Metric Tons Short Tons
1949 10,222 11,265
1953 25,973 28,622
1958 18,634 20,535
1962 23,250 25,622
1968 42,423 46,750
1972 46,368 51,098
1974 51,000 56,000
1977 55,484 61,204
1979 65,302 71,984
Since 1974 significant exploration and development has taken
place, and production is expected to increase at a higher rate.
Production figures are not available yet, but several new opera-
tions have begun since 1976, and a number of mines and mills are
in the planning and development stages.
As shown in Table 111-15, there are six mines involved with
molybdenum; three mines are producing now, and three have
exploration underway. The three producing mines are: one open-
pit in New Mexico, an underground and a combined open-
pit/underground in Colorado. One Colorado operation recovers
secondary products of tungsten and tin.
The two Colorado mines discharge to surface waters; one by way of
a mill water treatment system, the other by way of separate
treatment. The New Mexico mine has no discharge because
groundwater is not encountered.
All three active mines are associated with flotation mills which
are described in Table 111-16. The New Mexico facility treats
mill water for discharge by primary and secondary settling (two
tailings and one settling pond), and aeration for cyanide
removal. They are currently experimenting with hydrogen peroxide
for cyanide oxidation. The underground/open-pit Colorado opera-
tion uses a complex treatment system including settling, recycle,
ion exchange, electrocoagulation flotation, alkaline chlorination
and mixed-media filtration. The second Colorado mill accomp-
lishes no discharge by total recycle.
Aluminum
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In 1979, the major aluminum ore, bauxite, was mined by eight
companies located in Arkansas, Alabama, and Georgia (Reference
5). Arkansas accounted for 79 percent of the total bauxite mined
in 1979 (Reference 5). The only operations mining bauxite for
aluminum production are two operations located in Arkansas. In
both BPT and BAT effluent guidelines, the aluminum ore subcate-
gory applies only to the mining of bauxite for eventual metallur-
gical production of aluminum. Most bauxite mined at the two
Arkansas operations is refined to alumina (A1203_) by the
"Combination Process," which is classified as SIC 2819. A
gallium byproduct recovery operation is used at one Arkansas
operation. Domestic production of bauxite is listed below from
1974 to 1979 (References 5 and 18):
Production
Year Metric Tons Short Tons
1974 1,980 2,181
1975 1,800 1,983
1976 1,989 2,191
1977 2,013 2,217
1978 1,669 1,839
1979 1,821 2,006
Average annual production for the last 10 years is approximately
1,880,000 metric tons (2,070,000 short tons).
All production from the two domestic aluminum ore operations
originated from open pits. The sole underground bauxite mine
closed in late 1976. Bauxite ore used for refining alumina is
graded on silica content, and the percentage of domestic bauxite
shipments by silica content is listed below (Reference 5):
Silica (Si02_) Percentage of Total Domestic Shipments
Content (%) of Ore 1975 1976 1977 1978 1979
less than 8 46221
8 to 15 62 50 54 55 55
greater than 15 34 44 44 43 44
A pilot project proved the economic viability of alunite as an
alternate ore for production of alumina, but construction of a
commercial scale refinery in Utah has not begun. The mine and
refinery complex was expected to produce alumina, potassium
sulfate, and sulfuric acid when completed (Reference 19).
Both domestic bauxite ore operations require discharge of large
volumes of mine water, and there is no process water used for
crushing or grinding of the ore. The total daily discharge of
mine water attributable to bauxite ore mining is about 40,000
cubic meters (10.6 million gallons), with about 80 percent of the
discharge flow attributed to a single operation. Characteristics
of the two domestic bauxite operations are shown in Table II1-17.
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Tungsten
Tungsten mining and milling is conducted by numerous (probably
more than 50) facilities, the majority of which are very small
and operate intermittently. As illustrated below, tungsten
production was relatively constant between 1969 and 1979
(References 5, 18, and 20):
Domestic Production (Contained Tungsten)
Year Metric Tons Short Tons
1969 3,543 3,903
1970 4,369 4,813
1971 3,132 3,450
1972 3,699 4,075
1973 3,438 3,787
1974 3,350 3,690
1975 2,536 2,794
1976 2,646 2,915
1977 2,727 3,004
1978 3,130 3,448
1979 3,015 3,321
A profile of tungsten mining is presented in Tables 111-18 and
111-19. Table 111-18 describes the larger operations (annual
production greater than 5,000 metric tons ore processed/year),
and Table 111-19 describes smaller operations (production less
than 5,000 metric tons ore processed/year). The majority of the
mines are located in the western states of California, Oregon,
Idaho, Utah, and Nevada. Almost all are underground mines, and
many have no discharge of mine water.
Tables 111-20 and 111-21 profile tungsten mills. Processes used
are gravity separation and/or fatty acid and sulfide flotation.
One mill in California produces the majority of the tungsten
concentrate. Wastewater treatment methods vary, but may include
settling (tailing ponds) and recycle and/or evaporation. Most of
the active mills do not discharge, primarily because they are in
arid regions and need the water.
Mercury
During recent years, the domestic mercury industry has been
characterized by a general downward trend in the number of
actively producing mines or mine/mill operations. Historically,
mine output in the United States has come from a relatively large
number of small production operations. However, since 1969, the
number of actively producing mines has declined from 109 to just
4 in 1980 (Reference 19).
Domestic primary production during 1980 was 30,657 flasks (34.5
kilograms (76 pounds) per flask) (Reference 21). More than 75
percent was produced by a single mine/mill in Nevada which began
operation during 1975. Byproduct recovery was reported at a gold
45
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mining operation in Nevada. The total domestic production of
mercury during 1980 came from two mines in California and two
mines in Nevada.
The primary factor contributing to the recent depression of the
domestic mercury industry was a steady decline in the selling
price of mercury during the late 1960's and early 1970's.
Between 1968 and 1975, the selling price decreased to an average
of $117 per flask in New York. However, as of February 1982, the
price had risen to about $400 per flask. Additional factors
having an adverse impact include: (1) widely fluctuating prices
caused by erratic demand, (2) competition from low-cost foreign
producers, and (3) the low grade ore resulting in high production
costs. A descriptive summary of active mercury mine and mill
operations is presented in Tables 111-22 and 111-23.
The majority of U.S.-produced mercury is recovered by a flotation
process at one mill in Nevada. Ore processed in that mill is
mined from a nearby open pit. The flotation concentrate produced
is furnaced on site to recover elemental mercury. Wastewater
treatment consists of impoundment in a multiple pond system with
no resulting discharge. The majority of impounded wastewater
evaporates, although a small volume of clarified decant is
occasionally recycled.
A second operation, located in California, employs a gravity
concentration process. Ore is obtained from an open-pit mine and
the concentrate is furnaced on site to produce elemental mercury.
Wastewater is settled and recycled during the 9 months of the
year that the mill is active. During the remaining 3 months
(winter months), however, the mill is inactive, and a mine water
discharge from the settling pond often occurs as a result of
rainfall and runoff. This facility is presently inactive.
An additional number of small operations, located in California
and Nevada, operate intermittently. Ore is generally furnaced
directly without prior beneficiation. Water is not used except
for cooling in the furnace process.
Uranium
This category includes facilities which mine primarily for the
recovery of uranium, but vanadium and radium are frequently found
in the same ore body. Uranium is mined chiefly for use in gener-
ating energy and isotopes in nuclear reactors. Where vanadium
does not occur in conjunction with uranium/radium (nonradio-
active), it is considered a ferroalloy and is discussed as a
separate subcategory. Within the past 20 years, the demand for
radium (a decay product of uranium) has vanished due to the
availability of radioactive isotopes with specific characteris-
tics. As a result, radium is now treated by the industry as a
pollutant rather than as a product.
46
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Primary deposits of uranium ores are widely distributed in
granites and pegmatites. These black ores contain the tetrava-
lent minerals uraninite (U02.) and coffinite (U(Si04)1-x(OH)4x)
with pyrite as a common gangue mineral. Secondary, tertiary, and
higher order uranium deposits are found in relatively shallow
sandstones, mudstones, and limestones. These deposits are formed
by the transport of soluble hexavalent uranyl compounds (notably
carbonates) with the composition U30£ (i.e., U02_.2U03_) being
particularly stable. Transport of the uranyl compounds leads to
the surface uranium ores commonly found in arid regions,
including carnotite (K2(U02) 2.(V04 ) 2. 1 3H20), uranophane (Ca(U02)
(Si03)2.(OH)2_. 5H20), and autunite (Ca(U02.) 2 (P04 ) 2.. 1 0-1 2H20) . If
reducing conditions are encountered, as in subsurface sedimentary
deposits, tetravalent uranium compounds are redeposited.
The major deposits of high-grade uranium ores in the United
States are located in the Colorado Plateau, the Wyoming Basins,
and the Gulf Coast Plain of Texas. In 1976, New Mexico provided
46 percent; Wyoming, 32 percent; and Utah, Colorado, and Texas,
the remaining 22 percent of total U.S. production (Reference 22).
Total domestic production of uranium for 1977 was predicted to be
almost 9,100,000 metric tons (10,000,000 short tons) of ore, with
an average grade of 0.15 percent U308^ (Reference 22). Average
ore grade is down from 0.17 percent in 1975 and 0.18 percent in
1974, reflected in increases in the price of U30£ from $77 per kg
($35 per pound) in 1975 to $92 per kg ($42 per pound) in 1977
(References 23 and 24).
Tables 111-24, 111-25 and 111-26 present profiles of the uranium
mining and milling industry in the United States. The
information presented in Table II1-24 represents over 90 percent
of the total U.S. production of uranium ore. Depending on the
nature of the operation, each listing may represent anywhere from
one to 40 individual mines. Table 111-25 presents information
for all active uranium mills in the country. Table 111-26
presents available information for in situ uranium mines.
Following is a brief description of the uranium industry. A more
detailed account of the processes, water use, wastewater
generation, and treatment in the industry may be found in
Reference 24.
Mini ;g
Mining practice in the uranium industry is by open-pit or
underground. There were approximately 160 underground mines in
the United States as of September 1977 (Reference 11), and more
than 50 percent have fewer than five employees. There are 53
surface mines in the United States, and about 26 percent of these
employ fewer than five people. The actual number of active mines
at any given time will vary, depending on market conditions and
company status.
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Most mines ship ore to the mill by truck. The economics of
hauling unbeneficiated ore require that the distance from the
mine to the mill be no more than a few kilometers (approximately
one mile). However, certain high-grade ores (0.6 percent U308_)
may be shipped up to 200 kilometers (120 miles). The large
number of small mines often requires individual mills to purchase
ore from several different mines, both company and privately
owned. A single mill may be fed by as many as 40 different
mines.
Milling
As of February 1979, there were 20 active uranium mills in the
United States, ranging in ore processing capacity from 450 metric
tons per day (500 short tons per day) to 6,300 metric tons per
day (7,000 short tons per day). In addition, four of these mills
are practicing vanadium byproduct recovery, one mill is recover-
ing molybdenum concentrate as byproduct, and another intermit-
tently recovers copper concentrate. One mill, which historically
produced vanadium from uranium ore, is currently producing
several vanadium products from vanadium concentrate shipped from
a nearby mill. A complete discussion of the milling and
extraction technology used in this subcategory is presented in
Appendix A.
Byproduct vanadium recovery is practiced at three uranium mills,
At Mill 9401, an alkaline mill, purification of crude yellowcake
by roasting with soda ash (sodium carbonate) and leaching the
calcine with water generates a vanadic acid solution. This
solution, which contains about eight percent V205_, is stockpiled
and sold for vanadium recovery elsewhere. Mill 9403, which
operates an acid-leach circuit, recovers vanadium as a solvent-
exchange raffinate. Vanadium values in the raffinate are
concentrated and recovered by solvent exchange, precipitation of
ammonium vandates (from the pregnant stripping agent) with
ammonium chloride, filtration, drying, and packaging. Mill 9405,
which also operates an acid-leach circuit, recovers vanadium from
ion exchange circuit raffinate. The raffinate is treated with
sodium chlorate, soda ash, and ammonia to precipitate impurities
and is then directed to a solvent-extraction circuit. Pregnant
stripping solution from the solvent extraction circuit,
containing the vanadium values, is collected and shipped to a
nearby facility for further processing.
In addition to uranium and vanadium, Mill 9403 intermittently
recovers copper concentrate from uranium ore high in copper
values. Copper recovery includes a sulfuric acid leach, which
generates a pregnant liquor containing dissolved uranium as well
as copper. The dissolved uranium is recovered in a solvent
exchange circuit and then directed to the main plant for final
processing and recovery as yellowcake.
In Situ Recovery
48
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Eight operations in Southern Texas are practicing in situ
leaching of uranium, and seven more in situ leaching operations
are under development. Annual production from six of the eight
on-line facilities is estimated to be 671 metric tons (740 short
tons) of U308. (Reference 25). Typically, alkaline leach
solutions are pumped into a series of strategically placed
injection wells, recovered from a production well, and either
shipped to a nearby mill or recovered on site. The uranium is
concentrated using fixed bed ion exchange and conventional
yellowcake precipitation techniques.
Industry Trends
Although many uncertainties exist about the future use of nuclear
energy in this country, increases in yellowcake requirements are
expected within the next 10 years. The annual demand for
yellowcake is expected to grow from approximately 15,400 metric
tons (17,000 short tons) of U30£ in 1976 to 35,400 metric tons
(39,000 short tons) in 1985. As a result, the decreasing grade
of mined ore will require U.S. mill capacity to increase from
8,890,000 metric tons (9,800,000 short tons) of ore per year to
46,200,000 metric tons (50,800,000 short tons) of ore per year
(Reference 21). Because of recent developments, however, the
Nuclear Regulatory Commission estimates in 1979 indicated that
demand for mill capacity may be in the range of 22 to 27 million
metric tons in 1987 and approximately 36.5 million metric tons in
the year 2000 (Reference 26).
Projected increases in U3_08_ demand have resulted in:
(1) the exploration and expansion of known sandstone deposits;
(2) the exploration of new sandstone areas in Nevada, North and
South Dakota, Colorado, Wyoming, and Montana; and (3)
preliminary investigation of "hardrock" areas in Colorado,
Michigan, Wisconsin, Minnesota, the eastern and southwestern
United States, and Alaska.
Near-term industry growth patterns include three new uranium
mills, rated at 317,000 metric tons (350,000 short tons) per
year, 634,000 metric tons (700,000 short tons) per year, and
680,000 metric tons (750,000 short tons) of ore per year, and
sevev new in situ leach operations.
Increases in the demand and price of yellowcake will provide
added impetus for the extraction of lower grade ores, the refine-
ment of conventional milling processes, the development of new
mining and milling techniques (e.g., hydraulic mining of sand-
stones and new milling processes for hardrock ores), and the
development and expansion of nonconventional uranium sources such
as in situ leaching and phosphate byproduct recovery. Thus,
within a few short years, the nature of the uranium mining and
milling industry in this country may significantly change.
49
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Water Use and Wastewater Generation
Uranium ores are often found in arid climates, thus water is
conserved in milling uranium. Approximately 50 percent of the
total U.S. production of uranium ore is recovered from mines
which generate mine water. Mine water generation varies from
1.5 cubic meters (390 gallons) per day to 19,000 cubic meters
(5,000,000 gallons) per day. Some mines yield an adequate water
supply for the associated mill. Those mines which are too far
from the mill or which produce water in excess of mill
requirements usually treat the mine water to remove pollutants
and/or uranium values. Sometimes the treated water is
reintroduced into the mine for in situ leaching of values.
The quantity of water used in milling is approximately equal in
weight to that of the ore processed. Mills obtain process water
from nearby mines, wells and streams. The quantity of makeup
water required depends on the amount of recycle practiced, and on
evaporation and seepage losses. Eight of the 14 acid mills and
three of the four alkaline mills employ at least partial recycle
of mill tailing water. The remaining mills employ impoundment
and solar evaporation. Acid and alkaline mills, i.e., acid
leach, alkaline leach, are explained in detail in Appendix A.
Mine water treatment practices in the uranium industry include:
(1) impoundment and solar evaporation, i.e., evaporation by
exposure to the sun, (2) uranium recovery by ion exchange, (3)
flocculation and settling for heavy metal and suspended solids
removal, (4) BaCliZ coprecipitation of radium 226, and (5) radium
226 removal by ion exchange. Discharge is usually to surface
waters, which frequently have variable flows depending on
seasonal weather conditions.
All uranium mills in the United States impound tailings, which
are the primary source of process wastewater in large ponds.
Evaporation, seepage, and/or recycle from these ponds eliminate
all discharges. One acid mill, however, collects seepage from
its tailing pond and overflow from yellowcake precipitation
thickeners. This mill then treats the combined waste streams
(approximately 2,200 cubic meters (580,000 gallons) per day) to
remove radium 226 and total suspended solids (TSS) and discharges
to a nearby stream. This facility represents the only known
discharging uranium mill in the country.
Antimony
Antimony is recovered from antimony ore (stibnite) and as a
byproduct of silver and lead concentrates. This industry is con-
centrated in two states: Idaho and Montana. Currently, only one
operation (Mine/Mill 9901) recovers only antimony ore. The ore
is mined underground and concentrates are obtained by the froth
flotation process. There is no discharge from the mine, but
50
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wastewater from the mill flows to an impoundment. No discharge
of process wastewater to surface waters occurs.
A second facility, Mine/Mill 4403 recovers antimony as a
byproduct from tetrahedrite, a complex silver-copper-antimony
sulfide mineral. The antimony is recovered from tetrahedrite
concentrates in an electrolytic extraction plant operated by one
of the silver mining companies in the Coeur d'Alene district of
Idaho.
Antimony is also contained in lead concentrates and is recovered
as a byproduct at lead smelters usually as antimonial lead. This
source may represent about 30 to 50 percent of domestic
production in recent years.
In 1979, total U.S. mine production of antimony was 655 metric
tons (722 short tons). Production at facility 9901 in 1979 was
271 metric tons (299 short tons) of antimony, while production at
mine/mill 4403 was 384 metric tons (423 short tons) (Reference
5). In 1979, the total domestic mine production of antimony
concentrate was reported as 2,990 metric tons (3,294 short tons).
This concentrate contained 655 metric tons (722 short tons) of
antimony. Mine/mill 9901 is profiled in Table 111-27.
Titanium
The principal mineral sources of titanium are ilmenite (FeTi03_)
and rutile (Ti02_). Rutile associated with ilmenite in domestic
sand deposits is not separately concentrated typically. The
majority of all ilmenite concentrates (includes a mixed product
containing ilmenite, rutile, leucoxine and altered ilmenite)
produced domestically are from titanium dredging operations. The
remainder of the domestic production comes from a mine in New
York mining an ilmenite ore.
During recent years, domestic production of ilmenite concentrates
has substantially declined. U.S. production of ilmenite during
1968 was 887,508 metric tons (978,509 short tons), while five
years later in 1973 production had dropped to 703,844 metric tons
(776,013 short tons). Domestic production had dropped to 534,904
metric tons (589,751 short tons) in 1978. The production of
ilmenite in the U.S. has declined approximately 40 percent (39.73
percent) between 1968 and 1978. The price of domestically
produced ilmenite during early 1973 was approximately $22.64 per
metric ton (approximately $23 per long ton) and rose to $54.13
per metric ton ($55 per long ton) by July 1974. The selling
price of domestically produced ilmenite has essentially remained
at $54.13 per metric ton ($55 per long ton) since 1974, to the
present (early 1980). The selling price of domestically produced
ilmenite is not significant since the U.S. titanium industry is
nearly fully integrated, and most ilmenite concentrates are
consumed captively.
51
-------
A summary description of titanium mine/mill operations is pre-
sented in Table 111-28 and 111-29. As indicated, three of the
four active operations employ floating dredges to mine beach-sand
placer deposits of ilmenite located in New Jersey and Florida.
At these operations, concentration of the heavy titanium minerals
is accomplished by wet gravity and dry electrostatic and magnetic
methods (see Reference 1 for detailed process description). At
the remaining operation, located in New York, ilmenite is mined
from a hardrock, lode deposit by open-pit methods. A flotation
process is employed in the mill to concentrate the ore materials.
Wastewater treatment practices employed at titanium mine/mill
operations are designed primarily for removal of suspended solids
and adjustment of pH. In addition, peculiar to the beach sand
dredging operations in Florida is the presence of silts and
organic substances (humic acids, tannic acids, etc.) in these
placer deposits. During dredging operations, this colloidal
material becomes suspended, giving the water a deep "tea" color.
Methods employed for the removal of this material from water are
coagulation with either sulfuric acid or alum, followed by
multiple pond settling. Adjustment of pH is accomplished by
addition of either lime or caustic prior to final discharge.
Mine drainage from the single open-pit lode mine is settled prior
to discharge. Tailings from the flotation mill in which ore from
this mine is processed are collected and settled in an old mining
pit. Clarified decant from this pit is recycled to the mill for
reuse. Discharge from this pit to a river occurs only seasonally
as a result of rainfall and runoff during spring months.
One of the two beach-sand operations located in New Jersey is
inactive at present. Recycle of all wastewater for reuse was
practiced; consequently, no discharge occurred at this site.
Nickel
A relatively small amount of nickel is mined domestically, all
from one mine in Oregon (Mine 6106). This mine is open-pit, and
there is a mill at the site, but it only employs physical proces-
sing methods. The ore is washed and transmitted to an on-site
smelter. Mine and Mill 6106 is profiled in Table 111-30. As
shown below, production has decreased slightly from 1969 to 1980
(References 5, 18, 20, and 27):
52
-------
Production
Metric Tons Short Tons
15,483 17,056
14,464 15,933
15,465 17,036
15,309 16,864
16,587 18,272
15,086 16,618
15,421 16,987
14,951 16,469
13,024 14,347
12,263 13,509
13,676 15,065
13,302 14,653
Depending on the outcome of on-going exploration, nickel
production may increase in the next 5 to 10 years, and the Bureau
of Mines predicts a significant increase in production by 1985.
Nickel production is possible both from the Minnesota sulfide
ores and from West Coast laterite deposits similar to (but lower
in grade than) the deposit presently worked at Riddle, Oregon.
Both cobalt production and nickel recovery from laterite ores may
involve an increase in the use of leaching techniques.
Water used in beneficiation and smelting of nickel ore is
extensively recycled, both within the mill and from external
wastewater treatment processes. Most of the plant water is used
in the smelting operation since wet-beneficiation processes are
not practiced. Water is used for ore belt washing, for cooling,
and for slag granulation in scrubbers or ore driers. Water
recycled within the process is treated in two settling ponds
which are arranged in series. The first of these, 4.5 hectares
(11 acres) in area, receives a process water influx of 12.3 cubic
meters (3,256 gallons) per minute, of which 9.9 cubic meters
(2,, 624 gallons) per minute are returned to the process. Overflow
to the 5.7 hectare (14 acre) second pond amounts to 1.2 cubic
meters (320 gallons) per minute. This second pond also receives
runoff water from the open-pit mine site which is highly
seasonal, amounting to zero for approximately 3 months, but
reaching as high as 67,700 cubic meters (17.9 million gallons)
per day during the (winter) rainy season. The lower pond has no
surface discharge during the dry season. The inputs are balanced
by evaporation and subsurface flow to a nearby creek. A
sizeable discharge results from runoff inputs during wet weather.
Average discharge volume over the year amounts to 3,520 cubic
meters (930,000 gallons) per day.
Vanadium
This subcategory includes facilities which are engaged in the
primary recovery of vanadium from non-radioactive ore; however,
there is only one active facility in this subcategory, Mine/Mill
6107. The vanadium subcategory is profiled in Table VIII-31.
53
-------
At vanadium Mine/Mill 6107, vanadium pentoxide, V_205_, is obtained
from an open-pit mine by a complex hydrometallurgical process
involving roasting, leaching, solvent extraction, and
precipitation. The process is illustrated in Appendix A. In the
mill, a total of 6,200 cubic meters (1.6 million gallons) of
water are used in processing 1,270 metric tons (1,400 short tons)
of ore. This includes scrubber and cooling wastes and domestic
use.
Ore from the mine is ground, mixed with salt, and pelletized.
After roasting at 850 C (1562 F) to convert the vanadium values
to soluble sodium vanadate, the ore is leached and the solutions
are acidified to a pH of 2.5 to 3.5. The resulting sodium
decavanadate (Na6_V]_002_8) is concentrated by solvent extraction,
and ammonia is added to precipitate ammonium vanadate. This is
dried and calcined to yield a V2_05_ product.
The most significant effluent streams are from leaching and
solvent extraction, wet scrubbers or roasters, and ore dryers.
Together, these sources account for nearly 70 percent of the
effluent stream, and essentially all of its pollutant content.
Production of vanadium is summarized below (References 5 and 18):
Production
Year Metric Tons Short Tons
1973 3,737 4,117
1974 4,756 5,240
1975 4,731 5,213
1976 7,330 8,076
1977 6,866 7,565
1978 4,036 4,446
1979 5,302 5,841
54
-------
TABLE MM. PROFILE OF IRON MINES
MINE
1101
1102
1103
1104
1105
1106
1107
1108
1109
1110
1113
1112
1111
1114
1115
1116
1117
1118
1119
1120
LOCATION
(state)
MN
MN
MN
MN
UM
nfini
UU
ntri
Ml
Ml
Ml
PA
MN
mt\i
nnn
MN
MO
MO
Wl
UT
CA
WY
Ml
YEAR
OPENED
(original
facility) '
1957
1965
1948
1953
1967
1967
1959
1956
1964
1958
1976
1967
1955
1961
1968
1967
1946
1948
1962
1974
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
A
A
A
A
A
1
A
A
A
A
A
PRODUCT
Iron Ore
\
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
33,000,000
8,230,000
4,000,000
1,640,000
8.300.000
40,000,000
5,300,000
8.800.000
16,400,000
2,600,000
25,000,000
8,700,000
31,000,000
2,360,000
2,200,000
2,200,000
2,400.000
8,190,000
4,400.000
4,200,000
TYPE OF
MINE
OP
I
UG
(
i
3P
UG
UG
OP
i
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling
\
DISCHARGE
METHOD
To surface
To surface
None
To surface
To surface
None
To surface
To surface
To tailings pond
To tailings pond
or mill circuit
To tailings pond
or mill circuit
To surface
To surface
To surface
To surface
Seepage basin
To surface
None
To surface
To surface
DAILY
DISCHARGE
VOLUME
«n>f)
80 x 103
2.6 x 103
0
4.3 x103
48 x 103
0
13.4 x 103
15.8 x103
ne
na
0
26.5 x 103
52.9 x 103
5.4 x 103
na
0
na
0
1.7 x 103
0.7 x 103
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Coda: A active; I inactive; S seasonal; U.D. under development; EXP exploration underway.
-------
TABLE MM. PROFILE OF IRON MINES (Continued)
en
en
MINE
1121
1122
1123
1124
1125
1126
1127
1128
1129
1130
1131
1132
1133
1134
1135
1136
1137
1138
LOCATION
(state)
MN
MN
MN
MN
MN
MN
UT
NM
TX
NY
NY
WY
MN
MN
MN
Ml
CA
MN
YEAR
OPENED
(original
facility) .
1968
1973
1917
1933
1965
1965
1953
1938
1947
1942
1944
1900
1974
1974
1960
1940
1971
1976
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
A
A
A
I
A
A
I
A
A
PRODUCT
Iron Ore
i
ANNUAL
PRODUCTION
(metric tons'
of ore mined)
1,083,000
7,400,000
2,300,000
10,800,000
1,400,000
2,200,000
1,700,000
65,000
2,160,000
1,800,000
3,500,000
1,300,000
0
1,300,000
1,100,000
273.000
450,000
8,830,000
TYPE OF
MINE
OP
UG
OP
1
UG
OP
OP
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling
!
unk
Settling
unk
na
unk
unk
Neutralization;
Settling
unk
Settling
DISCHARGE
METHOD
To surface
In pit settling and
pump to surface
Same as above
To surface
unk
unk
None
None
To surface
unk
Surface
unk
na
unk
unk
To surface
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
22.5 x 103
67.4 x 103
11.3 x 103
11.3 x103
na
na
0
0
na
na
1.7x 103
na
1 '
456 x103
0
0
j
"To convert to annual short tons, multiply values shown by 1.10231
t To convert to daily gallons, multiply values shown by 264.173
Status Code: A active; I inactive; S seasonal; U.D. under development; EXP exploration underway
-------
TABLE 111-1. PROFILE OF IRON MINES (Continued)
MINE
1139
1140
1141
1142
1143
1144
1145
1146
1147
LOCATION
(state)
GA
nflni
MM
iwini
UIU
rnn
Ml
Ml
NV
MN
MN
YEAR
OPENED
(original
facility)
unk
unk
unk
1943
1957
1943
1960
unk
unk
STATUS
OF
OPERATION
unk
A
A
A
A
A
A
PRODUCT
Iron ore
i
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
na
i
1
2,400,000
1,700,000
104,000
600,000
375,000
TYPE OF
MINE
unk
\
I
UG
OP
OP
OP
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
unk
i
r
Settling
i
t
DISCHARGE
METHOD
unk
!
t
To surface
<
t
DAILY
DISCHARGE
VOLUME
Im3t)
na
1
1
i
0.5 x 103
12.3 x103
2.4 x 103
37.9 x 103
8.3 x 103
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Code: A active; I - inactive; S - seasonal; U.D. - under development; EXP - exploration underway
-------
TABLE 1112. PROFILE OF IRON MILLS
MILL
1101
1102
1103
1104
1105
1106
1107
1108
1109
1110
1113
1112
1111
1114
1115
LOCATION
(state)
MN
MN
MN
MN
MN
MN
Ml
Ml
Ml
PA
MN
MN
MN
MO
MO
YEAR
OPENED
(original
facility)
1957
1965
1948
1953
1967
1967
1959
1956
1964
1958
1976
1967
1955
1961
1968
STATUS
. OF
OPERATION
A
A
A
A
A
A
A
A
A
A
A
A
A
A
I
PRODUCT
Iron ore pellets
Iron ore pellets
Iron ore
I ron ore
Iron ore pellets
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
10,600,000
2,840,000
2,200,000
434,000
2,600,000
12,250,000
2,045,000
3,600,000
5,500,000
1.160,000
5,400,000
2,600,000
10,500,000
1,400,000
950,000
CONCENTRATION
PROCESS
USED
Magnetic sep.
Magnetic sep.
Jig; wash; heavy media
Jig; wash; heavy media
Magnetic sep.
Magnetic sep.
Magnetic sep.;
flotation
Flotation
Magnetic sep.;
flotation
Magnetic Sep.;
flotation
Magnetic sep.
Magnetic sep.
Magnetic sep.
Magnetic sep.;
flotation
Magnetic sep.;
flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling
DISCHARGE
METHOD
None
I
To surface
None
None
To si
i
irface
None
None
None
To surface
I
j
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
0
22.5 x 103
0
0
10.2 x 103
32.7 x 103
23 x 103
6.54 x 103
0
0
0
6.5 x 103
16.1 x 103
en
CO
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status coda: A active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE iil-2. PROFILE OF IRON MILLS (Continued)
MILL
1116
1118
1119
1120
1117
1121
1122
1123
1124
1125
1126
1127
1128
1129
1130
LOCATION
(ttate)
Wl
CA
WY
Ml
UT
MM
nfim
MN
MN
MN
MN
MN
UT
NM
TX
NY
YEAR
OPENED
(original
facility)
1967
1948
1962
1974
1946
1968
1973
1917
1933
1965
1965
1953
1938
1947
1942
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
A
A
A
A
A
A
PRODUCT
Iron ore pellets
Iron ore
Iron ore pellets
Iron ore pellets
Iron ore pellets
Iron or
e
Sinter-Iron ore
Iron ore
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
685,000
3.050,000
1.630,000
4,200,000
1,200,000
480,000
700,000
602,000
1,080,000
150,000
330,000
950.000
40,000
725,000
261,000
CONCENTRATION
PROCESS
USED
Magnetic sep.
Wash; jig; heavy media;
magnetic sep.
Magnetic sep.
Selective floc-
culation
Heavy media;
magnetic sep.
Wash
Wash; heavy media
Wash; heavy media; jig
Wash; heavy media
Wash
Wash; jig; heavy media
Wash
None
Wash
Magnetic sep.
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling
None
Settling
unk
DISCHARGE
METHOD
None
None
To surface
To surface
To surface
To surface
None
None
None
To surface
+
None
None
None
unk
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
Minimal
unk
na
5.7 x 103
0
0
0
na
*
0
0
0
na
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-2. PROFILE OF IRON MILLS (Continued)
MILL
1131
1132
1133
1134
1135
1138
1137
1139
1140
1141
1142
1143
1146
1146
1147
1148
1149
LOCATION
(state)
NY
WY
MN
MN
MN
MN
CA
GA
MN
MN
MN
MN
NV
MN
MN
MN
MN
YEAR
OPENED
(original <
facility)
1944
1900
1974
1974
1960
1976
1971
unk
unk
unk
1943
1957
1960
unk
i
»
STATUS
OF
OPERATION
A
A
1
A
A
A
A
unk
unk
unk
A
1
A
A
A
A
1
PRODUCT
Sinter -Iron ore
Iron ore
Iron ore
Iron ore
Iron ore
Iron ore pellets
1
r
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
943.000
500,000
0
100.000
195,000
2,640.000
na
i
t
673,000
0
104,000
337,000
182,000
1,430,000
0
CONCENTRATION
PROCESS
USED
Magnetic tep.
Jig; heavy media
-
Wash
Heavy media
Magnetic sep.
ur
!
ik
Gravity
-.
None
Screen; gravity
Gravity
Gravity
Gravity
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Settling and
Filtration
unk
na
unk
unk
Settling
unk
Settling
unk
None
(
Settling
DISCHARGE
METHOD
Filtrate to
surface
unk
na
unk
unk
None
None
unk
t
None
To surface
unk
None
t
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
61.8x10
na
closed
na
na
0
0
na
0
3.5 x 103
unk
0
0
" 0
na
CTi
O
"To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Statui code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-3. PROFILE OF COPPER MINES
MINE
2101
2102
2103
2104
2107
2108
2109
2110
2111
LOCATION
(ttate)
NV
AZ
NM
NM
AZ
AZ
AZ
AZ
AZ
YEAR
OPENED
(original
facility)
1910
1961
1973
1969
1913
1885
1951
1972
1968
1962
. STATUS
OF
OPERATION
A
A
A
A
I (mines)
A (leach)
A
A (OP)
UD (UG!
MOP)
A (ieach)
I
PRODUCT
Cu ore
Cement Cu
Cu ore
Cement Cu
Cu ore
Cement Cu
Cu ore
Cement Cu
Cement Cu
Cu ore
Cement Cu
Cu ore
Cu ore
Cement Cu
Cu ore
ANNUAL ,
PRODUCTION '
(metric tons*
of ore mined)
7,196,000
(1973)
5.457,000
8,200
13,974,000
(1973)
7,349,000
(1973)
3,994,000
(1972)
2.782,000
5,000
3.383,000
3,711,000
(1973)
1,480,000
(1973)
TYPE OF
MINE
OP;
Dump leaching
OP;
Vat leaching
OP;
Dump leaching
OP;
Dump leaching
OP;
UG;
Dump leaching
OP;
Dump leaching
OP;
UG
OP;
Dump leaching
OP
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Mine water to
leach circuit
Recycle to mill
None
Settling
unk
Total recycle
Recycle to mill
Total recycle
Mine water normally
used in mill circuit,
discharged at present
due to inactive status
of mine
DISCHARGE
METHOD
None
None
None
To surface
(intermittent)
None
None
None
None
To surface
DAILY
DISCHARGE
VOLUME
(m3 t,
0
0
0
680
(average)
0
0
0
0
NA
CT)
*To convert to annual short tons, multiply values shown by 1.10231
TTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
1 Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-3. PROFILE OF COPPER MINES (Continued)
MINE
2112
2113
2115
2116
2117
2118
2119
2120
2121
LOCATION
(state)
AZ
AZ
AZ
AZ
TN
AZ
AZ
MT
Ml
YEAR
OPENED
(original
facility)
1974
1917
1910
1955
1899
1942
1956
1955
1953
' STATUS
OF
OPERATION
UD
A
A
A
A(UG)
UD (OP)
A
A
A (OP)
(leach)
I (UG)
A
PRODUCT
Cu ore
Cu ore
Cu ore
Cu ore
Cement Cu
Cu ore
Cu ore
Cement Cu
Cu ore
Cu ore
Cement Cu
Cu ore
ANNUAL ,
PRODUCTION 1
(metric tons*
of ore mined)
635,000 (projected)
9,381,000
(1973)
1,411,000
8,894,000
31,000
1,836,000
16,653.000 (1973)
na
13,620,000
15.419.000
16,300
(1973)
3,281,000
TYPE OF
MINE
UG
OP
UG
OP;
Dump, vat
leaching
UG;
OP
OP;
Dump leaching
UG
OP;
UG;
Dump leaching
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
To be used in mill
Used in mill
Used in mill
Total recycle of
leach circuit. Mine
water used as
potable water.
Lime pptn;
aeration; settle.
Water from inactive
mine to tailing pond.
Mine water to
leach circuit
Mine water to
mill circuit
Lime pptn; settle;
pH adjustment;
partial recycle to
mill
Settle; alkaline
sedimentation;
secondary settling
DISCHARGE
METHOD
None
None
None
None
To surface
None
None
To surface
To surface
i (seasonal)
DAILY
DISCHARGE
VOLUME
(m3 t)
0
0
0
0
unk
0
0
35,960
(Combined d/c)
190
(sep mine water)
121,120
(Combined d/c)
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-3. PROFILE OF COPPER MINES (Continued)
O-i
U!
MINE
2122
2123
2124
2125
2126
2130
2131
2132
2133
LOCATION
(state)
UT
AZ
AZ
AZ
NV
NM
NV
NV
NV
YEAR
OPENED
(original
facility)
1906
1940
1915
1971
1953
1967
1969
1967
1974
STATUS
OF
' OPERATION
A (OP)
(leach)
UD(UG)
A
A
1
A
(due to close)
(in 1978)
A
UD
1
1
PRODUCT
Cu ore
Cement Cu
Cu ore
Cathode Cu
Cu ore
Cathode Cu
Cement Cu
Cement Cu
Cu ore
Cement Cu
Cu ore
Cu ore
Cement Cu
Cu ore
Cement Cu
Cu ore
ANNUAL ,
PRODUCTION
(metric tons*
of ore mined)
32,208.000
36,300
(1973)
1,854,000
6.620
6,086,000
(1973)
12.600 (1972)
7.400 (1973)
no production
in 1976
7,256,000
na
Confidential
na
na
0
TYPE OF
MINE
OP;
Dump leaching
UG
OP,
Dump leach
OP;
Dump; heap; vat
and in-situ leach-
ing
In-situ leaching
OP;
Vat and dump
leaching
OP;
UG
OP;
Heap leaching
OP;
Dump leaching
UG;
OP
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Mine water to leach
circuit with total
recycle
Mine water to leach
circuit - with total
recycle
unk
Total recycle
Used in mill
Used in mill circuit
unk
unk
To tailing pond
DISCHARGE
METHOD
None
None
None
None
None
To surface
None
None
None
DAILY
DISCHARGE
VOLUME
(m3 t)
0
0
0
0
0
combined d/c
0
0
0
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-3. PROFILE OF COPPER MINES (Continued)
MINE
2134
2135
2136
2137
2138
2139
2140
2141
2142
2143
LOCATION
(state)
ID
AZ
AZ
AZ
AZ
AZ
AZ
AZ
AZ
AZ
YEAR
OPENED
(original
facility)
1972
1964
unk
1974
1976
1971
1964
1959
1975
1977
STATUS
OF
OPERATION
A
A (OP; heap
iaach)
UD (vat leach)
I
I
A
A
A
I
A
unk
PRODUCT
Cu ore
Cathode Cu
Cement Cu
Cu ore
Cement Cu
Cu ore
Cement Cu
Cathode Cu
Cu ore
Cu ore
Cement Cu
Cu ore
Cement Cu
Cu ore
Cathode Cu
unk
ANNUAL ,
PRODUCTION
(metric tons*
of ore mined)
na
7.500
na
na
4,535.000
(design)
29.478,000
5,261,000
2,900
4,807,000
na
1,652,000
4,500
(capacity)
unk
TYPE OF
MINE
OP
OP;
Heap leaching;
Vat leaching
OP;
Dump leaching
OP;
Tailings leach
UG;
Vat leaching
OP
OP;
Dump leaching
OP;
Dump leaching
Leaching
unk
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
No minewater
Total recycle
unk
No minewater
Minewater to mill;
leach water evap.
unk
Used in mill
unk
Total recycle
unk
DISCHARGE
METHOD
None
None
None
None
None
None
None
None
None
unk
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-3. PROFILE OF COPPER MINES (Continued)
CTl
in
MINE
2144
2145
2146
2147
2148
2149
2150
2151
2154
LOCATION
(state)
AZ
AZ
AZ
AZ
AZ
AZ
UT
Ml
AZ
YEAR
OPENED
(original
facility)
1970
1965
1974
1957
1954
unk
1979
unk
unk
STATUS
OF
OPERATION
UD
A
A
I
A (leach)
MOP)
I
UD
EXP
EXP
PRODUCT
Cu ore
Cement Cu
Cu ore
Cathode Cu
Cu ore
Cu ore
Cu ore
Cement Cu
Cu ore
Cement Cu
Cu ore
Cu ore
Cu ore
ANNUAL ,
PRODUCTION
(metric tons*
of ore mined)
na
na
16,926.000 (1972)
na
na
17,777,000
na
1,500
na
na
unk
na
na
TYPE OF
MINE
UG;
In-iitu leaching
OP;
Vat leaching
OP
OP
OP;
Dump leaching
OP;
Vat leaching
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
unk
Used in mill
Used in mill
Used in mill
Leach circuit is
totaly recycled
unk
unk
unk
unk
DISCHARGE
METHOD
unk
None
None
None
unk
unk
unk
unk
unk
DAILY
DISCHARGE
VOLUME
(m3 t)
0
0
0
0
0
0
unk
na
na
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
1 Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-3. PROFILE OF COPPER MINES (Continued)
MINE
2155
2156
LOCATION
(state)
OR
ib
YEAR
OPENED
(original
facility)
1892
unk
STATUS
OF
OPERATION
A
1
PRODUCT
Cu, Au, Ag
ores
Cu ores
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
36,300
NA
TYPE OF
MINE
UG
NA
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
None
NA
DISCHARGE
METHOD
None
NA
DAILY
DISCHARGE
VOLUME
(m3*)
0
unk
O1
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE HI-4. PROFILE OF COPPER MILLS
MILL
2101
2102
2103
2104
2107
2108
2109
2111
2112
2113
2115
2116
LOCATION
(state)
NV
AZ
NM
NM
AZ
AZ
AZ
AZ
AZ
AZ
AZ
AZ
YEAR
OPENED
(original
facility)
1910
1961
1969
1913
1885
1951
1974
1962
1978
1924
1913
1910
STATUS
OF
OPERATION
A
A
A
A
I
A
A
I
UD
A
A
A
PRODUCT
Cu, Mo Cone.
Cu Cone.
Mo Cone.
Cu Cone.
Cu Cone.
Mo Cone.
unk
Cu Cone.
Mo Cone.
Cu Cone.
Cu Cone.
Cu Cone.
Cu Cone.
Cu Cone.
Cu Cone.
Mo Cone.
ANNUAL
PRODUCTION1
(metric tons*
of concentrate)
235,000 (1973)
92,000
250
421,000(1973)
na
na
na
59,000
na
62,000
na
41,000
(projected)
164,000(1973)
1,411,000
219,000
682
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settle in tailings
pond. Decant
partially recycled
to mill; remainder
settled in second
pond
Tailings pond w/
recycle of pond
decant to mill
unk
Total recycle
unk
Total recycle
Total recycle
Total recycle
Total recycle
Total recycle
Total recycle
Total recycle
DISCHARGE
METHOD
Second pond
overflow used
for agricultural
irrigation
None
None
To surface
'intermittent)
None
None
None
unk
None
None
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
0
680
0
0
0
na
0
0
0
0
cr>
i
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-4. PROFILE OF COPPER MILLS (Continued)
MILL
2117
2118
2119
2120
2121
2122
2123
2124
2126
2130
LOCATION
(state)
TN
AZ
AZ
MT
Ml
UT
AZ
AZ
NV
NM
__
YEAR
OPENED
(original
facility)
1899
1942
1956
1955
1954
1917
1940
unk
1953
(due to close
in 1978)
1967
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
A
PRODUCT
Cu Cone.
Fe Cone.
Zn Cone.
Cu Cone.
Cu Cone.
Mo Cone.
Cu Cone.
Cu Cone.
Ag Cone.
Cu Cone.
Mo Cone.
Cu Cone.
Mo Cone.
Ag Cone.
Cu Cone.
Mo Cone.
Cu Cone.
Cu Cone.
ANNUAL .
PRODUCTION '
(metric tons*
of concentrate)
73,000
771,000
9,000
440,000 (1973)
381,000
2.100
327,000
125,000
185
742,000(1973)
10,700 (1973)
9,500
180
1.9
69,000 (1973)
na
ns
Confidential
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Lime pptn;
aeration;
settling
Total recycle
Total recycle
See Mine Code
2120
Lime;pptrt; settle;
secondary settling;
poly electrolyte
addition
FeC)3 addition;
oxidation;
lime pptn; settling
Total recycle
unk
Total recycle
Partial recycle
DISCHARGE
METHOD
To surface
None
None
To surface
To surface
To surface
None
None
None
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
32,700
(combined
d/c)
0
0
35,960
(combined
d/c)
121,120
{combined
d/c!
32,200
(combined
d/c)
0
0
0
minimal
00
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status coda: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
1 Unless otherwise indicated, production data represent 1976 information.
-------
TABLE m-4. PROFILE OF COPPER MILLS (Continued)
MILL
2132
2133
2134
2137
2138
2139
2140
2141
2145
2146
2147
LOCATION
(state)
NV
NV
ID
AZ
AZ
AZ
AZ
AZ
AZ
AZ
AZ
YEAR
OPENED
(original
facility)
1967
1975
1973
1974
1976
1971
1964
1959
1969
1974
1957
STATUS
OF
OPERATiON
I
I
A
I
A
A
A
I
A
A
I
PRODUCT
Cu Cone.
Au Cone.
Ag Cone.
Cu Cow.
Cu Cone.
Ag Cone.
Cu Cone.
Cu Cone.
Cu.Mo, Ag
Cone.
Cu Cone.
Mo Cone.
Cu Cent.
Mo Cone.
Cu Cone,
Mo Cone,
Cu Cone.
Mo Cone.
Cu, Mo,
Ag Cone.
ANNUAL
PRODUCTION '
(metric tons*
of concentrate)
na
na
na
0
fit
na
na
131,000
(design!
369,000
87.000
2700
50,800
na
na
na
na
na
na
na
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Flotation
Flotation
,
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
unk
Total evaporation
Rscycle
Total recycle
Total recycle
Total recycle
Tots! recycle
TotaS recycle
Impoundment;
recycle planned
Total recycle
Total recycle
DISCHARGE
METHOD
unk
None
unk
None
None
None
None
None
None
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
na
0
0
0
0
Q
0
0
0
10
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-4. PROFILE OF COPPER MILLS (Continued)
MILL
2148
2150
2151
LOCATION
(state)
AZ
UT
Ml
YEAR
OPENED
(original
facility)
1954
To open in
1979
unk
STATUS
OF
OPERATION
1
UD
1
PRODUCT
Cu Cone.
Cu Cone.
Cu Cone.
ANNUAL
PRODUCTION 1
(metric tons*
of concentrate)
na
na
na
CONCENTRATION
PROCESS
USED
Flotation
unk
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
unk
unk
unk
DISCHARGE
METHOD
unk
unk
unk
DAILY
DISCHARGE
VOLUME
(m3t)
na
na
na
1
o
"To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S - seasonal; UD under development; EXP - exploration underway
Unless otherwise indicated, production data represent 1976 information.
-------
TABLE 111-5. PROFILE OF LEAD/ZINC MINES
MINES
3101
3102
3103
3104
3105
3106
3107
LOCATION
(state)
ME
MO
MO
NY
MO
PA
ID
YEAR
OPENED
(original
facility)
1972
1969
1969
1931
1973
1955
1887
STATUS
OF
OPERATION
1
A
A
A
A
A
A
PRODUCT
Zn/Cu ore
Pb/Zn ore
Pb/Zn/Cu ore
Zn/Pb ore
Pb/Zn/Cu ore
Zn ore
Zn/Pb ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
190,000
1,482,000
972,300
1,009.100
1,032,000
347,700
709,000
TYPE OF
MINE
UG
UG
UG
UG
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Minewater used as
mill feed
Minewater (67%)
used as mill feed;
remainder to mill
wastewater treatment
system
Minewater (62%)
used as mill feed.
remainder to mill
wastewater
treatment system
Minewater used as
millfeed
Minewater (25%)
used as mill feed;
remainder to
settling
Minewater (5%)
used as mill feed.
Minewater (95%)
combined w/ tailings
pond decant to
secondary settling
Minewater combined
w/ mill tailings.
smelter and refinery
wastewater for
treatment. Backfill
mines stopes w/
sand tails.
DISCHARGE
METHOD
None
To surface
To surface
None
To surface
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
0
7J570
3,115
0
8J300
107,900
22,500
(combined
flow)
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-5. PROFILE OF LEAD/ZINC MINES (Continued)
MINES
3108
3109
3110
3111
3112
3113
3114
3115
LOCATION
(state)
TN
MO
NY
TN
NM
CO
ID
Wl
YEAR
OPENED
(original
facility)
1957
1968
1915
1958
unk
1971
1939
1950
STATUS
OF
OPERATION
A
A
A
1
A
A
1
1
PRODUCT
Zn ore
Pb/Zn ore
Zn ore
Zn ore
Pb/Zn ore
Pb/Zn ore
Pb/Ag ore
Pb/Zn ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
354,500
1,013,000
93,700
90,700
122,400
184,500
TYPE OF
MfNE
UQ
UG
UG
UG
UG
UG
61,600
334,000
UG
UG
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Minewater used as
mill feed
Sanitary minewater
and mill water
pumped to sewage
lagoon. Process mine-
water pumped to mil!
tailings pond w/
partial recycle
Minewater used as
mil! feed
Settle in under-
ground sumps.
None
Minewater (50%)
used as mill feed.
No treatment of
excess. Backfill mine
w/ mill sand tails.
Minewater to mill
treatment system.
Sand tails used as
minefill.
Minewater (62%)
used as mill feed.
Remainder to
settling ponds.
DISCHARGE
METHOD
None
To surface
None
To surface
To surface
To surface
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
0
3,370
0
3,600
2,460
6,400
95
4,160
ro
To convert to annual short tons, multiply values shown by 1.10231
f To convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-5. PROFILE OF LEAD/ZINC MINES (Continued)
MINES
3116
3117
i
3118
3119
3120
3121
3122
LOCATION
(state)
CO
VA
VA
MO
ID
ID
MO
YEAR
OPENED
(original
facility)
1930
1928
1928
1954
1950
1940
1967
STATUS
OF
OPERATION
1
A
A
A
A
A
A
PRODUCT
Zn/Pb/Cu/
Ag ore
Zn ore
Zn ore
Pb/Cu ore
Pb/Zn ore
Pb/Zn/
Ag ore
Pb/Zn/
Cu ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
180.000
541,000
(includes production
from mine 3118)
see above
586,500
157.500
256.500
1.008,000
TYPE OF
MINE
UG
UG
UG
UG
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Lime ppt.
Settle in under-
ground mine sumps
Settle in under-
ground mine sumps
Mill feed (18%);
remainder to
multiple pond system
Minewater to mill
wastewater treatment
system. Backfill mine
stopes w/ sand tails
and cement.
Minewater (23%) to
mill wastewater
treatment systems.
Excess receives no
treatment. Backfill
mine stopes w/ mill
sand tails.
Minewater (9%) used
as milt feed;
remainder to multiple
pond system.
DISCHARGE
METHOD
To surface
To surface
To surface
To surface
To surface
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
3,300
6.280
39.000
4,920
2,200
4,700
26,000
CO
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive, S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE IM-5. PROFILE OF LEAD/ZINC MINES (Continued)
MINES
3123
3124
3125
3126
3127
3128
3129
3130
3131
3132
LOCATION
(state)
MO
NJ
NY
TN
TN
TN
UT
UT
Wl
Wl
YEAR
OPENED
(original
facility)
1960
1850
1940
1955
1965
1960
1890
1975
unk
unk
STATUS
OF
OPERATION
A
A
A
A
A
A
A (OP)
1 (UG)
1
A
A
PRODUCT
Pb/Cu/
Zn ore
Zn ore
Zn ore
Zn ore
Zn ore
Zr ore
Pb/Zn/Cu/
Ag/Au ore
Pb/Zn/Ag
ore
Pb/Zn ore
Pb/Zn ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
1,609,000
186,000
21,400
682,000
654,500
477.000
na
na
na
na
TYPE OF
MINE
UG
UG
UG
UG
UG
UG
OP;
UG
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Part of minewater
settled in multiple
pond systems. Part
used as mill feed
with remainder to
to settling pond.
None
None
Minewater used in
mill circuits as
required
Minewater (60%)
used as mill feed;
remainder receives
no treatment.
Multiple pond
system
Open pit minewater
to adjacent mine
(#2122). Under-
ground mine drainage
sold for irrigation.
Lime pptn;floc
addition, multiple
pond. Mill sand
tails for backfill.
None
None
DISCHARGE
METHOD
To surface
To surface
To surface
To surface
To surface
To surface
None
To surface
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
36,700
950
1,360
0-3,340
5,500
5,450
0
32,700
7,600
4,400
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-5. PROFILE OF LEAD/ZINC MINES (Continued)
MINES
3133
3134
3135
3136
3137
3138
3142
4404
3143
LOCATION
(state)
Wl
WA
WA
NV
AZ
CO
UT
CO
CO
YEAR
OPENED.
(original
facility)
unk
unk
unk
1977
1968
1880
1966
1921
1966
STATUS
OF
OPERATION
I
I
I
UD
I
I
A
A
A
PRODUCT
Pb/Zn ore
Pb/Zn ore
Pb/Zn ore
Pb/Zn ore
Zn/Cu ore
Pb/Zn/Cu/
Ag ore
Pb/Zn/Ag/
Cd/Au ore
Pb/Zn/Cu/
Au/Ag ore
Pb/Ag ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
0
273,000
(includes #31 35
prod)
(see above)
- 114,000
(pilot scale)
84,500
(1973)
89,100
196,000
369,100
- 54,500
TYPE OF
MINE
UG
UG
UG
UG
UG
UG
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
None
Settle in under-
ground sumps
No mine drainage
No mine drainage
Minewater used in
mill circuit.
Portion of minewater
to mill. Excess to
mill tailings pond w/
evap. and seepage.
Portion of minewater
to mill. Excess
combined w/ tailings
pond decant for
impoundment w/
evap. and seepage.
Portion minewater to
mill; excess to
impoundment. Other
mine portals receive
no treatment.
None
DISCHARGE
METHOD
To surface
To surface
None
None
None
To surface
(twice/yr)
None
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
na
2,725
0
0
0
na
0
3,800
na
I
01
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE lli-6. PROFILE OF LEAD/ZINC MILLS
MILL
3101
3102
3103
3104
3105
3106
3107
3108
3109
3110
LOCATION
(state)
ME
MO
MO
NY
MO
PA
ID
TN
MO
NY
YEAR
OPENED
(original
facility)
1972
1969
1969
1972
1973
1955
1946
1957
1968
1932
STATUS
OF
OPERATION
I
A
A
A
A
A
A
A
A
A
PRODUCT
Zn, Cu Cone.
Pb'Conc.
Zn Cone.
Pb Cone.
Zn Cone.
Cu Cone.
Pb, Zn
Cone.
Pb Cone.
Zn Cone.
Cu Cone.
Zn Cone.
Pb Cone.
Zn Cone.
Ag Cone.
Zn Cone.
Pb Cone.
Zn Cone.
Zn Cone.
ANNUAL
PRODUCTION
(metric tons*
of concentrate)
25,600 (1973)
228.600
41,600
92,400
16,000
9,800
113,100
60,800
12,200
45
54,500
24,000
41,500
345
16,100
75,000
11,100
11,800
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Lime pptn; multiple
pond, partial recycle
Alk. sed.; multiple
pond, biological
meander
Alk. sed.; multiple
pond; partial recycle
Alk. sed.
Alk. sed.; total
recycle
Alk. sed.; multiple
pond
Sed.; aeration;
flocculation, lime
pptn.; clarification;
high-density sludge
process
Alk. sed.
Alk. sed.; partial
recycle
Alk. sed.; multiple
pond
DISCHARGE
METHOD
To surface
To surface
To surface
To surface
None
To surface
To surface
To surface
To surface
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
1,430
15,140 (mill)
22,700 (total)
9,460 (mill-
water to pond)
9,840 (total)
6,800
0
5.680 (mill)
107,900 (total)
4,353 (mill)
23.650 (total)
216
3.760 (mill)
28,400 (total)
990 (mil!)
2,650 (total)
01
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE HS-6, PROFILE OF LEAD/ZINC MSLLS (Continued)
MILL
3113
3114
3115
3116
3118
3119
3120
3121
3122
3123
3126
3127
3130
3133
LOCATION
(state)
CO
ID
Wi
CO
VA
MO
ID
ID
MO
MO
TN
TM
UT
WI
YEAR
OPENED
(original
facility)
1971
1939
1950
1930
1928
1954
1950
1940
1967
1960
1975
1965
1975
1956
STATUS
OF
OPERATION
A
1
1
__
1
A
A
A
A
A
A
A
^
A
!
PRODUCT
Zn Cone.
Pb Cone.
Pb/Ag Cone.
Pb, Zn Cone
Zn Cone.
Pb Cone
Zn Cone.
Pb Cone.
Pb Cone.
Cu Conr.
Pb/Ag Cone
Zn Cone.
Pb/Ag Cone.
Zn Cone.
Pb Cone.
Zn Cone.
Pb Cone.
Zn Cone.
Cu Cone.
Zn Cone.
Zn Cone.
Pb, Zn,
Ag Cone.
Pb, Zn Cone.
ANNUAL
PRODUCTION
(metric tons*
of concentrate!
28,200
9,100
10,550
16,250
na
27,700
3,300
25.700
2,400
24,000
940
16,700
25,4bO
104,000
6,800
91,450
8,000
8,580
45,800
27,300
Confidential
na
^
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Gravity (jigging)
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Flotation
Heavy media;
Flotation
Flotation
Flotation
Flotation
WASTEWATER
TREATMENT
TECHNOLOGY
USED
Alk. sed,; multiple
pond; partial racycse
Aik. sed. w/
flocculant addition
Alk. sed.; multiple
pond
Alk. sed.; solar evao.
Afk. sed.
ASk. sed.
Alk. sed. w/
flocculant aodition
Aik. sed, w/
floeculant addition
Alk. sed. w/
partial recycle
Alk. sed. w/
total recycle
Alk. sed. w/
total recycle
Alk. acd.;
partial recycle
Impoundment
solar evap.
Settling
DISCHARGE
METHOD
To surface
To surface
To surface
To surface
To surface
To surface
To surface
To surface
None
None
None
None
None
Surface
DAILY
DISCHARGE
VOLUME
(m3t)
5,300
1,400
6,800 (mil!)
18,000 (total)
3,480 (total)
2,500 (3 mills-
combined flow)
6,750
2,040 (mill)
4,730 (total)
3,330 (mill)
5,980 (total)
O(mill)
2,600 (total)
0
0
0
0
2,900
.
"To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily galions, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-6. PROFILE OF LEAD/ZINC MILLS (Continued)
MILL
3134
3136
3137
3138
3139
3140
3141
3142
3143
4404
LOCATION
(state)
WA
NV
AZ
CO
IL
NM
TN
UT
CO
CO
YEAR
OPENED
(original
facility)
1950
1977
1968
1965
unk
1951
1913
1970
1966
1921
STATUS
OF
OPERATION
I
A
1
1
1
A
1
A
A
A
PRODUCT
Pb Cone.
Zn/Cd Cone.
Pb/Ag Cone.
Zn Cone.
Zn Cone.
Cu Cone.
Pb/Cu/
Ag Cone.
Zn Cone.
Zn Cone.
Pb Cone.
Pb, Zn Cone.
Zn Cone.
Pb/Ag/
Au Cone.
Zn/Cd Cone.
Pb Cone.
Ag Cone.
Pb/Ag/
Au Cone.
Zn Cone.
Cu Cone.
ANNUAL
PRODUCTION
(metric tons*
of concentrate)
na
~ 118,000
(pilot scale)
na
na
na
29,000
47,700
12,000
24,500
na
9.300
16,400
4,400
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Flotation
Gravity (jig);
Flotation
Flotation
Heavy media;
Flotation
Flotation
Flotation
Flotation
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Impoundment;
solar evap.
Impoundment;
solar evap. and
seepage
Impoundment;
solar evap. and
partial recycle
Impoundment;
solar evap.
unk
Impoundment;
evap. and seepage
Alk. sed.
Impoundment;
solar evap. and
seepage
Impoundment;
solar evap. and
partial recycle;
multiple pond
Impoundment;
solar evap. and
seepage
DISCHARGE
METHOD
None
None
None
Intermittent
to surface
(twice/yr)
To surface
None
To abandoned
mine
None
None
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
0
na
3,300
0
0
0
0
0
oo
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-7. PROFILE OF MISCELLANEOUS LEAD/ZINC MINES
NUMBER
OF
MINES
2
3
1
4
1
LOCATION
(state)
CO
ID
OR
TN
NM
YEAR
OPENED
(original
facility)
u
i
ik
P
STATUS
OF
OPERATION
A
A
A
A
A
PRODUCT
Pb/Zn ore
Pb/Zn ore
Pb/Zn ore
Zn ore
unk
ANNUAL
PRODUCTION
(metric tons*
-------
TABLE IH-8. PROFILE OF MISCELLANEOUS LEAD/ZINC MILLS
00
o
NUMBER
OF
MILLS
1
1
1
2
1
LOCATION
(state)
ID
MT
OR
TN
w;
YEAR
OPENED
(original
facility)
unk
1927
1975
unk
STATUS
OF
OPERATION
A
A
A
A
A
I
PRODUCT
unk
Zn Cone.
Zn Cone.
Pb Cone.
Zn Cone.
ANNUAL
PRODUCTION
(metric tons*
CONCENTRATION
PROCESS
of concentrate) I USED
(Employs 7
total)
na
(Employs 3
total)
na
(Employs 7
total)
28,200
- 614,000
na
unk
l
Flotation
Heavy media;
Flotation
Gravity (jigging);
Flotation
WASTE WAT£?< j
TREATMENT
TECHNOLOGIES
DISCHARGE
METHOD
USED J__
unk
\
p
urik
DAILY
DISCHARGE
VOLUME
!m3t)
na
I I
\
!
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status coda: A active; I inactive; S seasonal; UD under development; EXP - exploration underway
-------
TABLE 111-9. PROFILE OF GOLDMINES
MINE
4101
4102
4104
4105
4115
4123
4124
4125
4126
4127
4130
4154
LOCATION
(state)
:,. <
NV
CO
WA
SO
UT
NV
NM
CA
AK
AK
NV
NM
VEAR
xNED
tjr'.-^al
facil.iv)
1965
unk
1937
1877
1965
1973
unk
unk
1924
unk
1973
unk
STATUS
OF
OPERATION
A
A
A
A
A
A
S
I
S
S
I
UD
PRODUCT
Au ore
Pb, Zn, Au
ore
Au ore
Au ore
Au/Ag ore;
some copper
Au ore
Au c
i
re
i
Au ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
744,300
(1973)
162,000
(1973)
49,610
(1974)
1,416,387
(1973)
131,000
na
na
na
612.000m3**
(1975)
~ Same as above
~ 1530m3/day**
(seasonal)
680,000
TYPE OF
MINE
OP
UG
UG
UG
UG
OP
Placer
UG and
Placer
Placer; dred-
ging; Au re-
covered by
gravity +
amalgamation
Same as above
Placer;
mechanical
excavation
OP
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
None
None
mpoundment
None
None
None
Impoundment;
recycle
unk
Settling and
partial recycle
Settling and
partial recycle
Impoundment;
and solar evap.
None
DISCHARGE
METHOD
None
To surface
Indirect
seepaoe and
mil! makeup
Mill makeup
To surface
None
None
unk
Recycle;
To surface
Recycle;
To surface
None
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
A
V
3,788
144
1 1 ,500
na
0
0
na
minimal
minimal
0
unk
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
"Where placer mining is employed, productions are given in cubic meters of ore mined.
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-10. PROFILE OF GOLD MILLS (Continued)
oo
ro
MILL
4119
4120
4121
4122
4128
4129
4131
4154
LOCATION
(state)
NV
NV
NV
NV
NV
CO
NV
NM
YEAR
OPENED
(original
facility)
1975
1977
1969
1973
1975
1964
1976
unk
STATUS
OF
OPERATION
A
A
1
1
A
UD
A
UD
PRODUCT
Au/Ag
Au/Ag
Au
Au
Au/Ag
Au/Ag
Au/Ag
Au
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
na
na
81.200
troy oz
(1974)
23,100
troy oz
(1974)
na
Expect
~ 300,000
Expect
60-80x103
troy oz (Au)
30-40x103
troy oz (Ag)
unk
CONCENTRATION
PROCESS
USED
Cyanidation;
counter current
decantation
Flotation; cyanida-
tion of flotation
concentrate
Cyanidation;
agitation leach; heap
leach and Zn pptn
Cyanidation; heap
leach; carbon ad-
sorption and elec-
trowinning
Cyanidation; heap
leach; carbon ad-
sorption; electro-
winning
Flotation
Cyanidation;
heap leach; carbon
adsorption; elec-
trowinning
Crushing, cyanide
leach, carbon
adsorption, electro-
winning
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Recycle of spent
leach sol'n in mill;
impoundment of
tailings
Impoundment;
solar evap and
recycle
Impoundment and
recycle
Recycle
Impoundment and
recycle
Settling; partial
recycle
Impoundment and
recycle
Recycle
DISCHARGE
METHOD
Small volume
discharged
from tailings
pond to desert
floor
None
None
None
None
To surface
None
None
DAILY
DISCHARGE
VOLUME
-------
TABLE fll-10. PROFILE OF GOLD MILLS (Continued)
MILL
4119
4120
4121
4122
4128
4129
4131
4132
LOCATION
(state!
NV
NV
NV
NV
NV
CO
NV
NM
YEAR
OPENED
(original
facility)
1975
1977
1969
1973
1975
1964
1976
unk
STATUS
OF
OPERATION
A
A
1
1
A
UD
A
UD
PRODUCT
Au/Ag
Au/Ag
Au
Au
Au/Ag
Au/Ag
Au/Ag
Au
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
na
na
81,200
troy oz
(1974)
23,100
troy oz
(1974)
na
Expect
~ 300,000
Expect
60-80x103
troy oz (Au)
30-40x1 03
troy oz (Ag)
unk
CONCENTRATION
PROCESS
USED
Cyanidation;
counter current
decantation
Flotation; cyanida-
tion of flotation
concentrate
Cyanidation;
agitation leach; heap
leach and Zn pptn
Cyanidation; heap
leach; carbon ad-
sorption and elec-
trowinning
Cyanidation; heap
leach; carbon ad-
sorption; electro-
winning
Flotation
Cyanidation;
heap leach; carbon
adsorption; elec-
trowinning
Crushing, cyanide
leach, carbon
adsorption, electro-
winning
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Recycle of spent
leach sol'n in mill;
impoundment of
tailings
Impoundment;
solar evap and
recycle
Impoundment and
recycle
Recycle
Impoundment and
recycle
Settling; partial
recycle
Impoundment and
recycle
Recycle
DISCHARGE
METHOD
Small volume
discharged
from tailings
Bond to desert
floor
None
None
None
None
To surface
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
-10% of
total
0
0
0
0
na
0
0
(planned)
OD
CO
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
1To convert to grams, multiply value shown by 31.1
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-11. PROFILE OF MISCELLANEOUS GOLD AND SILVER MINES
oo
NUMBER
OF
MINES
1
na
39
47
48
41
31
36
!
LOCATION
(state)
AL
AK
AZ
CA
CO
ID
MT
NV
YEAR
OPENED
(original
facility]
unk
i
STATUS
OF
OPERATION
A: 1
A: na
S: na
A: 20
1: 19
A: 19
1: 28
A: 16
1: 32
A: 5
1: 36
A: 8
1: 23
A: 14
I: 22
PRODUCT
Au
Placet; Au
Placer; Ag
Au ore
Au/Ag ore
Ag ore
Au ore
Au/Ag ore
Ag ore
Placer Au
Placer Ag
Au ore
Au/Ag ore
Ag ore
Placer Au
Au ore
Au/Ag ore
Agore
Placer Au
Au ore
Au/Ag ore
Agore
Placer Au
Placer Ag
Au ore
Au/Ag ore
Ag ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
na
91 46 troy oz1
351 troy oz
15,708
0
15.273
1,805
2,221
90
2,809 troy oz
272 troy oz
na
na
110,935
226 troy oz
587
1,480
51,072
24 troy oz
4,020
19,204
26,511
143 troy oz
21 troy oz
1,803.183
11,434
229
TYPE OF
MINE
(If active)
OP
Dredging;
Hydraulic
and/or
mechanical
excavation
OP: 6
UG: 13
OP: 11
UG: 8
OP: 4
UG: 12
OP: 1
UG: 4
OP: 2
UG: 6
OP: 12
UG: 2
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
unk
Settling; recycle
None at most;
settling at some.
ur
i
k
I
Most are dry mines
DISCHARGE
METHOD
unk
Bleed to surface
To surface
ur
!
k
r
None
DAILY
DISCHARGE
VOLUME
(m3t)
na
~ 5-10% of
total volume
19,000-57.000
n
1
i
r
na
'To convert to ufr.uaf short tons, multiply values shown by 1.10231
tTo convert to daily gai'onj, multiply values shown by 264.173
To convert to grams, multiply value shown by 31.1
Status code: A - active; i - inisctiva; S - saMonal; UD - ur.cSar development; EXP - - exploration isr-darwav
-------
TABLE 111-11. PROFILE OF MISCELLANEOUS GOLD AND SILVER MINES (Continued)
NUMBER
OF
MINES
19
>__
I
:>
~
18
2
12
«
16
f
LOCATION
(stats)
NM
NC
OR
SD
UT
VA
WA
YEAR
OPENED
(original
faciHty)
i
ink
1
STATUS
OF
OPERATION
A: 19
A: 2
A: 1
!: 17
A: 1
!: 1
A: 6
i: 6
A: 1
A: 1
i: 15
PRODUCT
Ay ore
Au/Ag ore
Ag ore
Au/Ag ore
Au ore
Au/Ag ore
Agore
Au/Ag ore
Au/Ag ore
Au/Ag ore
Au/Ag ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
1,403,722
0
0
na
2,510
2,97''
47
na
na
na
na
TYPE OF
MINE
(If active)
OP: 8
UG: 11
WASTEWATEH
TREATMENT
TECHNOLOGIES
USED
ur
OP: 0
UG: 2 i
f
OP: 0
UG: 1
OP: 0
UG: 1
OP: 1
UG: 5
OP: 1
OP: 0
UG: 1
L
i
k
_ !
DISCHARGE
METHOD
ur
)
)k
t
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-12. PROFILE OF MISCELLANEOUS GOLD AND SILVER MILLS
NUMBER
OF
MILLS
5
4
8
1
1
8
1
1
LOCATION
(state)
AZ
CA
CO
ID
MT
NV
NM
OR
YEAR
OPENED
(original
facility)
ur
I
k
STATUS
OF
OPERATION
A: 3
I: 2
A: 2
I: 2
A: 5
I: 3
I
1
A: 4
1: 4
A
1
PRODUCT
Au/Ag
Au/Ag
Au/Ag
unk
unk
Au/Ag
Au/Ag
unk
ANNUAL
PRODUCTION
(troy oz.
of .
metal)1
n
i
a
CONCENTRATION
PROCESS
USED
u
i
-ik
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Ul
1
ik
DISCHARGE
METHOD
u
i
nk
DAILY
DISCHARGE
VOLUME
(m3 t)
n
\
a
oo
CTi
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
iTo convert to grams, multiply value shown by 31.1
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-13. PROFILE OF SILVER MINES
MiNE
4401
4402
4403
4406
4407
4408
4409
4410
4411
LOCATION
(state)
ID
CO
ID
ID
ID
MT
ID
ID
TX
YEAR
OPENED
(original
facility)
1947
1967
1921
1975
1976
1974
1920
1952
unk
STATUS
OF
OPERATION
A
A
A
A
A
A
UD
UD
UD
PRODUCT
Tetrahedrite
ore
Ag ore
Tetrahedrite
ore
Tetrahedrite
ore
Ag ore
Ag ore
Ag ore
Tetrahedrite
ore
Ag ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
164,045
(1976)
74,426
(1973)
180.000
(1973)
^ 97,950
na
- 67,500
na
- 76,500
(1972)
unk
TYPE OF
MINE
U
i
G
<
unk
UG
'
1
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Multiple pond
settling
Mechanical
lime ppt.
Settled in mill
tailings pond
None
unk
None
None
Settled in mill 4403
tailings pond
Settling (planned)
DISCHARGE
METHOD
To surface
To surface
To surface
Mill makeup
unk
None
To surface
To surface
To arroyo
tributory of
creek
DAILY
DISCHARGE
VOLUME
(m3t)
800
2,936
3,133
0
na
0
na
2,727
56.8**
co
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Coda: A active; I inactive; S seasonal; UD under development; EXP exploration underway
Allowable discharge (NPDES permit)
-------
TABLE 111-14. PROFILE OF SILVER MILLS
MILL
4403
4401
4402
4406
44Q7
4411
LOCATION
(state)
ID
ID
CO
ID
ID
TX
YEAR
OPENED
(original
facility)
1921
1947
1967
1975
1976
unk
STATUS
OF
OPERATION
A
A
A
A
A
UD
PRODUCT
Tetrahedrite
cone.; Ag + Cu
in pyrite cone.
Cu/Ag cone.
(tetrahedrite)
Pb/Ag cone.
Cu/Ag cone.
Ag
Ag
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
8,200
(1971)
4,522
(1973)
7,587
(1972)
Expect
- 2.700
na
unk
CONCENTRATION
PROCESS
USED
Flotation
Flotation
Flotation
Carbon-in-Pulp
Flotation
Cyanidatton; vat
leach; Zn pptn
Crushing, grinding,
cyanidation
WASTEWATEH
TREATMENT
TECHNOLOGIES
USED
Alk. sed.
Alk. sed. (multiple
pond)
Alk. sad.
Alk. sed. (multiple
pond!
Settling and re-
cycle
Recycle
DISCHARGE
METHOD
Decant to
surface
Decant to
surface; re-
cycle, seepage
Recycle
Decant to sur-
face seepage
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
3,133
na
955
na
na
0
(planned)
00
oo
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE MII-15 PROFILE OF MOLYBDENUM MINES
MINE
5
i ew
6102
U_
6103
61 1C
S_.» j
j...i
6111
6115
6165
t -
LOCATION
(st-itej
MM
CO
CO
,D
/Sf
AK
CO
NV
YEAR
OPENED
(original
facility)
1922
1922
7978
1983
(expected)
unk
unk
1980
STATUS
OF
OPERATION
A
A
A
UD
EXP
I (old
Pb/Zn) EXP
UD
PRODUCT
Mo ore
Mo,W,Sn ore
Mo ore
Mo ore
Mo ore
Mo ore
Mo ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
5,700,000
14,000,000
2,200,000
6.8 x 106
(capacity)
na
UG
OP
TYPE OF
MINE
OP
UG and OP
UG
OP
unk
na
unk
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
None
To mill treatment
system
Settling; flocculants,
aeration
To mill
tailing pond
unk
None
None
DISCHARGE
METHOD
None
To surface
None
unk
To surface
None
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-16. PROFILE OF MOLYBDENUM MILLS
MILL
6101
6102
6103
6110
6165
LOCATION
(state)
NM
CO
CO
ID
NV
YEAR
OPENED
(original
facility)
1922
1922
1976
1983
(expected)
1980
STATUS
OF
OPERATION
A
A
A
UD
UD
PRODUCT
MoS_ cone.
MoS_ cone.
W cofc.
Sn cone.
MoS- cone.
Mo$2 cone.
MoS2, MoO3
cone.
Cu cone.
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
10,831
27,000
na
8170
(capacity)
5447
908
CONCENTRATION
PROCESS
USED
Flotation
Flotation; mag.
and grav.
Flotation
Crushing,
concentration
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Tailings pond,
f locculation; settling
H2°2
Tailings pond;
recycle; 1 X;
chlorin; electro-
coagulation;
flotation
Tailings pond;
recycle
Tailing pond,
recycle
Evaporation,
and recycle
DISCHARGE
METHOD
To surface
To surface
None
None
None
DAILY
DISCHARGE
VOLUME
(m3t)
11,000
11,000
0
0
0
O
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-17. PROFILE OF ALUMINUM ORE MINES
MINE
5101
5102
LOCATION
(state)
AR
AR
YEAR
OPENED
(original
facility)
1942
1899
STATUS
Of
OPERATION
A
A
PRODUCT
Bauxite ore
Bauxite ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
900,000
~ 900,000
TYPE OF
MINE
OP
UG
(inactive)
OP
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Lime neut; settling
Lime neut.; settling
DISCHARGE
METHOD
To surface
To surface
DAILY
DISCHARGE
VOLUME
-------
TABLE ill-18. PROFILE OF TUNGSTEN MINES (PRODUCTION GREATER THAN 5000 MT ORE/YEAR)
I
MILL
61 04
6105
610S
6109
6112
I 6117
LOCATION
(state)
CA
NV
NV
CA
NC
NV
YEAR
OPENED
(original
facility)
1941
1947
1977
unk
unk
unk
STATUS
OF
OPERATION
A
A
A
A
I
A
PRODUCT
W,Mo,Cu ore
Wore
Wore
Wore
Wore
Wore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
6.4 x 106
1x104
na
1.5x 104
3x105
{capacity!
na
TYPE OF
MINE
UG
UG
UG
UG
UG
UG and OP
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Flocculants;
clarification
None
None
None
None
unk
DISCHARGE
METHOD
To mill
process and
to surface
To dry wash
None
On land
To surface
unk
DAILY
DISCHARGE
VOLUME
(m3t)
3.3 x 104
<4
0
<6
na
na
(£»
ro
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Code: A active; I inactive; S seasons!; UO under development; EXP exploration underway
-------
TABLE 111-19. PROFILE OF TUNGSTEN MINES (PRODUCTION LESS THAN 5000 MT ORE/YEAR)
MINE
6119
6120
6121
6122
8123
6126
6127
6128
6129
6130
LOCATION
(state)
CO
UT
NV
CA
to
ID
ID
CA
NV
YEAR
OPENED
(original
facility)
unk
NV j
<
STATUS
OF
OPERATION
EXP
1
UD
A
i
I1
I,EXP
!
A1
I1
PRODUCT
Wore
Wore
Wore
Wore
Wore
Wore
Wore
Wore
Wore
Wore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
(Expect 10-15}
na
na
~ 1000
na
i
P
TYPE OF
MINE
UG
OP and
UG
UG
UG
UG
UG
unk
unk
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Nona
None
None
None
unk
j
I
I
\
DISCHARGE
METHOD
None
None
None
None
unk
t
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-19. PROFILE OF TUNGSTEN MINES {PRODUCTION LESS THAN 5000 MT ORE/YEAR) (Continued)
MINE
6131
6132
6133
6134
6135
6136
6137
6138
6139
6140
6141
6142
6143
6144
6145
6146
6147
6148
6149
6150
6151
6152
LOCATION
(state)
CA
UT
MT
ID
CA
UT
UT
CA
CA
CA
CA
CA
CA
CA
CA
CA
CA
NV
ID
ID
MT
OR
YEAR '
OPENED
(original
facility)
unk
STATUS
OF
OPERATION
1
1
1
A1
1
1
A1
A1
1
unk
1
A1
unk
1
1
1
I1
I
A1
A1
PRODUCT
Wore
i
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
na
!
«noo
-500
<100
na
I
I
f
-200
-1000
TYPE OF
MINE
UG
UG
OP
UG
OP
UG
UG
unk
i
UG
unk
unk
UG
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
unk
T
None
unk
None
unk
None
unk
None
unk
None
None
DISCHARGE
METHOD
unk
I
1
None
unk
i
None
unk
To surface
To surface
None
unk
None
To surface
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-20. PROFILE OF TUNGSTEN MILLS (PRODUCTION LESS THAN 5000 MT/YEAR)
MILL
6119
6130
6131
6132
6134
6135
6145
6146
6147
6148
6149
6151
6152
6153
LOCATION
(state)
CO
NV
CA
UT
10
CA
CA
CA
CA
NV
ID
MT
OR
NV
YEAR
OPENED
(original
facility)
u
i
ik
1
STATUS
OF
OPERATION
UD
I1
I1
|1
I1
A1
A1
1
UD
I1
I1
A1
A1
I1
PRODUCT
W
\
cone
1
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
~40
na
na
na
Very small
na
«2
~6
na
na
na
~2
~ 10
na
CONCENTRATION
PROCESS
USED
Flotation
unk
unk
unk
Gravity
unk
Gravity
unk
'
r
Gravity
Gravity
Gravity
unk
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Recycle planned
unk
i
Recycle
unk
i
*
Settling pond
unk
Evap. and
settling pond
unk
DISCHARGE
METHOD
Backfill old
mine workings
unk
\
t
To surface
To sink hole
None
unk
DAILY
DISCHARGE
VOLUME
(m3 t)
n
i
i
»
0
na
0
na
'
~ 5 (est)
0
na
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
1Be$t available information; not contacted since July 1977.
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE Ifl-20. PROFILE OF TUNGSTEN MILLS (PRODUCTION LESS THAN 5000 MT/YEAR) (Continued)
MILL
81 C9
6124
6159
6163
6154
6155
6156
LOCATION!
(state)
CA
CA
NV
CA
MT
NV
UT
YEAR
OPENED
(original
facility)
unk
1978
(planned]
1974
unk
unk
unk
unk
STATUS
OF
OPERATION
UD
I
(restart
planned
1978)
I
(last report
Pb, Ag, Zn
prod.)
Al
A1
A
PRODUCT
W cone
W cone
W cone
W cone
(from tails!
W cone
W cone
W cone
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
-~500 (est)
500 (est)
"-100 (est)
~5G (est)
118
25
na
CONCENTRATION
PROCESS
USED
Gravity
Gravity
(centrifuge)
Gravity
(tables)
Gravity
(tables)
unk
Gravity
unk
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Tailings Pond;
recycle
Tailings pond;
recycle
Tailings pond;
evap,
Tailings pond:
recycle
unk
Settling pond;
recycle
unk
DISCHARGE
METHOD
None
None
None
None
ursk
None
unk
DAILY
DISCHARGE
VOLUME
*To convert to annual short tons, multiply values shown by 1,10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE 111-21. PROFILE OF TUNGSTEN MILLS (PRODUCTION GREATER THAN 5000 MT/YEAR)
"
MILL
6104
6105
6108
6112
6117
6157
S158
L
LOCATION
{state)
CA
NV
NV
NO
NV
NV
CA
YEAR
OPENED
(original
facility)
1941
1947
1977
unk
i
f
STATUS
OF
OPERATION
A
A
A
1
A
A
|1
PRODUCT
Wconc
MoSj cone.
Cu cone.
W cone.
W cone.
W cone.
W cone.
W cone.
W cone.
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
39.700
118
120
110
1,500 lest)
na
na
~50
450
(capacity)
CONCENTRATION
PROCESS
USED
Flotation
Flotation;
grav.
Flotation
Flotation
Flotation
Flotation
notation
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Lime; tailings
pond
None
Impoundment
Tailings ponds
Tailings pond;
recycle
Tailing: pond;
recycle
Tailings pond;
svap.
DISCHARGE
METHOD
None
To dry wash
None
unk
None
None
None
DAILY
DISCHARGE
VOLUME
(m3 t)
0
~200
0
unk
0
0
0
"to convert to annual short tons, multiply values shown by 1,10231
f To convert to daily gallons, multiply values shown by 264.173
'Bait available information; not contacted since July 1977.
Status code: A active; I inactive; S seasons!; UD under development; EXP exploration underway
-------
TABLE 111-22. PROFILE OF MERCURY MINES
MINE
9201
9202
(1 mine)
(2 mines)
LOCATION
(state)
CA
NV
NV
CA
YEAR
OPENED
(original
facility)
1970
1975
unk
unk
STATUS
OF
OPERATION
1
A
A
S
PRODUCT
Hg ore
Hg ore
Hg ore
Hg ore
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
27,000
222,000
na
na
TYPE OF
MINE
OP
OP
unk
unk
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
None
None
unk
unk
DISCHARGE
METHOD
None
None
unk
unk
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
na
na
co
*To convert to annual short tons, multiply values shown by 1.10231
^To convert to daily gallons, multiply values shown by 264.173
Status Code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
unk = unknown
OP = Open-Pit
-------
TABLE 111-23. PROFILE OF MERCURY MILLS
to
MILL
9201
9202
LOCATION
(state)
CA
NV
YEAR
OPENED
(original
facility)
1970
1974
STATUS
OF
OPERATION
I
A
PRODUCT
Hg
Hg
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
1,049
(1980)
CONCENTRATION
PROCESS
USED
Gravity
Flotation
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Flocculants, set-
tling pond; re-
cycle
Impoundment;
solar evap;
partial recycle
DISCHARGE
METHOD
Recycle when
active; inter-
mittent to sur-
face when in-
active
None
DAILY
DISCHARGE
VOLUME
(m3t)
541
Variable (de-
pends on
precip)
0
To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S - seasonal; UD under development; EXP exploration underway
-------
TABLE IM-24. PROFILE OF URANIUM MINES
o
o
MINE
9401
9402
9404
9404
9411
9412
9419
9408
9409
9410
9437
9445
9447
LOCATION
(state!
NM
NM
NM
NM
WY
WY
TX
CO
WY
WY
NM
NM
UT
YEAR
OPENED
(original
facility)
unk
1970
unk
1976
1963
1957
1973
1957
1972
unk
unk
1976
1972
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
I
A
A
A
PRODUCT
Uran. ore +
leachate
Uran
\
. ore
'
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
680,000
1,590,000
680,000
215,000
324,000
510,000
549,000
45,400
454,000
0
87,100
326,000
238,000
TYPE OF
MINE
UG;
in situ (H2O)
UG
OP
UG
OP
OP
OP
UG
OP
OP
UG
UG
UG
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
UIX; settling;
90% recycle
7 mines: U238 IX;
Impoundment;
solar evap.
2 mines: floccula-
tion; UIX; BaClj
Co-pptn; settling
Impoundment;
solar evap.
Impoundment;
solar evap.
BaCI2 Co-pptn;
settling
None
None
Flocculation; BaCl2
Co-pptn ; settling
None
None
BaCl2 Co-pptn;
settling
Settling; recycle to
mill
Alum, pptn; BaClj
Co-pptn; settling
DISCHARGE
METHOD
To surface
None
To surface
None
None
To surface
None
To surface
To surface
To surface
None
To surface
None
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
3,200
0
11,400
0
0
13,600
0
270
2.1
1,890
8,700
18,900
0
1.5
!
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status Code: A - active; I - inactive; S - seasonal,1 U.D. - under development; EXP - exploration underway
-------
TABLE IM-24. PROFSLE OF URANIUM MINES (Continued)
o
MINE
9448
9405
9450
9452
9413
9455
9460
9451
9402
9433
9438
9439
9440
9441
9443
9444
9446
9449
LOCATION
'natt)
NM
CO
WY
NM
WY
WA
WY
NM
NM
NM
NM
NM
WY
NM
NM
UT
NM
WY
YEAR
OPENED
(original
facility)
1976
urtk
unk
1969
1957
unk
1977
1969
unk
i
i
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
UD
UD
UD
UD
UD
UD
UD
UD
UD
UD
PRODUCT
Ursn. Ore
T
i
1
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
Proprietary
< 295,000
272,000
454,000
540,000
635,000
272,000
Uran. ieachatfe;
i
Urart. Ore 1 na
i .L i .
TYPE OF
MINE
OP and UG
UG
OP
UG
UG
OP
OP
In situ (HjO!
UG
t [ !
JG
i unk
r
450,000
OP
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
None
None
Nona
Flocculation; set-
tling; SaCl2 Co-
pptrs; U IX;
Re228 IX
BaCl2 Co-pptn;
settling
Impoundment;
solar evap.
BpCi2 Co-pptn;
SsttMng
Settling;
recycle
USX;
ursx
_ . . _. .. _
1
1- -- j
L .
1 "
, ,
_.
r
| Floccuiation; UIX;
BaCU Co-pptn;
settling
DISCHARGE
METHOD
f^one
Nona
None
To surface
To surface
None
To surface
None
unk
i
-| i
T
To surface
DAILY
DISCHARGE
VOLUME
(m3t)
0
0
0
6,540
1.36C
0
2,300
0
na
₯
5,500
na
na
37,300
To convert to annual -hort tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 284.173
Status Code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-25. PROFILE OF URANIUM MILLS
MILL
9401
9402
9403
9404
9411
9407
9419
9422
9423
9425
9427
9430
LOCATION
(state)
NM
NM
UT
NM
WY
WY
TX
CO
WA
WY
WY
UT
YEAR
OPENED
(original
facility)
unk
1958
1956
1955
1971
1957
1973
unk
1978
1972
unk
unk
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
A
A
A
UD
PRODUCT
U3°8
Vanad. cone.
U308
Mo cone.
U308
Copper cone.
Vanad. cone.
U30
1
B
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
2,0001
na
10,600
40(1973)
470
159(1973)
na
2,350
844
952
716
857
286 1
1180
543 1
na
CONCENTRATION
PROCESS
USED
Alk. leach
Acid leach
Acid & Alk. leach
Acid leach
Acid leach
Acid leach
Acid leach
Alk. leach
Acid leach
Acid leach
Acid leach
unk
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Impoundment;
solar evap.
Impoundment;
solar evap.
Impoundment;
BaC)2 co-pptn;
solar evap.; recycle
Impoundment;
solar evap; deep
well injection
Impoundment;
solar evap.
Impoundment;
solar evap.
Settling; recycle
Settling; recycle
Impoundment, evap.,
lime ppt, recycle
Impoundment;
solar evap.
Impoundment;
solar evap.
unk
DISCHARGE
METHOD
No
I
ne
1
unk
DAILY
DISCHARGE
VOLUME
(m3 t)
i
9
na
o
ro
*To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
0.18% U-jOg in ore assumed
Status code: A - active; I - inactive; S - seasonal; UD - under development; EXP - exploration underway
-------
TABLE 111-25. PROFILE OF URANIUM MILLS (Continued)
o
MILL
9442
9445
9447
9405
9450
9452
9413
9456
9460
9449
9472
LOCATION
(state)
WY
NM
UT
CO
WY
NM
WY
WA
WY
WY
UT
YEAR
OPENED
(original
facility)
1961
1976
1972
1952
unk
1977
1957
1978
1977
unk
1980
STATUS
OF
OPERATION
I
A
A
A
A
A
A
A
A
UD
A
PRODUCT
U308
U308
U308
U308
Vanad. cone.
U308
U3°8
U3°8
U3°8
U3°8
U3°8
U308
Vanad. cone.
ANNUAL
PRODUCTION
(metric tons*
of
concentrate)
0
8821
454
622
3,630
6861
1.7101
971 1
1.1401
444
410
~ 770
~ 2,700
CONCENTRATION
PROCESS
USED
Acid leach
Acid leach
Alk. leach
Acid leach
Acid leach
Acid leach
Acid leach
Acid; SX
Acid leach
Acid & heap
leach
Acid leach; SX
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Impoundment;
solar evap.
Recycle; impound-
ment; solar evap.
Settling; recycle
Settling; recycle;
solar evap.; floc-
culation; BaCl2
Co-pptn
Impoundment;
solar evap.
Impoundment;
solar evap.
Impoundment;
solar evap.
Lime ppt, BaCU,
recycle
Recycle;
Impoundment;
solar evap.
Impoundment;
recycle; solar
evaporation
Settling;
evaporation
in lined ponds
DISCHARGE
METHOD
None
None
None
To surface
None
None
None
None
None
None
None
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-26. PROFILE OF URANIUM (IIM-SITU LEACH) MIMES
MILL
9417
9424
9458
9459
9461
9462
9463
9464
9465
9466
9467
LOCATION
(state)
TX
TX
TX
TX
TX
TX
TX
TX
TX
TX
TX
YEAR
OPENED
(original
facility)
1975
1976
1976
1977
!
r
unk
tmk
STATUS
OF
OPERATION
A
A
A
A
A
A
A
A
UD
UD
UD
PRODUCT
Uran.
Leachate
»
ANNUAL
PRODUCTION
(metric tons*
of
concentrate) '
113
68
45
227
82
136
ns
i
t
CONCENTRATION
PROCESS
USED
Ammonia in-fitu
leach
Ammonia in-situ
!each; U238 IX
Ammonia in-situ
leach
!
}
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Recyete + deep wal!
injection
(
| ~l
*""!
DISCHARGE
METHOD
None
\
>
DAILY
DISCHARGE
VOLUME
(m3 t)
(
1
{
i
*To convert to annusi ihort torn, multiply saiues shown by '*. 10231
ITo convert to daily gallons, multiply values shown by 264.173
' Production given in terms of yellowcake.
Status code: A active; I inactive; S seasonal; UD under development; EXP exploration underway
-------
TABLE !!!-26. PROFILE OF URANIUM (IN-SITU LEACH) MINES (Continued)
MILL
9468
9468
9470
9471
LOCATION
(state)
TX
TX
TX
TX
YEAR
OPENED
(original
facility)
un
i
k
1
STATUS
OF
OPERATION
U
1
3
i
PRODUCT
Uran.
leachate
i
t
ANNUAL
PRODUCTION
(metric tons*
of
concentrate!
n
a
»
CONCENTRATION
PROCESS
USED
Ammonia
leach
1
in-situ
p
Tailings leach
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Recycle + deep well
injection
i
r
DISCHARGE
METHOD
None
i
r
DAILY
DiSCHARGE
VOLUME
(m3 t)
C
I
o
01
"To convert to annual short tons, multiply values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Production given in terms of yellowcake
Status code: A active; I inactive; S - seasonal; UD under development; EXP exploration underway
-------
TABLE 111-27. PROFILE OF ANTIMONY SUBCATEGORY
FACILITY
Mine 9901
Mill 9901
LOCATION
(state)
MT
MT
YEAR
OPENED
(original
facility)
1969
1970
STATUS
OF
OPERATION
A
A
PRODUCT
Sb ore
NaSbO3
ANNUAL
PRODUCTION
(metric tons*)
18,347
(1979)
272 as
antimony
(1979)
CONCENTRATION
PROCESS
USED
UG
Froth flotation;
caustic leaching
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
None
Impoundment
DISCHARGE
METHOD
None
None
DAILY
DISCHARGE
VOLUME
(n£t)
0
0
o
CTi
*To convert to annual short tons, multiple values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploraton underway
-------
TABLE 111-28. PROFILE OF TITANIUM MINES
MINE
9905
9906
9907
9908
9909
9910
9911
LOCATION
(state)
NY
FL
FL
FL
FL
NJ
NJ
YEAR
OPENED
(original
facility)
1943
1949
1954
1971
1975
1973
1962
STATUS
OF
OPERATION
A
A
A
A
1
A
1
PRODUCT
llmenite ore
llmenite
(containing
beach sands)
ANNUAL
PRODUCTION
(metric tons*
of ore mined)
900,000
7,243,000
7,243,000
na
na
6,585,000
na
TYPE OF
MINE
OP
Beach placer
dredging
i
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling
Q*« Mill QQOfi
Catt Mill QOAR
CM Mill QQ1 1
DISCHARGE
METHOD
To surf ace
DAILY
DISCHARGE
VOLUME
-------
TABLE 111-29. PROFILE OF TITANIUM DREDGE MILLS
MINE
9905
9906
9907
9908
9909
9910
9911
LOCATION
(State)
NY
FL
FL
FL
FL
NJ
NJ
YEAR
OPENED
(original
facility)
1343
1948
1955
1971
1975
1973
1962
STATUS OF
OPERATION
A
A
A
A
1
closed October
1979
A
I
PRODUCTS
llmenite
Magnetite
llmenite
Rutile
Zircon
Staurolite
I
llmenite
Rutile
Zircon
Leucoxine
Monozite
llmenite
Zircon
Monozite
llmenite
llmenite
Sand and
Gravel
ANNUAL _
PRODUCTION
(metric tons
of concentrate)*
~ 200,000
~ 130,000
176.865
126,980
~ 45,000
~ 23,000
~ 23,000
0
165,000
0
CONCEN-
TRATION
PROCESS
USED
Flotation
Dredging;
Wet Gravity
Separation;
Magnetic and
Electrostatic
Separation
WASTEWATER
TREATMENT
TECHNOLOGY
Settling;
partial recycle
Acid addition;
multiple pond
settling; lime
addition and
secondary settling
I
i
Coagulation with
alum or acid;
multiple pond
settling; neutralized
with caustic
solution
Coagulation with
alum; multiple
pond settling-
neutralized with
caustic solution
pH adjustment
with alum or
lime settling
Sent ing ponds
with recycle of
all process water
AVERAGE DAILY
DISCHARGE
VOLUME
(m3/day)
Variable
(depends
on pracip.)
25,927
8.170
13,626
4,883
14,269
0
I'
o
co
+ To convert to annual short tons, multiple values shown by 1.10231
t+To convert to daily gallons, multiple values shown by 264.1 73
Status Code: A - active; I inactive; S seasonal; UD under development; EXP exploration underway
Annual production based on 1978 production
-------
TABLE 111-30. PROFILE OF NICKEL SUBCATEGORY
FACILITY
Mine
6106
Mill
6106
LOCATION
(state)
OR
OR
YEAR
OPENED
(original
facility)
1954
1954
STATUS
OF
OPERATION
A
A
PRODUCT
Ni ore
Fertonickel
ANNUAL
PRODUCTION
(metric tons*)
3.4 x 106
ors
21,050
PROCESS
OP
Scraen, crush
WASTE WATER
TREATMENT
TECHNOLOGIES
USED
Settling ponds
Multiple pond set-
tling; partial recycle
DISCHARGE
METHOD
To mill/smelter
treatment
system
To stream
DAILY
DISCHARGE
VOLUME
(m3t)
3,500
(seasonal)
o
*To convert to annual short tons, multiply values shown by 1.10231
"To convert to daily gallons, multiply values shown by 264.173
Status code: A - active; I - inactive; S - seasonal: UD - under development; EXP - - exploration underway
-------
TABLE 111-31. PROFILE OF VANADIUM SUBCATEGORY
FACILITY
Mine 6107
Mill 6107
LOCATION
(state)
AR
AR
YEAR
OPENED
(original
facility)
1966
1967
STATUS
OF
OPERATION
A
A
PRODUCT
Vanadium
ore
V2°5
ANNUAL
PRODUCTION
(metric tons*)
363,000
4,500
PROCESS
OP
Roasting; leaching;
solvent extraction;
precipitation
WASTEWATER
TREATMENT
TECHNOLOGIES
USED
Lime neut.
and settling
Settling and
pH adjustment
DISCHARGE
METHOD
To surface
To surface;
multiple
discharges
DAILY
DISCHARGE
VOLUME
(mSt)
6,140
4,460
o
*To convert to annual short tons, multiple values shown by 1.10231
tTo convert to daily gallons, multiply values shown by 264.173
Status code: A active; I inactive; S seasonal; UD under development; EXP exploraton underway
-------
SECTION IV
INDUSTRY SUBCATEGORIZATION
During development of effluent limitations and new source
standards of performance for the ore mining and dressing
category, consideration was given to whether uniform and
equitable guidelines could be applied to the industry as a whole,
or whether different effluent limitations ought to be established
for various subparts of the industry. The ore mining and
dressing industry is diverse; it contains nine major SIC codes
and ores of 23 separate metals (counting rare earths as a single
metal). The wastewaters produced vary in quantity and quality,
and treatment technologies affect the economics of each operation
differently.
Because this category is complicated, concise descriptions of
potential subcategories are necessary to avoid confusion.
Therefore, the following definitions are given:
"Mine" is an active mining area, including all land and property
placed on, under or above the surface of such land, used in or
resulting from the work of extracting metal ore from its natural
deposits by any means or method, including secondary recovery of
metal ore from refuse or other storage piles derived from the
mining, cleaning, or concentration of metal ores.
"Mill" is a preparation facility within which the mineral or
metal ore is cleaned, concentrated or otherwise processed prior
to shipping to the consumer, refiner, smelter or manufacturer. A
mill includes all ancillary operations and structures necessary
for the cleaning, concentrating or other processing of the metal
ore such as ore and gangue storage areas, and loading facilities.
"Complex" is a facility where wastewater resulting from mine
drainage and/or mill processes is combined with wastewater from a
smelter and/or refinery operation and treated in a common
wastewater treatment system.
FACTORS INFLUENCING SELECTION OF SUBCATEGORIES
The factors that were examined as a possible basis for sub-
cateooc.i ^at i en are:
1. Designation as a mine or mill
2. General geologic setting
3. Type of mine (e.g., surface or underground)
4. Ore mineralogy
5. Type of mill process (beneficiation, extraction
process)
6. Wastes generated
7. End product
8. Climate, rainfall, and location
111
-------
9. Reagent use
10. Water use or water balance
11. Treatment technologies
12. Topography
13. Facility age
These factors have been examined to determine if the BPT sub-
categorization should be retained or if any modification would be
appropriate.
Designation as a Mine or Mill
It is often desirable to consider mine water and mill process
water separately. Many mining operations do not have an asso-
ciated mill and deliver ore to a mill located some distance away
which other mines also use. In many instances, it is advanta-
geous to separate mine water from mill process wastewater because
of differing water quality, flow rate, or treatability. Levels
of pollutants in mine waters are often lower or less complex than
those in mill process wastewaters. For many mine/mill opera-
tions, it is more economical to treat mine water separately from
mill water, especially if the mine water requires minimum treat-
ability, Mine water contact with finely divided ores (especially
oxidized ores) is minimal and mine water is not exposed to the
process reagents often added in milling. Wastewater volume
reduction from a mine is seldom a viable option whereas the
technology is available to reduce or eliminate discharge from
many milling operations. Therefore, development of treatment
alternatives and guidelines may be difficult for mines and mills.
Because many operations follow this approach, designation as a
mine or mill provides an appropriate basis to classify the
industry within subcategories.
That is not to say that mine water and mill water might not be
advantageously handled together. In some instances, use of the
mine wastewater as mill process water will result in an improved
discharge quality because of interactions of the process chemi-
cals and the mine water pollutants.
General Geologic Setting
The general geologic setting (e.g., shape of deposit, proximity
to surface) determines the type of mine (i.e., underground,
surface or open-pit, placer, etc.). Therefore, geologic setting
is not considered a basis for subcategorization.
Type of_ Mine
The choice of mining method is determined by the general geology;
ore grade, size, configuration, and depth; and associated
overburden of the ore body. Because no significant differences
resulted from application of mine water control and treatment
112
-------
technologies from either surface (open-pit) or underground mines,
mine type was not selected as a suitable basis for general
subcategorization in the industry.
Ore Mineralogy, Type of_ Beneficiation Process and Wastes
Generated
The mineralogy of the ore often determines the beneficiation
process to be used. Both of these factors, in turn, often
determine the characteristics of the waste stream, the treatment
technologies to be employed, and the effectiveness of a
particular treatment method. For these reasons, both ore
mineralogy and type of beneficiation processes are important
factors bearing on subcategorization. For example, pollutants
associated with uranium mining and milling, such as radium 226
and uranium, require treatment technologies not applicable to
lead, zinc, and copper facilities. Chemical reagents used in
froth flotation processes at lead, zinc, copper and other metals
facilities often contain cyanide and other pollutants which are
not used in the uranium mills.
On the other hand, many metals are often found in conjunction
with one another, and are recovered from the same ore body
through similar beneficiation processes. As a consequence, in
these instances wastewater treatment technologies and the effec-
tiveness of particular treatment methods will be similar, and,
therefore, one subcategory is justified. This is the case, for
example, with respect to the copper, lead, zinc, gold, silver,
platinum and molybdenum ores subcategory, where several metals
often occur in conjunction with each other and are recovered by
the froth flotation process. The methods for controlling these
metals (and commonly used reagents such as cyanide) in the
wastewater discharge is similar throughout the subcategory, and
establishing uniform effluent limitations for these facilities is
appropriate. In either case, treatment of total suspended solids
(i.e., settling) is similar.
Processing (or beneficiation) of ores in the ore mining and
dressing industry includes crude hand methods, gravity
separatic", froth flotation using reagents, chemical extraction,
and hyd> '-metallurgy. Physical processing using water, such as
grav^t^ eparation, discharge the suspended solids generated from
wash.ng, dredging, crushing, or grinding. The exposure of finely
*.. >vided ore and gangue to water also leads to solution of some
material. The dissolved and suspended metals content varies with
the ore being processed, but wastewater treatments are similar.
Froth flotation methods affect character of mill effluent in
several ways. Generally, pH is adjusted to increase flotation
efficiency. This and the finer ore grind (generally finer than
for physical processing) may have the secondary effect of sub-
stantially increasing the solubility of ore components. Reagents
used in the flotation processes include major pollutants.
113
-------
Cyanide and phenol compounds, for example, are used in several
flotation processes. Although their usage is usually low, their
presence in effluent streams have potentially harmful effects.
Ore leaching operations differ from physical processing and
flotation plants. The use of large quantities of reagents such
as strong acids and bases and the deliberate solubilization of
ore components (resulting in higher percent soluble metals
content) characterizes these operations. Therefore, the
characteristics of the wastewater quality as well as the
treatment and control technologies employed are different than
for physical processing and flotation wastewater.
The wastes generated as part of mining and beneficiating metal
ores are highly dependent upon mineralogy and processes employed.
This was considered in all subcategories.
End Product
The end products are closely allied to the mineralogy of the ores
exploited; therefore, mineralogy and processing were found to be
more advantageous methods of subcategorization.
Climate, Rainfall, and Location
There is a wide diversity of yearly climatic variations in the
United States. Unlike many other industries, mining and asso-
ciated milling operations cannot choose to locate in areas which
have desirable characteristics. Some mills and mines are located
in arid regions of the country and can use evaporation to reduce
effluent discharge quantity. Other facilities are located in
areas of net positive precipitation and high runoff conditions.
Treatment of large volumes of water by evaporation in many areas
of the U.S. cannot be used where topographic conditions limit
space and provide excess surface drainage water. A climate which
provides icing conditions on ponds will also make control of
excess water more difficult than in a semi-arid area. Climate,
rainfall, and location were, therefore, considered in determining
whether a particular subcategory can achieve zero discharge.
Reagent Use
Reagent use in many segments of the industry (for example, in the
cyanidation process for gold) can potentially affect the quality
of wastewater. However, the types and quantities of reagents are
a function of the mineralogy of the ore and extraction processes
employed. Reagent use, therefore, was included in the
consideration of beneficiation processes (e.g., cyanidation for
gold).
Water Use and/or Water Balance
114
-------
Water use or water balance is highly dependent on the choice of
process employed or process requirements, routing of mine waters
to a mill treatment system or discharge, and potential for use of
water for recycle in a process. Beneficiation processes play a
determining role in mill water balance and, therefore, water use
and/or water balance are considered with beneficiation processes.
Treatment Technologies
Many mining and milling establishments use a single type of
effluent treatment method. Treatment procedures vary within the
industry, but widespread adoption of differing technologies is
not prevalent. Therefore, it was determined that ore mineralogy,
mill process and waste characterization provide a more acceptable
basis for subcategorization than treatment technology.
Topography
Topographical differences between areas are beyond the control of
mine or mill operators, and these place constraints on the
treatment technologies employed. One example is tailing pond
location. Topographical variations can cause serious problems
with respect to rainfall accumulation and runoff from steep
slopes. Topography varie? widely from one area to another and,
therefore, is not a practical basis for subcategorization for
national regulations. However, topography is known to influence
the treatment and control technologies employed and the water
flow within the mine/mill facility. While not used for
subcategorization, topography has been considered in the
determination of effluent limits for each subcategory.
Facility Age
Many mines and mills have operated for the past 100 years. For
mining operations, installation of replacement equipment results
in minimal differences in water quality. Many mill processes for
concentrating ores in the industry have not changed considerably
(e.g., froth flotation, gravity separation, grinding and
crushing), but improvements in reagent use, monitoring and
control have resulted in improved recovery or the extraction of
values from lower grade ores. New and innovative technologies
have resulted in changes in the character of the wastes. This is
not a function of the age of the facilities, but is a function of
extractive metallurgy and process changes. Virtually every
facility continuously updates in-plant processing and flow
schemes, even though basic processing may remain the same. In
addition, most treatment systems employ end-of-pipe techniques
which can be installed in either old or new plants. Therefore,
age of a facility is not a useful factor for subcategorization in
the industry.
SUBCATEGORIZATION
115
-------
After a review of BPT subcategorization, and after a 36-month
data collection effort, it has been determined that the BPT
subcategorization (with minor modifications for BAT) adequately
represents the inherent differences in the industry. The sub--
categories are presented in Table IV-1. The proposed subcate-
gorization is based on:
1. Metal being extracted (referred to as Subcategory)
2. Differences between mine and mill wastewater (referred
to as Subdivision)
3. Type of mill process (referred to as subpart).
The Settlement Agreement approved by the U. S. District Court for
the District of Columbia requires the EPA to establish standards
for toxic pollutants and to review the best available technology
economically achievable (BAT) for existing sources in the ore
mining and dressing industry. The Settlement Agreement, refers
to the ore mining and dressing industry as major group 10, as is
defined in the Standard Industrial Classification Manual (SIC).
Each of the SIC codes in major group 10 were examined in
reviewing potential subcategorization using factors required by
the Act as contained in this section. The prominent factors
identified are: the difference between mine and mills; the ore
type, which is generally related to the SIC code/ the size of
facility (in tonnage); and most importantly, the wastewater
characteristics (pollutants found and the treatment employed for
removal of the pollutants).
All subcategories are subdivided according to mine drainage and
discharge from mills. Subcategories relating to the SIC codes
include: iron ore, aluminum ore, uranium ores, mercury ores,
titanium and antimony ores (the only representative of metal ores
not elsewhere classified SIC 1099). Molybdenum, nickel and
tungsten have been separated from FerroalloysSIC code 1061.
Nickel is a separate subcategory and molybdenum is combined with
several other metals.
Because of the similarity of the wastewater discharge from mills
and mine drainage, a large subcategory is maintained which is
applicable to ores mined or milled for the recovery of copper,
lead, zinc, gold, silver and platinum. Molybdenum was also added
to this group because of similarity in mill processes. The mine
drainage from this subcategory was identified as being of similar
pH with relatively high concentrations of heavy metals regardless
of the ore mined. The most commonly used mill process in the
subcategory is the froth flotation process. In this process,
similar reagents are used, and the wastewater from the mills is
characterized by high levels of total suspended solids and con-
centrations of heavy metals. Cyanide is generally used in the
mill processes. Many mills in this large subcategory produce two
or more metal ore concentrates. Because of the similarity of the
wastewater generated, the same wastewater treatment technologies
are applicable in this subcategory regardless of the type of ore
116
-------
mined and milled. Because of these factors, the subeategory
formerly (in BPT) called Base and Precious Metals is expanded and
renamed the Copper, Lead, Zinc, Gold, Silver, Platinum and
Molybdenum Ores Subcategory.
COMPLEXES
The subcategorization scheme has subdivisions for mines and
mills; complexes are not included. Because of the individuality
of complexes, regulation of them has been delegated to the
Agency's Regional offices and that practice will continue. As
discussed in Section V, Sampling and Analysis Methods, several
complexes have been sampled during BAT guideline development and
a separate guidance document has been prepared to aid the Regions
in preparation of permits commensurate with BAT.
117
-------
TABLE IV-1. PROPOSED SUBCATEGORIZATION FOR BAT - ORE MINING AND
DRESSING
SUBCATEGORY
Iron Ore
Copper, Lead, Zinc, Gold,
Silver, Platinum,
Molybdenum Ores
Aluminum Ore
Tungsten Ore
Nickel Ore
Vanadium Ore*
Mercury Ore
Uranium Ores
Antimony Ores
Titanium Ores
SUBDIVISION
Mine Drainage
Mills
Mine Drainage
Mills or Hydro-
metallurgical
Beneficiation
Mine Drainage
Mine Drainage
Mills
Mine Drainage
Mills
Mine Drainage
Mills
Mine Drainage
Mills
Mine Drainage
Mills, Mines and Mills
or In-Situ Mines
Mine Drainage
Mills
Mine Drainage
Mills
Mills with Dredge
Mining
PROCESS
Physical and/or Chemical Beneficiation
Physical Beneficiation Only (Mesabi Range)
Cyanidation or Amalgamation
Heap, Vat, Dump, In-Situ Leaching (Cu)
Froth Flotation
Gravity Separation Methods (incl. Dredge, Placer,
or other physical separation methods; Mine
Drainage or mines and mills)
(Physical Processes)
Ore Leaching
Gravity Separation, Froth Flotation, Other
Methods
Flotation Process
Vanadium extracted from non-radioactive ores
118
-------
SECTION V
SAMPLING AND ANALYSIS METHODS
The sampling and analysis program discussed in this section was
undertaken primarily to implement the Consent Decree and to
identify pollutants of concern in the industry, with emphasis on
toxic pollutants. A data base has been developed over several
years and consists of nine sampling and analysis programs:
1. Screen Sampling Program
2. Verification Sampling Program
3. Verification Monitoring Program
4. EPA Regional Offices Surveillance and Analysis Program
5. Cost Site Visit Program
6. Uranium Study
7. Gold Placer Mining Study
8. Titanium Sand Dredging Mining and Milling Study
9. Solid Waste Study
This section summarizes the purpose of each of these nine sam-
pling efforts, and identifies the sites sampled and parameters
analyzed. It also presents an overview of sample collection,
preservation, and transportation techniques. Finally, it
describes the pollutant parameters quantified, the methods of
analyses and the laboratories used, the detectable concentration
of each pollutant, and the general approach used to ensure
reliability of the analytical data produced. The raw data
obtained during these programs are included in Supplement A and
are discussed in Section VI, Wastewater Characterization.
SITE SELECTION
The facilities sampled were selected to represent the industry.
Considerations included the number of similar operations to be
represented; how well each facility represented a subcategory,
subdivision, or mill process as indicated by the available data;
problems in meeting BPT guidelines or potential problems in
meeting BAT guidelines; geographic differences; and treatment
processes in use. Successive sampling programs were designed
based on data gathered in previous programs.
Several complexes were sampled even though effluent guidelines
have not been developed for them. The mine and/or mill waters
were sampled separately from the refinery/smelter waters; there-
fore, data applicable to similar mines and/or mills was deve-
loped. Samples taken of refinery and/or smelter water (during
the same sampling program) were used to develop a separate
guidance document for regulation of complexes.
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Screen Sampling Program
Twenty facilities were chosen for initial site visits and sam-
pling to update data previously acquired by EPA and supplied by
industry, and to accomplish screen sampling objectives required
by the Settlement Agreement.
The Water Quality Control Subcommittee of the American Mining
Congress, consisting of representatives from the industry, were
presented with descriptions of candidate sites. The comments of
the subcommittee were considered and a list of sites to be
visited for screen sampling was compiled. At least one facility
in each major BPT subcategory was selected. The sites selected
were typical of the operations and wastewater characteristics
present in particular subcategories.
To determine representative sites, the BPT data base and industry
as a whole was reviewed. Consideration was given to:
1. Those using reagents or reagent constituents on the
toxic pollutants list.
2. Those using effective treatment for BPT control para-
meters .
3. Those for which historical data were available as a
means of verifying results obtained during screening.
4. Those suspected of producing wastewater streams which
contain pollutants not traditionally monitored.
After selection of the facilities to be sampled, a data sheet was
developed and sent to each of the 20 operations, together with a
letter of notification as to when a visit would be expected.
These inquiries led to acquisition of facility information and
establishment of industry liaison, necessary for efficient on-
site sampling. The information that resulted aided in the
selection of the points to be sampled at each site. These
sampling points included, but were not limited to, raw and
treated effluent streams, process water sources, and intermediate
process and/or treatment steps. Copies of the information
submitted by each company as well as the contractor's trip report
for each visit are contained in the supplements to this report.
Sites visited for screen sampling are listed below by subcategory
and facility code:
1. Iron Ore Subcategory - Mine/Mill 1105 and Mine/Mill 1108
2. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Ore Subcategory
- Copper OreMine/Mill 2120, Mine/Mill/Smelter/
Refinery 2122, Mine/Mill/Smelter/Refinery 2121,
and Mine/Mill/Smelter 2117
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- Lead and Zinc OreMine/Mill 3110, Mine/Mill/Smelter/
Refinery 3107, and Mine/Mill 3121
- Gold OreMine/Mill 4105
- Silver OreMine/Mill 4401
- Molybdenum OresMill 6101
3. Aluminum Ore Subcategory - Mine 5102
4. Tungsten Ore Subcategory - Mine/Mill 6104
5. Mercury Ore Subcategory - Mill 9202
6. Uranium Ores Subcategory - Mine 9408, Mine 9411, Mine
9402, and Mill 9405
7. Titanium Ore Subcategory - Mine/Mill 9905
These facilities were visited in the period of. April through
November 1977.
Verification Sampling Program
Sample get: ]_. After review of screen sampling analysis results
(which are summarized in Section VI), 14 sites were selected for
additional sampling visits. Three of these sites were visited to
collect additional analytical data on mine/mill/smelter/refinery
complexes sampled in screen sampling. Three others, including
two not sampled earlier, were visited to collect additional
analytical data on treatment systems which were determined to be
among the more effective facilities studied during the BPT
effort. Because most of the organic toxic pollutants were either
not detected or detected only at low concentrations in the screen
samples, emphasis was placed on "verification" sampling for total
phenol, total cyanide, asbestos (chrysotile), and toxic metals.
The following sites were visited in the period of August 1977
through February 1978:
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Ore Subcategory
- Copper OresMine/Mill/Smelter/Refinery 2122 (two
trips), Mine/Mill Smelter 2121, and Mine/Mill 2120
- Lead and Zinc OresMine/Mill/Smelter/Refinery 3107
(two trips), Mine/Mill 3101 (not screen-sampled),
and Mine/Mill 3103 (not screen-sampled)
Sample Set 2_. Six more facilities, all sampled earlier, were
revisited to collect additional data on concentrations of total
phenol, total cyanide, and/or asbestos (chrysotile), and to
confirm earlier measurements of these parameters. In the period
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of August 1977 through January 1978 the following sites were
visited:
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Ore Subcategory
only)
- Silver OresMine/Mill 4401 (asbestos only)
- Molybdenum OresMill 6101 (asbestos only)
2. Aluminum Ore SubcategoryMine 5102 (total phenol and
asbestos)
3. Uranium Ore SubcategoryMine 9408 (asbestos only) and
Mill 9405 (asbestos only)
Additional Sampling Program
After completion of verification sampling two additional sites
were sampled. At the first, a molybdenum mill operation, a
complete screen sampling effort was performed to determine the
presence of toxic pollutants and to collect data on the perfor-
mance of a newly installed treatment system. The second facil-
ity, a uranium mine/mill, was sampled to collect data on a
facility removing radium 226 by ion exchange. Samples collected
there were not analyzed for organic toxic pollutants. The
following sites were visited in the period of August through
November 1978.
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Ore Subcategory-(Molybdenum) Mine/Mill 6102
2. Uranium Ore Subcategory - Mine/Mill 9452 (not screen-
sampled)
Verification Monitoring Program
Verification monitoring was conducted at three facilities visited
during screen sampling to supplement and expand the data for
these facilities. The programs lasted from 2 to 12 weeks. Two
of the operations were chosen because they had been identified
during the BPT study as two of the more treatment efficient
facilities. Additional data on long term variations in waste
stream characteristics at these sites were needed to supplement
the historical discharge monitoring data, and to reflect any
recent changes or improvements in the treatment technology used.
The third operation was sampled to determine seasonal variability
in the raw and treated waste streams and to supplement existing
NPDES monitoring data.
For these monitoring efforts, contractor sampling of the
facilities was not economical due to the extended time periods
required. Therefore, industry cooperation was solicited, and all
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monitoring program samples were collected by industry personnel
and shipped to the contractor for analysis. Industry personnel
were provided detailed instructions on the propoer methods for
sample collection, preservation, and transportation. The methods
prescribed are the EPA mandated techniques used in the contractor
conducted sampling programs and described in the next subsection.
Facilities monitored during the period of September 1977 through
March 1978 included:
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Ore Subcategory
- Copper OresMine/Mill/Smelter/Refinery 2122 and
Mine/Mill 2120
- Lead and Zinc OresMine/Mill 3103
Additional information on the monitoring efforts conducted at
these facilities is provided in the supplements to this report.
EPA Regional Offices Surveillance and Analysis Program
The data labeled "Surveillance and Analysis" in the supplement to
this report was developed by the Agency's regional Sampling and
Analysis groups. Fifteen facilities were sampled; ten in the
western states of Colorado, Idaho, Wyoming, Montana, and Oregon,
one in Arkansas, and four in Missouri. Facilities visited during
the period of July through September 1977 were:
1. Copper, Lead, Zinc, Gold Silver, Platinum, and
Molybdenum Ore Subcategory
- Copper Mine/Mill 2120
- Lead/Zinc Mine/Mill 3107, Mine/Mill 3102, Mine/Mill
3103, Mine/Mill 3109, and Mine/Mill 3119
- Gold Mill 4102
- Silver Mills 4401 and 4406, Mine 4402 and Mine/Mill
4403
2. Nickel Ore Subcategory - Mine/Mill/Smelter/Refinery 6106
3. Vanadium Ore Subcategory - Mine/Mill 6107
4. Uranium Ore Subcategory - Mine/Mills 9405 and 9411
Cost Site Visit Sampling
The primary reason for these visits was to determine the cost of
implementing particular treatment technologies; therefore, sites
were selected by cost considerations. However, many of these
sites were sampled previously, and the data obtained from the
cost site visits serve to verify the original data.
Facilities visited during the period of September 1979 through
January 1980 were:
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1. Iron Ore Subcategory - Mine/Mill 1132
2. Copper, Lead, Zinc, Gold, Silver, Platinum, Molybdenum
Ore Subcategory
- Copper OreMine 2110 and Mine/Mill 2116
- Zinc OreMine/Mill 3106
- Lead/Zinc OreMine/Mill 3113 and Mine 3117
3. Tungsten OreMine/Mill 6104
Uranium Study
Wastewater sampling was conducted at five uranium mines and at
five uranium mills to expand the data base on current state-of-
the-art treatment technologies. Uranium mine wastewater treat-
ment technologies studied were barium chloride coprecipitation
and ion exchange. Wastewater treatment technologies studied at
the five mill sites included barium chloride coprecipitaiton and
lime precipitation (for metals removal). The following mine and
mill sites were visited during the study:
Mine 9408,
Mill 9414,
Uranium Ore Subcategory - Mine 9401, Mine 9402,
Mine 9411, Mine 9412, Mill 9404, Mill 9405,
Mill 9415, Mill 9416.
Gold Placer Mining Study
A sampling effort was conducted to evaluate treatment technol-
ogies at Alaskan placer mines. BAT regulations for gold placer
mining are reserved for further study. However, several gold
placer mining operations, all located in Alaska, were sampled to
determine performance capabilities of existing settling ponds
used to remove suspended and settleable solids. A summary report
has been issued and the data are included in Reference 1. The
operations visited as part of this effort are:
]. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum Subcategory - Gold Mines 4126, 4127, 4132,
4133, 4134, 4135, 4136, 4137, 4138, 4143, and 4144.
Titanium Sand Dredging Mining and Milling Study
A study at three titanium dredge mining and milling facilities
was conducted to obtain wastewater treatment data on this sub-
category of the ore mining and dressing industry. Facilities
visited during the period from December 1979 to January 1980
were:
1. Titanium Ore Subcategory - Mine/Mill 9906, Mine/Mill
9907 and Mine/Mill 9910.
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The data collected have been summarized in a compehensive report
on titanium sand dredging wastewater treatment practices
(Reference 2).
SoUd Waste Study.
The purposes of the solid waste study were to obtain updated
wastewater and analytical data on six subcategories and to
develop baseline data on the characteristics and amounts of solid
waste generated at the facilities selected. One facility in each
of the following subcategories was selcted:
1, Aluminum Ore (Mine 5101)
2. Tungsten Ore (Mine/Mill 6105)
3. Nickel Ore (Mine/Mill/Smelter/Refinery 6106)
4. Vanadium Ore (Mine/Mill 6107)
5. Mercury Ore (Mine/Mill 9202)
6. Antimony Ore (Mine/Mill 9901)
SAMPLE COLLECTION, PRESERVATION, AND TRANSPORTATION
Collection, preservation, and transportation of samples were
accomplished in accordance with procedures outlined in Appendix
III of "Sampling and Analysis Procedures for Screening of
Industrial Effluents for Priority Pollutants" (published by the
EPA Environmental Monitoring and Support Laboratory, Cincinnati,
Ohio, March 1977, revised April 1977) and in "Sampling Screening
Procedure for the Measurement of Priority Pollutants" (published
by the EPA Environmental Guidelines Division, Washington, D.C.,
October 1976). The procedures used are summarized in the
paragraphs which follow.
In general, four types of samples were collected:
Type ]_. A 24-hour composite sample, totaling 9.6 liters (2.5
gallons) in volume, was analyzed for the presence of metals,
pesticides and PCBs, asbestos, organic compounds (via gas chroma-
tography/mass spectroscopy (GC/MS) using the liquid/liquid
extraction or electron capture methods), and the classical para-
meters. Usually, this consisted of 200-ml (6.8 ounce) samples,
collected and composited at 30-minute intervals by an ISCO Model
1680 peristaltic pump automatic sampler.
When circumstances prevented the use of this sampler, 2.4-liter
(81.2 ounce) grab samples were collected and manually composited
(also non-flow proportioned) at 6-hour intervals. For example,
all tailing samples were composited because the high solids
content prevented collection of representative samples with an
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ISCO sampler. Also, in the case of one
mine/mill/smelter/refinery complex, two consecutive 24-hour
composite samples were usually collected to better characterize
the several waste streams involved.
Type 2. A 24-hour composite sample, totaling 1 liter (33.8
ounce) in volume, was analyzed for the presence of total cyanide.
This was a composite of four 250-ml (8.5 ounce) grab samples
collected at 6-hour intervals.
Type 3_. A 24-hour composite sample, totaling 0.47 liter (16
ounce) in volume, was analyzed for the presence of phenolics.
This was a composite of four 118-ml (4-ounce) grab samples,
collected at 6-hour intervals.
Type 4_. Two 125-ml (4.2 ounce) grab samples (one a backup sample
collected midway in the 24-hour sampling period) were analyzed
for the presence of volatile organic compounds by the "purge and
trap" method (discussed further under Sample Analysis later in
this section).
All sample containers were labeled to indicate sample number,
sample site, sampling point, individual collecting the sample,
type of sample (e.g., composite or grab, raw discharge or treated
effluent), sampling dates and times,, preservative used (if any),
etc.
Collection and Preservation
Screen, Verification, and Additional Sampling Programs
Whenever practical, all samples collected at each sampling point
were taken from mid-channel at mid-depth in a turbulent, well-
mixed portion of the waste stream. Periodically, the temperature
and pH of each waste stream sampled were measured on-site.
Each large composite (Type 1) sample was collected in a new 11.4-
liter (3-gallon), narrow-mouth glass jug that had been washed
with detergent and water, rinsed with tap water, rinsed with
distilled water, rinsed with methylene chloride, and air dried at
room temperature in a dust-free environment.
Before collection of Type 1 samples, new Tygon tubing was cut to
minimum lengths and installed on the inlet and outlet (suction
and discharge) fittings of the automatic sampler. Two liters
(2.1 quart) of blank water, known to be free of organic compounds
and brought to the sampling site from the analytical laboratory,
were pumped through the sampler and its attached tubing into the
glass jug; the water was then distributed to cover the interior
of the jug and subsequently discarded.
A blank was produced by pumping an additional 3 liters (3.2
quarts) of blank water through the sampler, distributed inside
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the glass jug, and poured into a 3.8-liter (1 gallon) sample bot-
tle that had been cleaned in the same manner as the glass jug.
The blank sample was sealed with a Teflon-lined cap, labeled, and
packed in ice in a plastic foam-insulated chest. This sample
subsequently was analyzed to determine any contamination con-
tributed by the automatic sampler.
Metals analyses were run by EPA and Calspan laboratories. During
collection of each Type 1 sample, the glass jug was packed in ice
in a separate plastic foam-insulated container. After the
complete composite sample had been collected, it was mixed to
provide a homogeneous mixture, and two 0.95-liter (1-quart)
aliquots were removed for metals analysis and placed in labeled,
new plastic 0.95-liter bottles which had been rinsed with
distilled water. One of these 0.95-liter aliquots was sealed
with a Teflon-lined cap, placed in an iced, insulated chest to
maintain it at 4 C (39 F), and shipped by air to EPA/Chicago for
plasma-arc metal analysis. Initially, the second sample was
stabilized by the addition of 5 ml (0.2 ounce) of concentrated
nitric acid, capped and iced in the same manner as the first, and
shipped by air to the contractor's facility for atomic-absorption
metal analysis. The Calspan analyses are reported herein because
atomic-adsorption is the preferred technique (except for some
beryllium analyses which were taken from plasma-arc data).
Because of subsequent EPA notification that the acid pH of the
stabilized sample fell outside the limits permissible under
Department of Transportation regulations for air shipment,
stabilization of the second sample in the field was discontinued.
Instead, this sample was acid-stabilized at the analytical
laboratory.
This procedure for obtaining metals samples was not used when the
waste streams sampled contained very high concentrations of
suspended solids. These solids were generally heavy and rapidly
settled out of solution. When samples to be analyzed for metal
content were collected from a high solids content stream, they
were manually collected and a separate composite sample made
(rather than being removed from the 9.6-liter (2.5-gallon) com-
posite). This was necessary to provide a representative sample
of the solids fraction.
After removal of the two 0.95-liter (1-quart) metals aliquots
from the 9.6-liter (2.5-gallon) composite sample (in the case of
low solids content samples), the balance of the sample was sealed
in the 11.4-liter (3-gallon) glass jug with a Teflon-lined cap,
iced in an insulated chest, and shipped to the Calspan laboratory
for further subdivision and analysis for non-volatile organics,
asbestos, conventional, and nonconventional parameters. Calspan
performed the extraction of organics to be analyzed by GS/MS
liquid/liquid detention and shipped the stable extracts to Gulf
South Research. If a portion of this 7.7-liter (2 gallon) sample
was requested by an industry representative for independent
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analysis, a 0.95-liter (1-quart) aliquot was placed in a sample
container supplied by the representative.
Sample Types 2 and 3 were stored in new bottles which had been
iced and labeled, one liter (33.8-ounce) clear plastic bottles
for Type 2, and 0.47-liter (16-ounce) amber glass for Type 3.
The bottles had been cleaned by rinsing with distilled water and
the samples were preserved as described below.
To each Type 2 (cyanide) sample, sodium hydroxide was added as
necessary to elevate the pH to 12 or more (as measured using pH
paper). Where the presence of chlorine was suspected, the sample
was tested for chlorine (which would decompose most of the cya-
nide) by using potassium iodide/starch paper. If the paper
turned blue, ascorbic acid crystals were slowly added and dis-
solved until a drop of the sample produced no change in the color
of the test paper. An additional 0.6 gram (0.021 ounce) of
ascorbic acid was added, and the sample bottle was sealed (by a
Teflon-lined cap), labeled, iced and shipped to Calspan for
analysis.
To each Type 3 (total phenol) sample, phosphoric acid was added
as necessary to reduce the pH to 4 or less (as measured using pH
paper). Then, 0.5 gram (0.018 ounce) of copper sulfate was added
to kill bacteria, and the sample bottle was sealed (by a Teflon-
lined cap), labeled, iced and shipped to Calspan for analysis.
Each -Type 4 (volatile organics) sample was stored in a new 125-ml
(4.2-ounce) glass bottle that had been rinsed with tap water and
distilled water, heated to 105 C (221 F) for 1 hour, and cooled.
This method was also used to prepare the septum and lid for each
bottle. Each bottle, when used was filled to overflowing, sealed
with a Teflon-faced silicone septum (Teflon side down) and a
crimped aluminum cap, labeled, and iced. Hermetic sealing was
verified by inverting and tapping the sealed container to confirm
the absence of air bubbles. (If bubbles were found, the bottle
was opened, a few additional drops of sample were added, and a
new seal was installed.) Samples were maintained hermetically
sealed and iced until analyzed by Gulf South Research.
Verification Monitoring Program
Sampling methods for the monitoring program were similar to those
used in the screening and verification efforts. However, the
monitoring samples were collected by industry personnel, in
contractor-supplied containers and per contractor instructions,
over time periods ranging from 2 to 12 weeks. Samples were
shipped to Calspan for analysis.
Surveillance and Analysis Program
As discussed previously, the samples for this program were
collected and analyzed by Agency regional personnel in three dif-
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ferent regions. Techniques were very similar to those described
above and EPA approved protocol was observed.
Cost Site Visit Sampling Program
As discussed earlier, this program was primarily designed to
collect cost data and sampling was conducted only during the
short site visit necessary to gather cost data. Therefore,
single grab samples were taken. For total metals and classical
pollutant analyses, a one-liter plastic bottle and cap were
rinsed several times with the stream to be sampled, and the
bottle was filled. For dissolved metals, a portion of this
sample was sucked by hand pump through a Millipore 0.45 micron
filter into a plastic vacuum flask to rinse the apparatus. This
water was discarded, and another quantity filtered, a portion of
which was used to rinse a half-liter sample bottle and cap, and
the remainder was poured into the rinsed bottle and sealed. The
bottles were tightened, and after fifteen minutes tightened again
and sealed with plastic tape.
The bottles were stored in a styrofoam chest with ice, and
shipped to Radian Corporation.
Sample Transportation
Bottled samples were packed in ice in waterproof plastic foam-
insulated chests which were used as shipping containers. Large
glass jugs were supported in custom fitted, foam plastic con-
tainers before shipping in the insulated chests. All sample
shipments were made by air freight.
Associated Data Collection
Drawings and other data relating to plant operations were
obtained during site sampling visits. This additional data
included detailed information on production, water use, waste-
water control, and wastewater treatment practices. Flow diagrams
were obtained or prepared to indicate the course of significant
wastewater streams. Where possible, control and treatment plant
design and cost data were collected, as well as historical data
for the sampled waste streams. Information on the use of rea-
gents or products containing chemicals designated as toxic
pollutants was also requested.
Uranium Study
Both 24-hour automatic composite samples and grab samples were
obtained, depending upon site conditions. It was necessary to
collect grab samples at several sites due to freezing conditions;
however, company personnel were consulted to ensure that the
water quality variations over a 24-hour period were insignificant
where grab samples were obtained. The samples were analyzed
using EPA methods.
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Gold Placer Mining Study
Four-to eight-hour composite samples were collected and shipped
to the laboratory for analysis of total suspended solids, mercury
and arsenic. All samples analyzed for metal parameters were
analyzed for total metal present within the sample. All analyses
were performed according to methods described in "Methods for
Chemical Analysis of Water and Waste," EPA 600/4-79-020, 1976 and
"Standard Methods for the Examination of Water and Wastewater,"
1976. Temperature, pH, conductivity, settleable solids, and flow
rate measurements were performed on-site using portable
instruments.
Titanium Sand Dredging Mining and Milling Study
Sampling and preservation methods employed at two of the
facilities studied followed the methods outlined for the
screening, verification, and additional sampling programs. The
samples were composited over three consecutive, 24-hour periods
at designated sites. The samples were analyzed for conventional
and toxic pollutants by Radian Corporation using EPA approved
methods. Grab samples for conventional and "priority metals"
were obtained from a third facility. These samples were also
analyzed by Radian Corporation.
Solid Waste Study
The actual waste streams sampled were first identified from the
available background data, and then subsequently determined from
contact with mine/mill personnel. The number of water and
wastewater sample sites chosen at each facility was dependent
upon the number of raw waste streams discharging to the treatment
system, the number and types of treatment systems utilized, and
the known characteristics of the wastewater.
Water and wastewater samples were taken from various locations
within the treatment system to obtain the most representative
samples from all segments of the system. Due to time and cost
limitations, all samples were taken as grab samples. Although
the differences between grab and composite samples cannot be
fully evaluated here, it is believed that due to the long
residence times and large ponds employed in most of the systems
studied, the differences between the sampling methods would be
slight in most cases.
Wherever possible, the samples were obtained from the middle of
the stream in a region of high turbulence to minimize solids
separation. In several instances, however, samples were taken
from clearwater areas, either because the system had low water
levels or was a zero discharge system without recycle. In these
cases, samples were obtained near the discharge structure or at a
point farthest from the pond influent.
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All samples were placed in sample bottles, preserved according to
the methods outlined in "Methods for Chemical Analysis of Water
and Wastes" EPA-600/4-79-020, and properly labeled. The
information on each bottle included the mine/mill name, the
sample site, date, and additional sampling information. This
information was also recorded in field notes. For samples sent
to IFB laboratories, the sample was assigned an EPA sample
control number, and the appropriate traffic report was completed.
Field data on pH, settleable solids, and temperature were
collected to augment the data base.
SAMPLE ANALYSIS
Sampling and Analytical Methods
As Congress recognized in enacting the Clean Water Act of 1977,
the state-of-the-art ability to monitor and detect toxic
pollutants is limited. Most toxic pollutants were relatively
unknown until only a few years ago, and only on rare occasions
has EPA regulated or has industry monitored or even developed
methods to monitor these pollutants. Section 304(h) of the Act,
however, requires the Administrator to promulgate guidelines to
establish test procedures for the analysis of toxic pollutants.
As a result, EPA scientists, including staff of the Environmental
Research Laboratory in Athens, Georgia, and staff of the Environ-
mental Monitoring and Support Laboratory in Cincinnati, Ohio,
conducted a literature search and initiated a laboratory program
to develop analytical protocols. The analytical techniques used
in this study were developed concurrently with the development of
general sampling and analytical protocols and were incorporated
into the protocols ultimately adopted for the study of other
industrial categories. See Sampling and Analysis Procedures for
Screening of_ Industrial Effluents for Priority Pollutants,
revised April 1977.
Because Section 304(h) methods were available for most toxic
metals, pesticides, cyanide and phenolics (4AAP), the analytical
effort focused on developing methods for sampling and analyses of
organic toxic pollutants. The three basic analytical approaches
considered by EPA are infrared spectroscopy (IS), gas chromato-
graphy (GO with multiple detectors, and gas chromatography/mass
spectrometry (GC/MS). Evaluation of these alternatives led the
Agency to proposed analytical techniques for 113 toxic organic
pollutants (see 44 FR 69464, 3 December 1979, amended 44 FR
75028, 18 December 1979) based on: (1) GC with selected detec-
tors, or high-performance liquid chromatography (HPLC), depending
on the particular pollutant and (2) GC/MS. In selecting among
these alternatives, EPA considered their sensitivity, laboratory
availability, costs, applicability to diverse waste streams from
numerous industries, and capability for implementation within the
statutory and court-ordered time constraints of EPA's program.
The rationale for selecting the proposed analytical protocols may
be found in 44 FR 69464 (3 December 1979).
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In EPA's judgment, the test procedures used repreesnt the best
state-of-the-art methods for toxic pollutant analyses available
when this study was begun.
EPA is aware of the continuing evolution of sampling and ana-
lytical procedures. Resource constraints, however, prevented the
Agency from reworking completed sampling and analysis efforts to
keep up with this constant evolution. As state-of-the-art tech-
nology progresses, future rulemaking will be initiated to evalu-
ate, and if necessary, incorporate these changes'.
Before analyzing ore mining and dressing wastewater, EPA defined
specific toxic pollutants for the analyses. The list of 65
pollutants and classes of pollutants potentially includes
thousands of specific pollutants, and the expenditure of
resources in government and private laboratories would be
overwhelming if analyses were attempted for all these pollutants.
Therefore, to make the task more manageable, EPA selected 129
specific toxic pollutants for study in this rulemaking and other
industry rulemakings. The criteria for selection of these 129
pollutants included frequency of occurrence in water, chemical
stability and structure, amount of chemical produced, and
availablity of chemical standards for measurement.
As discussed in Sample Collection, EPA collected four types of
samples from each sampling point: (1) a 9.6 liter, 24-hour com-
posite sample used to analyzed metals, pesticides, PCBs, asbes-
tos, organic compounds, and the classical parameters; (2) a 1-
liter, 24-hour composite sample used to analyze total cyanide;
(3) a 0.47-liter, 24-hour composite sample to analyze total
phenolics (4AAP); and (4) two 125-ml grab samples to analyze
volatile organic compounds by the "purge and trap" method.
EPA analyzed for toxic pollutants according to groups of
chemicals and associated analytical schemes. Organic toxic
pollutants included volatile (purgeable), base-neutral and acid
(extractable) pollutants, and pesticides. Inorganic toxic pollu-
tants included toxic metals, cyanide, and asbestos, (chrysotile
and total asbestiform fibers).
The primary method used in screening and verification of the
volatile, base-neutral, and acid organics was gas chromatography
with confirmation and quantification on all samples by mass
spectrometry,(GC/MS). Phenolics (total) were analyzed by the 4~
aminoantipyrine (4AAP) method. GC was employed for analysis of
pesticides with limited MS confirmation. The Agency analyzed the
toxic metals by atomic adsorption spectrometry (AAS), with flame
or graphite furnace atomization following appropriate digestion
of the sample. Samples were analyzed for total cyanide by a
colormetric method, with sulfide previously removed by distilla-
tion. Asbestos was analyzed by transmission electron microscopy
and fiber presence reported as chrysotile and total fiber counts.
EPA analyzed for seven other parameters including: pH, tempera-
132
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ture, TSS, VSS, COD, TOC, iron, aluminum, and radium 226 (total
and dissolved).
The high costs, time-consuming nature of analysis, and limited
laboratory capability for toxic pollutant analyses posed con-
siderable difficulties to EPA. The cost of each wastewater
analysis for organic toxic pollutants ranges between $650 and
$1,700, excluding sampling costs (based on quotations recently
obtained from a number of analytical laboratories). Even with
unlimited resources, however, time and laboratory capability
would have posed additional constraints. Efficiency is improv-
ing, but when this study was initiated, a well-trained technician
using the most sophisticated equipment could perform only one
complete organic analysis in an eight-hour workday. Moreover,
when this rulemaking study began only about 15 commercial labora-
tories in the United States could perform these analyses. Today,
EPA knows of over 50 commercial laboratories that can perform
these analyses, and the number is increasing as the demand does.
In plannning data generation for the BAT rulemaking, EPA con-
sidered requiring dischargers to monitor and analyze toxic pollu-
tants under Section 308 of the Act. The Agency did not use this
authority, however, because it was reluctant to increase the cost
to the industry and because it desired to keep direct control
over sample analyses in view of the developmental nature of the
methodology and the need for close quality control. In addition,
EPA believed that the slow pace and limited laboratory capability
for toxic pollutant analysis would have hampered mandatory sam-
pling and analysis. Although EPA believes that-available data
support the BAT regulations, it would have preferred a larger
data base for some of the toxic pollutants and will continue to
seek additional data. EPA will periodically review these regula-
tions, as required by the Act, and make any revisions supported
by new data.
Parameters Analyzed
Analyses varied for the different sampling programs and to
simplify the discussion, they are cited by category whenever
possible. The categories are:
Toxics
Organics (see Table V-l)
All
Total Phenolics (4AAP)
Metals (see Table V-2)
Total
Dissolved
Cyanide (total)
Asbestos
Total Fiber
133
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Chrysotile
Conventionals
Total Suspended Solids (TSS)
PH
Non-Conventionals
Temperature
Volatile Suspended Solids (VSS!
Chemical Oxygen Demand (COD)
Total Organic Carbon (TOO
Radium 226
Total
Dissolved
Total Phenolics (4AAP)
Total Settleable Solids
Miscellaneous Others
The 114 toxic organics (listed in Table V-l), the 13 toxic metals
(listed in Table V-2), the conventionals, and the
nonconventionals were analyzed in separate groups, with a few
exceptions as follows: phenolics, which were occasionally
analyzed without the other toxic organics; radium 226, which was
only analyzed at uranium facilities; and some exceptions to the
toxic metals group, to be noted in the following discussion of
specific programs.
Screen Sampling Program
The screen sampling program was designed to build the data base
on toxics. Therefore, all mine/mill samples (and most complex
samples) were analyzed for the toxics (except dissolved metals).
The raw data presented in Supplement 1 include only the
parameters detected.
All screen samples were also analyzed for the conventional and
nonconventional parameters, with the exception of uranium mines
and/or mills, which only analyzed for radium 226.
Verification Sampling Program
Sample Set 1_. From the six facilities visited in the screen
sampling program, only samples from Mine/Mill/Smelter/Refinery
3107 were screened for toxic organics. Samples from all
facilities were analyzed for total metals; two sites were
excluded from analysis of cyanide and total phenolics (4AAP).
Total fiber and chrysotile asbestos counts were performed on most
samples (primarily effluents).
All samples were screened for both conventional parameters, as
well as COD and TOC.
Sample Set 2.- This set of samples was taken to determine
specific parameters, as follows:
134
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Facility
Pb/Zn-Mine/Mill 3121
Ag-Mine/Mill 4401
Mo-Mill 6101
Al-Mine 5102
U-Mine 9408
Mill 9405
Additional Sampling Program
Parameter
CN
Asbestos
Asbestos
Phenolics,
Asbestos
Asbestos
Asbestos
Of the two facilities sampled (6102 and 9452), only samples from
Mine/Mill 6102 were screened for toxic organics, total fibers,
and chrysotile asbestos. No samples from either site were
analyzed for cyanide and total phenolics (4AAP).
Samples from both facilities were screened for the conventional
pollutants. Mine/Mill 6102 was analyzed for the nonconventional
pollutants VSS, COD, and TOC, whereas samples from Mine/Mill 9452
were tested for the nonconventional pollutants COD, chloride,
sulfate, uranium, vanadium, and radium 226.
Verification Monitoring Program
Under this program, no analysis of the toxic organics or asbestos
was initiated; however, all samples were screened for toxic
metals, cyanide and total phenol (4AAP). No sampling analysis of
parameters occurred.
conventional or
and
nonconventional
Surveillance and Analysis Program
Fourteen facilities were sampled under this program. Samples
from all ten were screened for total toxic metals and most sam-
ples were tested for total and dissolved toxic metals, cyanide,
and total phenolics (4AAP). Of the ten facilities, only samples
from Mine/Mill 6107 were analyzed for toxic organics. None of
the samples were tested for asbestos.
All samples were screened for conventional pollutants, as well as
a wide assortment of nonconventional pollutants, including
magnesium, manganese, cobalt, tin, barium, ammonia, settleable
solids, and others.
Cost Site Visit Program
The samples were
solved), but not
analyzed for all toxic metals (total and dis-
for toxic organics, or cyanide and asbestos.
The samples were screened for both conventionals, and a variety
of nonconventionals, such as alkalinity, settleable solids,
manganese, and iron.
Uranium Study
135
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Ten uranium mining and milling facilities were sampled during
this study. Influent, effluent and intermediate waste streams
were analyzed. The wastewater was analyzed for total suspended
solids (TSS), cadmium, zinc, arsenic, copper, uranium, molybde-
num, vanadium, chemical oxygen demand (COD), pH, sulfate and both
total and dissolved radium 226.
Gold Placer Mining Study
Conventional and nonconventional parameters including tempera-
ture, pH, conductivity, settleable solids, and total suspended
solids were measured. Two toxic metals (arsenic and mercury)
were analyzed for total metals present within the sample.
Titanium Sand Dredge Mining and Milling Study
Samples from two facilities were analyzed for toxic organics,
toxic metals (total and dissolved), cyanide and asbestos. Toxic
metals (total and dissolved), cyanide and asbestos were also
measured at a third facility. Conventional parameters and non-
conventional parameters such as chemical oxygen demand, total
organic carbon, total phenolics (4AAP), iron, manganese,
titanium, and oil and grease were also measured.
Solid Waste Program
Wastewater samples were analyzed for conventionals, toxic metals
(total and dissolved), and asbestos at each facility sampled.
Nonconventionals including chemical oxygen demand, total organic
carbon, total phenolics (4AAP), manganese, iron, settleable
solids, temperature, and others were measured.
Analytical Methods, Laboratories, arid Detection Limits
All parameters were analyzed using methods or techniques
described in:
1. "Sampling and Analysis Procedures for Screening of
Industrial Effluents for Priority Pollutants" (published
by EPA Environmental Monitoring and Support Laboratory,
Cincinnati, Ohio, March 1977, revised April 1977).
2. "Analytical Methods for the Verification Phase of the
BAT Review (Additional References)" (published by EPA
Environmental Guidelines Division, Washington, D.C.,
June 1977).
3. "Methods for Chemical Analysis of Water and Waste"
(published by EPA Environmental Monitoring and Support
Laboratory Cincinnati, Ohio, 1974, revised 1976).
4. "Standard Methods for the Examination of Water and Waste
Water" (published by the American Public Health
136
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Association, Washington, D.C., 14th edition, 1976).
5. Other EPA approved methods cited in "Guidelines
Establishing Test Procedures for the Analysis of
Pollutants" (Federal Register, Vol. 41, No. 232, 1
December 1976, pp. 52780-52786).
6. Asbestos analyses were performed using the method
outlined in "Preliminary Interim Procedure for Fibrous
Asbestos" (published by EPA, Athens, Georgia, undated)/
(Note: In accordance with a 13 December 1977 EPA
letter, anthracene was added to the sample by Gulf South
Research Institute for analysis of organic compounds by
GC/MS using the liquid/liquid extraction method.)
The choice of laboratories depended on the analysis to be
conducted, and laboratories were changed for different sampling
programs.
Detection limits (the lowest concentration at which a parameter
can be quantified) vary between sampling programs and even within
a program. The detection limit (DL) for a parameter depends on
the particular instrument used, the range of standards each set
of samples analyzed, the complexity of the sample matrix, and
optimization of the instrument response. Therefore, the
detection limits given herein are only indicators of the minimum
quantifiable concentrations. In fact, some data points are
reported below the listed detection limit.
Screen, Verification and Verification Monitoring Programs
The analysis methods, laboratories, and detection limits for
samples analyzed during these programs are given in Tables V-3
and V-4. (All of the parameters listed were not analyzed in all
four programs.)
Surveillance and Analysis Program
As discussed, this program was conducted by the EPA regional
laboratories. The methods used were EPA approved and the detec-
tion limits are approximately those shown in Table V-4.
Cost Site Visit Sampling Program
The cost site visit data were generated by Radian Corporation and
the methods and detection limits are given in Table V-5.
Uranium, Gold Placer Mining, Titanium Sand Dredging Mining and
Milling, and Solid Waste Programs
137
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During these programs many different laboratories were utilized
each using EPA approved analytical methods. The detection limits
reported are approximately those shown in Table V-4.
Quality Control
Quality control measures used in performing all analyses con-
ducted for this program complied with the guidelines given in
"Handbook for Analytical Quality Control in Water and Wastewater
Laboratories" (published by EPA Environmental' Monitoring and
Support Laboratory, Cincinnati, Ohio, 1976). As part of the
daily quality control program, blanks (including sealed samples
of blank water carried to each sampling site and returned
unopened, as well as samples of blank water used in the field),
standards, and spiked samples were routinely analyzed with actual
samples. As part of the overall program, all analytical instru-
ments (such as balances, spectrophotometers, and recorders) were
routinely maintained and calibrated.
The atomic-absorption spectrometer used for metals analysis was
checked to see that it was operating correctly and performing
within expected limits. Appropriate standards were included
after at least every 10 samples. Also, approximately 15 percent
of the analyses were spiked with distilled water to assure
recovery of the metal of interest. Reagent blanks were analyzed
for each metal, and sample values were corrected if necessary.
Total Phenolics (4AAP)
The quality control for total phenolics (4AAP) analysis included
demonstrating the quantitative recovery of each phenol
distillation apparatus by comparing distilled standards to
nondistilled standards. Standards were also distilled each
analysis day to confirm the distillation efficiency and purity of
reagents. Duplicate and spiked samples were run on at least 15
percent of the samples analyzed for total phenolics.
Cyanide
Similarly, recovery of total cyanide was demonstrated with each
distillation-digestion apparatus, and at least one standard was
distilled each analysis day to verify distillation efficiency and
reagent purity. Quality control limits were established.
During this program, problems were frequently encountered with
quality control and analysis of cyanide in mining wastewater
samples using the EPA approved Belack Distillation method. Both
industry experience and the contractor's laboratory experience
indicated problems in obtaining reliable results at the concen-
trations typically encountered in the metal ore mining industry.
Quality control for cyanide included analysis of duplicate
samples within a single laboratory and between two laboratories,
analysis of spiked samples, and analysis by two different
138
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methods. Analysis of duplicated samples often produced results
that varied by a factor of 10 (Table V-6). Two commercial
laboratories evaluated the analytical recovery of cyanide from
mill tailing pond decant samples which had been spiked with
cyanide. Samples were delivered to these laboratories immedi-
ately after collection to eliminate the possibility of cyanide
loss. The results of this quality control program are presented
in Table V-7.
A study of the analysis of cyanide in ore mining and processing
wastewater was conducted in cooperation with the American Mining
Congress to investigate the causes of analytical interferences
observed and to determine what effect these interferences had on
the precision of the analytical method. Samples of five ore
mining and processing wastewaters were obtained along with two
municipal wastewater effluents. Cyanide analyses were performed
by eight laboratories, including six AMC representatives, EPA's
EMSL laboratory in Cincinnati and Radian Corporation's chemical
laboratory.
The purpose of this study was to evaluate the EPA-approved method
and a modified method for the determination of cyanide. The
modified method employed a lead acetate scrubber to remove
sulfide compounds produced during the reflux-distillation step.
Sulfides have been suspected of providing an interference in the
colorimetric determination of cyanide concentrations. Also,
several samples were spiked with thiocyanate to ascertain if this
compound caused interference in the cyanide analysis.
A statistical analysis of the resultant data shows no significant
difference in precision or accuracy of the two methods employed
when applied to metal ore mining and milling wastewaters having
cyanide concentrations in the 0.2 mg/1 to 0.4 mg/1 range. Based
upon the statistical analysis, approximately 50 percent of the
overall error of either method was attributed to intralaboratory
error. This highlights the need for an experienced analyst to
perform cyanide analyses.
Initial cyanide concentrations were determined by the EPA
approved cyanide analysis method. Samples which were found to
contain less than 0.2 mg/1 cyanide were spiked with sodium
cyanide and potassium ferricyanide. Samples containing over 0.5
mg/1 cyanide were air sparged at pH 2 for 24 hours and filtered
through 0.45u membrane filters prior to raising the pH above 10.
The following table summarizes the initial and adjusted cyanide
concentrations.
139
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Initial CN Theoretical CN
by Approved Method After Adjustment
Sample (mg/1) (mg/1)
A 0,26 0.26 (no adjustment)
B 0.02 0.210
C 0.02 2.73
D 0.54* 0.54 (no adjustment)
E 0.02 0.389
F 84.00 2.5 - 3.5**
G 0.02 0.273
*Sample showed strong interferences during colorimetric
procedure.
**Results were not repeatable.
Sample D contained approximately 180 mg/1 of thiocyanate.
Samples B and E were spiked with 33 mg/1 and 100 mg/1 of thiocya-
nate, respectively. Sodium thiocyanate was used as the
thiocyanate source.
The results of this study are summarized in Table V-8. Based
upon these data, the following conclusions have been drawn for
ore mining and processing wastestreams containing 0.2 to 0.4 mg/1
cyanide:
1 . The overall relative standard deviation for the
EPA-approved method was 27.6 percent.
2. The overall relative standard deviation for the modified
method was 30.4 percent.
3. Accuracy as average percent deviation from the standards
was -12.6 for the approved method and -6.1 for the
modified method. Neither of these values is signifi-
cantly different from zero for this sample size.
4. The approved and modified methods work equally well for
the analysis of cyanide in ore mining and processing
wastewaters.
5. No major problems were demonstrated in the cyanide
analysis by either method for samples containing 30 to
TOO mg/1 of thiocyanate.
Based upon the relative standard deviations calculated, it can be
said that for an ore mining or processing wastewater sample
containing 0.2 mg/1 of cyanide, 95 percent of the analyses would
be between 0.08 and 0.32 mg/1 using the modified method and
between 0.09 and 0.31 mg/1 using the approved method. Over 99
percent of the analyses would be between 0.02 and 0.38 mg/1 using
the modified method and between 0.035 and 0.365 mg/1 using the
approved method (Reference 3). Accordingly, the Agency must
140
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allow for analytical measurement of up to .4 mg/1 for total
cyanide. (See discussion Section X).
PCBs and Pesticides
In analyses of pesticides and PCBs, extraction efficiencies were
determined. Known standards were prepared, extracted, and
analyzed to guarantee minimum extraction/concentration losses.
Calibrations of all analytical components were carried out with
high-quality pure materials. A calibration mixture was run daily
to ensure that retention time and instrumentation response for
each parameter analyzed did not change due to column and detector
aging.
141
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Table V-l
TOXIC ORGANICS
Compound Name
1. *acenaphthene (B)***
2. *acrolein (y)***
3. *acrylonitrile (V)
4. *benzene (V)
5. *benzidene (B)
6. *carbon tetrachloride (tetrachloromethane) (V)
^Chlorinated benzenes (other than dichlorobenzenes)
7. chlorobenzene (V)
8. 1,2,4-trichlorobenzene (B)
9. hexachlorobenzene (B)
^Chlorinated ethanes(including 1,2-dichloroethane,
1,1,1-trichloroethane and hexachloroethane)
10. 1,2-dichloroethane (V)
11. 1,1,1-trichlorethane (V)
12. hexachlorethane (B)
13. 1,1-dichloroethane (V)
14. 1,1,2-trichloroethane (V)
15. 1,1,2,2-tetrachloroethane (V)
16. chloroethane (V)
*Chloroalkyl ethers (chloromethyl, chloroethyl and
mixed ethers)
17. bis (chloromethyl) ether (B)
18. bis (2-chloroethyly) ether (B)
19. 2-chloroethyl vinyl ether (mixed) (V)
^Chlorinated naphthalene
20. 2-chloronaphthalene (B)
*Chlorinated phenols (other than those listed elsewhere;
includes trichlorophenols and chlorinated cresols)
21. 2,4,6-trichlorophenol (A)***
22. parachlorometa cresol (A)
23. ^chloroform (trichloromethane) (V)
24. *2-chlorophenol (A)
142
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Table V-l (Continued)
TOXIC ORGANICS
*Dichlorobenzenes
25. 1,2-dichlorobenzene (B)
26. 1,3-dichlorobenzene (B)
27. 1,4-dichlorobenzene (B)
*Dichlorobenzidine
28. 3,3'-dichlorobenzidine (B)
*Dichloroethylenes (1,1-dichloroethylene and
1,2-dichloroethylene)
29. 1,1-dichloroethylene (V)
30. 1,2-trans-dischloroethylene (V)
31. *2,4-dichlorophenol (A)
*Dichloropropane and dichloropropene
32. 1,2-dichloropropane (V)
33. 1,2-dichloropropylene (1,3-dichloropropene) (V)
34. *2,4-dimenthylphenol (A)
*Dinitrotoluene
35. 2,4-dinitrotoluene (B)
36. 2,6,-dinitrotoluene (B)
37. *1,2-diphenylhydrazine (B)
38. *ethylbenzene (V)
39. *fluoranthene (B)
*Haloethers (other than those listed elsewhere)
40. -chlorophenyl phenyl ether (B)
41 ."bromophnyl phenyl ether (B)
42. bis(2-chloroisopropyl) ether (B)
43. bis(2-chloroethoxy; methane (B)
*Halomethanes (other than those listed elsewhere)
44. methylene chloride (dichloromethane) (V)
45. methyl chloride (chloromethane) (V)
46. methyl bromide (bromomethane) (V)
47. bromoform (tribromomethane) (V)
48. dichlorobromomethane (V)
143
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Table V-l (Continued)
TOXIC ORGANICS
49. trichlorofluoromethane (V)
50. dichlorodifluoromethane (V)
51. chlorodibromomethane (V)
52. *hexachlorobutadiene (B)
53. *hexachlorocyclopentadiene (B)
54. *isophorone (B)
55. *naphthalene (B)
56. *nitrobenzene (B)
*Nitrophenols (including 2,4-dinitrophenol and dinitrocesol)
57. 2-nitrophenol (A)
58. 4-nitrophenol (A)
59. *2,4-dinitrophenol (A)
60. 4,6-dinitro-o-cresol (A)
*Nitrosamines
61. N-nitrosodimethylamine (B)
62. N-nitrosodiphenylamine (B)
63. N-nitrosodi-n-propylamine (B)
64. *pentachlorophenol (A)
65. *phenol (A)
*Phthalate esters
66. bis(2-ethylhexyl) phthalate (B)
67. butyl benzyl phthalate (B)
68. di-n-butyl phthalate (B)
69. di-n-octyl phthalate (B)
70. diethyl phthalate (B)
71. dimethyl phthalate (B)
*Polynuclear aromatic hydrocarbons
72. benzo (a)anthracene (1,2-benzanthracene) (B)
73. benzo (a)pyrene (3,4-benzopyrene) (B)
74. 3,4-benzofluoranthene (B)
75. benzo(k)fluoranthane (11,12-benzofluoranthene) (B)
76. chrysene (B)
77. acenaphthylene (B)
78. anthracene (B)
79. benzo(ghi)perylene (1,12-benzoperylene) (B)
80. fluorene (B)
81. phenathrene (B)
144
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Table V-l (Continued)
TOXIC ORGANICS
82. dibenzo (a,h)anthracene (1,2,5,6-dibenzanthracene) (B)
83. indeno (1,2,3-cd)(2,3,-o-phenylenepyrene) (B)
84. pyrene (B)
85. *tetrachloroethylene (V)
86. *toluene (V)
87. *trichloroethylerie (V)
88. *vinyl chloride (chloroethylene) (V)
Pesticides and Metabolites
89. *aldrin (P)
90. *dieldrin (P)
91. *chlordane (technical mixture and metabolites) (P)
*DDT and metabolites
92. 4,4'-DDT (P)
93. 4,4'-DDE(p,p'DDX) (P)
94. 4,4'-DDD(p,p'TDE) (P)
*endosulfan arid metabolites
95. a-endosulfan-Alpha (P)
96. b-endosulfan-Beta (P)
97. endosulfan sulfate (P)
*endrin and metabolites
98. endrin (P)
99. endrin aldehyde (P)
*heptach1or _and metabolites
100. hfM.,;achlor (P)
101. h'.-p each lor epoxide (P)
*hexachlorocyclohexane (all isomers)
102. a-BHC-Alpha (P) (B)
103. b-BHC-Beta (P) (V)
104. r»BHC (lindane)-Gamma (P)
105. g-BHC-Delta (P)
145
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Table V-l (Continued)
TOXIC ORGANICS
*polychlorlnated biphenyls (PCB's)
106. PCB-1242 (Arochlor 1242) (P)
107. PCB-1254 (Arochlor 1254) (P)
108. PCB-1221 (Arochlor 1221) (P)
109. PCB-1232 (Arochlor 1232) (P)
110. PCB-1248 (Arochlor 1248) (P)
111. PCB-1260 (Arochlor 1260) (P)
112. PCB-1016 (Arochlor 1016) (P)
113. *Toxaphene (P)
114. **2,3,7,8-tetrachlorodibenzo-p-dioxin (TCDD)
*Specific compounds and chemical classes as listed in the
consent degree.
**This compound was specifically listed in the consent degree,
= analyzed in the base-neutral extraction fraction
V = analyzed in the volatile organic fraction
A = analyzed in the acid extraction fraction
146
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Table V-2
TOXIC METALS, CYANIDE AND ASBESTOS
1. *Antimony (Total)
2. *Arsenic (Total)
3. *Asbestos (Fibrous)
4. *Beryllium (Total)
5. *Cadmium (Total)
6. *Chromium (Total)
7. *Copper (Total)
8. *Cyanide (Total)
9. *Lead (Total)
10. *Mercury (Total)
11. *Nickel (Total)
12. *Selenium (Total)
13. *Silver (Total)
14. *Thallium (Total)
15. *Zinc (Total)
^Specific compounds and chemical classes as listed in the
consent degree.
147
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TABLE V-3. POLLUTANTS ANALYZED AND ANALYSIS TECHNIQUES/LABORATORIES
SUBSTANCE
pH
total *uip*nd«d tolidt
volalik lutptndwf x>l»di
COD
TOC
radium 226 (toul)
radium 226 (dittoU-adl
antimony (total)
arMnic (total}
beryllium (total)
cadmium (tout)
cniomium {total)
copper ho 1*1)
lead (total)
mercury (total)
nickel 1 tot jill
*eler..um (tota;)
Silver (total)
thallium (total)
line (to til)
*tbettoi (fibrous)
cyanide (total)
phenol (total)
aldnn
dieldnn
chlordane
(technical mixture and metabolites)
' 4. 4 DDT
4.4' DDE (pjJ' DDX)
4.4 ODD rob4flMne
3,3'«}tchloroMniidin«
l,l-did>k)fDcThyiene
1 .2-trans-dich!oro*1hy)cn«
2.4^jic^kirophcnol
1 .Z-dtchloropropane .
1 .3-dtcMotoftroPYltrtt
(1,3-dichloropropene)
7,4-dimethylphenol
2.4-ffmitrotolucne
2.6!ol«n«
l,2
1 bis (chloromethyl) ether
j bis IcMofoeihyl) dhe>
| 2-ehloroelhyI vinyl tiller irrmed)
I 2.4.6 tnchlorophertoi
hTchlorophenol
j 1,2-drchrofoberwentr
LL J ]! acenaphthylen* ' |
Li. j jj anthracene 1 1
PT 1 jj 1.1? ben^operylene j j
PT j '; fluo^n.- [
LL ] 1] phpnanihrene J !
PT | 'j 1,2 5.6-diben?
-------
TABLE V-4. LIMITS OF DETECTION FOR POLLUTANTS ANALYZED
SUBSTANCE
pH 1
total suspended aelidi
volatile suipended solidi
COO
TOC
radium 226 (ton!)
radium 226 (dissolved)
antimony (total)
arsentc (total)
beryllium (total!
cadmium (to*.»l)
chromium (total)
copper {total )
l«ad (total)
[mercurv (total)
rttckel (total)
selenium (total)
silver (total)
thallium (total)
rinc (toul)
asbestos (fibrousi (total)
j cyanide (total)
phenol Itonl)
| aldrin
dieldrirr
[ chlordara
(technical mixture and metabolites)
4.4' DDT
4,4'.DD£ Ipjj' DDX]
4«' ODD ipjj' TDE)
oelhane
1 ,^.1-inchloroeUiane
hevachloroethane
1 ,» tJ:eM jfoethan*
1 .1 ,2-trichloroethane
1 .1 .2.2-le^afhloroethane
c^loro«thane
ini (chloromethyU ether
bit (chloroethyl) ether
2-ehloroethyl vinyl ether imiKed)
2-chloron»phthalpne
2 ,4 ,6-lnchlOTOphenol
parachlorometa crefol
chloroform (trichloromethane)
2-chlorophenol
' ,2-dichloroheniene
CONCENTRATION
-
lme/1
Img/l
2mi/l
Img/l
1 pCi/l
1PO/I
0.2 ran/1
0.002 mj/l
0.005 tng/l
0.002 rng/l
0.02 mg/l
0.01 mg/l
O.OS mg/t
O.OOOSmg/l"
0.02 mg/l
0,002 mg/l
0.01 tng/l
0.1 mg/l
0.005 mg/l
2.2 » 10s fiban/l
0.02 mg/l
0.002 mg/l
0.1 0g/l
O.Sug/1
1(1 1/'
1 «g/l
1(-g/l
lj.9/1
1*9/1
1 W9/I
IfS/l
O.Sjjg/l
0.5 ug/l
C.I M9"
o.i ue/i
0.1 t/g/l
0.1 pg/l
0.1 )ig/l
0.1 ug/l
1«g/l
1(ig/l
0.03 P9/I
no estimate
no estim&te
0.04 ftg/\
- 2S.Oig/i ,
1.00|ig/l
0. 10O.20 (j 8/1
SUBSTANCE
U-dkhlorobenzane
1 ,4-dichlorobenzenc
3^'^ichlorobcnzidine
l,!-dichloro«thyl«rie
1 ^-trant^lichlorotthylene
2.4<) ichlocopricnol
1 ^^jichloropropcne
1 .Sxlichloi'opropylvne
(1 ,3-dichloropropine)
2 ,4^im«thylphenol
2,4xjinitrotoluene
2.6-dinitrotoluene
1 .2romophenyl phenyl ether
bis (2-chloroisopropyl) ether
bis (2^hloroethoxy) methane
methylene chloride (dichloromethanc)
methyl chloride (chloromethanc)
methyl bromide (bromomethane)
bromoform (tribromomethane)
dichlorofaromomethane
trichlorofluoromethane
dichlorodifluororomethane
chlorod ibromomethane
hexachlorobuudiene
hexaohlorocydopentadMne
'sophorone
napthalene
nitrobenzene
2^iitrophenol
4^iitrophertol
2,4^initrophenol
4 ,6-dmmo-o-cresol
N-nitrosodimetriylamine
N ^itrosotjiphenylamine
N niuosodi-n-propylamine
pentachlorophenol
phenol
bii (2«thylhexyl) phthalate
butyl benzyl phthalate
d> n-butyl phthalate
diethyl phthalate
dimethyl phthalate
1 ,2-benzanthracene
benzo (a) pyrene (3,4-benzopvrene)
3.4 benzofluoiathene
11.1 2-benzof luoranthenc
chryjene
acenaphthylene
anthracene
1 ,12-benzoperylene
fluoiene
phenanthrene
1.2:5.6-dibenzarithracene
indeno (1.2.3-c^il pyrene
pyrene
2,3,? .8 tetrachlorodibenzo-p-dioxtn
(TCODI
tetrachloroethylene
toluene
trichloroethylene
vinyl chloride
tonaphene
CONCENTRATION
0.104.20 (((l/l
0.104.20 «g/l
~2B.OJ|9/I
0.50 ua/l
0.35 K a/1
- 1.0«g/l
0.02 ^ig/l
0.3S fig/l
0.40 utll
O.X ug/l
0.7S (ig/l
0.20 (ig/l
0.10 (K/l
0.20 (ig/l
0.06 ug/l
~ 0 50 jig/I
noflttimate
no astiiTMIa
0.08 MO/I
0.35 (ig/l
0.08 (ig/l
0.40 (ig/l
0.05 ug/l
- 0.10 utl\
~ 0.10 »g/l
~ 0.10 ug/l
0.08 ug/l
noattimate
0.10 (19/1
0.15 ug/l
0.85(11/1
1.0(11/1
no estimatt
twill not chromatogrtph
will not chromatognph
not etoblished
not established
not established
~ 50.0 ug/l
1.0 ug/l
0.20 ug/l
0.26 ug/l
OJ4).4ug/l
0.20 ug/l
0.35 ug/l
0.05 ug/l
~ 0.5 ug/l
~ 0.5 ug/l
- 0.5 uo/l
0.20 u«/l
~ 0.5 (19/1
0.05 ug/l
- 0.5 (ig/l
0.10 (ig/l
0.06 /ig/l
~ 0.5 ug/l
~ 0.5 ug/l
0.40 jig/!
no astimate
1.10 U9/'
0.35 ug/l
0.35 ug/l
-v 0.1 ug/l
not established
'vAi w^yidd by stomic-ebAO'ptkm ip«ctro*copy iniftg gaieoui hydride U flame tnethod!
** As ffin*lv]!«d by *tomtc-»Uofption spectroscopy utmg cold vipor tcehntque (a fUmelei* method)
149
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Table V-5
LIMITS OF DETECTION FOR POLLUTANTS ANALYZED FOR COST -
SITE VISITS BY RADIAN CORPORATION
Detection
Parameter Method of Analyses Limit (ppm)
Sb AAS* - hydride generation 0.005
As AAS - hydride generation 0.002
Be AAS - flameless - HGA** 0.001
Cd ICPES*** 0.005
Cr ICPES 0.005
Cu ICPES 0.005
Fe ICPES 0.004
Pb AAS - flameless - HGA 0.002
Hg AAS - flameless - cold vapor 0.001
Mn ICPES 0.001
Ni ICPES 0.020
Se AAS - hydride generation 0.005
Ag ICPES 0.005
Tl AAS - flameless - HGA 0.002
Zn ICPES 0.002
^Atomic Absorption Spectrophotometry
**Inductively Coupled Argon Plasma Emission Spectrometry
***Heated Graphite Analyzer
150
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TABLE V-6. COMPARISON OF SPLIT SAMPLE ANALYSIS* FOR CYANIDE
BY TWO DIFFERENT LABORATORIES USING THE BELACK
DISTILLATION/PYRIDINE-PYROZOLONE METHOD
Sample Description
1. Tailing Pond Influent
2. Tailing Pond Influent
3. Tailing Pond Influent
4. Tailing Pond Effluent
5. Tailing Pond Effluent
6. Tailing Pond Effluent
Analytical Result (mg Total CN/I)
Lab#1
0.02
0.05
0.08
0.04
0.04
0.03
Lab #2
0.06
<0.02
0.03
0.50
0.47
0.57
All samples collected at the same time at Mine/Mill 3121
151
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TABLE V-7. ANALYTICAL QUALITY CONTROL PERFORMANCE OF COMMERCIAL
LABORATORY PERFORMING CYANIDE* ANALYSES BY EPA APPROVED
BELACK DISTILLATION METHOD
SAMPLE DESCRIPTION
Distilled H.,O + 0.1 mg/l CN
as NaCN
Distilled H,O + 0.2 mg/l CN
as NaCN
Distilled H2O + 0.4 mg/l CN
as NaCN
Tailing Pond Decant (no spike)
Tailing Pond Decant Spiked with
0.1 mg/l CN as NaCN
Tailing Pond Decant Spiked with
0.2 mg/l CN as NaCN
Tailing Pond Decant Spiked with
0.4 mg/l CN as NaCN
Tailing Pond Decant Spiked with
1.0mg/ICNasK3Fe{CN)6
Tailing Pond Decant Spiked with
1.0mg/ICNasK4Fe(CN)6
CYANIDE AS
CN mg/l
0.125
0.250
0.200
0.069
0.144
0.119
0.136
0.028
0.032
0.027
0.079
% ANALYTICAL
RECOVERY OF SPIKE
125
125
50
-
85
44
29
3
3
3
7
*AII cyanide analyses are total cyanide
152
-------
Table V-8
SUMMARY OF CYANIDE ANALYSIS DATA FOR SAMPLES
OF ORE MINING AND PROCESSING WASTEWATERS
Number Theoretical Mean Standard Intralaboratory
of CN concentration Recovery Deviation Standard Deviation
Sample Method* Labs ES/I
B M 7 0.176 0.080 0,042
0.210
A 3
E M 7
0.389
A 3 0.338 0.064 0.025
C M 7 3.04 1.15 1.87
2.73
^ A ^
en A O
OJ
F M 7
A 3 3.27 0.283 0.431
A M 7 0.268 0.062 0.049
0.26
A 3
G M 7
0.273
A 3 0.285 0.018 0.053
DM7
***
A 2
mg/1
0.176
0.139
0.353
0.338
3.04
2.72
2.66
3.27
0.268
0.233
0.290
0.285
0.269
0.175
mg/1
0.080
0.059
0.117
0.064
1.15
0.89
1.23
0.283
0.062
0.073
0.030
0.018
0.123
0.108
*A = EPA Approved Cyanide Analysis Method
M = Modified Cyanide Analysis Method
**Results were not repeatable in the initial analysis
***Strong interferences were noted in the initial analysis
-------
154
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SECTION VI
WASTEWATER CHARACTERIZATION
The data base developed during the sampling program described in
Section V is presented in Supplement A and summary tables are
presented and discussed in this section. Also, a summary of
reagent usage at flotation mills, the largest users of mill
process chemicals, is presented to evaluate mill reagents as
potential sources of toxic pollutants. Special circumstances;
such as, the presence of certain toxic pollutants in mine water
as a result of backfilling mines with mill tailings, are
discussed at the end of this section.
SAMPLING PROGRAM RESULTS
The analytical results of the nine sampling programs discussed in
Section V are presented in Supplement A and were entered into a
computerized data base. Using this data base, summary data
tables were generated for the entire category; and each
subcategory, subdivision, and mill process (Tables VI-1 through
VI-18, which may be found at the end of this section). These
tables include raw and treated wastewater data; and the range of
pollutant concentrations observed is indicated by the mean and
median values, and the 90 percent 'and maximum values (defined
below).
All Subcategories Combined
Table VI-1 summarizes the BAT data base for all the mines and
mills in all subcategories in the ore mining and dressing point
source category. As indicated by the table, only 27 of the toxic
organics were detected in the category's treated wastewater.
Organic compounds are not found naturally with metal ores.
Introduction of organics during froth flotation mill processing
is discussed later in this section. Otherwise, the discussion of
toxic organics is left to Section VII, Selection of Pollutant
Parameters.
Toxic metals are naturally associated with metal ores and all of
the 13 toxic metals were found in wastewater from the category.
The concentrations of each metal varied greatly, as expected for
such a diverse category. Cyanide and asbestos, also toxic
parameters, were observed in many samples and in varied
concentrations.
The conventional parameters observed were primarily those
regulated by BPT effluent guidelines, that is TSS and pH. The
TSS values are very high in many raw samples because tailings
samples which typically run in the tens of thousands of mg TSS/1
are included in "raw" samples. Effluent TSS values vary, but are
generally low indicating good solids settling characteristics.
155
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Values of pH vary, but are often in the alkaline range (7 to 14).
This is because several mill processes operate at elevated pHs.
As indicated by discussions in Section III, pH, TSS, and metals
values are closely allied. The solubility of many metals varies
greatly with pH, and the status of the metals (dissolved v.
solubilized) affects the concentration of TSS. This relationship
is used by the industry for ore beneficiation and for wastewater
treatment.
Nonconventional parameters such as COD, TOC, volatile suspended
solids (VSS), and iron were also analyzed for many samples. The
concentrations of the organic related parameters, COD, TOC, and
VSS, were always low. Any organic compounds added in mill
processes are not indicated by these tests which are designed to
measure relatively large masses of organics (in the mg/1 range at
a minimum). Iron is common in metal ores and the summary table
reflects this.
The entire BAT data set is discussed below by subcategory,
subdivision, and as a mill process or mine drainage, and these
discussions more completely characterize mine/mill wastewaters.
In general it can be noted from Table VI-1 that organic compounds
are not the major concern in this category (a point discussed
thoroughly in Section III), metals are prevalent, pH values are
generally alkaline, and cyanide and asbestos are often present.
Iron Subcategory, M_ine_ Drainage Subdivision
Table VI-2 summarizes the data for iron mines. Many of the toxic
metals were not detected in the one or two available samples;
arsenic (.005 mg/1) copper (.090 and 120 mg/1), and zinc (.018
and .030 mg/1) are the exceptions. Asbestos fibers, both total
and chrysotile, were detected in relatively small amounts
compared to the rest of the category (see Table VI-1).
Generally, (comparing Tables VI-1 and VI-2) iron mine water is
characterized by low pollutant levels. This is true of most mine
water and is the reason for separate mine and mill subdivisions.
Iron Subcategory, Mill Subdivision, Physical and/or Chemical Mi11
Processes
As indicated in Table VI-3, several of the toxic metals were
present in the one or two raw samples taken, but most are removed
by existing treatment technologies (sedimentation) and were not
detected in discharge samples. Copper is the least affected by
current treatment methods. Asbestos was detected in relatively
high concentrations in the raw sample (compared to Table VI-1),
and in lower concentrations in the discharged sample. This indi-
cates that current treatment methods are removing a portion of
the asbestos; a conclusion supported by Table VI-3. The COD,
VSS, and TOC (indicators of gross organic pollution) are somewhat
higher than the rest of the industry (compared to Table VI-1),
but they are effectively removed by current technologies. Iron
156
-------
was detected in one raw sample, as expected for iron mills, but
was below detection in the discharge water. Several toxic
metals, asbestos, TSS, and some nonconventional parameters were
found in the raw wastewater of iron mills, but these parameters
were reduced during treatment and many do not appear in the
discharge water,
Copper/Lead/Zinc/Gold/Silver/Platinum/Molybdenum Subcategory,
Mine Drainage Subdivision
This subcategory includes more mines than any other subcategory
and more samples are available for characterization than for
other subcategories. As shown in Table VI-4, all of the toxic
metals were detected at least four times in sixteen raw samples.
High median concentrations (relative to the other metals
detected) of antimony, arsenic,, cadmium, chromium, copper, lead,
nickel, thallium, and zinc are shown in Table VI-4 for raw mine
drainage. In the discharged water, however, the metals
concentrations are lower, with the median values ranging from not
detected to 280 ug/1 (zinc).
Cyanide, asbestos, and phenolics are other toxic parameters
detected in this subdivision. Cyanide is used in the froth flo-
tation process and backfilling mines with mill tailings can cause
cyanide to pollute the mine water. Asbestos, being a mineral, is
found with many metal ores, although the concentrations reported
in Table VI-4 are relatively low (compared to Table VI-1) and
have a small range for samples taken at many types of mines.
Phenolics were detected at low concentrations.
Copper/Lead/Zinc/Silver/Gold/Platinum/Molybdenum Subcategory,
Cyanidation M_i_l_l_ Process
This subdivision was regulated as no discharge of process
wastewater in BPT effluent guidelines, therefore, few samples
were taken in BAT sampling programs and no discharge samples were
taken. It can be seen from Table VI-5 that many toxic para-
meters, including cyanide, were found in high concentrations in
this mill water; thereby supporting the BPT no discharge
requirement.
Copper/Lead/Z i nc/S i1ver/Go1d/P1at i num/Mo1ybdenum Subcategory,
MJJ_1_ Subdivision, Frgtji Flotation Mill Process
There were more samples of this mill process than of any others
because froth flotation is a widely used process with the
potential to generate wastewater polluted with many toxics. As
seen in Table VI-6, all of the toxic metals were detected in raw
mill water. The number of detections ranged from 7 to 78 out of
78 samples and median concentrations ranged from 1.1 ug/1 (mer-
cury) to 63,300 ug/1 (copper). These wide ranges are due to the
variations in the ore milled at different locations. Generally,
the metals concentrations are in the high range of values
157
-------
reported for the category as a whole (Table VI-1). The
discharged concentrations of metals are, generally, one or two
orders of magnitude lower than the raw values. The number of
toxic metals with median concentration over 20 ug/1 are reduced
from ten in raw samples to five in treated samples and, overall,
the concentrations are reduced by existing treatment.
Asbestos, cyanide, and phenolics were also detected in both raw
and discharged samples. Median values for all were above the
respective medians for the whole category (Table VI-1). All were
reduced by the existing treatment systems.
Nonconventional parameters and TSS were generally high (compared
to Table VI-1) and the pH range is great.
Generally, mill water and tailings from this mill process contain
a wider range and higher concentrations of pollutants, including
toxics, than other mill processes or mines in this category. The
various process reagents used in flotation are discussed later in
this section.
Copper/Lead/Zinc/Gold/Silver/Platinum/Molybdenum Subcategory,
Mill Subdivision, Heap/Vat/Dump/In-Situ Leaching
Very few samples were taken in this mill process because it is
regulated as no discharge of process water in BPT effluent guide-
lines. As can be seen in Table VI-7, the raw wastewater has high
concentration of several parameters, the reason for the no dis-
charge requirement. The one discharged sample reported is
actually treated recycle water which is not discharged.
Copper/Lead/Z i nc/Go1d/S i1ver/P1 a t i num/Mo1ybdenum Subcategory,
Placer Operations Recovering Gold
A study was conducted in 1978 to evaluate current wastewater
handling practices at gold placer mines. Eleven operations, all
located in Alaska, were sampled to determine performance capabil-
ities of existing settling ponds. Only two of the toxic metals
were monitored during the program, arsenic and mercury. Settle-
able solids were also monitored to provide an indication of
treatment pond performance. As can be seen in Table VI-8, the
settleable solids concentrations range from not detected to 500
ml/l/hr. However, many of the different samples are discharges
that had not been treated in settling ponds.
Aluminum Subcategory, Mine Drainage Subdivision
As shown in Table VI-9, aluminum mine drainage is low in most
pollutants. The toxic metals present in the discharge are in
relatively low concentrations (compared to Table VI-1) and are
chromium, copper, mercury, nickel, and zinc. Asbestos was
present in moderate concentrations (compared to Table VI-1) and
was not affected by the existing treatment methods. Acid pH
158
-------
levels were noted in the raw, but these increased to the alkaline
range (7 pH 14) after pH adjustment.
Tungsten Subcategory, Mill Subdivision
As shown in Table VI-10, 13 of the toxic metals were detected in
the raw wastewater. However, these are reduced during treatment
leaving only seven above 20 ug/1 in the discharge. Of these,
copper, lead, and zinc have high concentrations (compared to the
other discharge metals concentrations).
Asbestos and phenolics were detected in the raw samples; cyanide
was not. The values of asbestos are high relative to the
category as a whole (see Table VI-1). The effluent phenolics are
low relative to the values in Table VI-1 .
Mercury Subcategory, Mill Subdivision
As seen in Table VI-11, the toxic metals are found in high
concentrations in the raw wastewater in this subdivision, as are
asbestos and phenolics. That is why the applicable BPT regula-
tion is no discharge of process wastewater. The discharged
sample in Table VI-11 is actually treated recycle water.
Uranium Subcategory, Mine Drainage Subdivision
Uranium mine drainage, is, relative to mill water less polluted.
As seen in Table VI-12, many of the toxic metals were detected,
all but zinc in concentrations less than 65 ug/1. Only six were
detected in the treated samples, none greater than 50 ug/1.
Cyanide was not detected, and phenolics were detected at a low
concentration (10 ug/1). Asbestos was detected in both raw and
treated samples at moderate concentrations (as compared to Table
VI-1 ) .
Not listed in Table VI-12, but shown in the support data
(Supplement A), are radium 226 concentrations. Uranium ore is
radioactive and radium 226 is a radionuclide always associated
with uranium. It is one of the uranium decay series and has a
half life of 1,620 years. Raw mine water may have several
hundred to a thousand pico-Curies per liter (p Ci/1) of Ra 226,
but existing treatment is capable of reducing this to the BPT
guideline of 10 p Ci/1 (total, 30-day average).
Uranium Subcategory, Mill Subdivision
As seen in Table VI-13, several of the toxic metals are found in
both raw and treated wastewater. Treated wastewater in this
table is actually recycle water. The facilities do not
discharge. This recycle water is not treated specifically for
metals, and, therefore, little reduction occurs.
159
-------
Asbestos was found in both influent and effluent samples in
moderate concentrations (as compared to Table VI-1). Cyanide was
not detected and total phenol (4AAP) were detected at a low con-
centration (10 ug/1). As with mine drainage, mill water may have
several hundred to a thousand p Ci/1 Ra 226, Current treatment
at the single uranium mill discharging is reducing this to 10 p
Ci/1, the BPT limitation.
Titanium Subcategory, MijTe Subdivision
As can be seen in Table VI-14, the mine water from this
subcategory is relatively clean (relative to Table VI-1). Three
toxic metals (copper, lead, and zinc) were detected at 20 ug/1.
Relative to the category as a whole (Table ' VI-1), the asbestos
values are low. Total phenolics were detected at 30 ug/1.
Titanium Subcategory, Mill Subdivision
As shown in Supplement A (Support Data; Sample Points 1A and 2A,
for Mill 9905), seven toxic metals were detected in the raw
wastewater; all but selenium and lead at concentrations greater
than 200 ug/1. These concentrations were reduced by treatment,
leaving only five detected toxic metals ranging in concentration
from 20 to 100 ug/1.
Asbestos was detected at moderate concentrations (compared to
Table VI-1). Cyanide was not detected and phenolics were
detected at 10 ug/1 in raw and discharged samples.
Titanium Subcategory, Mills with Dredge Mining Subdivision
Table VI-15 summarizes the data for the titanium mills employing
dredge mining. Ten toxic metals were detected in the raw water,
at concentrations less than or equal to 80 ug/1. In the treated
effluent, six toxic metals were detected. Only zinc was detected
in concentrations greater than 10 ug/1.
COD and TOC concentrations in the raw water were generally
present in higher concentrations than the rest of the category
due to the presence of organic material in some of the ores. The
treatment processes used substantially reduced the concentrations
of both COD and TOC. The TSS concentration of the effluents were
less than 10 mg/1.
Vanadium Subcategory, Mine Drainage Subdivision
Table VI-16 illustrates the character of vanadium mine drainage.
Several toxic metals were present both in the raw and discharged
water. Discharge concentrations greater than 20 ug/1 were
reported for chromium, copper, lead, nickel, and zinc. Cyanide
and total phenolics were not detected. The asbestos values were
low relative to the category as a whole.
160
-------
Vanadium Subcateqory, Mill Subdivision
As seen in Table VI-17, many toxic metals were detected in both
the raw and discharged waters from this subdivision. Of the
metals, only mercury was reduced below the detection limit by the
existing treatment system. Cyanide was also reduced below the
detection limit, and no total phenolics were detected in raw or
discharged water.
Ant^imonv Subcategory^ Mill Subdivision
The data for this subcategory are presented in Table VI-18.
There is no discharge of treated wastewater from the single mill
in this subdivision. Relatively high concentrations of antimony
and arsenic are present in the raw and treated wastewater.
Phenolics were not detected in the raw or treated wastewater.
Asbestos was detected in moderate concentrations compared to
Table VI-1. The pH of the impounded water was greater than 12.0.
REAGENT USE £N FLOTATION MILLS
Froth flotation processes use various reagents in the porcess,
and these reagents are discharged with the tailings and mill
process water. Flotation reagents are a possible source of toxic
organics in an industry which, otherwise, has no known source of
toxic organics. Therefore, a survey was conducted to determine
the availability of toxic organics and other toxics in flotation
reagents.
The results of a nationwide survey of sulfide ore flotation mills
indicate that over 547,400 metric tons (602,000 short tons) of
chemical flotation reagents were consumed in 1975 (Reference 1).
Reagent use data supplied by 22 milling operations indicate that
63 different chemical compounds are used directly in sulfide ore
flotation circuits. These reagents are categorized as:
1. pH Modifier (Conditioner, Regulator)Any substance used
to regulate or modify the pH of an ore pulp or flotation
process stream. Examples of the most commonly used
reagents are lime, soda ash (sodium carbonate), caustic
soda (sodium hydroxide), and sulfuric acid.
2. Promoter (Collector)A reagent added to a pulp stream
to bring about adherence between solid particles and
air bubbles in a flotation cell. Examples of the most
common promoters are xanthate and dithiophosphate salts,
as well as saturated hydrocarbons (such as fuel oil).
3. FrotherA substance used in flotation processing to
stabilizeair bubbles, principally by reducing surface
tension. Common frothers are pine oil, cresylic acid,
amyl alcohol, MIBC, and polyglycol methyl ethers.
161
-------
4. ActivatorA substance which, when added to a mineral
pulp, promotes flotation in the presence of a collecting
agent. It may be used to increase the floatability of a
mineral in a froth or to refloat a depressed mineral. A
good example of an activating agent is copper sulfate,
used in the flotation of sphalerite.
5. DepressantA substance which reacts with the particle
surface to render it less prone to stay in the froth,
thus causing it to wet down as a tailing 'product
(contrary to activator). Examples of depressing agents
most commonly used are cyanide, zinc sulfate, corn
starch, sulfur dioxide, and sodium sulfite.
Table VI-19 summarizes reagent use for copper, lead, zinc,
silver, and molybdenum flotation mills which discharge process
wastewater. Comparing the reagents listed in Table VI-19 to the
list of toxic pollutants given in Section V, only the following
reagents are considered to be potential sources of one or more
toxic pollutants in mill process wastewater: copper, zinc,
chromium, and total phenolics (4AAP).
Copper
Copper sulfate addition to a flotation pulp containing sphalerite
(ZnS) is a good example of an activating agent. The cupric ions
replace zinc in the sphalerite lattice to permit better collector
attachment, thus allowing the mineral to be floated with a
xanthate (Reference 2). Copper ammonium chloride functions in
much the same manner and is used at one operation (Mill 3110)
because it is purchased as a waste byproduct from the
manufacturer of electronic circuit boards. Copper sulfate is
highly soluble in water and is added to the flotation circuit in
concentrations as high as 100 mg/1 (as Cu). Residual dissolved
copper in the tailings pulp stream readily forms copper hydroxide
precipitates at the alkaline pH common to most sulfide flotation
systems.
Zinc
The function of zinc sulfate is the depression of sphalerite when
floating galena and copper sulfides (Reference 3), and the
mechanism involved is very similar to that of copper sulfate
described above. Typically, dosage rates of 0.1 to 0.4 kilogram
of zinc sulfate per metric ton (0.2 to 0.8 pound per short ton)
of ore feed are used, often in conjunction with cyanide. These
dosage rates translate to dissolved zinc loads in the flotation
circuit of 5.2 to 65 mg/1 (as Zn). Residual zinc concentrations
from excessive zinc sulfate use are small compared to the total
zinc content of the tailings.
162
-------
Chromium
Sodium dichromate is used as a flotation reagent at only one of
the 22 flotation mills listed in Table VI-19. It functions as a
depressant for galena in copper/lead separations. Dosages of
this reagent are relatively small, and long term analyses of
treated effluent have not indicated the presence of chromium in
detectable concentrations.
Cyanide
Sodium cyanide and, to a lesser extent, calcium cyanide have
found widespread application within the industry as strong
depressants for iron sulfides and sphalerite. Cyanide also acts
as a mild depressant for chalcopyrite, enargite, bornite, and
most other sulfide minerals with the exception of galena (Refer-
ence 4). A secondary action of cyanide, in some instances, may
be the cleaning of tarnished mineral surfaces, thereby allowing a
more selective separation of the individual minerals (Reference
5). Typical cyanide reagent dosages range from 0.003 to 0.125
kilogram per metric ton (0.006 to 0.250 pound per short ton) of
ore feed and average 0.029 (0.058). Expressed in terms of water
use, cyanide dosages range from less than 1.0 to 50.4 milligrams
per liter (as sodium cyanide), with an average of about 11.
Sodium cyanide and calcium cyanide flotation reagents are the
sole source of cyanide in flotation mill effluents. Four flota-
tion mills (2122, 3121, 6101, and 6102) have effluent discharge
concentrations of 0.1 mg/1 total cyanide or greater. Mill 6102
is the largest consumer of cyanide in terms of dosage per unit of
ore feed and per unit of flotation circuit water feed. As a
result, Mill 6102 produces a raw discharge with total cyanide
concentrations of 0.2 to 0.4 mg/1. Cyanide dosages used at Mills
2122, 3121, and 6101 are consistent with amounts used throughout
the industry, and, for this reason, reagent use alone does not
appear to be the cause for high cyanide levels. The treatment of
cyanide-bearing wastewater and the chemistry of cyanide in mill
wastewater are discussed in Section VIII of this report.
Phenolic Compounds
"Reco" (sodium dicresyldithiophosphate) is used at Mill 2122 to
promote the flotation of copper sulfide minerals. Reco is
similar to American Cyanamid's AEROFLOAT 31 and 242 promoters,
which are used at Mills 3101, 3104, 3115, 4403, and 9202. These
reagents contain the cresyl group (CH3_.C6H3_.OH), a very close
relative of the toxic substance 2,4-dimethylphenol, which has
been detected in raw mill wastewater samples collected during the
toxic substance screen sampling program at Mills 2122 and 9202.
Mills 3101, 3104, 3115 and 4403 were not selected as sites for
screen and/or verification sampling of organic toxic pollutants
during this program.
163
-------
Cresylic acid is used as a flotation reagent at Mills 2117, 2121,
and 4403. Xylenols, C2H 5_, C6H4_OH or (CH3.)2.C6H3_.OH, are the
dominant constituents of commercial cresylic acids and include
the toxic pollutant, 2,4-dimethyphenol, which has not been
detected in raw or treated wastewater samples at Mills 2117 and
2121. Mill 4403 was not sampled for the organic toxic
substances. Nitrobenzenes are present in Aero 633, but
nitrobenzene was not detected in wastewater during this program.
However, screening and verification sample data strongly
implicate these phenol-based flotation reagents as the sources of
total phenol (4AAP) in mill process wastewaters. From a
practical standpoint, cresylic acid can be considered as 100
percent phenolic with the relative phenolic content of the other
phenol-containing reagents being considerably less. Phenolic
concentrations of 5.2 mg/1 and 5.0 mg/1 have been detected in the
mill tailing samples at Mill 2117, and treated effluent samples
were found to contain 0.30 mg/1 and 0.36 mg/1 on 2 consecutive
days. The large consumption of cresylic acid at Mill 2117 (0.035
kilogram/metric ton equivalent to 0.070 pound per short ton, of
ore) and the consistency of data substantiate cresylic acid as
being a significant source of phenolic compounds in flotation
mill process effluents.
Phenolic compounds were found to be the most prevalent toxic
organic species detected in the screen samples, but concentra-
tions did not exceed 0.03 mg/1 except at operations which are
known to employ one or more of the phenol based flotation
reagents previously discussed.
SPECIAL PROBLEM AREAS
Backfilling of Mines With Mill Sand Tailings
A review of sample data and historical monitoring data supplied
by the industry indicates the presence of significant
concentrations of cyanide in several mine water discharges.
Further examination revealed that the facilities with cyanide in
mine water backfilled mined-out stopes using mill sand tailings
from flotation circuits which use cyanide compounds as process
reagents.
A variety of undergound mining techniques are used throughout the
mining industry. Typical mining methods include room-and-pillar,
vein (or drift) mining, open stoping, pillar stoping, cut-and-
fill, and panel-and-fill. The selection of method(s) is
dependent on many factors, such as the type and shape of the ore
deposit, the depth of excavations, and the ground conditions.
Cut-and-fill, pillar stoping, and panel-and-fill techniques have
found common application in lead, zinc, and silver mines located
in Colorado, Utah, and the Coeur d'Alene Mining District in
Idaho. An inherent feature of these mining methods "is the
refilling of worked-out and abandoned stopes and other workings
164
-------
to prevent subsidence and cave-ins as mining progresses through
the ore body. For many years, waste rock from the mine
exploration crosscuts was used as fill material; however, the
development of hydraulic sandfill procedures has simplified the
backfill operation, In current practice, the coarse (sand)
fraction of the flotation-mill tailings is often segregated from
the tailings pulp stream by hydro-cyclones and pumped into the
mine for backfilling.
Nine mines (Mines 3107, 3113, 3120, 3121, 3130, 4104, 4105, 4401
and 4402) are known to practice hydraulic backfilling with mill
sand tailings. Eight of these nine mills use cyanide either as a
flotation reagent (Mills 3107, 3113, 3121, 3130 and 4401) or as a
leaching agent (Mills 4104, 4105, and 4402). The nature of the
mechanism by which cyanide depresses pyrite and sphalerite is
such that much of the cyanide added to the flotation circuit
associates with the depressed minerals in the tailings and ulti-
mately is leached into mine water during hydraulic backfill.
Mine 3130 is the only facility with a separate mine drainage
treatment system that periodically monitors for cyanide. Efflu-
ent monitoring data (summarized in Section VIII) include cyanide
analyses of five 24-hour composite samples collected during the
period of June 1977 through October 1977. The data indicate that
cyanide concentrations in the treated mine water did not exceed
0.2 mg/1 total cyanide for mills and mine/mills on a daily basis,
although the monthly average exceeded 0.1 rng/1 on one occasion.
Examination of raw (untreated) mine-water data from Mine 3130
indicates that cyanide is not effectively removed by the treat-
ment system, which consists of lime and flocculant addition,
followed by a series of two sedimentation ponds. This treatment
is not designed for destruction or removal of cyanide and, does
not provide sufficient residence time for natural aeration.
fore, the poor removals observed are'not surprising.
Total cyanide concentrations detected in five mine-water grab
samples collected to support BAT at Mine 3130 were found to range
from 0.04 to 0.16 mg/1. A 24-hour composite mine-water sample
collected at Mine 3107 was found to contain 0.4 mg/1 during
backfill operations.
Mine 4105, located in South Dakota, was visited during the
screening phase of this program. Analysis of mine water for
total cyanide indicated that, for the days when the contractor
sampled, concentrations were less than detectable. During pre-
vious visits to this facility, no cyanide was detected in mine
water samples.
165
-------
Table VI -1
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
NUMBER OF
SAMPLES
ACENAPHTHENE
ACROLEIN
ACRYLONITRILE
BENZENE
BENZIDENE
CARBON TETRACHLORIDE
CIILOROBENZENE
1.2. 3 TRICIU OROBENZENE
HE XACI ILOROB ENZ ENE
. 2-OICHLOROETHANE
. 1 . 1-TRICHLOROETHANE
HEXACHLOROETHANE
. 1-DICHLOROETHANE
. 1.2-TRICHLOROETHANE
.1.2 . 2-TETRACHLOROETHAN
CHLOROETHANE
BIS(CHLOROMETHYL) ETHER
B1S(2 aiLOROETHYL) ETHER
2-CHLOROETHYL VINYL ETHE
2 -CHLORONAPHTHALENE
2 4 .6-TRICHLOROPHENOL
PARACHLOROMETA CRESOL
CHLOROFORM
2 CHLOROPHENOL
1 . 2-DICHLOROBENZENE
1 , 3-DICHLOROBENZENE
1.4 DICIII OROBENZENE
3,3-OICHLOROBENZIDINE
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
32
32
32
32
32
32
32
32
32
32
RAW(UQ/L) *
NUMBER DETECTED VALUES ONLY *
DETECTED MEAN MED 90% MAX *
0
O
0
10
0
1
0
0
0
0
9
0
0
0
0
0
0
O
O
0
1
0
9
0
0
0
O
O
*
*
*
4.8922 4 10 1O *
*
1 1 1 1 *
*
*
*
*
8.7208 6.5811 10 10 *
*
*
*
*
*
*
*
*
*
11.667 11.667 11.667 11.687 *
*
7.6098 3.1623 12.6 35 *
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
29
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED (UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
0
O
0
3
0
0
1
0
O
0
s
0
0
0
O
0
0
0
0
0
O
0
8
0
O
0
0
0
8.3333 7 10.7 11
O.005 O.O03 O.OOS O.OO5
7.2649 8.6811 1O 1O
5 1281 3.1823 1O 1C
-------
Table VI -1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
en
-4
NUMBER OF
SAMPLES
1. 1-DICHLOROETHYLENE
1 , 2 -TRANS-DICHLOROETHYLE
2 . 4 - OI CHLOROPHENOL
1 . 2 -OICHLOROPROPANE
1.3-DICHLOROPROPENE
2 . 4-DIMETHYLPHENOL
2.4-OINITROTOLUENE
2,e-DtNITROTOLUENE
1 . 2-D1PHENYLHYDRAZINE
ETHYL-BENZENE
FLUOR ANTHENE
METHYL CHLORIDE
METHYL BROMIDE
BROMOFORM
DICHLOROBROMOMETHANE
TR I CHLOROFLUOROMETHANE
0 1 CHLOROO I F LUOROMETHANE
CHLORODI BROMOME THANE
HE XACHLOROBUT AD I ENE
HE XACHLOROC YCLOPENT AD I EN
ISOPHORONE
NAPHTHALENE
NITROBENZENE
2-NITROPHENOL
4-N1TROPHENOL
2.4-OINITROPHENOL
4 . 6 DINITRO O CRESOL
32
32
32
32
32
32
32
32
32
32
32
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
RAW(UG/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
2 6.5811 3.1823 8 6326 1O
0
1 10 1O 1O 1O
O
0
1 140 14O 140 140
0
0
0
4 8.7187 1 13.48 17.687
0
1 45 45 45 45
O
0
0
S 5.0325 2.O811 10 10
0
0
0
0
0
1 12.5 12.5 12.5 12.5
0
6
0
0
0
*
*
*
*
*
t
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED (UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 80% MAX
0
1 270 270 270
0
0
O
1 27O 27O 27O
0
0
0
3 6.8 4.8 8.64
O
O
0
0
2 6.6811 3.1623 8.8325
3 4.7208 2.0811 7.8487
0
0
0
0
0
O
0
0
0
0
0
270
270
1O
1O
10
-------
Table VI -1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
CO
RAW(UQ/L)
NUMBER OF
SAMPLES
N-NITROSODIMETHYLAMINC
N-N1TROSODIPHENYLAMINE
N-NITROSODI-N-PROPYLAMIN
PENTACHLOROPHENOL
PHENOL
BIS(2-ETHYLHEXYL) PHTHAL
BUTYL BENZYL PHTHALATE
DI-N-BUTYL PHTHALATE
OI-N-OCTYL PHTHALATE
D I ETHYL PHTHALATE
DIMETHYL PHTHALATE
BENZO( A) ANTHRACENE
BENZO(A)PYRENE
BENZO(B)FLUORANTHENE
BENZO(K)FLUORANTHENE
CHRYSENE
ACENAPHTHYLENE
ANTHRACENE
BENZO(G,H £ )PERVL£NE
FLUORENE
PHENANTHRENE
DI8tNZO( A, H) ANTHRACENE
INOENO( 1,2,3-C,D)PYRENE
PYRENE
TETRACHLOROETHYLENE
TOLUENE
TRICHLOROETHYLENE
33
33
33
33
33
33
33
33
10
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
NUMBER DETECTED VALUES ONLY
DETECTED MEAN * MED 90% MAX
0
0
0
t
a
15
2
13
3
18
0
0
0
0
0
0
0
0
0
1
0
0
0
0
2
9
0
to
its
20.16
10.75
18.489
10
24.414
to
7.75
399.28
10
78
13
0.6
10
10
10
10
4.6
2 . 08 1 1
to
143.2
39.833
18. S
28.1
to
69.4
10
9.7
388.3
10
160
100
21
56
10
90
»
*
» NUMBER OF
* SAMPLES
28
28
28
28
28
28
28
28
7
28
28
28
28
* 28
* 28
t 26
* 28
* 28
* 28
10 23
28
28
28
28
11 28
358O 28
28
TREATED
(UQ/L)
NUMBER DETECTED
DETECTED MEAN MED
0
O
0
0
3
IB
4
12
3
4
3
0
0
O
0
0
0
0
0
1
0
0
0
0
1
a
0
92.3
12.458
27-791
25.884
12. 187
7.87S
12.2
10
1. 1
2 . 6987
33. 4B
10
10
10
10
9.8
s. a
1O
1.1
1
VALUES ONLY
90% MAX
188.8
28
52.4
39.2
14.65
10
2O.33
10
1.1
5.28
2 tO
60
68
140
18. 6
to
23
1O
1.1
to
-------
Table VI -1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
NUMBER OF
SAMPLES
VIHVL CHLORIDE
ALDRIN
Oli-LDSIN
CHLORDANE
4,4-ODT
4.4-DD£
4. 4 -ODD
ENDOSULF AN -ALPHA
EWDOSULFAN B£TA
ENDOSULFAN SULFATE
ENDRIN
ENORIN ALDEHYDE
HEPTACHLOR
HEPTACHLOR E POX IDE
BHC -ALPHA
BHC BETA
BHC (LINDANE) -GAMMA
BUG-DELTA
PCB-1242 (AROCHLOR
PCB -1254 (AROCHLOR
PCB-1221 (AROCHLOR
PCB -1232 (AROCHLOR
PCB -1 248 (AROCHLOR
PCB- 1260 (AROCHLOR
PCB-1016 (AROCHLOR
TOXAPHENE
1242)
1254)
1221)
1232)
1248)
126O)
1016)
2.3,7 , 8-TETRACHLORODIBEN
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
a
8
a
8
8
32
33
RAM(UO/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 80% MAX
O
4 6.4156 6 9 1O
O
O
0
t 5555
1 6.6667 6.6667 6.66S7 6.6687
1 10 10 10 tO
0
0
0
0
1 7.5 7.5 7.5 7.S
0
B 5.2648 4.0811 7.5 1O
5 6.1325 5 8.75 10
4 6.2O72 5 8.6667 10
2 BBSS
0
0
0
0
0
0
0
0
0
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
»
*
*
*
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
6
6
6
6
6
27
28
TREATED (UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN M£0 80% WAX
0
2 6.5811 3.1823 8.8325
2 6.5811 3.1623 8.6325
0
C
0
0
0
O
O
t 555
0
2 3.5811 3,1623 8.8325
0
3 555
1 686
O
2 665
O
0
0
0
0
0
0
0
O
1O
to
5
to
B
g
B
-------
Table VI -1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
NUMBER OF NUMBER
SAMPLES DETECTED
CIS t-3-DICHLOROPROPYLEN
TRAN 1 . 3-DICHLOROPROPYLE
RAW(UQ/L)
DETECTED VALUES ONLY
MEAN MED 80% MAX
*
*
»
*
»
*
TREATED (UQ/L)
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MED SO* MAX
-------
Table VI-1 (Continued)
['ATA M1MMAKY
OfU IIIM(;r- !Mt
ML M.i"( * Ti. CUKI FS
if si. { t:<~. /L ) >
AMI rrr '. v
AK..I.MC (
ill K Yl 1 I III
C t P r I i i I ' (
CHF. Ul; i UP
c n p I F; K 1 1
CY AiilF L (
L F e ! > n o i
M F F- r. U 1 Y (
n r K 1 1 1 1
=5!. Li.sn u1'
S 1 L VF P (1
TIIALL HIM
7 I N C 1 T I i T
COD
IS.i
TOC
PH (tr. i ':
PIU NOI. ICS
IROil < TO 1
( T i. 1 A 1 )
: i; r ;. i >
( T n i f ( >
i m n >
(KM A L )
:: i A L )
T". 1 .1 1 )
M. )
1 0 [ t. L >
'J 1 A L )
( 1 !'l AL )
) T ii L )
' TOTM. t
M. )
J
( 'i / t P )
M. !
NHM,Li> OF Ml! nl >'
SAfiF-U'S PI II Ml I
P? f,
111 1 T- 6
o 4 1 .'
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R ';, ( II
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bf. 71)
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17 17
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t-1 > n 'j
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. i .'. 7 ? 'r.
i. . :.! 1 '
.<.«!<-;
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. 1 1 .i 7 2
1 2 .i . 7
IF. M'.L VAt.u;
MU. 1 «': ''.I
n'ci;'"
0.1 n . ;;
n . ,< ; 2 -- 0 .
n.M
^ . » ? 2
o . r i n
1.35 1
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: . 5
r . i '.. t. .
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1.17 1
0 . C 7 T.. ? n
1C', 'b 1
If
I f- (.'
7 n
''.11 C . 3
f, . 2 1 f. 1
:; oui Y
S M4X
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'i . b * 6 1
. ^} 1.21
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n ) * o . P <.
n f t 1 't . 2
f 1 S 1.5
.77 1.1
.21 1.21
2.1 300
7f, i i ^n (j
P-C QJ^-TP
7 r 0 750
.11 1 . S
1 ,' 2 0 . 7 K
"'": ?oi(i
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' 75
90
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37
13
37
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.5
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'b
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<|Q
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TRL6TFD O:'-/l>
HFAN
0.031
n . o o t )
.01115
.13623
.23161
. 13571
.131R1
.01261
. 2220J
.05^57
.015(7
.527f 7
.90209
1 1 .69B
21 .989
5 . 1 P 7 5
t .9711
.07378
. 6 2 6 1 "
:t TFCTFU
M C U 1 A r1
1 Pfirj-5
0.01P
0.005
0.005
0.035
0.06
O.OR
0.05
r .5
5.23
7.7
0.032
0.20°
VAIUC? ONI
0.1
0.6
0 . 0 1 U 9
0.06
0.532
n .51 P
O.Ub
0 . J7f
0.0306
0.966
0.112
O.C1
0 .P1
2.211
20.6
69.2
6
P. 16
0.21
2.192
Y
MAX
0.1
0.011
0. 077
l.P
1 .6
fl.t
0.9"9
0.25
1 .2fl
0. 1
0.01
0 .fll
11.1
53
157
6
P. 5
0.16
3.87
-------
ro
Table VI-2
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY IRON
SUBDIVISION MINE
MILL PROCESS MINE DRAINAGE
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL!
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL;
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
2
1
2
2
1
2
1
2
2
2
2
2
2
2
1
1
2
1
2
1
1
1
1
0
0
0
0
0
1
0
0
0
0
0
0
0
1
1
1
2
1
2
0
0
1
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.03
0.018
10
2
4.65
25
8.075
3500E3
1700E4
0.09
0.018
10
2
4.65
25
8.075
3500E3
1700E4
0.09
0.018
10
2
5
25
8. 15
3500E3
1700E4
0.09
0.018
10
2
5
25
8. 15
3500E3
1700E4
*
*
TREATED (MG/L)
* NUMBER OF NUMBER
* SAMPLES DETECTED
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
* 1
*
* 1
* 1
0
1
0
0
0
1
0
0
0
0
0
0
0
0
1
1
DETECTED
MEAN MEDIAN
0.005
0. 12
0.03
6
3
4
19
8
3800E3
4200E4
0.005
0. 12
0.03
6
3
4
19
8
3800E3
4200E4
VALUES ONLY
90% MAX
0.005
0. 12
0.03
6
3
4
19
8
3800E3
4200E4
0.005
0. 12
O.03
S
3
4
19
8
3800E3
4200E4
-------
Table VI-3
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY IRON
SUBDIVISION MILL
MILL PROCESS PHYSICAL AND/OR CHEMICAL
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
2
2
2
2
2
2
1
2
2
2
2
2
2
2
1
1
2
1
2
1
1
1
1
0
1
1
1
2
2
0
2
0
2
1
2
0
2
1
1
2
1
2
0
1
1
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.89
0.92
0.031
0.276
0.225
0 . 0505
2. 15
0.02
0.017
3.15
96
80
64500
22
7.775
73
3800E7
2300E8
0.89
0.92
O.O31
0.276
0.225
0 . 0505
2. 15
0.02
0.017
3. 15
96
80
64500
22
7.775
73
3800E7
2300E8
0.89
0.92
0.031
0.5
0.32
0.08
2.7
0.02
0.02
5.8
96
80
110000
22
7.9
73
3800E7
2300E8
0.89
0.92
0.031
0.5
0.32
0.08
2.7
0.02
0.02
5.8
96
80
110000
22
7.9
73
3800E7
2300E8
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF NUMBER
SAMPLES DETECTED
2
2
2
2
2
2
1
2
2
2
2
2
2
2
1
1
2
1
2
1
1
1
1
0
1
0
0
2
1
0
0
0
0
0
0
0
2
1
1
2
1
2
0
0
1
1
TREATED
(MG/L)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.005
0.014
0. 1
0.019
4
.03162
2.0158
11
7.675
4100E3
4300E4
0.005
0.014
0. 1
0.019
4
.03162
2.0158
11
7.675
4100E3
4300E4
0.005
0.018
O. 1
0.03
4
.03162
4
11
8. 1
4100E3
4300E4
0.005
0.018
0. 1
0.03
4
.03162
4
11
8. 1
41OOE3
43OOE4
-------
Table VI-4
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/MOLYBDENUM
SUBDIVISION MINE
MILL PROCESS MINE DRAINAGE
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
15
15
15
15
15
15
11
15
15
15
15
15
15
15
12
6
13
10
12
13
7
6
6
3
11
4
9
7
14
2
10
4
11
1
5
4
15
12
5
13
10
12
9
7
6
5
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
. 11633
.03727
. 00908
.02819
0.0216
.76418
0.0115
1 . 0909
O.O027
.07084
0.012
.01106
0. 1045
5.2516
24.289
16.454
195.2
7 . 6507
7. 1125
.00822
20.05
917E13
2424E7
0. 121
0.018
.00715
0.005
0.017
0.045
0.0115
0.287
0.002
0.059
O.O12
0.012
0.0715
0.31
7.9
3.2
20
3.75
7. 1
0.008
1 .44
3063E6
1000E5
O. 132
0. 1708
0.022
0. 124
0.065
4.285
0.02
5.496
0 . OO63
0. 184
O.O12
0.02
0.269
23.98
118. 15
70
1094.6
22.3
8. 16
0.016
133.4
550E14
1200E8
0. 132
0. 196
0.022
0. 124
0.065
7.3
0.02
5.87
0.0063
0.2
0.012
0.02
0.269
28. 15
154
70
1456
23
8.25
0.016
133.4
550E14
1200E8
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
TREATED (MG/L)
NUMBER OF NUMBER DETECTED
SAMPLES DETECTED MEAN MEDIAN
5
5
5
5
5
5
4
5
5
5
5
5
5
5
4
2
5
2
4
4
3
2
2
0
2
1
2
0
5
1
4
1
2
0
1
2
5
4
2
5
2
4
2
3
2
2
0.01
0 . 0064
.01055
0.056
0.035
.06275
0.049
0 . 32O5
0.03
0.309
3.762
27.25
2.5
11
3.5
8.2875
.00755
9.4117
4650E3
6450E4
0.01
0 . 0064
.01055
0.04
0.035
0.066
0.049
0.3205
0.03
0.309
O.53
14
2.5
10
3.5
8.225
. 00755
0.65
4650E3
6450E4
VALUES ONLY
90% MAX
0.01
0 . O064
0.013
0. 12
0.035
0.099
O.O49
0.601
0.03
0.476
13.99
77
3
20
6
9
0.0101
27.36
8200E3
7200E4
0.01
O . 0064
0.013
0. 12
0.035
O.099
0.049
0.601
0.03
O.476
13.99
77
3
20
6
9
0.0101
27.36
8200E3
7200E4
-------
Table VI-5
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/MOLYBDENUM
SUBDIVISION MILL
MILL PROCESS CYANIDATION
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOL ICS (4AAP)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
2
1
2
1
0
2
2
2
2
2
0
2
1
0
2
2
2
2
2
2
0
2
2
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0. 1
100
0.03
0.825
1 .44
3.85
0. 195
0.2775
0.0775
0. 1
2. 16
354
649
30149
11.5
8.825
1372E6
5527E7
0. 1
10O
0.03
0.825
1 .44
3.85
0. 195
0.2775
0.0775
0. 1
2. 16
354
649
30149
11.5
8.825
1372E6
5527E7
0.1
200
0.03
1.6
2.6
6.8
0.37
0.54
0. 15
0. 1
3.9
700
1290
60200
18
9
2700E6
1100E8
0.1
200
0.03
1 .6
2.6
6.8
0.37
0.54
0. 15
0. 1
3.9
700
1290
60200
18
9
2700E6
1100E8
* TREATED (MG/L)
t
* NUMBER OF NUMBER DETECTED VALUES ONLY
* SAMPLES DETECTED MEAN MEDIAN 9O% MAX
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
-------
Table VI-6
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/MOLYBDENUM
SUBDIVISION MILL
MILL PROCESS FLOTATION (FROTH)
RAW(UG/L)
NUMBER OF
SAMPLES
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
TOTAL FIBERS
ASBESTOS (CHRYSOTILE)
COD (MG/L)
VSS (MG/L)
TSS (MG/L)
TOG (MG/L)
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL-MG/L)
78
78
78
78
78
78
74
78
72
78
77
78
78
78
13
15
22
9
22
22
22
73
9
NUMBER
DETECTED 107=
13
78
55
61
63
78
31
69
48
72
50
43
7
78
13
15
22
9
22
22
22
65
9
44 . 7
28
0.35
0.81
20
307.2
40
100.7
0.2
72.8
12
11.84
1 .7
98
7.4E+09
1 .6E+09
22.2
345
38.6
3.4667
6.52
11 .5
1 .9
MEAn
1725
2853.4
75.401
637.99
4643.1
98854
282.91
20192
51 .63
3708.6
242.38
410.99
89.557
74137
1 .8E+12
2.3E+11
1126.8
10244
199538
14.864
8.7848
214.54
57.733
NED
167
800
75
170
1850
63300
180
2750
1 .1
2000
200
25! .67
8.1
5600
5.4E+11
4.8E+10
530
3750
164000
9.5
8.35
35.5
28.5
90%
1359
8100
150
1178
11000
292000
590
27300
22.2
9200
526.67
805
197.8
266400
3.5E+12
4.2E+1 1
2988
14794
444199
28.6
11 .61
402.5
159.53
*
*
*
*
*
*
*
*
k
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
TREATED (UG/L)
NUMBER OF
SAMPLES
59
59
59
59
59
59
51
59
58
59
50
59
59
59
14
14
15
8
21
15
21
52
6
NUMBER
DETECTED 1 07,
3
43
7
6
20
55
12
27
16
35
23
8
0
48
14
14
14
7
20
15
21
48
6
100
6
1
5
10
20
44
20
0.5
25
5
1 1
30
3.7E+06
1 .5E+05
4.4
.03162
1
3
6.21
9
0.095
MEAN
133.33
75.244
5.7143
7-3333
162.05
313.64
165
100.19
27.3
90.543
23.225
31 .375
258.12
6.1E+08
2.2E+08
15.655
1 .721 1
11.483
11 .667
7.8226
74.326
.56867
MED
100
12.5
3
5
30
70
!20
42.5
0.8
60
12.083
20
70
1 .9E+07
1 . 7E+06
12
1 .5
4.6
9.5
7.825
19
0.15
907.
170
285
12.1
11 .6
320
640
256
233
68
185
34
46
562
1 .9E+09
3.2E+08
26.9
3.15
17.333
20.5
8.8
216
1 .288
-------
Table VI-7
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/MOLYBDENUM
SUBDIVISION MINE/MILL
MILL PROCESS HEAP/VAT/DUMP LEACHING
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL) 1
ARSENIC (TOTAL) 1
BERYLLIUM (TOTAL) 1
CADMIUM (TOTAL) 1
CHROMIUM (TOTAL) 1
COPPER (TOTAL) 1
LEAD (TOTAL) 1
MERCURY (TOTAL) 1
NICKEL (TOTAL) 1
SELENIUM (TOTAL) 1
SILVER (TOTAL) 1
THALLIUM (TOTAL) 1
ZINC (TOTAL) 1
TSS 1
PH (UNITS) 1
IRON (TOTAL) 1
0
1
0
1
0
0
1
1
1
1
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.027
0.3
0.26
1.3
88
0.01
14.2
0.003
107
323
3.04
1860
0.027
0.3
0 26
1 .3
88
0.01
14.2
0.003
107
323
3.04
I860
0.027
0.3
0.26
1.3
88
0.01
14.2
0.003
107
323
3.04
1860
0.027
0.3
0.26
1 .3
88
0.01
14.2
0.003
107
323
3.04
1860
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
t
*
*
*
*
*
*
*
*
*
*
0
1
1
1
0
1
1
0
1
0
0
* 1 1
* 1 1
* 1 1
* 1 1
* 1 0
TREATED
(MG/L)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.002
0.002
0.003
0.023
100E-5
0.028
0.003
0.013
50
7.87
0.002
0.002
0.003
O.O23
100E-5
0.028
0.003
0.013
50
7.87
0.002
0.002
0.003
O.O23
100E-5
0.028
O.003
0.013
50
7.87
0.002
0.002
0.003
0.023
1QOE-5
0.028
0.003
0.013
50
7.87
-------
Table VI-8
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/MOLYBDENUM
SUBDIVISION MINE/MILL
MILL PROCESS GRAVITY SEPARATION
RAW(MGXL)
TREATED (MG/L)
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MEDIAN 90% MAX
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MEDIAN 90% MAX
ARSENIC (TOTAL)
MERCURY (TOTAL)
TSS
PH (UNITS)
11
11
11
6
11
11
11
6
1 . 1736
436E-6
18.598
7.2
0.2
100E-6
12.4
7. 15
4.78
.00134
59.34
7.9
5
0.0014
64. 1
7.9
*
*
*
*
10
10
10
6
10
10
10
6
0. 1729
150E-6
1.4462
7.2
0.05
100E-6
0.757
7.25
1 . 105
470E-6
5.45
7.9
1.2
50OE-6
5.7
7.9
I
oo
-------
Table VI-9
DATA SUMMARY
ORE MINING DATA
SUBCATEGORV ALUMINUM
SUBDIVISION MINE
MILL PROCESS MINf: DRAINAGE
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL*
ARSENIC (TOTAL)
BERYLLIUM (TOTAL!
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY 'TOTAL)
NICKrL (TOTAL)
SELENIUM (TOTAL!
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
vss
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
2
2
1
0
0
0
0
1
1
0
0
1
1
0
0
0
1
0
1
1
1
1
1
2
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.03
0.06
0.037
0.06
0.57
1 .6
2.8
2
3.05
0.005
5500E3
3500E4
0.03
0.06
0.037
0.06
O.57
1 .6
2.8
2
3.05
0.005
5500E3
3500E4
O.03
0.06
0.037
O.06
0.57
1 .6
2.8
2
3.05
0.005
5500E3
3500E4
0.03
0.06
0.037
O.06
0.57
1 .6
2.8
2
3.05
0.005
5500E3
350OE4
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF NUMBER
SAMPLES DETECTED
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
1
3
2
2
0
0
0
0
1
1
0
0
1
0
0
0
0
0
1
1
1
1
1
2
2
2
TREATED
(MG/1.)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.025
0.05
0.084
2
5
6
4
8.6
0.0245
1016E5
7500E5
0.025
0.05
O.OH4
2
5
"5
4
8.S
0.0245
1016E5
7500E5
0.025
0.05
0.084
2
5
G
4
8.6
0.044
2000E5
1400E6
0.025
0.05
0.084
2
5
6
4
8.6
0.044
2000E5
1400E6
-------
Table VI-10
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY TUNGSTEN
SUBDIVISION MILL
00
o
RAW(UG/L)
NUMBER OF
SAMPLES
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
TOTAL FIBERS
ASBESTOS (CHRYSOTILE)
COD (MG/L)
VSS (MG/L)
TSS (MG/L)
TOC (MG/L)
PK (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL-MG/L.)
2
2
2
2
2
2
\
2
2
2
2
2
2
2
1
1
1
1
2
1
2
2
1
NUMBER
DETECTED 107,
1
1
2
I
2
2
0
2
1
2
2
2
0
2
1
1
1
1
2
1
2
1
1
53
370
90
160
680
19000
1300
2
890
11
210
10000
1 .3E+12
3.7E-M1
300
4400
29000
220
9.9
63
660
MEAN
53
370
420
260
940
22000
3050
2
1345
30.5
295
17000
1 .3E+12
3.7E+1 1
300
4400
257000
220
9.9
63
660
MED
53
370
90
160
680
19000
1300
2
890
11
210
10000
1 .3E+12
3.7E+11
300
4400
29000
220
9.9
63
660
*
_ _ *
TREATED
* NUMBER OF NUMBER
90% * SAMPLES DETECTED 107,
53 *
370 *
618 *
320 *
1096 *
23800 *
*
4100 *
2 *
1618 *
42.2 *
346 *
*
21200 *
1.3E-M2 *
3.7E+11 *
300 *
4400 *
393800 *
220 *
9.9 *
63 *
660 *
1 3
1 15
1 0.7
1 36
1 15
1 14000
1 220
0
0
0
1 35
0
1 1 100
1 160
1 9.2
1 15
(CJG/L)
MEAN
3
15
0.7
36
15
14000
220
35
1100
160
9.2
15
MED
3
1-5
0.7
36
15
1 4000
220
35
1 100
160
9.2
15
907.
3
15
0.7
36
15
1 4000
220
35
1 100
160
9.2
15
-------
oo
Table VI-11
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY MERCURY
SUBDIVISION MILL
MILL PROCESS FLOTATION (FROTH)
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
1
1
1
1
1
1
0
1
1
1
0
1
THALLIUM (TOTAL) 1 1
ZINC (TOTAL) 1 1
COD 1 1
VSS 1 1
TSS 1 1
TOC 1 1
PH (UNITS) 1 1
PHENOLICS (4AAP) 1 1
ASBESTOS (CHRYSO) (F/L) 1 1
TOTAL FIBERS (F/L) 1 1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
53
1. 1
O.O9
0.56
0.46
0.85
1
230
1.6
0.01
0.2
2.4
60
4300
139000
21
8
0.92
1500E8
1300E9
53
1 . 1
0.09
0.56
0.46
0.85
1
230
1 .6
0.01
0.2
2.4
60
4300
139000
21
8
0.92
1500E8
1300E9
53
1 . 1
0.09
0.56
0.46
0.85
.1
230
1.6
0.01
0.2
2.4
60
4300
139000
21
8
0.92
1500E8
1300E9
53
1. 1
0.09
0.56
0.46
0.85
1
230
1 .8
0.01
0.2
2.4
60
4300
139000
21
8
0.92
150OE8
1300E9
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
.
*
*
*
*
*
*
*
*
*
*
*
<'
0
0
1
0
0
0
* 1 0
* 1 1
* 1 1
* 1 0
* 1 1
* 1 1
* 1 1
* 1 1
* 1 1
* 1 1
TREATED (MG/L)
DETECTED
MEAN MEDIAN
0.2
0. 11
0.006
0.015
0.05
0.05
0.04
22
IS
13
8.3
0.22
57OOE4
7700E5
0.2
0. 11
0.006
O.015
0.05
0.05
0.04
22
16
13
8.3
0.22
5700E4
7700E5
VALUES ONLY
90% MAX
0.2
0. 11
0.006
0.015
0.05
O.05
O.04
22
16
13
8.3
0.22
5700E4
7700E5
0.2
0. 11
0.006
O.O15
0 05
O.05
0.04
22
16
13
8.3
0.22
5700E4
7700E5
-------
co
Table VI-12
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY URANIUM
SUBDIVISION MINE
MILL PROCESS MINE DRAINAGE
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOL I CS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
3
17
3
16
4
14
3
4
3
4
5
3
3
17
15
2
18
2
13
3
1
3
2
1
16
0
13
3
14
0
3
1
1
3
0
0
17
15
2
18
2
13
1
1
3
2
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.05
0.0195
. 0038 1
.04333
.01673
0.09
O . 0038
0.06
.02333
. 04306
22.504
23.5
144.58
8.5
7.6519
0.01
0.319
1050E5
1950E6
0.05
0.007
0.003
0.045
0.0075
0.05
0.0038
0.06
0.028
0.02
7
23.5
21
8.5
8.05
0.01
0.319
1100E5
1950E6
0.05
0.0832
0.0092
0.05
0.075
0.18
O.OO38
0.06
0.037
0. 158
104.2
28
415.94
9
8.655
0.01
0.319
1900E5
2300E6
0.05
O. 17
0.01
O.O5
0.11
0. 18
O.O038
0.06
0.037
0. 19
140.5
28
1639.5
9
8.825
0.01
0.319
1900E5
23OOE6
*
*
*
*
,
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
TREATED (MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
3
13
3
13
3
11
3
3
3
3
5
3
3
13
12
2
1>3
2
9
3
1
2
2
0
11
0
10
2
8
0
1
1
0
3
0
0
12
12
2
13
1
9
1
1
2
2
DETECTED
MEAN MEDIAN
.O0798
0.0038
0.0425
.00575
0.05
O . 009 1
.03633
.01983
10. 169
1 .5
33. 185
10
7.8833
0.01
0.054
4000E4
5000E5
0.006
0.003
O.O425
O.OO6
0.05
0 . 009 1
0.048
0.014
8.95
1 .5
27
10
7.9
0.01
0.054
4000E4
5000E5
VALUES ONLY
90% MAX
0.0228
0.0069
0.06
0.011
0.05
O . OO9 1
0.051
0 . 0666
33.5
2
75.8
10
8.5
0.01
0.054
5300E4
570OE5
0.024
0.007
O.O6
O.011
0.05
O . O09 1
0.051
0 . 078
38
2
83
10
8.5
0.01
0.054
5300E4
570OE5
-------
oo
CO
Table VI-13
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY URANIUM
SUBDIVISION MILL
MILL PROCESS ARID LOCATIONS
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOG
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
4
10
6
12
8
12
2
8
4
8
6
6
4
12
5
1
5
1
6
2
7
-) 'I
2
9
2
11
8
10
1
5
1
8
6
3
2
12
5
1
5
1
6
2
5
1
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.516
4 . 2602
0.274
. 14791
1.738
0.9966
0.046
1 . 9O76
0.036
2.3422
O. 1705
0.069
1 .205
26. 176
95.206
20
19134
24
6.43
0.0085
1462. 1
2300E4
290OE5
0.516
0.243
0.274
0. 1
1.575
0.485
0.046
1 .3
0.036
2.835
0. 1525
0.056
1 .205
22.365
26
20
64
24
7.45
0.0085
1660
2300E4
2900E5
1.03
10.6
0.295
0.4068
3.7
3.4
0.046
4. 18
0.036
3.68
0.49
0. 1
1 .24
59. 13
386
20
95450
24
8.3
0.01
2040
2300E4
2900E5
1.03
1O. 6
0.295
0.423
3.7
3.4
0.046
4. 18
O.O36
3.68
0.49
0. 1
1 .24
60.9
386
20
95450
24
8.3
0.01
2040
2300E4
2900E5
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF NUMBER
SAMPLES DETECTED
5
12
7
13
10
14
3
10
5
10
7
7
5
14
7
2
9
2
9
3
7
2
2
3
11
3
11
5
11
0
5
1
8
6
4
2
13
6
2
9
2
9
2
5
2
2
TREATED
(MG/L)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.299
. 11518
0.0072
.03545
0 . 0406
0. 192
0.3888
0. 14
.82637
.06383
0.016
0.79
4.729
59.505
6
55.611
21.5
6.65
0.01
1 .4164
1750E5
1750E6
100E-5
0.029
0.01
0.029
0.028
0.1
0.2
0. 14
0.955
0.0215
0.0195
0.79
2.52
10.5
6
26
21 .5
7.7
0.01
0.4
1750E5
1750E6
0.895
0.65
0.011
0.0746
0. 1
0.84
0.959
0. 14
1 .28
0.213
0.023
0.84
11 .06
279
10
157
27
8.45
0.01
3.87
2000E5
2300E6
0.895
0.75
0.011
0.077
0.1
O.9
0.959
0. 14
1 .28
0.213
0.023
0.84
11.1
279
10
157
27
8.45
0.01
3.87
2000E5
2300E6
-------
co
Table VI-14
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY TITANIUM
SUBDIVISION MINE
MILL PROCESS MINE DRAINAGE
RAW(MG/L) *
*
NUMBER OF NUMBER DETECTED VALUES ONLY * NUMBER OF NUMBER
SAMPLES DETECTED MEAN MEDIAN 90% MAX * SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
* 1 0
* 1 O
* 1 0
* 1 0
* 1 C
* 1 1
*
*
*
*
*
*
*
*
*
0
1
0
0
0
0
0
1
1
* 1 0
* 1 0
* 1 1
* 1 1
* 1 1
* 1 1
* 1 i
TREATED
(MG/L)
DETECTED
MEAN MEDIAN
0.02
0.02
0.02
2
8
7.95
0.03
140000
1900E3
0.02
0.02
0.02
2
8
7.95
0.03
140000
1900E3
VALUES ONLY
90% MAX
0.02
O.O2
0 02
2
8
7.95
0.03
140000
1900E3
0.02
0.02
0.02
2
8
7.95
0.03
140000
190OE3
-------
Table VI-15
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY TITANIUM
SUBDIVISION MILLS WITH DREDGE MINING
MILL PROCESS PHYSICAL AND/OR CHEMICAL
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
:-'
OD
cn
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
9
9
9
9
9
9
6
9
9
9
9
3
g
c
6
3
6
%
6
6
1
3
0
0
7
9
0
4
2
3
3
5
0
9
6
9
6
9
5
9
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.002
.00867
.04743
.02733
0.0375
0.006
0.023
0.031
0.007
.03122
1076.7
341 .44
485
5.9
0 . 0066
3. 1924
0.002
0.009
0.03
0.016
0.042
0.006
0.023
0.029
0.009
0.021
1060.5
160
560
5.7
0,007
1 .928
0.002
0.01
0.08
0.063
0.058
0.011
0.033
0.036
O.011
0.071
1900
1100
750
6.6
0.007
6.287
0.002
0.01
0.08
0.063
0.058
0.011
0.033
0.036
0.011
0.071
1900
1100
750
6.6
0.007
6.287
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
*
*
*
*
+
*
*
*
*
*
*
*
*
*
*
*
*
9
9
9
9
9
9
9
9
9
9
9
3
9
9
9
9
9
9
8
9
1
1
0
0
0
1
0
5
0
1
1
0
0
2
0
8
9
8
8
9
1
9
1
1
TREATED
(MG/L)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.002
0 . O058
0.005
100E-5
0.003
.02675
14
3.5625
5. 1875
5.9444
0.01
. 20533
3300E3
2700E3
0.002
0.006
0 005
100E-5
0.003
0.008
14
2.9
5.25
6.8
0.01
0. 171
3300E3
27OOE3
0.002
0.008
0.005
100E-5
0.003
0.071
17
9
6
7.6
0.01
0.5
3300E3
2700E3
0.002
0.008
0.005
100E-5
O.003
0.071
17
9
8
7.6
0.01
0.5
3300E3
2700E3
-------
Table VI-16
DATA SUMMARY
ORE MINING DATA
oo
SUBCATEGORY VANADIUM
SUBDIVISION MINE
MILL PROCESS NO MILL PROCESS
RAW(UG/L)
NUMBER OF NUMBER
SAMPLES DETECTED 101
ANTIMONY (TOTAL) 1 1 18
ARSKNIC (TOTAL) 1 1 130
BERYLLIUM (TOTAL) 1
CADMIUM (TOTAL) 1
CHROMIUM (TOTAL) 1
COPPER (TOTAL) 1
CYANIDE (TOTAL) 1
LEAD (TOTAL) 1
MERCURY (TOTAL) 1
NICKEL (TOTAL) 1
SELENIUM (TOTAL) 1
SILVER (TOTAL) 1
THALLIUM (TOTAL) 1
ZINC (TOTAL) 1
16.333
16.833
120
41 .833
3
317.5
1
446.67
6.3333
3
2
1476.7
PHENOLICS (4AAP) 1 0
IRON (TOTAL-MG/L) 1 1 69.133
MEAN
18
130
16.333
16.833
120
41 .833
317.5
1
446.67
6.3333
3
2
1476.7
69.133
MED
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
69.133
907,
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
69.133
*
*
*
*
*
A-
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF NUMBER
SAMPLES DETECTED
1
1
1
1
0
1
0
1
1
0
1
1
1 0
1 1
TREATED
10X
2
5
1
8.2
29.333
20.5
171 .33
59.333
12
1
159.17
.86467
(UG/L)
MEAN
2
5
1
8.2
29.333
20.5
171.33
59.333
12
1
159.17
.86467
MED
2
5
1
8.2
29.333
20.5
17'. 33
59.333
1 9
1
159.17
.86467
90%
2
5
1
8.2
29.333
20.5
171 .33
59.333
12
*
V'.9.'7
.56467
-------
Table VI-17
DATA SUMMARY
ORE MINING DATA
SUBCATEGORY VANADIUM
SUBDIVISION MILL
MILL PROCESS FLOTATION (FROTH)
oo
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NTCKEL (TOTAL)
SuLENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
PHENOLICS (4AAP)
IRON (TOTAL)
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
1
3
2
3
3
3
3
3
0
3
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
.03733
.29889
0.041
.18456
1 .6493
4.2113
0.29
5.7773
0. 1425
.78883
.86767
.03089
.29211
34. 165
135.24
.04167
.37333
0.038
0.0245
.47117
.06533
0.29
0.3175
0. 1425
.44667
0. 14
.02667
0.002
1 . 4767
69. 133
.05233
.39333
. O6867
.51233
4.3567
12.527
O.29
16.717
0.284
1 .8207
2 . 4567
0.063
.87333
100.82
334.9
.05233
.39333
.06867
.51233
4 . 3567
12.527
0.29
16.717
O.284
1 . 8207
2 . 4567
0.063
.87333
100.82
334.9
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
0
3
0
3
3
2
3
3
0
3
TREATED
(MG/L)
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
. 0206?
0. 108
.05317
.02384
. 14189
.03669
.55928
.13636
.08367
0.009
. 11667
. 11664
.55572
J.C14
0.014
0 03(5
0.025
.06133
.04225
0.3265
.09475
0.079
0.009
0.0015
0. 1115
0 . 5505
C.046
;"! . 3O5
0. 1225
.03833
0.3:15
04~33
1 . 18
0.255
O. 16
O.016
0.3475
. 15917
.86467
0.048
0.305
0. 1225
.03833
0.33S
.04733
1 . 13
0.255
0. 16
0.016
0.3475
. 15917
.36467
-------
TABLE V8-1 8 SUMMARY OF REAGENT USE (BY FUNCTION) (N ORE FLOTATION MILLS*
REAGENT
_ . .
Lime
Caustic soda
Soda Ash
Sulfuric Acid
Copper Sulfate
Copper Ammonium Chloride
Sodium Sulfhydrate
DESCRIPTION
CaOorCa(OH)2
Sodium Hydroxide,
NaOH
Sodium Carbonate,
Na2 C03
H2S04
CuSO4 or
CisS04 5H2O
CuNH2CI
Sodium Hydrosulfide
Ma SH 2H2O
FUNCTION
NUMBER OF
MILLS
WHERE
USED**
MODIFIERS
Alkaline pH regulator and depressant for
galena, metallic gold, iron sulfides, cobalt,
and nickel sulfide. Has flocculating effect
on ore slimes.
Alkaline pH regulator
Alkaline pH regulator w/slime dispersing
action.
Acidic pH regulator.
ACTIVATORS
Universal activator for sphalerite. Also used
for the reactivation of minerals depressed by
cyanide.
Activator for sphalerite. Purchased as a
waste by-product from the manufacture
of electric circuit boards.
Activator for copper sulfide mineral?
15
3
3
2
13
1
1
USUAL DOSAGE
kg/metric ton
ore feed
0.054 - 14.2
0.00015-0.025
0.54- 12.12
0.018-4.3
0.06 - 2.32
0.13 f
0.0094
c»
oo
* Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
1 Expressed as soluble copper metal.
**Reagent usage data supplied by 22 milling operations.
-------
TABLE VI-18 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
(Continued)
CO
4 H2O or
ZnSO4 7H2O
Na2 Cr2 O7
Corn starch
so2
Phosphorus Penta-
sulfide
P2S5
H2°2
Strong depressants for the iron sulfides,
arsenopyrite, and sphalerite. Mild
depressant for chalcopyrite, enargite,
bornite and most other sulfide minerals
w/ exception of galena.
Depressant for pyrite and sphalerite
while floating lead and/or copper.
Depressant for sphalerite while floating
lead and/or copper minerals. Often
used in conjunction w/ cyanide.
Depressant for galena in copper-lead
separations. Excess depresses copper
sulfides and iron sulfides.
Depressant for galena and molybdenite
while floating copper sulfides
Depressant for galena and activator for
copper sulfides. Often used in conjunc-
tion w/ starch.
Depressant for copper and lead while
floating molybdenite.
Depressant for copper sulfides in
copper-molybdenite separations.
13
2
7
1
4
2
4
1
0.003 - 0.065
0.2 - 7.46
0.1 - 1.35
0.022
0.0005 - 0.071
0.156-0.406
0.0001 - 0.47
0.016
'Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
* Reagent usage data supllied by 22 milling operations.
-------
TABLE VI-18 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
(Continued)
REAGENT
Sodium Silicate
AERO Depressant
610,633
Jaguar Mud
Xanthates:
AERO 301, 325, 343, 355
Dow Z-3, Z-4, Z-6, Z-1 1 ,
Z-14.
DESCRIPTION
Na2O: nSiO2
Composition unknown
Contains ~ 1.5%
phenolics
Colloidal material
Sodium or potassium
salts of xanthic acid.
SNa
R-O-C^
S
or
SK
R-O-C'
S
where R is an alkyl
group of 2-6 carbon
atoms.
FUNCTION
DEPRESSANTS
Depressant for quartz and other siliceous
gangue minerals. Also acts as slime
dispersant.
Depressant for graphitic and talcose
gangue. Also acts as gangue dispersants
useful in sand-slime separation.
Depressant for gangue materials
COLLECTORS/PROMOTERS
Strong promoters for all sulfide minerals.
Essentially non-selective in the absence of
modifiers.
NUMBER OF
MILLS
WHERE
USED**
5
3
1
17
USUAL DOSAGE
kg/metric ton
ore feed
0.031 - 2.08
0.001 -0.16
0.016
0.0003 - 0.40
o
*Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
**Reagent usage data supplied by 22 milling operations.
-------
TABLE VI-18 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
(Continued)
REAGENT
Dow Z-200
Fuel Oil
Vapor Oil
Tar Oil
Minerec
DRESSENATE TX-65W
SOAP
M.I. B.C. (Methyl
Isobutyl
Carbinol)
Methanol
Pine Oil
Cresylic Acid
DESCRIPTION
Isopropyl Ethyl-
Thionocarbamate
Saturated Hydrocarbons
Composition Unknown
Composition Unknown
Synonomous with
Methyl Amyl Alcohol
(CH3)2 CHCH2CHOHCH3
CH3OH
Composed primarily of
terpene hydrocarbons,
terpene ketones, and
terpene alcohols.
Higher homologs of
phenol, C6H5 OH,
particularly cresols,
CH3 C6H4 OH,
and xylenols,
C2H5 C6H4 OH,
or
(CH3)2 C6H4 OH
FUNCTION
Promoter for copper sulfides and activated
sphalerite w/ selectivity over iron sulfides.
Promoters, usually used for readily float-
able minerals, such as molybdenite.
Promoter.
Promoter.
FROTHERS
Alcohol type frothers are used for the
flotation of sulfide minerals where a
selective, fine textured froth is desired.
Frother
Frother, widely used in sulfide flotation.
It exhibits some collecting properties,
especially for such readily floatable
minerals as talc, graphite and molybdenite.
Pine oil produces a tough, persistent froth
and has a tendency to float gangue.
A powerful frother exhibiting some
collecting properties. Produces froth of
variable texture and persistence, and
tends to be non-selective.
NUMBER OF
MILLS
WHERE
USED**
3
4
1
1
10
1
5
3
USUAL DOSAGE
kg/metric ton
ore feed
0.004-0.10
0.0013-0.78
0.01
0.41
0.008-0.17
0.00005
0.015-0.175
0.003 - 0.034
* Copper, lead, zinc, silver, and molybedenum concentrators which discharge process wastewater (data available).
**Reagent usage data supplied by 22 milling operations.
-------
TABLE VI-18 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
(Continued)
REAGENT
Aliphatic
Dithiophosphates:
Sodium AEROFLOAT
AEROFLOAT 211, 249,
3477
AEROFLOAT 31 and 242
"Reco"
ARMAC "C"
DESCRIPTION
B.OX /S
sp\
where R is an alkyl
group of 2-6 carbon
atoms.
Aryi Dithiophosphoric
Acids
R - O S
R - O ^S H
where R is an aryl group
(benzene-based).
Sodium Dicresyl
Dithiophosphate
R.0^/S
R-O' ^SNa
where R is the cresyl
group:
CH3-C6H3-OH
Acetate Salt of
Aliphatic Amines
FUNCTION
Promoters of variable selectivity, and
strength for the flotation of sulfide materials.
Sometimes used in conjunction with
xanthates for improved precious metal
recoveries.
Promoters for copper, lead, zinc and
silver sulfide minerals. Has frothing
properties.
Promoter, selective to copper sulfide
minerals. Very similar to AEROFLOAT
31 and 242.
Promoter, very selective cationic collector.
NUMBER OF
MILLS
WHERE
USED**
5
4
1
1
USUAL DOSAGE
kg/metric ton
ore feed
0.015-0.043
0.012-0.05
0.016
0.0005
. i-- ___.,.,
ro
'Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
* Reagent usage data supplied by 22 milling operations,
-------
TABLE VI-H8 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
(Continued)
REAGENT
Polyglycols:
DOWFROTH 200, 250
AEROFROTH65
Diphenyl Guanidine
UCON-R-23
UCON-R-133
SYNTEX
AEROFLOC
AERODRI 100
VALCO 1801
NALCOLYTE 670
SEPARAN IMP-10
SUPER F LOG 3302
Flocculants (unspecified)
DESCRIPTION
Polyglycol Methyl
Ethers (i.e. Poly-
propylene glycol methyl
ether)
CH3 (OC3H6)X OH
HN-C (NHC6H5)2
Composition unknown
Composition unknown
Composition unknown
An ionic. Cationic, or
Nonionic Organic
Polymers.
FUNCTION
Frothers, for metallic flotation, w/ froth
persistency and selectivity against non
metals.
Frother
Frother
Frother
Frother
FLOCCULANTS
Used as dewatering aids or filtration
aids for thickening or filtering ore
pulps, concentrates, and tailings.
NUMBER OF
MILLS
WHERE
USED**
8
1
1
1
1
9
USUAL DOSAGE
kg/metric ton
ore feed
0002-0.17
0.00005
0.035
0.015
0.017
0.00015-0,051
<£}
CO
"Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
**Reagent usage data supplied by 22 milling operations.
-------
194
-------
SECTION VII
SELECTION OF POLLUTANT PARAMETERS
The Agency has studied ore mining and dressing wastewaters to
determine the presence or absence of toxic, conventional and non-
conventional pollutants. The toxic pollutants are of primary
concern to the development of BAT effluent limitations guide-
lines. One hundred and twenty-nine pollutants (known as the 129
priority pollutants) were studied pursuant to the requirements of
the Clean Water Act of 1977 (CWA). The 129 priority pollutants
are included in the 65 classes of toxic pollutants referred to in
Table 1, Section 307(a)(1) of the CWA.
EPA conducted sampling and analysis at facilities where BPT
technologies are in place; therefore, any of the priority
pollutants present in treated effluent discharges are subject to
regulation by BAT effluent limitations guidelines. The
Settlement Agreement in Natural Resources Defense Council,
Incorporated, v^ Train, 8 ERC 2120 (D.D.C. 1976), modified 12 ERC
1833 (D.D.C. 1979) provides a number of provisions for the
exclusion of particular pollutants, categories and subcategories.
The criteria for exclusion of pollutants are summarized below:
1. Equal or more stringent protection is already provided
by an effluent limitation and guideline promulgated pursuant
to Section(s) 301, 304, 306, 307(a), or 307(c) of the CWA.
2. The pollutant is present in the effluent discharge
solely as a result of its presence in the intake water taken
from the same body of water into which it is discharged.
3. The pollutant is not detectable in the effluent within
the category by approved analytical methods or methods
representing the state-of-the-art capabilities.
4. The pollutant is detected in only a small number of
sources within the category and is uniquely related to only
those sources.
5. The pollutant is present in only trace amounts and is
neither causing nor likely to cause toxic effects.
6. The pollutant is present in amounts too small to be
effectively reduced by technologies known to the
Administrator.
7. The pollutant is effectively controlled by the tech-
nologies upon which are based other effluent limitations and
guidelines.
195
-------
DATA BASE
Table VII-1 presents a summary of the data gathered for this
study. The sources of data are screen sampling, verification
sampling, verification monitoring, EPA Regional sampling, engi-
neering cost site visits, gold placer mining study, titanium sand
dredges study, uranium study, and the solid waste study. These
data are presented in complete form in Supplement A. The summary
table and extensive information about the sampled industries,
based on the criteria listed above, are used to determine which
pollutant parameters are excluded from regulation.
SELECTED TOXIC PARAMETERS
Several conventional and non-conventional pollutants were found
at all the facilities sampled; the 129 priority pollutants
occurred on a less frequent basis. The 13 metals listed as
priority pollutants, cyanide and asbestos were found at many of
the facilities. Six of the 13 metals were detected at levels too
low to be effectively reduced by the technologies known to the
Administrator. Eighty-seven (87) priority organic pollutants
were not found in the treated effluents during sampling. Of the
remaining 27 organic pollutants, 17 were found in the effluent of
only one or two sources and always at or below 10 ug/1, nine
pollutants were detected at levels too low to be effectively
reduced by technologies known to the Administrator, and one was
uniquely related to the source at which it was found.
The priority pollutants which were identified for controll by BAT
include arsenic, asbestos, cadmium, copper, lead, mercury,
nickel, zinc, and cyanide. The conventional parameters to be
regulated are pH and TSS. The non-conventional parameters are
COD, total and dissolved iron, radium 226 (dissolved and total),
ammonia, aluminum, and uranium. The priority pollutants,
conventionals, and non-conventionals for control in BAT are
displayed in Table VII-2. All 114 of the toxic organic
pollutants were excluded from regulation. The toxic metals were
excluded on a case-by-case basis within certain subcategories and
subdivisions. The reasons for exclusion are displayed in Table
VII-3.
EXCLUSION OF TOXIC POLLUTANTS THROUGHOUT THE ENTIRE CATEGORY
Pollutants Not Detected by_ Approved Methods
The toxic organic compounds are primarily synthetic and are not
naturally associated with metal ore. As shown in Table VII-1, 27
of the 114 toxic organics were detected during sampling, while 87
toxic organics were not detected in treated wastewater during
sampling. Therefore, the 87 toxic organics not detected are
excluded by Criterion 3 (the pollutant is not detectable by
approved analytical methods).
196
-------
Of the 27 toxic organics detected, 17 were detected at at least
one facility and always at or below 10 ug/1, which is the limit
of detection set by the Agency for the toxic organics in these
sampling and analysis programs.
1. Chlorobenzene
2. 1,1,1-Trichloroethane
3. Dichlorobromomethane
4. Chloroform
5. Fluorene
6. Ethylbenzene
7. Trichlorofluoromethane
8. Diethyl Phthalate
9. Tetrachloroethylene
10. Toluene
11. -BHC (Alpha)
12. -BHC (Beta)
13. -BHC (Delta)
14. Aldrin
15. Dieldrin
16. Endrin
17. Heptachlor
Thus, it follows that
under Criterion 3.
these 17 compounds are subject to exclusion
Pol_lut ants Detected But Preservt i_n Amounts Too Low to be
Effectively Reduced by_ Known Technologies
Toxic Organic Pollutants
There were 10 organic pollutants detected during the n,ine sam-
pling programs discussed in Section V at levels above 10 ug/1.
In general, the concentrations of nine of these pollutants are so
low that they cannot be substantially reduced. In some cases
this is because no technologies are known to further reduce them
beyond BPT; in other cases, the pollutant reduction cannot be
accurately quantified because the analytical error at these low
levels can be larger than the value itself. The following nine
are thus excluded from regulation because they were
amounts too low to be effectively reduced by tech-
pollutants
present in
nologies known to the Administrator (Criterion 6):
(1) Benzene
(2) 1,2-Trans-Dichloroethylene
(3) Phenol
(4) Bis(2-Ethylhexyl) Phthalate
(5) Butyl Benzyl Phthalate
(6) Di-n-Butyl Phthalate
(7) Di-n-Octyl Phthalate
(8) Dimethyl Phthalate
(9) Methylene Chloride
In addition, contamination during sample collection and analysis
has been documented for particular organic pollutants including
these nine, as discussed below.
Six of the 10 toxic organics detected are members of the
phthalate and phenolic classes. During sample collection,
automatic composite samplers were equipped with polyvinyl
chloride (Tygon) tubing or original manufacturer supplied tubing.
197
-------
Phthalates are widely used as plasticizers to ensure that the
Tygon tubing remains soft and flexible (References 1 and 2),
These compounds, added during manufacturing, have a tendency to
migrate to the surface of the tubing and leach into water passing
through the sampler tubing. In addition, laboratory experiments
were performed to determine if phthalates and other priority
pollutants could be leached from tubing used on composite
samplers. The types of tubing used in these experiments were:
1. Clear tubing originally supplied with the sampler at the time
of purchase
2. Tygon S-50-HL, Class VI (replacement tubing)
Results of analysis of the extracts representing the original and
replacement Tygon tubing are summarized in Table VI1-4. The data
indicate that both types contain bis(2-ethylhexyl) phthalate and
the original tubing leaches high concentrations of phenol.
Although bis(2-ethylhexyl) phthalate was the only phthalate
detected in the tubing in these experiments, a similar experiment
conducted as part of a study pursuant to the development of BAT
Effluent Limitations Guidelines for the Textiles Point Source
Category found dimethyl phthalate, diethyl phthalate, di-n-butyl
phthalate, and bis(2-ethylhexyl) phthalate in tubing "blanks"
(Reference 3).
Three of the volatile organic compounds (benzene, 1,2-transdi-
chloro-ethylene, and methylene chloride) were detected as a
result of the analysis of grab samples. The volatile nature of
these compounds suggests contamination as a possible source,
especially considering the relatively low concentrations detected
in the samples. More importantly, all of the compounds may be
found in the laboratory as solvents, extraction agents or aerosol
propellants. Thus, the presence and/or use of the compounds in
the laboratory may be responsible for sample contamination. This
type of contamination has been addressed in other studies
(Reference 4). In a review of a set of volatile organic blank
analytical data, inadvertent contamination was shown to have
occurred; the prominent compounds were benzene, toluene, and
methylene chloride.
The contamination by the volatiles as discussed above may be due
to the changing physical environment during the collection of
samples. The volatile sample is collected in a 45- to 125-ml
vial. During collection in the field, the sample vial is filled
completely with the wastewater, sealed (so that no air is present
in the vial at that moment) and chilled to 4 C until the time of
analysis. The volume of the water sample will decrease as it
cools from ambient conditions to 4 C, inducing an internal pres-
sure in the vial less than atmospheric. In addition, teflon
chips were used as lid liners to prevent contamination of the
sample by any compounds present in the lid. Experience in the
field has shown that it is difficult to ensure a tight seal at
198
-------
the time of collection because the teflon is not pliable. The
combination of the poor seal and the formation of the vacuum may
encourage contamination from the ambient laboratory atmosphere
where, as previously mentioned, volatile organic compounds are
prevalent. Methylene chloride, in particular, is .used in the
analytical procedure as a solvent (References 4, 5); this may
explain the detection of and high concentrations of this volatile
in 10 to 25 of the treated water samples (Table VII-1).
The presence of the three volatile organic compounds may be
attributed to sampling and analytical contamination, and as such,
they cannot be conclusively identified with the wastewater.
Toxic Metal Pollutants
Six toxic metal pollutants were detected.during the nine sampling
programs. Like the toxic organic pollutants, the concentrations
of these pollutants were so low that they cannot be substantially
reduced by known technologies. Each of the six metals is
discussed in more detail below.
Antimony. Antimony removal is discussed in Section VIII and in a
report by Hittman Associates (Reference 6). The conclusion of
these discussions is that antimony is very difficult to remove in
wastewater treatment. Using seven state-of-the-art technologies,
Hittman Associates could not attain lower than 500 ug/1 in the
effluent. Table VII-1 indicates that the maximum concentration
observed in effluent from this category was 200 ug/1. Therefore,
Criterion 6 (the pollutant is present in amounts too small to be
effectively reduced by known technologies) is applicable and
antimony is excluded from regulation in this category.
Thallium. Thallium removal is discussed in a report prepared
examining the analysis protocol for this toxic metal (Reference
7). The conclusion that may be drawn from this discussion is
that the procedure used in the analysis of thallium is subject to
interferences which prevent its conclusive identification in
wastewater samples. Thallium is, therefore, excluded from BAT
regulation since it cannot be conclusively identified in waste-
water samples by approved analytical procedures (Criterion 3).
Selenium. There are little data in the literature on selenium
removal from industrial wastewater, treatment methods for
selenium wastes, or costs associated with removal of selenium
from industrial wastewater (Reference 8). Selenium is present in
trace amounts in metallic sulfide ores. Generally the selenium
is released in the smelting and refining process and is not
liberated during mining and milling. Although 37 samples of 73
contained detectable selenium, the mean value reported was 0.059
mg/1. Ninety percent of the samples in which selenium was
detected contained 0.112 mg/1 or less. Most of the samples con-
tained very low levels of selenium as indicated by a median value
of only 0.015 mg/1. No specific treatment data or application of
199
-------
specific treatment could be found in the ore mining and milling
industry. Pilot-scale treatment studies have reported removals
ranging from 10 to 84 percent using conventional technologies,
but only cation and anion exchange used in combination achieved
high removal efficiencies (Reference 8). The removals obtained
by utilizing conventional technology are not consistent enough to
base regulations upon them, and cation and anion exchange, while
possibly providing additional treatment, are judged too costly
for this industry. Consequently, selenium is excluded from
regulation since it is found at levels too low to be effectively
reduced by known technologies (Criterion 6).
Silver. Most data available on treatment of waste streams
containing silver represent attempts at recovery of this valuable
metal from the photographic and electroplating industries. In
these industries there has been an ample economic incentive to
develop recovery/removal technology because: (1) the silver is
valuable and may be reused; and (2) the concentration levels
present favor the economic recovery of silver from the waste
stream. Four basic methods for silver removal from wastewater
are discussed in Reference 8: (1) precipitation, (2) ion
exchange, (3) reductive exchange, and (4) electrolytic recovery.
Levels to 0.1 mg/1 have been reported by various investigators
but most of these in bench- or pilot-scale systems. In addition,
waste streams in this industry are high in solids, effluent flow
rates are very high, and treated effluent levels are already low
(mean of treated effluent samples 0.015 mg/1; maximum 0.04). It
has been concluded that the concentration levels present are too
low to be effectively reduced by known technologies (Criterion
6).
Beryllium. There are little data available in the literature for
beryllium removal in wastewater from the ore mining and dressing
industry. Only one domestic facility mines beryllium ore and
uses water in a beneficiation process (Mine/Mill 9902). This
mill uses a proprietary leach process with raw wastewater having
beryllium at a concentration of 36 mg/1 at pH 2.6 (Reference 9).
This facility has no discharge. However, when TSS in the
impoundment is reduced from 116,000 to 44,000 mg/1 (after a short
settling period), beryllium is reduced to 25 mg/1. Since
beryllium is relatively insoluble, it is believed that reduction
to an effluent level of TSS of 20 mg/1 after lime precipitation
and settling would result in a substantial reduction of beryllium
levels.
In a related industry, primary beryllium refining, some data are
available which indicate effluent beryllium levels of 0.09 mg/1
are possible by lime precipitation and multiple pond settling
(Reference 10). Of 73 effluent samples analyzed during BAT
screening for beryllium, only 10 samples had detectable beryllium
concentrations with a mean of 0.005 mg/1 and maximum
concentration "of 0.011 mg/1. These are the levels which are
achieved by known technologies and since the pollutant is present
200
-------
only in trace amounts, it has been concluded that the present
concentration level would not be effectively reduced and,
therefore, this pollutant is excluded (Criterion 6).
Chromium. Data acquired during this study measured total
chromium levels (as opposed to dissolved) regardless of valence
state. Extensive literature references are available for treat-
ment of wastewater with respect to either trivalent or hexavalent
chromium. However, these references predominantely address waste
streams from the electroplating industry, dyes, inorganic pig-
ments, and metal cleaning operations. The natural mineral,
chromite (FeC^O^), has chromium in the trivalent form. Of 75
treated effluents for which chromium measurements were made, only
26 had detectable total chromium concentrations with a median
value of 0.035 mg/1 (for detected values only). Trivalent
chromium is effectively removed by the BPT treatment, lime
precipitation and settling at the pH range normally encountered
in treatment systems (pH 8 to 9) associated with this industry.
Consequently, it has been concluded that the concentration levels
present at most facilities would not be effectively reduced
further by the known technologies (Criterion 6).
Pollutants Detected in Treated Effluents a_t a Small Number of_
Facilities and Uniquely Related"to Those Facilities
The toxic organic pollutant, 2,4-dimethylphenol, was detected in
the effluent at only one facility (9202) during the screen
sampling program. AEROFLOAT^, used as a flotation agent in ore
beneficiation at this facility, is a precursor of 2,4-dimethyl-
phenol. However, since the compound was identified only in one
facility, it is excluded under Paragraph 8(l)(iii) of the Revised
Settlement Agreement,
EXCLUSION OF TOXIC POLLUTANTS BY SUBDIVISION AND MILL PROCESS
The toxic parameters which did not qualify for exclusion
throughout the entire category (i.e., toxic metals, cyanide, and
asbestos) were evaluated for potential exclusion in each subcate-
gory. Table VI-1 summarizes the sampling data for the toxic
metals, asbestos and cyanide, by subcategory, subdivision and
mill process. The number of representative samples taken and a
summary of influent and effluent data are provided in the table.
This table was reviewed and the data evaluated for possible
exclusion from regulation based on the criteria previously dis-
cussed. In particular, Criterion 3 (not detected in treated
effluents by approved analytical methods) and Criterion 6 (pres-
ent at levels too low to be effectively reduced by known technol-
ogies) were used to exclude certain toxic metals and cyanide from
particular subdivisions and mill processes. The toxic pollutant
parameters chosen for exclusion and the exclusion criteria are
summarized by category, subdivision, and mill process in Table
VII-3.
201
-------
CONVENTIONAL POLLUTANT PARAMETERS
Total Suspended Solids (TSS)
Total suspended solids (or suspended solids) are regulated for
all subcategories under BPT effluent limitations. High suspended
solid concentrations result as part of the mining process, and by
crushing, grinding, and other processes commonly used in milling.
Dredging and gravity separation processes also produce high
suspended solids. Effluent limitations are proposed for total
suspended solids under BCT.
EH
This parameter is regulated for every subcategory under BPT
effluent guidelines; BCT effluent limitations will apply in the
same manner. Acid conditions prevalent in the ore mining and
dressing industry may result from the oxidation of sulfides in
mine waters or discharge from acid leach milling processes.
Alkaline-leach milling processes also contribute waste loading
and can adversely affect receiving water pH.
BOD, Oil and Grease
These conventional parameters are not regulated under BPT
effluent guidelines and were not found in significant
concentrations during development of the data base. They are not
applicable to an industry which deals primarily in inorganic
substances.
NON-CONVENTIONAL POLLUTANT PARAMETERS
Settleable Solids
Solids in suspension that will settle in one hour under quiescent
conditions because of gravity, are settleable solids. This
parameter is most useful as an indicator of the operating
efficiency of sedimentation technologies, particularly
sedimentation ponds, and is recommended for use as such to
establish effluent limitations for gold placer mines.
Iron
Iron is very common in natural waters and is derived from common
iron minerals in the substrata. The iron may occur in two forms:
inherently increases iron levels present in process and mine
drainage. The aluminum ore mining industry also contributes
elevated iron levels through mine drainage. Iron, both total and
dissolved, is regulated for segments of the industry under BPT
effluent limitations and effluent guidelines are developed for
iron under BAT effluent limitations.
202
-------
Radium 226
Radium 226 is a member of the uranium decay series and, as
discussed in Section III, it is always found with uranium ore.
As a result of its long half-life (1,620 years), radium 226 may
persist in the biosphere for many years after its introduction
through effluents or wastes. Therefore, because of its radio-
logical consequences, concentrations of this radionuclide must be
restricted to minimize potential exposure to humans. It is regu-
lated under BPT effluent limitations because of the radiological
consequences and because data indicate that control of radium 226
also serves as a surrogate control for other radionuclii and is
regulated under BAT effluent limitations.
Ammonia
Ammonia compounds (e.g., ammonium hydroxide) may be used as
precipitation reagents in alkaline leaching circuits in uranium
mills. The sodium diuranate which results from leaching, recar-
bonization and precipitation is generally redissolved in sulfuric
acid to remove sufficient sodium to meet the specifications of
American uranium processors. The uranium values are precipitated
with ammonia to yield a yellowcake low in sodium. By-product
ammonium sulfate and excess ammonia remaining may flow to waste-
water treatment downstream. Consequently, ammonia is regulated
under BPT and will be regulated under BAT.
CONVENTIONAL AND NON-CONVENTIONAL PARAMETERS SELECTED
A review of the data collected subsequent to BPT effluent
guidelines development serves to confirm parameter selections
made for BPT. No new parameters were discovered in significant
quantitities in any subcategory. Therefore, development of BAT
regulations for conventional and non-conventional parameters will
be for the same parameters regulated under BPT. Table VII-2
illustrates the parameters to be regulated by subcategory and
subpart.
SURROGATE/INDICATOR RELATIONSHIPS
The Agency believes that it may not always be feasible to
directly limit each toxic which is present in a waste stream.
Surrogate/indicator relationships provide an alternative to
direct limitation of toxic pollutants. A surrogate relationship
occurs between a toxic pollutant and a set of commonly regulated
parameters when the concentration(s) of the regulated param-
eter^) are used to predict the concentration of the toxic pollu-
tant. When the concentration(s) of the regulated parameter(s)
are used to predict whether or not the toxic pollutant level will
be reduced, it is an indicator relationship. In the first
instance, the regulated parameter(s) are called surrogates and in
the second, they are called indicators.
203
-------
The advantage of the surrogate/indicator relationship is that, by
regulating certain conventional and non-conventional parameters,
toxic pollutants are controlled to the same degree as if they had
been directly controlled. Only those toxics whose concentrations
can be quantitatively predicted based on knowledge of the
concentration of one or more regulated parameters can be
indirectly limited in this manner. Surrogates and indicators are
discussed more fully in the Federal Register, Vol. 44, No. 166,
pp. 34397-9.
Statistical Methods
Surrogate/indicator relationships were developed for several of
the priority pollutant metals which were selected for regulation.
The statistical methodology used in the development of these
relationships included the following phases:
1. exploratory data analysis
2. model estimation
3. model verification
The objective of the exploratory data analysis phase was to
assess the likelihood of accurately specifying the chemical and
physical relationships between the priority pollutant metals and
the potential surrogate/indicator parameters, given the
limitations of the data available for the analysis. Summary
statistics, plots, and correlations were examined. The model
estimation phase quantified the relationships which were identi-
fied during the exploratory data analysis phase by using regres-
sion analysis. The model verification phase assessed the valid-
ity of the models by applying them in a simulated regulatory
situation. The relationships were tested on a separate set of
data from that used in the estimation phase.
Relationships
A statistical analysis of pollutant concentrations in ore mining
wastewaters indicates a relationship between TSS and the
following toxic pollutants:
1. chromium
2. copper
3. lead
4. nickel
5, selenium
6. zinc
7. asbestos
Therefore, when treatment technologies are employed for reducing
TSS there is a reduction in the levels of these toxics. The
relationship and indicated control was used in selecting
technologies considered for BAT, technologies which reduce TSS,
as discussed further in Section VIII.
204
-------
Additonal Paragraph 8_ Exclusion
As discussed in Section X, additional paragraph 8 exclusions were
made during the selection of BAT options and BAT effluent
limitations. These exclusions included the decision to regulate
asbestos (chrysotile) by limiting the discharge of TSS as
discussed in Section X. The reader is referred to the additional
information and supporting data found in a separate report
entitled, "Development of Surrogate/Indicator Relations in the
Ore Mining and Dressing Point Source Category." Also, cyanide is
not regulated because the Agency cannot quantify a reduction in
total cyanide by use of any technology known to the
Administrator. The Agency concluded that limitations on copper,
lead, and zinc would ensure adequate control of arsenic and
nickel. Finally, EPA excluded uranium mills from BAT because the
pollutants found in the discharge are uniquely related to a
single sources.
205
-------
Table VII-1
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
NUMBER OF
SAMPLES
ACENAPHTHENE
ACROLEIN
ACRYLONITRILE
BENZENE
BENZIOENE
CARBON TETRACHLORIDE
CHLOROBENZENE
1.2. 3-TRICHLOROBENZENE
HEXACHLOROBENZENE
,2-DICHLOROETHANE
. 1 , 1-TRICHLOROETHANE
HEXACHLOROETHANE
, 1-OICHLOROETHANE
ro . 1 . 2-TRICHLOROETHANE
§> , 1 . 2. 2-TETRACHLOROETHAN
CHLOROETHANE
BIS(CHLOROMETHYL) ETHER
BIS(2-CHLOROETHYL) ETHER
2-CHLOROETHYL VINYL ETHE
2 - CHLORONAPHTHAL ENE
2 , 4 . 6-TRICHLOROPHENOL
PARACHLOROMETA CRESOL
CHLOROFORM
2-CHLOROPHENOL
1 . 2-DICHLOROBENZENE
1 , 3 DI CHLOROBENZENE
1 . 4 DICHLOROBENZENE
3, 3-OICHLOROBENZIDINE
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
32
32
32
32
32
32
32
32
32
32
RAW(UGXL) *
NUMBER DETECTED VALUES ONLY *
DETECTED MEAN MED 90% MAX *
0
0
0
10
0
1
0
0
0
0
9
0
0
0
0
0
0
0
0
0
1
0
9
O
0
0
O
0
*
*
*
4.8922 4 10 10 *
*
1 1 1 1 *
*
*
*
*
6.7208 6. S811 10 10 *
*
*
*
*
11.667 11.667 11.667 11.687
7.6096 3.1623 12.5 35 *
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
23
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED (UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
0
0
0
3
0
0
1
0
0
0
5
0
0
0
O
0
0
0
0
O
0
0
8
0
0
0
0
0
8.3333 7 1O. 7 11
0.005 O.005 O.OO5 0.005
7.2649 6.5811 10 10
5.1281 3.1823 10 10
-------
Table VII-1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
NUMBER OF
SAMPLES
1, 1-DICHLOROETHYLENE
1 , 2-TRANS-DICHLOROETHYLE
2 . 4-DICHLOROPHENOL
1 . 2-DICHLOROPROPANE
1 . 3-DICHLOROPROPENE
2.4-DIMETHYLPHENOL
2.4-DINITROTOLUENE
2.6-DINITROTOLUENE
1,2-DIPHENYLHYDRAZINE
ETHYLBENZENE
*LUORANTHENE
METHYL CHLORIDE
METHYL BROMIDE
BROMOFORM
DICHLOROBROMOMETHANE
TRICHLOROFLUCROMETHANE
DICHLORODIFLUOROMETHANE
CHLORODIBROMOMETHANE
HEXACHLORQBUTADIENE
HEXACHLOROCYCLOPENTADIEN
ISOPHORONE
NAPHTHALENE
NITROBENZENE
2-NITROPHENOL
4-NITROPHENOL
2.4-DINITROPHENOL
4,6-DINITRO-O-CRESOL
32
32
32
32
32
32
32
32
32
32
32
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
RAW(UG/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
2 6.5811 3.1623 8.6325 10
0
1 10 1O 10 10
0
0
1 140 140 140 140
0
0
0
4 6.7187 1 13.48 17.667
0
1 45 45 45 45
0
0
0
5 5.0325 2.0811 10 10
0
0
0
0
0
1 12.5 12.5 12.5 12.5
0
0
0
0
0
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED (UG/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 9O% MAX
0
1 270 270 270
0
O
O
1 270 27O 270
0
0
0
3 6.6 4.9 9.64
O
0
0
0
2 6.5811 3.1623 8.6325
3 4.7208 2.0811 7.9487
0
O
0
0
0
0
0
0
0
0
0
27O
270
1O
10
10
-------
Table VII-1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
RAW(UQ/L) *
NUMBER OF
SAMPLES
N-NITROSODIMETHYLAMINE
N-NITROSODIPHENYLAMINE
N-NITROSODI -N-PROPYLAMIN
PENTACHLOROPHENOL
PHENOL
BIS(2-ETHYLHEXYL) PHTHAL
BUTYL BENZYL PHTHALATE
DI-N-BUTYL PHTHALATE
DI-N-OCTYL PHTHALATE
DI ETHYL PHTHALATE
DIMETHYL PHTHALATE
BENZO( A) ANTHRACENE
BENZO(A)PYRENE
BENZO( B ) FLUORANTHENE
BENZO(K)FLUORANTHENE
CHRYSENE
ACENAPHTHYLENE
ANTHRACENE
BENZO(G,H, I )PERYLENE
FLUORENE
PHENANTHRENE
DIBENZO( A, H) ANTHRACENE
INDENO( 1,2, 3-C, D)PYRENE
PYRENE
TETRACHLOROETHYLENE
TOLUENE
TRICHLOROETHYLENE
33
33
33
33
33
33
33
33
10
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
NUMBER DETECTED
DETECTED MEAN MED
O
0
0
1
2
15
2
13
3
16
0
0
0
0
0
0
0
0
0
1
0
0
0
0
2
9
0
10
118
20.16
10.75
18.489
10
24.414
10
7.75
399.28
10
76
13
0.5
10
10
10
10
4.5
2 . 08 1 1
VALUES ONLY
90% MAX
10
143.2
39.833
16.9
26. 1
10
59.4
10
9,7
368.3
10
160
100
21
56
10
90
10
11
3560
»
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
7
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED
(UQ/L)
NUMBER DETECTED
DETECTED MEAN MED
0
0
0
0
3
18
4
12
3
4
3
0
0
0
0
0
0
0
0
1
0
0
0
0
1
6
0
92.3
12.458
27.791
25.864
12. 167
7.875
12.2
10
1. 1
2 . 5967
33.45
10
10
10
10
9.6
5.8
10
1.1
1
VALUES ONLY
90% MAX
166.8
26
52.4
39.2
14.55
10
20.35
10
1.1
5.28
21O
50
66
140
16.5
10
25
1O
1. t
10
-------
Table VII-1 (Continued)
ORE MINING DATA
ALL SUB CAT EGORIES
DUMBER OF
SAMPLES
VINYL CHLORIDE
ALQRIN
CIELDRIN
CHLORDANE
4.4-DDT
4, 4 -DDE
4,4-DOO
ENDOSULFAN- ALPHA
ENDOSULFAN-BETA
ENDOSULFAN SULFATE
ENDRIN
ENDRIN ALDEHYDE
HEPTACHLOR
HEPTACHLOR EPOXIDE
BHC-ALPHA
BHC-BETA
BHC { LINDANE) -GAMMA
BHC- DELTA
PCB-1242 (AROCHLOR 1242)
PCS- 1254 (AROCHLOR 1254)
PC8-1221 (AROCHLOR 1221)
PCB-1232 (AROCHLOR 1232)
PCS -1248 (AROCHLOR 1248)
PCB-1260 (AROCHLOR 1260)
PCB-1016 (AROCHLOR 1016)
TOXAPHENE
2,3,7 , 8-TETRACHLORODIBEN
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
9
9
9
9
9
32
33
RAW(UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
O
4 6.4156 5 9 10
0
0
0
1 5555
1 6.6667 6.6667 6.6667 6.6667
1 10 10 10 10
0
0
0
0
1 7.5 7.5 7.5 7.5
0
S 5.2649 4.0811 7.5 10
5 6.1325 5 8.75 10
4 6.2072 5 8.6667 10
2 5555
0
0
0
0
0
0
0
0
0
*
*
*
*
*
*
*
*
*
*
*
*
*
if
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
NUMBER OF
SAMPLES
28
28
28
28
28
23
28
28
28
28
23
28
28
28
28
28
28
28
28
28
e
6
6
6
6
27
28
TREATED (UG/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
0
2 8.5811 3.1623 8.6325
2 6.5811 3.1623 8 . 8325
0
0
0
0
O
0
0
1 S 5 5
0
2 8.5811 3.1323 8.3325
0
3 555
1 555
0
2 555
0
0
0
0
0
0
O
0
0
10
10
S
10
6
5
5
-------
Table VII-1 (Continued)
DATA SUMMARY
ORE MINING DATA
ALL SUBCATEGORIES
RAW(UG/L) * TREATED (UO/L)
NUMBER OF NUMBER DETECTED VALUES ONLY * NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MED 90% MAX * SAMPLES DETECTED MEAN MED 90% MAX
CIS 1-3-DICHLOROPROPYLEN *
TRAN 1,3-DICHLOROPROPYLE *
ro
t>
o
-------
Table VII-1 (Continued)
ro
UA 1« SUMM4HY
OFT. IIINIK'
Ml SUIT « 1 : I OKI F S
'"<''' """->
NUMi.'i'" OF niirn.r rtir^Ti_i vAnirs ONI.
SA"H_rs rirn.rtH: rt/v.\ MM M>, r..u
AM; t rr I.Y
t' l k f l ! I U t'
C ADM in1 f
CHKiiMUil
COf-ITK (1
c v t N r; ' (
LF Ml < (0 1
HER f.lU Y (
JICKLl (1
"»M. i ;"I U"
si L vf r ( T
HULL i UN
( T i. I ft |. )
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( 1 n ' u L ^
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1 I 0 T .< L )
m /> L )
in M. j
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i a f A t >
HIM)
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< T 0 T i!.. 1 1' ii. (';?!=
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Hi 1 1' 0 f- '.-.:, 1 5 . H
f- r, ; i . ^ r ' ** i. r . r i
h f. 70 i. . 1 'l ri 7 1 . 3 5
P 7 M . i(M"" . C01 'i5
« f- 70 .--.'-iff, ; . 5
Fa -, ( [i . i ^ c r- r- . 1 L
H» ; ' ..' " or i i . r i
T .' . : ( i. ' j 1.17
i ^ f. i n f 3 ' . M P ^ n . p 7 ij
. ? r n .' a .: . 7 ; i o . " h
17 * 7 3 1 ':> f. . f- 1 C
6 I- 1'is r6t>
.'. ') "- f, . .'' n ? 1 7
7 ;' 71 . 1 1 , 1 7 ,? " . n i
,"i' i '- '>; i . 7 (. . PH-
II . l
0 . f ? 5
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P ? 1 . b
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0.013
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If £. (j «
1 ? ^ *
130 >
0.02
11."
1.5 *
1.1 «
1.20
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q
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P2
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"b
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5f
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? 1
TfU ATF[)
< MO/LI
(JF. TFCTFJ)
MFAN MCOlAf
0-031
. 1531°
0.0051
.01115
.13623
.2316*
.13571
.131S)
.01261
.2220J
. 05957
.015t7
.5?7f 7
.90209
1 1 .69P
2 1 . y n 9
5 . 1P75
fr .?711
. 0 7 3 7 P
. 6 2 { I "
mnr. -5
o.oie
0.0 H5
0.035
C . f!6
0.08
0.05
«OOI -6
0.07
0 . 0 I 5
n.oi^
0.71
0.06?
11
9.5
5.25
7. 7
O.P3?
n.ro0
VAl lirj ONI Y
"OT M«X
0.1
0.6
0.0109
0.06
0.332
0.516
0.135
0.376
0.0306
0.966
0.112
0.01
0 .PI
2 2'i1
20.6
6°. 2
6
B.16
0.21
2-192
0.1
0.011
0.077
l.P
1.6
0.6
0.959
0.25
1 .28
0.9
n.o»
0.81
11.1
53
157
6
P. 5
0.16
3.87
-------
TABLE VII-2. POLLUTANTS CONSIDERED FOR REGULATION
ro
i'
ro
Subcetegory
Iron Or«
Copper, Lead,
Zinc, Gold,
Silver, Platinum,
Molybdenum
Aluminum
Tungsten
Mercury
Uranium
Antimony
Titanium
Nickel
Vanadium
Sub-
Division
Mines
Mills
Mines
Mills
Mines
Mines
Mills
Mines
Mills
Mines
Mills, In
Situ Leach
Mines
Mills
Mines
Mills
Mills w/
Dredges
Mines
Mills
Mines
Milt Process
Phys/Chem
Phys (Mesabi!
Cyanidation or Amalgamation
Heap, Vat, Dump, In Situ Leach
Froth Flotation
Gravity Separation
Mills |
Toxic Pollutants
Sb
E
E
As
E
E
tChrysotile
G
G
Be
E
E
Cd
E
E
Cr
E
E
Cu
E
E
CN
E
E
Pb
E
E
Hi)
E
E
Ni
E
E
Se
E
E
Ag
E
E
Tl
E
E
Zn
E
E
Total Phenolics
E
E
Conven-
tional
_TSSJ
G
G
PH
G
G
Nonconventiona s
COD
Fe - Total or Diss.
G
G
Ra
226
Al
U
Settlaabl* Solidi
Zero Discharge at 8PT
E
E
G
E
G
E G
G
G
G
G
E
E
E
G
E
G
G
Zero Discharge at BPT
Zero Discharge at BPT
E
E
E
E
G
E
G
G
G
G
G
G
E
E
E
£
G
E
G
G
E
£
E
E
G
i
G
G
G
E
E
E
G
E
E
G
G
E
E
E
G
G
E
E
E
G
E
E
E
E
E
E
E
E
E
E
E
E
G
E
G
G
G
E
E
E
G
G
G
G
G
G
G
G
G
^ G
G
G
G
G
Zero Discharge at BPT
E
E
E
E
G
G
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
G
E
E
E
E
E
E
G
G
E
E
G
G
G
G
G
G
G
G
G
Reserved
E
E
E
E
E
E
G
G
G
E
E
E
f
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
G
c
E
E
E
E
E
E
E
E
E
E
G
E
E
E
E
G
G
G
G
a
G
G
G
Reserved
Reserved
_
E c Excluded from Guideline Development
G ** GuidGiines to be Considered
-------
TABLE VSI-3. PRIORITY METALS EXCLUSION CRITERIA BY SUBCATEGORY. SUBDIVISION,
MILL PROCESS
Co
Subcategory
Iron Ore
Capper, Lead,
Zinc, Gold,
Silver, Platinum,
Molybdenum
Aluminum
Tungsten
Mercury
Uranium
Antimony
Titanium
Nickel
Vanadium
Sub
Division
Mines
Mills
Minas
Mills
Mines
Mines
Mills
Mines
Mills
Mines
Mills. In
Situ Leach
Mines
Mills
Mines
Mills
Mills w/
Dredges
Mines
Mills
Mines
Mills
Mil! Process
Phys/Chem
Phys (Mesatai)
Cyanidstion of Amalgamation
Heap, Vat, Dump, InSitu Leach.
Froth Flotation
Gravity Separation
Toxic PoNntanb
Sb
3
3
As
6
6
Chryntlle
Be
3
3
Cd
3
3
Cr
3
6
Cu
6
6
CM
3
3
Pb
3
3
Hg
3
3
Ni
3
3
Se
3
3
Ag
3
3
Tl
3
3
Zn
6
6
«
o
r-
3
3
Zero Discharge at BPT
3
6
6
3
S 3 ! 6
6
6
Zero Discharge aft BPT
Zero Discharge at BPT
6
3
3
6
3
6
3
3
6
3
6
6
6
3
3
3
3
3
6
6
3
3
3
3
6
3
3
3
6
3
3
G
3
3
3
3
3
Zaro Discharge at BPT
3
3
3
e
3
6
3
3
3
3
3
3
3
3
3
3
3
3
6
6
6
3
3
3
3
6
3
3
6
6
6
6
3
S
6
,
6
6
6
3
3
3
6
S
S
3
3
6
3
3
3
3
3
3
6
3
3
6
3
3
3
3
3
6
6
3
e
6
6
6
6
E
6
6
Reserved
"-
EXCLUSION CRITERIA
protection B already
l»OT«fcd by EPA's
2. The pollutant is present
as a nsuht of its presence
3. The pollutant is not
ibyj
5- The pollutant a present m
f'- The poHHtent is effectively
imiuulle«i by treating
other poHiMants.
-------
Table VII-4
TUBING LEACHING ANAYSIS RESULTS
Micrograms/Liter
Component Originaj. Tygon
Bis (2-ethylhexyl) Phthalate
Acid Extract 915 N.D.
Base-Neutral Extract 2,070 885
Phenol
Acid Extract 19,650 N.D.
Base-Neutral Extract N.D. N.D.
N.D. - Not Detected
214
-------
SECTION VIII
CONTROL AND TREATMENT TECHNOLOGY
This section discusses the techniques for pollution abatement
applicable to the ore mining and milling industry. General
categories of techniques are: in-process control, end-of-pipe
treatment, and best management practices. The current or poten-
tial use of each technology in this and similar industries and
the effectiveness of each are discussed.
Selection of the optimal control and treatment technology for
wastewater generated by this industry is influenced by several
factors:
1. Large volumes of mine water and mill wastewater must be
controlled and treated. In the case of mine water, the
operator often has little control over the volume of water
generated except for diversion of runoff from surface mine
areas.
2. Seasonal and daily variations in the amount and charac-
teristics of mine water are influenced by precipitation,
runoff, and underground water contributions.
3. There are differences in wastewater composition and
treatability caused by ore mineralogy, processing
techniques, and reagents used in the mill process.
4. Geographic location, topography, and climatic conditions
often influence the amount of water to be handled, treatment
and control strategies, and economics.
5. Pilot plant testing and acquisition of empirical data
may be necessary to determine appropriate treatment
technologies for the specific site.
6. The availability of energy, equipment, and time to
install the equipment must be considered. Selection of BAT
by mid-1980 will give the industry three years to implement
the technology.
IN-PROCESS CONTROL TECHNOLOGY
This section discusses process changes available to existing
mills to improve the quality or reduce the quantity of wastewater
discharged from mills. The techniques are process changes within
existing mill?
Control of_ Cyan_ide
Cyanide is a commonly used mill process reagent, used in froth
flotation as a depressant and in cyanidation for leaching.
215
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Froth Flotation
In the flotation of complex metal ores, depressing agents assist
in the separation of one mineral from another when flotabilities
of the two minerals are similar for any given combination of
flotation reagents. Cyanide is a widely used depressant, either
in the form of crude calcium cyanide flake or sodium cyanide
solution. Alkaline cyanides are strong depressants for the iron
sulfides (pyrite, pyrrhotite, and marcasite), arsenopyrite, and
sphalerite. They also act as depressants, to a lesser extent,
for chalcopyrite, enargite, tannantite, bornite, and most other
sulfide minerals, with the exception of galena (Reference 1).
Cyanide, in some instances, cleans tarnished mineral surfaces,
thereby allowing more selective separation of the individual
minerals (Reference 2).
In flotation, cyanide has primarily been used to aid in the
separation of galena from sphalerite and pyrite. It also has
been used to separate silver and copper sulfides from pyrite,
nickel and cobalt sulfides from copper sulfides, and molybdenum
sulfide from copper sulfide.
In beneficiation of base metal ores by flotation, the rate of
cyanide addition to the circuit must be varied to optimize both
the percentage recovery and concentrate grade of the various
metals recovered (References 1, 2, 3, and 4). The addition of
either too much or too little cyanide can result in loss of
recovery and reduction in the grade of concentrate. For example,
in selective flotation of copper, lead/zinc, and copper/lead/zinc
ores, the addition of too little cyanide will result in the flo-
tation of pyrite, thereby reducing the copper, lead, and zinc
concentrate grades. Also, too little cyanide will result in
flotation of zinc in the lead circuit, which produces a lower
lead concentrate grade.
Cyanide control is also desirable from a waste treatment
standpoint. Excess cyanide use subsequently requires more copper
sulfate when zinc is activated for flotation. This not only
represents uneconomical use of reagents, but also increases the
waste loading of both copper and cyanide. Reagent use at various
domestic base metal flotation mills and comments relative to the
efficiency of cyanide use in these mills are described in Section
VI, Summary of Reagent Use in Flotation Mills.
Many mills have replaced valve operated reagent feeders for
cyanide addition with metered feeders, such as the Clarkson or
Geary feeder, which maintain constant flow of a controlled
solution of cyanide. The use of these metered feeders influence
the amount of cyanide fed to the process by insuring that the
proper amount required is added and, thereby, reducing the
possibility of "overshooting" the correct dosage. Also, some of
these same mills have imposed restrictions on which personnel can
adjust these automatic feeders to eliminate the arbitrary
216
-------
increase in dosage that can overshoot the minimum amount required
to produce the most efficient separation.
The degree of sophistication of in-process control of cyanide
varies widely in the category. The greatest degree of sophisti-
cation is used at copper/lead/zinc Mill 3103. A Courier online
X-ray analyzer performs analyses at 10-minute intervals of the
mill heads and tails and of the concentrates, heads, and tails of
the individual flotation circuits. Analytical results are com-
puted by a Honeywell 316 computer and automatically printed and
charted. The mill operator may then adjust the rate of reagent
addition based on these analytical results. For example, the
rate of cyanide addition is decreased when the copper content of
the copper circuit tailings increases during a time increment
(usually two hours). Conversely, the rate of cyanide addition is
increased when the iron content of the lead and zinc concentrates
increases. In this manner, the mill operator is able to optimize
the reagent use, percentage recovery, and grade of concentrate
produced^ Several mills (3103, 3105, 3122, 3123, 2117, and 2121)
have on-stream analytical capabilities.
Laboratory analysis provides adequate control in the milling of
simple, "clean" ores. However, the greater the complexity and
variability of the ore being milled, the more advantageous it
becomes to a mill operator to have on-stream analytical
capabilities.
The prevailing practice for in-process control of reagent
addition consists of manual sampling and laboratory analysis of
heads, tails, and concentrates. Typically, samples are collected
at two-hour intervals and analyses are begun immediately.
Approximately two hours are required before analytical results
are available. This method is slower and produces less informa-
tion than the more sophisticated method previously described.
Many small mills have limited analytical capabilities and the
control of reagent addition depends on the experience of the mill
operator. According to mill operators and site visit data,
cyanide addition in excess of the amount required is generally
used with limited analytical control.
Control of the rate of reagent addition depends on the attention
given to the analytical results by the mill operator. The atti-
tude, conscientiousness, and experience of the mill operator have
a significant effect on the degree of control maintained over
reagent usage. The efficiency of reagent usage impacts the over-
all efficiency and economy of the mill, as well as the character
of the wastewater generated, and operators must remain aware of
this.
217
-------
Cyanidation
Cyanide is also used prominently in processing lode gold and
silver ores by a leaching process (the cyanidation process) which
uses dilute, weakly alkaline solutions of potassium or sodium
cyanide. In-process control of cyanide at cyanidation mills
involves recycle of the spent leach solutions. This control
practice is, therefore, beneficial in two respects. First, the
cyanide wasteload is greatly reduced, making treatment more eco-
nomical. Second, since a fraction of the cyanide is recovered
for reuse, the cost of reagent is reduced. The BPT effluent
guidelines for cyanidation mills is no discharge of process
wastewatcr.
Alternatives tc> Cyanide I_n_ Flotation
Cyanide is believed to function primarily as a reducing agent in
the depression of pyrite in xanthate flotation operations. In
1970, Miller (Reference 5) investigated alternative reducing
agents and found that, in terms of effectiveness and cost, sodium
sulfite compared quite favorably with cyanide as a pyrite
depressant. In particular, it was found that cyanide exerted
some depressant effect on chalcopyrite and pyrite, but sodium
sulfite did not. The sodium sulfite alternative appeared to be
applicable to copper ore flotation operations.
Some mills use sulfite or sulfides instead of cyanide. Mill 3101
is an example of a copper/zinc flotation mill which uses sodium
sulfite and no cyanide. There was no measurable effect on
recovery or grade of concentrate. At Mill 6104, copper and
molybdenum minerals are separated in froth flotation with sodium
bisulfide used as a copper depressant, provide another
alternative to the use of cyanide for the depression of copper
minerals in selective flotation.
An EPA-sponsored study to identify and evaluate alternatives to
sodium cyanide was initiated in May 1978 (References 6 and 7) and
alternatives were identified by a literature search. Points
taken into consideration were: ability to depress pyrite,
selectivity of depressant, theory of performance, inferred and
specific environmental aspects, state of development as a practi-
cal depressant, and cost. Fourteen alternatives were identified,
three of which were carried into the evaluation phase of the
study. The compounds selected for bench-scale evaluation were:
sodium monosulfide (Na2S), sodium sulfite (Na2S03_), and sodium
thiosulfate (Na2S203_). Three types of ore (copper, copper/lead/
zinc, and zinc) were chosen for the flotation experiments. All
of the ores contained pyrite.
The results of this study are summarized in Table VIII-1. The
most effective depressant in the copper ore experiments was
sodium cyanide. Sodium sulfite at 0.504 kg/metric ton (1.008
pound/short ton) of ground ore approached the effectiveness of
218
-------
the cyanide at the natural pH level, natural meaning the pre-
vailing pH of the ground ore plus water. At elevated pH (10 to
12), sodium sulfite and sodium monosulfide surpassed cyanide in
the amount of copper recovered, but these were less effective in
depressing the pyrite.
When dealing with copper/lead/zinc ore, it is desirable to float
the copper and lead initially, while depressing the iron and
zinc. At the natural pH level, the sodium sulfite equaled the
cyanide in recovery of copper and lead and was superior to the
cyanide in depressing iron and zinc. At pHs of 10 to 12, sodium
sulfite surpassed the cyanide in the recovery of copper and lead
and nearly equaled the cyanide in depression of iron and zinc.
Sodium monosulfide resulted in good recoveries of copper and
lead, but not as good as other alternatives. It was ineffective
in depressing the pyrite.
In the experiments with the zinc ore, only sodium sulfite and
sodium monosulfide were studied. Zinc/pyrite ore is one of the
most difficult ores to float and the study confirms this. Tech-
niques used to improve the floatability of zinc ore were not
applied in the experiments. At the natural pH level, the low
level of sodium sulfite surpassed sodium cyanide in the recovery
of zinc and was slightly less effective in the suppression of
iron. At elevated pH values, all of the alternatives studied,
including the absence of a depressant, out-performed sodium cya-
nide. Sodium monosulfide was the most effective alternative
under the high pH conditions.
In summary, bench-scale tests indicate that sodium sulfite is a
potential substitute for sodium cyanide. Also, sodium monosul-
fide is fairly effective at high pH. However, these are bench
scale tests, and full-scale operations in this industry rarely
equate directly to bench-scale results. Typically, extensive
bench and pilot scale testing with the particular ore to be
milled are conducted by an operator before the decision to con-
vert is made. Even then, weeks or months of adjustments may be
necessary to optimize the new process.
Reagent cost estimates are given in Table VIII-1, and the
difference in cost is negligible. However, reagent cost is only
one of the economic considerations. Components of the cost are
(1) reagent costs, (2) downtime, (3) laboratory process
simulation costs, (4) equipment cost, and (5) optimization costs.
The last are probably the highest. Interviews with operators
revealed that downtime may be only a few days, but optimizing the
process may take a year and concentration grades, they fear,
would never reach current standards. The financial penalties can
be severe, as evidenced by one mill's report on smelter penalties
for offgrade lead concentrates (mill process 700 TPD, Reference
7):
219
-------
1. For every 0.1 percent of copper in excess of 1.0 percent, the
penalty could amount to $96,000 per year.
2. For every 0.1 percent of iron in excess of 4 percent, the
penalty could amount to $264,000 per year.
The study concludes that conversion costs are complex and cannot
be accurately estimated (Reference 7).
Therefore, cyanide substitution should not be the basis for
selection of BAT effluent guidelines, as the cost of substitution
cannot be calculated and an economic analysis cannot be con-
ducted. However, if a particular mill can meet BAT effluent
guidelines by reagent substitution and maintain concentrate
quality, that option is available.
Alternatives to Use of_ Phenolic Compounds As_ Mill Reagents
Several phenolic compounds are used in this industry. The most
common is cresylic acid, which is essentially 100 percent
phenolics, and is used as a frothing agent at several base and
precious metals flotation mills (e.g., 2117 and 4403). A
frother, pine oil, used in sulfide mineral flotation, is composed
essentially of terpene alcohols, terpene ketone, and terpene
hydrocarbons. These terpene compounds are not phenolics, but
some phenolics are likely to be present as byproducts of the
steam distillation process used to produce them. Several
collectors (promoters), such as Reco and AEROFLOAT, also contain
phenolic radical groups, In isolated instances, depressants
containing phenolics have been used. At one mill (2120), a
phenolic compound (Nalco 8800) is used as a wetting agent for
dust control during secondary ore crushing. In this latter case,
nonphenolic wetting agents, including olefinic compounds and
petroleum-based sulfonates, are being considered for use.
The flotation reagents and dosages used vary widely from mill to
mill (refer to Table VI-19). Reagent and dosage rate selection
is a complex process that often takes years to optimize and is
continuously reevaluated at individual mills. Considerations
include reagent cost and availability, compatibility with other
reagents, effect on concentrate grade and metal recoveries,
consistency of the ore body, and environmental impact of chemical
residuals in the wastewater discharge. Selection of dosage rate
is essentially a trial and error process of optimizing concen-
trate grade and metal recoveries and is dependent upon in-process
control.
The chemistry of flotation is complex and reagent substitution
may have repercussions throughout a circuit. However, a large
number of nonphenolic frothers are promising as alternatives to
phenol-based or phenol-containing compounds. Among the most pop-
ular ncnphenolic frothers are methyl isobutyl carbinol (MIBC) and
polyglycol rr.cthyl ethers. Frothers are generally nonselective,
220
-------
generic-ally related compounds, with fairly predictable charac-
teristics. As such, substitution within this class of reagents
(frothers) should not be difficult. For example, Mill 2121 has
recently discontinued use of cresylic acid in a silver flotation
circuit by substitution with polyglycol methyl ethers.
Collectors are much more selective than frothers, and their
effectiveness is highly dependent upon their compatibility with
associated modifiers, promoters, activators, and depressants.
Possible alternatives to phenolic collectors are dithiophosphate
salts and dithiophosphoric acids with alkyl groups in place of
phenol groups. Substitution of these reagents for phenol con-
taining collectors may be feasible without serious complications
or economic consequences; however, the consequences of substitu-
tion are site dependent and require extensive experimentation at
each mill.
In-Process Recycle of Waste? Streams
In-process recycle of concentrate thickener overflow and/or
recycle of filtrate produced by concentrate filtering is prac-
ticed at a number of flotation mills (e.g., 2121, 3101, 3102,
3108, 3115, 3116, 3119, 3123, and 3140). In addition, several
mills (2120, 6101, and 6157) use thickeners to reclaim water from
tailings prior to the final discharge of these tailings to ponds.
Water reclaimed in this manner is used as makeup water in the
mill. In-process recycle of waste streams produced by concen-
trate dewatering is incorporated primarily as a process control
measure for the recovery of metals which would otherwise be lost
in the tailings. This latter practice is intended as a safeguard
in the event of concentrate filter malfunctions, which would
allow large quantities of metals to pass through the filter. To
avoid this, these process waters are generally returned to the
flotation circuits.
These practices conserve water and recover metals which would
otherwise be in the wastewater discharge. The in-process recycle
of concentrate thickener overflow and/or filtrate produced by
concentrate filtering reduces the volume of wastewater discharged
by 5 to 17 percent. Likewise, mills which use thickeners to
reclaim water from tailings reduce both the new water requirement
and the volume of wastewater discharged by 10 to 50 percent.
In-process recycle to reduce wastewater volume can improve the
performance of existing treatment systems. For example, as the
volume of wastewater discharged from a mill decreases, the
retention time within the tailing pond increases. As a result,
conditions favorable to settling of solids, formation of metal
precipitates, and degradation of flotation reagents and cyanide
(by chemical, physical, and biochemical mechanisms) are enhanced.
Therefore, the in-process recycle of wastewater can be an effec-
tive means of improving the capabilities of existing tailing
ponds.
221
-------
Recycle of spilled reagent can also be an advantage. At Mill
3101, the occurrence of spills and overflow from flotation cells
results from the milling of a higher grade ore than the reagent
dosage is optimized for. A system has been implemented to
collect spills and return them to the flotation circuit. This
control practice not only improves the quality of treated waste-
water, but the percentages of metals recovered as well.
Use of_ Mine Water as_ Makeup in. the Mill
A large number of mine/mill operations use mine drainage as
makeup in the mill (e.g., 4103, 4104, 4105, 3101, 3102, 3103,
3104, 3105, 3106, 3108, 3110, 3113, 3118, 3119, 3122, 3123, 3126,
3127, 3138, 3142, 6102, 6104, 9402, and 9445). In some
instances, the entire process water requirement of the mill is
obtained from mine drainage.
From a wastewater treatment aspect for facilities allowed to
discharge, a great advantage is gained by this practice. First,
this practice either eliminates the requirement for a mine water
treatment system or greatly reduces the volume of wastewater
discharged to a single system. As discussed previously, reducing
the volume of wastewater flow to an existing treatment system can
be an effective means of enhancing the capabilities of that sys-
tem. Second, in situations where mine water contains relatively
high concentrations of soluble metals, its use in the mill pro-
vides a more effective means for the removal of these metals than
could generally be attained by treatment of the mine water alone.
This is due to reduced metals solubility in the alkaline condi-
tions maintained in flotation and most mill circuits. Therefore,
use of mine water as makeup in a mill can be considered a control
practice which improves the quality of mine and mill treated
wastewater.
Techniques for Reduction of_ Wastewater Volume
Pollutant discharges from mining and milling sites may be reduced
by limiting the total volume of discharge, as well as by reducing
pollutant concentrations in the wastestream. Volumes of mine
discharges are not, in general, amenable to control, except inso-
far as the mine water may be used as input to the milling process
in place of water from other sources. Techniques for reducing
discharges of mill wastewater include limiting water use, exclud-
ing incidental water from the waste stream, recycle of process
water, and impoundment with water lost to evaporation or trapped
in the interstitial voids in the tailings.
In most of the industry, water use should be reduced to the
extent practical, because of the existing incentives for doing so
(i.e., the high costs of pumping the high volumes of water
required, limited water availability, and the cost of water
treatment facilities). "Incidental water enters the waste stream
directly through precipitation and through the resulting runoff
222
-------
influent to tailing and settling ponds. By their very nature,
the water-treatment facilities are subject to precipitation
inputs which, due to large surface areas, may amount to substan-
tial volumes of water. Runoff influxes are often many times
larger, however, and may be controlled to a great extent by
diversion ditches and (where appropriate) conduits. Runoff
diversion exists at many sites and is under development at
others.
Complete Recycle - Zero Discharge
Mill Water-
Recycle of process water is currently practiced where it is
necessary due to water shortage, where it is economically
advantageous because of high make-up water costs, or the cost to
treat and discharge. Some degree of recycle is accomplished at
many ore mills, either by reclamation of water at the mill or by
the return of decant water to the mill from the tailing pond or
secondary impoundments. Recycle is becoming, and will continue
to become, a more frequent practice. The benefits of recycle in
pollution abatement are manifold and frequently are economic as
well as environmental. By reducing the volume of discharge,
recycle may not only reduce the gross pollutant load, but also
allow the employment of abatement practices which would be
uneconomical on the full waste stream. Further, by allowing
concentrations to increase in some instances, the chances for
recovery of certain waste components to offset treatment cost
or, even, achieve profitabilityare substantially improved. In
addition, costs of pretreatment of process waterand, in some
instances, reagent usemay be reduced.
Recycle of mill water almost always requires some treatment of
water prior to its reuse. In many instances, however, this may
entail only the removal of solids in a thickener or tailing
basin. This is the case for physical processing mills, where
chemical water quality is of minor importance, and the practice
of recycle is always technically feasible for such operations.
In flotation mills, chemical interactions play an important part
in recovery, and recycled water may, in some instances, pose
problems. The cause of these problems, manifested as decreased
recoveries or decreased product purity, varies and is not, in
general, well-known, being attributed at various sites and times
to circulating reagent buildup, inorganic salts in recycled
water, or reagent decomposition products. Experience in arid
locations, however, has shown that such problems are rarely
insurmountable. In general, plants practicing bulk flotation on
sulfide ores can achieve a high degree of recycle of process
waters with minimal difficulty or process modification. Complex
selective flotation schemes can pose more difficulty, and a fair
amount of work may be necessary to achieve high recovery with
extensive recycle in some circuits. Problems of achieving suc-
223
-------
cessful recycle operation in such a mill may be substantially
alleviated by the recycle of specific process streams within the
mill, thus minimizing reagent crossover and degradation. The
flotation of non-sulfide ores (such as scheelite) and various
oxide ores using fatty acids, etc., has been found to be quite
sensitive to input water quality. Water recycle in such
operations may require a high degree of treatment of recycle
water. In many cases, economic advantage may still exist over
treatment to levels which are acceptable for discharge, and
examples exist in current practice where little or, no treatment
of recycle water has been required.
A large number of active mills employ recycle of process
wastewater and achieve zero discharge. The following list is not
all inclusive, but serves to illustrate the degree to which this
practice has been adopted in the ore milling industry:
Iron Mills Copper Mills
1101 1118 2103 2118
1102 1122 2108 2139
1103 1123 2109 2140
11051129 2113 2141
1106 1138 2115 2146
1112 2116 2147
Lead/Zinc Mercury
3105 2126 9202
3123 2143
Copper ore leaching (heap, dump, in-situ) operations practice
recycle in order to reuse the acid and to maximize the extraction
of copper values by hydrometallurgical methods.
Technical limitations on recycle in other ore leaching operations
center on inorganic salts. The deliberate solubilization of ore
components, most of which are not to be recovered, under recycle
operations can lead to rapid buildup of salt loads incompatible
with subsequent recovery steps (such as solvent extraction or ion
exchange). In addition, problems of corrosion or scaling and
fouling may become unmanageable at some points in the process.
The use of scrubbers for air-pollution control on roasting ovens
provides another substantial source of water where recycle is
limited. At leaching mills, roasting will be practiced to
increase solubility of the product material. Dusts and fumes
from the roasting ovens may be expected to contain appreciable
quantities of soluble salts. The buildup of salts in recycled
scrubber water may lead to plugging of spray nozzles, corrosion
of equipment, and decreased removal effectivenes as salts
crystallizing out of evaporating scrubber water add to particu-
late emissions.
Impoundment and evaporation are techniques practiced at many
mining and milling operations in arid regions to reduce
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discharges to, or nearly to, zero. Successful employment depends
on favorable climatic conditions (generally, less precipitation
than evaporation, although a slight excess may be balanced by
process losses and retention in tailings and product) and on
availability of land consistent with process-water requirements
and seasonal or storm precipitation influxes. In some instances
where impoundment is not practical on the full process stream,
impoundment and treatment of smaller, highly contaminated streams
from specific process may afford significant advantages.
Total and partial recycle have become more common in recent
years. Facilities that use recycle are often in arid regions
because of the scarcity of available water. Many facilities both
in arid and humid regions recycle their process wastewater.
Mine Water
Complete recycle of mine drainage is generally not a viable
option because often an operator has little control over water
which infiltrates the mine. Except for small amounts of water
used in dust control, cooling, drilling fluids, and transport
fluids for sluicing tailings back to the mine for backfill, water
is not widely used in the actual mining. In some cases, mine
drainage is used by the mill as process water in beneficiation.
However, the volume of mine drainage may exceed the mill's
requirement for process water, making complete use of mine
drainage unachievable.
Other Process Changes
Mill 4105 has, as a result of environmental regulation,
discontinued the use of mercury (amalgamation) for the recovery
of gold. The process change used consists of incorporation of a
cyanidation circuit, described as a carbon-in-pulp circuit. This
process technology is described in detail in Appendix A.
At uranium Mill 9405, a process change has recently been
implemented specifically as a pollution control measure.
Yellowcake precipitation with sodium hydroxide, rather than
ammonia, is now used to reduce ammonia levels entering the
receiving stream. Although only limited experience has been
gained, plant personnel have noted a reduction in product grade
resulting from the process change. However, product grade is
apparently still within acceptable limits.
Mill 9403 has indirectly eliminated all wastewater discharges
recently by eliminating their resin-in-pulp circuit (see Section
III). This change was not based solely on environmental consid-
erations, but resulted from a variety of factors which included
ore characteristics, process economics, and pollution control
requirements.
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END-OF-PIPE TREATMENT TECHNIQUES
This subsection presents discussions of several end-of-pipe
techniques which are used in industry or are applicable to the
treatment problems encountered. These technologies were
considered as possible BAT technologies. However, it should be
noted that at many facilities in the industry, implementation of
additional technology beyond BPT will not be necessary to meet
the limitations based on BAT technology. The reasons for this
are facility specific and may include low-waste loading due to
clean ore, extremely well managed treatment systems, existing
systems exceeding BPT requirements, extensive reuse of
wastewater, and water conservation practices. The description of
the candidate BAT technologies includes the discussion of the
processes involved and their degree of use in the industry,
treatability data collected by Agency contractors, and finally,
historical data where available.
Technique Description
Secondary Settling
Ponds are used in the industry for settling. Tailings ponds
receive relatively .high solids loading and therefore require
frequent cleaning or enlargement. Primary settling ponds for
mine drainage used to meet BPT effluent guidelines have larger
surface areas, receive larger solids loadings than secondary
ponds, and may not require cleaning or dredging. Secondary
settling ponds are sometimes used to provide better solids
removal by plain (nonchemical aided) sedimentation.
In theory, several ponds in a series will not remove any more
solids than one large pond of equal size, since the theoretical
detention time in the two situations are identical. However,
many sediment ponds currently in use in this industry have not
been designed, operated, and maintained so as to optimize set-
tling efficiency. Therefore, in practice, providing secondary
settling in a series of ponds has been demonstrated to provide
additional reduction of suspended solids in this industry.
For example, short circuiting in the primary pond (either tailing
or settling), too much depth in the primary pond, shock hydraulic
loads (such as precipitation runoff), and an improper discharge
structure in the primary pond are all cases where secondary
settling ponds can remove significant quantities of solids by
plain settling.
Coagulation and Flocculation
Coagulation and flocculation are terms often used interchangeably
to describe the physiochemical process of suspended particle
aggregation resulting from chemical additions to wastewater.
Technically, coagulation involves the reduction of electrostatic
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surface charges and the formation of complex hydrous oxides.
Time required for coagulation is short; only what is necessary
for dispersing the chemicals in solution. Flocculation is the
physical process of the aggregation of wastewater solids into
particles large enough to be separated by sedimentation, flota-
tion, or filtration. Flocculation typically requires a detention
of 30 minutes.
For particles in the colloidal and fine supracolloidal size
ranges (less than one to two micrometers), natural stabilizing
forces (electrostatic repulsion, physical repulsion by absorbed
surface water layers) predominate over the natural aggregating
forces (van der Waals) and the natural mechanism which tend to
cause particle contact (Brownian motion). The function of chemi-
cal coagulation of wastewater may be the removal of suspended
solids by destabilization of colloids to increase settling velo-
city, or the removal of soluble metals by chemical precipitation
or adsorption on a chemical floe.
The inorganic coagulants, or flocculants, commonly used in
wastewater treatment are aluminum salts such as aluminum sulfate
(alum), lime, or iron salts such as ferric chloride. Hydroxides
of iron, aluminum, or (at high pH) magnesium form gelatinous
floes which are extremely effective in enmeshing fine wastewater
solids. These hydroxides are formed by reaction of metal salt
coagulants with hydroxyl ions from the natural alkalinity in the
water or from the addition of lime or another pH modifier.
Sufficient natural iron and/or magnesium is normally present in
wastewater of this industry so that effective coagulation can be
achieved by merely raising the pH with lime addition. Lime and
metal salt coagulants also act to destabalize colloidal solids,
neutralizing the negatively charged solids by adsorption of
cations.
Polymeric organic coagulants, or polyelectrolytes, can be used as
primary coagulants or in conjunction with lime or alum as a
coagulant aid. Polymeric types function by forming physical
bridges between particles, thereby causing them to agglomerate.
Polymers also act as filtration aids by strengthening floes to
minimize floe shearing at high filtration rates.
Coagulants are added upstream of sedimentation ponds, clarifiers,
or filter units to increase the efficiency of solids separation.
This practice has also been shown to improve dissolved metals
removal due to the formation of denser, rapidly settling floes,
which appear to be more effective in adsorbing and absorbing fine
metal hydroxide precipitates. The major disadvantage of
coagulant addition to the raw wastewater stream is the production
of large quantities of sludge, which must remain in perpetual
storage within tailing ponds. Coarser mineral materials thicken
as particulate (nonflocculant) suspensions, yet most materials
(especially pulps, precipitates, slimes, tailings, and various
wastewater treatment sludges) are flocculant suspensions and
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behave quite differently. Sedimentation is the only process
occurring during thickening of particulate suspensions, with the
weight of the particles borne solely by hydraulic forces. Two
physically different processes occur during thickening of
flocculant suspensions: sedimentation of separate floes and
consolidation of the flocculant porous medium, in which the
weight of the particles is borne partially by mechanical means
and partially by hydraulic forces. In efforts to reduce the
solids load on primary sedimentation units, several mine/mill
wastewater treatment systems add chemical coagulants after the
larger, more readily settled particles have been removed by a
settling pond or other treatment. Polyelectrolyte coagulants are
usually added in this manner.
In most cases, chemical coagulation can be used with minor
modifications and additions to existing treatment systems,
although the cost for the chemicals is often significant.
However, a model coagulation and flocculation system may consist
of a mixing basin, followed by a flocculation basin, followed by
a clarifier or settling pond and possibly a filtration unit. The
purpose of the mixing basin is to disperse the coagulant into the
waste stream; the reason for the flocculation basin is to
increase the collisions of coagulated solids so that they
agglomerate to form settleable or filterable solids. This is
accomplished by inducing velocity gradients with slowly revolving
mechanical paddles or diffused air.
A low capital cost alternative to the model system and one that
is well suited to the industry involves introduction of the
coagulant directly into wastewater discharge lines, launders, or
conditioners (in the flotation process). The coagulated waste-
water is then discharged to a sedimentation pond or tailing pond
to effect flocculation and sedimentation of the coagulated
solids. The advantages of this system, as opposed to the model
treatment facility, are minimization of treatment units and capi-
tal expenditures, and treatment simplicity resulting in reduced
maintenance and increased system reliability. Disadvantages of
this system are lack of control over the individual treatment
processes and potentially reduced removal efficiency.
The effectiveness and performance of individual flocculating
systems must be analyzed and optimized with respect to mixing
time, chemical-coagulant dosage, retention time in the
flocculation basin (if used) and peripheral paddle speed,
settling (retention) time, thermal and wind-induced mixing, and
other factors.
Coagulation and flocculation are used at several facilities in
this industry. Coagulants (polymers) are presently used for
wastewater treatment at Mine/Mills 4403, 3121, 3120, and 1108 and
at Mine 3130. In the past, flocculants have also been employed
at Mine/Mills 2121 and 3114. At Mine/Mill 1108, the tailing pond
effluent is treated with alum, followed by polymer addition and
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secondary settling to reduce suspended solids from approximately
200 mg/1 to an average of 6 mg/1. At Mine/Mill 3121, initiation
of the practice of polymer addition to the tailings has greatly
improved the treatment system capabilities. Concentrations of
total suspended solids (TSS), lead, and zinc in the tailing-pond
effluent have been reduced over concentrations previously
attained, as shown in the tabulation below (company-supplied
data):
Parameter
TSS
Pb
Zn
Effluent Levels (mg/1)
Attained Prior to
Use of Polymer
Effluent Levels (mg/1)
Attained Subsequent
to Use of Polymer
Mean
39
0.51
0.46
Range Mean
15 to 80 14
0.24 to 0.80 0.29
0.23 to 0.86 0.38
Range
4 to 34
0.14 to 0.67
0.06 to 0.69
Similarly, the use of a polymer at Mine 3130 reduced treated
effluent concentrations of total suspended solids, lead, and zinc
over concentrations attained prior to use of the polymer, as
shown in the following (company supplied data);
Effluent Levels (mg/1) Effluent Levels (mg/1)
Attained Prior to Use Attained Subsequent to
of Polymer and Secondary Use of Polymer and
Parameter Settling Pond* Secondary Settling Pond*
TSS
Pb
Zn
Mean
19
0.34
0.45
Range Mean
4 to 67 2
0. 1 1 to 1.1 0.08
0.23 to 1.1 0.32
Range
0.02 to 6.2
less than 0.05 to 0.10
0.18 to 0.57
*Secondary settling pond with 0.5 hour retention time.
Filtration
Filtration is accomplished by the passage of water through a
physically restrictive medium with the resulting deposition of
suspended particulate matter. Typical filtration applications
include polishing units and pretreatment of input streams to
reverse osmosis and ion exchange units. Filtration is a versa-
tile method in that it can be used to remove a wide range of
suspended particle sizes.
Filtration processes can be placed in two general categories; (1)
surface filtration devices, including microscreens and
diatomaceous-earth filters and (2) granular media filtration, or
in-depth filtration devices such as rapid sand filters, slow sand
filters, and granular media filters.
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Microscreens are mechanical filters which consist of a
horizontally mounted rotating drum. The periphery of the drum is
covered by fabric woven of stainless steel or polyester, with
aperture sizes from 23 to 60 micrometers. Microscreens have
found fairly widespread process application for concentrate
dewatering, but are less used in wastewater treatment
applications because of sensitivity to solids loadings and the
relatively low filtration rates required to prevent chemical floe
shearing and subsequent filter penetration.
Diatomaceous earth (DE) filters have been applied to the
clarification of secondary sewage effluent at pilot scale and
they produce a high quality effluent. However, they are
relatively expensive and appear unable to handle the solids
loadings encountered in this industry.
Next to gravity sedimentation, granular media filtration is the
most widely used process for the separation of solids from
wastewater. Most filter designs use a static bed with vertical
flow, either downward or upward, using gravity or pressure as the
driving force.
Slow sand filters are single, medium-gravity granular filters
without a means of backwashing. The filter is left in service
until the head loss reaches the point where the applied effluent
rises to the top of the filter wall. Then the filter is drained
and allowed to partially dry, and the surface layer of sludge is
manually removed. Such filters require very large land areas and
considerable maintenance. For these reasons, they are not com-
petitive tertiary treatment processes other than for small pack-
age plants.
Rapid sand filters are much the same as slow sand filters in that
they are composed of a single type of granular medium which is
drained by gravity and hydrostatic pressure. The primary differ-
ence between the two is the provision for backwashing of the
rapid sand filter by reversing the flow through the filter.
During filter backwashing, the media (bed) is fluidized and
settles with the finest particles at the top of the bed. As a
result, most of the solids are removed at or near the surface of
the bed. Only a small portion of the total voids in the bed are
used to store particulates, and head loss increases rapidly.
Despite this disadvantage, rapid sand filters are relatively
common in potable water supply treatment plants.
Effective filter depth can be increased by the use of two or more
types of granular media. Granular media filters typically use
coal (specific gravity about 1.6), silica sand (specific gravity
about 2.6), and garnet (specific gravity about 4.2) or ilmente
(specific gravity about 4.5), with total media depths ranging
from about 50 cm (20 inches) to about 125 cm (48 inches).
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Pressure filters are often advantageous in waste treatment
applications for the following reasons: (1) pressure filters can
operate at higher heads than are practical with gravity filter
designs, thus increasing run length and operational flexibility;
(2) the ability to operate at higher head losses reduces the
amount of wash water to be recycled; and (3) steel shell package
units are more economical in small and medium-size plants
(Reference 8).
Whenever possible, designs should be based on pilot filtration
studies using the actual wastewater. Such studies are the only
way to assure: (1) representative cost comparisons between dif-
ferent filter designs capable of equivalent performance (i.e.,
quantity filtered and filtrate quality); (2) selection of optimal
operating parameters such as filter rate, terminal head loss, and
run length; (3) effluent quality; and (4) determining effects of
pretreatment variations. Ultimate clarification of filtered
water will be a function of particle size, filter medium
porosity, filtration rate, and other variables.
Granular media filtration has consistently removed 75 to 93
percent of the suspended solids from lime treated secondary
sanitary effluents containing from 2 to 139 mg/1 of suspended
solids (Reference 9). One lead/zinc complex is currently oper-
ating a pilot-scale filtration unit to evaluate its effectiveness
in removing suspended solids and nonsettleable colloidal metal
hydroxide floes from its combined mine/mill/smelter/ refinery
wastewater. Preliminary data indicate that the single medium
pressure filter operated at a hydraulic loading of 2.7 to 10.9
1/sec/m2 (4 to 16 gal/min/ft2) is capable of removing 50 to 95
percent of the suspended solids and 14 to 82 percent of the
metals (copper, lead, and zinc) contained in the waste stream.
Final suspended solids concentrations which have been attained
are within the range of less than 1 to 15 mg/1. Optimum filter
performance has been attained at the lower hydraulic loadings;
performance at the higher hydraulic loadings appears to degrade
significantly.
A full-scale granular media filtration unit is currently in
operation at molybdenum Mine/Mill 6102. The filtration system
consists of four individual filters, each composed of a mixture
of anthracite, garnet, and pea gravel. This system functions as
a polishing step following settling, ion exchange, lime precipi-
tation, electrocoagulation, and alkaline chlorination. Since its
startup in July 1978, the filtration unit has been operating at a
flow of 63 liters/second (1,000 gallons/minute), and monitoring
data from November 1979 to August 1980 have demonstrated signifi-
cant reductions of TSS, Mo, Cu, Pb, Cd, and Zn. Suspended solids
concentrations have been reduced from an average 34.7 mg/1 to
less than 11.3 mg/1. Zinc removals from 0.2 mg/1 (influent) to
0.05 mg/1 (effluent) and iron removals of 0.2 mg/1 (influent) to
0.09 mg/1 (effluent) have also been achieved.
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A pilot-scale study of mine drainage treatment in Canada has
demonstrated the effectiveness of filtration. Pre-filtration
treatment consisted of lime precipitation, flocculation, and
clarification. Polishing of the clarifier overflow by sand
filtration further reduced the concentration of lead (extract-
able) from 0.25 mg/1 to 0.12 mg/1, zinc from 0.37 mg/1 to 0.19
mg/1, copper from 0.05 mg/1 to 0.04 mg/1, and iron from 0.23 mg/1
to 0.17 mg/1. For further discussion of this subject, refer to
the discussion of Pilot- and Bench-Scale Treatment Studies, later
in this section.
Also, slow sand filters are used on a full-scale basis at Mine/
Mill 1131 to further polish tailing pond effluent prior to final
discharge.
Recovery of metal values contained in suspended solids may, in
some cases, offset the capital and operating expenses of filter
systems. For example, filtration is used to treat uranium mill
tailings for value recovery through countercurrent washing. In
this instance, the final washed tail filter cake is reslurried
for transport to the tailing pond.
Adsorption
Adsorption on solids, particularly activated carbon, has become a
widely used operation for purification of water and wastewater.
Adsorption involves the interphase accumulation or concentration
of substances at a surface or interface. Adsorption from
solution onto a solid occurs as the result of one of two
characteristic properties for a given solvent/solute/solid
system. One of these is the lyophobic (solvent-disliking)
character of the solute relative to the particular solvent. For
example, the more hydrophilic a substance is, the less likely it
is to be adsorbed, and the reverse is true.
A second characteristic property of adsorption results from the
specific affinity of the solute for the solid. This affinity may
be either physical (resulting from van der Waal's forces) or
chemical (resulting from electrostatic attraction or chemical
interaction) in nature.
The best known and most widely employed adsorbent at present is
activated carbon. The fact that activated carbon has an
extremely large surface area per unit of weight (on the order of
1,000 square meters per gram) makes it an extremely efficient
adsorptive material. The activation of carbon in its manufacture
produces many pores within the particles, and it is the vast area
of the walls within these pores that accounts for most of the
total surface area of the carbon. In addition, due to the
presence of carboxylic, carbonyl, and hydroxyl group residuals
fixed on its surfaces, activated carbon also can exhibit limited
ion exchange capabilities.
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Granular activated carbon is generally preferred to the powdered
form, due to dust, and handling problems which accompany the
latter. The commercial availability of a high activity, hard,
dense, granular activated carbon made from coal, plus the
development of multiple-hearth furnaces for on-site regeneration
of this type of carbon, have drastically reduced the cost of
granular activated carbon for wastewater treatment. Although
powdered carbon is less expensive, it can only be used on a once-
through basis and, subsequently, must be removed from the waste
stream in some manner (e.g., filtration or settling).
A number of carbon-contacting system designs have been employed
in other industries. Basic configurations include upflow or
downflow, by gravity or pump pressure, with fixed or moving beds,
and single (parallel) or multi-stage (series) unit arrangements.
The most important design parameter is contact time. Therefore,
the factors which are critical to optimum performance are flow
rate and bed depth. These factors, in turn, must be determined
from the rate of adsorption of impurities from the wastewater.
Activated carbon presently finds application in purification of
drinking water and treatment of domestic, petroleum-refining,
petrochemicals, and organic chemical wastewater streams. Com-
pounds which are readily removed by activated carbon include
aromatics, phenolics, chlorinated hydrocarbons, surfactants,
organic dyes, organic acids, higher molecular weight alcohols,
and amines. This technology also removes color, taste, and odor
components in water. In addition, the potential of activated
carbon to adsorb selected metals has been evaluated on both pure
solutions and wastewater streams. The removal efficiencies range
from slight to very high, depending on the individual metals,
example of metals removal by activated carbon is presented in the
tabulation that follows. This list is a summary of removal
capabilities observed at three automobile wash establishments
employing carbon adsorption for wastewater reclamation.
Metal Concentration (mg/1-total)
Initial Final
Cd 0.015 to 0.034 less than 0.005
Cr 0.01 to 0.125 less than 0.01
Cu 0.04 to 0.15 less than 0.01 to 0.02
Ni 0.045 to 0.16 less than 0.01 to 0.04
Pb 0.32 to 1.32 less than 0.02
Zn 0,382 to 1.49 0.02 to 0.417
In general, the literature indicates significant quantities of
from wastewater by activated carbon. Removal of Cu, Cd, and Zn
appears to be highly variable and dependent upon wastewater
characteristics, while metals such as Ba, Se, Mo, Mn, and W are
reported to be only poorly removed by activated carbon. The
removal mechanism is thought tc involve both adsorption and
filtration within the carbon bed.
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In addition to metals, other waste parameters of major interest
in the ore mining and dressing industry are cyanide and
phenolics. The use of granular carbonaceous material to catalyze
the oxidation of cyanide to cyanate by molecular oxygen has
been demonstrated (References 10, 11, and 12). The efficiency of
cyanide destruction in this manner is reportedly improved when
the cyanide is present as a copper cyanide complex (Reference
10). Application of this technology to treatment of copper-
plating waste having an initial cyanide concentration of 0.315 to
4.0 mg/1 has resulted in a final effluent concentration of 0.003
to 0.011 mg/1. Flow rate through the carbon bed was found to be
0.45 l/sec/m3 (0.2 gpm/ft3).
Phenolics have also been demonstrated to be readily removed by
activated carbon in many industrial applications. However,
little information is available relative to removal of phenolics
at concentrations characteristic of milling wastewater (i.e.,
less than 3 mg/1).
Cyanide Treatment
Depressing agents are commonly used in the flotation of metal
ores to assist in the separation of minerals with similar float-
abilities. As discussed previously, cyanide, either as calcium
cyanide flake or as sodium cyanide solution, is widely used as a
depressant for iron sulfides, arsenopyrite, and sphalerite during
flotation of base metals, and ferroalloys. Cyanide is also used
in processing lode gold and silver ores by the cyanidation pro-
cess, a leaching process.
The use of cyanide in these milling processes results in its
presence in mill tailings and wastewater. The maximum
theoretical concentration of total cyanide in untreated mill
wastewater, based on reported reagent consumption and water use,
is approximately 1.3 mg/1 for flotation operations and 114 mg/1
for gold cyanidation operations. In practice, however, cyanide
levels below the theoretical maximum are observed. (Refer to
Section VI, Wastewater Characteristics.)
An additional source of cyanide-bearing wastewater is underground
mines which backfill stopes with the sand fraction of mill
tailings. Residual cyanide is found in tailings from flotation
circuits using sodium cyanide as a depressant. At least two
lead/ zinc facilities cyclone these tailings to separate the
heavy sand fraction from slimes and then sluice the sands to
backfill mined-out stopes. Overflow from the backfilled stopes
introduces cyanide to the mine drainage.
The dissociation of simple cyanide salts in water and the
subsequent hydrolysis of the cyanide ion leads to the formation
of hydrocyanic acid (HCN). The relative amounts of free cyanide
ion to HCN are dependent on pH. For example, at pH 7, the ratio
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of cyanide (CN-) to HCN is 0.005 to 1; at pH 11, the ratio of CN-
to HCN is 50 to 1.
In addition to the presence of free cyanide ion and hydrocyanic
acid, it has been suggested that the predominant cyanide species
found in flotation mill wastewater are metal-cyanide complexes.
Willis and Woodcock (References 13 and 14) have demonstrated the
presence of copper-cyanide complexes in flotation circuits, with
cupro-cyanides (Cu(DN)3~2 and/or CuCN being the predominant
complexes formed.
Although only the presence of copper-cyanide complexes in
flotation circuits has been shown, the presence of other
transition metals in the float circuit may present situations
favorable to the formation of additional metal-cyanide complexes.
These additional complexes include zinc cyanides (Zn(CN)4~2
and/or Zn(CN?), and iron-cyanides (Fe(CN)6~2 and/or Fe(CN)6~3).
Indirect evidence for their existence has been presented by two
domestic mills. Both operations have inferred the existence of
iron cyanide complexes in mill tailings based on the presence of
residual cyanide in the effluents from laboratory and pilot-plant
treatability studies (Reference 15 and 16).
Three options available to eliminate cyanide from mill effluents
are: (1) in-process control, (2) use of alternative depressants,
and (3) treatment. The particular option or combination depends
on process type, existing controls, the availability and
applicability of alternatives, plant economics, and personal
preference of the plant operator. In-process control and
alternative reagent use have been discussed; treatment is
discussed here.
Sophisticated technology for the destruction of cyanide is not
employed at most domestic mine/mill operations which use cyanide.
Such technology is generally not necessary because in-process
controls and retention of mill tailings in tailing ponds have
reduced cyanide concentrations to less than detectable levels in
the final effluents. The mechanism of cyanide decomposition
within a tailing pond is thought to involve photo-decomposition
by ultraviolet light (Reference 17) and biochemical oxidation.
For this reason, elevated levels of cyanide in the final effluent
(tailing-pond decant) are some times observed during winter
months, when daylight hours are at a minimum and ice sometimes
covers the tailing pond.
Because of increasingly stringent regulation of cyanide in
industrial wastewater discharges during recent years, a number of
domestic and foreign mine/mill operations have investigated and
implemented sophisticated technology for cyanide destruction.
Treatment technologies which have been investigated and/or
employed by various industries for the destruction of cyanide are
listed below:
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1. Chemical oxidation
- Alkaline chlorination (calcium, sodium, or magnesium
hypochlorite)
- Gaseous chlorine
- Permanganate
- Ozone
- Hydrogen or sodium peroxide
2. Electrolysis
3. Biological Degradation
4. Carbon-Bed Oxidation
5. Destruction by Gamma Irradiation
6. Physical Treatment
- Ion exchange
- Reverse osmosis
7. Ferrocyanide Precipitation
Of the technologies listed above, alkaline chlorination, hydrogen
peroxide, and ozonation appear to be best suited for use in the
ore mining and dressing industry. They most readily lend
themselves to the treatment of high volume, relatively low
concentration waste streams at reasonable cost. Free cyanide and
cadmium, copper, and zinc-cyanide complexes can be destroyed by
these treatment technologies. However, it is uncertain in the
ore industry whether cyanide complexes (such as nickel cyanide
and iron cyanide) are attacked or destroyed by chlorine or ozone.
Thus, the effectiveness of these technologies is dependent on the
specific nature of the wastewater treated.
Alkaline Chlorination Theory. The kinetics and mechanisms of
cyanide destruction have been described in the literature
(References 18 through 23). Destruction is accomplished by
oxidation of free cyanide (CN~) to cyanate (CNO~) and,
ultimately, to C02_ and N2_. Destruction of metal-cyanide
complexes (e.g., CuCN) is accomplished by oxidation of the
complex anion to form the metal cation and free cyanide. The
probable reactions in the presence of excess chlorine are:
C12 + CM- + 2NaOH > CNO- + H20 + 2NaCl
and
3C12 + 2CuCN + BNaOH > 2NaCNO + 2Cu(OH)2 + 6NaCl
+ 2H20
Rapid chlorination at a pH above 10 and a minimum of 15-minutes
contact time are required to oxidize 0.45 kilogram (1 pound) of
cyanide to cyanate with 2.72 kilograms (6 pounds) each of sodium
hydroxide (caustic soda) and chlorine. If metal-cyanide
complexes are present, longer detention periods may be necessary.
An alternative chlorination technique involves the use of sodium
hypochlorite (NaOCl) as the oxidant. Reactions with sodium
hypochlorite are similar to those of chlorine except that there
is no caustic requirement for destruction of free cyanide in the
oxidation stages. However, alkali is required to precipitate
236
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metal- cyanide complexes as hydroxides. Reactions of the free
cyanide and the metal-cyanide complex with hypochlorite are:
NaOCl + CN- CND- + NaCl
and
3NaOCl + 2CuCN + 2NaOH + H20 2NaCNO + 2Cu(OK)2 +
3NaCl
To oxidize cyanide to cyanate, a 15 percent solution of sodium
hypochlorite is required at a dosage rate ranging from 2.72
kilograms (6 pounds) to 13.5 kilograms (30 pounds) of sodium
hypochlorite per 0.45 kilogram (1 pound) of cyanide is required
to oxidize cyanide.
Complex destruction of cyanate requires a second oxidation stage
with an approximate 45~minute retention time at a pH below 8.5.
The theoretical reagent requirements for this second stage are
1.84 kilograms (4.09 pounds) of chlorine and 0.51 kilogram (1.125
pounds) of caustic per 0.45 kilogram (1 pound) of cyanide. Actual
reagent consumption and choice of reagent will be dependent on
process efficiency, residual chlorine levels from the first
oxidation stage, optimization through pilot-scale testing,
temperature, etc. The overall reaction for the second stage is:
3C12 + 2CNO- + 6 NaOH > 2HC03~ + N2 + 6NaCL + 2H20
Note that the intermediate reaction product, carbon dioxide,
reacts with alkalinity in the water to form bicarbonate.
Advantages to the use of alkaline chlorination include relatively
low reagent costs, applicability of automatic process control,
and experience in its use in other industries (e.g.,
electroplating). Major disadvantages are the potential health
and pollution hazards associated with its use, such as worker
exposure to chlorine gas (if gas is used) and cyarogen chloride
(byproduct gas), the potential for production of harmful
chloramines and chlorinated hydrocarbons, and the presence of
high chlorine residual levels in the treated effluent.
Ozonation Theory. Because of the disadvantages associated with
alkaline chlorination, ozonation is receiving a great deal of
attention as a substitute technique for cyanide destruction.
Oxidation of cyanide to cyanate with ozone requires approximately
0.9 kilogram (2 pounds) of ozone per 0.45 kilogram (1 pound) of
cyanide, and complete oxidation requires 2.25 kilograms (5
pounds) of ozone per 0.45 kilogram (1 pound) of cyanide. Cyanide
oxidation to cyanate is very rapid (10 to 15 minutes) at pH 9 to
12 and practically instantaneous in the presence of trace amounts
of copper. Thus, the destruction of cyanide to cyanate in mill
wastewater containing copper cyanide complexes can be expected to
proceed rapidly.
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The reaction mechanism for the destruction of cyanide to cyanate
is generally expressed as:
CN- -i- 03 > CNO- + 02
The reaction mechanism for the subsequent reaction, destruction
of cyanate, has not been positively identified. Proposed
mechanisms include.(Reference 24 and 25):
2CNO- + 03 + H20 > 2HC03 + N2 + 302
CNO- + OH- + H20 > C03 -2 + NH3
or
CNO- + NH3 > NH2~ CO- NH2
Regardless of the actual mechanism, destruction of cyanate can be
accomplished in approximately 30 minutes (Reference 26).
Hydrogen Peroxide Theory. Two processes for the oxidation of
cyanide with hydrogen peroxide (H2_02_) have been investigated on a
limited scale. The first process involves the reaction of
hydrogen peroxide with cyanide at alkaline pH in the presence of
a copper catalyst. The following reactions are observed:
CN- + H202 > CNO- + H20
CNO- + 2H202 > NH4 + C03~2
The second process, known as the Kastone process, uses a
formulation containing 41 percent hydrogen peroxide, trace
amounts of catalyst and stabilizers and formaldehyde. The
cyanide wastes are heated to 129 C (248 F), treated with three to
four parts of oxidizing solution and two to three parts of a 37
percent solution of formaldehyde per part of sodium cyanide, and
agitated for one hour. Principal products from the reaction are
cyanates, ammonia, and glycolic acid amide. Complete destruction
of cyanates requires acid hydrolysis (Reference 23).
Cyanide Treatment Practices. As discussed, treatment
specifically designed for cyanide treatment has not been widely
installed in this industry. However, investigations of treatment
techniques specific to cyanide reduction have been conducted.
Extensive laboratory and pilot-plant tests on cyanide destruction
in mill wastewater were conducted by molybendum Mill 6102 for the
development of a full scale waste treatment system, which was
brought on-line during July 1978. Testing was aimed at
identifying treatment to achieve a 1 July 1977 permit limitation
of 0.025 mg/1.
Ozonation tests in the laboratory showed substantial destruction
of cyanide, as the data in Table VIII-2 show. The target level
of less than 0.025 mg/1 of cyanide was not achieved, however, and
tests of ozonation under pilot-plant conditions showed even less
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favorable results. Ozonation did, however, result in substantial
removals of cyanide and manganese.
Laboratory chlorination tests (also at Mill 6102) indicated that
removal of cyanide to 0.025 mg/1 or less could be achieved under
the proper conditions. As the data in Table VIII-3 show,
chlorine doses in excess of stoichiometric amounts were required,
and pH was found to be a major determinant of treatment
effectiveness. Results on a pilot-plant scale were less
favorable, but improved performance in the full-scale treatment
system is anticipated through use of a retention basin in which
additional oxidation of cyanide by residual chlorine can take
place.
Mill 6102 has built a treatment system employing lime precipi-
tation, electrocoagulation-flotation, ion exchange, alkaline
chlorination, and mixed media filtration. This is followed by
final pH adjustment. The alkaline chlorination system includes
on-site generation of sodium hypochlorite by electrolysis of
sodium chloride. The hypochlorite is injected into the waste-
water following the electrocoagulation-flotation process and
immediately preceeding the filtration unit. At this point in the
system, some cyanide removal has been realized incidental to the
lime precipitation-electrocoa,gulation treatment. The first four
months of operating data show the concentration of cyanide at
0.09 mg/1 prior to the electrocoagulation unit. Concentrations
of cyanide progressively decreased from a 0.04 mg/1 (electrocoa-
gulation effluent) to less than or equal to 0.01 mg/1 after fil-
tration, and less than 0.01 mg/1 after the final retention pond.
Mill personnel expect this removal efficiency to continue
throughout the optimization period of the system. The problem of
chlorine residuals at elevated levels has not been resolved.
Control and treatment of cyanide at a Canadian lead/zinc
operation, Mill 3144, is achieved by segregation of the cyanide
bearing waste streams and subsequent destruction of the cyanide
by alkaline chlorination. Waste segregation is practiced to
reduce the volume and solids loading of the cyanide-bearing waste
streams. The waste stream treated in the alkaline chlorination
treatment plant comprises about 30 percent of the total tailings
volume ultimately discharged, or approximately 1,000 cubic meters
(300,000 gallons) of tailings per day. This waste stream has an
initial : yani.de concentration of approximately 60 to 70 mg/1.
If;..- initial design of the alkaline chlorination treatment plant
was based on extensive laboratory and pilot-plant studies. Final
design modifications, made in 1975, were based on a requirement
to comply with an amended discharge permit limitation of 0.5 mg/1
cyanide (total). Because some cyanide is present in the 3,000
cubic meters (700,000 gallons) per day of tailings not treated,
the alkaline chlorination treatment plant is operated with the
goal of destroying the cyanide contained in that portion of
-------
wastewater being treated. The composite waste stream should then
meet the permit limitation.
The treatment plant includes three FRP (fiber-reinforced-plastic)
tanks, measuring 3 by 3 meters (10 by 10 feet), which operate in
series, as reaction chambers, Chlorine is added at a rate of 540
to 680 kilograms (1,200 to 1,500 pounds) per day from a
chlorinator having a capacity of 900 kilograms (2,000 pounds) per
day of chlorine. The pH of the combined waste stream is
maintained between 11 and 12 by lime addition to the incoming
waste stream. Process controls include pH and ORP
(oxidation/reduction-potential) recorders and a magnetic flow
meter.
The average cyanide concentration of the total tailing effluent
discharge between July and December 1975 was 0.18 mg/1 of cyanide
(total). This compares to the average concentration of 4.72 mg/1
of cyanide (total) in the discharge prior to installation of the
treatment plant. Performance data for the alkaline-chlorination
treatment plant at Mill 3144 are presented in the following table
(industry data):
Source Total Cyanide (mg/1)*
Chlorination-Plant Feed 68.3
Chlorination-Plant Discharge 0.13
Mine Drainage (Overflow from
Backfill) 0.06
Total Combined Tailings 0.07
*Average of daily samples taken during September 1975.
Government regulations in the USSR presently limit the discharge
of cyanide waste from ore milling operations to 0.1 mg/1 of
cyanide. A more stringent limitation of 0.05 mg/1 applies when
cyanide-bearing wastewater is discharged to surface waters
inhabited by fish (Reference 27).
Two references in the literature described the use of alkaline
chlorination in the Soviet Union for treatment of cyanide in ore-
milling wastewater (References 28 and 29). At one mill, the
effluent, containing 95 mg/1 of cyanide ion, was treated with
chlorine at a dosage rate of approximately 10 parts chlorine to 1
part cyanide. It was claimed that the cyanide was completely
destroyed. A second mill treated the overflow from copper and
lead thickeners with calcium hypochlorite. The overflow con-
tained more than 45 mg/1 of cyanide ion. The cyanide concentra-
tion of the treated effluent was reported to be less than 1 mg/1,
although difficulties were experienced in the oxidation stage.
Similar difficulties with the use of calcium hypochlorite have
been reported elsewhere (Reference 17).
Research conducted by the Air Force Weapons Laboratory on the
destruction of iron cyanide complexes has resulted in development
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of a pilot-scale process to treat electroplating and photographic
processing wastes. Briefly, the process employs ozonation at
elevated temperatures and ultraviolet irradiation to reduce cya-
nide concentrations in the effluent to below detectable levels
(Reference 30} .
Reduction of cyanide in tailing pond decant water using hydrogen
peroxide has been practiced on an experimental basis at Mill
6101. Although earlier monitoring data had shown cyanide to be
reliably absent from the effluent (less than 0.02 mg/1), recent
data using EPA approved analytical procedures indicated that,
during the colder months, elevated levels of cyanide (up to
approximately 0.09 mg/1) may occur. To reduce these levels, mill
6101 has experimented with a very simple peroxide treatment sys-
tem with modest success.
Treatment is provided by dripping a hydrogen peroxide solution
from a drum into the channel carrying wastewater from the tailing
pond to the secondary settling pond. Mixing is by natural turbu-
lence in the channel, and the peroxide addition rate is manually
adjusted periodically based on the effluent cyanide concentra-
tion. Results to date indicate that this simple treatment system
achieves cyanide removals on the order of 40 percent.
Treatment of Phenols
Several phenolic compounds are used as reagents in floation
mills. These compounds, which include Reco, AEROFLOAT 31 and
242, AERO Depressant 633, cresylic acid, and diphenyl guanidine,
find specific application as promoters and frothers in the flota-
tion process.
Phenol-based compounds can be a significant source of phenolics
in mill process wastewater. The degree of control exerted over
reagent addition, uniformity in ore grade, and mill operator
preferences could affect residual phenol levels in the flotation
circuit. The surface-active nature of frothers and promoters,
coupled with the volatility of many phenolic compounds, further
compli cat ;;; the theoretical prediction of phenol concentrations
in flot"' ion-mill wastewater streams.
A n-^re dev Tiled account of reagent use, including phenolic-based
x-o^pounds and their presence in mill process wastewater, is
provided in Section VI, Wastewater Characteristics.
Several methods are available for treating phenolic wastes,
including:
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1. Chemical Oxidation
Chlorine dioxide
Hydrogen peroxide
Ozone
Potassium permanganate
2. Biological Oxidation
3. Carbon Adsorption
4. Aeration
5. Ultraviolet Irradiation
6. Incineration
7, Recovery
The specific technology applied depends on the chemical
characteristics of the waste stream, the discharge concentrations
required, and the economics of implementation. However, the low-
concentration, high-volume phenolic wastes generated in this
industry are best treated by chemical oxidation or aeration.
Chemical Oxidation Theory. Chemical oxidizing agents react with
the aromatic ring of phenol and phenolic derivatives, resulting
in its cleavage. This cleavage produces a new organic compound
(a straight-chain compound), which still exerts a chemical oxygen
demand (COD). Complete destruction of the organic compound
(conversion to C02^ and H20) and reduction in COD requires either
additional chemical oxidation or other treatment (e.g.,
biological oxidation).
Complex wastewater may require an additional oxidizing agent. As
an example, hydrogen peroxide will react with sulfides,
mercaptans, and amines in addition to phenolic compounds. The
total consumption of oxidizing agent is dependent on the type
and concentration of oxidizable species present in the waste, the
reaction kinetics, and the end products desired (i.e., straight-
chain organic compounds or carbon dioxide and water). Therefore,
it is difficult to predict actual reagent consumption without
treatability studies. Some general guidelines may be given,
however, for the various oxidizing agents available.
Chlorine Dioxide. Chlorine gas is considered unacceptable as an
oxidizing agent because of the potential for forming toxic
chlorinated phenols, as well as the potential safety hazards
involved in handling the gas. As an alternative, chlorine
dioxide (C102J can be generated on-site from chlorine gas or
hypochlorite and used as a relatively safe oxidizing agent.
Chlorine dioxide reacts with phenol to form benzoquinone (C6H4_0£)
within the pH range of 7 to 8 and at a reagent dosage of 1.5
parts of C102^ per part of phenol. At a pH above 10 and a dosage
of 3.3 parts of C102^ per part of phenol, maleic acid and oxalic
acid are formed, rather than benzoquinone.
Hydrogen Peroxide. In the presence of a metal catalyst (e.g.,
Fe++, Fe+ + + , A1 + + + , Cu + + , and Cr + +), hydrogen peroxide (H2_020
effectively oxidizes phenols over a wide range of temperatures
242
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and concentrations. Investigations into the use of hydrogen
peroxide show phenol removal efficiencies of 98+ percent at a
dosage rate of one to two parts of H2_02_ per part phenol and at an
optimum pH between 3 and 5. Wastewater containing substituted
phenols, such as cresylic acid, can increase the required perox-
ide dosage to four parts of H202_ per part of substituted phenol.
An approximate five-minute retention time is required to
partially oxidize simple phenols. Either batch or continuous
operation may be employed, with batch treatment preferred at
flows less than 190 to 380 cubic meters (50,000 to 100,000
gallons) per day (Reference 31).
Ozone. Ozone, a very strong oxidizing agent, attacks a variety
of materials, including phenols. Because of its poor
selectivity, ozonation is generally used as a polishing step
after conventional treatment processes which remove gross
suspended solids and nontoxic organic compounds.
Ozone will completely oxidize phenols to carbon dioxide and water
if a sufficient retention time and enough ozone are provided. In
practice, however, the reaction is allowed to proceed only to
intermediate, straight-chain organic compounds. Ozone
requirements for the partial destruction of phenols range from
one to five parts per part of phenol. The actual ozone demand is
a function of phenol concentration, pH, and retention time. For
example, reduction of phenol in a particular wastewater from
2,500 mg/1 to 25 mg/1, at a pH of 11, has been found to require
1.7 parts of ozone per part of phenol and a 60-minute retention
time. The same waste, when treated at a pH of 8.1, required 5.3
parts of ozone per part of phenol and a 200-minute retention time
to achieve similar reductions in phenol (Reference 63). The
efficiency of ozonation appears to increase with decreasing
phenol concentration. Operating data from a full-scale ozonation
system treating 1,500 cubic meters (400,000 gallons) per day of
wastewater from a Canadian refinery show ozone requirements of
one part per part of phenol to reduce phenol reductions to 0.003
mg/1 at dosage rates ranging from 1.5 to 2.5 parts of ozone per
part of phenol (References 31 and 32).
Potassium Permanganate. Paint-stripping and foundry wastes
containing 60 to 100 mg/1 of phenol have been treated with
potassium permanganate (KMnO£). Phenol reductions to less than 1
mg/1 are reported. The destruction of simple phenol by potassium
permanganate is expressed as:
3C6H60 + 28KMn04 + 5H20 18C02. + 28KOH + 28Mn02
Based on this expression, the theoretical dosage for complete
destruction of phenol is 15.7 parts of KMn04_ per part of phenol.
However, dosages of only six to seven parts of KMn04 per part of
phenol have been reported to be effective in destroying the
aromatic-ring structure. Optimum pH for permanganate destruction
of phenol is between seven and ten (References 31 and 33).
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A disadvantage to the use of potassium permanganate is the
generation of a manganese dioxide precipitate, which settles as a
hydrous sludge. Clarification and sludge disposal may be
required, resulting in additional equipment and maintenance
costs.
Aeration Theory. Limited phenol removal may be obtained in ponds
or lagoons by simple aeration. In general, forced aeration is
more effective in reducing phenol levels than is passive
aeration. Field studies have shown that, at an initial
concentration of 15 mg/1 of phenol, wastewater phenol levels can
be reduced to approximately 1 mg/1 after 30 hours of forced
aeration and after 70 hours of passive aeration (Reference 31},,
The mechanisms for removal of phenols in ponds is not well
understood, but probably includes degradation by biological
action and ultraviolet light, and simple air stripping.
£l]^Q,2l. Treatment Pracj:icejs. The only treatment for phenols used
in the ore mining and dressing industry is aeration. In practice,
phenol reduction is incidental to treatment of more traditional
design parameters (i.e, heavy metals, suspended solids, etc.).
At Mill 2120, phenol concentrations in the tailing-pond influent
and effluent average 0.031 mg/1 and 0.021 mg/1, respectively.
Similar results are noted at Mill 2122, where phenol
concentrations in the tailing-pond influent and effluent average
0.26 mg/1 and 0.25 mg/1, respectively. Data from samples
collected at Mill 2117 show phenol reductions from 5.1 mg/1 of
phenol in the raw tailings to 0.25 mg/1 of phenol in-the tailing-
pond overflow.
Sulfide Precipitation of Metals
The use of sulfide ions as a precipitant for removal of heavy
metals can accomplish more complete removal than hydroxide pre-
cipitation. Sulfide precipitation is widely used in wastewater
treatment in the inorganic chemicals industry for the removal of
heavy metals, especially mercury. Effective removal of cadmium,
copper, cobalt, iron, mercury, manganese, nickel, lead, zinc, and
other metals from mine and mill wastes show promise by treatment
with either sodium sulfide or hydrogen sulfide. The use of this
method depends somewhat on the the availability of methods for
effectively removing precipitated solids from the waste stream,
and on removal of the solids to an environment where reoxidation
is unlikely.
Several steps enter into the process of sulfide precipitation:
1. Preparation of sodium sulfide. Although this product is
often in abundence as a byproduct it can also be made by the
reduction of sodium sulfate, a waste product of acid-leach
milling. The process involves an energy loss in the partial
oxidation of carbon (such as that contained in coal) as
follows:
?44
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Na2S04 + 4C Na2S + 4CO (gas)
2. Precipitation of the pollutant metal (M) in the waste
stream by an excess of sodium sulfide:
Na2S + MS04 MS (precipitate) + Na2SO4_
3. Physical separation of the metal sulfide in thickeners
or clarifiers, with reducing conditions maintained by excess
sulfide ion.
4. Oxidation of excess sulfide by aeration:
Na2S + 202. Na2S04
This process usually involves iron as an intermediary and, as
illustrated, regenerates unused sodium sulfide to sodium sulfate.
In practice, sulfide precipitation can be best applied when the
pH is sufficiently high (greater than about eight) to assure
generation of sulfide, rather than bisulfide ion or hydrogen
sulfide gas. It is then possible to add just enough sulfide, in
the form of sodium sulfide, to precipitate the heavy metals
present as cations. Alternatively, the process can be continued
until dissolved oxygen in the effluent is reduced to sulfate and
anaerobic conditions are obtained. Under these conditions, some
reduction and precipitation of molybdates, uranates, chromates,
and vanadates may occur; however, ion exchange may be more
appropriate for the removal of these anions.
Because of the toxicity of both the sulfide ion and hydrogen
sulfide gas, the use of sulfide precipitation may require both
pre- and post-treatment and close control of reagent additions.
Pretreatment involves raising the pH of the waste stream to
minimize evolution of H2J5, which could cause odors and pose a
safety hazard to personnel. This may be accomplished at
essentially the same point as the sulfide treatment, or by
addition of a solution containing both sodium sulfide and a
strong base (such as caustic soda). The sulfides of many heavy
metals, such as copper and mercury, are sufficiently insoluble to
allow essentially complete removal with low residual sulfide
levels. Treatment for these metals with close control on sulfide
concentrations could be accomplished without the need for
additional treatment. Adequate aeration should be provided to
yield an effluent saturated with oxygen.
Sulfide precipitaition is presently practiced at most mercury-
cell chloralkali plants to control mercury discharges. In this
application, treatment with sodium sulfide is commonly followed
by filtration and typically results in the reduction of mercury
concentrations from 5 to 10 mg/1 to 0.01 to 0.05 mg/'l. Although
lead is also present in the waste streams treated with sulfide,
its concentration is not often measured. The limited available
245
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data indicate that lead is also removed effectively by sulfide
precipitation (Reference 34).
Sulfide is also used in the treatment of chemical-industry waste
streams bearing high levels of chromates. It is reported that,
after sulfide treatment and sedimentation, levels of hexavalent
chromium consistently below 1 m.icrogram per liter and total
chromium between 0.5 and 5 micrograms per liter are achieved in
the effluent (Reference 35). Other sources report that sulfide
precipitation can achieve effluent levels of 0.05 mg/1 arsenic,
0.008 mg/1 cadmium, 0.05 mg/1 selenium, and "complete" removal of
zinc (References 35, 36, 37, and 38).
Sulfide precipitation is not employed in the domestic metal-ore
mining and milling industry at present. However, the use of
sulfide for removal of copper, zinc, and manganese from acid-mine
drainage has been evaluated both theoretically and experimentally
(References 35 and 39). A field study of mine drainage treatment
in Colorado demonstrated that greater than 99.8 percent removal
of metals was attained by treatment which consisted of partial
neutralization with lime, followed by sulfide addition and
settling. The treated effluent attained in this manner contained
0.2 mg/1 zinc, 0.4 mg/1 manganese, and less than detectable
concentrations of copper and arsenic. However, it was also noted
that the standard neutralization with lime and settling produced
similar results.
Asbestos Treatment
The term "asbestos" has many definitions (Reference 40). The
EPA's Effluent Guidelines Division chose to define asbestos as
chrysotile for the purpose of this program. (For the rationale
for this choice refer to page nine of Reference 40.)
The main source of asbestos fibers in this industry is the
milling and beneficiation of copper, iron, nickel, molybdenum and
zinc ores. The total-fiber counts made from samples collected at
Data on asbestos fiber removal in this industry is very limited.
However, the physical treatment processes used and consistent
fiber morphology make data from municipal and other industries
applicable. Two good data sources are the Duluth, Minnesota
potable water treatment plant and the chlorine/caustic
(chloralkalai) industry.
Duluth Treatment Plant. Extensive pilot-scale studies on removal
of asbestiform minerals from Lake Superior water were conducted
by Black and Veatch Consulting Engineers under joint sponsorship
of the EPA and the Army Corps of Engineers (References 41, 42,
and 43). Filtration processes investigated included filtration,
pressure diatornaceous earth (DE) filtration, and vacuum DE
filtration.
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Forms of asbestiform minerals evaluated in these studies include
amphibole and chrysotile. A basic difference between these forms
which influences treatment is that the treatment for the
amphibole form usually carries important conclusions and
recommendations from the pilot-scale studies on granular
filtration, including:
1. Granular filtration is successful in removal of asbesti-
form fibers.
2. Sedimentation prior to filtration increases filter run
length (i.e., time between backwashes) but does not increase
fiber removal. (Note that untreated water for these studies
was Lake Superior water, clean water relative to mine/mill
wastewater. Sedimentation does remove some asbestos fiber
in this industry and would be necessary to prevent filter
clogging and frequent filter backwashing.)
3. Two-stage flash-mixing followed by flocculation is
recommended for conditioning raw water prior to filtration.
4. Alum is a more effective coagulant than ferric
for Duluth raw water.
chloride
5. Nonionic polyelectrolyte is most effective in preventing
turbidity breakthrough.
6. A positive-lead mixed media filter designed to operate
7. Backwash water should be discharged to a sludge lagoon,
and supernatant should be returned to the treatment plant.
8. For large capacity plants, granular filtration is
recommended over DE filtration. For small plants, DE
filtration should also be considered.
Two kinds of DE filtration processes were studied in the pilot-
scale studies, pressure filtration and vacuum filtration. Flow
through both filtration systems ranged from 0.0006 to 0.001
m3/sec (10 to 20 gal min). Both kinds of DE filters were oper-
ated in various ways to evaluate conditioning of DE with alum,
cationic polymers, and anionic polymers. Single-step and twostep
precoat were studied. Conditioned DE was used in filter precoat,
as well as for body feed. Various grades of DE, from fine to
coarse, were evaluated. Details of pilot-plant DE testing are
reported in References 41, 42, and 43,
Vacuum DE filtration was found unsuitable for treating the raw
Lake Superior water being tested because of the formation of air
bubbles in the filter during filtration of cold water. (This is
not to say that vacuum filtration would not be effective on
warmer waters.) For the conditions experienced during the pilot-
scale studies on potable Lake Superior water, it was concluded
247
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that pressure DE filtration is effective (i.e., reduces fiber
content to 4 x 10* fibers/liter or less) with the addition of
chemical aids to the precoat, the body feed, and the raw water.
Conditioning steps found to effect high removal of fibers con-
sisted of the following:
1. Alum coatings or plain precoat with cationic polymer
added to raw water
2. Anionic polymer added to precoat and alum-coated body
feed
3. Filter precoat with medium-grade DE
4. Conditioning of DE with alum or soda ash, or with
anionic polymer
5. Fine grade of DE for body feed
6. DE filter flow rate of approximately 41 liters/min/m2
(1 gal/min/ft2).
Pilot plant studies have been conducted on asbestos removal using
synthetic asbestos suspensions containing approximately the same
concentrations and size distributions as typical asbestos mine
effluents (i.e., about 1012 fibers/liter) (Reference 44). In
addition, pilot plant tests have been conducted on samples of
asbestos-laden water from three locations in Canada. Treatment
processes studied include plain sedimentation, sand filtration,
mixed media (sand and anthracite) filtration, and DE filtration.
Effectiveness of the various treatment methods at pilot plants
using synthetic asbestiform (chrysotile) particle-laden water is
summarized in Table VIII-4 (Reference 44).
Data on pilot-plant asbestos removal from wastewater at specific
locations are summarized in Tables VIII-5 and VIII-6. The
authors conclude that sedimentation followed by mixed-media
filtration is very effective for removing most of the asbestos
from mine and processing plant effluents. Diatomite filters are
even more effective, but may be more than what is necessary
except where wastewater streams are discharged into water used as
a drinking water source (Reference 44).
Based on the pilot-scale studies discussed previously (Reference
42), a 114,000-mVday (30,000,000-gal/day) granular filtration
plant was designed and constructed for treating the water supply
for the city of Duluth, Minnesota. This plant became operational
in November 1976 (Reference 45). Figure VIII-1 presents a
schematic diagram of the principal portion of this full-scale
plant. The major components of the system are the mixed media
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filtration beds containing anthracite, sand, and ilmenite. Each
of the four filters has a capacity of 28,400 mVday (7,500,000
gal/day) at a design loading rate of 204 liters/min/m2 (5.0
gal/rnin/f t2) .
Prior to filtration, raw Lake Superior water is flocculated and
settled to increase fiber removal efficiency as well as to
provide suspended solids separation. Coagulation facilities are
three rapid-mix chambers. Anionic polymer is added in the first
chamber, alum and caustic soda in the second chamber, and non-
ionic polymer in the third chamber. Flocculation and sedimen-
tation are carried out in adjacent tanks. Effluent from the
sedimentation basin flows directly to the mixed media filters
described previously.
Filters are backwashed when head loss becomes greater than
approximately 2.4 m (8 feet) or when effluent turbidity is
greater than 0.2 JTU (Jackson Turbidity Units). Filter backwash
water is sent to a storage tank and then to a settling lagoon,
where alum and polymer are added to increase solids settling.
Supernatant from the settling basin is sent back through the
treatment system. Settled sludge from the settling basin is
mechanically transferred to a sludge lagoon for further settling.
Sludge is periodically removed from the lagoons and disposed of
in a sanitary landfill. Decant water from the sludge lagoons is
also returned to the treatment influent. The frequent freeze/
thaw cycles experienced in northern Minnesota enhance water
separation from the asbestos laden sludge. This phenomenon may
not occur in other regions.
The Duluth plant is being monitored closely. Unpublished data
fibers/liter by the full-scale mixed media filtration plant.
Chlorine/Caustic Industry. Treatment for the removal of asbestos
from wastewater is practiced at a significant and growing number
of facilities which produce chlorine and caustic soda by
electrolysis in diaphragm cells. Treatment practices used
include both sedimentation and filtration, with flocculants
frequently used to enhance the efficiency of either process. At
the chlorine/caustic facilities, asbestos removal is practiced on
segregated, relatively low-volume waste streams which generally
have high levels of suspended solids, consisting mostly of
asbestos. Due to uncertainties in the analytical procedure,
asbestos concentrations are not generally reported explicitly and
must be inferred from TSS data.
At one chlorine/caustic facility in Michigan (Reference 34), a
pressure leaf filter is used, together with flocculants, to treat
a 1.5 liter/second (25 gal/min) stream to remove (by filtration)
approximately 102 kg (225 Ib) of asbestos per day. Effluent per-
formance of this system, reported by the facility as "no detect-
able asbestos discharge," was verified by sample analysis, which
shows a reduction of asbestos (total fiber) content from a con-
249
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centration of greater than 5 x 109 fibers/liter in the filter
influent to less than detectable (approximately 3 x 105
fibers/liter) in the filtered discharge.
Another facility removes asbestos from an intermittent flow
totaling about 37.8 m3 (10,000 gal)/day by sedimentation in a
concerete sump with a volume of 327 m3 (11,550 ft3) and com-
partments to provide separate surge and settling chambers. With
the addition of flocculants, this system reduces TSS (of which a
significant fraction is asbestos) from about 3,000 mg/1 to 30
mg/1.
Other facilities report the use of sedimentation technology (with
varying degrees of efficiency), or the elimination of asbestos
discharges by wastewater segregation and impoundment. Plans have
been announced for additional use of filtration within the
chlorine/caustic industry.
Review and comparison of pilot-scale results from treatment of
raw water with granular filtration and DE filtration discussed
earlier (Reference 42) indicate that similar results can be
obtained from the two systems. Data in Tables VII1-5 through
VIII-7 (from Reference 44), however, indicates that DE filtration
may be more effective. An economic analysis of both types of
systems has been conducted (Reference 42).
Practices i_n This Industry. No treatment systems in use in this
industry are operated specifically for asbestos removal.
However, asbestos is a suspended solid and, as discussed in
Section VI, correlates very well with TSS in wastewater generated
in this industry. Therefore, asbestos data taken at facilities
designed and operated for TSS removal is indicative of the
industries' current asbestos removal practices.
The sampling program was not designed to establish treatment
efficiencies and samples are grabs and 24 hour-composites taken
over short terms. However, the data obtained generally demon-
strates the effectiveness of asbestos removal by existing
facilities.
Table VIII-8 is a comparison of the total-fiber and chrysotile
asbestos contents of the influent and effluent streams associated
anions, such as Cl~ and S04~z. Anions adsorb along with
treatment systems at the facilities surveyed. The data indicate
better removal of asbestos in mill water than in mine water.
For mill treatment systems consisting primarily of tailing ponds
and settling or polishing ponds, some facilities have
demonstrated reductions of 104 to 105 fibers/liter. At all
milling facilities surveyed, reduction by at least a factor of 10
is realized, but the most common reduction factors range from 103
to 104 fibers/liters. Examination of these treatment systems
indicates several factors in common: high initial suspended
250
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solids loading, effective removal of suspended solids, large
systems or systems with long residence times, and/or the presence
of additional settling or polishing ponds. Comparison of screen
sampling data with verification sampling data at some facilities
suggests that the asbestiform fiber content of the wastewater may
be quite variable from time to time.
Seven mine water treatment systems exhibited two common
characteristics: (1) generally low total-fiber and chrysotile
asbestos counts, and (2) low to no removal of the fibers in their
treatment systems. This can be explained by two factors: first,
fibers tend to be liberated by milling processes, as compared to
mining activities alone; and, second, mine waste streams tend to
have considerably lower suspended-solids values than mill waste
streams. Because of this, there is less opportunity for inter-
action between the fibrous particles and the suspended solids and
their simultaneous removal by subsequent settling.
Ion Exchange
Ion exchange is basically a process for transfer of various ionic
species from a liquid to a fixed media. Ions in the fixed media
are exchanged for soluble ionic species in the wastewater.
Cationic, anionic, and chelating ion exchange media are available
and may be either solid or liquid. Solid ion exchangers are
generally available in granular, membrane, and bead forms (ion
exchange resins) and may be employed in upflow or downflow beds
or column, in agitated baskets, or in cocurrent or countercurrent
flow modes. Liquid ion exchangers are usually employed in equip-
ment similar to that employed in solvent-extraction operations
(pulsed columns, mixed settlers, rotating-disc columns, etc.).
In practice, solid resins are probably more likely candidates for
end-of-pipe wastewater treatment, while either liquid or solid
ion exchangers may, potentially be utilized in internal process
streams.
Individual ion exchange systems do not generally exhibit equal
affinity or capacity for all ionic species (cationic or anionic)
and, then may not be suited for broad-spectrum removal schemes in
wastewater treatment. Their behavior and performance are usually
dependent on pH, temperature, and concentration, and the highest
removal efficiencies are generally observed for polyvalent ions.
In wastewater treatment, some pretreatment or preconditioning of
wastes to reduce suspended solid concentrations and other
parameters is likely to be necessary.
Progress in the development of specific ion exchange resins and
techniques for their application has made the process attractive
for a wide variety of industrial applications in addition to
water softening and deionization. It has been used extensively
in hydrometallurgy, particularly in the uranium industry, and in
wastewater treatment (where it often has the advantage of
allowing recovery of marketable products). This is facilitated
9r i
c _> i
-------
by the requirements for periodic stripping or regeneration of
ionic exchangers.
Disadvantages of using ion exchange in treatment of mining and
milling wastewater are relatively high costs, somewhat limited
resin capacity, and insufficient specificity (especially, in
cationic exchange resins for some applications). Also, regenera-
tion produces a waste, and its subsequent treatment must be
considered.
For recovery of specific ions or groups of ions (e.g., divalent
heavy-metal cations, or metal anions such as molybdate, vanadate,
and chromate), ion exchange is applicable to a much broader range
of solutions. This use is typified by the recovery of uranium
from ore leaching solutions using strongly basic anion exchange
resin. Additional examples are the commercial reclamation of
chromate plating and anodizing solutions, and. the recovery of
copper and zinc from rayon-production wastewaters. Chromate
plating and anodizing wastes have been purified and reclaimed by
ion exchange on a commercial scale for some time, yielding eco-
nomic as well as environmental benefits. In tests, chromate
solutions containing levels in excess of 10 mg/1 chromate,
treated by ion echange at practical resin loading values over a
large number of loading/elution (regeneration) cycles, consis-
tently produced an effluent containing no more than 0,03 mg/1 of
chromate.
High concentrations of ions other than those to be recovered may
interfere with practical removal. Calcium ions, for example, are
generally collected along with the divalent heavy metal cations
of copper, zinc, lead, etc. High calcium ion concentrations.,
therefore, may make ion exchange removal of divalent heavy metal
ions impractical by causing rapid loading of resins and
necessitating unmanageably large resin inventories and/or very
frequent elution steps.
Less difficulty of this type is experienced with aniori exchange.
Available resins have fairly high selectivity against the common
anions, such as Cl~ and S04~2. Anions adsorbed along with
uranium include vanadate, molybdate, ferric sulfate anionic
complexes, chlorate, cobalticyanide, and polythionate anions.
Some solutions containing molybdate prove difficult to elute and
have caused problems.
Ion exchange resin beds may be fouled by particulates,
precipitation within the beds, oils and greases, and biological
growth. Pretreatrnent of water, as discussed earlier, is therefore
commonly required for successful operation. Generally, feed
water is required to be treated by coagulation and filtration
for removal of iron and manganese, CQ2^ H2S, bacteria and algae,
and hardness. Since there is some latitude in selection of the
ions that are exchanged for the contaminants that are removed,
post treatment may or may not be required.
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In many cases, calcium is present in mining and milling waste-
water in appreciably greater concentrations than the heavy metal
cations to be removed. Ion exchange, in those cases is expensive
and little advantage is offered over lime precipitation. For the
removal of anions, however, the relatively high costs of ion
exchange equipment and resins may be offset partially or totally
by the recovery of a marketable product. This has been
demonstrated in the removal of uranium from mine water (Refer to
"Treatability Studies, Uranium/Vanadium Mill 9401," in this
section).
Removal of molybdate ion from ferroalloy ore milling wastewater
has been investigated in a pilot plant study (Reference
Historical Data, Molybdenum/Tungsten/Tin Mine/Mill 6102 in this
section). Treating raw wastewater containing up to 24 mg/1 of
molybdenum, the pulsed-bed ion exchange pilot plant produced
effluent consistently containing less than 2 mg/1. Continuous
operation was achieved for extended periods of time, with results
indicating profitable operation through sale of the recovered
molybdenum and the procedure was put into full-scale operation.
The application of this technique at any specific site depends on
a complex set of factors, including resin loading achieved,
pretreatment required, and the complexity of processing needed to
produce a marketable product from eluant streams.
Radium 226 Removal
Radium 226 is a product of the radioactive decay of uranium. It
occurs in both dissolved and insoluble forms and, in this
industry, is found predominantly in wastewater resulting from
uranium mining and milling. Two treatment techniques are used in
this industry, and they represent state-of-the-art technology.
They are barium chloride coprecipitation and ion exchange.
Barium Chloride Coprecipitation. Coprecipitation of radium with
a barium salt (usually, barium chloride) has typically been used
for radium removal from uranium mining and milling waste streams
in the United States and Canada (References 46 and 47). The
removal mechanism involves precipitation of dissolved radium as
the sulfate in the first step which results in a residual
concentration of dissolved radium at this stage of approximately
20 ug/1. The dissolved radium concentration is then further
reduced by coprecipitation, whereby radium sulfate molecules are
incorporated into nascent crystals of barium sulfate. Dissolved
radium concentrations can be less than or equal to 1 to 3
picocuries (picograras)/liter (pCi/1). Effective settling is
necessary for removal of coprecipitated radium.
Dosages of 10 to 300 mg Ba/liter are generally required,
depending upon the characteristics of the waste stream. It has
been reported that 0.03 mole/liter of sulfate is required for
effective removal of radium (Reference 48).
253
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The removal of radium by adsorption of barite (the mineral form
of barium sulfate) has also been demonstrated in the laboratory.
More than 90 percent of the radium in uranium mine or mill
wastewater has been removed in this manner by passing the waste
stream through barite in a packed column (References 49, 50 and
51).
A number of facilities in the domestic uranium mining and milling
industry use the barium chloride coprecipitation process for
removal of radium, and this technology was used as the basis for
BPT effluent limitations. A summary of facilities which effec-
tively employ this technology is presented in Table VIII-9.
Ion Exchange. At uranium Mill 9452, a unique mine-water
treatment system exists which uses radium 226 ion exchange in
addition to flocculation, barium chloride coprecipitation,
settling, and uranium ion exchange. The mine water to be treated
is pumped from an underground mine to a mixing tank, where
flocculant is added. The water is then settled in two ponds in
series, before barium chloride is added. After barium chloride
addition, the water is mixed and flows to two additional settling
ponds (also in series). The decant from the final pond is
acidified before it proceeds to the uranium ion exchange system.
The uranium ion exchange column effluent is pumped to the radium
226 ion exchange system. After treatment for removal of radium
226, the final effluent is pumped to a holding tank for either
recycle to the mill or discharge.
The total treatment system at Mine/Mill 9452 is capable of
removing radium 226 from levels of 955 picocuries/liter (total)
and 93.4 picocuries/liter (dissolved) to 7.18 picocuries/liter
(total) and less than 1 picocurie/liter (dissolved). This per-
formance represents 99.2 percent removal of total radium 226 and
greater than 99 percent removal of dissolved radium 226.
Ammonia Stripping
High concentrations of ammonia in facility wastewater can be
effectively removed by air stripping processes. In this mass
transfer process, air and water are contacted in a packed or wet
column. Water is sprayed from the top of the column and allowed
to trickle down the wood or plastic media in packed columns, or
fall as droplets in wet columns. Air is conducted in a counter-
current mode (from bottom to top of column) or a cross-flow mode
(entering from the sides, rising and venting from the top of the
column) through the system by one or more fans or blowers.
The efficiency of ammonia removal by this system depends on pH,
temperature, gas-to-liquid flow ratio, ammonia concentration, and
turbulence of flow at the gas-liquid interface. Strippers are
operated at a pH of 10.8 to 11.5, which reduces the concentration
ammonia (NH3_). Proper design of the stripping unit considers
254
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ammonia concentration and temperature to determine column sizing
and gas/liquid flow rates.
Advantages of this system include its simplicity of operation and
control. Some disadvantages are inefficiency at low temperatures
(including freezeups at temperatures below 0 C), and formation of
calcium carbonate scale on tower packing material (due to lime
addition necessary for pH elevation). A further environmental
consideration is the quantity of ammonia gas discharged to the
atmosphere and its eventual impact on the concentration of
ammonia in rainfall.
A variation of the ammonia stripping process, which is currently
in its developmental stages, is a closed-loop system. This
system recovers the stripped ammonia gas by absorption into a low
pH liquid. The gas (initially air) passes from the stripping
unit to an absorption unit where its ammonia content is reduced,
and then returns to the stripping unit in a continuous closed
loop operation. This type of system allows for recovery of the
stripped ammonia and recycling of the absorption liquid in a
second closed loop. Thus, the system avoids discharge of ammonia
to water supplies as well as to the atmosphere. Also, since the
equilibrium condition for the gas stream in a closed system is
low in carbon dioxide, the problem of calcium carbonate deposits
on the stripping media is avoided. Further description of these
processes can be found in References 52 and 53.
Ammonia used in a solvent extraction and precipitation operation
at one milling site is removed from the mill waste stream by air
stripping. The countercurrent flow air stripper used at this
plant operates with a pH of 11 to 11.7 and an air/liquid flow
ratio of 0.83 m3 of air per liter of water (110 ft3 of air per
gallon of water). Seventy-five percent removal of ammonia is
achieved, reducing total nitrogen levels for the mill effluent to
less than five mg/1, two mg/1 of which is in the form of
nitrates. Ammonia may also be removed from waste streams through
oxidation to nitrate. Aeration will accomplish this oxidation
slowly, and ozonation of chemical oxidants will do it quickly.
However, these procedures are less desirable because the nitrogen
still enters the receiving stream as nitrate, a cause of
eutrophication.
PILOT AND BENCH SCALE TREATMENT STUDIES
Numerous pilot- and bench-scale wastewater treatment studies have
been performed throughout the ore mining and dressing industry.
These treatment studies were conducted to determine the
following: the effects of combining various waste streams on
wastewater treatability; the feasibility of employing a unit
treatment process or system for removal of specific pollutants;
the effluent quality attainable with a unit treatment process or
system; the optimal operating conditions of a treatment process;
and the engineering design parameters and the economics of
255
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building, operating, and maintaining a unit process or treatment
facility. The studies were conducted by industry, the University
of Denver, and EPA.
Copper Mine/Mill 2120
At copper Mine/Mill 2120, a bench-scale study of pH adjustment
and settling was conducted by the company to determine the
effects of combined treatment of water from an underground mine,
barren leach water, and mill tailings. The individual and com-
bined wastewater streams were treated with milk of lime to pH
values ranging from 6 to 11 and allowed to settle for 20 minutes.
The analytical results presented in Tables VIII-10 through VIII-
13 demonstrate the heavy metals removal attained. Metals removal
in all experiments were similar except for zinc, which was much
more effectively removed by combined treatment of the wastewater.
The solids produced by treatment of the mine water or barren
leach water were observed to be light and easily resuspended by
turbulence caused by wind action. However, when these waters
were treated in combination with tailings, the solids which
settled were more dense and not as easily resuspended by wind
generated turbulence. This led to the conclusion that the heavy
metal precipitates produced by lime treatment of the combined
wastewater streams are stabilized by the mill tailings.
Molybdenum Mine/Mill 6102
Molybdenum Mine/Mill 6102, which has historically- discharged
wastewater from its tailing pond system intermittently, has
performed extensive bench- and pilot-scale evaluations of tech-
niques for improving the quality of the wastewater discharged.
In addition to conventional precipitation technology the
following methods have been evaluated: ion exchange, alkaline
chlorination, ozonation, and electrocoagulation flotation.
Molybdenum recovery by ion exchange was evaluated in an extensive
pilot-plant study. This mill recycles water extensively and high
levels of molybdenum, on the order of 20 mg/1, are commonly found
in the discharge. Treatment of mill water in a pilot-scale,
pulsed-bed, counter-flow ion exchange unit achieved substantial
reductions in molybdenum concentrations, as demonstrated by the
summarized results below.
256
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Test Date ( 1975) _____ __ Mo(mg/l)
Effluent Eluate*
7/24 and 7/25 20.5 1.18 16,140
7/28 and 7/29 23.0 0.91 16,045
7/29 and 7/30 22.4 1.38 16,568
7/31 and 8/1 24.4 1.76 18,090
8/1 and 8/2 19.5 1.14 12,930
8/5 and 8/6 ^2_JJ 1 .38 17,484
Average 22.0 1.29 16,230
*Pregnant recovery fluid, see glossary,
For the period studied, service time was 41 minutes, resin-pulse
volume averaged 1.73 liters (1.83 quarts), and flow-rate feed was
121 to 125 liters (32 to 33 gallons) per minute. Effluent
concentrations of molybdenum were consistently below 2 mg/1. The
high concentrations achieved in the ion exchange eluate allow
economical recovery of the molybdenum, defraying a substantial
fraction (or possibly all) of the costs of the ion exchange
operation. On the basis of pilot-plant testing results, a
decision was made to install a full-scale ion exchange unit.
Laboratory tests of precipitation technology at. this site
indicate that, at the low effluent temperatures which prevail,
conventional precipitation technology would not be effective in
removing heavy metals at retention times considered economical.
Electrocoagulation flotation was evaluated as an alternative, and
the pilot-scale unit was run to define optimum operating condi-
tions and performance capabilities. Performance achieved at
various operating pH levels is summarized below:
Parameter _____ _ Concentration (mg/1)
pH (units)
Iron
Manganese
Zinc
Copper
Cadmium
Cyanide
Feed
__
25
6.
1 .
0.
0.
0.
Effluent a
to 3
3 to
4 to
59 to
03 to
22 to
5
6.6
1 .6
0.
0.
0.
74
04
33
8.
1 .
1 .
0.
0.
0.
0.
Effluent b
5
9
6
1
1
0
1
5
1
3
9.
0.
0.
0.
0.
0.
0.
Effluent c
2
6
5
04
10
02
07
10.
0.
0 .
0.
0.
0.
0.
4
8
1
04
09
01
06
Cyanide destruction and removal techniques were also evaluated in
conjunction with the electrocoagulation flotation pilot-plant.
Removal by ferric hydroxide sorption was found to be ineffective.
Ozonation did not consistently reduce cyanide to the desired
levels; however, substantial reductions were achieved at elevated
values of pH. Chlorinatiori using sodium hypochlorite in excess
(by a factor of 40) was found to be effective in reducing cyanide
concentrations to less than or equal to 0.02 mg/1.
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On the basis of pilot-plant test results, it was decided to
install full-scale electrocoagulation flotation treatment,
augmented by alkaline chlorination using sodium hypochlorite
(generated on-site by electrolysis of sodium chloride). The
treatment system treats a continuous bleed stream, with a capa-
city of 126 liters per second (2,000 gallons per minute), from
the mill water system. Post treatment is planned, as required,
to provide additional retention time for cyanide decomposition
and to decompose chlorine residuals, probably by addition of
sodium sulfide.
Actual performance capabilities of the full-scale system, which
has been on-line at this facility since July 1978, are presented
later in this section under the subheading Historical Data
Summaries.
Lead/Zinc Mine/Mill/Smelter/Refinery 3107
Facility 3107, a lead/zinc mine/mill/smelter/refinery complex,
has recently been investigating the feasibility of additional
treatment using filtration. A pilot-scale pressure filtration
unit is treating 9.5 to 31.5 liters per second (150 to 500 gal-
lons per minute) of treatment system effluent using granulated
slag as the filtration medium. Full-scale designs currently
under consideration will provide a maximum effluent total
suspended solids concentration of 5 mg/1 with 100 percent
reliability (industry report).
Lead/Zinc Mine/Mill 3144
Laboratory and pilot plant studies were conducted at lead/zinc
Mine/Mill 3144 in 1973 to define an effective treatment for the
destruction of cyanide. Preliminary laboratory tests were con-
ducted using calcium hypochlorite as an oxidizing agent.
Although this agent effectively destroyed cyanide contained in
the mill wastewater, the use of hypochlorite in a full-scale
operation was deemed inefficient and uneconomical (Reference 17).
As a result, a second series of tests was conducted using chlo-
rine gas as the oxidizing agent. Based on the results of these
tests, construction of a full-scale chlorination plant was
initiated in mid-August 1973. Startup operation of the full
scale plant began in December 1973. Monitoring data indicate
that the full scale plant effectively reduces cyanide (total)
from an average of 68.3 mg/1 in the raw waste to an average of
0.13 mg/1 in the treated effluent. The design and operating
characteristics of the full-scale plant have been previously
described in Technique Description, Cyanide Treatment earlier in
this section.
Canadian Mine Drainage Study
During 1973 and 1974, a pilot treatment plant was operated at a
mill located in New Brunswick, Canada, to demonstrate the
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treatability of base-metal mine water discharge using
conventional treatment technology and to define the factors
critical to the optimization of treatment. Treated effluent
polishing techniques were also evaluated and a final report of
the project was published (Reference 54). Several earlier
reports described the treatment plant design, optimization, and
capabilities and the development of flocculant addition and
sludge handling and dewatering methods (References 55, 56, 57,
and 58).
The pilot-plant treatment included provisions for two-stage lime
addition, coagulation, mechanical clarification, and sludge
recycle. Effluent polishing techniques employed included addi-
tional settling or sand filtration. Treatments of three acidic,
metal-bearing mine drainages were evaluated in the pilot-plant.
The characteristics of these three mine drainages are shown in
Table VIII-14. As indicated, the individual mine drainages
greatly differ in acidity and total metal content. Results of
treatment studies are summarized in Table VIII-15.
The principal findings of this pilot-plant project are summarized
as follows:
1. The following metal levels were attained as clarifier
overflow concentrations, on an average basis, for the
various drainages treated during periods of steady
operation:
Metal Concentrations (mg/1)
Extractable (total) Dissolved
Pb
Zn
Cu
Fe
2. Polishing of the clarifier overflow by sand filtration
and bucket settling further reduced the above extractable
total metal levels. Levels attained on an averaged basis
were:
Metal Concentration (mg/1)
Extractable (total)
Pb 0.12
Zn 0.19
Cu 0.04
Fe 0.17
3. The initial acidity and total metal concentrations had
little effect on the final effluent quality, but greatly
influenced the volume and density of sludge produced; these
factors affected the quantity of neutralizing reagent
required.
0.25
0.37
0.05
0.28
0.24
0.26
0.04
0.22
259
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4. Optimization of the coagulant (polymer) addition by
experiments run on the wastewater in question was determined
to be the process most critical to obtaining low
concentrations of metals in the clarifier overflow. The
beneficial effect of polymer addition was clearly
demonstrated in the treatment of mine 3 drainage, where
polymer additions increased settling rates fourfold (1.8 to
7.4 m/hr, equal to 5.9 to 24.3 ft/hr) and reduced the total
metal concentrations of the clarifier overflow sixfold. In
the case of mine 2 drainage, polymer addition reduced metal
concentrations by an additional 30 to 50 percent and
increased settling rates fivefold (0.45 to 2.4 m/hr, equal
to 1.5 to 7.9 ft/hr) during once-through operation of the
clarifier (i.e., no sludge recycle).
5. No performance advantages were found in two-stage lime
neutralization compared to single-stage lime neutralization.
The sensitivity of the process was found to be a function of
solid/liquid separation and not pH, provided that the pH was
maintained within one pH unit of the optimum.
6. Sand filtration and quiescent settling were shown to be
effective methods of further reducing metal values in
clarifier overflow and reducing the variability in these
levels.
University of Denver Mine-Drainage Study
The University of Denver, in cooperation with EPA and the State
of Colorado Department of Health and the Department of Game,
Fish, and Parks, has conducted field experiments to evaluate the
treatability of metal-bearing mine drainage from mines in the
highly pyritic districts of the San Juan Mountains of
southwestern Colorado (References 35 and 39). Charactersitics of
this mine drainage are tabulated below:
Parameter Concentration (mq/1)
pH (units) 2.6 to 3.1
Fe 336 to 800
Cu 51.6 to 128
Mn 4.52 to 19.0
Al 20.8 to 62.5
Zn 122 to 294
Pb 0.04 to 0.50
Ni 0.19 to 0.51
As 6.01 to 22.0
Cd 0.44 to 1.0
Sulfate 1 ,400 to 3,820
The study was conducted specifically to evaluate the capabilities
of the treatment scheme depicted in Figure VIII-2. This
treatment consisted of a two-stage process of chemical addition
260
-------
5.0
6.4
ND
12.7
6.4
ND
ND
LT 0.13
ND
5.5
6.4
ND
0.3
0.5
ND
ND
0. 13
ND
5.0
5.5
ND
30.0
6.8
LT 0.5
ND
0.29
ND
5.0
5.6
ND
30.0
7. 1
LT 0.3
ND
0. 19
ND
5.0
6.5
ND
0.2
0.4
ND
ND
0. 13
ND
to the mine drainage, followed by settling. The first stage
consisted of lime addition, and the second stage involved sulfide
addition. The pH was very easy to control with the first stage
achieving pH 5.0, and the second stage achieving pH 6.5 (the
ambient pH of the region). A second finding of the field studies
was that moderate, wind-induced turbulence in the settling pond
would maintain hydroxide floes in suspension, while the sulfide
precipitates settled immediately. During the field experiment,
pH was varied in the two stages of treatment. The results are
tabulated below:
Exp. 1 Exp. 2 Exp. 3 Exp. 4 Exp. __5
pH (Stage 1 )*
pH (Stage 2)*
Fe (total)
Zn
Mn
Cu
Al
Ni
Cr
*Value in pH units
LT = less than or equal to
ND = below detection limit
The experiment 5 results tabulated above, were reported to define
the standard design condition, which was held at steady-state for
an extended period. For this condition, the following effluent
concentrations of additional metals were attained:
Hg 0 mg/1
Cd 0.008 mg/1
As 0 mg/1
Lead/Zinc/Gold Mine 4102
Drainage from lead/zinc/gold Mine 4102 enters a precipitous
avalanche area which is too small for construction of settling
ponds of adequate size for conventional pH adjustment and set-
tling of mine water. As a result, research has been conducted at
this facility to design a satisfactory treatment system.
Various techniques, such as adsorption, reverse osmosis, and ion
exchange, were initially considered for treatment of the mine
drainage. However, chemical precipitation (conventional lime
precipitation) was ultimately chosen as the treatment process.
As indicated in Table VIII-16, this method has been demonstrated
to effectively precipitate dissolved metals present in the mine
water at elevated pH (i.e., pH 9.2).
After conventional lime treatment was selected as the alter-
native, an evaluation of technologies for removal of suspended
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solids was initiated. On the basis of the test results, the
EnviroClear and Lamella Gravity Settler techniques (commercial
package treatment systems) were determined to be efficient and
practical means of treatment which require a minimum amount of
space.
A number of sludge dewatering or sludge thickening methods were
investigated, including conventional sludge filtration (both the
drum filter and the frame filter press); centrifugation; the
Parkson Corporation's "Magnum Press"; Carborundum Company's "New
Sludge Filtering System"; Aerodyne Corporation's "Filtration
Cylinder"; and Enviro-Clear Company's "New Belt Filter." Sludge
recycle and use of coagulants were also considered as methods to
enhance settling and sludge dewatering. Although some of the
methods investigated were technologically feasible, no final
choice of a sludge dewatering or sludge-handling method was
identified as preferable on an economic basis.
Lead/Zinc Mine 3113
Bench scale studies were conducted by mine personnel at lead/zinc
Mine 3113 in 1975 through 1976 to evaluate the effectiveness of a
proposed treatment system to handle 6,400 m3 (1.7 million
gallons) of mine water drainage per day which is discharged
without treatment. The basic treatment scheme investigated
consisted of lime addition to pH 10 to 11, followed by
sedimentation. Sludge thickening and polyelectrolyte addition
were also evaluated. The results of these tests are presented in
Table VIII-17.
In subsequent bench-scale tests, mine-water samples were shipped
to the manufacturer of a high rate settling device (Lamella
Gravity Settler) to evaluate the effectiveness of this device
when used in conjunction with lime and polyelectrolyte. The best
overflow quality and sludge dewatering properties were attained
with 1.0 to 1.8 mg/1 polymer. Lime requirements were reduced by
15 percent by presettling prior to lime addition.
Based on the results of the treatment studies, a full-scale mine
water treatment scheme was developed. Mine water drainage would
be pumped to a lined holding lagoon with a theoretical retention
time of 24 hours (value = 6,400 m3 or 1.7 million gallons). Mine
water would be pumped from the holding lagoon to a tank, where
lime would be added to raise the pH to the range of 10.5 to 11.0.
Overflow from the lagoon would flow by gravity to a flash-mix
tank, where coagulant would be added to the stream prior to
passage to a pair of Lamella Gravity Settlers (in parallel).
Overflow from the settling units would flow to a polishing lagoon
with a theoretical retention time of approximately 6 to 9 hours
with sulfuric acid neutralization prior to discharge.
EPA Treatability Studies
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In August 1978, comprehensive studies of the treatability of
wastewater streams from ore mining and milling facilities were
initiated by Calspan Corporation under contract to EPA (Contract
68-01-4845). The primary purpose of this program was to delin-
eate the capabilities of BAT alternative treatment technologies
for mine and mill waters, technologies for the treatment of
uranium mill wastewater, and to expand the data for technologies
for which little or no empirical information was available. In
addition, the operating conditions were varied at each site for
the pilot-scale system used in the studies. This was done to
clarify engineering and economic considerations associated with
designing and costing full-scale versions of the treatment
schemes investigated.
The studies were performed at seven ore mining and milling
facilities and the results are summarized in this document (Refer
to Table VIII-18). A detailed discussion of these studies, their
analytical results, and the experimental designs and procedures
is presented in Reference 59. A discussion of two treatability
studies at Facility 2122 is presented in this section, under the
heading ADDITIONAL EPA TREATABIITY STUDIES.
All EPA-sponsored pilot scale treatability studies were conducted
on-site using a 2.4-meter (8 foot) by 12.2-meter (40 foot)
semitrailer designed specifically for performance of pilot- and
bench-scale wastewater treatment studies in the field. This
mobile treatment plant provides the following unit processes,
either individually or in combination, on a pilot-scale: flow
equalization, primary sedimentation, secondary sedimentation, pH
adjustment, chemical addition (polymer, lime, ferrous sulfate,
sodium hydroxide, barium chloride, sodium hypochlorite, and
others) coagulation, granular media pressure filtration, ozona-
tion, aeration, alkaline chlorination, ultrafiltration, flota-
tion, ion exchange and reverse osmosis. A schematic diagram of
the basic system configurations used are presented in Figures
VIII-3, VIII-4 and VIII-5.
Fifteen parameters, (pH, total suspended solids, and 13 toxic
metals) were monitored at all sites. Additional parameters such
as iron, aluminum, molybdenum, vanadium, radium 226, uranium,
phenols, and cyanide were monitored at appropriate sites.
Results of the testing at various facilities follow.
Lead/Zinc Mine/Mill 3121. At this facility, lead/zinc ore is
mined from an underground mine and concentrated in a mill by the
froth flotation process. Mine drainage is combined with mill
tailings for treatment in a tailing pond. A coagulant (polymer)
is added to the combined waste stream to improve settling in the
tailing pond. However, the tailing pond provides limited reten-
tion time, and the tailing pond decant generally contains rela-
tively high concentrations of metals. Therefore, the pilot-scale
treatment schemes investigated at this facility consisted of add-
on or polishing technologies for improved removal of metals from
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the tailing-pond effluent. Pilot-scale unit processes used
included lime addition for pH adjustment, coagulant (polymer)
addition as a settling and filtration aid, secondary settling,
and dual-media filtration.
The study at Mine/Mill 3121 was performed in two segments, the
first segment during warm weather (August), and the second seg-
ment during cold weather (March). Th-is scheduling was deliberate
since cyanide and copper concentrations in the tailing pond
decant at this facility are generally much higher during the cold
months than during the warm months. A major goal of the second
segment of this study was to determine the removal of copper by
effluent polishing techniques when relatively high concentrations
of both copper and cyanide were present. In addition, the capa-
bilities of alkaline chlorination and ozonation for destruction
of cyanide in the tailing pond decant were studied during March.
For reasons which will be discussed, the cyanide destruction
studies could not be completed.
A characterization of the wastewater influent (i.e., the tailing
pond decant) sent to the pilot-scale treatment system during the
period of study is presented in Tables VIII-19 and VIII-20. As
illustrated, pH, TSS, and total metals concentrations varied over
a wide range during the August study. This variability appeared
to be related to the schedule of mill operation (Refer to Table
VIII-21). During periods when the mill was not operating, only
mine water was being discharged into the tailing pond. (However,
this facility does not use lime treatment; alkaline mill water
provides pH adjustment for mine water). During the March study,
the concentrations of the parameters of interest in the decant
were generally much higher and less variable than during August.
Two basic experimental designs were employed to investigate metal
removal by effluent polishing. Initially, direct filtration of
the tailing pond decant was investigated. Subsequently, a second
set of experiments was conducted to determine the improvement in
metals removal attained by lime addition, coagulant (polymer)
addition, and settling prior to dual-media filtration.
A summary of the treated effluent concentrations attained is
presented in Tables VIII-22 (August study) and VIII-23 (March
study). Results of experiments to evaluate cyanide destruction
by ozonation are presented in Table VII1-24. Conclusions and
observations made on the basis of these results are summarized
below:
1. During the August study the metal removal efficiency of
filtration was found to be dependent on the pH maintained in
the tailing pond system. Metal removal efficiency improved
with increasing pH.
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2. Filtration consistently reduced the total suspended
solids concentrations of tailing-pond decant to 1 mg/1 or
less during the August study.
3. Also during the August study, filtration consistently
reduced total metal concentrations to levels well below BPT
limitations when the pH of the waste stream was in the range
of 7.7 to 11 .3. Zinc concentrations in excess of 0,5 mg/1
were observed in filtrates when the pH was below 7,7.
4. The results of this study demonstrated the chemical'
addition/settle/filtration treatment scheme to be very
effective for removal of TSS and metals. Furthermore, this
treatment system was 30 to 90 percent more efficient in the
removal of TSS and metals than filtration alone. Total
copper was reduced from an initial concentration of 0.15 to
0.19 mg/1 to a final concentration of 0.12 to 0.16 mg/1
attained by filtration alone.
5. While the results of these studies indicate that copper
can be removed from the wastewater of Mine/Mill 3121, some
additional observations need to be made. First, the
concentration of dissolved copper in the tailing pond decant
during the August study ranged from 0.010 .to 0.040 mg/1.
During the March study the dissolved copper concentration
was 0.010 to 0.050 mg/1. These concentrations are low
relative to the 30-day average dissolved copper
concentrations, typically 0.08 to 0.22 mg/1, reported by
Mine/Mill 3121 for NPDES monitoring purposes. Second, the
concentration of cyanide in the tailing pond decant during
August was 0.07 to 0.08 mg/1. During March the
concentration of cyanide ranged from 0.04 to 0.125 mg/1.
The cyanide concentration during August was typical of the
warm weather months, but the concentrations for March were
much lower than normal for the time of year (see Table VIII-
20). On the basis of these low dissolved copper and cyanide
concentrations, it does not appear that a copper cyanide
complex could have been present at any significant
concentration during either of the two treatability studies
conducted at Mine/Mill 3121. Therefore, there is no
evidence that copper can not be removed from this facility.
6. Results of cyanide destruction by ozonation indicate the
greatest degree of cyanide destruction occurred at the
highest ozone dosage rate. Even at a 10:1 ratio of ozone to
cyanide the efficiency of cyanide destruction was only 43
percent. This is due to the initial low concentration of
cyanide (i.e., 0.115 mg/1) and the problems involved in the
mass transfer of small amounts of ozone gas and contact with
cyanide in a dilute solution.
7, During the March study the cyanide concentration in the
tailing pond decant decreased to 0.04 mg/1 before the
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experiments to evaluate the destruction of cyanide by
alkaline chlorination and ozonation could be completed.
With an initial total cyanide concentration of only 0.04
mg/1 it was considered to be impractical to continue these
experiments. Therefore, only limited results for cyanide
destruction by ozonation were obtained and none were
obtained for alkaline chlorination.
Lead/Zinc
3107
Mine/Mill/Smelter/Refinery
generated from mining, milling, smelting,
at this lead/zinc complex are combined in
pond, and the effluent from this pond is
a physical/chemical treatment plant by
aeration, flocculation, and clarification,
_ Wastewater streams
and refining activities
a common impoundment
subsequently treated in
lime precipitation,
in conjunction with
high- density sludge recycle. Treated effluent from this
facility is characterized as being alkaline with relatively high
concentrations of zinc, cadmium, lead, and total suspended solids
(refer to Table VIII-25).
The pilot-scale treatment schemes investigated at this facility
focused on end-of-pipe polishing technologies for improved
removal of suspended solids from the treated effluent. The unit
processes investigated were dual-media granular pressure
filtration arid supplementary sedimentation. The use of polymers
and flocculation as settling aids was also investigated.
A characterization of wastewater treated in the pilot-scale
system is presented in Table VIII-26. It is evident from a
review of this table that metals in this waste stream are
components of the suspended solids, since the dissolved metal
concentrations are very low relative to total metal
concentrations.
A summary of results for the treatment schemes investigated is
presented in Table VIII-27. Treatment efficiencies are reported
only for BPT control parameters and other parameters present at
significant levels in the raw wastewater.
Results of this study indicate that the suspended solids
(especially, the metal hydroxide floes) in the effluent from the
physical/chemical treatment plant are filterable and not subject
to shear in the filters. Total suspended solids were consis-
tently removed to less than 1 mg/1 by all three filter configura-
tions investigated and over the range of hydraulic loadings
employed (i.e., 117 to 880 mVmVday, 2 to 15 gpm/ft2).
Correspondingly, metals were effectively removed by filtration.
Secondary settling reduced suspended solids by 81 percent from an
average of 16 mg/1 to 3 mg/1, with metal removals ranging from 38
percent (an average of 0.13 mg/1 to 0.18 mg/1) for lead to 72
percent (an average of 2.9 mg/1 to 0.79 mg/1) zinc (Table VIII-
27). A theoretical retention time of 11 hours was employed for
secondary settling experiments. The effluent quality produced by
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secondary settling was not as good as that produced by dual-media
filtration, probably due to the poor settling characteristics of
metal hydroxides (especially, zinc hydroxide).
A non-ionic polymer did not appear to enhance the settleability
or filterability of the wastewater treated. However, sufficient
time was not available for process optimization (i.e., polymer
and dosage selection, flocculation time, agitation intensity
during flocculation, etc.).
Lead/Zinc Mine 3113. Drainage from this lead/zinc mine flows
primarily from extensive inactive mine workings. Occasionally
mine water from an active mine is discharged via the mine
drainage system. Mine 3113 drainage is characterized as acidic,
with high concentrations of heavy metals, especially iron and
zinc (refer to Table VIII-28). The experiments conducted at this
site were to determine the quality of effluent which could be
attained by treatment of the mine drainage with lime
precipitation, flocculation, aeration, and mixed media filtration
processes. At present, this drainage is discharged without
treatment.
The character of the mine water treated during the period of
study is presented in Table VIII-29. It should be noted that the
pH is low; Cd, Cu, and Zn concentrations are practically all dis-
solved; and Fe is less than one half dissolved. Results of the
pilot-scale treatability studies are summarized in Table VIII-30.
Experimental treatment systems E, G, and I in Table VIII-30 were
designed to investigate the efficiency of lime addition,
aeration, polymer addition, flocculation, and settling at various
pHs. With the exception of zinc, the efficiencies of metals
removal were practically identical and independent of the pH.
However, in the case of zinc, a relationship was found between
the efficiency of removal and pH, and the greatest removal of
zinc occurred at the highest pH (10.5). In contrast to the
improved efficiency of zinc removal, the total suspended solids
concentration increased with increasing pH. This was the result
of the increased lime dosages required to attain the higher pH
levels.
Experimental systems identified as A and C in Table VIII-30 were
designed to investigate anticipated improvements in treatment
efficiency by using aeration to oxidize ferrous (Fe+2) ion to the
ferric (Fe+3) state. A ferric hydroxide precipitate is formed
(in system C); however, only slightly improved iron removal (4.8
to 4.0 ug/1) was observed. No improvement in TSS or other toxic
metals removal was observed. Heavy reddish-brown sediments
(yellowboy) were observed in the mine-drainage discharge ditch,
indicating that the iron present was mostly oxidized.
Experimental systems identified as C and E in Table VIII-30 were
used to investigate the extent to which the treatment efficiency
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of the basic lime and settle treatment system could be improved
by incorporating polymer addition and flocculation. Review of
the experimental results demonstrates that polymer addition,
followed by flocculation, greatly improved the capabilities of
the basic lime and settle system. Most notable was the threefold
improvement in removal efficiency of total suspended solids,
zinc, and iron.
A dual-media, granular filtration step was used with all systems
as a final polishing step (systems D, F, H, and J). The
incremental improvements in removal of total suspended solids and
total metals resulting from filtration are also represented in
Table VIII-30. Results indicate filtration is an effective
polishing treatment showing significant (14 to 5 mg/1) improve-
ments in TSS in all cases and general improvement in metals
concentrations.
On the basis of the experimental results, the treatment scheme
producing optimum removals of suspended solids and heavy metals
from acid mine drainage at Mine 3113 consisted of adjustment of
pH to 10.5 with lime, flocculant addition, flocculation, sedi-
mentation, and filtration (experimental system J in Table IX-30).
Other, less rigorous treatment schemes with lower lime dosages
and without filtration would not reliably produce the excellent
effluent quality attained with system J.
Aluminum Mine 5102. At this site, bauxite is mined by open-pit
methods. Watewater (approximately 17,000 m3, or 4.5 million
gallons, per day) emanates as runoff and as drainage from the
open-pit mine. This wastewater is generally characterized as
acidic (with a pH of 2.2 to 3.0}, with total iron and aluminum
concentrations in the range of 50 to 150 mg/1 and 50 to 200 mg/1,
respectively. The treatment system used for this mine water con-
sists of lime addition and sedimentation in a multiple pond
system.
in
The character of treated mine water (influent to pilot-scale
treatment system) during the period of study is presented in
Table VIII-31. As indicated, concentrations of the 13 toxic pol-
lutant metals were found to be either below detection limits or
only slightly above detection limits of atomic adsorption
spectrophotometric analysis. Other parameters (TSS, iron, and
aluminum) were present at higher concentrations, but were well
below BPT limitations.
The basic pilot-scale unit treatment processes investigated at
Mine 5102 consisted of lime addition, aeration, polymer addition,
flocculation, sedimentation, and dual-media filtration. Results
of the treatability studies are summarized in Table VIII-32.
These results are of limited value because the waste stream being
treated was already of a very high quality (low pollutant
concentrations).
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Findings of the treatability studies at Mine 5102 are summarized
below:
1. Elevated pH, in the range of 8.2 to 10.7, did not
significantly improve or degrade removal efficiency of
aluminum or iron, as illustrated by the results shown in
Table VIII-32. This is to be expected from the low
dissolved metal concentrations in the mine water influent to
the pilot-scale treatment units (Table VIII-31).
2, The use of a polymer improved settling performance
during lime addition experiments, species, i.e., arsenate
(As04~3), molybate (Mo04~2), selenite )Se03-2), and
vanadates (HV04~2, V04~3, etc.). These metal species are
highly soluble at alkaline pH and cannot be removed by
precipitation as hydroxides or carbonates.
3. Filtration consistently produced effluent total sus-
pended solids concentrations of 1 mg/1 or less. Total
aluminum and iron concentrations were equal to the
corresponding dissolved metal concentrations, indicating
essentially complete removal of particulate metal compounds.
The use of hydraulic loadings of 117 to 880 mVm2/day (2 to
15 gpm/ft2) and effective filter media sizes over the ranges
of 0.35 mm to 0.7 mm, respectively, did not significantly
alter the quality of effluent attained.
Uranium Mill 9402. This mill, like all domestic uranium ore
mills, is located in an arid region and attains zero point
discharge of wastewater. This is accomplished by recycling and
using evaporation ponds. It was selected as a wastewater
treatability site for several reasons. It is possible that
regulations will be developed to protect groundwater quality
pursuant to other statutes. Such regulations could mandate the
use of seepage control devices and practices which would elimi-
nate loss of wastewater from tailing ponds by seepage. This
could ultimately necessitate end-of-pipe discharges of wastewater
at uranium mills presently attaining zero discharge. For these
reasons, EPA found it desirable to investigate treating
wastewater generated at uranium mills.
As discussed in Section III, two basic processes are employed at
uranium mills, acid leaching and alkaline leaching. The pH in
wastewater produced in an acid leach circuit are different from
those in wastewater produced in an alkaline leach circuit.
Therefore, uranium mills representative of both processes were
selected. Mill 9402 was selected as a representative acid leach
mill; Mill 9401, discussed later, is an alkaline leach mill.
The wastewater treatability study conducted at the uranium Mill
9402 was performed in two phases. The initial phase was
performed during the period of 4 to 9 November 1978 and the
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second phase during the period of 3 to 12 December 1978, The
basic pilot-scale treatment scheme employed during the initial
phase consisted of lime precipitation (pH adjustment), aeration,
flocculant (polymer) addition, barium chloride coprecipitation,
flocculation, single-stage or two-stage settling, and mixed-media
filtration. Reagent dosages and operating parameters employed
during the initial phase of this study were chosen largely on the
basis of previous laboratory studies conducted by the Australian
Atomic Energy Commission (Reference 49) and by Calspan Corpora-
tion (Reference 59).
During all experimental runs the inclined settling tank used in
the pilot plant was operated in a manner similar to a sludge
blanket clarifier (see Figure VIII-6). 'However, during the
initial phase of the study, the sludge produced was light and
unconsolidated. For this reason, it was impossible to maintain a
surface layer of clarified supernatant within the settling tank
and the effluent contained high concentrations of total suspended
solids. As a result, the total concentrations of certain metals
in the effluent also tended to be high, although the dissolved
concentrations were relatively low.
During the second phase of the study an attempt was made to
improve this situation by incorporating sludge recycle and a
metered sludge bleed into the pilot-scale system (see Figure
VIII-4). This modification was added specifically to consolidate
and thicken the sludge. Sufficient time was not available to
optimize this process, but it was successful enough to allow con-
tinuous operation of the pilot plant. During the final experi-
mental runs, the sludge recycle produced a noticeably thicker
sludge and made it possible to maintain a 10 to 15-centimeter (4
to 6-inch) layer of clarified supernatant in the settling tank.
A summary of the physical/chemical characteristics of the raw
wastewater (i.e., tailing pond seepage) at Mill 9402 which was
used in the treatability studies is presented in Tables VIII-33
and VIII-34. Examination of these tables reveals that this
wastewater is very acidic and contains very high concentrations
of total dissolved solids, dissolved metals, and radium 226
(total and dissolved). A comparison of these two tables further
indicates that the character of the wastewater during the second
phase of the study (December) was somewhat different from the
wastewater character during the initial phase of the study
(November). Specifically, several parameters including TSS, Mo,
Fe, Mn, and Al were present at much higher concentrations during
the second phase of the study. The higher concentrations of
these parameters were not found to have a readily apparent impact
on the treatment system capabilities.
A review of the raw wastewater character at Mill 9402 demon-
strates the metals present at high concentration to be Cu, Pb,
In, Ni, Cr, V, Mo, Fe, Al, and Mn. Addition of lime to form
metal hydroxide precipitates is known to be an effective removal
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mechanism for all of these metals except V and Mo (References 38,
60, 61 and 62). However, the high concentrations of Fe and Al in
the wastewater suggested other possible removal mechanisms for
these latter two metals. Hem (1977) has reported that the
solubility of vanadium and molybdenum may be controlled by
precipitation of iron vanadates and molybdates over the pH range
of 3 to 9 for vanadate and 5.3 to 8.3 for molybdate (Reference
63). Michalovic e_t aj_ (1977) conducted laboratory studies and
reported that excess ferric hydroxide formed by oxidation and
hydrolysis of ferrous sulfate was found to consistently yield
vanadate (+5) concentrations of less than 4 mg/1 (Reference 64).
The removal mechanism proposed involved precipitation/coagulation
as given below:
2Fe+3 + 3 (VOj)-1 + 30H- > Fe(V03)3 + Fe(OH)3
The pH was adjusted by addition of lime and a final pH of 7 was
reported to provide the best results. Similarly, Kunz, et. al.
(1976) reported that the results of screening tests showed that
ferrous sulfate provided the most efficient removal of vanadium
anions (Reference 65). Concentrations of vanadium of less than 5
mg/1 were attained when the pH was kept between 7.5 and 9 for V+4
precipitation and between about 6 and 10 for V+5 species. Again,
the removal mechanism was thought to involve simultaneous
precipitation of Fe(V03_)2^ and Fe(OH). Similar removal mechanisms
have been reported for Mo (Reference 66). However, the optimum
pH for Mo removal using iron salts is much lower than required
for V removal (i.e., about pH 3 to 4 for Mo). During laboratory
studies significant removal of Mo has also been attained with
aluminum hydroxide at a pH of about 4.5 (Reference 66).
On the basis of the wastewater (i.e., tailing pond seepage)
characteristics and the literature summarized above, it was
anticipated that the treatment scheme chosen for pilot scale
testing would provide effective removal of most of the metals of
concern. However, the removal of Mo and total dissolved solids
(TDS) were expected to present some problems. The pilot scale
experiments conducted were designed to investigate the effect of
variable lime and barium chloride dosages on removal of metals,
TDS, and coprecipitat ion of radium 226, respectively. A summary
of the st ay results is presented in Table VIII-35.
Gener-.ilJ.j', p>l values greater than the range 8.2 to 9 were
recnured f.^r optimum removal of TDS and most metals. However,
opcimum removal of molybdenum occurred at the lowest pH range
investigated, pH 5.8 to 6.1, and improved removal of this metal
would probably require operation at an even lower pH. A barium
chloride dosage of 51 to 63 mg/1 was required for optimum removal
of radium 226.
The basic treatment scheme employed did not demonstrate effective
removal of ammonia because it was not designed for removal of
this parameter.
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As described previously, the major operating problem encountered
was maintaining control over suspended solids and. sludge removal
in the sedimentation unit. The metals which appeared to be the
most sensitive to this problem were zinc, uranium, and radium
226.
Reduction of effluent TSS concentrations would also improve the
total metal concentrations for all the metals subject to
precipitation or coprecipitation removal mechanisms.
Uran i um/Vanad i um Mill 9401. This mill is located in an arid
region of New Mexico. An alkaline leaching process is employed
at this mill to selectively leach uranium and vanadium from ore.4
This facility achieves no end-of-pipe discharge of process
wastewater by: (!) net evaporation (due to location in an arid
region), (2) loss of water as seepage, and (3) recycle. The
rationale for selection of this uranium mill as a treatability
site was dicussed for uranium Mill 9402, which uses acid leach;
Mill 9401 uses alkaline leach.
Clarified water from the tailing pond is passed through an ion
exchange column prior to recycle to the mill leaching circuit.
The purpose of the ion exchange unit is to recover solubilized
uranium present in the recycle stream. This is apparently an
economically feasible process for this mill, since the mill has
continued to operate and recover uranium which would otherwise be
lost by this approach.
Treatability experiments conducted on the water recycled from the
tailing pond focused on removal/recovery of dissolved uranium and
removal of other dissolved components, especially metals. As
indicated in Table VIII-36, the metals of highest concentration
(other than U) and, therefore, of interest are arsenic, molyb-
denum, radium 226, selenium, and vanadium. The highly alkaline
and oxidized character of the mill wastewater and the existence
of the metals in soluble form indicates their presence as anionic
species, i.e., arsenate (As04-3), molybate (Mo4~2), selenite
(SeOj"2), and vanadates (HV04-2, V04~3, etc.). These metals
species are highly soluble at alkaline pH and cannot be removed
by precipitation as hydroxides or carbonates.
Therefore, the treatability experiments conducted at this site
focused on two basic treatment schemes. The first involved pas-
sing the wastewater through an ion exchange column containing
amberlite IRA-430 resin to remove uranium. Ion exchange was
followed by coprecipitation with ferrous sulfate, alum, or lime
in conjunction with H2S04_, polymer addition and flocculation, and
aeration. Results of preliminary bench scale experiments indi-
cated that of the three chemical reagents investigated, (ferrous
sulfate, alum and lime), ferrous sulfate provided the most effec-
tive removal of metals and was employed during all subsequent
pilot scale experiments. Therefore, the basic pilot-scale treat-
ment scheme consisted of ion exchange followed by ferrous sulfate
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addition/pH adjustment/aeration, barium chloride addition for
coprecipitation of radium 226, polymer addition, flocculation,
sedimentation, and dual media filtration. A schematic represen-
tation of the pilot-scale treatment system used is presented in
Figure VIII-5,
The second treatment scheme investigated used the mixture, in
varying proportions, of wastewater from an acid leach uranium
mill (Mill 9402) and the alkaline leach wastewater of Mill 9401.
Bench-scale treatment units were used.
Results of the ferrous sulfate/barium chloride coprecipitation
system investigated at pilot-scale are presented in Table VIII-
37. Results of the acid mill/alkaline mill wastewater admixture
treatment scheme investigated at bench-scale are presented in
Table VII1-38. A summary of the raw wastewater character (i.e.,
tailing pond recycle water) is presented in Table VIII-36.
The results summarized in Table VIII-38 indicate that ion
exchange removed approximately 97 to 99 percent of the uranium
present in the waste stream while 98 percent removal of radium
226 was attained by barium chloride coprecipitation with a BaC12,
dosage of 15 to 60 mg/1. The effectiveness of removing vanadium,
molybdenum, and selenium increased with decreasing pH. At pH
8.0, approximately 80 percent of the vanadium and 50 percent of
the molybdenum and selenium were removed. The TDS concentration
remained high (in excess of 20,000 ug/1) in the effluent because
(1) the metals precipitated were at comparedly (compared to TDS)
insignificant concentrations and (2) dissolved solids in the form
of Fe (S04J, BaCl^ and polymer were added to the water as part of
the treatment.
Because acid and alkaline leach uranium mills are sometimes
located in close proximity, the mixture of wastewater from these
two types of mills for neutralization and treatment may be a
feasible alernative. Therefore, admixture experiments were con-
ducted to investigate the degree of neutralization and the
removal of molybdenum, selenium, and vanadium attained. These
metals were of special interest since they are extremely diffi-
cult to r^juove from wastewater.
Enhar;:; ratals removal was observed under all conditions
studied. s,\.;-Limum removal of metals was achieved at the highest
:,i.:.io of acid to alkaline wastewater investigated (i.e., 5:3
ratio by volume). Even at a ratio of 5:4 acid to alkaline waste-
water the removal efficiency of both Mo and V exceeded 97 per-
cent. At admixture ratios of 5:4 and 5:3 by volume the final pH
attained was 4.3 and 3.9, respectively. The amount of iron
remaining in solution following admixture and the final pH sug-
gests that a lime barium chloride addition treatment scheme would
be very effective for subsequent treatment of the acid/alkaline
wastewater mixture (see discussion of treatability study con-
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ducted at Mill 9402). Time did not permit investigation of
lime/barium chloride precipitation, however.
It is notable that the admixture treatment scheme was the only
scheme investigated which resulted in the effective removal of
molybdenum. In view of the high cost required for neutralization
of acid or alkaline uranium mill wastewater (if such treatment
were ever required), admixture provides an additional advantage.
EPA-Sponsored Studies at Complex Facilities
During the month of September 1979, two pilot-scale treatability
studies were conducted by Frontier Technical Associates, Inc.
under contract to the EPA (Contract No. 68-01-5163). These
studies were conducted to gather data on treatment at mine/mill/
smelter/refinery complexes. The studies were conducted on-site
using an EPA mobile laboratory truck and company owned, pilot
scale treatability equipment. This equipment included a 90
gallon batch lime mix tank, dual media filter column, flow tray,
100-gallon filtrate holding tank, and associated pumps, piping,
valves, and instrumentation. Figure VIII-7 is a schematic of the
pilot-plant configuration.
The test operating conditions were varied at each site. Tests
included combinations of pH adjustment by lime addition, second-
ary settling, dual media filtration, and dosing with hydrogen
peroxide for cyanide treatment.
Samples were monitored for pH, total suspended solids, cyanide,
phenols, and the 13 toxic metals.
Copper Mine/Mill/Smelter/Refinery 2122. The wastewater treatment
plant at this facility treats the combined waste streams from two
mills, a refinery (including a refinery acid waste stream), a
smelter, and the facility sanitary wastewater. Existing
treatment includes lime addition, polymer addition, flocculation,
and settling. The pilot-scale treatment schemes investigated at
this facility consisted of polishing technologies for improved
metals removal.
A characterization of the untreated mine/mill/smelter/refinery
wastewater during the period of the treatability study is pre-
sented in Table VIII-39 and a summary of the treated (existing)
wastewater characteristics is given in Table VIII-40. The
influent wastewater data was taken from analyses of daily compo-
site samples (each daily non-flow proportional composite was
composed of periodic grab samples taken over a three- to eight
hour period). Treated effluent quality is the average of indi-
vidually analyzed grab samples taken periodically each day over a
two- to ten-hour period. Treated effluent samples taken from the
facilities treatment plant represent the influent to the pilot
plant system.
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Sixteen treatability runs were completed during the test period
(Table VIII-41). Test runs 01 through 05 used dual-media
filtration at varying flow rates. Test run 06 used batch lime
addition and one-hour settling. Test runs 07 through 09 included
batch lime addition and flocculation followed by dual-media
filtration at varying flow rates. Tests 10 through 13 represent
one dual- media filter run at a fixed flow rate for an extended
time (samples were taken after five minutes, six hours, 12 hours,
and 18 hours). Runs 14, 15 and 16 were batch lime treated,
followed by dual-media filtration and varying dosages of hydrogen
peroxide for cyanide removal.
The following observations were made:
1. Dual-media filtration (tests 1-5) consistently achieved
total suspended solids concentrations of less than or equal
to 4 mg/1; total copper concentrations less than or equal to
0.11 mg/1; total lead concentration less than or equal to
0.018 mg/1; and total iron concentrations less than or equal
to 0.09 mg/1. Zinc was less efficiently removed (influent
mean = 0.309 mg Zn/1).
2. Lime addition with flocculation and secondary settling
(one test run, test 6) achieved a total suspended solids
concentration of 8 mg/1; total copper concentration of 0.25
mg/1; total lead concentration of 0.04 mg/1; and total zinc
concentration of 0.17 mg/1.
3. Dual-media filtration with lime addition to a pH of
approximately 9.0 (tests 7 through 9) achieved total
suspended solids concentrations less than or equal to 1
mg/1; total copper concentrations less than or equal to
0.005 mg/1; and total zinc concentrations less than or equal
to 0.043 mg/1,
4. During the 18 hour filter run (tests 10 through 13) no
solid breakthrough occurred and metals concentration were
relatively steady.
5. Lime addition to pH 10 and filtration (test 16) showed
no improvement over the pH 9 tests.
No significant changes were observed in any of the remaining
pollutants measured. A summary of pH, TSS, Cu, Pb, and Zn
treatability study effluent concentrations is displayed in Table
VIII-41.
Copper Mine/Mi 11/Smelter/Ref inery 2121. The wastewater treatment
system at this facility receives water from a smelter, a
refinery, and a sanitary sewer. The combined flow is discharged
to a tailing pond. Decant from the tailing pond passes through a
series of five stilling ponds before final discharge. Table
VIII-42 characterizes the facility's wastewater discharge during
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the pilot-scale treatability study conducted by Frontier
Technical Associates. The quality was very good and represents
the influent to the treatability study.
The study conducted included three test runs. Tests 01 and 02
were dual media filtration tests at hydraulic loadings of 6.5
gpm/ft2 and 9.3 gpm/ft2 respectively. Test 0.3 combined lime
addition to a pH of 8.8 with dual media filtration at a hydraulic
loading of 9,1 gpm/ft2.
The results of sample analyses indicate no significant change in
the already low concentrations of most pollutants. TSS levels
dropped from an average of 4.1 mg/1 to an average of 1.3 mg/1.
In test 03, the pH dropped to 7.7 through the filter column,
while causing partial clogging of the filter. Lime was visible
in the filter effluent. This phenomenon presumably is due to the
low solubility of lime in the facility wastewater which is high
in sodium and calcium chloride salts. A summary of treatability
study effluent sampling data is presented in Table VIII-43.
HISTORICAL DATA SUMMARY
This subsection presents long-term monitoring data gathered from
individual facilities in several subcategories. Facilities
considered here include those for which long-term data are
available and which are regularly achieving or surpassing BPT
limitations by optimizing their existing treatment systems.
Iron Ore Subcategory
Mine/Mill 1108 is located in the Marquette Iron Range in northern
Michigan. The ore body consists primarily of hematite and
magnetite and is mined by open-pit methods. Approximately
8,800,000 metric tons (9,700,000 short tons) of ore are mined
yearly. The concentration plant produces approximately 2,800,000
metric tons (3,100,000 short tons) of iron ore pellets annually.
Wastewater is presettled prior to treatment with alum and a long
chain polymer to promote flocculation and improve settling
characteristics. The treated water is polished in additional
small settling ponds prior to discharge.
Table VIII-44 is a summary of industry supplied monitoring data
for the period January 1974 through April 1977. As indicated, pH
is well controlled and always in the range of 6 to 9. Alum and a
polymer have been used on a continuous basis since 1975 to
improve TSS removal at this facility. Since that time, TSS
control has been excellent and has exceeded 30 mg/1 only three
times on a daily maximum basis. On a monthly average basis, this
facility has exceeded 20 mg TSS/1 only once since 1975.
Dissolved iron concentration averages approximately 0.36 mg/1.
Since 1975, dissolved iron concentration has not exceeded 2.0
mg/1 on a daily basis or 1.0 mg/1 on a monthly basis.
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Copper, Lead, Z_inCj_ Gold, Silver, Platinum, and Molybdenum Ore
Subcategory
Copper/Silver Mine/Mill/Smelter/Refinery 2121
This facility is located in northern Michigan, with copper and
silver ore extracted by underground methods. Mine production in
1976 was approximately 3,281,000 metric tons- (3,617,000 short
tons) of ore with 125,000 metric tons (138,000 short tons) of
copper concentrate, and 185 metric tons (204 short tons) of'
silver concentrate. The primary mineral form is chalcocite.
Concentration is accomplished by the froth flotation process.
The smelter and refinery contribute wastewater to the combined
treatment system. Wastewater also originates from power
generation, sewage treatment, and collection of storm runoff.
Wastewater from the above sources is combined in a tailing pond
and decanted to a series of small settling basins before final
discharge. The alkaline pH of the treatment system is maintained
by the alkaline nature of the discharge from the mill as well as
by the addition of lime to the slimes fraction of the tailings.
The limed slimes are combined with all other wastewater sources
in a mixing basin and then pumped into the tailing pond. The
mine water contribution to the total discharge ranges from 0 to
4,500 mVday (0 to 1.2 million gal/day), and this mine waste
stream is released into the tailing 'pond on a seasonal basis.
The total pond discharge volume averages approximately 79,000
mVday (approximately 21 million gal/day).
Discharge monitoring data supplied by the company for a 58-month
period between March 1975 and December 1979 are presented in
Table VII1-45. This summary presents data derived from monthly
averages for all parameters. The data presented for pH and TSS
represent almost continuous daily monitoring throughout the
reporting period. For these two parameters, the values shown are
based on approximately 1,500 measurements. These data indicate
that, for the parameters monitored, effluent performance is
consistently far below BPT effluent standards, even when the
maximum values reported are considered.
Wastewater treatment practices at Mine/Mill/Smelter/Refinery 2121
which are employed to attain its high quality effluent are:
1. Supplemental lime addition for improved coagulation,
metals removal, and pH control
2. Use of a multiple pond system for improved settling con-
ditions and system control
3. Sufficient pond volume to provide adequate retention
time for sedimentation of suspended particulates and metals
4. Provision of a tailing pond design resulting in rela-
tively efficient and undisturbed sedimentation conditions
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5. Use of a decant configuration which effectively controls
pond levels without disturbing settled solids in the
vicinity of the decant towers
6. Mixing all waste streams prior to entry into the tailing
pond system to reduce the possibility of thermal
stratification and pH fluctuations.
Copper Mine/Mill 2120
This mine/mill facility is located in southwest Montana. The ore
body consists primarily of chalcocite and enargite, mined only by
open-pit methods at present. Underground mines at this facility
are inactive, but mine water is continuously pumped. The mill
employs the froth flotation process to produce copper concen-
trate, while cement copper is produced by dump leaching of low
grade ore. In 1976, ore production was 15,419,000 metric tons
(17,000,000 short tons), and 327,000 metric tons (360,000 short
tons) of copper concentrate were produced. Approximately 16,000
metric tons (17,600 short tons) of cement copper are produced
annually.
Schematics of the wastewater treatment system employed at Mine/
Mill 2120 are presented in Figures VIII-8 and VIII-9. .Figure
VIII-8 portrays the system configuration as it existed during the
period (i.e., September 1975 through June 1977) when the data
presented in Table VIII-46 were collected.
Wastewater streams routed to the tailing pond system for
treatment include underground mine water, excess leach circuit
solution, and mill tailings. The mine water is acidic because
sulfuric acid is added to prevent iron-deposit fouling of pipes
and pumps used for mine dewatering. The acidic leach circuit
waste stream results as a 3 percent bleed from a 190,000 m3 (50
million gallons) of solution recycled through the leach circuit
daily. Reportedly, this bleed is used because seepage into the
dump leach system necessitates discharge of excess water.
Additional lime is added to the mill tailings to neutralize the
acidity of the mine water and leach solution. These three waste
streams are thoroughly mixed prior to combined discharge into the
tailing pond. Prior to 1977 the tailing pond decant was largely
recycled to the mill for use as process water, but tailing pond
overflow was discharged when effluent quality permitted. When
tailing pond decant was discharged, the average daily discharge
volume was 11,000 m3 (3 million gallons).
When tailing pond overflow quality did not permit discharge,
wastewater was reintroduced into the mill circuit and subse-
quently mixed with open-pit mine water for additional treatment
in a second treatment system (i.e., the "barrel pond" system).
This treatment system consists of a three celled settling pond
where the influent wastewater is limed and polymer is added to
enhance flocculation and settling. A relatively high pH is
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maintained through this treatment system, but a final pH adjust-
ment is made when necessary by addition of sulfuric acid.
Average discharge volume from this treatment system is
approximately 25,000 m3 (6.5 million gallons) per day.
Figure VIII-9 shows modifications which have been made to the
treatment system at Mine/Mill 2120 since June 1977. Although
direct discharge of tailing pond decant has not occurred during
the past two years, discharge could occur if excess water condi-
tions warrant. Notable among the changes made to the treatment
system is the addition of a pond for secondary settling of tail-
ing pond decant before recycle. In addition, open pit mine
drainage has been directed to the tailing thickeners to avoid
surge and overflow. However, mine/mill personnel report that
overflow from the surge pond still occurs intermittently and is
still combined with treated effluent from the barrel pond system
for final discharge. As will be discussed, this latter practice
has an adverse impact on the quality of the combined discharge
stream.
Tailing pond effluent monitoring data supplied by industry are
presented in Table VI11-46. These data have been summarized for
the period September 1975 to June 1977 on the basis of both daily
averages and averages of monthly means. It is noted that the pH
of the tailing pond effluent falls outside of the BPT limits much
of the time, but a high pH level is reportedly maintained to
improve pH values downstream of the discharge point with the
consent of the state.
Practices which have been identified as essential to the
attainment of consistent and reliable treatment in the tailing-
pond system are:
1. Maintenance of pH slightly in excess of 9.
2. Maintenance of an earthern dike (baffle) within the
tailing pond to prevent short circuiting and reduce wind
induced turbulence.
3. Discontinuation of tailing pond discharge during upset
conditions.
Effluent monitoring data describing the quality of effluent
discharged from the second treatment system (i.e., the barrel
pond system) are presented in Table VIII-47, a summary for the
period January 1975 to September 1977.
It is important to note that the data presented in Table VIII-47
do not accurately reflect the capabilities of the barrel pond
treatment system. The reason for this, as indicated in Figure
VIII-9, is that untreated wastewater is often combined with
treated effluent prior to final discharge. (The effluent
monitoring station is located downstream of the point where these
279
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waste streams are combined.) This practice adversely impacts the
quality of final discharge and is considered to be primarily
responsible for the BPT violations. (The exception is pH, which
reportedly is purposely maintained at a high level to improve
acid conditions in the receiving stream.) Industry personnel
report that actions are presently being taken to eliminate the
necessity for this practice.
Lead/Zinc/Copper Mine/Mill 3105
This underground mine is located in Missouri. Galena, sphal-
erite, and chalcopyrite (lead, zinc, and copper minerals) are the
primary minerals recovered. Ore production began in 1973, and
reported mine production was 1,032,000 metric tons (1,137,700
short tons) in 1976.
Mining and milling wastewater streams are treated separately.
The mill operates in a closed loop system; tailings are treated
in a tailing pond, and the pond decant is recycled back to the
mill. Some mine water is used as makeup water in the mill
flotation process. Excess mine water, averaging 8,300 cubic
meters (2.1 million gallons) per day is treated by sedimentation
in a 11.7 hectare (29 acre) settling pond.
Effluent monitoring data for the mine water treatment system at
Mine 3105 are presented in Table VII1-48. This data summary is
based on NPDES monitoring reports submitted for this facility for
the period January 1974 through January 1978.
The mine is an example of low solubilization of heavy metals due
to the mineralization of the ore body. More specifically, the
ore body is low in pyritic minerals and exists in a dolomitic
host rock. Mines exhibiting low solubilization potential are
characterized by mine waters of near neutral to slightly alkaline
pH.
Mine water treatment at Facility 3105 illustrates that simple
sedimentation at mines exhibiting low solubilization potential
may be sufficient to achieve water quality superior to BPT limi-
tations, by effective removal of suspended solids and associated
particulate metals (see Table VIII-48).
Lead/Zinc/Silver Mine 3130
This facility is located in Utah and produces ore with economic
mineral values of sphalerite, galena, and tetrahedrite in a
quartz and calcite matrix. Production at this facility is
confidential. No discharge occurs from the associated mill by
virtue of process wastewater impoundment and solar evaporation.
Mine water pumped from this operation averages 32,700 m3 (8.64
million gallons) per day. The mine water treatment system con-
sists of lime and coagulant addition, followed by multiple-pond
sedimentation. Backfilling the mine with the sand tailing
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fraction from the milling circuit is practiced. Since the asso-
ciated mill utilizes sodium cyanide in the flotation process, the
sand tailings used for backfilling contribute cyanide to the mine
water discharge. This cyanide is not effectively removed by the
treatment system. The problem of cyanide in mine water resulting
from operations which practice cut and fill techniques has been
discussed in Section VI.
Tables VII1-49 and VII1-50 summarize data on raw and treated
waste streams by industry for the period June 1977 to October
1977. A new treatment system was recently brought online, so a
great deal of data are not available. Examination of raw waste
data indicates that the mine water contains high concentrations
of metals. As shown in Table VIII-5Q, the careful control of pH,
use of a settling aid (i.e., polymer), and use of a multiple pond
settling system have resulted in effective removal of metals and
suspended solids during the period reported.
Zinc/Copper Mine/Mill 3101
This mine/mill facility is located in Maine. Ore mined from
underground contains sphalerite and chalcopyrite (also minor
amounts of galena). Zinc and copper concentrates are produced in
the mill by the flotation process. In 1973, mine production
totaled 209,000 metric tons (231,000 short tons) of ore. Zinc
and copper concentrate production from the mill totaled 25,600
metric tons (28,200 short tons). Operations were suspended at
this facility in October 1977 due to the depressed copper and
zinc markets.
For the most part, mine water was used in the mill flotation
circuit. Mill tailings and any mine water not used in the mill
were discharged to a primary tailing pond having an area of about
20.2 hectares (50 acres). Decant from this pond flowed into an
auxiliary pond, approximately 3.2 hectares (8 acres) in area, to
a pump pond approximately 0.81 hectare (2 acres) in area, and was
discharged. The pH of the final discharge was continually moni-
tored and adjustments were made to optimize removal of metals
(especially zinc, iron, and manganese), and to maintain the pH
within limits specified by state and federal permits.
Tailings discharged from the mill flotation circuits had a pH in
the range of 9.9 to 11.7. This was largely due to the use of
lime as a depressant in the zinc flotation circuit. Additional
lime was occassionally added to the tailings. On weekends, when
the mill was not operating, lime was added to the excess mine
water, which was discharged to the tailing pond system. During
the coldest months of the year (January, February, and March),
problems were encountered with maintenance of the final effluent
pH within the required 6 to 9 range. During this period, the 30-
day average pH is often as high as 10.7. For this reason, no
lime, other than that, used in the mill flotation circuits, was
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added to either the tailings or the excess mine water during
these months.
Because available mine water did not provide the total volume of
water required in the mill, part of the treatment system effluent
was recycled. Approximately 56 percent of the mill feed water
was obtained in this manner.
Other wastewater control technologies included the segregation of
spills from the copper and zinc flotation circuits and control of
surface drainage with ditches and surface grading.
Effluent data submitted by the company for the period of January
1974 to August 1977 are summarized in Tables VIII-51 and VIII-52.
These data consistently demonstrate achievement of effluent
quality superior to that specified by BPT guidelines, with the
exception of pH. Severe pH excursions occur in the winter
months, and this phenomenon is not clearly understood.
Comparison of Tables VIII-51 and VIII-52 illustrates the
improvements in water quality as it passed through a multiple
pond system. Note the reductions in the percentage of time the
quality is out of compliance at the tailing pond decant compared
to the final discharge. The merits of the multiple pond treat-
ment system are further substantiated by the reduced average
concentrations and variability illustrated by the data describing
the secondary pond effluent.
Wastewater treatment practices at Mine/Mill 3101 considered to be
important to consistent and reliable attainment of a high quality
effluent are:
1. Maximum utilization of mine water in the mill flotation
circuits, thus minimizing wastewater flows to be treated
2. Supplemental lime addition (after flotation) for optimum
metals precipitation
3. Use of the multiple pond system for improved sedimenta-
tion conditions and improved system control
4. Provision of ponds of sufficient size (volume) to pro-
vide adequate sedimentation conditions and long-term storage
capacity
5. Segregation and recycle of spills and washdown water in
the mill
6. Combined treatment of mine and mill wastewater streams
for improved metals removal
Lead/Zinc Mine/Mill 3102
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This facility is located in Missouri and produces the largest
output of lead concentrate and the second largest output of zinc
concentrate in the United States. Approximately 1,482,000 metric
tons (1,635,000 short tons) of ore are mined annually at this
facility, with sphalerite, galena, and chalcopyrite as the pri-
mary ore minerals. In 1975, 228,600 metric tons (252,100 short
tons) of lead concentrate and 41,600 metric tons (45,900 short
tons) of zinc concentrate were produced at the flotation mill.
Wastewater treatment consists of alkaline sedimentation of
combined mining and milling wastewater streams in a multiple pond
system. The exclusive use of mine water as the process and
potable water supply for the mill reduces the hydraulic loading
percent. Since the mine produces more water than 'the mill can
use, the excess mine water is discharged to the tailing pond for
treatment.
The mill slime tailings are discharged to the main tailing pond
after separation (by hydrocyclones) of the sand fraction for dam
building. The tailing pond now occupies about 32.4 hectares (80
acres) and will occupy 162 hectares (400 acres) when completed.
The decant from this pond flows into a small stilling pool, then
through a series of shallow meanders, to a polishing pond of
approximately 6.1 hectares (15 .acres), and is subsequently
discharged.
A summary of effluent monitoring data for the period of December
1973 through September 1974 is presented in Table VIII-53. These
data indicate that all parameters analyzed are several orders of
magnitude lower than BPT limitations.
Zinc/Lead/Copper Mine/Mill 3103
This facility is located in Missouri and
The minerals of principal value are
chalcopyrite. Zinc, lead, and copper
by the flotation process in the mill. I
totaled 972,300 metric tons (1,072,400
metric tons (102,000 short tons) of lead
metric tons (10,800 short tons) of
produced at the mill.
has an underground mine.
galena, sphalerite, and
concentrates are produced
n 1976, mine production
short tons), while 92,400
concentrate, and 9,800
copper concentrate were
Mine and mill wastewater streams are combined for treatment at
this facility, as indicated in Figure VIII-10. Wastewater treat-
merit consists of alkaline sedimentation in a multiple pond
settling system.
Wastewater discharge volume is minimized by the extensive use of
mine water and tailing pond recycle as flotation makeup water in
the mill. Combined influent flow to treatment averages 10,900 m3
(2.88 million gallons) per day, of which 5,450 m3 (1.44 million
gallons) per day are recycled when the mill is operational.
Throughout most of the year, the lime added to the flotation
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circuit is considered (by plant personnel) to be sufficient to
produce a wastewater pH high enough for effective heavy metals
removal. However, during cold winter months, as much as 0.9 to
1.8 metric tons (1 to 2 short tons) of additional lime are added
daily to the mill tailings to suppress rising heavy metal
concentrations, especially zinc, in the final effluent.
Summaries of effluent monitoring data are presented in Tables
VIII-54 and VIII-55. These summaries are based solely on
analytical data provided by industry. These data reveal several
important points relative to the treatment system performance and
capabilities at Mine/Mill 3103:
1. The effluent from the secondary settling pond (Table
VIII-55) was far below BPT limitations (monthly mean),
sometimes by an order of magnitude for all control
parameters (pH, TSS, lead, zinc, copper, cadmium, mercury,
and cyanide) for the period February 1974 through November
1977.
2. The tailing pond effluent was in compliance with BPT
limitations (monthly mean) for pH, TSS, and copper 100
percent of the time during the same 46-month period. Only
two of the 40 observations, or 5 percent, were above the
limitations for both lead and zinc (not necessarily in the
same sample).
3. Both the mean and the standard deviation (variability)
of all metals data were significantly less in the effluent
from the second settling pond than in the effluent from the
tailing pond.
The factors contributing to the effluent quality attained at
Mine/Mill 3103 are:
1. The multiple pond treatment system;
2. Extensive use of mine water and tailing pond recycle in
the mill;
3. Combined treatment of excess mine water, concentrate
thickener overflow, and mill slime tailings; and
4. Supplemental lime addition for metals removal when
necessary.
Lead/Zinc Mine/Mill 3104
This facility is located in northern New York State. Ore mined
from an underground mine contains sphalerite and galena. Zinc
and lead concentrates are produced by the flotation process in
the mill. Mine production was 1,009,100 metric tons (1,110,000
short tons) of ore in 1973, while the mill produced 113,100
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metric tons (124,400 short tons) of lead and zinc concentrates
that year.
Approximately 6,820 m3 (1.8 million gallons) of wastewater per
day are treated by alkaline sedimentation at this facility. The
tailing pond has a total impoundment area of 32 hectares (80
acres). This area is divided into three ponding areas, which are
4 hectares (10 acres), 15 hectares (37 acres), and 13 hectares
(33 acres) in area, respectively. Recent modifications at this
operation include partial recycle of treated effluent during
summer months and the use of all mine water as mill feed, thus
eliminating mine water discharge,
Table VIII-56 summarizes tailing pond effluent data for this
treatment system for the period January 1974 to December 1977.
An examination of these data indicates that total metal values
are well within the BPT limits even when the maximum values
reported are considered. The TSS concentrations average approxi-
mately 7 mg/1, with a maximum reported monthly value of 16 mg/1.
The factors contributing to the effluent quality attained at
Mine/Mill 3104 are:
1. Maximum utilization of mine water in the mill, which
reduces the volume of wastewater requiring treatment, and
2. A tailing pond configuration designed to minimize short
circuiting, with provision of adequate impoundment volume to
promote effective sedimentation.
These practices have eliminated the discharge of mine water and,
thus, reduced the total volume of wastewater to be treated and
discharged. Although the pH attained in the tailing pond is not
considered to be optimum for metals removal, the alkalinity of
the mill tailings is sufficient to reduce dissolved metals to
levels consistently better than BPT limitations without supple-
mental lime addition or extensive pH control.
Zinc Mill 3110
This flotation mill is located in central New York and benefi-
ciates an ore which contains sphalerite and pyrite as the major
minerals in a dolomitic marble. Minor constituents of lead,
cadmium, copper, and mercury are also present. In 1976, the mill
recovered 118,000 metric tons (13,000 short tons) of zinc concen-
trate from 93,900 metric tons (103,300 short tons) of ore. An
average of 830 m3 (220,000 gallons) per day of wastewater is
pumped from the mine to the mill water supply reservoir for use
as mill makeup water. The mill water supply is augmented by
other fresh water sources as required. The mill discharges 990
tailing deposition area. Mill water flows over and percolates
through the deposited tailings and is collected in a 3.2-hectare
(8 acre) settling pond. Decant from this pond flows by gravity
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into a 1.2 hectare (3 acres) pond, followed by a third 8.1 hectare
(20 acre) settling pond. (The third pond was not constructed by
the operators, but exists due to a beaver dam.) Due to the
influence of surface drainage, daily discharge volume from the
treatment pond system averages 2,650 cubic meters (650,000
gallons) per day.
Table VII1-57 presents a summary of company monitoring data for
the period January 1974 to April 1977. These data represent grab
samples collected once monthly for 40 months. All parameters
analyzed were well below BPT limitations throughout the
monitoring period.
Molybdenum Mine 6103
This operation, located in Colorado, is a recently opened
underground mine yielding molybdenum ore at the rate of
approximately 2,200,000 metric tons (2,425,000 short tons) per
year. A discharge of 9,100 m3 (2.4 million gallons) per day is
treated by spray cooling, and suspended solids are removed in a
multiple pond system with the aid of flocculants prior to
discharge. The mill which recovers molybdenite by flotation, is
located some distance from the mine and is connected to the mine
by a long haulage tunnel. Extensive recycle is practiced, and
there is no wastewater discharge at the mill site.
Table VIII-58 summarizes the limited data provided by the company
for the period July 1976 to June 1977. In general, TSS
concentrations are well below 20 mg/1. Effluent metal values are
reduced substantially below BPT limitations.
Molybdenum Mill 6101
This facility, which uses the flotation process to concentrate
molybdenum ore, is located in mountainous terrain in New Mexico.
Ore is obtained from a large open-pit mine, with production at
5,700,000 metric tons (6,300,000 short tons) per year. The
flotation mill produces an alkaline tailings discharge which
flows approximately 16.1 kilometers (10 miles) to the tailings
disposal area, where sedimentation in primary and secondary
settling ponds takes place. The average discharge volume from
this treatment system is 11,000 cubic meters (4.6 million
gallons) per day.
summarizes effluent monitoring data for the period
through December ^976. Values reported are
below BPT limitations. Recently, this operation
peroxide addition to the tailing pond decant
cold and inclement weather for the control of
cyanide discharges on an experimental basis. The effect of this
treatment, not reflected in the data presented in Table VIII-59,
is, according to mine personnel, the reduction of cyanide
Table VIII-59
January 1975
substantially
used hydrogen
stream during
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concentrations from approximately 0.05 mg/1 to less than 0.02
mg/1.
Molybdenum Mine/Mill 6102
This facility is located in Colorado and employs both open pit
and underground mining methods. Approximately 14,000,000 metric
tons (15,400,000 short tons) of ore containing molybdenum, tungs-
ten, and tin are processed each year. The ore is beneficiated at
the site by a combination of flotation, gravity separation, and
magnetic separation methods to produce concentrates of molybde-
num, tungsten, and tin.
A daily average of 3,800 cubic meters (1 million gallons) of mine
water is pumped from the underground workings to the mill tailing
ponds. Three tailing ponds receive the mill tailings discharge,
and most of the clarified effluent is recycled to the mill. The
system of tailing ponds, impoundment, and extensive recycle has
been used to achieve zero discharge throughout most of the year.
Heavy snowmelts flowing to the treatment system have necessitated
a discharge during the spring of most years. Extensive runoff
diversion works have been installed to reduce spring discharge
volume. The treatment system includes ion exchange for
molybdenum removal, electrocoagulation flotation removal of heavy
metals, alkaline chlorination for the destruction of cyanide, and
mixed media filtration. A continuous bleed through this
treatment system will replace the previous seasonal discharge to
limit the required capacity and, thus, the capital costs.
Full scale operation of the treatment system described above was
initiated during July 1978. This treatment system is designed to
treat 7.6 cubic meters (2,000 gallons) per minute; however, at
the date of sampling, the system had been operated at only 3.8
cubic meters (1,000 gallons) per minute. The following
discussion of this treatment system reflects its performance
during the first four months of its operation.
The treatment facility houses all the aforementioned unit pro-
cesses and is located below the series of tailing ponds. Feed
for the system is a bleed stream from a final settling pond whose
characteristics are presented in Table VIII-60.
The wastewater is treated first in an ion exchange unit (pulsed
bed, counter-flow type) to remove molybdenum. This ion exchange
unit uses a weak-base amine-type anion exchange resin for optimum
molybdenum adsorption. The influent is acidified to
approximately pH 3.5, since molybdenum adsorption is reported to
be most efficient at a pH in the range of 3.0 to 4.0 (Reference
67). Initial results indicate that an influent molybdenum
concentration of 5.6 mg/1 is reduced to 1.3 mg/1 in the ion
exchange effluent. Molybdenum recovery from the eluant
(backwash) has not been practiced to date. When the system is
optimized, molybdenum recovery is planned. However, several
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problems with the columns (most notably, excessive pressure at
flow exceeding 3.8 cubic meters, or 1,000 gallons, per minute)
have impeded the assessment of the actual treatment capability of
this unit process.
The ion exchange effluent is treated by electrocoagulation
flotation for removal of heavy metals. This process involves the
formation of a metal hydroxide precipitate (by addition of lime),
which is then conditioned in an electrocoagulation chamber via
contact with hydrogen and oxygen gases, generated by electrol-
ysis. The resulting slurry is mixed with a polymer flocculant
and floated in an electroflotation basin by small bubbles of
oxygen and hydrogen. The floated material is skimmed off and
discarded. To date, the effluent from this process has been
monitored only for TSS, iron (total), and cyanide. The extent to
which these parameters have been removed by the electrocoagula-
tion flotation process is indicated by the following:
Concentration (mg/1)
Influent to Effluent from
ParameterElectrocoagulation Electrocoagulation
TSS
Fe (Total)
Cyanide
127
1 .8
0.09
65
0.5
0.04
Total system effluent monitoring data indicate that effective
removal of zinc and manganese is also attained (refer to Table
VIII-60). Efficient dewatering and handling of the sludge which
results from this process have not been optimized and this
problem has not been resolved.
Effluent from the electrocoagulation flotation process is treated
by alkaline chlorination for destruction of cyanide and then
polished by mixed-media filtration prior to final discharge. The
sodium hypochlorite used for the alkaline chlorination is
generated on-site by the electrolysis of sodium chloride. The
hypochlorite is injected into the waste stream prior to the
filtration step. The first four months of data indicate that
influent cyanide levels (clear pond bleed) range from less than
0.01 to 0.20 mg/1 while the treatment-system effluent
concentrations of cyanide range from less than 0.01 to 0.04 mg/1.
After the treatment plant effluent passes through a final
retention pond (residence time of approximately 2 hours), the
cyanide levels are consistently below 0.01 mg/1. The retention
pond was added to the system to ensure adequate contact time for
the oxidation reaction to occur. Since the system is still in
the process of optimization, it is expected that dosage levels
for the hypochlorite will be optimized, and that possible
problems with high levels of residual chlorine will be
eliminated.
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Mixed-media filtration was incorporated into the treatment scheme
to provide effluent polishing for optimum removal of suspended
solids and metals.
In spite of difficulties which have been encountered the overall
performance of the treatment system has been good (refer to Table
VIII-61). Plant personnel report that the effectiveness of the
treatment system at this time has generally exceeded their
expectations based on pilot plant experience.
Aluminum 0_re Subcategory
Open-pit Mine 5102 is located in Arkansas and extracts bauxite
for metallurgical production of aluminum. Approximately 900,000
metric tons (approximately 1,000,000 short tons) of ore are mined
annually at this site. A bauxite refinery which produces alumina
(A12_0:3) in a variety of forms and which recovers gallium as a
byproduct is located on site, but no wastewater from the refining
operation is contributed to the mine water treatment system.
Bauxite mining at this operation occurs over a large expanse of
land, and several mines may be worked at one time. Because of
the long distance between mine sites, several mine water
treatment plants have been constructed. There are three mine
water discharge points averaging 1.0,900 cubic meters (2.8 million
gallons) per day, 14,100 cubic meters (3.7 million gallons) per
day, and 7,000 cubic meters (1.9 million gallons) per day,
respectively. The associated wastewater treatment systems
consist of lime addition and settling. Monitoring data for each
of the discharges are presented in Tables VIII-62 through VIII-
64. Each of the three discharges consistently meets BPT daily
average and monthly maximum total suspended solids
concentrations.
Mine 5101 is an open pit mine located adjacent to Mine 5102 in
Arkansas. Bauxite is mined at this facility for the production
of metallurgical aluminum. Approximately 900,000 metric tons
(1,000,000 short tons) of ore are mined yearly. The ore is
hauled directly to the refinery. There are presently three
active discharge streams with separate treatment systems employ-
ing similar treatment technologies. Lime addition and settling
are used to treat the acid mine drainage of Mine 5101. Portable,
semi-portable, and stationary treatment systems are all currently
being used at this mine. Monitoring data for each of the dis-
charges are presented in Tables VIII-65 through VIII-67. Each of
the discharges consistently met BPT limitations for total sus-
pended solids and aluminum during the monitoring period.
Tungsten Ore Subcategory
Tungsten Mine/Mill 6104
This operation is located in California in mountainous terrain at
elevations of 2,400 to 3,600 meters (8,000'to 11,000 feet). A
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complex tungsten, molybdenum, and copper ore is mined at the rate
of 640,000 metric tons (700,000 short tons) per year. A large
volume of mine water, 38,000 m3 (10 million gallons) per day,
flows by gravity from the portal of this underground mine.
Approximately 20 percent of this flow is used in the mill. The
remainder is treated for suspended solids removal in a clarifier
and discharged to a stream. The mill at this site uses several
stages of flotation to yield concentrates of molybdenum and
copper, and a tungsten concentrate which is further processed by
leaching and solvent extraction to yield purified ammonium
paratungstate. All mill effluent flows to a series of three
tailing ponds which have no surface discharge.
Table VlII-68 summarizes treated mine water effluent monitoring
data for the period from Jaunary 1976 through December 1976. As
the data show, this effluent contains extremely low
concentrations of most pollutants. The treatment system provides
effective control of TSS and, consequently, of most metals
present in the effluent. Molybdenum occasionally occurs at
measurable concentrations.
Uranium Ore Subcateqory
Uranium Mine 9408
This operation recovers uranium from a hard-rock, underground
mine in Colorado. The principal uranium mineral found in the
vein-type deposits is pitch blende, -in association with carbo-
nates and pyrite. The ore contains an average of 0.3 percent
U3_08_ and must be shipped approximately 200 kilometers (190
miles) to the associated mill. Therefore, it is crushed and
sorted on-site to increase grade. The ore finally shipped
contains an average of 0.6 percent U3_08_.
Approximately 3,500 cubic meters (940,000 gallons) per day of
mine water and a small volume of sanitary wastes are combined and
directed to the wastewater treatment plant. Treatment consists
of chlorination with sodium hypochlorite to disinfect the sani-
tary wastes, coagulation with an anionic polymer, barium chloride
coprecipitation for radium removal, and settling. Settling takes
place in a series of two concrete-lined basins and four ponds
with a combined capacity of 4,700 m3 (1,250,000 gallons).
Settled solids are periodically removed and trucked to the asso-
ciated mill for recovery of residual uranium and subsequent
disposal in mill tailings.
A summary of company reported effluent monitoring data is pre-
sented for the period April 1975 through January 1977 in Table
VI-69. All parameters are well below BPT limitations. The data
demonstrate correlation between control of suspended solids and
total Ra 226.
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Concern has been expressed over the applicability and efficiency
of this treatment system to mine water from the more common
sandstone deposits. However, similar facilities treating mine
water from sandstone deposits are achieving effective removal of
radium 226 also.
Nickel Ore Subcategory
Nickel Mine/Mill 6106
This facility is located in Oregon and produces ferronickel
directly by smelting 3,401,000 metric tons (3,746,500 short tons)
per year of lateritic ore from an open-pit mine. Mine area run-
off, ore and belt wash water, and smelter wastewater are combined
and treated in a series of two settling ponds. A considerable
volume of water is recycled to the smelter from the second of
these ponds, and surface discharge from the third pond occurs
intermittently, depending on seasonal rainfall.
Available monitoring data submitted by the company for the period
of January 1976 through December 1980 are summarized in Table
VIII-70. Because discharge from the ponds is intermittent, the
data represent the quality of surface water in the final settling
pond from which discharge occurs.
Figure VIII-11 is a plot of the concentrations of selected
effluent constitutents versus time. These data illustrate the
importance of seasonal meteorological conditions to many
facilities. At this site, mine runoff during the rainy season
(approximately November through April) significantly increases
flow through the settling ponds, thus reducing residence time,
adversely affecting secondary settling and increasing the
concentration of TSS in the effluent.
Vanadium Ore Subcategory
Vanadium Mine 6107
Mine 6107 is an open pit vanadium mine located in Arkansas.
Opened in 1966, this mine annually produces approximately 363,000
metric tons (400,000 short tons) from a non-radioactive vanadium
ore.
Mine area runoff, waste pile runoff, and seepage from the waste
pile are collected and treated in a common system. Mine area and
waste pile runoff are diverted to the wastewater treatment plant.
Seepage from the waste pile is collected in several small ponds
and pumped to the treatment plant. At the treatment plant, lime
is mixed with the wastewater to adjust the pH to within the range
of 6.0 and 9.0. Wastewater from the treatment plant flows into a
large settling pond which was formerly an active pit. Depending
upon the water quality, the effluent from the settling pond is
either discharged or recycled to the treatment plant.
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A summary of discharge monitoring data from Mine 6107 between.
July 1978 and December 1980 is presented in Table VIII-71, Rela-
tively low concentrations of suspended solids were consistently
reported. The average total iron concentration was 0.65 mg/1.
The pH ranged from 5.4 to 9.3 with only two excursions from the
range of 6.0 to 9.0.
Titanium Ore Subcateqory
Titanium Mine/Mill 9906
Mine/Mill 9906 is a titanium dredge mining and milling operation
located in Florida and adjacent to titanium Mine/Mill 9907.
Ilmenite ore from a placer deposit is dredged from a man-made
pond. Humphrey spirals located on a floating barge behind the
dredge are used to concentrate the heavy minerals in the ore.
The lighter minerals are returned directly to the dredge pond.
Electrostatic and magnetic separation methods are utilized to
further concentrate the ilmenite.
Excess mine water, runoff, and mill wastewater from the caustic
pond overflow are combined and treated in a common system. The
first step in the treatment process consists of lowering the pH
to approximately 4.0 with a strong acid to assist in coagulation
of the organic material. The wastewater then flows through a
series of settling ponds, after which the pH is adjusted upward
to meet discharge limitations. The wastewater then flows through
a series of small ponds before final discharge.
A summary of reported effluent monitoring data is presented in
Table VIII-72. The average discharge rate was approximately
26,000 cubic meters (6.85 million gallons) per day. The average
TSS concentration was less than 10 mg/1. The pH concentration
ranged from 4.0 to 10.0 and averaged 7.0 with few excursions.
ADDITIONAL EPA TREATABILITY STUDIES
Copper Mill 2122
Tailings from bulk copper flotation circuits located in two
copper mills at this facility are discharged to a 2,145~hectare
(5,300-acre) tailing disposal area for treatment. Due to the
design and mode of operation of this tailing disposal area, the
effluent quality attained is often very poor. Wind disturbances,
short-circuiting of the settling pond (the area actually covered
by standing water is 101 hectares, or 250 acres and the depth of
water is only a few centimeters over much of this area), and a
floating-siphon effluent system that at times pulls solids off
the bottom of the pond are all factors which frequently produce
high total suspended solids-concentrations in the tailing pond
effluent. For this reason, the unit processes investigated at
this facility were flocculant (polymer) addition, flocculation,
secondary settling, and filtration. Lime addition was also
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investigated to determine possible benefits derived in terms of
metals removal. Experiments employing various combinations of
these unit processes were designed primarily to evaluate improve-
ments in treatment efficiency attainable by addition of polishing
treatment to the existing tailing pond system.
The treatability study at Mill 2122 was performed during two
different time segments. These time periods were 5 to 15
September 1978 and 8 to 19 January 1979. During the September
phase of the study all of the unit processes identified above
were investigated. The purpose of the final phase was to further
investigate the capabilities of dual-media filtration for removal
of TSS and metals from the tailing pond decant at this site in
addition conducting cyanide destruction studies.
Company personnel at Mill 2122 have previously reported that
cyanide concentrations in the tailing pond decant are high enough
to cause problems only during the winter months (i.e.,. December
to March). For this reason, the second treatability study at
Mill 2122 was scheduled for early January which was expected to
be an optimum time to conduct cyanide destruction studies. How-
ever, the concentration of cyanide in the decant remained very
low (i.e., less than 0.05 mg/1) throughout the January study
period. For this reason, it was decided to spike the tailing
pond decant with cyanide prior to the cyanide destruction experi-
ments. Two unit processes, alkaline chiorination and ozonation
were investigated for cyanide destruction capabilities.
Initially, four species of cyanide (i.e., calcium cyanide, sodium
cyanide, ferrocyanide, and ferricyanide) were used for spiking,
independently of one another, to investigate the impact of the
chemical form of cyanide on the destruction technology
capabilities. However, experiments with ferricyanide and
ferrocyanide were discontinued after quality control results
indicated almost no analytical recovery of cyanide from control
samples spiked with these species and analyzed by the EPA
approved Belack distillation method. All samples collected for
cyanide analysis were analyzed within 24-hours by a local
commercial laboratory.
Influent to the pilot plant was taken from the tailing pond
effluent line. This line is used for recycle as well as for dis-
charge. The character of the tailing pond effluent during the
periods of study is presented in Tables VIII-73 and VIII-74. As
can be seen from these tables, the concentrations of total sus-
pended solids and total metals in the recycle water were highly
variable during the period of the study. The consistently low
concentrations of dissolved metals observed indicate that metals
present in the tailing-pond discharge (i.e., recycle water) were
contained in the suspended solids. This is further evidenced by
the high correlation (r = 0.99) between total copper and TSS
concentrations in the wastewater (see Figure VIII-12). This
relationship suggests that any polishing treatment which effec-
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tively removes the suspended solids will also effectively remove
the metals.
Results of the pilot-scale treatability studies are presented in
summary form in Tables VIII-75, VIII-76, VIII-77, and VIII-78. A
review of the results presented in Table VIII-75 indicates that
secondary settling at a theoretical retention time of 10.4 hours
was sufficient to produce effluent total metal concentrations
well below BPT limitations. In a full scale system, even longer
times (24 to 72 hours) would be recommended to reliably achieve
this limit. A larger pond would also provide protection against
surge loads and short-circuiting.
Polymer and lime addition, followed by flocculation prior to
settling (2.8 hour retention time), produced effluent suspended-
solids concentrations comparable to those achieved by secondary
settling with a longer retention time (10.4 hours). The observed
improvement in efficiency of removal of TSS at shorter retention
time is attributed to the addition of polymer. This is further
evidenced by the fact that treatment schemes employing lime addi-
tion and settling resulted in much higher suspended solids levels
in the effluent when a polymer was not employed.
Experiments employing lime addition were conducted to investigate
its effect on dissolved metal precipitation and suspended solids
settleability. However, because dissolved metal concentrations
in the tailing pond recycle water were already very low (i.e.,
less than 0.04 mg/1), this treatment provided little or no
benefit.
Three dual-media, downflow, pressure filters consisting of
different filter-media sizes and depths, were evaluated over a a
range of hydraulic loadings of 117 to 880 mVm2/day (2 to 15
gpm/ft2). All three filters employed consistently produced
filtrates with suspended solids concentrations of less than 10
mg/1 throughout the range of 30 to 50 mg/1. On two occasions,
however, filter performance was adversely impacted by shock
loads. At these times, the suspended solids concentrations of
the tailing pond recycle water being treated ranged upwards to
several percent solids, and the filtrate concentrations attained
were 13 and 30 mg/1.
Because dual media filtration at hydraulic loadings of 293 to 880
mVm2/day (5 to 15 gpm/ft2) demonstrated consistently good
removal of suspended solids during eight-hour runs, it was
desired to investigate filter performance at a high hydraulic
time. The results of this experiment are presented in Table
VIII-75. As indicated, the TSS concentration of the tailing pond
decant averaged 33 mg/1 during this experiment. Total suspended
solids concentrations of 7 to 12 mg/1 were attained in the filter
effluent during the first four hours of the run. However, solids
breakthrough began to occur between the 4th and 7th hours of the
run. Therefore, at a hydraulic loading of 13 gpm/ft2 the
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frequency of backwash required appears to be much greater than
the frequency required for a loading of 10 gpm/ft2 or less.
Subsequent filtration experiments planned during January were
long runs at 5 and 10 gpm/ft2 to determine the frequency of
backwashing required. An experiment at a loading of 4 gpm/ft2
was initiated, but was terminated after one hour because solids
breakthrough occurred almost immediately due to extremely high
concentration of TSS (i.e., 1,200 mg/1) in the tailing pon'd
decant being filtered. The use of dual-media filtration for
effluent polishing is generally effective when the influent TSS
concentrations are no greater than 35 to 50 mg/1. However, at
the very high TSS concentrations which frequently occurred at
Mill 2122 filtration is not feasible. Because high concentra-
tions of TSS persisted in the tailing pond decant during the
remainder of the final study period, no further filtration
experiments were attempted.
To investigate alkaline chlorination a series of bucket tests
were conducted to maximize the number of dosages, pH values and
contact times which could be employed over a short period of
time. As previously mentioned, meaningful results were not
obtained from initial experiments in which ferricyanide or
ferrocyanide were used as spikes due to the lack of quantitative
analytical recovery of these cyanide species from untreated spike
samples. For this reason, experiments with these cyanide species
were discontinued.
Results of bucket tests in which sodium cyanide was used to spike
the wastewater are summarized in Table VIII-70. These data
indicate the most efficient destruction of cyanide occured at the
highest hypochlorite dosages employed, i.e., 20 and 50 mg/1. At
these dosages, significant differences between the various pH
levels and contact times employed were not evident. At the lower
hypochlorite dosages, 5 and 10 mg/1, good destruction of cyanide
appeared to be achieved at pH 9. However, these data must also
be viewed with caution as a quality control program conducted in
conjunction with the sodium cyanide spike experiments indicated
erratic and unreliable analytical recoveries (see Section V).
During the alkaline chlorination study the destructability of
total phenol (4AAP) present in the tailing pond decant was
observed. The data presented in Table VIII-70 indicate that
effective destruction of total phenol (4AAP) occurred only at the
highest hypochlorite dosage, 50 mg/1.
The literature indicates that oxidation of phenolic compounds
with chlorine species may produce highly toxic chlorophenols.
Although the production of these compounds was not investigated
during this study, their potential production should be evaluated
if full scale alkaline chlorination of a phenol containing waste
stream is seriously considered.
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Experiments to evaluate the destruction of cyanide by ozonation
were conducted using a continuous flow pilot scale treatment
system. Variables evaluated during the ozonation experiments
included the weight ratio of ozone to cyanide maintained to the
contact chamber, pH, and contact time. The results of these
experiments are presented in Table VIII-78. These results
indicated that destruction of cyanide occurred to varying degrees
in the contact chamber. At ratios of 5:1 or greater, pH and
contact time did not appear to be significant factors. Again,
however, these results must be viewed with some caution due to
the cyanide analytical problem mentioned previously (also see
Section V of this report).
The results for destruction of total phenol (4AAP) by ozonation
are also presented in Table VIII-78. These results do not reveal
a definite trend, although the most effective removal was
indicated at the highest ozone dosages, i.e., 8 and 24 mg
To summarize, the treatability study conducted at Mill 2122
demonstrated the effectiveness of polishing technologies (i.e.,
secondary settling or dual-media filtration) for removal of sus-
pended solids when the initial TSS concentration was in the range
of 30 to 50 mg/1. Much higher TSS concentrations often occur in
the tailing pond decant at Mill 2122 due to the manner in which
the pond is operated and effluent is withdrawn. Under conditions
of high TSS loading, secondary settling with the use of a floccu-
lating aid (i.e., polymer) would be a more practical effluent
polishing technology than filtration. Lime addition provided
little benefit. However, the effective removal of TSS by the
polishing technologies also resulted in effective removal of
metals. Cyanide was apparently destroyed by both hypochlorite
and ozone. At high dosages of these oxidants, total phenol
(4AAP) were also removed.
CONTROL AND TREATMENT PRACTICES
Control and Treatment of_ Wastewater ajt Placer Mines
Placer mining sites generally have limited area available for
construction of treatment facilities. In addition, the lifetime
of a given mining site is generally very short (1 to 5 years).
However, as mining methods improve and economics of gold recovery
become more favorable, the same area may be remined several times
by different miners. The BPT effluent limitations governing the
placer mining of gold were reserved because of insufficient
economic data (Reference 68). A discussion of control practices
at placer mines using gravity separation processes is presented
here.
Placer mining consists of excavating waterborne or glacial
deposits of gold bearing gravel and sands which can be separated
by physical means. This separation is classified as gravity
separation milling (reference Section IV). Since many placer
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deposits are deeply buried, bulldozers, front-end loaders, and
draglines are being used for overburden stripping, sluicebox
loading, and tailing removal operations. However, where water
availability and physical characteristics permit, dredging or
hydraulic methods are often favored based on cost.
Gold has historically been recovered from placer gravels by
purely physical means. Gravity separation is accomplished in a
sluicebox. Typically, a sluicebox consists of an open box to
which a simple rectangular sluiceplate is mounted on a downward
incline. A perforated metal sheet is fitted onto the bottom of
the loading box, and riffle structures are mounted on the bottom
of the sluiceplate. These riffles may consist of wooden strips,
or steel or plastic: plats which are angled away from the
direction of flow in a manner designed to create pockets and eddy
currents for the collection and retention of gold.
During actual sluicing operations, pay gravels (i.e., goldbearing
gravels) are loaded into the upper end of the sluicebox and
washed down the sluiceplate with water, which enters at right
angles to (or against the direction of) gravel feed. Density
differences allow the particles of gold to settle and become
entrapped in the spaces between the riffle structures, while the
less dense gravel and sands are washed down the sluiceplate.
Eddy currents keep the spaces between riffle structures free of
sand and gravel, but are not strong enough to wash out the gold.
Wastewater from placer mining operations consists primarily of
the process water used in the gravity separation process.
Recovery of placer gold by physical methods generally involves no
crushing, grinding, or chemical reagent usage. As a result, the
primary waste parameters requiring removal are the suspended
and/or settleable solids generated during washing (i.e.,
sluicing, tabling, etc.) operations.
Arsenic is present at relatively high concentrations in some of
the sediments being mined by placer methods. However, this
arsenic occurs primarily in particulate form and can be removed
by effective settling prior to discharge of the wash (sluice)
water.
Current best treatment practice in this segment of the industry
is the use of a dredae pond or a sedimentation pond. In some
instances, the discharge of wastewater through old tailings
achieves a filtering effect. The treatment effectiveness
achieved by selected placer mining operations using this
technology is indicated in Table V1II-79. Data provided here are
documented in Reference 69, "Evaluation of Wastewater Treatment
Practices Employed at. Alaskan Gold Placer Operations" (July
1979).
Most of the over 250 active placer mines are located in Alaska.
Some have estimated the actual number of placer mines at over
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500. EPA Region X has issued National Pollution Discharge
Elmination System (NPDES) permits to many placer miners in the
State of Alaska which identify required settling pond designer
effluent limitations, as indicated by the following (excerpted
from a Region X NPDES permit):
"a. Provide settling pond(s) which are designed to contain
the maximum volume of process water used during any one day's
operation. Permittee shall design single and/or multiple ponds
with channeling, diversions, etc., to enable routing of all
uncontaminated waters around such treatment systems and also to
prevent the washout of settling ponds resulting from normal high
water runoff. Choice of this alternative requires no
monitoring."
or
"b. Provide treatment of process wastes such that the fol-
lowing effluent limitations be achieved. The maximum daily
concentration of settleable solids from the mining operation
shall be 0.2 milliliter of solids per liter of effluent. This
shall be measured by subtracting the value of settleable solids
obtained above the intake structure from the value of settleable
solids obtained from th.e effluent stream."
Few (if any) placer mining operations have ponds "which are
designed to contain the maximum volume of process water used
during any one day's operation." The actual retention capacity
of the few existing settling ponds or pond systems at placer
mining operations is typically two hours or less. However, as
indicated in Table VIII-79, many of the operations which have
installed settling ponds are producing an effluent which contains
less than 1.0 ml/1/hr of settleable solids. Reductions of
suspended solids attained at the operations are highly variable
because of different flows, particle size distribution, working
hours per day, etc.
Two practices were identified at placer mining operations. The
first is the use of any screening device which effectively
classifies (size separation) the paydirt prior to washing. The
second is the use of multiple settling ponds.
The practice of screening greatly reduces the volume of water
required for washing by eliminating the need for great hydraulic
force to move large rocks and boulders through the sluice box.
This increases retention time and improves settling conditions
within a given settling pond by reducing the volume of wastewater
requiring treatment.
The use of a number of smaller ponds in series appears to be more
effective than a single larger pond at any given site.
Generally, a limited area is available to placer miners for
construction of a settling pond. As a result, ponds are
298
-------
generally small in size relative to the volume of wastewater to
be treated. Therefore, it is not unusual that these ponds are
severely short circuited. For this reason, the use of a number
of small ponds in series serves to reduce hydraulic surges,
offset short circuiting and reduce the velocity of flow, thereby
improving conditions for removal of settleable solids. Also,
some miners use sluice box tailings to construct dikes between
several ponds in series. In passing from one pond to another,
the wastewater must filter through these dikes. This practice
provides very effective removal of settleable solids in most
instances.
Multiple-settling pond systems have been used at placer Mines
4114, 4133, 4136, 4138, 4139, 4140, and 4141. Screening devices
to classify paydirt prior to washing or sluicing have been used
at placer mines 4133, 4136, 4138, and 4141. As indicated in
Table VIII-79, all of the placer mines which employ multiple
ponds and screening were capable of producing a treated effluent
having less than 1.0 ml/l/hr of settleable solids. Mine 4142
also employs two ponds in series; however, these ponds were being
short circuited and, as a result, were not as effective as they
could have been.
A report prepared for the State of Alaska, Placer Mining
Wastewater Settling Pond Demonstration Project, confirms that in
theroy and practice, for settling placer mine wastewater
discharges, an effective holding time of four hours of quiescent
settling will reduce settleable solids to below detectable
levels. For a pond to provide the equivalent of four hours
quiescent settling, the pond generally must be designed for a
holding time of more than four hours.
Control of. Mine Drainage
It is a desirable practice to minimize the volume of water
contaminated in a mine because the volume to be treated will be
less. Best practices for mine drainage control result from
careful planning and assessment of all phases of mining
operations. Mining techniques used, water infiltration control,
surface water control, erosion control, and regrading and
revegetation of mined land are all essential considerations when
planning for mine drainage control. In the past, inadequate
planning resulted in a significant adverse impact on the environ-
ment due to mining. In many instances, extensive and costly
control programs were necessary.
The types of mining operations (planned or existing) used to
recover metal ores differ in many respects from those of the coal
mining industry. This is important to note when considering the
information available on mine drainage control in these indus-
tries. Mine drainage problems in the coal industry appear to be
more widespread than those in the metal ore mining and dressing
category. This is primarily because of the number of mines
299
-------
involved, geographic location, age, disturbed area, and geology
of the mined areas. There is an abundance of literature
describing the problem of mine drainage from both active and
abandoned coal mines. The discussions which follow present the
limited available information on mine drainage control in metal
ore mines. However, references to practices employed in coal
mining operations which may be applicable to metal ore mining are
also presented.
Water-Infiltration Control
Diversion of water around a mine site to prevent its contact with
possible pollution forming materials is an effective and widely
applied control technique. Flumes, pipes, ditches, drains, and
dikes are used in varying combinations, depending on the geology,
geography, and hydrology of the mine area. This technique can be
applied to many surface mines and mine waste piles.
Regrading, or recontouring, of some types of surface mines, and
surface waste pile can be used to modify surface runoff, decrease
erosion, and/or prevent infiltration of water into the mine area.
There are many techniques available, but they are highly
dependent on the geography and hydrology of the land and the
availability of cover or fill materials. This practice, along
with the establishment of a stable vegetative cover, is currently
being used experimentally at one eastern metal ore mine to
decrease erosion and stabilize soil on an abandoned waste pile.
Use of regrading techniques at the larger open-pit mines may be
limited only to the disturbed area surrounding the pit or to
stabilization of some steep slopes.
Mine sealing techniques and procedures for sealing boreholes and
fracture zones are more frequently applied to inactive or
abandoned mines. Internal sealing by the placement of barriers
within an underground mine can be used in an active mine with
caution. Mine sealing practices are used either to prevent water
from entering a mine or to promote flooding of an abandoned mine
to decrease oxidation of pyritic materials. No data on the use
or efficiency of mine sealing techniques in the metal ore mining
and dressing industry were available for use in this report.
Control Practices in the Ore Mining and Dressing Industry
Most of the metal-ore mines examined in this report (both
underground and open-pit) practice some measure of mine drainage
control. These practices involve controlled pumping of mine
drainages and application of a variety of treatment technologies,
or use in a mill process. Use of mine water as makeup water in
mill circuits is a desirable management practice and is widely
implemented in this industry. In many areas of the West, water
availability is limited, and water conservation practices are
essential for mine/mill operations. Mine water which has been
300
-------
adequately treated is suitable for discharge to surface waters,
and this practice is also common to this industry.
Regrading and revegetation of areas disturbed by mining is
practiced at some operations, but is primarily directed at
stabilization of tailing areas and, in some instances, waste or
overburden piles. Documentation of the use and effectiveness of
these practices is limited to uranium mining at this time.
Prevention o£ Control of_ Seepage from Treatment Ponds
Uranium mill wastewater is characterized by very high salinity
and the presence of radioactive parameters. Therefore, at least
four western states either have requirements or are developing
requirements for seepage control to protect limited groundwater
supplies (Reference 70).
Under certain conditions, unlined tailing or settling ponds may
represent an acceptable level of environmental control for
disposal of uranium mine water or milling wastewater (Reference
70). With proper siting, ponds could, in some instances, be
located to take advantage of the properties of native soils in
mitigation of the adverse effects of seepage. Many uranium
deposits and milling facilities, however, are not located where
the natural soils provide sufficient uptake of waterborne
pollutants and prevention of contamination of groundwater.
Seepage rates and soil uptake of pollutants depends on the soil's
chemical and physical properties, the design and construction of
the pond itself, and the geological conditions prevalent at each
site. Unlined ponds are best used under the following
circumstances:
1. Deep groundwater table and/or soils exhibiting permea-
bilities sufficiently low to minimize the volume of seepage,
2. Native soils with significant capacity to remove and fix
pollutants from seepage,
3. Arid climates, and
4. Geological and hydrological conditions at the site pre-
cluding contamination of aquifers or other bodies of water
which are useful as water supplies.
The seepage rates from unlined uranium mill ponds depend upon the
characteristics of the tailings, soils, underlying geoglogy, and
hydrologic conditions prevalent at the site. Soils and tailing
deposits exhibiting high permeabilities may permit high seepage
rates, especially if sandy soils underlie the pond. If ponds are
located on soils containing high proportions of natural clay or
on impervious rock (such as shale), seepage rates can be reduced
301
-------
substantially. Permeability values as low as 10~6 to 10~8 cm/sec
(down to 0.006 gal/min/acre) can often be achieved under these
circumstances.
Pond Liner Technology
Prevention of seepage from impoundment systems can be achieved by
the use of liners. Pond liners fall into two general categories:
natural (clay or treated clay) and synthetic (commonly, polyvinyl
chloride (PVC), polyethylene (PE), chlorinated polyethylene
(CPE), or Hypalon).
Pond liners installed to date have usually been in new ponds
which are used only for evaporating mill wastewater. Lining of
tailing disposal ponds has not been practiced to a great extent
for the following reasons:
1. Tailing ponds are usually larger than evaporation ponds.
Large investments must be made for lining tailing ponds.
The cost for lining a tailing pond may account for 60 to 90
percent of the pond capital cost. Where liners are
installed and seepage is prevented, pond surface area must
be increased in order to evaporate the wastewater;
2. Thicker liners may be required for tailing ponds than
for evaporation ponds,-
3. Reliable information on the long term performance of
liners in tailing pond applications is lacking; and
4. Tailings themselves often prevent seepage as they are
deposited in the ponds.
Natural (Clay) Liners. Clays can be effectively used in sealing
ponds because of a layered structure and the ability of certain
clay minerals to exchange cations with wastewater seeping
through. Some clays, usually commercially identified as bento-
nite (montmori1lonite), absorb water molecules between layers,
resulting in a swelling of the clay structure. Under confined
conditions, such as the case of a pond liner, swelling will be
retarded, but the clay particles will be pressed tightly
together. The amount of space between the particles is reduced,
resulting in a decrease in permeability. The water which does
permeate the clay will lose cations by ion exchange, preventing
these contaminants from seepage into the groundwater.
According to Reference 70, the effective use of untreated clays
for seepage control is limited to situations where the liner will
be in contact with relatively fresh water. If high levels of
dissolved salts, strong acids, or alkalies come into contact with
the clay, ion exchange reactions between the wastewater and the
clay will take place. This may remove some of the heavy metal
ions present, but a loss of exchangeable ions from the clay
302
-------
results in a reduction of the swelling capacity of the clay and
in eventual failure of the seal.
To improve the swelling characteristics of natural clays and make
them more effective as pond liners, a clay may be treated with
polymeric materials. The use of treated clay improves the
sealing properties of the clay, and also permits a reduction in
the amount required compared to untreated clay. Although
complete containment of pond wastewater cannot be obtained with
clay liners, permeabilities as low as 10~6 to 10~8 cm/sec (0.6 to
0.006 gal/min/acre) are achievable with treated clays (Reference
70) .
Clay liners have the advantages of being easy to install with
commonly used machinery, and they are relatively inert to most
chemical constituents. One supplier of treated clay claims its
product is effective in sealing ponds containing up to 20 percent
dissolved solids. This product may be used in constructing a
liner for a new pond or may be used to control lateral seepage
through use of a slurry trench technique.
Commerical experience with treated clay liners is minimal. Mill
9446 uses a treated clay (variety unknown) to mitigate pond
seepage. Plans call for American Colloid Company (Skokie,
Illinois) to install a treated clay liner for a uranium project
in Colorado (Reference 70).
Synthetic Pond Liners. Synthetic pond liners-may also be used to
control seepage from uranium mill ponds. These types of liners
have an advantage over natural clay or treated clay liners,
because they possess much lower permeability values than
polyester reinforced Hypalon liners). Flexible synthetic liners,
however, exhibit several disadvantages also:
1. Performance is highly dependent upon the quality of the
foundation and substrate. Structural failures may result
from poor initial design, poor substrate compaction, seismic
disturbances, water buildup beneath the liner, inappropriate
liner choice, or poor installation technique;
2. They are susceptible to degradation due to the chemical
err-, ronment or exposure to the elements; and
3. They are more prone to puncture and tear during instal-
lation and may pose difficulties in field handling.
The most common synthetic liners used in the uranium industry for
pond seepage control are PVC, PE, CPE, and Hypalon (Reference
69). They are used, alone or in conjunction, in thicknesses of
0.25 to 1.5 mm (10 to 60 mils). Different materials exhibit
varying degrees of strength, flexibility, weatherability, and
resistance to chemical attack.
303
-------
No data are available on the long term performance of synthetic
liners in uranium mill pond applications. However, tests
conducted on these liners in sanitary landfill applications
indicate little loss of tensile strength or tear or puncture
resistance. Some increase in permeability has been noted, with
the thermoplastic types (CPE, PVC, and Hypalon) tending to swell
and soften.
Four currently operating uranium mills in the United States are
using synthetic liners for seepage control. Mills 9422 and 9456
use Hypalon liners with thicknesses of 1.5 and 0.91 mm (60 and 36
mil), respectively. Mills 9402 and 9404 have decant pond liners
constructed of 0.25 mm (10 mil) PVC bottoms and 0.51 mm (20 mil)
CPE dike slopes. The PVC/CPE liners carry 15- and 10-year
warranties, respectively.
It is common practice to cover synthetic liners with a layer of
native soil to protect the liner from sunlight and to prevent
damage to the liner from earthmoving equipment, if and when the
pond requires dredging. Heavy-duty liners are preferred in
uranium mill tailing pond applications, because they minimize
posible mechanical damage when the pond is cleaned, generally
have longer life and better aging properties, and are more resis-
tant to the rocky soils present at many uranium mill sites
(Reference 70).
Recently, two secured landfills in New York State and one in Ohio
have been constructed for disposal of hazardous industrial wastes
(Reference 71). These secured landfills make use of two layers
of low permeability, natural clay liners, with a synthetic liner
(reinforced Hypalon) placed in between. The Ohio facility is
also equipped with an external, underdrain-type, leachate
collection and monitoring system. Although this type of
installation is very expensive, it represents state-of-the-art of
liner technology. Where disposal of toxic liquids or solid
wastes is necessary or groundwater supplies must be protected,
this approach may represent the only viable alternative.
Other Seepage Control Methods
Other methods for mitigating seepage from uranium-mill ponds have
been used successfully in the United States and Canada. These
methods have been used to control both underseepage from mill
ponds and lateral seepage through tailing dams or permeable
subsoils.
Underseepage from an existing tailing pond at Mill 9401, located
in New Mexico, is being controlled by collection and recycle of
contaminated groundwater. Wells in the collection system pump
groundwater contaminated by pond seepage from 12- to 18-meter
(40- to 60-foot) depths back to the pond. Downgradient of the
collection wells, injection wells are used to pump well water to
dilute groundwater which might be contaminated by uncollected
304
-------
pond seepage (Reference 70). A system such as this can be
effective only when specific favorable subsurface conditions
prevail.
It is common practice to collect lateral seepage through existing
tailing dams in a catch basin or sump for subsequent disposal or
return to the pond (Reference 70). Visits to a number of
facilities as part of this study, as well as visits during
previous efforts, indicate that at least seven other mill
facilities practice some form of seepage collection. The system
in existence at uranium Mill 9401 is described above. Uranium
Mill 9402 (also located in New Mexico) collects seepage from the
tailing pond in a dam toe pond. Seepage occasionally appearing
in an adjacent arroyo, due to precipitation events, is collected
and pumped back to the pond.
Uranium Mill 9405 (an acid-leach facility) located in Colorado,
collects seepage from its tailing pond and overflow from
yellowcake precipitation thickeners and treats the combined waste
stream to remove radium 226 and TSS prior to discharge.
Copper Mill 2121 collects seepage from its tailing pond and
conveys it to a secondary settling pond, from which it is dis-
charged. Lead/zinc Mill 3103, located in Missouri, also collects
seepage and discharges it into a secondary settling pond. Mill
3123, located in Missouri, collects seepage at the toe of the
tailing dam and pumps it back to the tailing pond.
Gold Mill 4101 intercepts seepage in a collection sump and pumps
it back to the mill for reuse. This facility does not discharge
to surface waters. Gold Mill 4105 has recently designed and
installed a seepage collection system which takes seepage from
the base of its tailing dam and pumps a volume in excess of 0.76
The use of lateral seepage control methods, such as the slurry
trench technique, can be most effective when an impermeable layer
exists beneath the impoundment. Otherwise, lateral seepage may
escape containment by migrating through permeable materials under
the dam. The lateral seepage curtain should extend down to the
impermeable layer.
Lateral seepage of tailing pond water through the subsoil at a
uranium r; . 11 in Eastern Ontario, Canada, is controlled by a grout
curtair constructed of clay, bentonite, and cement (Reference
70). 7r,^ crrout slurry was injected into the subsurface alluvium,
dcvn co .-.n impervious bedrock layer, and forms an underground
barrier to the lateral flow of seepage to a nearby recreational
lake. Monitoring the concentrations of dissolved radium 226 in
the groundwater has demonstrated the effectiveness of the cur-
tain. The effectiveness of this method is attributed to
increased flow-path length and ion exchange with the montmorillo-
nite clay in the grout.
305
-------
TABLE VIII-1. ALTERNATIVES TO SODIUM CYANIDE FOR FLOTATION CONTROL
ORE
copper with
pyrite
copper/lead/zinc
with pyrite
zinc with
pyrite
pH
LEVEL
natural
10 to 12
natural
10 to 12
natural
10 to 12
DEPRESSANT
Sodium cyanide
Sodium sulfite
Sodium cyanide
Sodium sulfite
Sodium
monosulfide
Sodium cyanide
Sodium sulfite
Sodium cyanide
Sodium sulfite
Sodium
monosulfide
Sodium cyanide
Sodium sulfite
Sodium cyanide
(none)
Sodium
monosulfide
Sodium sulfite
DEPRESSANT USAGE*
kg/metric ton
ground ore
0.196
0.504
0.196
0.006
0.312
0.196
0.504
0.196
0.504
0.003
0.196
0.005
0.002
-
0.003**
0.504
Ib/short ton
ground ore
0.392
1.008
0.392
0.013
0.624
0.392
1.008
0.392
1.008
0.006
0.392
0.010
0.004
-
0.006
1.008
PERCENTAGE
RECOVERY
copper
62.0
56.6
60.9
65.4
67.5
73.6
74.6
58.2
74.2
73.4
IMA
NA
NA
NA
NA
NA
lead
NA
NA
NA
NA
NA
78.8
74.9
75.0
80.3
75.9
NA
NA
NA
NA
NA
NA
zinc
NA
NA
NA
NA
NA
NA
NA
NA
NA
NA
7.97
11.6
9.11
11.2
13.1
12.9
PERCENTAGE
DEPRESSION
iron
84.1%
81.9
82.6
69.5
66.4
85.7
87.3
87.4
83.9
84.2
93.3
91.2
92.6
91.9
91.4
91.2
zinc
NA
NA
NA
NA
NA
77.8
82.9
77.2
72.7
53.3
NA
NA
NA
NA
NA
NA
DEPRESSANT COST
per metric ton
ground ore
SO. 18
$0.14
SO. 18
< S0.01
$0.11
$0.18
$0.14
$0.18
$0.14
<$0.01
$0.18
-
<$0.01
$0.14
per short ton
ground ore
$0.16
$0.13
$0.16
<$0.01
$0.09
$0.16
$0.13
$0.16
$0.13
<$0.01
$0.16
-
< $0.01
$0.13
GO
o
Ol
Based on References 37 and 38
NA = not applicable
* Based on laboratory-scale experiments using 500 grams of ground ore.
**0.312 kg/metric ton ground ore not studied
-------
TABLE VIII-2.
RESULTS OF LABORATORY TESTS OF CYANIDE DESTRUCTION
BY OZONATION AT MILL 6102
pH
5.2
7.4
8.1
9.3
9.4
10.2
11.4
12.7
Final Cyanide Concentration
0.36
0.09
0.06
0.05
0.04
0.03
0.02
0.04
(mg/l)
Based on ozone dosage = 10 times stoichiometric;
15-minute contact time; and initial cyanide concentration of 0.55 mg/l.
307
-------
TABLE VIII-3. RESULTS OF LABORATORY TESTS AT MILL 6102 DEMONSTRATING
EFFECTS OF RESIDENCE TIME, pH, AND SODIUIV HYPOCHLORITE CONCEN-
TRATION ON CYANIDE DESTRUCTION WITH SODIUM HYPOCHLORITE
pH:
NaOC1 Concentration:
Residence Time:
30 minutes
60 minutes
90 minutes
8.8
10mg/l
-
-
-
20 mg/l
0.08
0.05
0.07
10.6
10 mg/l
0.04
0.03
0.04
20 mg/i
0.03
0.02
0.02
11.0
10 mg/l
0.03
0.03
0.03
20 mg/l
0.01
0.02
0.02
Initial cyanide concentration was 0.19 mg/l.
308
-------
TABLE VIII-4. EFFECTIVENESS OF WASTEWATER-TREATMENT ALTERNATIVES
FOR REMOVAL OF CHRYSOTILE AT PILOT PLANTS
TREATMENT METHOD
Sedimentation
Sedimentation plus
Sedimentation plus
Earth Filtration
Sedimentation plus
Earth Filtration
Mixed-Media Filtration*
Uncoated-Diatomaecous-
Alum-Coated-Diatomaceous-
FIBER CONCENTRATION (fibers/liter)
Raw Water
4x1012
4x1012
4x1012
4x1012
Treated Water
5x1011tto1x1011**
1 xlO9
3x106
1 x105
Source: Reference 44
* Dual-media filtration with a column containing 25 mm (1 in.) of
anthracite and 320 mm (12.5 in.) of graded sand.
t After 1 hour of sedimentation
** After 24 hours of sedimentation
309
-------
TABLE VIII-5. EFFECTIVENESS OF WASTEWATER-TREATMENT ALTERNATIVES
FOR REMOVAL OF TOTAL FIBERS AT ASBESTOS-CEMENT
PROCESSING PLANT
TREATMENT METHOD
Sedimentation (for 24 hours)
Sedimentation (for 24 hours) plus Sand Filtration
FIBER CONCENTRATION (fibers/liter)
Raw Water
o*
5x 103
Q*
5x 109
Treated Water
9.3 x 109t
Q**
3.2 x 10a
Source: Reference 44
Corresponding turbidities arc:
*620 JTU's
t 1.0 JTU's
** 0.38 JTU's
310
-------
TABLE VIII-6. EFFECTIVENESS OF WASTEWATER-TREATMENT ALTERNATIVES
FOR REMOVAL OF TOTAL FIBERS AT ASBESTOS, QUEBEC,
ASBESTOS MINE
TREATMENT METHOD
Mixed-Media Filtration
Uncoated-Diatomaceous-Earth Filtration
Coated-Diatomaceous- Earth Filtration
FIBER CONCENTRATION (fibers/liter)
Raw Water
1 x 109
1 x 109
1 x 109
Treated Water
3x107
3x10G
8x104
Source: Reference 44
311
-------
TABLE VIH-7. EFFECTIVENESS OF WASTEWATER-TREATMEISIT ALTERNATIVES
FOR REMOVAL OF TOTAL FIBERS AT BAIE VERTE, NEWFOUNDLAND,
ASBESTOS MINE
TREATMENT METHOD
Sedimentation
Sedimentation plus Dual-Media Filtration
Sedimentation plus Uncoated-Diatomaceous-
Earth Filtration
Sedimentation plus Alum-Coated-
Diatomaceous-Earth Filtration
FIBER CONCENTRATION (fibers/liter)
Raw Water
1x109(1 x 1011)
1 x 109(1 x 1011)
1x109
1 x 109
Treated Water
1 x 109(1 x 1010)
1 x 108 (1 x 109)
2x106
< 1 x 105
Source: Reference 44
Parentheses enclose results for a second sample.
312
-------
TABLE VIH-8. COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS FOR TOTAL FIBERS AND
CHRYSOTILE AT SEVERAL FACILITIES SURVEYED
FACILITY
4401
(Mine-Water Settling Pond)
4401
(Tailing Pond)
5102 (Mine-Water Treatment System)
5102 (Mine-Water Treatment System)
2122 (Tailing Pond)
2122 (Tailing Pond)
2122 (Tailing Pond)
2122 (Tailing Pond)
2122 (Tailing Pond)
2121
(Treatment System)
2120
(Tailing Pond'i
2120
(Tailing Pond)
2120
(Mine-Water Treatment System)
2117
(Treatment Plant)
TYPE OF
SAMPLING
S
S
S
V
S
V
S
V
V
S
S
V
S
S
INFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
3.8 x 107
7.1 x 1011
3.5 x 107
3.6 x 107
2.5 x1012
ND
6.1 x 1012
ND
ND
3.0 x 1011
1.2x1012
1.3 x 1013
4.6 x 107
2.5 x 108
CHRYSOTILE
1.1 x 107
1.1 x1011
5.5 x 106
5.5 x 106
4.3 x1011
ND
6.2 x1011
ND
ND
5.5 x1010
3.1 x1011
1.7 x 1012
1.8x106
5.5 x 106
EFFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
5.7 x107
2.1 x 109
1.4x 109
I.Ox 108
4.3 x 109
6.3 x 109
3.7 x107
2.7 x 108
1.3 x107
8.2 x 106
1.2x 109
7.8 x 10?
7.2 x 107
3.4 x 106
CHRYSOTILE
1.1 x106
1.8 x 108
2.0 x 108
3.3 x 106
6.7 x 108
<2.2x 105
8.2 x 106
<2.2 x 105
9.1 x 105
5.5 x 105
3.0 x 108
1.2 x 107
8.2 x106
<2.2x105
TREATMENT
REDUCTION FACTOR
TOTAL FIBER
-
>102
INC
INC
~103
-
~105
-
-
104-105
103
105106
-
~102
CHRYSOTILE
10
~103
INC
-
~103
-
~105
-
-
105
103
~105
-
>10
CO
i*
CO
S = screen sampling
V = verification phase
INC = increase
ND = no data
-------
TABLE VIN-8. COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS FOR TOTAL FIBERS AND
CHRYSOTILE AT SEVERAL FACILITIES SURVEYED (Continued)
FACILITY
2117
(Treatment Plant)
2117
(Tailing Pond)
1105
(Mine-Water Settling Pond)
1108 (Tailing Pond)
9202
(Tailing Pond)
6101 (Tailing Pond)
6101
(Tailing Pond)
3107
(Treatment System)
3121 (Tailing Pond)
3103
(Tailing-Settling Pond)
3101
(Tailing-Settling Pond)
3110
(Tailing Pond)
9905
(Settling Pond)
TYPE OF
SAMPLING
S
S
S
S
S
S
V
S
S
V
V
S
S
INFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
1.6 x 107
1.9 x 1011
1.6 x 107
2.3 x 1011
1.2 x 1012
3.8 x 1011
5.8 x 1011
2.2 x 1010
1.8 x 1011
2.1 x 1011
2.4 x 1010
9.0 x 1011
7.1 x 109
CHRYSOTILE
<6.8x 105
5.5 x1010
3.8 x 106
3.8x 1011
1.5x 1011
2.7 x1010
2.7x 1011
1.4 x 109
2.2 x1010
8.2 x1010
3.2 x 109
2.6 x 1011
1.1 x 109
EFFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
8.8x 106
9.2 x 106
4.2x 107
4.3x 107
7.7 x 108
3.3 x1010
8.7 x 107
4.1 x 108
1.6 x 109
9.9 x 106
1.9x 107
3.4 x 108
1.5x 108
CHRYSOTILE
5.5 x 105
4.4 x 105
3.8 x 106
4.1 x 106
5.7 x 107
2.0 x 109
9.7 x 106
4.1 x 107
<3.3x 105
1.1 x 106
2.7 x 106
2.4 x 107
1.3x 106
TREATMENT
REDUCTION FACTOR
TOTAL FIBER
<10
104-105
-
~104
103-104
10
~104
~102
102
104105
103
103
10
CHRYSOTILE
-
105
~ j
~105
103104
10
104105
~102
105
104
103
104
~103
CO
S = screen sampling
V = verification phase
INC = increase
ND = no data
-------
TABLE VIH-8. COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS FOR TOTAL FIBERS AND
CHRYSOTILE AT SEVERAL FACILITIES SURVEYED (Continued)
FACILITY
9402
(Mine-Water Treatment System)
9408
(Mine-Water Treatment System)
9408
(Mine-Water Treatment System)
9405
(Mill Settling Pond)
9405
(Mill Settling Pond)
9411
(Mine-Water Treatment System)
6104
(Mine-Water Treatment System)
TYPE OF
SAMPLING
S
S
V
S
V
S
S
INFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
1.2 x 108
1.6 x 109
1.4 x 108
2.9 x 108
1.0 x 108
2.3 x 109
7.7 x 106
CHRYSOTILE
5.2 x 106
1.9 x 108
1.5 x 107
2.3 x 107
<2.2x 105
1.1 x 108
<2.1 x 105
EFFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
4.3x 108
2.3x 109
7.3 x 106
1.2x 109
6.6 x 107
5.7 x 108
3.3x 107
CHRYSOTILE
5.3 x 107
2.0 x 108
<2.2 x 105
1.5 x 108
<2.2 x 105
2.7x 107
8.2 x 106
TREATMENT
REDUCTION FACTOR
TOTAL FIBER
-
-
10 102
INC
<10
<10
INC
CHRYSOTILE
INC
-
~102
INC
-
<10
INC
Co
K-J
en
S = screen sampling
V = verification phase
INC = increase
ND = no data
-------
TABLE VHI-9. EFFLUENT QUALITY ATTAINED BY USE OF BARIUM SALTS FOR
REMOVAL OF RADIUM FROM WASTEWATER AT VARIOUS URANIUM
MINE AND MILL FACILITIES
CO
t1
Ol
OPERATION
94031
Mills 94051
9405*8
9411
941 1*9
Mines 94121'6
9408
9408*
94523
AMOUNT
(mg/l)
OF BaC!2
ADDED
7.42
9.5
9.5*
5
10
10.4
55
55
45
RADIUM CONCENTRATIONS (picocuries/l)
BEFORE BaCI2 TREATMENT
Total
111 (±1.1)
15.9 (±1.6)
39.2 (±3.9)
35.4 (±0.3)4
56.9 (±5.7)
48.9 (±0.2)
123.6 (±1.5)7
142 (±14)
955
Dissolved
_
33.3 (±3.3)
15.5 (±2.0)5
60.2 (±6.0)
4.7 (±0.1)
37.7 (±0.3)4
120 (±12)
93.4
AFTER BaCI2 TREATMENT
Total
4.09 (±0.41)
0.0
5.05 (±0.5)
8.4 (±0.1 )4
<2
10.9 (±0.2)
2.1 (±0.23)7
1.12 (±0.11)
7.18
Dissolved
_
<2
0.2 (±0.1 )5
1.6 (±0.1)
0.6 (±0.1 )4
<0.9
<1
PERCENTAGE
OF RADIUM
REMOVED
Total
96.3
>93.7
87.1
76.3
>96
77.7
98.3
99
99.2
Dissolved
_
>93.9
98.7
_
66.0
98.4
>99
>99
1. Data obtained from single grab sampling and analysis (April, 1976).
2. Calculated value based on average flow and annual BaCI2 usage.
3. Includes ion exchange treatment; facility visited August 1978.
4. Data obtained from analysis of two grab samples (April, 1976).
5. Company data for February 1975 (Average of 12 grab samples).
6. Final discharge to dry watercourse.
7. Colorado Dept. of Health data for period January 1973 through February 1975 (Average of 24 samples analyzed for
"extractable" Ra 226).
8. Data obtained from composite of two grab samples representing two separate influent points (May, 1977).
9. Note that the dosage has doubled apparently enhancing the treatment system efficiency.
* Updated data obtained during sampling trips occuring April-May, 1977. All samples, unless otherwise indicated, are 24-hr
composites.
t Dosage rates are assumed to remain the same as previous rates.
( ) Parenthetical values indicate analytical accuracy.
-------
TABLE VIII-10. RESULTS OF MINE WATER TREATMENT* BY LIME ADDITION AT
COPPER MINE 2120
TREATMENT
PH
6.2
8.5
10.3
11.5
TREATMENT
PH
6.2
8.5
103
11.5
TOTAL METAL CONCENTRATION (mg/l)
Fe
11.6
0.45
0.12
0.17
Cu
0.26
0.10
0.04
0.07
Zn
15.0
0.25
0.56
0.30
Pb
0.01
<0.01
0.01
0.01
Cd
0.19
0.02
0.02
0.01
As
0.002
0.006
0.003
0008
Hg
< 0.0005
0.0006
<0.0005
< 0.0005
DISSOLVED METAL CONCENTRATION (mg/l)
Fe
7.2
0.05
0.03
0.03
Cu
0.25
0.03
0.03
0,04
Zn
14.6
0.14
0.10
0.23
Pb
<0.01
<0.01
<0.01
<0.01
Cd
0.19
0.02
0.02
0.01
As
-
-
-
Hg
-
-
'Bench-scale experiments; raw data not provided; a measure of effectiveness is obtained by comparison
to values shown at initial ph.
317
-------
TABLE VIII-11. RESULTS OF COMBINED MINE AND MILL WASTEWATER TREATMENT*
BY LIME ADDITION AT COPPER MINE/MILL 2120
SAMPLE
Mine water +
mill tailings
Mill tailings
(control)
SAMPLE
Mine water +
mill tailings
Tailings (control)
TREATMENT
PH
6.5
7.0
8.0
9.0
10.0
11.0
-
TREATMENT
PH
6.5
7.0
8.0
9.0
10.0
11.0
-
TOTAL METAL CONCENTRATIONS (mg/l)
Fe
0.35
0.06
0.05
0.05
0.05
0.10
0.06
Cu
1.09
0.33
0.06
0.04
0.05
0.04
0.09
Zn
22.8
5.4
0.29
0.09
0.06
0.04
0.06
Pb
0.01
<0.01
<0.01
0.01
0.01
<0.01
0.01
Cd
0.32
0.18
0.04
0.02
0.01
0.01
0.01
As
0.006
0.009
0.002
0.004
0.004
0.006
0.003
Hg
0.0006
< 0.0005
< 0.0005
< 0.0005
< 0.0005
0.0008
< 0.0005
DISSOLVED METAL CONCENTRATIONS (mg/l)
Fe
0.06
0.03
0.02
0.02
0.02
0.02
0.01
Cu
1.00
0.26
0.06
0.04
0.05
0.03
0.04
Zn
22.1
5.4
0.29
0.09
0.06
0.04
0.06
Pb
<0.01
<0.01
<0.01
<0.01
<0.01
<0.01
<0.01
Cd
0.31
0.18
0.04
0.02
0.01
0.01
<0.01
As
-
-
-
-
-
-
Hg
-
-
-
-
-
-
'Bench-scale experiments. 5 parts mine water:9 parts mill tailings.
Raw data not provided; a measure of effectiveness of treatment is obtained by comparison to values
shown at initial pH.
313
-------
TABLE VIII-12. RESULTS OF COMBINED MINE WATER + BARREN LEACH SOLUTION
TREATMENT* BY LIME ADDITION AT COPPER MINE/MILL 2120
1
SAMPLE
Mine +
barren
leach
water
oAlVlrLt
Mine +
barren
leach
water
TREATMENT
PH
6.1
7.7
9.8
11.5
TO c A TIV/I C M T
PH
6.1
7.7
9.8
11.5
TOTAL METAL CONCENTRATIONS (mg/l)
Fe
533.0
15.6
0.10
0.07
Cu
0.68
0.08
0.04
0.05
Zn
150.0
1.90
0.12
0.96
Pb
0.01
0.01
0.01
0.01
Cd
1.48
0.22
0.01
0.01
As
0.007
0.004
0.004
0.002
DISSOLVED METAL CONCENTRATIONS (mg/l)
Fe
532.0
14.0
0.10
0.04
Cu
0.68
0.03
0.03
0.04
Zn
150.0
1.90
0.12
0.83
Pb
0.01
0.01
0.01
0.01
Cd
1.47
0.21
0.01
0.01
As
-
'Bench-scale experiments. 5 parts mine water:1.5 parts barren leach solution
Raw data not provided; a measure of effectiveness of treatment is obtained by comparison to values
shown at initial pH.
319
-------
TABLE VIII-13. RESULTS OF COMBINED MINE WATER + BARREN LEACH SOLUTION +
MILL TAILINGS TREATMENT* BY LIME ADDITION AT COPPER
MINE/MILL 2120
SAMPLE
Mine water +
mill tailings +
barren leach water
Tailings (control)
SAMPLE
Mine water +
mill tailings +
barren leach water
Taiiings (control)
TREATMENT
pH
6.2
7.0
7.9
9.3
10.0
11.1
10.6
TREATMENT
pH
6.2
7.0
7.9
9.3
10.0
11.1
10.6
TOTAL METAL CONCENTRATIONS (mg/l)
Fe
191.0
75.7
0.59
0.14
0.09
0.05
0.03
Cu
0.48
0.08
0.06
0.05
0.05
0.04
0.04
Zn
86.0
18.0
50.0
0.08
0.12
0.05
0.05
Pb
0.01
0.01
0.01
0.01
0.01
0.01
0.01
Cd
0.70
0.33
0.04
0.01
0.01
0.01
<0.01
As
0.008
0.004
0.002
0.002
0.004
0.002
0.002
DISSOLVED METAL CONCENTRATIONS (mg/l)
Fe
175.0
68.2
0.13
0.07
0.06
0.03
<0.01
Cu
0.48
0.08
0.04
0.05
0.05
0.04
0.03
Zn
83.0
17.2
0.37
0.07
0.12
0.05
0.04
Pb
0.01
0.01
0.01
0.01
0.01
0.01
0.01
Cd
0.60
0.31
0.04
0.01
0.01
0.01
<0.01
As
-
-
-
-
* Bench-scale experiments. 5 parts mine water: 1.5 parts barren leach solution:9 parts mill tailings.
Raw data not provided; a measure of effectiveness of treatment is obtained by comparison to values
shown at initial pH.
320
-------
TABLE VIII-14. CHARACTERISTICS OF RAW MINE DRAINAGE TREATED DURING
PILOT-SCALE EXPERIMENTS IN NEW BRUNSWICK, CANADA
PARAMETER
pH*
Sulfate
Acidity (as CaCOg)
Cu
Fe
Pb
Zn
Suspended Solids
CONCENTRATION (mg/l)
MINE 1
Mean
2.6
10,100
6,511
10.0
1,534
3.9
1,158
172
Range
2.4 to 3.2
1,860 to 14,892
4,550 to 9,650
4.8 to 22.3
8.5 to 3,21 1
0.9 to 10.3
142 to 1,61 5
70 to 645
MINE 2
Mean
2.7
4,454
4,219
47.2
718
1.2
538
65
Range
2.3 to 2.9
2,354 to 7,290
2,600 to 7,000
24.3 to 76.0
350 to 1.380
0.3 to 3.2
390 to 723
10 to 190
MINE 3
Mean
3.0
1.121
746
19.4
77
1.3
114
31
Range
2.8 to 3.3
729 to 1.790
70 to 1,530
1.0 to 52.0
24 to 230
0.1 to 5.0
18 to 185
5 to 90
* Value in pH units
Source: Reference 54
321
-------
TABLE VIII-15. EFFLUENT QUALITY ATTAINED DURING PILOT-SCALE MINE-
WATER TREATMENT STUDY IN NEW BRUNSWICK. CANADA
MINE STREAM
EXTRACTABLE METAL (mg/l)
LEAD
ZINC
COPPER
IRON
COMPARISON OF EFFLUENT QUALITIES DURING ALL PERIODS OF STUDY -
OPERATING PARAMETERS VARIED
1
2
3
Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Basin-Settled
Sand-Filtered
0.18
(0.05 to 0.39)
0.18
(0.08 to 0.25)
0.12
(0.07 to 0.15)
0.35
(0.05 to 0.62)
0.26
(0.01 to 0.50)
0.31
(0.01 to 0.50)
0.13
(0.05 to 0.62)
0.11
(0.05 to 0.36)
0.10
(0.05 to 0.36)
0.41
(0.13 to 0.87)
0.23
(0.20 to 0.25)
0.26
(0.14 to 0.38)
0.52
(0.07 to 1.42)
0.26
(0.03 to 0.60)
0.28
(0.03 to 0.58)
0.64
(0.14 to 1.45)
0.29
(0.01 to 0.74)
0.21
(0.01 to 0.75)
0.04
(0.02 to 0.10)
0.03
(0.01 to 0.05)
0.03
(0.02 to 0.04)
0.06
(0.03 to 0.19)
0.03
(0.02 to 0.07)
0.03
(0.02 to 0.04)
0.10
(0.01 to 0.30)
0.05
(0.01 to 0.30)
0.04
(0.01 to 0.30)
0.30
(0.1 4 to 0.65)
0.16
(0.08 to 0.29)
0.23
(0.09 to 0.63)
0.54
(0.12 to 2.51)
0.20
(0.04 to 0.41)
0.11
(0.02 to 0.20)
0.47
(0.09 to 1.40)
0.26
(0.01 to 0.61)
0.22
(0.01 to 0.89)
COMPARISON OF EFFLUENT QUALITIES DURING PERIODS OF OPTIMIZED STEADY OPERATION
1
2
3
Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Basin-Settled
Sand-Filtered
0.18
(0.01 to 0.35)
0.21
(0.16 to 0.25)
0.15
(0.14 to 0.15)
0.44
(0.25 to 0.62)
0.29
(0.01 to 0.50)
0.29
(0.17 to 0.42)
0.15
(0.09 to 0.25)
0.11
(0.05 to 0.18)
0.08
(0.05 to 0.22)
0.33
(0.13 to 0.52)
0.29
(0.28 to 0.30)
0.39
(0.38 to 0.39)
0.45
(0.27 to 0.69)
0.22
(0.03 to 0.60)
0.15
(0.03 to 0.28)
0.35
(0.14 to 0.90)
0.22
(0.13 to 0.40)
0.12
(0.03 to 0.1 8)
0.04
(0.03 to 0.06)
0.04
(0.03 to 0.04)
0.03
(0.03 to 0.03)
0.05
(0.03 to 0.07)
0.03
(0.02 to 0.03)
0.03
(0.02 to 0.03)
0.06
(0.03 to 0.11)
0.07
(0.02 to 0.15)
0.03
(0.02 to 0.04)
0.19
(0.10 to 0.26)
0.18
(0.15 to 0.21)
0.20
(0.14 to 0.26)
0.42
(0.14 to 0.45)
0.17
(0.04 to 0.29)
0.13
(0.11 toO. 17)
0.26
(0.09 to 0.60)
0.24
(0.10 to 0.30)
0.14
(0.08 to 0.23)
Source: Reference 54
322
-------
TABLE VIII-16. RESULTS OF MINE-WATER TREATMENT BY LIME ADDITION AT
GOLD MINE 4102
WASTE STREAM
Raw mine
drainage
Treated mine
water after
adjustment of pH
with lime
PH
6.0
5.9
7.4
8.1
9.2
CONCENTRATION (mg/l) OF PARAMETER
Pb
Total
Dissolved
Dissolved
Dissolved
Dissolved
2.2
0.02
<0.01
0.01
<0.01
Cu
0.02
0.01
0.01
0.01
0.01
Zn
9.8
9.6
3.84
0.65
0.02
Fe
40
0.3
0.4
0.4
0.2
323
-------
TABLE VIH-17. RESULTS OF LABORATORY-SCALE MINE WATER TREATMENT
STUDY AT LEAD/ZiNC MINE 3113
UNIT TREATMENT
PROCESSES EMPLOYED
No treatment (control)
Lime addition to
pH 10, sedimentation**
Lime addition to
pH 11, sedimentation**
EFFLUENT CONCENTRATIONS ATTAINED*
(mg/l)
PH^|
7.0 to 7.5
10
11
Cu
N.A.
0.022 to
0.033
0.22
Pb
N.A.
0.05 to
0.12
0.05
Zn
20
0.1 25 to
0.19
0.095
Fe
60
N.A.
N.A.
Results shown are based on jar tests.
*AII metals concentrations are based on "total" analyses.
tValue in pH units
**Theoretical retention times were variable arid not always specified.
N.A. = Not analyzed
324
-------
TABLE Vlll-18. ERA-SPONSORED WASTEWATER TREATABILITY STUDIES CONDUCTED
BY CALSPAN AT VARIOUS SITES IN ORE MINING AND DRESSING INDUSTRY
SITE IDENTIFICATION
Mine/Mill 3 121 (Pb/ZN)*
Mine/Mi!) /Smelter/
Refinery 3107 (Pb/Zn)*
Mill 2122 (Cu)*
Mine3113(Pb/Zn!*
Mine 5102 (Al)
Miil 9401 (U)»*
Mill 9402 (U)*»
PERIOD OF STUDY
August 3-1 3, 1978
March 19 29, 1979
August 14-19, 1978
September 5-1 5,1 978
January 8 19. 1979
September 22-29, 1978
October 10-16, 1978
October 23-30, 1978
November 1-10. 1978
December 4-14, 1978
COMMENTS
Polishing Treatments (Filtration, Secondary
Settling), Lime Precipitation and Cyanide
Destruction Technology (Ozonation, Alkaline
Chiorination) Investigated
Polishing Treatments (Filtration, Secondary
Settling) Investigated
Polishing Treatments (Filtration, Secondary
Settling), Lime Precipitation and Cyanide
Destruction Technology (Ozonation, Alkaline
Chiorination) Investigated
Lime Precipitation, Aeration, Polymer Addition,
Floccuiation, Sedimentation, Filtration
Investigated
Polishing Treatments (Filtration, Secondary
Settling) and Lime Precipitation Investigated
Bench-Scale and Pilot-Scale Investigation of
IX, pH Adjustment with H2SO4, Ferrous
Sulfate Coprecipitation, Barium Chloride
Coprecipitation, Settling, and Filtration.
Treatment Scheme Employing Lime Addition,
Aeration, Barium Chloride Addition, Settling,
and Filtration Investigated
* Operations in base and precious metals subcategory
"Operations in uranium ore subcategory
325
-------
TABLE VIII-19. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT LEAD/ZINC MINE/MILL 3121
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
NUMBER OF
OBSERVATIONS
75
13
i
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
7.8
4.5
<0.5
0.001
<0.002
0.002
<0.01
0.10
0.21
0.0002
<0.02
0.002
0.01
<0.01
0.74
RANGE
6.7 - 9.1
1 - 10
<0.5
<0. 001-0. 025
<0.002
0.005-0.011
<0.01
0.02-0.16
0.18-0.25
<0. 0002-0. 0005
<0.02
<0. 002-0. 004
<0. 01-0. 05
<0.01
0.25-1.25
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
n.
\
a.
r
RANGE
n.a.
i
i
Based on observations made in period 6 through 10 August 1978.
n.a. = Not Analyzed.
326
-------
TABLE VIII-20. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO
PILOT-SCALE TREATMENT TRAILER) AT LEAD/ZINC MINE/MILL
3121 DURING PERIOD OF MARCH 19-29,1979
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd (total)
(diss.)
Cr
Cu (total)
(diss.)
Pb (total)
(diss.)
Hg
Ni
Se
Ag
T1
Zn (total)
(diss.)
Fe (total.)
(diss.)
CN
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
8.9
10
<0.10
<0.0020
-------
TABLE VIII-21. OBSERVED VARIATION WITH TIME OF pH AND COPPER AND
ZINC CONCENTRATIONS OF TAILING-POND DECANT AT LEAD/ZINC
MINE/MILL 3121
DATE
(1973)
August 6
August 7
August 7
August 8
August 8
August 8
August 8
August 8
August 9
August 9
August 10
August 10
August 10
POLLUTANT PARAMETER CONCENTRATION (mg/l)
pH*
6.9 to 7.4
6.9 to 7.7
7.3 to 7.7
6.7 to 7.7
7.6 to 7.8
7.4 to 7.6
7.2 to 7.6
7.1 to 7.4
8.2 to 8.3
8.2 to 8.3
8.5 to 8.6
8.5 to 8.9
8.6 to 9.1
Cu
0.004
0.05
0.02
0.08
0.08
0.09
0.11
0.14
0.14
0.12
0.16
0.14
0.13
Zn
1.1
1.2
1.3
0.94
0.75
0.76
0.79
0.80
0.55
0.44
0.49
0.28
0.25
COMMENTS
Mill not operating
Mill not operating
Mill not operating
Mill startup 4:30 PM, Aug. 7
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
"Value in pH units
328
-------
TABLE VIII-22. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH
PILOT-SCALE UNIT TREATMENT PROCESSES AT MINE/MILL 3121
DURING AUGUST STUDY
UNIT TREATMENT
PROCESS EMPLOYED
Secondary Settling (approx.
11- to 22-hr theoretical
retention time)
Lime Addition to pH 9.2, Polymer
Addition, Flocculation, Secondary
Settling (approx. 2.6-hr theoretical
retention time)
Lime Addition to pH 9.2, Polymer
Addition, Flocculation, Secondary
Settling (approx. 2.6-hr theoretical
retention time). Filtration
Lime Addition to pH 11.3, Polymer
Addition, Flocculation, Secondary
Settling (approx. 2.6-hr theoretical
retention time)
Lime Addition to pH 11.3, Polymer
Addition, Flocculation, Secondary
Settling (approx. 2.6-hr theoretical
retention time). Filtration
Filtration
Filtration
EFFLUENT CONCENTRATION* ATTAINED (mg/l)
pHt
8.2 - 8.5
9.2
9.2
11.3
11.3
7.4**
(6.7 to
7.8)
8.3**
(7.7 to
9.1)
TSS
3
17
1
n.a.
<1
<1»*
K1 to
2)
<1*»
(0-1)
Cu
0.11
0.05
0.02
0.03
0.02
0.02**
«0.01 to
0.04)
0.05**
(0,03 to
0.06)
Pb
0.10
0.08
0.04
0.05
0.06
0.09**
(0.03 to
0.12)
0.035**
(0.01 to
0.06)
Zn
0.24
0.38
0.16
0.13
0.08
0.61**
(0.24 to
1.1)
0.044**
(0.02 to
0.06)
* All metals concentraticsis are based on "total" analyses
^ Value in pH units
** Average concentration- attained
( ) Range of concentrations attained
n.a. = Not analyzed
329
-------
TABLE VIII-23. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH
PILOT-SCALE UNIT TREATMENT PROCESSES AT
MINE/MILL 3121 DURING MARCH STUDY.
UNIT TREATMENT
PROCESS EMPLOYED
Filtration
Lime Addition to pH 10.5,
Polymer Addition, Flocculation,
Secondary Settling (approx. 2.6-hr
theoretical retention time)
Lime Addition to pH 10.5,
Polymer Addition, Flocculation,
Secondary Settling (approx. 2.6-hr
theoretical retention time).
Filtration
EFFLUENT CONCENTRATIONS ATTAINED (mg/l)*
PH*
8.9-9.0
10.5
10.3
TSS
7
(3-10)
24
(20-29)
2
K1-3)
Cu
0.14
(0.12-0.16)
0.13
(0.11-0.14)
0.053
(0.040-0.070)
Pb
0.12
(0.050-0.22)
0.12
(0.060-0.16)
0.040
(0.030-0.040)
Zn
1.4
(0.96-1.8)
1.0
(0.52-1.4)
0.26
(0.040-0.48)
T pH Units
* Values given are mean and range, in ( ), concentrations
**AII metal concentrations are based on "total" analyses
330
-------
TABLE VII1-24. RESULTS OF OZONATION FOR DESTRUCTION OF CYANIDE
Initial CN concentration = 0.11
0, Dosage
mg/min Ratio 0,/CN
1.4
0.48
0.36
0.7
0
10:1
10:1
10:1
5:1
-
Retention Time
(min)
10
30
45
10
10
Final CN Concentration
(rng/1)
0.066
0.080
0.081
0.108
0.115
00
CO
-------
TABLE VIII-25. EFFLUENT FROM LEAD/ZINC MINE/MILL/SMELTER/
REFINERY 3107 PHYSICAL/CHEMICAL-TREATMENT PLANT
PARAMETER
pH**
TSS
Cd
Pb
Zn
Hg
CONCENTRATION img/D*
Average
8.78
15
0.16
0.15
4.5
0.0033
Range
8.5 to 8.9
8.3 to 26
0.044 to 0.58
0.08 to 0.26
1.8 to 8.5
0.0010 to 0.023
Based on industry data collected during the
period of December 1974 through April 1977.
*AII metals concentrations are based on "total"
analyses
**Value in pH units
332
-------
TABLE VIII-26. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT LEAD/ZINC MINE/MILL/SMELTER/REFINERY 3107
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
NUMBER OF
OBSERVATIONS
64
11
12
1
'
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
8.5
16
<0.5
0.0024
<0.002
0.12
0.010
0.031
0.13
0.0006
0.030
0.002
<0.01
0.012
2.9
RANGE
8.1 - 8.7
13 - 20
<0.5
0.0015-0.0030
<0.002
0.075 - 0.16
<0.010 - 0.010
0.020 - 0.045
0.090 - 0.17
0.0003-0.0012
<0.02 - 0.060
<0. 002-0. 003
<0.01
0.010 - 0.025
1.8 - 4.2
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
n.
i
a.
!
0.036
n.a.
0.021
0.073
n.
i
a.
r
0.055
RANGE
n.
i
a.
t
0.025-0.050
n. a.
0.020-0.030
<0. 02-0. 17
n.
\
a.
i
0.030-0.12
Based on observations made in period 14 through 19 August 1978.
n.a. = Not analyzed.
333
-------
TABLE VIII-27. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH PILOT-SCALE
UNIT TREATMENT PROCESSES AT MINE/MILL/SMELTER/REFINERY 3107
UNIT TREATMENT
PROCESS EMPLOYED
Secondary Sedimentation
Polymer Addition, Secondary
Sedimentation
Filtration
EFFLUENT CONCENTRATION' ATTAINED (mg/l)
PH*
7.8
8.1
8.5**
(8.1 to
8.7)
TSS
3
6
<1«*
«1)
Cd
0.065
0.060
0.035* *
(0.01 5 to
0.070)
Cu
0.020
0.015
0.016**
(0.01 to
0.02)
Pb
0.080
0.070
0.061**
(0.030 to
0.09)
Hg
n.a.
n.a.
n.a.
Zn
0.79
1.0
0.042**
(0.01 5 to
0.080)
* All metals concentrations are based on "total" analyses
1 Value in pH units
"Average concentrations attained
( ) Range of concentrations attained
n.a. Not analyzed
334
-------
TABLE VIII-28. CHARACTER OF DRAINAGE FROM LEAD/ZINC MINE 3113
PARAMETER
PH*»
TSS
TDS
so4 =
Cd
Cu
Fe
Pb
Ag
Zn
Hg
As
CONCENTRATION (mg/1)»
Average
4.2
111
1687
813
0.13
0.60
90
0.070
0.01
44
<0.001
0.021
Range
2.9 to 7.5
86 to 322
214 to 9,958
485 to 3,507
<0.01 to 0.50
0.18 to 1.6
3.6 to 522
0.01 to 0.35
<0.005to0.20
1.5 to 76
< 0.001
<0.01 to 0.040
Based on industry data collected during the
period of 1970 through 1978.
*AII metals concentrations are based on "total"
analyses
**Value in pH units
335
-------
TABLE VIII-29. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT LEAD/ZINC MINE 3113
POLLUTANT
PARAMETER
PH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
A9
T1
Zn
Fe
so4 =
NUMBER OF
OBSERVATIONS
53
7
1
i
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
3.2
112
<0.1
0.013
<0.003
0.23
0.011
1.5
0.088
<0.0002
0.074
0.006
n .a.
<0.05
71
69
1063
RANGE
3.0 - 3.3
104 - 124
<0.1
0.005 - 0.030
<0.003
0.22 - 0.24
0.010 - 0.015
1.3 - 1.6
0.033 - 0.12
<0.0002
0.060 - 0.090
*
0.003 - 0.010
n.a.
<0.05
67 - 74
50 - 80
925 - 1320
"DISSOLVED" CONCENTRATION
(mg/1)
MEAN
X
n .a.
n.a.
n. a.
n. a.
n.a.
0.23
n. a.
1.4
n.a.
n.a.
n.a.
n.a.
n.a.
n.a .
57
25
n.a.
RANGE
n.a.
n. a.
n.a.
n.a.
n.a.
0.23 - 0.24
n.a.
1.4 - 1.5
n.a.
n .a.
n.a.
n.a.
n.a.
n.a.
56 - 60
22 - 29
n.a.
Based on observations made in period 24 through 28 September 1978.
n.a. = Not Analyzed.
336
-------
TABLE VIII-30. SUMMARY OF PILOT-SCALE TREATABILITY STUDIES AT MINE 3113
CO
CO
I
EXPERIMENTAL
SYSTEM
A
B
C
D
E
F
G
H
I
J
UNIT TREATMENT
PROCESS EMPLOYED
'-;« . ..'iiWon to pH ~ 9.5, Sedimentation
Lime Addition to pH ~ 9.5, Sedimentation,
Filtration
Lime Addition to pH ~ 9.5, Aeration,
Sedimentation
Lime Addition to pH ~ 9.5, Aeration,
Sedimentation, Filtration
Lime Addition to pH ^9.5, Polymer
Addition, Flocculation, Sedimentation
Lime Addition to pH ~ 9.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration
Lime Addition to pH ~- 8.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation
Lime Addition to pH ~ 8.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration
Lime Addition to pH ~ 10.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation
Lime Addition to pH ~ 10.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration
EFFLUENT CONCENTRATSONS* ATTAINED (mg/l)
PH*
9.3*"
(9.1 to
9.7}
8.8"*
(8.4 to
9.0)
9.7**
(9.7 to
9.8}
9.5**
(9.4 to
9.6)
9.3**
(8.8 to
9.8)
8.9**
(8.1 to
9.5)
8.4**
(7.5 to
8.3)
7.9**
(7.3 to
8.3)
10.5**
(10.2 to
10.7!
10.2**
(9.5 to
10.5)
TSS
33
<2**
«1to
3)
35
1
10*'
(10)
<1**
«1)
6
<1**
KD
15
<1**
«1to
1)
Cd
0.025
0.016**
K0.005 to
0.033
0.02
0.005
0.015**
( 0.015)
0.009**
(0.005 to
0.015)
0.02
0,012**
(0.010 to
0.015)
0.005
<0.005**
K0.005)
Cu
0.10
0.02**
(0.01 1 to
0.030)
0.11
0.020
0.05**
( 0.05)
0.015**
(0.01 to
0.02)
0.02
<0.01**
«0.01 to
0.01)
0.02
0.013**
(0.01 to
0.015)
Pb
<0.02
<0.02**
K0.02)
0.02
<0.02
<0.02**
K 0.02 to
0.02)
0.025**
« 0.02 to
0.04)
0.08
<0.02**
« 0.02 to
0.02)
<0.02
-------
TABLE VIII-31. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT ALUMINUM MINE 5102
POLLUTANT
PARAMETER
PH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
Al
Fe
Phenol
NUMBER OF
OBSERVATIONS
(TOTAL/
DISSOLVED)
20
21
21
19
21
21
21
21
21
21
21
19
21
21
21/12
21/12
21/12
5
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
6.5
3
<0.2
<0.0005
<0.005
0.005
<0.01
<0.01
<0.05
<0.0002
<0.03
<0.002
<0.01
<0.03
0.01
0.49
0.15
0.007
RANGE
5.8 - 7.1
<1 - 5
<0.2
<0.0005
<0.005
<0. 005-0. 010
<0. 01-0. 010
<0. 01-0. 015
<0.05
<0.0002
<0. 03-0. 040
<0.002
<0. 01-0. 015
<0.03
0.005-0.020
<0.2-0.7
0.020-0.24
<0. 002-0. Oil
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
1
n
i
.a.
r
<0.01
<0.02
<0.02
n. a.
RANGE
n.
i
a.
F
<0. 01-0. 010
<0. 02-0. 020
<0. 02-0. 020
n.a.
Based on observations made in period 10 through 16 October 1978.
n.a. = Not Analyzed.
338
-------
TABLE Vlll-32. SUMMARY OF PILOT-SCALE TREATABILITY STUDIES AT MINE 5102
UNIT TREATMENT
PROCESS EMPLOYED
No Treatment (Control)
Lime Addition to pH ~ 8.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition to pH «« 8.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation,
Filtration
Lime Addition to pH~9.0,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition to pH~ 9,0,
Aeration, Polymer Addition,
Flocculation, Sedimentation,
Filtration
Lime Addition to pH «*10.4,
Aeration, Flocculation,
Sedimentation
Lime Addition to pH ~10.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition topH~10.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation,
Filtration
Filtration
EFFLUENT CONCENTRATION* ATTAINED
PH'
6.5**
(5.8 to 7.1)
8.2
8.0**
(7.7 to 8.4)
9.0
8.5
10.4
10.2
10.0**
{9.8 to 10.2)
6.5**
(5.8 to 6.8)
TSS
3»*
«1 to 5)
1
<1**
«1)
7
<1
34
5
< 1**
«1to1)
< 1**
«1 to!)
Al
0.49**
«0.2to0.70)
<0.2
<0.2**
«0.2)
<0.2
0.2
0.2
0.3
<0.27**
« 0.2 to 0.4)
<0.2**
«0.2to0.2)
Fe
0.15**
(0.020 to 0.24)
0.15
0.067**
(0.06 to 0.07)
0.06
0.08
0.2
0.23
0.073**
'(0.07 to 0.08)
0.040**
(0.2 to 0.10)
* All metals concentrations are based on "total" analyses
Value in pH units
"Average concentrations attained
( ) Range of concentrations attained
339
-------
TABLE VIII-33. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT MILL 9402 (ACID LEACH MILL WASTEWATERl
1
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
V
Mo
fe
Mn
A,
so4
TDG
R3226
U
PERIOD OF
OBSERVATIONS
(DATES)
11/3-8/1978
I
j
i
1
i
..
NUMBER OF
OBSERVATIONS
5
3/4
3/4
"TOTAL" CONCENTRATION
(mg/li
MEAN
X
I .7
40
<0.5
2.5
0.03
0.05
0.67
4.0
0.93
<0.0002
1 .4
2.0
<0.1
<0.2
6.1
100
10
1900
118
786
21760
...
99.7±i%
17
RANGE
1.6-1.9
24-48
-------
TABLE VIII-34. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER AT MILL 9402 (ACID LEACH MILL WASTEWATER)
POLLUTANT
PARAMETER
pH
TSS
c,
Cu
Pb
Ni
V
Mo
Fe
Mn
Al
so4
JDS
Ra226
U
PERIOD OF
OBSERVATIONS
(DATES)
12/5/78-
12/11/78
i
NUMBER OF
OBSERVATIONS
8
3
3
3
3
3
3
3
3
3
3
3
3
3
3
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
1.6
168
0.86
3.2
1.3
1.0
109
17
3,670
287
1,840
19,600
--
154.8±1%
19. 8
RANGE
1.4-1.8
142-215
0.73-0.98
2.9-3.6
1.1-1.5
0.95-1.1
97-110
15-20
3200-4100
260-310
1,640-2,220
18,400-20,400
-
(168)±1%
16-24.1
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
--
--
0.82
3.1
1.3
0.93
93
17
2,32f
157
1,330
--
32,200
129±1%
20±12%
RANGE
--
--
0.71-0.93
2.9-3.4
1.2-1.3
0.87-1.0
89-98
15-19
2,100-2,600
150-160
1,170-1,480
--
29,700-34,200
(miro-isst!
(16±12°O-(24±12
1
343
-------
TABLE VIII-35. SUMMARY OF WASTEWATER TREATABILITY RESULTS USING A LIME
ADDITION/BARIUM CHLORIDE ADDITION/SETTLE PILOT-SCALE
TREATMENT SCHEME
PARAMETER
TDS
Cu
Pb
Zn
Ni
Cr
Fe
Mn
Al
V
Mo
Ra226
U
IMH3
OPTIMUM CONDITIONS
FOR REMOVAL
pH of 9-9.5
pH9.5
pH 8.2-9.5
pH 6.8-9.5; Final TSS < 40
pH 5.8-9.5
pH 5.8-9.5
pH 8.2-9.5
pH 8.2-9.5
pH 5.8-9.5
pH 5.8-9.5
pH 5.8-6.1
BaCl2 dosage of 51-63 mg/l;
Final TSS < 200
Final TSS < 50
None attained
REMOVAL EFFICIENCY
(%)
42-78
91-98
52-90
98 to > 99
90-96
87-96
98 to > 99
92 to > 99
96 to > 99
97 to > 99
73
96-97
78-97
0-25
FINAL CONCENTRATION
ATTAINED (mg/l)
Total
-
0.11-0.18
<0.20
0.10
< 0.040
< 0.050
0.80-32
0.88-4.9
0.90-17
0.20-1.3
4.6
3.9-4.0*
0.30-2.5
123-292
Dissolved
5590-9740
0.060-0.15
< 0.014
< 0.020-0.030
< 0.040
< 0.040
0.10-1.0
0.434.2
0.50-5.0
£0.20
2.2
1.0-2.3*
0.20-0.50
-
Values in picocuries per liter (pc/l)
342
-------
TABLE Vlll-36. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT MILL 9401
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
Mo
V
U
Ra 226**
NUMBER OF
OBSERVATIONS
20
6
6
5/4
6
6
6
6
6
5
6
5/4
6
6
6
6
5
5/3
5
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
10.0
945
<0.50
4.6
<0.010
< 0.020
<0.050
0.060
<0.10
0.0002
<0.10
19
<0.10
<0.10
0.023
106
26
58.6
163
RANGE
9.9-10.1
156-1528
<0.50
4.0 - 5.0
<0.010
< 0.020
< 0.050
0.040 - 0.080
<0.10
0.0002
<0.10
17-20
<0.10
<0.10
< 0.020 - 0.030
95-110
24-27
55-63
30-677
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
-
-
-
4.25
-
< 0.020
-
-
-
-
-
19
-
-
<0.020
108
27
39
29
RANGE
-
-
-
4.0 - 5.0
-
< 0.020
-
-
-
-
-
16-20
-
-
<0.020
100-110
25-27
8-57
18-48
* pH units
**pCi/l
-------
TABLE Vlii-37. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH PILOT SCALE TREATMENT
SYSTEM AT MILL 9401
i (MIT TDC ATRVIPIVIT
PROCESS EMPLOYED
Fe (SO4) (1000 mg/l). Aeration,
Non-Ionic Polymer, BaCU
(60 mg/l), Flocculation,
Sedimentation, Filtration
IX, Fe (SO4) (330 mg/l).
Aeration, Non-Ionic Polymer,
Bad 2 (15 mg/l), Flocculation,
Sedimentation, Filtration
IX, Fe(S04) (500 mg/l),
Aeration, Non-Ionic Polymer,
BaCI2 (120 mg/l), Flocculation,
Sedimentation, Filtration
EFFLUENT CONCENTRATIONS ATTAINED (mg/D*
pH«*
10.0
9.0
8.0
TSS
1744
2.0
468
TDS
21700
24300
23200
so4-
9380
10500
9830
AS
<5.0
(d)
<5.0
(d)
<5.0
(d)
SE
9.6
(d)
8.5
(d)
8.5
(d)
Mo
90
55
55
V
8.0
9.0
5.0
U
59
±13%
6.0
±9%
0.72
±18%
Ra 226f
3.1
±2%
2.5
±6%
44
±7.3%
C..O
-£=>
* Metals are total metals unless otherwise indicated by a (d) = dissolved
**pH units
f PC1/I
-------
TABLE VIII-38. RESULTS OF BENCH-SCALE ACID/ALKALINE MILL WASTEWATER NEUTRALIZATION
CO
c*
THEORETICAL PARAMETER CONCENTRATION Img/ll
ACID/ALKALINE
MILL WASTEWATER
MIX RATIO
BY VOLUME
i (theoretical concentration* - (measured concentration)
t^enrfeticaf concentration
(100)
-------
TABLE VIII-39. CHARACTERIZATION OF INFLUENT TO WASTEWATER TREATMENT
PLANT AT COPPER MINE/MILL/SMELTER/REFINERY 2122
(SEPTEMBER 5-7, 1979)
POLLUTANT
PARAMETER
pH
TSS
Fe
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
Tl
Zn
CN
Tot. Phenolics
NUMBER OF
OBSERVATIONS
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
TOTAL CONCENTRATION
(mg/l)
MEAN
2.46
297
17
0.015
5.16
< 0.0005
0.092
0.146
9.43
5.56
0.014
0.160
0.036
0.028
< 0.004
1.73
-------
TABLE VIII-40. CHARACTERIZATION OF EFFLUENT FROM WASTEWATER
TREATMENT PLANT (INFLUENT TO PILOT-SCALE TREATMENT
PLANT) AT COPPER MINE/MILL/SMELTER/REFINERY 2122
(SEPTEMBER 5-10,1979)
POLLUTANT
PARAMETER
pH
TSS
Fe
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
Tl
Zn
CN
Tot. Phenolics
NUMBER OF
OBSERVATIONS
14
14
14
14
14
14
14
14
14
14
14
14
14
14
14
14
14
14
TOTAL CONCENTRATION
(mg/l)
MEAN
8.2
14.7
0.66
0.014
1.95
<0.0005
0.044
0.054
0.374
0.253
0.002
0.146
0.110
< 0.020
0.004
0.309
<0.02
0.41
RANGE
7.2-8.65
3.0-61.0
0.06-1.1
0.009-0.032
1.0-4.0
<0.0005
0.028-0.065
0.035-0.077
0.120-0.650
0.005-0.410
<0.001 -0.005
0.110-0.190
0.021-0.230
< 0.01-0.036
< 0.002-0.007
0.120-0.730
<0.02
0.02-5.2
DISSOLVED CONCENTRATION
(mg/l)
MEAN
-
0.04
< 0.01 3
1.6
< 0.0005
0.039
0.047
0.079
< 0.003
< 0.002
0.143
0.100
< 0.019
< 0.005
0.154
-
-
RANGE
-
-
0.03-0.06
<0.005-0.023
0.9-2.4
< 0.0005
0.021-0.070
0.029-0.067
0.018-0.210
< 0.002-0.005
< 0.001 -0.003
0.099-0.200
0.020-0.180
< 0.01-0.033
< 0.002-0.008
0.018-0.660
-
-
-------
TABLE VI11-41. SUMMARY OF TREATED EFFLULiVT QUALITY FROM PILOT-SCALE
TREATMENT PROCESSES AT COPPER MINE/MILL/SMELTER/REFINERY
2122 (SEPTEMBER 5-10, 1979)
TEST
NUMBER
01
02
03
04
05
06
07
08
09
TREATMENT APPLIED
Filtration*
0.15m3/min/m2
(3.6 gpm/ft2)
Filtration
0.26 m3/min/m2
(6.3 gpm/ft2)
Filtration
0.37 m3/-nin/m2
(9.1 gpm/ft2)
Filtration
0.43 m3/min/m2
(10.6 gpm/ft2)
Filtration
0.50 m3/min/m2
(12.2 gpm/ft2)
One hour settling test
Settling -45 min
Filtration with Lime
Addition to pH 9.1
0.38 m3/min/m2
(9.4 gpm/ft2)
Filtration with Lime
Addition to pH 9.0
0.38 m /min/m2
(9.3 gpm/ft2)
Filtration with Lime
Addition to pH 9.5
0.37 m3/min/m2
(9.5 gpm/ft2)
I
EFFLUENT CONCENTRATION (TOTAL, mg
pH*
7.2
7.6
8.1
8.1
8.3
8.5
8.95
8.95
9.05
TSS
4
3
2
2
2
8
<1
1
1
Cu
0.084
0.095
0.094
0.093
0.100
0.250
0.060
0.070
0.064
Pb
0.004
0.015
0.010
0.015
0.009
0.043
0.004
0.005
0.017
Zn
0.310
0.260
0.120
0.140
0.210
0.170
0.031
0.043
0.023
/I) I
Fe
0.09
0.08
0.06
0.07
i
I
s
0.06
0.54
0.04
I
0.04
0.07
Dual media filtration (anthrafilt and silica sand)
*Field pH values
34 8
-------
TABLE VIII-41. SUMMARY OF TREATED EFFLUENT QUALITY FROM PILOT-SCALE
TREATMENT PROCESSES AT COPPER MINE/MILL/SMELTER/REFINERY
2122 (SEPTEMBER 5-10, 1979) (Continued)
TEST
NUMBER
10
11
12
13
16
TREATMENT APPLIED
Filtration
5 mins
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration
6 hours
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration
12 hours
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration
18 hours
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration with Lime
Addition to pH 10.0
0.37 m3/min/m2
(9.1 gpm/ft2)
EFFLUENT CONCENTRATION (TOTAL, mg/l)
PH»
___
8.45
8.3
8.6
9.75
TSS
1
1
<1
1
2
Cu
0.058
0.073
0.044
0.110
0.056
pb
0.011
0.002
0.018
0.011
< 0.002
Zn
0.038
0.110
0.097
0.038
0.110
Fe '
0.06
0.07
0.04
0.03
0.03
Dual media filtration (anthrafilt and silica sand)
* Field pH values
349
-------
TABLE VIIU2. CHARACTERIZATION OF EFFLUENT FROM WASTEWATER
TREATMENT SYSTEM (INFLUENT TO PILOT-SCALE TREATMENT
PLANT) AT COPPER MINE/MILL/SMELTER/REFINERY 2121
(SEPTEMBER 18-19, 1979)*
POLLUTANT
PARAMETER
pH**
TSS
Fe
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
Tl
Zn
Tot. Phenol ics
NUMBER OF
OBSERVATIONS
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
2
TOTAL CONCENTRATION
(mg/l)
MEAN
7.5
4.1
0.27
< 0.003
< 0.002
< 0.0005
< 0.008
< 0.008
0.023
0.004
< 0.001
0.054
< 0.005
<0.01
< 0.003
0.015
0.01
RANGE
7.4-7.65
3.9-4.2
0.25-0.29
<0.003
<0.002
<0.0005
<0.005-0.013
<0.005-0.015
0.022-0.025
0.003-0.006
< 0.001
0.053-0.055
<0.005
<0.01
<0.003
0.011-0.019
0.007-0.013
DISSOLVED CONCENTRATION
(mg/l)
MEAN
-
0.088
<0.003
<0.006
<0.0005
<0.008
<0.008
0.022
0.006
<0.001
0.055
< 0.005
<0.01
< 0.003
0.032
-
RANGE
-
-
0.071-0.110
< 0.003
< 0.002-0.01 5
< 0.0005
<0.005-0.013
<0.005-0.013
0.021-0.023
0.005-0.009
<0.001
0.053-0.057
< 0.005
<0.01
< 0.003
0.018-0.049
-
* Grab samples
**pH units
350
-------
TABLE VIII-43. SUMMARY OF TREATED EFFLUENT QUALITY FROM PILOT-SCALE
TREATMENT PROCESSES AT COPPER MINE/MILL/SMELTER/REFINERY
2121 (SEPTEMBER 18-19, 1979)
TEST
NUMBER
01
02
03
TREATMENT APPLIED
Filtration
0.26 m3/min/m2
(6.5 gpm/ft2)
Filtration*
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration with Lime
Addition to pH 8.8
0.37 m3/min/m2
(9.1 gpm/ft2)
EFFLUENT CONCENTRATION (TOTAL, mg/l)
pH»
7.2
7.3
7.7
TSS
1.1
1.6
1.1
Cu
0.038
0.020
0.020
Pb
0.005
0.004
0.004
Zn
0.024
0.011
0.016
Fe
0.21
0.11
0.17
Dual media filtration (anthrafilt and silica sand)
* Field pH values
351
-------
TABLE VIII-44. HISTORICAL DATA SUMMARY FOR IRON ORE MINE/MILL 1108 (FINAL DISCHARGE)
,
PARAMETER*
pH
pH
TSS
TSS
Fe (dissolved)
Fe (dissolved)
--
MONITORING
PERIOD
Jan. 74 -Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74- Apr. 77
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
3 - 5/mo
1/mo.
(mo. ave.)
4-31/mo.
3 5/mo.
(mo. ave.)
3 - 5/mo
NUMBER OF
OBSERVATIONS
35
148
35
804
35
147
MEAN
7.1
(ave. of mo. means)
7.1
(daily ave)
5.9
(ave. of mo. means)
6.1
(daily ave.)
0.36
(ave. of mo. means)
0.35
(daily ave.)
STANDARD
DEVIATION
0.2
0.3
4.8
6.8
0.47
0.58
RANGE
6.7 - 7.7
6.5 - 8.0
2-28
1.0-70
0.05 - 1.92
0.01 -3.6
01
ro
-------
TABLE VISI-45. HISTORICAL DATA SUMMARY FOR COPPER/SILVER MINE/MILL/SMELTER/REFINERY 2121
(FINAL TAILINGS POND DISCHARGE)
PARAMETER*
Flow (m3/day)
pH (Std. Units)
TSS (mg/l)**
Copper (rng/t)
Zinc (mg/S)
Chlorides (mg/!)
MONITORING
PERIOD
March 1975 - Dec. 1979
March 1975 - Dec. 1979
March 1975 - Dec. 1979
March 1975 - Dec. 1979
March 1975 -Dec. 1979
March 1975 - Dec. 1979
FREQUENCY OF
OBSERVATION
Continuous
1/montht1"
1 /month1" f
1 /month
1 /month
1 /month
NUMBER OF
OBSERVATIONS
58
58
59
59
57
MEAN»**
79,107
7.8
5.5
<0.023
<0.026
866
STANDARD
DEVIATION
36,714
0.4
2.9
0.015
0.019
262
RANGEt
21,953-168,400
7.2-9.0
2-29
0.003 - 0.065
<0.002 - 0.09
354-1,494
OJ
tn
oo
* AH metals expressed as total metals unless otherwise specified.
** TSS values measured by turbidity and converted to mg/l.
t Range of concentrations based on maximum and minimum data between March 1975 and April 1977, and average concentrations between
May 1977 and December 1979.
tt Monthly averages based on almost continuous daily monitoring.
***Mean of monthly averages.
-------
TABLE VIII-46. HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL 2120 (TAILING-POND
OVERFLOW: TREATMENT OF UNDERGROUND MINE, MILL, AND LEACH-
CIRCUIT WASTEWATER STREAMS)
PARAMETER*
pH
pH
TSS
TSS
Cu
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
Sept. 75 - June 77
FREQUENCY OF
OBSERVATION
15-31/mo.
1/mo.
(mo. ave.)
15-31/mo.
1/mo.
(mo. ave.)
15-31/mo.
1/mo.
(mo. ave.)
1 - 5/mo.
1/mo.
(mo. ave.)
15-31/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
538
22
536
22
536
22
82
22
535
22
MEAN
(mg/l)
9.53
(daily ave.)
9.23
(ave. of mo. means)
9.6
(daily ave.)
10
(ave. of mo. means)
0.065
(daily ave.)
0.07
(ave. of mo. means)
<0.01
(daily ave.)
<0.01
(ave. of mo. means)
0.177
(daily ave.)
0.23
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.02
1.07
10.6
3
0.081
0.05
0
0
0.556
0.43
RANGE
(mg/l)
6.4 - 12.0
7.2-11.0
1 -132
5-19
0.01 -0.88
0.02 - 0.27
none
none
0.01 - 7 1
0.02-1.71
CO
01
'All metals expressed as total metals unless otherwise specified.
-------
CO
en
01
TABLE VIII-47. HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL 2120 (TREATMENT
SYSTEM - BARREL POND - EFFLUENT: TREATMENT OF MILL WATER
AND OPEN-PIT MINE WATER
PARAMETER*
pH
pH
TSS
TSS
Cu
Cu
Pb
Hg
Zn
Zn
MONITORING
PERIOD
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Apr. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
FREQUENCY OF
OBSERVATION
25-31 /mo.
1/mo.
25-31 /mo.
1/mo.
25-31 /mo.
1/mo.
3-5/mo.
3-5/mo.
25-31 /mo.
1/mo.
NUMBER OF
OBSERVATIONS
953
33
954
33
957
33
33
30
957
33
MEAN
(mg/l)
10.8
(daily ave.)
10.7
(ave. of mo. means)
12
(daily ave.)
12
(ave. of mo. means)
0.09
(daily ave.)
0.10
(ave. of mo. means)
0.015
(ave. of mo. means)
0.0003
(ave. of mo. means)
0.14
(daily ave.)
0.15
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.1
0.6
10
4
0.09
0.05
0.012
0.0001
0.16
0.09
RANGE
(mg/l)
2.9-13.1
9.7-12.4
0-120
7-22
< 0.01 - 0.98
0.04 - 0.27
< 0.01 - 0.06
< 0.00005 - 0.0006
< 0.01 - 2.00
0.04 - 0.36
*AII metals expressed as total metals unless otherwise specified.
-------
TABLE VIII-48. HISTORICAL DATA SUMMARY FOR LEAD MINE/MILL 3105 (TREATED
MINE EFFLUENT)
CJ
en
a-,
PARAMETER*
pH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
49
43
49
49
49
49
MEAN
(mg/l)
8.1
(ave. of mo. means)
3.0
(ave. of mo. means)
<0.02
(ave. of rno. means)
0.006
(ave. of mo. means)
0.043
(ave. of mo. means)
0.026
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
0.23
2.0
0.0
0.001
0.023
0.022
RANGE
(mg/l)
7.4 - 8.5
1 -9
na
< 0.005 -0.010
0.01 -0.12
0.005-0.11
*A!i metals expressed as total metals unless otherwise specified.
na = not applicable
-------
TABLE VIII-49. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE 3130
(UNTREATED MINEWATER)
on
i
PARAMETER*
pH
TSS
TSS
CN
CN
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Aug. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
FREQUENCY OF
OBSERVATION
unk
6/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.J
6/mo.
1/mo.
(mo. ave.)
6/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
unk
18
3
3
3
20
3
20
3
MEAN
(mg/l)
7.7
42.6
44.2
(ave. of mo. means)
0.11
0.11
(ave. of mo. means)
0.93
1.00
(ave. of mo. means)
1.25
1.35
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
n.a.
20.2
na
na
na
0.357
na
0.52
na
RANGE
(mg/l)
7.2 - 8.3
5.6 - 90.2
33.2-61.9
0.04 - 0.16
0.04-0.16
0.30-1.60
0.70-1.40
0.45 - 2.50
1.09-1.88
* All metals expressed as total metals unless otherwise specified.
unk = unknown
na = not applicable
-------
TABLE VIII-50. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE 3130 (TREATED EFFLUENT)
CO
en
oo
PARAMETER*
pH
pH
TSS
TSS
Cu
Cu
CN
CN
Pb
Pb
Zn
Zn
MONITORING
PERIOD
June 77 - Sept. 77
June 77 - Sept. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
June 77 - Oct. 77
June 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
FREQUENCY OF
OBSERVATION
unk
unk
8/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.)
8/mo
1/mo.
(mo. ave.)
8/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
unk
4
23
3
3
3
5
5
25
3
25
3
MEAN
(mg/l)
8.8
8.8
(ave. of mo. means)
2.12
2.1
(ave. of mo. means)
0.12
0.12
(ave. of mo. means)
0.064
0.064
(ave. of mo. means)
0.079
0.08
(ave. of mo. means)
0.32
0.33
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
na
na
1.34
na
na
na
na
na
0.021
na
0.085
na
RANGE
(mg/l)
8.5-8.9
8.6 - 8.9
<0.1 -6.1
1.8-2.4
0.05 - 0.25
0.05 - 0.25
0.022-0.150
0.022 -0.150
< 0.05 -0.13
0.067 - 0.097
0.18-0.57
0.24 - 0.39
*AII metals expressed as total metals unless otherwise specified.
unk = unknown
na = not applicable
-------
TABLE VIII-51. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3101 (TAILING-
POND DECANT TO POLISHING PONDS)
PARAMETER*
pH
pH
Cu!
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Oct. 75 Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
FREQUENCY OF
OBSERVATION
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
197
44
196
44
196
44
197
44
MEAN
(mg/l)
9.3
(daily ave.)
9.2
(ave. of mo. means)
0.054
(daily ave.)
0.055
(ave. of mo. means)
0.025
(daily ave.)
0.040
(ave. of mo. means)
0.193
(daily ave.)
0.294
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.5
1.3
0.072
0.060
0.049
0.049
0.264
0.401
RANGE
(mg/l)
7.0-11.3
6.7-11.2
0.002 - 0.490
0.008 - 0.323
0.004 - 0.525
0.004 - 0.303
0.010-2.20
0.038-2.15
*AII metals expressed as total metals unless otherwise specified.
na - not applicable
-------
TABLE VIII-52. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3101 (FINAL
DISCHARGE FROM POLISHING POND)
PARAMETER*
pH
pH
TSS
Cu
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
FREQUENCY OF
OBSERVATION
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
101
44
14
100
44
98
44
101
44
MEAN
(mg/l)
8.1
(daily ave.)
8.3
(ave. of mo. means)
8.4
(daily ave.)
0.024
(daily ave.)
0.020
(ave. of mo. means)
0.009
(daily ave.)
0.020
(ave. of mo. means)
0.211
(daily ave.)
0.176
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.3
1.3
na
0.027
0.017
0.008
0.018
0.123
0.096
RANGE
(mg/l)
6.6-11.2
6.9 - 10.9
0.3 - 26.0
0.002-0.124
0.006 - 0.076
0.004 - 0.052
0.005 - 0.082
0.026 - 0.548
0.028 - 0.390
CO
en
o
*AII metals expressed as total metals unless otherwise specified.
na = not applicable
-------
TABLE VIII-53. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/M'LL 3102 (FINAL DISCHARGE)
PARAMETER*
pH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 Sept. 74
Dec. 73 - Sept. 74
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
10
10
10
10
10
10
MEAN
(mg/l)
7.9
(ave. of mo. means)
2.6
(ave. of mo. means)
<0.02
(ave. of mo. means)
0.001
(ave. of mo. means)
0.002
(ave. of mo. means)
0.01
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
0.3
3.8
0
0.001
0.002
0.02
RANGE
(mg/l!
7.6 - 8.4
0-10
na
< 0.001 - 0.003
< 0.001 - 0.007
< 0.001 - 0.07
*AII metals expressed as total metals unless otherwise specified.
na = not applicable
-------
TABLE VI11-54. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3103 (TAILING-POND
EFFLUENT TO SECOND SETTLING POND)
CO
01
r-o
PARAMETER*
pH
TSS
Cu
Pb
Zn
MONITORING
PERIODt
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
32**
35
40
40
40
MEAN
(mg/l)
7.84
<1.63
0.023
0.154
0.309
STANDARD
DEVIATION
(mg/l)
0.26
0.8
0.013
0.0685
0.174
RANGE
(mg/l)
7.4 - 8.5
<1.0-4.0
0.005 - 0.058
0.038 - 0.33
0.030 0.79
*AII metals expressed as total metals unless otherwise specified.
tThe following months were excluded due to missing data: Oct. 1975, June 1975,
Sept. 1975, Oct. 1976, and Jan. 1977.
**Several illegible data points excluded.
-------
TABLE VIM-55. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3103 (EFFLUENT
FROM SECOND SETTLING POND)
PARAMETER*
pH
TSS
Cu
Pb
Zn
MONITORING
PERIOD*
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
33**
38
39
40
40
MEAN
(mg/l)
7.93
<1.55
0.013
0.058
0.110
STANDARD
DEVIATION
(mg/l)
0.22
1.48
0.007
0.028
0.086
RANGE
(mg/l)
7.5 - 8.4
1.0-8.0
0.002 - 0.034
0.01 -0.122
0.018 - 0.440
CO
CTl
oo
*AII metals expressed as total metals unless otherwise specified.
tThe following months were excluded due to missing data: Oct. 1974, June 1975, July 1975, Sept. 1975, Oct. 1975, and Jan. 1977
'Several illegible data points excluded.
-------
TABLE VIII-56. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3104 (TAILJNG-
LAGOON OVERFLOW)
PARAMETER*
pH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
Apr. 77 - Dec. 77
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
FREQUENCY OF
OBSERVATION
4-5/mo. {mo. ave.)
4-5/mo. (mo. ave.)
1/mo.
4-5/mo. (mo. ave.)
4-5/mo. (mo. ave.)
4-5/mo. (mo. ave.)
NUMBER OF
OBSERVATIONS
48
48
9
48
48
48
MEAN
(mg/l)
7.8
6.8
<0.1
0.06
0.07
0.13
STANDARD
DEVIATION
(mg/l)
0.8
2.9
0
0.04
0.02
0.07
RANGE
(mg/l)
6.8-10
2.9-16
na
0.01 -0.14
0.03-0.12
0.03 - 0.42
*AII metals expressed as total metals unless otherwise specified.
na = not applicable
-------
TABLE VIII-57. HISTORICAL DATA SUMMARY FOR LEAD/ZING MILL 3110
(TAILING-LAGOON OVERFLOW)
PARAMETER*
pH
TSS
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
40
40
40
40
40
MEAN
(mg/i)
7.2
5.5
0.02
0.05
0.20
STANDARD
DEVIATION
(mg/l)
0.4
5.1
0.02
0.02
0.18
RANGE
(mg/l)
6.3 - 8.3
0.4 - 20
0.001 -0.111
0.001 -0.106
0.01 - 0.35
CO
CTl
on
*AII metals expressed as total metals unless otherwise specified.
-------
TABLE VIII-58. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE 6103 (TREATED EFFLUENT)
PARAMETER*
pH
TSS
As
Cd
Cu
Pb
Mo
Zn
MONITORING
PERIOD
July 76 - June 77
July 76 - June 77
Apr. 77 - June 77
July 76- June 77
July 76 - June 77
Apr. 77 - June 77
July 76- June 77
July 76 -June 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
12
12
3
12
12
3
12
12
MEAN
-------
TABLE VIII-59. Hu JSICAL DATA SUMMARY FOR MOLYBDENUM MINE 6101 (TREATED EFFLUENT)
PARAMETER*
pH
TSS
COD
Cd
Cu
CN
Mo
Se
Zn
MONITORING
PERIOD
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
1972
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
1972
Jan. 75 - Dec. 76
FREQUENCY OF
OBSERVATION
0-5/mo.
0-5/mo.
0-5/mo
0-5/mo
0-5/mo.
0-5/mo.
0-5/mo
0-5/mo.
0-5/mo.
NUMBER OF
OBSERVATIONS
47
50
33
44
4
51
43
4
48
MEAN
(mg/l)
7.57
(daily ave.)
7.6
(daily ave.)
28.5
(daily ave.)
<0.02
(daily ave.)
<0.02
(daily ave.)
<0.02
(daily ave.)
1.84
(daily ave.)
< 0.005
(daily ave.)
0.077
(daily ave.)
STANDARD
DEVIATION
(mg/l)
0.38
7.5
12
na
0.01
0.021
0.38
na
0.18
RANGE
(mg/l)
6.5 - 9.0
1 -34
8-52
< 0.01 - < 0.02
< 0.02 - 0.03
< 0.02 - 0.083
1.1 -2.9
none
< 0.01 - 0.90
CO
*AII metals expressed as total metals unless otherwise specified.
na = not applicable
-------
TABLE VIII-60. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/MILL 6102 (CLEAR
POND BLEED STREAM - INFLUENT TO TREATMENT SYSTEM)
CO
O)
co
PARAMETER*
TSS
Cd
Cu
Fe
Mn
Mo
Pb
Zn
Cyanide
MONITORING
PERIOD
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
FREQUENCY OF
OBSERVATION
1-5/week
1-5/week
1-5/week
1 -5/week
1-5/week
1-5/week
1-5/week
1 -5/week
1-5/week
NUMBER OF
OBSERVATIONS
33
34
35
33
35
35
35
35
35
MEAN
(mg/l)
65.4
0.02
0.055
2.2
7.3
5.6
0.01
0.84
0.06
STANDARD
DEVIATION
(mg/l»
180
0.022
0.016
2.4
1.2
1.3
0.002
0.68
0.06
RANGE
(mg/l)
10-1070
<0.01 -0.15
<0.05-0.10
0.60 - 13.0
5.4-12.2
3.0 - 8.3
<0.01 - 0.02
0.20 - 2.7
<0.01 - 0.2
All metals expressed as total metals unless otherwise specified
-------
TABLE VII! >i. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/MILL 6102 (FINAL
DISCHARGE FROM RETENTION POND - TREATED EFFLUENT)
CO
crj
l-O
PARAMETER*
TSS
Cd
h'
Cu
Fe
Mn
Mo
Pb
Zn
Cyanide
MONITORING
PERIOD
July - Oct. 1978
July -Oct. 1978
July -Oct. 1978
July -Oct. 1978
July -Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
FREQUENCY OF
OBSERVATION
1-5/week
1-5/week
1 -5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1 -5/week
NUMBER OF
OBSERVATIONS
36
37
37
37
36
37
37
37
37
MEAN
(mg/l)
9
0.01
<0.05
0.4
0.26
2.5
<0.01
0.29
<0.01
STANDARD
DEVIATION
(mg/l)
11.4
0.006
0
0.32
0.17
1.8
0
0.16
0
RANGE
(mg/l)
<5-60
<0.01 -0.03
no range
0.15-2.0
0.1 -0.7
0.5 - 7.0
no range
<0.05-0.8
no range
*All metals expressed as total metals unless otherwise specified.
-------
TABLE VIII-62. HISTORICAL DATA SUMMARY; BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 008;
JANUARY 1979- DECEMBER 1980
PARAMETER**
Flow (m3/day»*
pH (Std. Units)
TSSfmg/Ott
Iron (mg/Ott
Aluminum Img/l)**
MONITORING PERIOD
Jan. 1979 - Dec. 1980
Jan. 1979 - Dec. 1980
Jan. 1979 -Dec. 1980
Jan. 1979 - Dec. 1980
Jan. 1979 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
1/week
1/week
I/week
NUMBER OF
OBSERVATIONS***
-
21
21
21
21
MEAN*
10,860
-
3.7
0.52
1.04
STANDARD
DEVIATION
7.470
-
2.5
0.46
0.50
RANGE
0-38.153
4.5 - 8.8
0.5- 13.3
0.08 - 2.61
0.22 - 4.02
OJ
~-w
o
* Mean flow and standard deviation include zero discharge months. No discharge during July. August and September 1980.
** All metals expressed as total metals unless otherwise specified.
***Only one sample collected during September 1979, December 1979, June 1980 and October 1980.
t Mean of monthly averages.
tt Mean, standard deviation and range for discharging months only. Number rounded to significant figures.
-------
TABLE VIII-63. HISTORICAL DATA SUMMARY; BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 009;
FEBRUARY 1979 - DECEMBER 1980
CO
I
PARAMETER*
Flow (m^/day)
pH (Std. Units)
TSS(mg/l)tt
Iron (ing/I)
Aluminum (mg/\f*
MONITORING PERIOD
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
1/week
1/week
I/ week
NUMBER OF
OBSERVATIONS**
-
21
21
21
21
MEANf
14,120
-
2.2
0.19
0.67
STANDARD
DEVIATION
8,930
-
0.92
0.08
0.37
RANGE
5,791 - 43,868
6.2 - 8.8
0.5- 11.2
0.02 - 0.6
0.12-2.25
* All metals expressed as total metals unless otherwise specified
** No data for April 1980 and July 1980
t Mean of monthly averages
tt Numbers rounded to significant figures
-------
TABLE Vlll-64. HISTORICAL DATA SUMMARY: BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 010;
FEBRUARY 1979 - DECEMBER 1980
Co
PARAMETER*
Flow (m3/day)tt
PH (Std. Units)
TSS (mg/0***
Iron «mg/l***
Aluminum (mg/l)*»*
MONITORING PERIOD
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979- Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
I/week
1/week
I/ week
NUMBER OF
OBSERVATIONS**
19
19
19
19
MEAN*
7,040
-
4.4
0.25
1.47
STANDARD
DEVIATION
4,845
-
1.6
0.09
0.69
RANGE
0- 17.714
4.4 - 8.8
0.8- 14.4
0.02 - 0.78
0.48 - 5.0
* All metals expressed as total metals unless otherwise specified
** Only one sample collected during February and March 1980
**'Mean, standard deviation and range for discharging months only. Numbers rounded to significant figures
t Mean of monthly averages
ft Mean flow and standard deviation include zero discharge months. No discharge during July, August, September and November 1980,
-------
TABLE VIII-65. HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT; DISCHARGE 001;
JUNE 1978 - DECEMBER 1980
GO
^vl
GO
PARAMETER*
Flow (m3/day)tt
TSS(mg/l?tt
pH (Std. Units)
Aluminum (mg/l)tt
iron (mg/l)1"t
MONITORING PERIOD
June 1978- Dec. 1980
June 1978 Dec. 1980
June 1978- Dec. 1980
Oct. 1978 - Dec. 1980
Oct 1978- Dae. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
1/week
1-3/month
1-3/ month
NUMBER OF
OBSERVATIONS
31
31
27
27
MEAN*
S,5«0
5.2
-
<0.36
<0.23
STANDARD
DEVIATION
2,900
2.8
-
0.32
0.36
RANGE
0-12,536
0 - 30.0
5.2 - 8.7
<0.1 - 1.50
< 0.01 - 5.2
*AII metals expressed as total metals unless otherwise specified.
t Mean of monthly averages.
t+Numbers rounded to significant figures.
-------
rABLE VIII-66. HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT; DISCHARGE 007;
JANUARY 1978- DECEMBER 1980
PARAMETER**
Flow (m3/0ay)*
TSS(mg/l)t
pH (Std. Units)
Aluminum (mg/l)t
Iron (mg/l)1'
MONITORING PERIOD
Jan. 1978 - Dec. 1980
Jan. 1978 - Dec. 1980
Jan. 1978 -Dec. 1980
Jan. 1978 - Dec. 1980
Jan. 1978 - Dee. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1-4/month
I/week
1 -4/ month
1-4/month
NUMBER OF
OBSERVATIONS
-
16
16
9
9
MEAN^t
370**
6.8
-
<0.15
<0.76
STANDARD
DEVIATION
534**
5.0
-
0.06
0.60
RANGE
0 - 4,360
0- 32
6.0 8.9
<0.1 -0.29
<0.01 -2.6
* Mean flow and standard deviation include zero discharge months.
"There were 20 months during the monitoring period in which no discharge occurred.
t Mean, standard deviation and range fcr discharging months only. Numbers rounded to significant figures.
ttMean of monthly averages.
-------
TABLE VIII-67. HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT; DISCHARGE 009;
JANUARY 1980 - SEPTEMBER 1980
PARAMETER**
Flow (m3/dayi*
TSS(mg/!)tt
pH !Std. Units!
Aluminum (mg/IS*^
Iron (mg/l)tt
MONITORING PERIOD
Jan. 1980 -Sept. 1980
Jan. 1980 -Sept. 1980
Jan. 1980 - Sept. 1980
Jan. 1980 -Sept. 1980
Jan. 1980 -Sept. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
3-4/month
1/week
1-2/month
1-2/ month
NUMBER OF
OBSERVATIONS
_
7
7
7
7
MEAN*
1,060
5.9
-
<0.27
0.27
STANDARD
DEVIATION
1,050
2.6
-
0.14
0.14
RANGE
0-8,176
0- 26
6.1 - 8.2
<0.1 -0.64
<0.05- 0.53
* Mean flow and standard deviation include zero discharge months. No discharge occurred during January or September 1980.
**AII metals expressed as total metals unless otherwise specified.
t Mean of monthly averages.
ttMean, standard deviation and range for discharging months only. Numbers rounded to significant figures.
-------
TABLE VIII-68. HISTORICAL DATA SUMMARY FOR TUNGSTEN MINE 8104 (TREATED EFFLUENT)
PARAMETER*
PH
TSS
Cu
Mo
Zn
MONITORING
PERIOD
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1-4/mo.
(mo. ave.)
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
35
24
24
24
24
MEAN
(mg/l)
7.98
'ave. of mo. means)
10.8
(ave. of mo. means)
<0.02
0.074
<0.01
STANDARD
DEVIATION
(mg/l)
0.27
4.4
na
0.17
na
RANGE
(mg/I)
7.3 - 8.6
4-20
none
< 0.02 - 0.74
none
CO
-»J
en
*Aii metals expressed as tot a! metals unless otherwise specified.
na - not applicable
-------
TABLE VIII-69. HISTORICAL DATA SUMMARY FOR URANIUM MINE 9408 (TREATED MINE WATER)
PARAMETER*
pH
TSS
Ra226
(dissolved)
U238
Ba
Mo
Pb
Zn
MONITORING
PERIOD
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 -Jan. 77
Apr. 75 -Jan. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
4/yr.
1/mo.
4/yr.
NUMBER OF
OBSERVATIONS
21
18
19
21
21
7
19
7
MEAN
(mg/l)
8.3
6
0.55
2.0
0.52
0.65
0.06
0.02
STANDARD
DEVIATION
(mg/l)
0.3
4
0.45
1.3
0.38
0.53
0.11
0.02
RANGE
(mg/l)
7.8 - 8.7
1 -10
0- 1.2
0.2-4.2
0.06-1.6
0.05 - 1.3
0.01 - 0.50
0.008 - 0.08
CO
*AII metals expressed as total metals unless otherwise specified.
-------
TABLE VIII-70.
HISTORICAL DATA SUMMARY FOR NICKEL ORE MINE/MILL 6106 - TREATED
EFFLUENT; DISCHARGE 001; JANUARY 1976 - DECEMBER 1980
CO
^J
oo
PARAMETER*
Flow (m3/day)*
pH (Std. Units) ***
TSStmg/l)*"*
Chromium (mg/l)*"*
Manganese (mg/l) ***
«»*
Phosphorous (P) (mg/l)
*»*
Settleable Solids (mg/l)
MONITORING PERIOD
Jan. 1976- Dec. 1980
Jan. 1976 -Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 -Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Daily
I/week
1/week
I/month
1 /month
1/month
I/week
NUMBER OF
OBSERVATIONS
59
31
32
32
32
31
30
MEANtt
3,520**
8.5
18.6
0.034
0.023
0.05
<0.05
STANDARD
DEVIATION
5,560**
0.30
15.4
0.017
0.016
0.017
0
RANGE
0 - 67,700
8.1 -9.59
0.4- 138
Not available
Not available
Not available
<0.05-0.05
* Mean flow and standard deviation include zero discharge months.
**There are 28 months during the monitoring period in which no discharge occurred.
t All metals expressed as total metals unless otherwise specified.
ttMean of monthly averages.
***Mean, standard deviation and range for discharging months only. Number rounded to significant figures.
-------
TABLE VIII-71. HISTORICAL DATA SUMMARY; VANADIUM MINE 6107; MINE AREA RUNOFF WASTE
PILE RUNOFF; DISCHARGE 005; JULY 1978 - DECEMBER 1980
PARAMETER*
Flow (m3/day)**
TSS(mg/l)<>»*
Iron (mg/l)'*»
pH (Std. Unite)
MONITORING PERIOD
July 1978 - Dec. 1980
July 1978 - Dec. 1980
July 1978 - Dec. 1980
July 1980 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
I/ week
I/week
5/week
NUMBER OF
OBSERVATIONS
31
27t
27t
28
MEAN
6,140
4
0.65
-
STANDARD
DEVIATION
5,760
3
0.23
-
RANGE
0-51,098
<1 - 12
0.15-2.9
5.4 - 9.3
OJ
* All metals expressed as total metals unless otherwise specified.
** Mean flow and standard deviation include zero discharge months. No discharge during September 1979, July 1980 and September 1980.
***Mean, standard deviation and range for discharging months only. Numbers rounded to significant figures.
t No values recorded for April 1980.
tt Mean of monthly averages.
-------
TABLE VIII-72. HISTORICAL DATA SUMMARY FOR TITANIUM MINE/MILL 9906;
TREATED MINE/MILL WATER; OCTOBER 1975 - DECEMBER 1979*
CO
oa
o
PARAMETER
Flow (m3/day)
pH (Std. Units;
TSS (mg/l)
Settleable Solids (mg/l)
MONITORING PERIOD
Oct. 1975 -Dec. 1979
Oct. 1975 -Dec. 1979
Oct. 1977 - Dec. 1979
Oct. 1979 -Dec. 1979
NUMBER OF
OBSERVATIONS
54
54
27
3
MEAN"
25,927
7.0
8.1
<0.05
STANDARD
DEVIATION
9,176
-
1.1
-
RANGE
1,060-68,130
4.0 - 10.0
2.4 - 22
-
* Summarized from Mine/Mill 9906 Discharge Monitoring Reports from October 1975 - December 1979
"Mean of monthly averages
-------
TABLE VIII-73. CHARACTERIZATION OF RAW WASTEWATER (TAILING-POND EFFLUENT
AS INFLUENT TO PILOT-SCALE TREATMENT TRAILER) AT COPPER
MILL 2122 DURING PER IOD OF 6-14 SEPTEMBER 1978
r -
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
TOC
phenol
NUMBER OF
OBSERVATIONS
26
27
23
l
1
3
2
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
8.0
2554
-------
TABLE VIII-74. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TREATMENT TRAILER) AT BASE AND PRECIOUS METALS MILL 2122
DURING PERIOD OF 8-19 JANUARY, 1979.
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
Fe
COD
TOC
CN
Total
Phenolics
PERIOD OF
OBSERVATIONS
(DATES)
Jan. 8-19, '79
NUMBER OF
OBSERVATIONS
16
10
--
3
--
--
3
3
3
--
3
3
3
--
3
3
5
13
9
14
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
8.8
213
--
0.03
--
--
0.07
0.44
0.11
--
0.06
0.025
0.04
--
0.03
23
39
17
0.03
0.58
RANGE
8.5-8.9
25-1200
--
0.006-0.08
--
--
0.04-0. 13
0.04-1.24
0.06-0.19
--
0.02-0.14
0.02-0.03
<0. 03-0. 08
--
0.01-0.08
0.42-68
32-52
6-24
0.003-0.060
0.23-0.81
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
--
--
--
--
--
<0.02
0.08
0.06
--
0.02
--
0.03
--
0.01
0.04
--
--
--
RANGE
--
--
--
--
--
<0.02
0.02-0.21
0.05-0. OS
--
0.02
--
0.03
--
<0. 01-0. 03
0.02-0.07
--
--
--
382
-------
TABLE VIII-75. SUMMARY OF PILOT-SCALE TREATABILITY STUDIES PERFORMED AT
MILL 2122 DURING PERIOD OF 6-14 SEPTEMBER 1978
UNIT TREATMENT
PROCESS EMPLOYED
Sedimentation (2.8-hr theor.
retention time)
Sedimentation (10.4-hr theor.
retention time)
Polymer Addition, Lime Addition
Flocculation, Sedimentation
(2.8-hr theor. retention time)
Polymer Addition, Lime Addition
Flocculation, Sedimentation
(2.8-hr theor. retention time).
Filtration
Polymer Addition, Lime Addition
Flocculation, Sedimentation
(2.8-hr theor. retention time).
Polymer Addition, Lime Addition
Flocculation, Sedimentation
(2.8-hr theor. retention time)
Filtration
Filtration
EFFLUENT CONCENTRATION* ATTAINED (mg/l)
pH*
7.9
7.7
9.3
9.1**
(9.0 to
9.2)
9.9
9.9
8.0**
(7.8 to
8.2)
TSS
50
18
21
<1»*
(CD
52
1
7.5**
K1 to
30)
Cr
0.035
0.035
0.04
0.03**
(0.03)
0.035
0.035
0.03**
(0.02 to
0.04)
Cu
0.05
0.045
0.04
0.033**
(0.03 to
0.04)
0.035
0.02
0.032**
(0.01 to
0.055)
Pb
0.09
0.08
0.09
0.07**
(0.06 to
0.09)
0.06
0.06
0.075**
(0.05 to
0.11)
Ni
0.07
0.04
0.05
0.047**
(0.04 to
0.05)
0.04
0.05
0.05**
K0.02to
0.11)
Zn
0.03
0.05
0.03
0.027**
(0.025 to
0.030)
0.02
0.02
0.06**
(0.03 to
0.18)
* All metals concentrations are based on "total" analyses
* Value in pH units
**Average concentrations attained
( ) Range of concentrations attained
383
-------
TABLE Vlli-76, PERFORMANCE OF A DUAL MEDIA FILTER WITH TIME-FILTRATION
OF TAILING POND DECANT AT MILL 2122
3 2
Hydraulic Loading on Filter = 762 m /m /day (13 gpm)
Initial TSS concentration = 33 mg/1
Time Elapsed
t + 15 min
o
t + 2 hr 15 min
o
t + 4 hr 15 min
o
t + 7 hr 15 min
o
t + 10 hr 15 min
o
Final TSS Concentration
mg/1
12
7
11
23
31
384
-------
TABLE VIII-77. RESULTS OF ALKALINE CHLORINATION BUCKET TESTS FOR
DESTRUCTION OF PHENOL AND CYANIDE IN MILL 2122 TAILING
POND DECANT*
*Unspiked cyanide concentration =<0.01-0.06 mg/1
*Initial phenol concentration = 0.232-0.808 mg/1
*1 mg/1 cyanide added to wastwater as NaCN
NaOCl Dose 5 mg/1
pH
Contact Time (min)
0
15
30
60
120
NaOCl Dose 10 mg/1
PH
Contact Time (min)
0
15
30
60
120
NaOCl Dose 20 mg/1
, ___
Contact Time (min)
0
15
30
60
120
NaOCl Dose 50 mg/1
Contact Time (min)
0
15
30
60
120
9 I 10
CN~
0.189
0.20
0.020
0.095
0.040
Total
Phenolics
-
-
-
-
-
CN
0.159
0.080
0.55
0.051
0.050
Total
Phenolics
0.536
0.648
0.592
-
-
11
CN"
-
0.462
0.484
0.505
0.386
Total
Phenolics
-
0.776
0.704
-
-
9
CN"
0.190
0.095
0.080
0.079
0.088
Total
Phenolics
-
-
-
-
-
10
CN"
2.95
3,29
4.180
4.170
2.850
Total
Phenolics
0.664
0.488
0.536
-
-
11
CN"
-
0.294
0.284
0.396
0.305
Total
Phenolics
-
0.768
0.488
-
-
9
CN"
v$
Total
Phenolics
Xs
10
CN"
0.750
0.012
0.003
0.001
0.001
Total
Phenolics
0.808
0.680
0.396
-
-
11
CN"
1.09
0.039
0.012
0.011
0.011
Total
Phenolics
0.608
0.452
0.696
-
-
9
CN"
0.88
0.05
0.02
0.01
0.004
Total
Phenolics
0.336
0.228
0.088
-
0.216
10
CN"
1.2
0.001
0.008
0.003
<0.001
Total
Phenolics
0.720
0.058
0.084
0.080
-
11
CN" ! Total
UN Phenolics
>^
^0^
All cyanide and phenol concentrations are mg/1.
385
-------
TABLE VIII-78. RESULTS OF PILOT-SCALE OZONATION FOR DESTRUCTION OF
PHENOL AND CYANIDE IN MILL 2122 TAILING POND DECANT*
CO
oo
CT1
0_ Dosage
mg/min
2.4
0.8
6
2
1
24
8
4
2
Ratio O./CN
2:1
2:1
5:1
5:1
5:1
10:1
10:1
10:1
10:1
Initial pH
10
9
10
11
10
10
9
10
8.5
Retention Time
10
30
10
30
60
5
15
30
60
Final Concentration (mg/1)
CN
0.610
0.460
0.035
0.040
0.016
-
0.020
0.065
0.035
Total Phenolics
0.416
0.256
0.312
0.428
0.272
0.052
0.026
0.200
0.474
Initial Phenol Cone. = 0.232-0.808 mg/1; Cyanide added as NaCN to provide an initial cone.
of 1 mg CN/1 of wastewater.
-------
TABLE VIII-79. EFFLUENT QUALITY ATTAINED AT SEVERAL PLACER MINING OPERATIONS
EMPLOYING SETTLING-POND TECHNOLOGY
MINE
4142
4141
4114
4140
4139
4136
4135
4126
4127
4132
4133
4134
DATE
SAMPLED
July 14, 1977
July 20, 1977
July 14, 1977
July 17, 1977
July 12, 1977
Aug. 26, 1978
Aug. 24, 1978
Aug. 15, 1978
Aug. 15, 1978
Aug. 18, 1978
Aug. 19, 1978
Aug. 21, 1978
INFLUENT
pH*
-
-
-
7.3
7.4
-
-
6.5
6.7
6.6
7.9
-
TSS
(mg/l)
-
-
24,000
1,130
9,000
64,100
2,890
14,800
39,900
1,540
2,260
-
Settles ble
Solids
(ml/l/hr)
1.5
17
1.8
1.7
13
45
7.5
260
550
1.6 - 2.0
0.7 - 1.6
1.6
As
(mg/l)
-
-
-
-
1.2
3.9
0.04
1.3
5.0
0.05
1.5
-
Hg
(mg/l)
-
-
-
-
0.004
0.001
0.020
0.0002
0.0014
<0.0002
0.0002
-
EFFLUENT
pH*
-
-
8.5
7.4
-
-
6.8
6.4
6.5
7.7
-
TSS
(mg/l)
2080
120
<0.1
220
230
150
474
76
5,700
1040
170
1420
Settleable
Solids
(ml/l/hr)
0.3
<0.1
<0.1
<0.1
0.15
<0.1
0.7 - 0.9
<0.1
2.5
0.4 - 0.8
<0.1
0.4
As
(mg/l)
0.27
0.031
-
0.057
0.012
<0.002
0.022
0.25
1.2
0.05
0.06
0.28
Hg
(mg/l)
< 0.0002
<0.0002
-
< 0.0002
< 0.0002
< 0.0002
< 0.0002
0.0002
0.0005
<0.0002
0.0002
< 0.0002
GO
00
'Value in pH units
-------
Figure Vlll-1. POTABLE WATER TREATMENT FOR ASBESTOS REMOVAL AT
LAKEWOOD PLANT, DULUTH, MINNESOTA
CHEMICAL ADDITION
(ALUM, CAUSTIC SODA,
POLYMERS)
LAKE WATER
t
MIX
TANKS
FLOCCULANTS
FLOCCULATION
SEDIMENTATION
Co
00
O3
I
MULTI-MEDIA
FILTRATION
BACKWASH
WATER
SUPER-
NATANT
»>
POTABLE WATER TO
STORAGE AND
DISTRIBUTION
SLUDGE TO
SANITARY
LANDFILL
-------
Figure VIII-2. EXPERIMENTAL MINE-DRAINAGE TREATMENT SYSTEM FOR
UNIVERSITY OF DENVER STUDY
MINE
DRAINAGE
LIME-
CONTACT TANK
(NEUTRALIZATION
STAGE)
CONTACT TANK
(SULFIDE-TREATMENT
STAGE)
Q.
SULFIDE
SOLUTION
PUMP
fSULFIDE
MIXING
TANK
DISTRIBUTION TROUGH
\ III
TREATED
EFFLUENT
389
-------
Figure Vlll-3. CALSPAN MOBILE ENVIRONMENTAL TREATMENT PLANT CONFIGURATION
EMPLOYED AT BASE AND PRECIOUS METAL MINE AND MILL OPERATIONS
min
/mini
"*" 3,785-liter
(1.000 -gallon)
PRIMARY
SETTLING/EQUALIZATION
TANK f^
tl
BYPASS
I
t
su
TA
I
MP
NK
OVERFLOW
TO
WASTE
OVERFLOW
TO
WASTE
102liter
(27-gallon)
FLOCCULATION
GRAVITY TANK
OVERFLOW
-FLOCCULATOR BYPASS f^] jl
SECONDARY
TREATMENT
BYPASS
OVERFLOW
TO
^ WASTE
SLUDGE
TO WASTE"*
390
-------
Figure VIII-4. MOBILE PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT
URANIUM MILL 9402
0
INFLUENT
(END-OF-PIPE)
t
V BARIUM
^CHLORIDE
STOCK
PUMP RUMP LIQUID A
^. .Pi POLYMER
-------
Figure VIII-5. PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT URAWUM
MILL 9401
INFLUENT
OVERFLOW
TO WASTE
O
END-OF-PIPE)
J^%
J
1
T
SURGE
TANK
r-C*3-
h^
t
I
1
i
IX
CO
LUM
NS
1
!
IX
ELUATE
HOLDING
TANK
V
~l
1
, !
i
P
I
IX
ELUANT
FLOCCULATION
TANK
GRAVITY
OVERFLOW
SLUDGE
TO WASTE
OVERFLOW
TO WASTE
AERATOR METERING
(EJECTOR) PUMP
SECONDARY
TREATMENT
BYPASS
OVERFLOW
TO WASTE
BACKWASH
TO WASTE
392
-------
Figure VI11-6. MODE OF OPERATION OF SEDIMENTATION TANK DURING PILOT-SCALE
TREAT ABILITY STUDY AT MINE/MILL 9402
OVERFLOW WEIR
DISCHARGE
BAFFLE
INFLUENT
393
-------
Figure VIII-7. FRONTIER TECHNICAL ASSOCIATES MOBILE TREATMENT PLANT CONFIGURATION
EMPLOYED AT MINE/MILL/SMELTER/REFINERIES #2121 & #2122
LIME
SLURRY
PRESSURE
Q INDICATOR
BACKWASH
PORTS
INFLUENT
90 GAL.
BATCH
REACTION
TANK
HXJ-
t
DRAIN
100 GAL*
FILTRATE
HOLDING
TANK
* 90 gallons = 0.34 cubic meters
100 gallons =& 0.38 cubic meters
DRAIN
-------
Figure VIII-8. DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT AT COPPER
MINE/MILL 2120*
GO
to
01
OVERFLOW
POLYMER
OVERFLOW
OPEN-PIT
MINE WATER
MISC. WASTES I
LIME
/ SURGE \
I POND 1
MIX
BOX
! >
I TAILING
I POND
UNDERGROUND
MINE WATER
LEACHATE
DISCHARGE
CONTINUOUS WATER FLOW
INTERMITTENT WATER FLOW
PROPOSED WATER FLOW
OVERFLOW
DISCHARGE
Water flow configuration representative of period during which historical monitoring data was generated.
-------
Figure VIN-9. DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT AT BASE
AND PRECIOUS METALS MINE/MILL 2120 (COPPER)**
OPEN-PIT
MINE WATER
CO
DISCHARGE
INTERMITTENT WATER FLOW
PROPOSED WATER FLOW
DISCHARGE
**Water flow configuration represents modifications made to system as of September, 1979
-------
Figure VIII-10. SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT FACILITIES
AT LEAD/ZINC MINE/MILL 3103
MINEWATER PUMPAGE
94.6 I/sec (1,500gal/min)
*_
MINE SUMP
TO SMELTER
31.5 I/sec
(500 gal/min)
82.0 I/sec (1,300 gal/min)
L
MILL-WATER
STORAGE RESERVOIR
RECYCLE
63.1/0 I/sec
(1,000/0 gal/min)
113.5/0 I/sec
(1,800/0 gal/min)
CONCENTRATOR
MILL TAILINGS
94.6/0 i/sec
(1,500/0 gal/min)
RAINWATER
est. 17.7 I/sec
(est 280 gal/min)
12.6/94.6 I/sec
(200/1,500 gal/min)
CONCENTRATE
THICKENERS
EVAPORATION
AND SEEPAGE
est 13.2 I/sec
(est 210 gal/min)
RAINWATER
est 43.8 I/sec
(est 695 gal/min)
OVERFLOW
18.9/0 I/sec
(300/0 gal/min)
67.5/99.0 I/sec
(1,070/1,570 gal/min)
SMALL
STILLING POOL
DISCHARGE
11.3 I/sec
(1,765 gal/min)
CONTINUOUS OR
SEMI-CONTINUOUS FLOW
INTERMITTENT FLOW
XXX/XXX FLOW DURING MILL
OPERATION/FLOW DURING
MILL SHUTDOWN
397
-------
140
120
1 100-
S> 80
^ 60
X
< 40
20.
0.
= 80-
I
OT ^
P
ui 40
20--
-
OK
o
20--
16--
<«
Q E 12
Uj O
-------
Figure VHI-12. PLOT OF TSS CONCENTRATIONS VS. COPPER CONCENTRATIONS IN
TAILING-POND DECANT AT MINE/MILL/SMELTER/REFINERY 2122
D RND - TDTRL CDPPER CDNCENTRRTIDN5
+ RND = DISSOLVED CDPPER CDNCENTRHTIDNS
H.2 --
z
a
IT
a:
a
a:
LJ
a.
a.
H.I -.
H.B5T-
B B
»-
7S
IBS
IZS
I SB
175
zra
T55 CQNCENTRRTIDN CMB/L)
-------
-------
SECTION IX
COST, ENERGY, AND NON-WATER QUALITY ISSUES
DEVELOPMENT OF COST DATA BASE
General
Generalized capital and annual costs for wastewater treatment
processes at ore mining and dressing facilities have been estab-
lisned. Costing has been prepared on a unit process basis for
each ore category. Assumptions regarding the costs, cost
factors, and methods used to derive the capital and annual costs
are documented in this section. All costs are expressed in 1979
dollars (Engineering News Record construction index=3140; 13
December 1979, Reference 1).
The estimates were based on assumptions pertaining to system
loading and hydraulics, treatment process design criteria, and
material, equipment, manpower, and energy costs. These assump-
tions are documented in detail in this section.
Fourth quarter 1979 vendor quotations were obtained for all major
equipment and packaged systems. Construction costs were based on
standard cost manual figures (see References 2 and 3) adjusted to
December 1979.
The wastewater treatment processes studied are as follows:
Secondary Settling Ponds
Flocculation
Ozonation
Al ka1ine-Ch1orination
Activated Carbon Adsorption
Hydrogen Peroxide Oxidation
Chlorine Dioxide Oxidation
Potassium Permanganate Oxidation
Ion Exchange
Granular Media Filtration
pH Adjustment
Recy J.te
Ev- iration Ponds (total evaporation)
T??hle iX-1 indicates the processes studied for each ore
j"u>category.
CAPITAL COST
C_ap_i_tai_ Cost of_ Facilities
Settl ing Ponds, * Construction costs for settling ponds were based
upon assumptions (specifically documented later in this section)
401
-------
regarding the retention time and geometry of the ponds. Costs
for excavation and back filling were assumed to be
Process Tankage. Mixing tanks, flocculation tanks, wet wells,
ozone contactors and slurry tanks are sized for retention times
appropriate to the particular process. These retention times are
documented under the treatment process discussions later in this
section. Construction cost estimates for tankage were then based
on a factor of $300/yd3 of concrete (installed).
Reagent Storage Facilities. Cost estimates for tanks and bins
used for reagent storage were based on vendor quotations. Sizing
of the storage containers was based on dosage rates and backup
supply assumptions which are documented in the treatment process
discussions later in this section.
Buildings. Space requirements for housing treatment process
equipment were based on vendor quotations. Building construction
costs were developed from the methodology of References 2 and 3.
Piping. Unless otherwise stated, only local piping cost was
included in the capital cost estimates, and installed costs were
established from References 2 and 3. Long runs of interprocess
piping have not been included due to their site-specific nature.
Lagoon and Tank Liners. Where required, lagoon or tank lining
material was costed at two dollars per square foot (installed).
Structural Steel. Handrails and gratings, where required, were
costed at one dollar per pound (installed) of fabricated steel
equipment.
Capital Cost of_ Equipment
All equipment costs were obtained from vendor quotations.
Instrumentation and electrical packages (installed) were assumed
to be a percentage of the equipment costs. The percentages
documented in the individual treatment process discussions later
in this section, varied with the process in question.
Capital Cost of_ Installation
Unless otherwise stated, installation costs for equipment were
included in the vendor quotations. Construction costs for
facilities, including concrete, steel, ponds, tanks, piping and
electrical, were estimated on an installed basis.
Capital Cost of_ Land
Land costs were estimated at $4,000/acre unless otherwise stated.
402
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Capital Cost of Contingency
Unless otherwise stated, a contingency cost of 20 percent was
added to the total capital costs generated. This was intended to
cover taxes, insurance, over-runs and other contingencies.
ANNUAL COST
Annual Cost of_ Amortization
Initial capital costs were amortized on the basis of a 10 percent
annual interest rate with assumed life expectancy of 30 years for
general civil and structural equipment and 10 years for
mechanical and electrical equipment. Capital recovery factors
were calculated using the formula:
n
CRF = (r) (1+r)
n
(1+r) - 1
where CRF = capital recovery factor
r = annual interest rate
and n = useful life in years.
Annual cost of amortization was computed as:
C = B (CRF)
A
where C = annual amortization cost
A
and B = initial capital cost.
Annual Cost of_ Operation and Maintenance
Maintenance. Annual maintenance costs were assumed to be three
percent of the initial total capital cost unless otherwise noted.
Operation. Operating personnel wages were assumed to be
$13.50/hr. including fringe benefits, insurance, etc. Estimated
weekly operator manhours were established depending on the pro-
cess and the hydraulic flow rate. These manpower estimates are
documented in the individual treatment process discussions later
in this section.
Reagents. The following prices were used to estimate annual
costs of chemicals:
Polymer $ 2.00/lb.
Sodium Hydroxide $160.00/ton
Sodium Hypochlorite $ 0.40/lb.
403
-------
Hydrated Lime
Activated Carbon
Hydrogen Peroxide (70% cone.)
Sulfuric Acid (66 Be)
Ferrous Sulfate (400 Ib. drum
Chlorine Dioxide (5% cone, in
Potassium Permanganate (dry)
dry powder)
55 gal. drum)
$
?
$
$
$
$
$
65.
0.
0.
0.
0,
9.
0.
00/ton
50/lb.
35/lb.
04/lb
52/lb.
10/gal.
59/lb.
Reagent dosages are documented in the treatment process
discussions in this section.
Annual Cost of Energy
The cost of electric power was assumed to be three cents per
kilowatt-hour. Facilities were assumed to operate 24 hours per
day, 365 days per year.
Monitoring Costs
Additional wastewater monitoring costs were estimated as $7,000
per year for ozonation and alkaline chlorination systems, and
$10,000 per year for the remaining technologies except recycling.
These figures were intended to account for those added monitoring
costs associated only with the technologies described in this
section.
TREATMENT PROCESS COSTS
Secondary Settling
Capital Costs. The cost of
widely, depending on local
Figure IX-1 depicts the
estimates.
constructing settling ponds can vary
topographic and soil conditions.
typical layout assumed for these
The costs and required sizes of settling ponds were developed as
a function of hydraulic load. The basins were sized for a 24-
hour retention time with an anticipated 10 percent safety factor
(for sediment storage). It was assumed that lagoons and settling
ponds are rectangular in shape, with the bottom length twice the
bottom width. The dikes (berms) were constructed with a 2.5:1
slope. In all cases, the water depth was assumed to be 16 feet
and a one-foot freeboard was provided. Water was presumed to
flow by gravity.
For estimating purposes, it was assumed that 60 percent of the
total basin volume required excavation and backfilling (estimated
corrugated steel, and a total length of 200 feet was allowed
(estimated cost: $17.30/ft.). However, it was recognized that
longer runs of process interconnecting piping may be necessary in
individual cases.
404
-------
Complete capital cost estimates included costs for land,
excavation and backfilling, piping, installation of piping,
concrete pad for piping support, and pond liners (where
necessary to prevent seepage). The capital cost curve in Figure
IX-2 expresses the total capital cost as a function of hydraulic
flow rate for secondary settling ponds. A contingency cost
factor of 20 percent was included in these estimates. Figure IX-
3 expresses the estimated settling ponds line cost.
Annual Costs. Annual maintenance costs for secondary settling
ponds were assumed at $2,000, with additional monitoring costs of
$10,000/year. Amortization was based on a 30-year life
expectancy at 10 percent annual interest (CRF=0.10608). The
annual costs displayed in Figure IX-2 as a function of hydraulic
flow rate are the sum of the amortization, monitoring, and
maintenance costs. Annual costs for pond liners are shown in
Figure IX-3.
Flocculant Addition
Capital Costs. Capital costs were estimated for flocculation
systems consisting of the equipment shown in Figure IX-4. A
complete, installed mechanical package, which included the
flocculant preparation and feed equipment, was based on vendor
quotations. This package, designed for use with dry polymer,
included storage tank, feeder, wetting equipment, aging tank with
mixer, transfer pump, electrical and instrumentation package, and
installation. Piping, tanks, and metering pumps were corrosion
resistant. The remaining capital costs included site
preparation, enclosure, and civil work (i.e., grading, concrete,
super structure) as well as heating equipment (electrical heater,
installed). In addition, the total capital cost included a 20
percent contingency cost factor.
The systems were sized based on hydraulic flow rate/ conse-
quently, total capital cost is expressed as a function of waste-
water flow rate (Figure IX-5). A flocculant dosage of one part
per million was used. A one- to five-minute mixing time, and a
30-day reagent storage capacity were assumed.
Local electrical and piping connections were included in the cost
estimates. However, long runs of process interconnecting piping
and electrical power lines, if necessary, will need to be
estimated on a site-specific basis.
Annual Costs. Amortization of capital cost for flocculation
systems assumed a 10 percent annual interest rate with life
expectancies of 30 years for construction (CRF = 0.10608) and 10
years for mechanical and electrical equipment (CRF = 0.16275),
Operator hours were estimated at 13.3 hours per week (1/3 time),
and operator wages were calculated at $13.50 per hour including
benefits. Additional cssts were estimated as follows: annual
maintenance as three percent of capital cost; chemicals at a
405
-------
price of $2.00 per pound for dry polymer; energy at a rate of
$0.03 per kilowatt-hour; and additional monitoring at $10,000 per
year. An annual cost curve has been generated (Figure IX-5)
expressing the total of the above expenses as a function of
wastewater flow rate.
Ozonation
Capital Costs. The ozonation systems estimated in this section
were defined by the flow diagram shown in Figure IX-6. The
system equipment supply included air compressor with inlet
filter/ silencer, after cooler, refrigerant cooling system,
dessicant drying system, ozone generator, cooling tower, and
concrete ozone contact chamber, located indoors near the contact
chamber.
Equipment costs for the ozonation systems were based on vendor
quotation. Building construction costs were based on vendor
definition of special requirements with cost factors developed
from References 2 and 3. Installation costs were based on the
same references. A concrete cost factor of $300/yd3 (installed)
served as a basis for the ozone contact chamber costs.
The ozonation system design estimates were based on an ozone
dosage of five mg/1 and a contact time of 15 minutes.
Total capital cost figures included equipment, installation,
building construction, contactor tankage, and a 20 percent con-
tingency factor. The capital cost graph in Figure IX-7 expresses
the total capital cost as a function of flow in million gallons
per day.
Operating Costs. Amortization of capital costs was based on a 10
percent annual interest rate, a 30-year life expectancy for
construction (CRF = 0.10608), and a 10-year life expectancy for
equipment (CRF = 0.16275). Maintenance costs were assumed to be
three percent of the initial capital investment annually.
Operator manhours were estimated at 20 hours per week for systems
treating less than 10 million gallons per day, 30 hours per week
for 10 to 100 million gallons per day systems, and 40 hours per
week for systems treating greater than 100 million gallons per
day of wastewater.
Operator wages were costed at $13.50/hour including benefits.
Energy costs were based on a rate of $0.03 per kilowatt hour (3
cents). Electric power required for ozone generation was assumed
to be 10 to 12 kwh per pound of ozone generated. The annual cost
curve in Figure IX-7 depicts the sum of the above annual costs as
a function of flow in million gallons per day. Monitoring costs
of $7,000 per year should be added to the cost obtained from the
curve.
Alkaline Chlorination
406
-------
Capital Costs. The alkaline-chlorination system cost estimates
were generated based on the use of sodium hydroxide and sodium
hypochlorite as alkalinity and chlorine sources, respectively.
System definition is represented by flow schematic in Figure IX-
8.
Total capital cost estimates included storage facilities, mixing
tank with liner, mixers, electrical and instrumentation package,
reagent feed pumps, local piping and contingency costs (at 20
percent). Figure IX-9 includes a graph of total capital cost as
a function of hydraulic flow rate.
Cost estimates for chemical storage tanks, chemical feed pumps,
and mixing equipment were obtained from vendor quotations. The
two chamber mixing tank was estimated at $300/yd3 installed; and
electrical and instrumentation package costs were estimated at 20
percent of the total equipment cost.
In considering the capital costs, several system design
assumptions were made, including:
1. Sodium hydroxide dosage of 30 mg/1 and sodium hydro-
chlorite dosage of 10 mg/1
2. Mixing tanks sized for a two-minute retention time
3. Reagent storage capacity sized for a 30-day supply of
each chemical
4. Sodium hydroxide and sodium hypochlorite estimated to
5. Use of turbine-type mixers (carbon steel construction),
reciprocating (plunger) chemical feed pumps, carbon steel
sodium hydroxide handling and storage equipment, and
fiberglass sodium hypochlorite handling and storage
equipment
Annual Costs. Capital recovery was amortized over a 10-year
period for equipment and a 30-year period for construction. A 10
percent annual interest rate was used for both equipment and
construction. (Equipment CRF = 0.16275, Construction CRF =
0.10608). Annual maintenance costs were assumed to be three
percent of the initial capital investment. Operator manhours
were estimated at 10 hours/week and were costed at a rate of
$13.50/hour including benefits. Energy costs were developed at a
rate of $0.03/kilowatt hour; chemical costs were based on the
dosages previously mentioned (30 mg/1 NaOH; 10 mg/1 NaOCl);
chemical prices (delivered) were estimated at $160.00/ton (2,000
pounds) for caustic soda (NaOH) and $0.40/pound for sodium hypo-
chlorite (NaOCl); and additional monitoring costs of $7,000/year
were assumed. Figure IX-9 includes the annual cost curve.
Ion Exchange
407
-------
Capital Costs. The flow schematic for the ion exchange system is
exhibited in Figure IX-10. This is a combination cation-anion-
mixed bed process with a pretreatment (filtration) step. The
system costed consists of skid-mounted package units including
raw waste filters in steel tanks, cation exchangers, a
degasifier, anion exchanger, and mixed bed exchangers. Acid is
provided for regeneration of cation exchangers and caustic soda
for anion exchangers. These waste solutions are mixed and
require disposal. (All units are housed in a structure.)
Total capital costs include equipment, installation, building
construction, and 20 percent contingency. The capital cost curve
in Figure IX-11 relates this total capital cost to hydraulic flow
rate.
Supply and installation cost estimates for all equipment were
obtained from vendor quotations. Units were sized for hydraulic
loading according to vendor recommendations. Building construc-
tion costs (including concrete foundations) were estimated based
on vendor space requirement quotes and the costing methodology of
References 2 and 3.
Annual Costs. Amortization of initial capital investment was
based on a 10 percent annual interest rate at a 10-year life
expectancy for equipment (CRF = 0.16275) and a 30-year life
expectancy for construction (CRF = 0.10608). Annual maintenance
costs were estimated at three percent of initial capital cost.
Reagents were costed at $0.1 I/pound for caustic soda and
$0.03/pound for sulfuric acid. Electric power was costed at
$0.03/KWH. Operator hours were estimated at 20 hours per week
for plants treating less than 2.5 MGD, at 30 hours per week for
plants treating 2.5 to 10.0 MGD, and at 40 hours per week for
plants treating 10.0 to 35.0 OMGD. Operator wages and benefits
were estimated to total $13.50/hour. Additional monitoring costs
of $10,000/year were assumed. Figure IX-11 displays total annual
costs as a function of daily flow rate.
Granular Media Filtration
Capital Costs. Figure IX-12 depicts the basic granular media
filtration system proposed for cost estimates. Industrial,
gravity flow deep bed, granular media filters were selected. The
filters would be contained in prefabricated, portable steel
filter units. Treated effluent would discharge through a con-
crete backwash wastewater basin where the filtered solids would
settle. The supernatant would then be pumped back to the filters
for treatment.
All piping is carbon steel, valves are the butterfly type, and
the pumping equipment consists of vertical turbine pumps of
carbon steel construction. Pump impellers are bronze with
stainless steel shafts. Filter media consists of plastic filter
bottom, gravel, sand, and anthracite.
408
-------
Vendor quotations obtained for filters, pumps, and air blowers
(for backwash) included site preparation, installation, local
piping, and instrumentation and electrical package. Systems were
sized for a hydraulic loading of 10 gpm/ft2.
Total capital costs included a 20 percent contingency factor.
Figure IX-13 displays capital cost as a function of daily waste-
water flow.
Annual Costs. Initial capital investment was amortized at a 10
percent annual interest rate over a period of 10 years for
equipment (CRF = 0.16275) and 30 years for construction (CRF =
0.10608).
Costs estimated under annual costs include: (1) maintenance
estimated at three percent of annual capital cost; (2) operator
manhours established at 20 hours per week for systems treating
one to five million gallons per day (MGD) and 30 hours per week
for systems treating 10 to 100 MGD of wastewater; (3) electricity
computed at a rate of $0.03 KWH; and (4) additional monitoring at
$10,000 per year. Figure IX-13 includes the annual cost curve
for these systems.
p_H Adjustment
Capital Costs. System costs for pH adjustment by hydrated lime
addition were developed. A schematic representation of the
system is displayed in Figure IX-14. Major system components
include lime storage and feed equipment, slurry tanks, feed pump,
mixing tankage, and mixing equipment. The dry lime is stored in
a steel silo which is equipped with a screw type feeder. The
feed ratios of lime and water are preset and are started based on
level in the steel lime slurry tanks. A vertical type turbine
pump will pump the slurry into the wastewater mixing tanks. The
tanks are reinforced concrete structures containing 3 turbine
mixers of carbon steel construction. Mixers are for tank top
mounting.
Costs of lime storage and feed equipment as well as mixer costs
were obtained from vendor quotations. Mixing tankage and lime
slurry tankage costs were based upon installed costs of lined
concrete tanks. Electrical and instrumentation package installed
costs were assumed to be 20 percent of the equipment costs.
Cost estimates were completed based upon a 50 mg/1 dosage of 93
percent hydrated lime. A 30-day supply of lime was assumed for
the design of storage facilities. Lime slurry tankage was sized
for a 24-hour detention time, while mixing tanks were sized for a
two minute detention time for flows of up to 10 mgd, and a one
minute detention time for flows greater than 10 mgd.
Total capital cost estimates included storage bins, feeder
equipment, concrete and lining material for slurry and mixing
409
-------
tanks, mixers, slurry pumps, electrical and instrumentation,
installation, and contingency (at 20 percent). Figure IX-15
represents total capital cost as a function of wastewater flow.
Annual Costs. Annual costs estimated for the pH adjustment
process included the following: (1) amortization calculated at a
10 percent annual interest rate for a 10-year life expectancy for
equipment (CRT = 0.16275) and a 30-year life expectancy for
construction (CRF = 0.10608); (2) annual maintenance costs
estimated at three percent of the initial capital investment; (3)
operator manhours established at 10 hours per week and costed at
$13.507 hour including benefits, insurance, etc; lime costs based
on a price of $65/ton (2,000 Ibs.); (4) cost of energy estimated
at $0.03/KWH, (3 cents); and (5) additional monitoring costs of
$10,000/year. Figure IX-16 displays annual cost as a function of
daily wastewater flow.
Recycle
Capital Costs. Cost estimates were prepared for installation of
systems to provide for 25, 50, 75, and 100 percent recycle of
wastewater. Figure IX-17 represents the equipment and tankage
requirements on which the estimates were based. Recycle is
accomplished by collecting the effluent wastewater in a concrete
tank. Pumps are provided to return all or a portion of the flow
back to the mine or mill operations for reuse. Any quantity
greater than the recycle rate would overflow into the receiving
stream.
Recycle pumps are vertical turbine type complete with weather
proof motor for outdoor installations. Collection sewer and pump
discharge piping were not included in the costing.
Pumping equipment costs were based on vendor quotations. Wet
well costs were based on $300/yd3 installed concrete cost. Local
piping, valves, and fittings were costed based on vendor
definition and costing methodology taken from Reference 2.
Structural steel requirements for railings, gratings, etc. were
costed at a rate of one dollar per pound (installed). Electrical
and instrumentation package costs (installed) were estimated at
30 percent of the total equipment cost.
Pumping equipment selection was based on hydraulic flow
requirements assuming 75 feet total dynamic head requirement.
Wet well sizing was based on a 10-minute retention time.
Total capital cost estimates included concrete tankage, pumps and
motors, piping, valves, fittings, structural steel, electrical
and instrumentation, installation, and contingency (at 20
percent). Capital cost expressed as a function of hydraulic flow
rate is graphed in Figure IX-18. Cost curves are shown for 25,
50, 75, and 100 percent recycle.
410
-------
Annual Costs. Annual costs for wastewater recycle systems were
assumed to include the following: (1) amortization calculated at
10 percent annual interest over 10 years for equipment (CRF =
0.16275) and 30 years for construction (CRF = 0.10608); (2)
annual maintenance at three percent of total capital costs; (3)
operator manhours calculated at $13.50 per hour (including
benefits, insurance, etc.) for 20 hours per week
(4) energy
75 percent
which the
computed at $0.03/KWH based on pumping horsepower at
efficiency and 75 feet total dynamic head, for
following formulae apply:
Horsepower = Wastewater flow (gpm) x 75 feet
0.75 x 3960
Annual Energy Cost = Horsepower x 0.746 KW/HP x 24 hrs/day
x 365 days/yr x $0.03/KWH;
and (5) additional monitoring at $10,000 per year. Total annual
cost curves for 25, 50, 75, and 100 percent recycle systems are
shown in Figure IX-19.
Evaporation Pond
A lined evaporation pond was costed for the only known dis-
charging uranium mill (Mill 9405). The pond was estimated to
require 380 acres of land area. Land costs were assumed to be
$l,000/acre for this site alone. In addition, the pond was
assumed to be located ten miles from the site for purposes of
costing pump station and piping requirements. Piping distance
was based upon statements of the company concerning the location
of available land. Total capital and annual cost figures for
this pond are documented in Table IX-10.
ACTIVATED CARBON ADSORPTION
Capital Costs
Systems have been costed for activated carbon adsorption of
phenolic compounds. Figure IX-20 provides the equipment defini-
tion for these systems. Carbon contactor vessels are constructed
of carbon steel. A backwash system is provided to remove sus-
pended solids from the carbon contactors.
Carbon contactors are designed for 30-minute retention time and
(100 Ibs) of carbon for 0.23 kg (0.5 Ibs) of phenol. A total
phenol (4AAP) concentration of 0.4 mg/1 was assumed for system
sizing.
Total capital costs included equipment, installation, and con-
tingency. Figure IX-21 graphically represents this capital cost
as a function of hydraulic flow rate.
Annual Costs
411
-------
Annual costs for activated carbon adsorption include capital cost
amortization, maintenance, operation, energy, taxes and
insurance, and off-site regeneration of carbon. Amortization was
calculated at 10 percent annual interest rate over a 10 year life
expectancy for equipment (CRF = 0.16275) and a 30 year life
expectancy for construction (CRF = 0.10608). Annual maintenance
costs were estimated at three percent of the initial capital
investment. Operator manhours were established at 2,000 hours
per year and costed at $13.50/hour including benefits. Activated
carbon costs were based on a price of $0.50/lb.; energy was esti-
mated at $0.03/KWH (3 cents); and taxes and insurance were esti-
mated at two percent of the initial capital investment.
Figure IX-22 is a graphic display of the annual costs associated
with activated carbon adsorption of phenolic compounds.
HYDROGEN PEROXIDE TREATMENT
Capital Costs
Cost estimates have been prepared for systems which oxidize
phenolic compounds by the addition of hydrogen peroxide in the
presence of ferrous sulfate catalyst. The design assumptions
included the use of a 6:1:1 ratio of hydrogen peroxide:ferrous
sulfate:phenol. The total phenol (4AAP) concentrations was
assumed to be 0.4 mg/1.
Figure IX-23 is a schematic flow diagram of the system design.
Oxidation basins are sized for five-minute retention time.
Mixers assisted by an air buffing system are provided in the
include air compressor, oxidation basins, mixers, clarifiers,
sludge pumps, hydrogen peroxide storage tank, reagent pumps,
instrumentation and localized piping as well as installation and
contingency costs.
Figure IX-24 relates total capital costs
oxidation systems to hydraulic flow rate
for hydrogen peroxide
rate.
Annual Costs
Annual costs associated with hydrogen peroxide oxidation systems
have been estimated. Included in the estimates are capital cost
amortization, maintenance, operation, energy, and chemical costs.
Amortization was based on a 10 percent annual interest rate, a 10
year life expectancy for equipment (CRF = 0.16275) and a 30 year
life expectancy for construction (CRF = 0.10608). Maintenance
costs were estimated at three percent of the initial capital
investment annually. Operator manhours were established as 2,000
hours per year at a rate of $13.50/hour including benefits.
Energy costs were based on a rate of $0.03/KWH. Hydrogen
peroxide was costed at $0.35/lb for a 70 percent concentration,
sulfuric acid at $0.04/lb, and ferrous sulfate at $0.52/lb.
412
-------
Taxes and insurance were estimated at two percent of the initial
capital investment.
Figure IX-25 displays annual costs as a function of hydraulic
flow rate for hydrogen peroxide treatment systems.
CHLORINE DIOXIDE TREATMENT
Capital Costs
Systems for the oxidation of phenolic compounds by chlorine
dioxide addition have been estimated. Design assumptions include
chlorine dioxide dosage of 6 mg/1, retention time of 10 minutes
in the contact tank, and a 30-day reagent storage capacity.
Figure IX-26 is a schematic flow diagram of the system including
reagent storage tank, enclosure metering pump, contact tank, dis-
charge pump, ejector, and filter.
Capital costs include equipment, construction, installation,
localized piping and electrical work, instrumentation and
contingencies. Figure IX-27 graphically displays capital costs
for these systems as a function of hydraulic flow rate.
Annual Costs
Figure IX-28 shows the annual costs associated with chlorine
dioxide oxidation systems as a function of hydraulic flow rate.
Annual costs include capital cost amortization, maintenance,
operation, energy, and chemical costs. Amortization was calcu-
lated at a 10 percent annual interest rate over a 10 year life
expectancy for equipment (CRF = 0.16275) and a 30 year life
expectancy for construction (CRF = 0.10608). Maintenance costs
were estimated at three percent of the initial capital investment
annually. Operator manhours were estimated at 2,000 hours per
year at a rate of $13.50/hour including benefits. Energy costs
were based on a rate of $0.03/KWH, (3 cents). Chlorine dioxide
costs were estimated as $9.10/gal (5 percent cone.). Taxes and
insurance were estimated to be two percent of the initial capital
investement.
POTASSIUM PERMANGANATE OXIDATION
Capital Costs
Cost estimates have been prepared for the installation of systems
which oxidize phenolic compounds by the use of potassium
permanganate. The system definition is shown schematically in
Figure IX-29.
Design assumptions include one hour retention time in the
oxidation basins, neutral pH conditions, clarifier overflow rate
permanganate dosage was estimated at 7:1 ratio of potassium per-
413
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manganate to phenolics. A 0.4 mg/1 concentration of total phenol
(4AAP) was assumed.
Capital costs include equipment, installation, construction,
localized piping and electrical work, instrumentation, and con-
tingency costs. Figure IX-30 displays total capital costs as a
function of hydraulic flow rate.
Annual Costs
Capital recovery was amortized over a 10-year period for
equipment and a 30 year period for construction. A 10 percent
annual interest rate was used (Equipment CRF = 0.16275,
Construction CRF = 0.10608). Annual maintenance costs were
assumed to be three percent of the capital investment. Operator
manhours of 2,000 hours per year were costed at $13.50/hour
including benefits, etc. Energy costs were estimated at
$0.03/KWH. Potassium permanganate was costed at $0.59/lb (dry).
Taxes and insurance were estimated to be two percent of the total
capital cost. Figure IX-31 shows the total cost estimates for
these systems.
MODULAR TREATMENT COSTS FOR THE ORE MINING AND DRESSING INDUSTRY
Tables IX-2 through IX-10 list unit treatment process costs for
each facility studied. Costs are given in terms of a) Capital
Cost ($1,000), b) Annual Costs ($1,000), and c) Cost: cents/ton
of ore mined.
For purposes of these tabulations, the capital and annual cost
curves of this section to which the additional costs of
monitoring must be added where applicable were used.
NON-WATER QUALITY ISSUES
Solid Waste
Solid wastes generated during the ore mining and milling
processes are currently being investigated by EPA for possible
regulations under the Resource Conservation and Recovery Act
(RCRA). Solid wastes from mining and milling operations include,
but are not limited to: overburden, tailings, mine and mill
wastewater treatment sludges, lean ore, etc. The EPA has
sponsored several studies (References 4, 5, and 6) in response to
Section 8002, p and f of RCRA. These studies have examined the
sources and volumes of solid wastes generated, present disposal
practices, and quality of leachate generated under test condi-
tions. To date, leachate tests have been performed on approxi-
mately 370 ore mining and milling solid wastes. Solid wastes
from all of the ore mining and dressing subcategories have been
examined and only 11 samples (approximately three percent) were
found which exceeded the RCRA EP (extraction procedure) criteria
(References 4 and 5). The vast majority (approximately 97
414
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percent) of the ore mining and milling solid wastes are not
hazardous (EP toxic).
In addition, Section 7 of the Solid Waste Disposal Act Amendments
of 1980 has exempted, under Subtitle C of the RCRA, solid waste
from the extraction, beneficiation, and processing of ores and
minerals. This exemption will remain in effect until at least
six months after the administrator submits a study on the adverse
environmental effects of solid wase from mining. This study is
required to be submitted by 21 October 1983.
415
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TABLE IX-1. COST COMPARISONS GENERATED ACCORDING TO TREATMENT PROCESS AND ORE CATEGORY
Iron Ore
Copper Ore
Lead/Zinc Ores
Gold/Silver Ores
Aluminum Ore
Ferroalloy Ores
Mercury Ore
Titanium Ore
Uranium Ore
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Miils
Mines
Mines
Mills
Mines
Mills
Mine/Mill
Mines
Mills
If
V) CO
X
X
X
X
X
X
X
X
X
X
X
X
X
X
Fiocculant
Addition
X
X
X
X
X
X
-X
X
X
X
X
X
X
Ozonation
I
X
X
X
X
X
X
X
Alkaline
Chlorination
X
X
X
X
X
X
X
Ion Exchange
X
X
X
X
X
Granular Media
Filtration
X
X
X
X
X
X
X
X
X
X
X
X
X
pH Adjustment
X
X
X
X
X
X
X
X
X
X
X
RECYCLE
25%
X
X
X
X
50%
X
X
X
X
75%
X
X
X
X
100%
X
X
X
X
X
Activated
Carbon
X
X
X
Hydrogen
Peroxide
X
X
X
Chlorine
Dioxide
X
X
X
Potassium
Permanganate
X
X
X
c
o
'i
a«
S 2 §
KLU £
X
-------
TABLE !X-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 !bs)
p. 1 of 9
Type of Mine: Iron Ore
i-line
Code -
Location
Type
1101-MN
Mine
1101-MN
Mill
1102-MN
Mine
110i!-MN
Mill
1103-MN
i-line/Mill
11U4-MN
Mine
1104-MN
Mill
Ore
Production
(1000 torcv
year)
36,376
36,376
^,072
9,072
4,409
1,808
1,808
Ua ter
Uis-
charqed
(MOD)
21.13
0
0.69
0
0
1.05
5. 94
a, Capital Cost ($1000)
TREATMENT TECHNOLOGIES AW COSTS: b. Annual Cost ()1000)
c. Cost: t/ton of ore mined
Second.
Settling
a. 340
b. 44.8
c. 0.12
a.
b. -
c.
a. 84
b. 19.0
c. 0.21
a.
b. -
c.
a.
b. -
c.
a. 97
h. 20.2
c. 1-12
a. 187
b. 29
c. 1.60
Floc-
cula-
tion
110
160
0.44
-
65
30
0.33
-
-
70
32
1.77
90
58
3.21
Ozon-
ation
-
-
-
-
-
-
-
Alkal.
fhlfif
ination
-
-
-
-
-
-
-
Ion
-
-
-
-
-
-
-
Mixed
Media
Hltr.
2000
310
0.85
-
150
47
0.52
-
-
206
55
3.04
808
140
7.74
pH
-
-
-
-
-
-
-
R e c y c 1 e
25X
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
Vocess
Control
-
-
-
-
-
-
-
REMARKS
-------
CO
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. "2 of 9
Type of Mine: Iron Ore
Mine
Code -
Location
Type
1105-MN
1105-HN
Mill
1106-MN
Mine/Mil
1107-MI
Mine
1107-MI
Mill
1 108-MI
Mine
1 108-MI
Mill
1 109-MI
Mine
Ore
Production
(1000 ton;,
year)
9,149
9,149
44,092
5,842
5.842
9.700
9.700
18.078
Hater
Dis-
charged
(MOO)
12.70
0
0
3.53
2.69
4.17
8.71
N.A.
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMO COSTS: b. Annual Cost (UOOO)
c. Cost:
-------
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
p. 3 of 9
Mine
Code -
Location
Type
1109-MI
Mill
1110-PA
Mine
1110-PA
Mill
1111-MN
Mine
1112-MN
Mine
1112-MN
Mill
1113-MN
Mine/mill
1114-MO
Mine
Ore
Productior
(1UOO ton*
year)
18,078
2.866
2.866
34.172
9.590
9.590
27,558
2,601
Water
Dis-
charged
(MOD)
5.94
N.A.
1.71
14.00
7.00
0
0
1.43
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AM1) COSTS: b. Annual Cost (HOOO)
c. Cost: (/ton of ore mined
Second.
Settling
a. 186
b. 29.2
c. 0.16
a.
b. -
c.
a. 115
b. 22
c. 0.77
a. 280
b. 38.5
c. 0.11
a. 210
1). 31.5
c. 0.33
a.
b. -
c.
a.
b. -
c.
a. 107
l>- 21.1
c. 0.81
Floc-
cula-
tion
90
58
0.32
-
75
37
1.29
100
110
0.32
91
65
0.68
-
-
72
35
1.34
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
800
140
0.77
-
300
70
2.44
1500
230
0.67
900
150
1.56
-
-
253
62
2.38
PH
-
-
-
-
-
-
-
-
Recycle
25*
-
-
-
-
-
-
-
-
50*
-
-
-
-
-
-
-
-
75*
-
-
-
-
-
-
-
-
100*
-
-
-
-
-
-
-
-
'rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS
-pi
1'
VO
-------
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
p. 4 of 9
Mine
Code -
Location
Type
1114-MO
Hiti
lilS-MO
Mine
1115- MO
Mill
1116-HJ
Mine/Mil!
1117-SJT
Mine /Mill
1518-CA
Mine/Mill
Ore
Production
(1000 tonv
year)
2,601
2. 425
2,425
2.425
2,645
9,028
Ha ter
i)is-
charqed
(MGD)
1.71
NA
4.17
0
MA
0
a. Capital Cost ($1000)
THtAlMfNT TECHNOLOfilES AM') COSTS: b. Annual Cost (HOOO)
c. Cost: I/ton of ore mined
Second.
Settling
A. 115
b. 22
c. 0.85
a.
b. -
c.
a- 161
b. 26.5
c- 1.09
a.
b. .
c.
a.
b. .
c.
a.
b. -
c.
Floc-
cula-
tion
75
37
1.42
-
85
50
2.06
-
-
Ozon-
ation
-
-
-
-
-
-
Alkal,
Chlor-
inatiori
-
-
-
-
-
-
Ion
-
-
-
-
-
-
Mixed
Media
Filtr.
300
70
2.69 ,
-
605
110
4.54
-
-
-
PH
Adjust
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
50%
-
-
-
-
-
-
75%
-
-
-
-
-
-
100%
-
-
-
-
-
-
'rocess
Control
-
-
-
-
-
-
REMARKS
o
-------
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 !bs)
Type of Mine: Iron Ore
p. 5 of 9
iline
Code -
Location
Type
1119-WY
Mine
1119-HY
Mill
1120-MI
Mine
1120-MI
Mill
1121-MN
Mine
1121-MN
Mill
1122-MN
Mine
1122-MN
Mill
Ore
Production
(1UOO tons/
year)
4.850
4.850
4.630
4,630
1.194
1.194
8.157
8.157
Water
Dis-
charged
(MGO)
0.45
Minimal
0.18
Minimal
5.94
1.44
17.80
1
0
a. Capital Cost (!>1000)
TREATMENT TECHNOLOGIES AMO COSTS: b. Annual Cost (>1000)
c. Cost:
-------
ro
ro
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 6 of 9
Type of Mine: Iron Ore
itine
Code -
Location
Type
1123- MN
Mine
1123-MN
Mill
1124-MN
Mine
1124-MI
Mill
1125-MN
Mine/Mill
1126-MN
Mine/Mill
1127- UT
Mine/Mill
1128-NM
Mine/Mill
Ore
Product lor
(1UOO tons/
year)
2,535
2,535
11,905
11.905
1,543
2,425
1,874
71.65
Ha ter
Uls-
charqed
/MfifM
2.98
0
2.98
0
NA
NA
0
0
a. Capital Cost ($1000)
TREATMENT TECHNOLORIF.S AMfl COSTS: b. Annual Cost (>1000)
c. Cost:
-------
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn 2000 Ibs)
Type of Mine: Iron Ore
p. 7 of 9
dine
Code -
Location
Type
1129-TX
Mine
1129-TX
Mill
11 30- NY
Mine/Mill
1131-NY
Mine
1131-NY
Mill
1132-WY
Mine/Mill
1133-MN
Mine
1134-MN
^line/Mill
Ore
Production
(1UOO tons/
year)
2.380
2,380
1,984
3.858
3,858
1,433
0
1,433
Hater
Dis-
charged
(MGO)
NA
0
NA
0.44
16.36
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (>1000)
c. Cost: t/ton of ore mined
Second.
Caff] Inn
a.
b. .
c.
a.
b. .
c.
a.
b. .
c.
a. 74
b. 18.1
c. 0-47
a. 306
b. 38
c. 0.98
a.
b. -
c.
a.
b. -
c.
a.
b. -
C.
Floe-
tlon
_
.
.
60
28
0.73
100
110
2.85
-
-
-
Ozon-
_
_
_
-
-
-
.
Alkal.
fhlnr-
1 nation
_
_
_
-
-
-
-
Ion
_
_
.
-
-
-
-
Mixed
MnHia
Filtr.
_
«
_
108
41
1.06
1650
250
6.48
-
.
.
PH
_
_
m
-
_
.
Recycle
25%
_
_
_
-
-
.
_
50*
_
-
.
.
.
75t
.
_
-
-
.
,
1001
-
-
_
.
'rocess
ontrol
^
-
-
,
.
REMARKS
Operation
closed
OJ
-------
TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
p. 8 of 9
Mine
Code -
Location
Type
1135-MN
Mine/
Mill
1136-MI
kline
IH37-CA
Mine/
Mill
1138-MN
Mine/
Mill
1139-GA
Mine
1140-MN
Mine
1141-MN
Mine
1142-MN
Mine/
Mill
Ore
Production
(1UOO tons/
year)
1,212
301
496
9735
0
0
0
NA
Water
Dis-
charged
(MGO)
NA
120.46
0
0
NA
NA
NA
1
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES ANO COSTS: b. Annual Cost (J1000)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a. 820
b. 96
c.31.89
a.
b. -
c.
3.
b. -
c.
a.
b. -
c
a.
b. -
c.
a.
b. -
c.
o ' ii. .
' r. .
Floc-
cula-
tlon
-
140
860
285. 71
-
-
-
-
-
-
Ozon-
atlon
-
-
-
-
-
-
' -
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
:xchan.
-
-
-
-
-
-
-
Mixed
Media
Filtr,
-
6000
1060
352.16
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50«
-
-
-
-
-
-
-
-
75$
-
-
-
-
-
-
_
100%
-
-
-
-
-
-
-
'rocess
Control
-
-
-
-
-
-
REMARKS
Operation
assumed
closed
i;
It
-ft.
-------
TABLE IX-2, COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 !bs)
Type of Wne: Iron Ore
p. 9 of 9
-
s-iine
Code -
location
Typt
114 3- MN
Mine/
Mill
1I44-MI
Mine
1145-NV
Mine
1146-MN
Mine
1147-MN
Mine
1148-MN
Mill
1149-MN
Mill
i | a. Capital Cost ($1000)
Ore | Hater
Production Uis-
, , - i charged
(lUCO tonyl
vesr)
2,648
1,874
115
661
413
1297
0
(MGD)
1.06
3.25
0.63
10.00
2.19
0
NA
TREATMENT TECHNOLOGIES AMH COSTS: b. Annual Cost (HQOO)
c. Cost: i/ton of ore mined
Second.
fettling
a. 98
b. 20.2
c. 0.76
a, 148
b. 25.2
c 1.34
a. 83
b. 18.8
C. 16. 35
a, 240
b. 34.6
c. 5.23
a. 126
b. 23
c. 5.57
a.
b. -
c.
a.
b. -
,c'
b.
'-
Flcc-
cuU
tion
70
32
1.21
82
43
2.30
64
28
24. 35
98
82
12.41
79
37
8.96
-
Ozon-
aticn
,
-
-
-
-
-
Alkal.
Chlor-
Ion
Lxchan.
1 nation!
~1
-
-
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
200
55
2.08
500
95
5.07
145
46
40.00
1200
185
27.99
360
79
19.13
-
PH
Adjust.
-
-
-
-
-
Recycle
25%
-
-
-
-
-
50%
-
-
-
-
-
75% j
-
-
-
-
-
100%
-
-
-
-
-
Process
Control
-
-
-
-
-
i
!
REMARKS
I
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Bate and Precious
Metals (Copper)
P.I of 7
Mine Code -
Location Type
2101- NV
Mine/Mill
2102-AZ
Mine/Mill
2103-NM
Mine/Mill
2104-NM
Mine/Mill
2107-AZ
Mine/Mill
2108-AZ
Mine/Mill
2109-AZ
Mine/Mill
2110 AZ
Mine
Ore
Production
(1000
tons/year)
7,932
6.015
15,403
8.101
4,402
3,066
3,729
4,090
Water
Discharged
(MGD)
0
0
0
0.18
0
0
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Co»t ($10001
ta. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 61
b. 17
<= 0.21
a. 61
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
55
27
0.33
-
-
-
-
Ozona
tion
-
-
-
32
27
0.33
-
-
-
-
Alkal.
Chlorin-
ation
-
' -
-
48
27
0.33
-
-
-
-
Ion
Exchan.
-
-
-
850
250
3.09
-
-
-
-
Mixed
Media
Filtr.
-
-
-
50
33
0.41
-
-
-
-
pH
Adjust.
-
-
-
26
23
0.28
-
-
-
-
Recycle
25%
-
-
-
7
16
0.20
-
-
-
-
50%
-
-
-
9.5
16.5
0.20
-
-
-
-
75%
-
-
-
13
17
0.21
-
-
-
-
100%
-
-
-
17
18.5
0.23
-
-
-
-
Activated
Carbon
-
-
-
120
67
0.83
-
-
-
-
Hydrogen
Peroxide
-
-
-
140
91
1.12
-
-
-
-
Chlorine
Dioxide
-
-
-
108
83
1.02
-
-
-
-
Potaulum
Permang-
anate
-
-
-
160
99
1.22
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Operation
inactive
Operation
presently
inactive
-pa
ro
CTi
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Base and Precious
Metals (Copper)
.p. 2 of 7
Mine Code
Location Type
2111-AZ
Mine/Mill
2112-AZ
Mine/Mill
2113-AZ
Mine/Mill
2115-AZ
Mine/Mill
2116-AZ
Mine/Mill
2117-TN
Mill
2118-AZ
Mine/Mill
2119-AZ
Mine/Mill
Ore
Production
(1000
tons/year)
1,631
670
10,340
1,555
9.804
2,024
18,357
15,013
Water
Discharged
(MOD)
NA
0
0
0
0
8.50
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 220
b. 32
c. 1.58
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
95
75
3.71
-
-
Ozona-
tion
-
-
-
-
-
540
157
7.76
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
230
237
11.71
-
-
ton
Exchan.
-
-
-
-
-
11000
2760
136.36
-
-
Mixed
Media
Filtr.
-
-
-
-
-
1050
180
8.89
-
-
pH
Adjust.
-
-
-
-
-
58
70
3.46
-
-
Recycle
25%
-
-
-
-
-
60
31
1.S3
-
-
50%
-
-
-
-
-
90
43
2.12
-
-
75%
-
-
-
-
-
130
57
2.81
-
-
100%
-
-
-
-
-
170
70
3.45
-
-
Activated
Carbon
-
-
-
-
-
1650
805
39.77
-
-
Hydrogen
Peroxide
-
-
-
-
-
330
195
9.63
-
-
Chlorine
Dioxide
-
-
-
-
-
285
185
9.14
-
-
Potassium
Permang-
anate
-
-
-
-
-
1100
345
17.05
-
-
Process
Control
-
-
-
-
-
-
-
Remarks
Inactive
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mind:
Base and Precious
Metals (Copper)
p. 3 of 7
Mine Code -
Location Type
2120-MT
Mine
2120-MT
Mill
2121-Mi
complex
2122-UT
Mill
2123-AZ
Mine/Mil!
2124AZ
Mine/Mill
2125-AZ
Mine
2126-NV
Mine/Mill
Production
tons/year)
17,000
17,000
3,617
35,500
2,047
6,710
o
8,000
Water
(MCD)
0.05
9.50
32
8.50
0
0
0
0
TREATMENT TECHNOLOGIES AND COSTS: ». Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton 01 ore mined
Settling
a. 58
b. 17
c. 0.10
a. 235
b. 34
c. 0.20
a 400
b. 50
c. 1.38
a. 220
b. 32
c. 0.09
a.
b. -
c.
a.
b. -
c.
a.
b
c.
a.
b. -
c.
lation
45
26
0.15
97
80
0.47
120
210
5.81
95
75
0.21
-
-
tion
20
24
0.14
600
177
1.04
1800
470
1299
540
157
0.44
-
-
Alkal.
at ion
42
25
0.15
260
257
1.51
610
737
20.38
230
237
0.67
_
-
-
Exchan.
-
12000
3110
18.29
42000
13010
359.69
11000
2760
7.77
_
-
-
Mixed
Filtr.
18
28
0.16
1150
190
1.12
2500
400
11.06
1050
180
0.51
_
-
-
PH
Adjust.
23
22
0.13
70
75
0.44
125
190
5.25
68
70
0.20
_
-
--
Recycle
25%
-
65
33
0.19
1S5
70
1.94
60
31
0.08
_
-
-
50%
-
100
46
0.27
270
125
3.46
90
43
0.12
_
-
-
75%
-
145
62
0.36
380
170
4.70
130
57
0.16
_
-
-
100%
-
180
75
0.44
485
230
6.36
170
70
0.20
__
-
-
Carbon
-
1800
905
5.32
5200
2705
74.79
1650
805
2.27
_
-
-
Peroxide
-
340
205
1.21
580
400
11.06
330
195
0.55
_
-
-
Dioxide
-
300
195
1.15
540
390
10.78
285
185
0.52
_
-
-
Potassium
nate
-
1200
360
2.12
3200
820
22.67
1100
345
0.97
_
-
-
Control
-
-
-
-
_
-
-
Remarks
Already
meeting
BAT.
j
|
Temporarily !
inactive
j
5
ro
co
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Typa of Mine: Base and Precious
Metals (Copper)
p. 4 of 7
Mine Code
Location Type
2130-NM
Mine/Mill
2131 -NV
Mine
2132-NV
Mine/Mill
2133-NV
Mine/Mill
2134-ID
Mine
21 34-1 D
Mil!
213S-AZ
Mine
2136-AZ
Mine
Ore
Production
11000
tons/year)
NA
NA
NA
0
NA
NA
8.27
NA
Water
Discharged
IMGDI
Minimal
0
0
0
0
NA
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alttal.
Chlorin
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
pH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Mine dis-
charges to
mill - very
small or zero
discharge
Closed
permanently
Partial
recycle
Operation
inactive
ro
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Base and Precious
Metals (Copper)
p. 5 of 7
Mine Coda
Location Type
2137-AZ
Mine/Mill
2138-AZ
Mine/Mill
2139-AZ
Mine/Mill
2140-AZ
Mine/Mill
2141-AZ
Mine/Mill
2142-AZ
Mine
2143-AZ
Mine
2144-AZ
Mine
Ore
Production
(1000
tons/year)
NA
5,000
32,494
5,800
5,300
1,820
NA
NA
Water
Discharged
IMGD)
0
0
0
0
0
0
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: 4 /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin
at ion
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Suspected
inactive
GO
o
-------
TABLE »X-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Bat* and Precious
Metals (Copper)
p. 6 of 7
Min* Cod*
Location Type
2145-AZ
Mine/Mill
2146-AZ
Mine/Mill
2147-A2
Mine/Mill
2148-AZ
Mine/Mill
2149-AZ
Mine
2150-UT
Mine/Mill
2151 Ml
Mine/Mill
2152-NM
Mill
Ore
Production
(10OO
tons/year)
NA
NA
19,600
NA
NA
NA
NA
0
Water
Discharged
(MGOI
0
0
0
0
0
NA
NA
NA
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: *" /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Temporarily
inactive
Probably
inactive
Under devel-
opment -
0 discharge
likely
Pilot-scale
production
Temporarily
inactive
00
-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED {1 tn = 2000 Ibs)
Type of Mint:
But and Procious
Metals (Copper)
p. 7 of 7
Mine Cod*
Loulion Type
2164-AZ
Mini
Ore
Production
11000
toni/year)
NA
Water
Discharged
(MOD)
NA
TREATMENT TECHNOLOGIES AND COSTS: ». Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
Flo ecu-
lation
-
Ozona-
tton
-
Alkal.
Chlotin
at ion
-
Ion
Exchan.
-
Mixed
Media
Filtr.
-
PH
Adjust.
-
Recycle
25%
-
50%
-
75%
-
100%
-
Activated
Carbon
-
Hydrogen
Peroxide
-
Chlorine
Dioxide
-
Potassium
Permang-
anate
-
Process
Control
-
Remarks
Under
develop-
ment or
exploration
-p.
CO
INJ
-------
TABLE i/ 4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
Type of Mine:
Bass and Precious
Metals (Lead-Zinc)
p. 1 of 8
Location Type
3101
Mine/Mill
3102-MO
Mine/Mill
3103-MO
Mine/Mill
3104-NY
Mill
3105-MO
Mine
3196-PA
Mine
3106-PA
Mill
Production
tons/year)
206
1,634
1,072
1,112
1,138
383
383
3107-ID
Complex
732
Water
JMGDI
0.38
5.94
2.58
1.78
2.19
28.53
1.50
5.94
VR-.ATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: f /Ion or ore mined
Sealing
a. 70
b. 17.5
c. 8.50
a.180
b. 29
c. 1.77
a. 135
b. 24
c. 2.24
a. 115
b. 22
c. 1 .98
a. 124
b. 23
c. 2.02
a. 390
b. 48
c. 12.53
8.110
b. 21.4
c. 5.59
8.180
b. 29
c. 3.71
lation
60
27
13.11
90
60
3.67
80
40
3.73
75
37
3.33
78
38
3.34
too
190
4S.61
75
35
9.14
90
60
7.67
tion
50
35
16.99
400
122
7.47
195
74
6.90
145
59
5.31
180
66
5.80
1680
422
110.18
130
54
14.10
400
122
15.60
Alkal.
ation
54
32
15.53
195
172
10.5
120
85
7.92
92
65
5.85
110
75
6.59
585
667
174.15
88
59
15.40
195
172
21.99
Exchan.
1200
330
160.19
7500
2210
135.25
4000
1210
112.87
3000
860
77.34
-
-
2700
760
198.43
7500
2210
282.61
Mixed
Filti.
90
40
19.42
800
140
8.57
400
83
7.74
300
70
6.29
340
75
S.59
2400
370
96.61
260
65
16,97
800
140
17.90
PH
Adjust.
30
25
12.14
60
56
3.43
45
38
3.54
42
34
3.06
43
36
3.16
115
170
44.38
41
33
8.62
60
56
7.16
Recycle
25%
9.5
16
7.77
50
27
1.65
30
22
2.05
23
20
1.80
-
-
21
20
5.22
50
27
3.45
50%
14
16.5
8.01
73
37
2.26
44
27
2.58
32
24
2.16
-
-
30
23
6.00
73
37
4.73
75%
18
17.5
8.50
100
47
2.88
60
31
2.89
44
27
2.43
-
-
40
25
6.53
100
47
6.01
100%
24
19.5
9.47
130
57
3.49
75
36
3.36
56
30
2.70
-
-
52
28
7.31
130
57
7.29
Carbon
-
1260
600
36.72
700
305
28.45
540
235
21.13
-
-
460
205
53.52
1260
600
76.73
Peroxide
-
290
165
10.10
225
130
12.13
203
120
10.79
-
-
195
115
30.03
290
165
21.10
Dioxide
-
250
160
9,79
190
125
11.66
170
115
10.34
-
_
168
110
28.72
250
160
20.46
Potassium
anate
-
870
270
16.52
520
185
17.26
410
160
14.39
-
-
380
150
39.16
870
270
34.53
Control
-
-
-
-
-
-
-
-
Remsrfcs
Closed
CO
00
-------
TABLE 1X4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Ban and Precious
Metals (Lead-Zinc)
p. 2 of 8
Mini Cod« -
Location Type
3108-TN
Mill
3109 -MO
Mine/Mill
3110-NY
Mill
3111-TE
Mine
3112-NM
Mine
3113-CO
Mine
3113-CO
Mill
311 4-1 D
Mine/Mill
Production
(1000
tons/year)
391
1,117
103
100
135
203
203
68
Water
Discharged
(MGD)
0.05
7.50
0.58
0.95
0.66
1.69
1.40
0.42
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cod ($1000)
c. Cost: i /ton or ore mined
Second.
Settling
a. 55
b. 16.5
c. 4.22
a. 210
b. 32
c. 2.86
a. 80
b. 18.6
c. 18.0C
a. 94
b. 19.9
c. 19.90
a. 84
b. 18.9
c. 14.00
a. 112
b. 22.0
c. 10.84
a. 105
b. 21.0
c. 10.34
a. 74
b. 18.0
c. 26.47
Floccu
lation
45
24
6.14
95
67
6.00
68
28
27.18
70
32
32.00
65
30
22.22
75
34
16.75
75
34
16.75
61
27
39.71
Ozona-
tion
20
24
6.14
490
147
13.16
64
36
34.95
92
43
43.0
69
38
28.15
140
57
28.08
120
53
26.11
52
33
48.53
Alkal.
Chlorin
ation
42
25
6.39
220
217
19.43
62
36
34.95
72
47
47.00
64
39
28.89
92
62
30.54
85
58
28.57
56
33
48.53
Exchan.
460
150
38.35
9000
2610
233.66
1600
450
436.89
_
2700
810
399.01
2500
730
359.61
1400
360
529.41
Mixed
Fillr.
18
28
7.16
930
155
13.88
140
46
44.66
190
54
54.00
145
47
34.81
280
67
33.00
250
61
30.05
95
41
60.29
PH
Adjust.
23
22
5.63
66
65
5.82
33
27
26.21
36
28
28.00
34
26
19.26
42
33
16.26
40
32
15.76
31
26
38.23
Recycle
26%
3.5
15
3.84
55
30
2.69
13
17
16.50
_
23
20
9.85
20
19
9.36
11
16
23.53
50%
5
16
4.10
80
43
3.85
17
18
17.48
_
32
23
11.33
28
22
10.84
15
16.5
24.26
75%
7
16.5
4.22
125
55
4.92
24
19
18.4!
^
-
45
27
13.30
39
25
12.32
19
17.5
25.74
100%
8
17
4.35
150
66
5.91
30
22
21.36
-
55
30
14.78
49
28
13.79
24
20
29.41
Carbon
52
47
12.02
1500
725
64.91
250
115
111.65
_
-
-
450
195
96.06
200
97
142.65
Peroxide
124
85
21.74
315
186
16.56
160
99
96.12
_
-
-
190
115
56.65
155
95
139.71
Dioxide
86
75
19.18
275
180
16.11
134
94
91.26
_
-
-
165
108
53.20
126
92
135.29
Potaeiium
anate
125
89
22.76
1030
320
28.65
240
120
116.50
_
-
-
360
150
73.89
205
115
169.12
Control
-
-
-
-
-
-
-
Remark*
OJ
-------
TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Base and Precious
Metals (Lead-Zinc)
p. 3 of 8
Location Type
3115
Mine/Mill
3116-CO
Mine
3118-VA
Mine
3118-VA
Mine
3118-VA
Mill
3118-VA
Mill
3118-VA
Mine/Mill
3119-MO
Mine/Mill
Or*
Production
tons/year)
372
198
596
596
596
596
596
647
Water
(MGDI
4.7
0.87
1.80
2.60
0.01
0.20
14.00
1.78
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: 4 /ton or ore mined
Settling
a. 160
b. 26
c. 6.99
a. 91
b. 19.6
c. 9.90
a. 118
b. 22.1
c. 3.71
a. 135
b. 23.7
c. 3.98
a. 55
b. 16.5
c. 2.77
a. 60
b. 16.8
c. 2.82
a. 280
b. 38
c. 6.38
a.115
b. 22.0
c. 3.40
lation
90
51
13.71
69
32
16.16
75
35
5.87
80
40
6.71
35
25
4.19
55
26
4.36
100
110
18.46
75
35
5.41
tion
320
98
26.34
84
42
21.21
146
60
10.07
196
75
12.58
15
21
3.52
33
27
4.53
860
242
40.60
145
59
9.12
Alkal.
ation
166
147
39.52
70
45
22.73
96
67
11.24
120
86
14.42
42
25
4.19
48
28
4.70
350
357
59.89
92
65
10.04
Exchan.
6000
1610
432.80
-
2900
810
135.91
3900
1110
186.24
330
120
20.13
880
250
41.95
16000
4510
756.71
2900
810
125.19
Mixed
Filtr.
640
120
32.26
180
51
25.76
310
72
12.08
400
88
14.77
10
17
2.85
54
33
5.54
1500
230
38.60
310
72
11.13
PH
Adjust.
53
50
13A4
36
28
14.14
42
34
5.70
46
38
6.38
18
22
3.69
28
23
3.86
85
100
16.78
42
34
5.26
Recycle
25%
43
25
6.72
23
20
3.36
30
22
3.69
1.7
15
2.52
6.9
16
2.68
87
40
6.71
23
20
3.09
50%
60
32
8.60
32
23
3.86
42
26
4.36
2.5
16
2.68
9.5
16.5
2.77
130
59
9.90
32
23
3.55
75%
84
39
10.48
45
26
4.36
58
30
5.03
3.1
16.5
2.77
12
17
2.85
170
80
13.42
45
26
4.02
100%
110
45
12.10
60
30
5.03
73
36
5.87
4.2
17
2.85
16
18
3.02
240
100
16.78
60
30
4.64
Carbon
-
-
-
20
40
6.71
125
70
11.74
2500
1255
210.57
540
235
36.32
Peroxide
-
-
-
120
85
14.26
140
90
15.10
400
245
41.11
203
120
18.55
Dioxide
-
-
-
70
70
11.74
110
82
13.76
350
240
40.27
170
115
17.78
Potassium
anat*
-
-
115
85
14.26
165
100
16.78
1650
465
78.02
410
160
24.73
Control
-
-
-
-
-
-
-
Remarks
Closed.
Annual pro-
duction based
on 250
working days
Ore produc-
tion includes
Mine 31 17
Ore produc-
tion includes
Mine 31 17
00
en
-------
TABLE 1X4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Typs of
Base and Precious
Metals (Lead-Zinc)
p. 4 of 8
Location Type
31 20- ID
Mine/Mill
3121-ID
Mine
3121-ID
Mill
3122-MO
Mine
3123-MO
Mine
3124-NJ
Mine
3125-NY
Mine
3127-TN
Mine
Ore
Production
tons/year)
174
283
283
i,111
1,774
205
24
721
Water
(MGD)
1.24
1.24
1.58
6.90
9.60
0.25
0.37
1.45
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: 4 /ton or ore mined
Settling
a.102
b. 20.7
c. 11.90
a. 102
b. 20.7
c. 7.31
a. 111
b. 21.6
c. 7.63
a. 205
b. 30.9
c. 2.78
a. 230
b. 34.0
c. 1.92
a. 64
b. 17.3
c. 8.44
a. 72
b. 17.8
c.74.17
a. 108
b. 21.2
c. 2.94
lation
71
33
18.97
71
33
11.66
75
34
12.01
95
65
5.85
98
80
4.51
57
27
13.17
60
28
116.67
75
34
4,72
tion
110
49
28.16
110
49
17.31
136
55
19.43
460
139
12.51
600
177
9.98
37
29
14.15
47
32
133.33
125
53
7.35
Alkal.
ation
80
52
29.88
80
52
18.37
88
60
21.20
210
197
17.73
260
257
14.49
50
30
14.63
54
31
129.17
85
56
7.76
Exchan.
2300
650
373.56
-
2900
780
275.62
_
-
-
-
-
Mixed
Filtr.
230
60
34.48
230
60
21.20
270
66
23.32
900
160
14.40
1165
191
10.77
65
35
17.07
90
38
158.33
255
63
8.74
pH
Adjust.
39
31
17.82
39
31
10.95
41
34
12.01
63
62
5.58
71
76
4.28
28
23
11.22
30
24
100.00
40
31
4.29
Recycla
25%
19
19
10.92
-
22
20
7.07
_
-
-
-
-
50%
26
21
12.07
-
30
23
8.13
_
-
-
-
-
75%
37
24
13.79
-
42
26
9.19
_
-
-
-
-
100%
47
27
15.52
-
52
29
10.25
-
-
-
-
Carbon
410
180
103.45
-
500
215
75.97
_
-
-
-
-
Peroxide
190
113
64.94
-
200
120
42.40
_
-
-
-
-
Dioxide
160
104
59.77
-
165
110
38.87
_
-
-
-
-
Potassium
anata
340
145
83.33
-
390
150
53.00
_
-
-
-
-
Control
-
-
_
-
-
-
-
Rtmnrkl
C..)
a?
-------
TABLE 1X^4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Beta and Precious
Metalj (Lead-Zinc)
p. 5 of 8
Mine Cod*
Location Type
3128-TN
Mine
3130-UT
Mine
3131-WI
Ming
3132-W!
Mine
3133-WI
Mine
3133-WI
Mill
3134-WA
Mine
3135-WA
Mine/Mil!
Or*
Production
(1000
toni/yaar)
526
NA
NA
NA
0
0
301
ft
Water
Discharged
(MGD)
1.45
8.50
2.00
1.16
NA
0.76
0.71
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Colt ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.108
b. 21.2
c. 4.03
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
a. 86
b.19.0
c. 6.31
a.
b.
c.
Floccu-
lation
75
34
6.46
-
-
-
-
-
66
30
9.97
-
Ozona-
tion
125
53
10.07
-
-
-
-
-
73
40
1329
-
Alkal.
Chlorin-
ation
85
56
10.64
-
-
-
-
-
65
41
13.62
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Madia
Filtr.
255
63
11.98
-
-
-
-
-
166
50
16.61
-
PH
Adjust.
40
31
5.89
-
-
-
-
-
35
28
9.30
-
Recycle
26%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Inactive
Presently
inactive
Presently
inactive
Presently
inactive
Presently
inactive
* Production
included
with Mine
3134
-p-
OJ
-------
TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Base and Precious
Metals (Lead-Zinc)
p. 6 of 8
Mine Code
Location Type
3136-NV
Mine/Mill
3137-AZ
Mine/Mill
3138-CO
Mine
31 39-1 L
Mill
3140-NM
Mill
3141-TN
Mine
3141-TN
Mill
3142-UT
Mine/Mill
Ore
Production
(1000
tons/year)
126
93.2
98.2
NA
144
0
0
216
Water
Discharged
IMGDI
0
0
NA
0.87
0
0
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($10001
b. Annual Cost ($1000)
c. Cost: i /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Pilot scale
operation
Discharges
twice
yearly
Presently
inactive
Operation
closed
00
00
-------
TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Base and Precious
Metali (Lead-Zinc)
p. 7 of 8
Mine Code
Location Type
3143-CO
Mine
3143-CO
Mill
4103-CO
Mine/Mill
UKA-WI
Mine
UKB-WI
Mine
UKC-TN
Mine/Mill
UKD-TN
UKD-TN
Mill
Ore
Production
tons/year)
60
60
0
NA
NA
NA
NA
NA
Water
(MGD)
NA
0
NA
1.00
1 90
2.00
4.49
1.06
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: i /ton or ore mined
Settling
a.
b. -
c.
a.
b. -
c.
b. -
c.
a.
b. -
c.
a.
.
'
c.
a. 120
b. 22.5
c.
a. 163
b. 27
a. 95
b. 20
c.
lation
-
79
37.5
89
51
70
32
tion
-
155
62
310
103
98
44
Alkal.
ation
-
100
71-
170
137
74
48
Exchan.
-
3200
910
_
2200
610
Mixed
Fillr
-
330
75
630
110
208
58
pH
Adjust.
-
43
35
55
50
38
29
Recycle
25%
-
26
21
17
18
50%
-
37
24
_
24
20
75%
-
50
28
_
32
22
100%
-
63
32
_
41
26
Carbon
-
580
255
375
165
Peroxide
-
210
125
180
105
Dioxide
-
176
117
155
100
Potassium
anata
-
440
165
_
310
135
Control
-
-
Remarks
Operation
closed
Presently
inactive
Presently
inactive
Under-
ground
Mine
Under-
ground
Mine
CO
vo
-------
TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Bass and Preciou*
Metals (Lead-Zinc)
p. 8 of 8
Mine Coda -
Location Type
? (lOO)-TN
Mine/Mill
7 (1021-CO
Mine/Mill
Ore
Production
11000
tons/year)
NA
NA
Water
Discharged
IMGDi
0
NA
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
-
-
Floccu-
lation
-
-
O«> na-
tion
-
-
Alkal.
Chlorin-
ation
-
-
Ion
Exchan.
-
-
Mixed
Media
Filtr.
-
-
PH
Adjust.
-
-
Recycle
25%
-
-
50%
-
-
75%
-
-
100%
-
-
Activated
Carbon
-
-
Hydrogen
Peroxide
-
-
Chlorine
Dioxide
-
-
Potassium
Permang-
anate
-
-
Process
Control
-
-
Remarks
-£»
-£>
O
-------
TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
Base and Precious
Metals (Gold-Silver)
p. 1 of 4
Minn Cod* -
Location Type
4101-NV
Mine/Mill
4102CO
MinQ
4102-CO
Mill
4104-WA
Mine/Mill
4105-SD
Mine
4115-UT
Mine/Mill
4116-AZ
Mino/Mill
I
4117-NV
Mine/Mi!)
i
Of 8
Production
(1000
tons/yearj
320
179
287
55
1,530
145
NA
80
mater
Discharged
(MGO)
0
1.00
0.36
0
3.04
Minimal
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($10001
b. Annual Cost ($1000)
c. Cost: 4 /ton or ore mined
Second.
Settling
a.
b. -
c.
a. 96
b. 20
c. 11.17
a. 72
b. 17.8
c. 6.20
a.
b. -
c.
a. 145
b, 24.7
c. 1.58
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
70
32
17.88
60
28
9.76
82
43
2.76
-
Ozona-
tion
95
43
24.02
50
31
10.80
220
79
5.06
-
Alkal.
Chlorin-
ation
74
47
26.26
53
32
11.15
130
97
6.22
-
Ion
Exchan.
2100
600
335.20
1200
330
114.98
-
-
Mixed
Media
Filtr.
200
55
30.73
90
38
13.24
480
92
5.90
-
PH
Adjust.
36
28
15.64
30
24
8.36
48
44
2.82
-
Recycle
25%
9.5
16
5.57
-
-
50%
15
17
5.92
-
75%
18
18
6.27
-
-
100%
23
20
6.97
-
-
Activated
Carbon
180
88
30.66
-
-
Hydrogen
Peroxide
152
95
33.10
-
-
Chlorine
Dioxide
122
91
31.70
-
-
Potassium
Permang-
anate
195
108
37.63
-
-
Process
Control
-
-
Remarks
-------
TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Base and Precious
Metals (Gold-Silver)
p. 2 of 4
Mine Code
Location Type
4118-NV
Mine/Mill
4119-NV
Mine/Mill
4120-NV
Mine/Mill
4121-NV
Mine/Mill
4122-NV
Mine/Mill
4123-NV
Mine
4124-NM
Mine
4126-AK
Mine/Mill
Ore
Production
(1000
tons/year)
NA
NA
NA
0
0
NA
0
NA
Water
Discharged
(MGD)
0
Minimal
0
0
0
0
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: i /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
pH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remarks
Presently
inactive
Presently
inactive
Temporarily
inactive
-pi
-e*
ro
-------
TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine
Ba*e and Precious
Metalj (Gold-Silver)
p. 3 of 4
Mine Code -
Location Type
4127-AK
Mine/Mill
4128-NV
Mine/Mill
4129-CO
Mine
4129-CO
Mill
4130-NV
Mine
4131-NV
Mine/Mill
4401 -ID
Mine
4401 and
4406-ID
Mills
Ore
Production
11000
tons/year)
612 (m3)
730
NA
0.18-0.36
0
2,004
181
181 (4401)
108 (4406)
Water
Discharged
(MGD)
NA
0
NA
NA
0
0
0.21
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 63
b. 17.1
c. 9.45
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
55
28
15.47
-
Ozona-
tion
-
-
-
-
-
-
34
28
15.47
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
48
28
15.47
-
Ion
Exchan.
-
. -
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
57
34
18.78
-
pH
Adjust.
-
-
-
-
-
-
27
23
12.71
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remark*
Under
exploration
Under
exploration
Inactive
Wastewater
in combined
tailings
pond
-p.
-p>
CO
-------
TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Typa of Mine:
Bale and Pracioul
Metais (Gold-Silver)
p. 4 of 4
Mine Code -
Location Type
4402-CO
Mino
4403-1 D
Mill
4404-CO
Mine
4406-1 D
Mina
4407-ID
Mine
4408-MT
Mine
4409-1 D
Mine
4410-SB
Mine
Ore
Production
(1000
tons/year)
74.4
198
407
108
684
74
NA
84
Water
Discharged
(MGD)
0.78
0.83
1.00
0
NA
0
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Settling
a. 88
b.1 9.4
c. 26.08
a. 90
b.1 9.5
c. 9.85
» 96
b-20
« 4.91
a.
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
lation
68
30
40.32
69
30.5
15.40
70
32
7.86
~
tion
78
40
53.76
85
41
20.71
95
44
10.81
~
Alkal.
at ion
69
43
53.76
74
47
11.55
Exchan.
-
1900
510
257.58
Mixed
Filtr.
165
50
67.20
175
53
26.77
200
55
13.51
pH
Adjust.
35
27
36.29
34
28
14.14
36
28
6.88
Recycle
25%
-
16
17
8.59
50%
-
22
19
9.60
-
75%
-
29
21
10.61
-
100%
-
38
24
12.12
~
-
Carbon
-
320
135
68.18
~
-
-
Peroxide
-
180
105
53.03
~~
-
-
Dioxide
-
145
98
49.50
~
-
-
Potassium
anate
-
282
127
64.14
~~
-
-
Control
-
-
-
-
R*nwki
-------
TABLE IX-6. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Aluminum Ore
P- 1 of 1
Hine
Code -
Location
Type
5101-AR
Mine
5102-AR
Mine
Ore
Production
(1UOO tons.
year)
1,200
872
Hater
Dis-
charged
(MfiO)
1.90
3.67
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AM') COSTS: b. Annual Cost UlOOO)
c. Cost: t/ton of ore mined
Second.
Settling
a. 120
b. 22.4
c. 1.87
a. 152
b. 25.8
c. 2.%
a.
b.
c.
a.
b.
c.
a.
1).
c.
a.
b.
c.
a.
b.
c.
a.
1).
C .
rioc-
cula-
tion
77
37
3.08
85
47
5.39
Ozon-
ation
-
-
Alkal.
Chlor-
ination
-
-
Ion
Exchan.
-
-
Mixed
Media
Filtr.
340
75
6.25
546
102
11.70
PH
Adjust.
-
-
Recycle
25%
-
-
50%
-
-
75%
-
-
100%
-
-
'rocess
Control
-
-
REMARKS
01
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
P- 1 of 8
Type of Mine: Ferroalloy
Mine
Code -
Location
Type
6101-NM
Mill
6102-CO
Mill
6103-CO
Mine
6104- CA
Mine
6105- NY
Mine/Mill
6 106- OR
Mine/
..SlUdUejl
6107-AZ
Mine
6108-NV
Mine/Mill
Ore
Production
(1000 ton*
year)
6,283
15.430
2,425
705
11
1.322
361
NA
Ha ter
Dis-
charged
(MOD)
2.90
2.90
2.87
8.71
0
Minimal
5.54
0
a. Capital Cost ($101)0)
TREATMENT TECHNOLOGIES AMO COSTS: b. Annual Cost (J1000)
c. Cost:
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 2 of
Type of Mine: Ferroalloy
dine
Code -
Location
Type
6109- CA
Mine/Mil
6110-ID
Mine
6111-AK
Mine
6112-NC
Mine/Mill
611 3- NM
Mine
6114-NV
Mine
61 15- CO
Mine/Mill
6116-SC
Mine/ Mi 11
1
Ore
Production
(1000 tons;
year)
16
NA
NA
ca.330
50
0
NA
NA
Ha ter
Uis-
charqed
(MOD)
Minimal
NA
NA
NA
0
NA
NA
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES Am COSTS: b. Annual Cost (HOOO)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
li. -
c.
Floc-
cula-
tlon
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
505!
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
'recess
Control
-
-
-
-
-
-
-
-
REMARKS
Exploratory
operations
underway
Temporarily
inactive-under
exploratior
Exploratory
operations
underway
Inactive
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Ferroalloy
p. 3 of 8
i'1 i ne
Code -
Location
Type
6117-NV
Mine/Mill
6118-MN
Mine
6119-CO
Mine
6120-UT
Mine
6121-NV
Mine
6122-CA
Mine
6123-ID
Mine
6124-CA
Mine
Ore
Production
(1000 tonv
year)
NA
NA
NA
0
NA
ca. 11
NA
ca.33
Ha ter
Dis-
charged
(MGD)
0
NA
NA
0
0
0
NA
NA
a. Capital Cost ($1000)
TKEATMLNT TECHNOLOGIES ACT COSTS: b. Annual Cost (»1000)
c. Cost: {/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
rioc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
Ion
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
I'H
Adjust,
-
-
-
-
-
-
-
-
Recycle
25X
-
-
-
-
-
-
-
-
50X
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
'rocess
Control
-
-
-
-
-
-
-
-
REMARKS
Exploratory
operations
Inactive
Inactive
-p.
-p>
oo
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 4 of 8
Type of Hine: Ferroalloy
rtine
Code -
Location
Type
6125-CA
Mine/Mill
61 26- ID
Mine
6127-10
Mine
6128-CA
Mine
6129-NV
Mine
6130-NV
Mine
6131-CA
Mir.e/Mil
6S32-UT
Mine/Mir
Ore
Production
(1000 tons/
year)
NA
NA
NA
NA
NA
NA
NA
NA
Hater
Uis-
charqed
(MGD)
NA
Inter-
mittent
NA
NA
0
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMI COSTS: b. Annual Cost (HOOO)
c. Cost: tf/ton of ore mined
Second.
Settling
a.
b.
c.
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
a.
b.
c.
a.
b. -
c.
a.
b.
c.
a.
().
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
Ion
:xchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
pH
-
-
-
-
-
-
-
-
Recycle
25$
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100S
-
-
-
-
-
-
-
-
rocess
lontrol
-
-
-
-
-
-
-
REMARKS
Presently
inactive
Exploration
underway
Inactive
Inactive
-
"
MD
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Ferroalloy
p. 5 of 8
Mine
Code -
Location
Type
6133-MT
Mine
6134-ID
Mine
6135-CA
Mine
6136-UT
Mine
6137-UT
Mine
6138-CA
Mine
6139-CA
Mine
6140-CA
Mine
Ore
Production
(1UOO tons;
year)
NA
NA
NA
NA
NA
NA
NA
NA
Ha ter
Dis-
charged
(MGP.)
NA
0
NA
NA
NA
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AM') COSTS: b. Annual Cost (HOOO)
c. Cost: t/ion of ore mined
Second.
Settling
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
a.
b. -
c.
a.
1).
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Cblor-
i nation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25X
-
-
-
-
-
-
-
-
50Z
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
'rocess
Control
-
-
-
-
-
-
-
-
REMARKS
Inactive
Inactive
-pi
en
o
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Ferroalloy
p. 6 of 8
Mine
Code -
Location
Type
6141-CA
Mine
6142-CA
Mine
6143- CA
Mine
6144-CA
Mine
6145- CA
Mine/Mill
6146-CA
Mine/Mill
6147-CA
Mine
6148-NV
Mine/Mill
Ore
Production
(1000 tons-
year)
NA
NA
NA
NA
0.11
0.5
0.11
NA
Ha ter
Uis-
cliarqed
(MGD)
NA
NA
NA
NA
0
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMf) COSTS: b. Annual Cost (HOOO)
c. Cost:
-------
-p.
en
(V)
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 7 of 8
Type of Mine: Ferroalloy
iline
Code -
Location
Type
6149-ID
Mine/Mil!
6) 50- ID
Mill
6151-MT
Mill
6152-OR
Mine
61S3-NV
Mi 1 i
6154-MT
Mill
6155-NV
Mill
6156-UT
Mill
1
Ore
Production
(1UOO tons/
year)
NA
NA
0.002
1
NA
NA
0.028
NA
Ma tor
Dis-
charged
(MGO)
NA
NA
Minimal
NA
NA
NA
0
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (11000)
c. Cost: it ton of ore mined
Second.
Settlipg
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
1). -
c.
a.
b. -
c.
a.
h. -
c.
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
ion
-
-
-
-
-
-
-
-
M xed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust,
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
- .
-
-
-
-
-
Vocess
Control
-
-
-
-
-
-
-
-
REMARKS
Inactive
M
-------
TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Ferroalloy
p. 8 of 8
Mine
Code -
Location
Type
6157-NV
Mill
6158-CA
Mill
6159-NN
Mill
6160-TX
Mill
6161-NM
Mill
6162-SC
Mill
6163- CA
Mill
Ore
Production
(1UOO tonv
year)
0.055
(cone. )
0.496
0.110
(cone. )
NA
8
(cone. )
NA
0.055
(cone. )
Hater
Dis-
charged
(MGO)
0
0
0
0
0
0
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMH COSTS: b. Annual Cost (*1000)
c. Cost: i/ton of ore mined
Second.
Settling
a.
b.
r .
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
t). -
c.
a.
b.
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
Aikal.
Chlor-
i nation
-
-
-
-
-
-
-
Ion
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
Pll
Adjust.
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
50%
-
«
-
-
-
-
-
75%
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
'rocess
Control
-
-
-
-
-
-
REMARKS
'reduction based
on mill capacity
-------
TABLE IX-8. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 1 of 1
Type of Mine: Mercury
tn
Mine
Code -
Location
Type
9201 -CA
Mine/Mill
9202-NV
Mine/Mill
~1
Ore
Production
(1000 ton^
year)
30
17.5
Hater
Dis-
charqed
(MGD)
0
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AW COSTS: b. Annual Cost (»1000)
c. Cost:
-------
TABLE IX-9. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Titanium
p. 1 of 1
en
01
Mine
Code -
Location
Type
9905-NY
Mine
9906-FL
Mill
9907-FL
Mill
9903-FL
Mine/Mill
9909-FL
Mine/Mill
9910-NJ
Mill
9911-NJ
Mine/Mill
1
Ore
Production
(1UOO tonv
year)
464
726U
.7260
NA
HA
6600
NA
Hater
Dis-
charged
(MOD)
0.70
6.84
1.63
NA
NA
3.77
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AUH COSTS: b. Annual Cost (UOOO)
c. Cost:
-------
TABLE IX-10. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 1 of 5
Type of Mine: Uranium, Radium,
Vanadium
LTI
01
it 1 ne
Code -
Location
Type
9401-NM
Mine
9401-NM
Mill
9402-NM
Mine 35. 31
9402-NM
Mine 17,22
24,30,30
9402-NM
Mill
9403-UT
Mill
9404- NM
Mill
9405-CO*
Mill
* In
( )** fo
Ore
Production
(1UOO tons,
year)
750
1,270
1J25
2409
274*
2,490
439*
Hd ter
Dis-
charged
(MGO)
0.85
(1.86)**
3.00
0.00
(1.94)**
(0.41)**
(1.38)**
1.00
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMf> COSTS: b. Annual Cost (»1000)
c. Cost: (/ton of ore mined
Second.
Settling
a. 90
b. 19.5
c. 2.60
a. 120
b. 22
c. 1.72
a.
b. -
c.
a.
b. -
c.
a. 120
b. 22
c. 0.91
a. 74
b. 18
c. 6.57
a. 110
b. 21
c. 0.84
a. 96
b. 20
c- 4.55
Floc-
cula-
tion
68
30
4.00
-
-
-
-
-
- .
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
Ination
-
-
-
-
-
-
-
-
Ion
Exchan.
1900
510
68.00
-
-
-
-
-
-
-
Mixed
Media
Filtr.
180
52
6.93
-
-
-
-
-
-
-
PH
Adjust.
35
28
3.73
-
-
-
42
35
1.45
30
25
9.12
39
31
1.24
37
29
6.61
Recycle
25%
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
60
30
2.3(
-
-
60
30
1.25
25
20
7.30
47
27.5
1.10
40
26
5.92
Vocess
Control
-
-
-
-
-
-
-
-
REMARKS
No point
discharge
0 Discharge
Ref. p. 111-71
Mill receives
ore from
other mines
No point
discharge
M
Single
discharging
uranium mill
dicates production obtained from mill capacity - assume 365 working days/year. +See also page 5 for off site evaporation
r 0 discharge mills, flow indicated = volume discharged to treatment or recycle system. Pot>d design.
-------
TABLE IX-10, COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 2 of 5
Type of Mine: Uranium, Radium,
Vanadium
Mine
Code -
Location
Type
9407-WY
Mill
9409-WY
Mir.e
'9409-WY
Mill
9410-WY
Mine
9411-WY
Mill
9413-MY
Mine
9413-WY
Mill
9419-TX
Mill
"1
Ore
Production
(1000 tons/
year)
724
500
1,086
0
357
595
603
1,046
Hater
Dis-
charged
(MOO)
(1.11)**
0.50
(1.22)**
2.30
(0.39)**
0.36
(2.04)**
4.65
a. Capital Cost ($10UO)
TREATMENT TECHNOLOGIES AM1) COSTS: b. Annual Cost (*1000)
t. Cost:
-------
TABLE IX-10. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
tn
00
p. 3 of 5
Type of Mine: Uranium, Radium,
Vanadium
Mine
Code -
Location
Type
9422-CO
Mill
9423-HA
Mill
9425-WY
Mill
9427-WY
Mill
9430- UT
Mill
9437-NM
Mine
9442-WY
Mill
9443-NM
Mine
Ore
Product lor
(IUOO tons;
year)
165
183
NA
346*
NA
96
563*
NA
Water
Dis-
charged
(MGD)
NA
(0.14)**
0
(0.33)**
NA
5.03
(0.28)**
1.45
* Indicates production o
( )** for 0 discharge mills,
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMH COSTS: b. Annual Cost (>1000)
c. Cost: {/ton of ore mined
Second.
Settling
a.
b. ~
c.
a.
b. -
c.
a.
b. -
c.
b! 17.5
c. 5.06
a.
b. -
c.
a. 180
b. 28
c. 29.17
a. 66
b. 17.4
c. 3.09
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
90
54
56.21
-
-
btalne3~Trom mill
flow Indicated =
Ozon-
ation
-
-
-
-
-
-
-
capacity
volume (
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
6500
1710
L781.25
-
-
Mixed
Media
Filtr.
-
-
-
-
-
700
130
-
-
PH
Adjust
-
-
-
29
24
6.94
-
58
52
54.17
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
23
19
5.49
-
-
-
-
'rocess
Control
-
-
-
-
-
-
-
REMARKS
Lined evapor.
pond : no
discharge
No point
discharge
No point
discharge
No point
discharge
Under
development
t - assume 365 working days/year.
discharged to treatment or recycle system.
-------
en
vo
TABLE IX-10. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 4 of 5
Type of Mine: Uranium, Radium
Vanadium
iline
Code -
Location
Type
9445-NM
Mill
9446-NM
Mill
9447-UT
Mine
9447-UT
Mill
9449-WY
Mine
9449-WY
Mill
9450-UY
Mill
9452-NM
Mine
Ore
Production
(1000 tonsy
year)
540
607*
262
273*
750
NA
483*
500
Water
Dis-
charged
(MGD)
0
(0.63)**
minimal
(0.31)**
9.8
0
(0.55)**
1.80
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMO COSTS: b. Annual Cost (>1000)
c. Cost: (/ton of ore mined
Second.
Settling
a.
b. -
c.
a. 80
b. 18.8
c. 3.10
a.
b. -
c.
a. 70
b. 17.5
c. 6.41
a. 240
b. 34
c. 4.53
a.
b. -
c.
a. 80
b. 18.5
c. 3.83
a.
1). -
c.
Floc-
cula-
tlon
-
-
-
-
98
80
10.67
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
ination
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
1200
190
25.33
-
-
-
PH
Adjust
-
35
27
4.45
-
-
71
78
10.4
-
34
26
5.38
-
Recycle
25X
-
-
-
-
-
-
-
-
50*
-
-
-
-
-
-
- .
-
751
-
-
-
-
-
-
-
-
loot
-
-
-
-
-
29
22
4.5
-
Yocess
Control
-
-
-
-
-
-
-
-
REMARKS
i No point
discharge
No point
discharge
No point
i discharge
indicates production obtained from mill
( )** for 0 discharge mills, flow indicated =
capacity - assume 365 working days/year.
volume discharged to treatment or recycle system.
-------
TABLE IX-10. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
p. 5 of 5
Type of Mine: Uranium, Radium,
Vanadium
iline
Code -
Location
Type
9452-NM
Mill
9456-WA
Mill
New Mine
"A" - NM
Mine
9460-WY
Mine
9460-WY
Mill
9463-TX
Mill
Union
Carbide-
Rifle Mil
9405-CO
Mill
1
Ore
Production
(1000 tons/
year)
1,448*
764*
0
300
NA
NA
NA
439
Wa I.er
Uis-
charqed
(MGD)
(1.04)**
(0.63)*"
0.87
0.87
0
NA
NA
1.00
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMU COSTS: b. Annual Cost (3,1 000)
c. Cost: i/ton of ore mined
Second.
Settl ing
a. 95
b. 20
c. 1.38
a.
b. -
c.
a.
b. -
c.
a. 90
b. 19.5
c. 6.50
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a -23, 900
b. 2,550
c. 580
Floc-
cula-
tion
-
-
-
68
30
10. OC
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
ination
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
1900
510
170.00
-
-
-
-
Mixed
Media
rntr.
-
-
-
180
52
17.33
-
-
-
-
PH
Adjust.
37
30
2.07
-
-
36
28
9.33
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
40
26
1.79
-
-
-
-
-
-
-
Vocess
Control
-
-
-
-
-
-
-
-
REMARKS
No production
at present
Evaporation
pond design
cr>
o
YribTcates""production obtained from"mi 11
)** for 0 discharge mills, flow indicated =
capacity - assume 365 working days/year.
volume discharged to treatment or recycle system.
-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l)
Waste Type by Segment
Iron Mining and Milling
(1101. 1105,1109,1120)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds - Settled
Solids
Copper Mining and Milling
(2101,2104,2118,2119,
2120,2121,2122.2126,
2139,2147.2164)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds Settled
Solids
Arsenic
Range
<0.0005-
0.161
<0.0005-
0.005
<0.0005-
0.010
<0.0005-
0.550
0.002-
0.050
< 0.002-
0.0155
0.006-
0.055
0.0026-
0.065
Barium
Range
0.025-
0.715
0.105-
0.51
0.13-
0.39
0.02-
0.41
0.058-
0.62
0.032-
0.11
0.04-
2.8
< 0.001-
2.0
Cadmium
Range
<0.008-
0.016
<0.008
<0.008-
0.021
<0.008
< 0.008-
0.17
< 0.008-
0.071
< 0.008-
0.022
< 0.008-
0.039
Chromium
Range
<0.001-
0.003
<0.001-
0.009
<0.001-
0.076
<0.001-
0.013
<0.001-
<0.04
<0.001-
0.052
<0.001-
0.057
<0.001-
0.110
Lead
Range
<0.084
<0.084
<0.084
0.084-
0.112
<0.06-
0.840
<0.08-
0.084
<0.084-
0.840
<0.06-
0.084
Mercury
Range
<0.0005-
0.001
<0.0005-
0.001
<0.0005-
0.001
<0.0005-
0.001
<0.0005-
0.002
< 0.0005-
0.001
< 0.0005-
0.001
< 0.0005-
0.002
Selenium
Range
0.001-
0.009
0.003-
0.009
0.003-
0.015
0.001-
0.021
0.0005-
0.079
0.006-
0.056
0.006
0.104
< 0.001 -
0.105
Silver
Range
<0.002-
0.008
<0.002-
<0.003
<0.002-
0.01
<0.002
<0.002-
0.024
< 0.002-
0.010
< 0.002-
0.012
< 0.002-
0.021
CTl
-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Lead/Zinc Mining and Milling
(3104,3106,3107,3110,
3113,3122,3123,3126)
Mine Waste Rock
Fresh Tailings
Tailing Ponds - Settled
Solids
Mine Water Ponds -
Settled Solids
Gold/Silver Mining and Milling
(4101,4105,4119.4121,
4402, 4407)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds - Settled
Solids
Aluminum Mining (5101)
Mine Water Ponds
Settled Solids
Mine Waste Rock
Arsenic
Range
<0.0041-
0.047
< 0.002-
< 0.023
<0.002-
0.043
< 0.0044-
0.008
<0.0005-
0.027
<0.002-
0.103
0.004-
0.017
0.007-
0.369
< 0.025
< 0.025
Barium
Range
0.052-
0.270
0.051-
0.865
0.016-
1.68
0.47-
1.5
0.095-
2.90
0.160-
3.25
0.009-
1.73
<0.001-
1.90
0.15-
1.19
0.34
Cadmium
Range
0.039-
0.650
<0.008-
0.17
<0.008-
0.36
0.013-
0.040
< 0.003-
0.098
<0.008-
0.087
<0.008-
0.170
<0.008-
0.3
< 0.005-
0.01
< 0.005
Chromium
Range
0.004-
0.180
0.003-
0.090
0.007-
0.140
0.057-
0.060
<0.001-
0.009
<0.001-
0.087
<0.001-
0.120
<0.001-
0.074
<0.05
<0.05
Lead
Range
<0.084-
23.0
<0.084-
16.0
<0.06-
8.8
<0.084-
0.100
<0.06-
11.0
< 0.060-
4.600
< 0.060-
17.0
< 0.060
100.0
0.14-
0.23
0.13
Mercury
Range
<0.0005-
0.018
<0.0005-
0.041
<0.0001-
0.102
0.023-
0.027
< 0.0005-
0.001
< 0.0005
<0.0005-
0.009
<0.0005-
0.033
< 0.0003
< 0.0003
Selenium
Range
< 0.001-
0.032
0.006-
0.050
<0.001-
0.106
0.007-
0.039
<0.001-
0.041
0.002
0.050
0.009-
0.672
0.013
0.194
< 0.025
< 0.025
Silver
Range
< 0.002-
0.014
<0.002-
0.251
<0.002-
0.180
< 0.002-
0.007
< 0.002-
0.034
0.002-
0.049
0.002-
0.130
< 0.002-
0.065
< 0.001
< 0.001
IN5
-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Molybdenum Mining and
Milling (6101, 6102, 6103!
Low Grade Ore
Mine Waste Rock
Fresh Tailings
Tailing Ponds - Settled
Solids
Wastewater Treatment
Sludge
Mine Water Pond -
Settled Solids
Tungsten Mining and Milling
(6104, 6105)
Mine Waste Rock
Tailing Pond - Settled
Solids
Dry Tailings
Mine Water Pond Settled
Solids
Arsenic
Range
<0.005-
<0.006
<0.005
<0.0005-
0.019
<0.005-
0.017
0.026
0.048
<0.001-
< 0.002
0.0218-
0.075
<0.001-
0.02
< 0.002-
<0.003
Barium
Range
0.058-
0.155
0.08-
0.19
0.14-
0.27
0.09-
0.2
0.039
0.74
0.22-
0.4
0.395-
0.59
0.2-
0.4
0.31-
0.38
Cadmium
Range
<0.008
<0.008
<0.008
<0.008
0.064
< 0.008
0.015-
0.02
0.017-
0.027
<0.01-
0.01
0.011-
0.015
Chromium
Range
0.001-
<0.002
<0.001-
0.004
<0.001-
0.01
<0.001-
0.018
0.21
0.12
< 0.001-
0.07
0.0085-
0.032
<0.04
<0.001-
0.002
Lead
Range
<0.084
<0.084
<0.084
<0.084-
0.19
<0.084
<0.084
<0.05-
< 0.084
<0.06-
< 0.084
<0.05
< 0.084
Mercury
Range
<0.0005
<0.0005
<0.0005
<0.0005-
0.0018
<0.0005
<0.0005
<0.0005
< 0.0005-
0.0005
0.0001-
0.0004
< 0.0005-
<0.0018
Selenium
Range
0.004-
0.018
0.002-
0.010
0.002-
0.043
0.003-
0.023
0.055
0.006
0.0199-
0.052
0.0448-
0.173
0.041-
0.046
0.0048-
0.0212
Silver
Range
<0.002
<0.002
<0.002-
0.004
<0.002-
0.015
0.03
0.011
< 0.002-
<0.01
< 0.002-
<0.01
<0.01
< 0.002-
0.002
O-l
CO
-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Vanadium Mining and Milling
(6107)
Mine Waste Rock
Mine Water Pond - Settled
Solids
Tailing Pond - Settled
Solids
Mill Wastewater Ponds -
Settled Solids
Nickel Mining, Milling and
Smelting (6106)
Mine Waste Rock
Low Grade Ore
Mine/Smelter Wastewater
Settled Solids
Mercury Mining and
Milling (9202)
Fresh Tailings
Tailing Ponds - Settled
Solids
Mine Waste Rock
Arsenic
Range
< 0.001
<0.001
<0.001
<0.001-
0.54
0.02
<0.001
0.001
0.17
0.26
0.1
Barium
Range
0.13
0.15
1.31-
1.69
0.03-
0.3
0.1
<0.1
0.2
0.76
0.76
0.76
Cadmium
Range
<0.01
0.04
<0.01
<0.01-
0.37
<0.01
<0.01
<0.01
<0.005
<0.005
0.01
Chromium
Range
<0.04
<0.04
<0.04
<0.04-
1.9
<0.04
<0.04
<0.04
<0.05
<0.05
0.06
Lead
Range
<0.05
<0.05
<0.05
<0.05
<0.05
<0.05
<0.05
0.14
0.11
<0.1
Mercury
Range
<0.0001
<0.0001
<0.0001-
0.0002
<0.0001-
0.0036
<0.0001
<0.0001
0.0001
0.0019
0.041
0.14
Selenium
Range
<0.001
0.008
<0.001-
0.004
0.001-
0.03
0.001
<0.001
0.001
<0.015
<0.015
<0.015
Silver
Range
<0.002
0.004
<0.002-
0.046
0.008-
0.25
<0.002
0.032
0.004
<0.001
<0.001
<0.001
Oi
-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Uranium Mining
(9402, 9403, 9404, 9405,
9408,9409,9411,9412,
9423, 9447, 9451, 9455,
9460)
Mine Waste Rock
Low Grade Ore
Mine Water Ponds -
Settled Solids
Titanium Dredge Mining
and Milling (9906)
Fresh Tailings
Mine Water Pond -
Settled Solids
Mill Wastewater Pond -
Settled Solids
Antimony Mining and
Milling (9901)
Fresh Tailings
Tailing Pond - Settled
Solids
Arsenic
Range
< 0.0005-
0.031
< 0.0005-
0.023
< 0.0005-
0.057
< 0.001
0.02
0.001
0.21
0.25
Barium
Range
0.001-
1.29
0.059-
0.83
0.26-
48.0
<0.1
<0.1
<0.1
1.85
1.0
Cadmium
Range
< 0.008-
0.040
< 0.008-
0.040
< 0.008-
0.040
<0.01
<0.01
<0.01
0.02
< 0.005
Chromium
Range
< 0.001 -
0.056
0.004-
<0.02
<0.001-
0.060
<0.04
<0.04
<0.04
0.06
<0.05
Lead
Range
<0.06-
0.100
<0.06-
< 0.084
<0.06-
0.100
<0.05
<0.05
<0.05
0.14
0.14
Mercury
Range
< 0.0005-
0.046
< 0.0005-
<0.014
< 0.0005-
0.009
0.0003
0.0004
0.0005
< 0.0003
0.0003
Selenium
Range
< 0.0005-
0.085
< 0.0005-
0.154
< 0.0005-
0.073
0.007
0.008
0.006
<0.015
<0.01&
Silver
Range
< 0.002-
0.06
0.005-
0.018
< 0.002-
0.026
0.03
0.016
0.015
< 0.001
<0.001
-------
Figure IX-1. SECONDARY SETTLING POND/LAGOON - TYPICAL LAYOUT
01
CTl
WASTEWATER
INFLUENT
PLAN
2.5:1
.01
"I
2.5:1
2.5=1
EFFLUENT
PIPE
SECTION
WASTEWATER INFLUENT
OUTLET PIPE
EFFLUENT PIPE
EXISTING
GRADE
-------
Figure IX-2. ORE MINING WASTEWATER TREATMENT SECONDARY SETTLING
POND/LAGOON COST CURVES
10000
100
WASTEWATER FLOW - MGD
(I) WASTEWATER SAMPLE ANALYSIS NOT INCLUDED
1000
10000
REVISED 2/29/80
-------
COST IN THOUSAND DOLLARS
01
Co
2
O
~n T»
Oc
*"w
iS
CO
H
CO
CO
c
m
IN)
\
ro
ID
X.
03
O
O
O
O
o
o
o
w
o
3
s
**
^
3fc
^-
o
8
(
><
CO
b
CO
o
c
33
33
m
m
m
>
m
33
H
33
m
m
Z
H
CO
m
O
T3
O
Z
D
CO
Z
o
-------
Figure IX-4. FLOCCULANT (POLYELECTROLYTE) PREPARATION AND FEED-FLOW SCHEMATIC
HOPPER
FEEDER
WETTING CHAMBER
MIXER
WATER 4r
SUPPLY
MIXING TANK
METERING
PUMP
WASTE WATER
INFLUENT
MIXER
EFFLUENT,
AGING TANK
MIXING TANK.
-------
Figure IX-5. ORE MINE WASTEWATER TREATMENT FLOCCULANT (POLYELECTROLYTE)
PREPARATION & FEED SYSTEM COST CURVES
IQOr
REVISED 2/29/80
1.0 10.0
WASTEWATER FLOW - M G D
(I) WASTEWATER SAMPLE ANALYSIS NOT INCLUDED
100,0
-------
Figure IX-6. OZONE GENERATION AND FEED FLOW SCHEMATIC
WASTEWATER
INFLUENT
AIR
coo
COMPRESSOR -7
j . /
->
*n >r
>y »i
FILTER
SILENCER CQO
WA"
SUF
JNG WATER ELECTRIC COOLING
ro WASTE POWER SUPPLYn WATER
T /-COMPRESSOR T
| / AFTER COOLER | |
~Vj 1 * f J OZONE "\_^
1 % J 1 f ^GENERATORj *
REFRIGERANT | |
COOLING
ING SYSTEM
PER
'PLY
VT
4
COOLING /
WATER /
OZONE /
CONTACTOR^ T|
E
i
mmt
1
^E/
FFI
r
mtm
t
VTE
_UE
:D
:NT
DESSICANT
DRYING SYSTEM
-------
Figure IX-7. ORE MINE WASTEWATER TREATMENT OZONE GENERATION & FEED SYSTEM COST CURVES
REVISED 2/29/80
6/4/80
o
o
DOLLARS
_
o
COST
10
DESIGN FLOW IN M.G.D.
NOTES! I) OZONE DOSAGE ASSUMED AT 5mfl/l.
2) O WASTEWATER SAMPLE ANALYSIS NOT INCLUDED.
100
1000
-------
Figure IX-8. ALKALINE-CHLORINATION FLOW SCHEMATIC
00
CAUSTIC
SODA
STORAGE
TANK
RAW
WASTES
PUMP
PUMP
SODIUM
HYPO-
CHLORITE
STORAGE
TANK
MIXER
^TREATED
EFFLUENT
MIXING TANK
-------
CAPITAL COST IN THOUSAND OF DOLLARS
5 8
C/)
m
o
ro
00
O
m
I
m
I
"U
rn
CO
to
z
o
r~
o
m
o
m
-n
i
I ^
o
o
(O
CD
X
CD
o
cz
33 m
m
m
33
m
H
2
m
Z
H
>
r~
>
rn
O
I
p =
o
z
-------
Figure IX-10. ION EXCHANGE FLOW SCHEMATIC
ACID
STORAGE TANK
CAUSTIC
STORAGE TANK
-pi
*-J
en
RAW
WASTED v v
T T
m-
i
ir 1 ir
* 1
* <"*
7I_
IT
1 i
FILTERS i !
' CATION i
EXCHANGERS
I
. ^
w
i
-*
DEGASIFIE
i *
._A
r
i > '
(^
Tl
R !,
EXC
x
^-
i i
.__!>.*-
r
i
i
i i
i
1
r
T
1
j
i
MMION I
HANGERS
^
h
i
^T
i
1
U
E
^
p
r
i
i A
i
TT
IIXED BED
XCHANGERS
i i
TREATED
EFFLUENT
WASTE
STORAGE
TANK
\
s
T
A/AST E
TORAG
TANK
i
i
i
i
E !
1
S
WASTE
TORAGE
TANK
TREATED
" WASTES
NEUTRALIZATION
TANK
-------
Figure IX-11. ORE MINE WASTEWATER TREATMENT ION EXCHANGE COST CURVES
COST IN MILLIONS OF DOLLARS _
o c
- o 'o c
CAPITAL
j*
_^r~
X^
^
^
-^^
x^
x
x
x
>
x
'
c
*
s
0
^
r
S"
^
i
g^
r v
X
^
^£
f^^*L -t
\
>
X
^r
j
r"
X
j(
'
X
.
r1
/
r
/
/
/
/
»
/
j
f
f/
(1)
ir ANNUAL COST
/
/
/
f
}
/
/
/
r
1.0 10.0
WASTEWATER FLOW - MGD
100.0
(I) WASTEWATER SAMPLE ANALYSIS NOT INCLUDED
REVISED 2/29/80
476
-------
Figure IX-12. GRANULAR MEDIA FILTRATION PROCESS FLOW SCHEMATIC
WASTEWATER
INFLUENT
SETTLED
WATER
RECYCLE
PUMP
GRAVITY FILTER (GRANULAR MEDIA)
TREATED
1
EFFLUENT
CLEAR
WELL
BACKWASH
WASTEWATER
BASIN
BACKWASH
PUMP
SOLIDS TO
DECANT
-------
Figure IX-13. ORE MINE WASTEWATER TREATMENT GRANULAR MEDIA
FILTRATION PROCESS COST CURVES
COST IN MILLIONS OF DOLLARS
D - b
CAPIT/
COST
j+
s
/
^^
J
/
r
*^
\
-
s
^
*
f
*
>
,/
f
jj
_^^^
^
/
/
^s^
^^
/
^
-rf
^_
'K,
/
s
\
s
'
\
s
s
s
/
s
^
*
*
»
k
/
^
/
r
/
r
s
s
^
s
/
^ ANNUAL COST
s
/
(I
!
^
^
f
*
w
I.O 10
WASTEWATER FLOW - MGD
IOO.O
(I) WASTEWATER SAMPLE
ANALYSIS NOT INCLUDED
REVISED 2/29/80
478
-------
Figure IX-14. pH ADJUSTMENT FLOW SCHEMATIC
HYDRATED
LIME
FEEDER t/WWx
WATER
SUPPLY"
FEED
PUMP
D
A
J
MIXER
LIME SLURRY
TANKS
WASTEWATER
INFLUENT
NEUTRALIZED
WASTES
MIXING TANK
-------
Figure IX-15. ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT CAPITAL COST CURVES
-pi
co
o
IOOO
£E
<
§
u.
o
IOO
o
I
z to
h-
en
O
o
O.OI
O.I
I 10
WASTEWATER PLOW - M G D
IOO
IOOO
REVISED 2/29/80
-------
Figure IX-16. ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT ANNUAL(1) COST CURVES
CO
10.000
_
§ IOOO
u.
o
o
z
13
o
100
o
o
10
0.01
O.I I 10
WASTEWATER FLOW - M G D
(I) WASTEWATER SAMPLE ANALYSIS NOT INCLUDED.
100
IOOO
REVISED 2/29/80
-------
Figure IX-17. WASTEWATER RECYCLE FLOW SCHEMATIC
RECYCLE FLOW TO MINING
AND MILLING OPERATION
X
00
ro
D
WASTEWATER
INFLUENT
D
PUMP STATION
WET WELL
NOTE:
NUMBER OF PUMPS DEPENDS ON THE
RECYCLE FLOW
EFFLUENT TO
>RECEIVING
STREAM
-------
Figure IX-18. ORE MINE WASTEWATER TREATMENT RECYCLING CAPITAL COST
CURVES
I.O
CO
(T
-J
o
o
u_
o
tr>
z
o
.0!
to
O
IOO %
75%
\
25%
50%
^
I.O IO.O
WASTEWATER FLOW - MGD
IOO.O
483
-------
Figure IX-19. ORE MINE WASTEWATER TREATMENT RECYCLING ANNUAL COST CURVES
oo
.I
I.O IO.O
WASTEWATER FLOW - MOD
IOO.O
-------
Figure IX-20. ACTIVATED CARBON ADSORPTION FLOW SCHEMATIC
WASH WATER DRAIN |
oo
WASTEWATER
INFLUENT
PUMP
STATION
0
EXISTING
BPT POND
WASH
WATER
PUMPS
D
CLEAR
WATER
WELL
J
ACTIVATED CARBON FILTERS
-------
co
Figure IX-21. ACTIVATED CARBON ADSORPTION CAPITAL COST CURVE FOR PHENOL REDUCTION IN
BPT EFFLUENT DISCHARGED FROM BASE AND PRECIOUS METAL ORE MILLS
IOOOO
cr
_
o
O
U.
O
(f)
a
z
O
I
o
o
1000
IOO
10
0.01
O.I 1.0
WASTEWATER FLOW - MGD
10.0
100.0
-------
Figure IX-22. ACTIVATED CARBON ADSORPTION ANNUAL COST CURVE FOR PHENOL REDUCTION
IN BPT EFFLUENT DISCHARGED FROM BASE AND PRECIOUS METAL ORE MILLS
CO
en
cr
_j
_i
O inno
i*^ 1 w^w
i ,
LL.
Q
Z
§
O
X
inn
4»»
~"
h-
CO
O
in
0
mmm~*~
.01
"
-
f
«
*
*
*
c
_4*
^^
.1
_->
^^
^^
J
jr
s
'
/
^-
-
1
0
k7
^^
l^1
A
'
/
/
/
4
A
l(
X
D.O
/
/
f
s
f
10
WASTEWATER FLOW-MOD
-------
Figure IX-23. CHEMICAL OXIDATION - HYDROGEN PEROXIDE FLOW SCHEMATIC
AIR
COMPRESSOR
INFLUENTi
-£>
00
CO
OXIDATION BASINS
r- pH ADJUSTMENT
TO
DISCHARGE
SLUDGE TO
DISPOSAL
CLARIFIERS
n
PUMP
SULFURIC
ACID
FERROUS SULFATE
DISSOLVER
FROM
TANK
TRUCK
HYDRO6EN PEROXIDE
STORAGE TANK
-------
Figure IX-24. HYDROGEN PEROXIDE TREATMENT CAPITAL COST CURVE FOR PHENOL REDUCTION
IN BPT EFFLUENT DISCHARGED FROM BASE & PRECIOUS METAL ORE MILLS
10000
o>
0.01
1.0 10.0
WASTEWATER FLOW - M G D
100.0
-------
-P.
U3
O
Figure IX-25. HYDROGEN PEROXIDE TREATMENT ANNUAL COST CURVE FOR PHENOL REDUCTION
IN BPT EFFLUENT DISCHARGED FROM BASE & PRECIOUS METAL ORE MILLS
10000
1000
CO
cc
o
0
U_
O
CO
13
O
£ 100
CO
O
o
0.01
1.0 10.0
WASTEWATER FLOW - M G D
100.0
-------
Figure IX-26. CHEMICAL OXIDATION-CHLORINE DIOXIDE FLOW SCHEMATIC
FROM
TANK
TRUCK
n.
STORAGfc
TANK
METERING
PUMP
DILUTION
IN-LINE
FILTER
fl-n
EJECTOR
WASTE WATER
INFLUENT'
T
/
^X
-TO DISCHAR6E
CONTACT TANK
-------
Figure IX-27. CHLORINE DIOXIDE CAPITAL COST CURVE FOR PHENOL REDUCTION IN BPT EFFLUENT
DISCHARGED FROM BASE AND PRECIOUS METAL ORE MILLS
10000
o
0
LL
O
(O
ZD
O
o
(J
1000
100
10
0.01
1.0
10.0
100.0
WASTEWATER FLOW - MGD
-------
vo
CO
Figure IX-28. CHLORINE DIOXIDE ANNUAL COST CURVE FOR PHENOL REDUCTION IN BPT EFFLUENT
DISCHARGED FROM BASE AND PRECIOUS METAL ORE MILLS
10000
o
o
u_
o
Q
z
ID
O
I
h-
co
o
o
1000
100
10
0.01
O.I 1.0
WASTEWATER FLOW - M G D
10.0
100.0
-------
Figure IX-29. CHEMICAL OXIDATION-POTASSIUM PERMANGANATE FLOW SCHEMATIC
FROM
TANK
TRUCK
H20
INFLUENT-
STORAGE
TANK
D
D
OXIDATION BASINS
TO
DISCHARGE
SLUDGE
TO
DISPOSAL
CLARIFIERS
DISSOLVER
-------
en
Figure IX- 0. POTASSIUM PREMANGANATE TREATMENT CAPITAL COST CURVE FOR PHENOL
REDUCTION IN BPT EFFLUENT DISCHARGED FROM BASE AND PRECIOUS METAL
ORE MILLS
10000
V)
(T
O
Q
U_
O
C/)
Q
Z
<
0)
1>
O
I
I-
cn
O
o
1000
100
10
-t4-
0.01
O.I 1.0 10.0
WASTEWATER FLOW - MGD
100.0
-------
-Pi
IO
CT)
CO
o
o
Figure IX-31. POTASSIUM PERMANGANATE TREATMENT ANNUAL COST CURVE FOR PHENOL
REDUCTION FROM BASE AND PR ECIOUS METAL ORE MILLS
10000
CO
o:
o
Q
LL
O
CO
Q
CO
ID
O
X
1000
100 -
10
0.01
O.I 1.0 10.0
WASTEWATER FLOW - MGD
100.0
-------
SECTION X
BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE
The effluent limitations which must be achieved by 1 July 1984,
are based on the best control and treatment technology employed
by a specific point source within the industrial category or
subcategory, or by another industry where it is readily
transferable. Emphasis is placed on additional treatment
techniques applied at the end of the treatment systems currently
employed for BPT, as well as improvements in reagent control,
process control, and treatment technology optimization.
Input to BAT selection includes all materials discussed and
referenced in this document. As discussed in Section VI, nine
sampling and analysis programs were conducted to evaluate the
presence/absence of the pollutants (toxic, conventional, and
nonconventional). A series of pilot-scale treatability studies
was performed at several locations within the industry to
evaluate BAT alternatives. Where industry data were available
for BAT level treatment alternatives, they were also evaluated.
Consideration was also given to:
1. Age and size of facilities and wastewater treatment
equipment involved
2. Process(es) employed and the nature of the ores
3. Engineering aspects of the application of various types
of control and treatment techniques
4. In-process control and process changes
5. Cost of achieving the effluent reduction by application
of the alternative control or treatment technologies
6. Non-water quality environmental impacts (including
energy requirements)
This level of technology also considers those plant processes and
control and treatment technologies which at pilot-plant and other
levels !" ve demonstrated both technological performance and
economic 'lability at a level sufficient to justify investiga-
tion,
The Clean Water Act requires consideration of costs in BAT
selection, but does not require a balancing of costs against
effluent reduction benefits (see Weyerhaeuser v. Costle, 11 ERC
2149 (DC Cir. 1978)). In developing the proposed BAT, however,
EPA has given substantial weight to the reasonableness of costs
and reduction of discharged pollutants. The Agency has
considered the volume and nature of discharges before and after
497
-------
application of BAT alternatives, the general environmental
effects of the pollutants, and the costs and economic impacts of
the required pollution control levels. The regulations proposed
are, in fact, based on the application of what the Agency deems
to be Best Available Control Technology Economically Achievable,
with primary emphasis on significant effluent reduction
capability.
The options considered are limited only by their ability to meet
BPT Effluent Guidelines (as a minimum), technical feasibility in
the particular subcategory, and obviously extreme (high) cost.
The options presented represent a range of costs so as to assure
that affordable alternatives remain after the economic analysis.
SUMMARY OF BEST AVAILABLE TECHNOLOGY
Zero discharge limitations are established for the following
subcategories and subparts:
Iron Ore
Mills in the Mesabi Range
Mercury Ore
Mills
Cu, Pb, Zn, Au, Ag, Pt, and Mo Ores
Mills using the cyanidation process or the amalgamation
process to recovery gold or silver
Mills and mine areas that use leaching processes to
recover copper
Subcategories and subparts permitted to discharge subject to
limitations are:
Nonconventional
Pollutants Toxics
Subcategory and Subpart Controlled Controlled
Iron Ore
Mine Drainage Fe (dissolved)
Mills (physical methods) Fe (dissolved)
498
-------
Subcateqory and Subpart
Nonconventionjal
Pollutants
Controlled
Toxics
Controlled
Subcategory and Subpart
Aluminum Ore
Mine Drainage
Uranium, Radium, and
Vanadium Ores
Mine Drainage
Mercury Ore
Mine Drainage
Titanium Ore
Mine Drainage
Mills
Dredges
Tungsten Ore
Mine Drainage
Mills
Cu, Pb, Zn, Au, Ag,
Pt, and Mo Ores
Mine Drainage (not
placer mining)
Mills (froth
flotation)
Controlled
Controlled
Fe, Al
COD, Ra226 (dis-
solved) Ra226
(total), U
Zn
Fe
Fe
Hg
Zn
Cd, Cu, Zn
Cd, Cu, Zn
Cd, Cu, Zn,
Pb, Hg,
Cd, Cu, Zn,
Pb, Hg,
499
-------
The specific effluent limitations guidelines for the subcate-
gories and subparts permitted to discharge are:
Toxic Pollutants
Copper
Zinc
Lead
Mercury
Cadmium
Daily
Maximum
mg/1
30-Day
Average
mg/1
0.30
1.0(1
0.6
0.002
0. 10
5)*
0
0
0
0.001
0.05
15
5 (0.75)*
3
Nonconventional Pollutants
Iron (dissolved)
Iron (total)
Aluminum
COD
Radium 226 (dissolved)
Radium 226 (total)
Uranium
Daily
Maximum
mg/1
30-Day
Average
mg/1
2.0
2.0
2.0 .
10 (pCi/1)
30 (pCi/1)
4
1 .0
1 .0
1 .0
500
3 (pCi/1)
10 (pCi/1)
2
*Limitations applicable to mine drainage for the copper,
lead, zinc, gold, silver, platinum, and molybdenum ores
subcategory.
GENERAL PROVISIONS
Several items of discussion apply to options in more than one
subcategory. To avoid repetition, these items are discussed here
and referred to in the discussion of the options.
Relief From No Discharge Requirement
Facilities which are not allowed to discharge under "normal"
conditions may do so as a result of:
1. An overflow or increase in volume from a precipitation
event if the facility is designed, constructed and
maintained to contain a 10-year, 24-hour rainfall design
(storm provision); or
2. If they are located in a "net precipitation" area.
These exemptions are discussed in greater detail below.
500
-------
Storm Provision
The Agency recognizes that relief is necessary as a practical
matter for many discharges within the ore mining and dressing
point source category during and immediately after some
precipitation events. It would be unreasonable to require
facilities to construct retention structures and treatment
facilities to handle runoff resulting from extreme rainfall con-
ditions which could statistically occur only rarely. Further, it
must be emphasized that the regulations for the ore mining and
dressing point source category do not require any specific
treatment technique, construction activity, or other process for
the reduction of pollution. The effluent limitations guidelines
limit the concentration of pollutants which may be discharged,
while allowing for an excursion from the normal requirements when
precipitation causes an overflow or increase in the volume of a
discharge from a facility properly designed, constructed, and
maintained to contain or treat a 10-year, 24-hour rainfall.
This relief applies to the excess volume caused by precipitation
or snow melt, and the resulting increase in flow or shock flow to
the settling facility or treatment facility. While there has
been criticism of the relief adopted by the Agency, the few
alternatives suggested by environmental groups and industry are
substantially less satisfactory in light of the data available to
the Agency. This is discussed in detail in the preamble to the
BPT regulation (43 FR 29771) and in the clarification of
regulations (44 FR 7953).
The general relief in the BPT regulation states that:
"Any excess water, resulting from rainfall or snow melt,
discharged from facilities designed, constructed and
maintained to contain or treat the volume of water which
would result from a 10-year, 24-hour pecipitation event
shall not be subject to the limitations set forth in 40 CFR
440." 43 FR at 29777-78, 440.81(c)(1978 ) .
The term "ten-year, 24-hour precipitation event" is defined,
in turn, as:
"the maximum 24-hour precipitation event with a probable re-
occurrence interval of once in 10 years as defined by the
National Weather Service and Technical Paper No. 40,
'Rainfall Frequency Atlas of the U.S.,' May 1961, and
subsequent amendments, or equivalent regional or rainfall
probability information developed therefrom." 43 FR at
29778, 440.82(d).
Under BAT, similar relief is granted: for existing sources that
are designed, constructed, and maintained to contain or treat the
maximum volume of process wastewater discharged in a 24-hour
period, including the volume which would result from a 10-year,
501
-------
24-hour precipitation event, or snow melt of equal volume, any
excess wastewater discharged shall not be subject to the
limitations set forth in 40 CFR 440.
In determining the maximum volume of wastewater which would
result from a 10-year, 24-hour precipitation event at any
facility, the volume must include the volume that would result
from runoff from all areas contributing runoff to the individual
treatment facility, 'i.e., all runoff that is not diverted from
the active mining area, runoff which is not diverted from the
mill area, and other runoff that is allowed to commingle with the
influent to the treatment system.
In general, the following will apply in granting relief:
1. The exemption as stated in the rule is available only if
it is included in the operator's permit. Many existing
permits have exemptions or relief clauses stating
requirements other than those set forth in the rule. Such
relief clauses remain binding unless and until an operator
requests a modification of his permit to include the
exemption as stated in the rule.
2. The storm provision is an affirmative defense to an
enforcement action. Therefore, there is no need for the
permitting authority to evaluate each tailings pond or
treatment facility now under permit.
3. Relief can be granted to deep mine, surface mine, and
ore mill discharges.
4. Relief is granted as an exemption to the requirements
for normal operating conditions (i.e., without overflow,
increase in volume of discharge, or discharge from a by-pass
system caused by precipitation) with respect to an increase
in flow caused by surface runoff only.
5. Relief can be granted for discharges during and immedi-
ately after any precipitation or snow melt. The intensity
of the event is not specified.
6. The relief does not grant, nor is it intended to imply
the option of ceasing or reducing efforts to contain or
treat the runoff resulting from a precipitation event or
snow melt. For example, an operator does not have the
option of turning off the lime feed to a facility at the
start of or during a precipitation event, regardless of the
design and construction of the wastewater facility. The
operator must continue to operate his facility to the best
of his ability.
7. Under the regulation, relief is granted from all efflu-
ent limitations contained in BAT.
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8. In general, relief can not be granted to treatment
facilities which employ clarifiers, thickeners, or other
mechanically aided settling devices. The use of
mechanically aided settling usually is restricted to
discharges which are not affected by runoff.
9. In general, the relief was intended for discharges from
tailings ponds, settling ponds, holding basins, lagoons,
etc. that are associated with and part of treatment
facilities. The relief will most often be based on the
construction and maintenance of these settling facilities to
"contain" a volume of water.
10. The term "treat" for facilities allowed to discharge
means the wastewater facility was designed, constructed, and
maintained to meet the daily maximum effluent limitations
for the maximum flow that would result from a 10-year, 24-
hour rainfall. The operator has the option to "treat" the
volume of water that would result from a 10-year, 24-hour
rainfall in order to qualify for the rainfall exemption, or
as mentioned in paragraph 9 above, the second option is to
"contain" the volume of wastewater.
11. The term "maintain" is intended to be synonymous with
"operate." The facility must be operated at the time of the
precipitation event to contain or treat the specified volume
of wastewater. Specifically, in making a determination of
the ability of a facility to contain a volume of wastewater,
sediment and sludge must not be permitted to accumulate to
such an extent that the facility cannot in fact hold the
volume of wastewater resulting from a 10-year, 24-hour
rainfall. That is, sediment and sludge must be removed as
required to maintain the specific volume of wastewater
required by the rule, or the embankment must be built up or
graded to maintain a specific volume of wastewater required
by the rule.
12. "Contain" and "maintain" for facilities which are
allowed to discharge do not mean providing for draw down of
the pool level of the facility allowed to discharge. As an
example, the volume can be determined from the top of the
stage of the highest dewatering device to the bottom of the
pond at the time of the precipitation event. There is no
requirement that relief be based on the facility which is
allowed to discharge being emptied of wastewater prior to
the rainfall or snow melt upon which the relief is granted.
The term "contain" for facilities which are allowed to
discharge means the wastewater facility's tailings pond or
settling pond was designed to include the volume of water
that would result from a 10-year, 24-hour rainfall.
13. The term "contain" for facilities which are not allowed
to discharge means the wastewater facility was designed,
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constructed, and maintained to hold, without a point source
discharge, the volume of water that would result from a 10-
year, 24-hour rainfall, in addition to the normal amount of
water which would be in the wastewater facility.
Net Precipitation Areas
The general relief or exemption for the requirement of "no
discharge of process wastewater" as promulgated for ore mining
and dressing in BPT Guidelines (43 FR 29771) states that:
"In the event that the annual precipitation falling on the
treatment facility and the drainage area contributing
surface runoff to the treatment facility exceeds the annual
evaporation, a volume of water equivalent to the difference
between annual precipitation falling on the treatment
facility and the drainage area contributing surface runoff
to the treatment facility and annual evaporation may be
discharged subject to the limitations set forth in paragraph
(a) of the section." Paragraph (a) refers to limitations
established for mine drainage.
Relief for net precipitation areas is included in the BAT
regulation.
Commingling Provision
The general provision as promulgated for ore mining and dressing
in the BPT Guidelines (43 FR 29771) states that:
"In the event that waste streams from various subparts or
segments of subparts in part 440 are combined for treatment
and discharge, the quantity or quality of each pollutant or
pollutant property in the combined discharge that is subject
to effluent limitations shall not exceed the quantity or
quality of each pollutant or pollutant property that would
have been discharged had each waste stream been treated
separately. The discharge flow from a combined discharge
shall not exceed the volume that would have been discharged
had each waste stream been treated separately."
This consideration is also appropriate for BAT and the regulation
states:
For existing sources which as of the date of this proposal
have combined for treatment waste streams from various
subparts or segments of subparts in Part 440, the quantity
and quality of each pollutant or pollutant property in the
combined discharge that is subject to effluent limitations
shall not exceed the quantity and quality of each pollutant
or pollutant property that would have been discharged had
each waste stream been treated separately. The discharge
flow from a combined discharge shall not exceed the volume
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that would have been discharged had each waste stream been
treated separately.
BAT OPTIONS CONSIDERED FOR TOXICS REDUCTION
As discussed in Section VII, many toxic pollutants found in this
category are related to TSS (that is, as TSS concentrations are
reduced during treatment, observed concentrations of certain
toxic metals are also reduced). In order to remove these toxics,
suspended solid removal technologies can be used. The technolo-
gies are secondary settling, coagulation and flocculation, and
granular media filtration. They are applicable throughout the
category for suspended solids reduction and associated toxic
metals reduction, and are discussed ' here to avoid repetition
during description of options. Dissolved metals are not
controlled further by physical treatment methods or additional
suspended solids removal.
Secondary Settling
This option involves the addition of a second settling pond in
series with the existing pond (as described in Section VIII).
The technique is used in many ore subcategories. The most
prevalent configuration is a second pond located in series with a
tailings pond.
Examples of the use of secondary and tertiary settling ponds can
be seen at lead/zinc Mills 3101, 3102, 3103 and at Mill 4102
(Pb/Zn/Au/Ag). This last facility uses a secondary pond to
achieve an effluent level of 4 mg/1 TSS, as determined during
sampling (Reference 1). Secondary settling ponds (sometimes
called polishing ponds) are also used in settling solids produced
in the coprecipitation of radium with barium salts at uranium
mines and mills. (See Section VIII, End-of-Pipe Techniques,
Secondary Settling; and Historical Data Summary, lead/zinc Mills
3101, 3102, 3103, and 4102.)
Coagulation and Flocculation
Chemically aided coagulation followed by flocculation and
settling is described in Section VIII. It is used by facilities
in several subcategories of the industry for solids and metals
reduction.
At Mine/Mill 1108, the tailing pond effluent is treated with alum
followed by polymer addition and secondary settling to reduce TSS
from 200 mg/1 to an average of 6 mg/1. At Mine 3121, polymer
addition has greatly improved the treatment system capabilities.
A TSS mean concentration of 39 mg/1 (range 15 to 80) has been
reduced to a mean of 14 mg/1 (range 4 to 34), a reduction of 64
percent. Similarly, polymer use at Mine 3130 reduced treated
effluent total suspended solids concentrations from a mean of 19
mg/1 (range 4 to 67 mg/1) to a mean of 2 mg/1 (range of 1 to 6.2
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mg/1). It should be pointed out that these effluent levels are
attained by the combination of settling aids and a secondary
settling pond (Section VIII).
Granular Media Filtration
This option uses granular media such as sand and anthracite to
filter out suspended solids, including the associated metals (as
discussed in Section VIII). This technology is used at one
facility (Mine 6102) and has been pilot tested at other mines and
mills and is used in other industry categories.
Granular-media filtration can consistently remove 75 to 93
percent of the suspended solids from lime-treated secondary
sanitary effluents containing from 2 to 139 mg/1 of suspended
solids (Reference 2). In 1978, lead/zinc
Mine/Mill/Smelter/Refinery 3107 was operating a pilot-scale
filtration unit to evaluate its effectiveness in removing
suspended solids and nonsettleable colloidal metal-hydroxide
floes from its wastewater treatment plant. Granulated slag was
used as the medium in some of the tests. Preliminary data
indicate that the single medium pressure filter was capable of
removing 50 to 95 percent of the suspended solids and 14 to 82
percent of the metals (copper, lead and zinc) contained in the
waste stream. Final suspended solids concentrations which have
been obtained are within the range of 1 to 15 mg/1. Optimum
filter performance was attained at lower hydraulic loadings (2.7
A full-scale multi-media filtration unit is currently in
operation at molybdenum Mine/Mill 6102. The filtration system is
used as treatment following settling (tailing pond), ion exchange
(for molybdenum removal), lime precipitation, electrocoagulation,
and alkaline chlorination. Since startup in 1978, the filtration
unit has been operating at a flow of 63 liters/second (1000 gpm)
and monitoring data show TSS reductions to an average of less
than 5 mg/1. Zinc removal and iron reduction have also been
achieved (see Treatment Technology - Section VIII).
A pilot-scale study of mine drainage treatment in Canada has also
demonstrated the effectiveness of filtration (Section VIII).
Polishing of clarifier overflow by sand filtration resulted in
reduction of the concentration of lead and zinc (approximately 50
percent) and removal of iron (approximately 40 percent) and
copper.
In addition to the above, a full-scale application of slow sand
filters is employed at iron ore Mine/Mill 1131 to further polish
tailing pond effluent prior to final discharge.
Besides the application at the various facilities described
above, a series of pilot-scale tests was performed at a number of
facilities in the ore mining category as part of the investi-
gation of BAT technologies described. These studies were con-
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ducted at Mine/Mill 3121, Mine/Mill/Smelter/Refinery 3107, Mill
2122 (two studies on tailing pond effluent), Smelter/Refinery
2122 (wastewater treatment plant), Mine/Mi1I/Smelter/Refinery
2121, Mine 3113, Mine 5102, Mill 9401, Mill 9402 (two studies)
(Reference Section VIII). In each case, filtration (among other
technologies) was evaluated and produced average effluent levels
of TSS consistently below 10 mg/1, and usually below 5 mg/1 on an
average basis.
Partial Recycle
This option consists of the recycle and reuse of mill process
water (not once-through mine water used as mill process water).
One of the principal advantages of recycle of process water is
the volume of wastewater to be treated and discharged is reduced.
Although initial capital costs of installation of pumps, piping,
and other equipment may be high, these are often offset by a
reduction in costs associated with the treatment and discharge.
Many facilities within this industry practice partial recycle
including lead/zinc Mills 3105 (67 percent), 3103 (40 percent),
3101 (all needs met by recycle), gold Mill 4105 (recycle of
treated water), molybdenum Mill 6102 (meets needs of mill),
nickel Mill and Smelter 6106, vanadium Mill 6107, and titanium
Mill 9905. In-process recycle of concentrate thickener overflow
and/or filtrate produced by concentrate filtering is practiced by
a number of flotation mills including 2121, 3101, 3102, 3108,
3115, 3116, 3119, 3123, and 3140. In addition, Mills 2120, 1132,
6101, and 6157 employ thickeners to reclaim water from tailings
or settling ponds prior to the final discharge of these tailings
to tailings ponds.
The practices described above are beneficial with respect to
water conservation and recovery of metals which might be lost in
the wastewater discharge. These practices are also significant
with respect to wastewater treatment considerations. The in-pro-
cess recycle of concentrate-thickener overflow and/or filtrate
produced by concentrate filtering reduces the volume of waste-
water discharged by 5 to 17 percent at mills which employ these
practices. Likewise, those mills which reuse water reclaimed
from tailings reduce both new water requirements and the volume
discharged by 10 to 50 percent. The advantage of any practice
which reduces the volume of wastewater discharged can be viewed
in terms of economy of treatment and enhancement of treatment
system capabilities (i.e., increased retention time of existing
sedimentation basins).
The use of mine drainage as makeup in the mill is a practice that
also deserves mention here as a method of reducing discharge
volume to the environment. A large number of facilities in the
ore mining and dressing point source category employ this
practice (see Section VIII).
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In general, there are four benefits resulting from adoption of
this practice. They are:
1. Recovery of raw materials in processing;
2. Conservation of water;
3. Reduction of discharge to tailings ponds; and
4. Increase in performance of tailings ponds.
Implementing recycle within a facility or treatment process may
require modification. Modification will be specific to each
facility and each operator will have to make his own determina-
tions.
100 Percent Recycle - Zero Discharge
This option consists of complete recycle and reuse of process
water with no resulting discharge of wastewater to the environ-
ment. Many facilities in the industry have demonstrated that
total recycle of process water is technically feasible. All of
the iron ore mills in the Mesabi Range have demonstrated the
viability of this option. Total recycle systems are also
demonstrated by iron ore Mill 1105, rare earth Mill 9903 and
mercury Mill 9201. Forty-six mills using froth flotation in the
Copper, Lead, Zinc, Gold, Silver, Platinum and Molybdenum Ores
Subcategory presently achieve zero discharge including 31 copper,
five lead/zinc, five primary gold, one molybdenum and four
primary silver mills.
There are two methods of water reclamation that are practiced in
a number of mills. They are in-process recycle and end-of-pipe
recycle. In-process recycle may involve recycle of overflow from
concentrate thickeners, recycle of filtrate from concentration
filters, recycle of spilled reagents or any combination of these.
End-of-pipe recycle involves recycle of overflow from a tailings
thickener or recycle from the tailings pond itself.
For facilities practicing mining and milling, it can be argued
that in many cases the combined treatment of mine and mill waste-
water is beneficial from a discharge standpoint, however, the
feasibility of combining the mine and mill streams will depend on
the magnitudes of:
1, The flow of mine drainage; and
2. The process water makeup flow required for the mill.
In order to achieve zero discharge at many mills, it will be
necessary to treat mine water separately and use part of it as
makeup water in the mill.
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In comments to EPA concerning existing facilities, some
facilities claimed that zero discharge by recycle would
contribute to the buildup of reagents in the process water. It
was further alleged that this buildup would interfere with the
metallurgy of the process. To date, no data have been submitted
demonstrating this contention. The cost of preventing runoff
from entering tailings or sedimentation ponds and geographical
locations in areas which have net precipitation are also concerns
with this option for existing sources.
SELECTION AND DECISION CRITERIA
Summary of_ Pollutants to be Regulated
In Section VII, Selection of Pollutant Parameters, the effluent
data obtained during sampling and analysis for each of the 129
toxic pollutants were reviewed by subcategory and subpart for
further consideration in regulation development. In summary, all
114 of the toxic organic, and six of the toxic metal pollutants
were excluded from further consideration under provisions in
Paragraph 8 of the Settlement Agreement as shown below.
Toxic Pollutants
Organics
129
- 87 (Not Detected)
- 17
Organics - 17 (Detected at levels below EPA's
nominal detection limit)
Organics - 9 (Detected at levels too low to be
effectively treated)
Organic - 1 (Uniquely related to the facility
in which it was detected)
Metals - 6 (Detected at levels too low to be
effectively treated)
9 (Remaining for consideration)
7 Toxic Metals (arsenic, copper, lead, zinc,
cadmium, mercury and nickel) Asbestos and
Cyanide
The seven toxic metals, asbestos and cyanide were considered for
regulation in subcategories and subparts where these pollutants
were detected during sampling and analysis and were not excluded
under Paragraph 8 as discussed in Section VII. Chemical oxygen
demand, total iron, dissolved iron, total radium 226, dissolved
radium 226, aluminum, and uranium, are regulated in the
subcategories and subparts in which they were regulated under
BPT.
Subcategories
Detected Are
Agreement
and Subparts in Which Toxic Pollutants Were Not
Excluded Under Paragraph 8 of the Settlement
There were subcategories and subparts in which all of the toxic
pollutants were excluded from further consideration in regulation
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development (refer to Table VII-2, Pollutants Considered for
Regulation). These include:
1. Iron ore mine drainage and mill process wastewater (not
in the Mesabi Range);
2. Aluminum mine drainage;
3. Titanium mine drainage (lode ores); and
4. Titanium mines/mills employing dredging of sand
deposits.
Consequently, for these subcategories and subparts, BAT effluent
limitations are the same as BPT effluent limitations since there
are no toxic pollutants to be controlled.
Subcategories and Subparts Which Were Not Permitted to Discharge
Under BPT
No discharge of wastewater was specified for facilities in the
following subcategories and subparts under BPT:
1. Iron Ore Mills in the Mesabi Range;
2. Copper, Lead, Zinc, Silver, Gold, Platinum and
Molybdenum Mines and Mills that leach to recover copper;
3. Gold Mills that use cyanidation; and
4. Mercury Mills.
Facilities in these subcategories and subparts have achieved the
goal of the Clean Water Act and no additional reduction of toxics
is possible. Therefore, the BAT effluent limitations are the
same as under BPT.
Subcategories and Subparts Where BAT Limitations Are Developed
There were subcategories and subparts in which some of the toxic
pollutants were detected and are not excluded from regulation
development. These include:
1. Copper, Lead, Zinc, Gold, Silver, Platinum, and
Molybdenum mine drainage, mill process water from facilities
using froth flotation, and placer mines;
2. Tungsten mine drainage and mill process water;
3. Mercury mine drainage;
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4. Uranium mine drainage; and
5. Titanium mill process water.
Recycle was considered as an option for froth flotation mills in
the copper, lead, zinc, gold, silver, platinum, and molybdenum
ores subcategory. This option was rejected for froth flotation
mills because of the extreme costs that would be incurred during
retrofit, the downtime required to retrofit existing equipment
and the impact of changing the metallurgical process.
Recycle was also considered as an option for placer mines
recovering gold. The placer mining industry consists primarily
of small operations located in remote areas in Alaska. Placer
mining involves recovery of gold and other heavy mineral deposits
by washing, dredging, or other hydraulic methods. Chemical
reagents are not used in the processing of the deposits. For
this reason, the pollutants of primary concern are suspended or
settleable solids which may result during recovery.
Arsenic and mercury were found in placer effluents during recent
studies because (U arsenic occurs naturally in abundance in many
areas of Alaska; and (2) mercury has been used extensively in the
past by placer miners for recovery of gold in sluiceboxes, and
mercury residuals are undoubtedly present in old deposits
presently being reworked by modern day miners. Results of this
study indicate that effective removal of the total suspended and
settleable solids by settling also resulted in effective removal
of arsenic and mercury.
At a few placer mines it may be technically feasible to recycle
water for reuse in sluicing gold-bearing sediments. However, the
location of most of the operations, the fact that electric power
is not available to run pumps and the magnitude of the costs and
energy requirements mitigate against this practice. As a result,
EPA has selected settleable solids limitations based on settling
ponds as the means for controlling discharges from placer mining
operations. The choice of settleable solids frees the operators
from having to ship samples from remote locations to laboratories
for analysis. The analytical method is undemanding, inexpensive,
short-term duration test that can be performed by large and small
operators alike.
The settelable solids data from placer mining facilities included
two separate studies of existing placer mines in Alaska and other
studies performed by EPA and by departments of the State of
Alaska. However, the actual data for effluent from existing
settling ponds associated with gold placer mines is limited
because many of the mines, including mines in the data base, have
no settling facilities. Of the remaining mines which have
settling ponds, it was identified that the majority, if not all,
of the existing ponds for which we have data, were undersized,
filled with sediment, short circuited, or otherwise poorly
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operated to remove settlable solids from the wastestreams of
placer mines. The data from well constructed, operated, and
maintained settling ponds is limited to demonstration projects
and a few existing settling ponds which may not be truly
representative of gold placer mining operations (e.g., mines
located outside of the boundries of streams or floating dredge
operations).
Cost comparisons for two treatment technologies (primary settling
followed by secondary settling and primary settling with
flocculation) were perfomred including the subsequent cost per
ton of ore mined. However, no economic analysis was perfomred
for the gold placer mining subpart because no data are available
that would enable the Agency to perform cash flow analysis of
placer mine operations.
Limitation for gold placer mines are reserved in the proposed BAT
rulemaking in the absence of information regarding the economic
impact of regulating gold placer mines and to allow the Agency to
seek additional data on the effluent from well operated settling
ponds assocaited wi.th gold placer mines.
The method used to compute the achievable levels for the toxic
metals is summarized below and presented in greater detail in
Supplement B. The data obtained during sampling and analysis as
well as that supplied by industry were reviewed and effluent
levels achievable were computed for each toxic metal considered
for regulatory development in Section VII. As discussed in
Section VII, TSS removal technologies also remove metals, so the
following TSS removal measures were considered for metal removal:
1. Secondary settling;
2. Coagulation and flocculation; and
3. Granular media filtration.
Eighteen facilities throughout the ore mining and dressing
industry were identified as using multiple settling ponds; 14
facilities using coagulation and flocculation; and one facility
using granular media filtration. The entire BAT and BPT data
base was searched and screened to obtain 17 facilities with data.
Of these 17 facilities, seven were eliminated because it was
believed that they were not operated properly (e.g., short
circuiting in the settling ponds was observed) or no raw
(untreated) wastewater data were available to compare with
treated effluent.
The facility treated effluent mean values were ranked for each of
the 10 remaining facilities for each pollutant from largest to
smallest. Since each facility used only one of the candidate BAT
treatment technologies, the facility mean also represented a
treatment technology mean value. When examining the ranked mean
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values, it was observed that mean values for facilities using
secondary settling bracketed those for facilities using
flocculation and granular media filtration. This variation
indicated that the differences between facilities were greater
than the differences between treatment technologies. Possibly,
differences existed between the true performance capabilities of
the treatment technology; however, on the basis of available
data, one cannot discern such differences.
The 10 facilities were then further reduced to six by eliminating
facilities whose raw (untreated) waste contained low pollutant
concentrations. This was done to ensure that only those
facilities which demonstrated true reduction would be included in
the analysis. Data for a particular pollutant were excluded if
the median raw wastewater concentration was less than the average
facility effluent concentration of any other facility. Of the
six facilities, five use secondary settling and one uses granular
media filtration. Since there were no discernable differences in
the levels achievable by the three technologies (based on
available data), the least costly alternative was selected for
establishing effluent limitations, secondary settling.
Achievable levels were computed by using the average of the
facility averages for each pollutant to represent, the average
discharge. The data used were from the five facilities using
secondary settling (two copper, two lead/zinc, and one silver)
that remained following the screening procedures described above.
The data base indicates that within-plant effluent concentrations
were approximately log normally distributed. The 30-day average
maximum and daily maximum effluent limits were determined on the
basis of 99th percentile estimates. The 30-day limits were
determined by using the central limit theorem. The achievable
levels computed for each of the metals and TSS are shown below:
30-Day Daily
Average Maximum
Arsenic 0.01 0.05
Cadmium 0.005 0.01
Copper 0.05 0.20
Lead 0.04 0.14
Mercury 0.001 0.002
Nickel 0.10 0.40
Zinc 0.20 0.80
TSS* 10 25
*TSS limitations were computed, but TSS
would be limited under BCT.
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The limitations derived from the data analysis for some pollutant
metals were more stringent than the BPT limitations.
Having computed achievable levels for the candidate BAT
technologies, EPA then completed an environmental assessment,
which analyzed the environmental significance of toxic pollutants
currently discharged from facilities in this industry and also
those toxic pollutants known to be discharged from this industry
at BPT and expected to be discharged based on the computed
achievable levels. The basis for determining the environmental
significance of toxic pollutants in current discharges is a
comparison of average plant effluent concentrations with Ambient
Water Quality Criteria (WQC) for the protection of human health
and aquatic life published by the EPA's Criteria and Standards
Division (CSD) in November 1980. Because WQC for the protection
of aquatic life were not developed for all of the Section
307(a)(l) toxic pollutants, the average plant effluent concentra-
tions for pollutants lacking these WQC are compared with pollu-
tant-specific toxicity data reported in the Ambient Water Quality
Criteria Documents. The environmental significance of toxic
pollutants in post-BAT discharges is determined by comparing the
achievable levels with WQC or, for those pollutants lacking WQC
for the protection of acquatic life, with EPA toxicity data.
Based on a review of the sampling and analysis data available for
this industry, the only environmentally significant pollutants
after applying the median dilution from the average receiving
stream flow available (to this industry) are cadmium and arsenic.
The concentration of cadmium currently being discharged from this
industry (BPT) is the lowest of any industry known to discharge
cadmium. In addition, the additional BAT reductions are small
relative to the levels present in raw (untreated) waste streams.
In preparing the environmental assessment, the Agency also
compared raw waste mass loadings to those of BPT and those
expected by achievable levels. It was found that the industry's
current discharge is less than 10 percent of the industry's raw
waste load. This is due to the installation and proper
maintenance of the Best Practicable Technology at many plants.
In other information and data available to the Agency at the time
this Development Document was written, the BAT limitations
proposed or being considered for metals in discharges from other
industries are less stringent than those promulgated for BPT in
the Ore Mining and Dressing Industry. In examining limitations
for Coil Coating, Metal Finishing and Inorganic Chemicals, the
following observations may be made with respect to the
limitations:
1. Based on the use of similar technology (lime precipi-
tation and settling), proposed BAT limitations on copper,
cyanide, mercury and zinc for Coil Coating are less
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stringent than those promulgated as BPT for Ore Mining and
Dressing. Cadmium and lead limitations are more stringent.
2. Based on the use of similar technology (lime precipi-
tation and settling), proposed BAT limitations for the
following pollutants being considered for Inorganic
Chemicals are less stringent than in Ore Mining BPT in the
following subcategories.
- Sodium Dichromate - zinc (copper, lead, cyanide, cadmium
and mercury are not regulated)
Copper Sulfate - copper, cadmium and zinc (mercury and
cyanide are not regulated; lead is more stringent)
Nickel Sulfate - copper, cadmium, and zinc (mercury and
cyanide are not regulated; lead is more stringent)
Sodium Bisulfite - copper, mercury, and zinc (cadmium and
cyanide are not regulated; lead is more stringent)
Sodium Hydrosulfite - copper, lead and zinc (cadmium,
cyanide and mercury are not regulated)
Aluminum Fluoride - copper (cadmium, cyanide, lead, and
mercury are not regulated; zinc is more stringent)
Titanium Dioxide - cadmium, copper and zinc (cyanide and
mercury are not regulated; lead is more stringent)
3. Based on the use of similar technology (lime precipita-
tion and settling), limitations for the Common Metals
Subcategory in Metal Finishing on copper and zinc are less
stringent than those promulgated for Ore Mining and
Dressing. Cadmium and lead limitations are more stringent.
EPA also considered the solubility of compounds in which the
toxic metals are found in the industry effluent. The metals
present in the wastewater from ore milling are found primarily in
the ore and gangue discharged by the mill. The metals present in
mine drainage similarly consist primarily of ore and gangue. The
metals in ore and gangue are primarily sulfides and to a lesser
extent oxides. The metals as mined generally exist as sulfides:
as example, chalcopyrite (CuFeS2), bornite (Cu5FeS4), covellite
(CuS), and chalcocite (Cu2S) for copper; galena (PbS) for lead;
sphalerite (ZnS) for zinc; and argentite (Ag2S), proustite
(Ag3AsS3) and stephanite Ag5S4Sb for sliver. The metals may also
exist as
for copper;
zinc; and
example,
(Zn2Si04)
oxides: as example, tenorite
pyromorphite Pb5Cl(P04)3
wulfenite (PbMo04) for
hemimorphite, Zn4(Si207) (OH2)»H20
for zinc (Reference 5). The oxides
(CuO) and cuprite (Cu20)
for lead; zincite (ZnO) for
molybdenum or silicates, as
and willemite
and silicates are
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less soluble than the sulfides and the sulfides are far less
soluble than the hydroxides as shown in the table below.
Solublility Products of Toxic Metals
Solubility Product
Toxic Metal Metal Hydroxide Metal Sulfide
Cadmium, Cd
Copper, Cu
Lead, Pb
Mercury, Hg
Zinc, Zn
13.6
18.6
16. 1
25.4
15.7
26. 1
35.2
26.6
52.2
25.2
Source: Development Document for Effluent Limitations
Guidelines and Standards for the Inorganic
Chemicals Manufacturing Point Source Category,
EPA 440/1-79/007, June 1980.
By contrast, in many other industrial point source categories
including, for example, the coal coating, metal finishing,
inorganic chemical categories, toxic metals are introduced with
raw waste as dissolved species. The toxic metals in those
industries are present in the wastewater either in the dissolved
form or as metal hydroxide. As indicated by the above table
metal hydroxides are of comparatively high solubility.
The metal hydroxides (solids) present in the wastewater of these
other industries have a greater potential of redissolving into
solution and being available to the environment to be assimilated
thereby posing a greater danger to human and aquatic species.
The metals (solids) present in the wastewater from ore milling
consist primarily of the minerals associated with the ore and
gangue discharged by the mill. The metals (solids) present in
mine drainage similarly consists of ore and gangue. The metals
in the ore and gangue are generally the sulfides of the metal and
are of relatively low solubility.
After considering the environmental assessment, the BAT
limitations proposed for other industries, the physical form the
toxic metals, and economic factors, the Agency has concluded that
nationally applicable regulations based on secondary settling or
any of the other candidate BAT technologies are not warranted in
the Ore Mining and Dressing Point Source Category.
Cyanide Control and Treatment
As discussed in Section VIII, cyanide compounds are used in froth
flotation process of copper, lead, zinc, and molybdenum, ores.
In addition, the cyanidation process is used for leaching of gold
and silver ores. Consequently, residual cyanide is found in mill
tailings and wastewater streams from these mills. Cyanide is
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also found in
which backfill
tailings.
low concentrations
mine stopes with
in mine drainage at facilities
the sand fraction of mill
Of the control and treatment technologies available for cyanide,
consideration was given to the following options: in-process
control, chemical oxidation (alkaline chlorination, hydrogen
peroxide oxidation, ozonation), and natural oxidation and the
incidental removal occurring in existing treatment systems
(tailing ponds). These options were judged to be most applicable
to the high flow volume and comparatively low concentrations of
cyanide in the wastewater streams typical in this category.
Another alternative which was considered was the substitution of
other reagents for cyanide compounds in froth flotation
processes. Bench-scale testing indicated that this alternative,
although technically feasible, would require extensive testing in
actual production of circumstances with specific ores. In
addition, it would be difficult in these cases to predict
downtime, loss of recovery (if any), and costs associated with
process modifications.
Alkaline Chlorination
This method was described in
operating cost assumptions
Basically, oxidation of cyanide
accomplished by infusion of
stream at a pH greater than 10,
detail
in Section VIII, while
are outlined in Section IX.
by alkaline chlorination may be
gaseous chlorine into the waste
or by the addition of sodium
hypochlorite (NaOCl) as an oxidant along with an alkali such as
sodium hydroxide (NaOH). The alkali achieves pH adjustment and
precipitation of metal hydroxides formed from the breakdown of
metal-cyanide complexes.
Pilot-scale tests of alkaline chlorination treatment at Mill 6102
showed reduction of effluent cyanide concentrations from 0.19
mg/1 to less than 0.1 mg/1 at pH values greater than 8.8. In
addition, Mill 3144 achieved reduction of effluent cyanide
concentrations to an average of 0.18 mg/1 from 4.72 mg/1
following the installation of a full-scale alkaline-chlorination
treatment system.
Ozonation
Oxidation of cyanide by ozonation is also accomplished at ele-
vated pH (9 to 12). Copper appears to act as a catalyst in this
process, which suggests that waste streams containing copper
cyanide complexes may be treated more effectively by ozonation.
Pilot-scale testing of ozonation at Mill 6102 showed reduction of
cyanide concentration from 0.55 mg/1 to less than 0.1 mg/1 at pH
greater than 7.4.
Hydrogen Peroxide Treatment
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Hydrogen peroxide (H202) has also been tested on a limited basis
as an oxidant for cyanide treatment in milling wastewater
streams. This process also requires an alkaline pH and can be
enhanced by a copper catalyst. Mill 6101 has achieved approxi-
mately 40 percent removal of cyanide during periods of elevated
effluent levels (up to approximately 0.09 mg/1) by hydrogen
peroxide oxidation.
Process Control
One characteristic of the froth flotation process which poten-
tially affects effluent wastewater quality is the latitude avail-
able to the mill operator at the upper end of the dosage applica-
tion spectrum. That is, while the addition' of less than the
necessary quantities of cyanide reagent may lead to loss of
recovery or reduced product purity, the addition of more than the
necessary quantities of cyanide reagent is not accompanied by
penalties to the same degree, except of course, the cost of the
additional reagent.
Close attention to mill feed characteristics and careful and
frequent analysis of its mineral content can result in reduction
of cyanide dosage to that actually required. In recent years,
on-line analysis techniques and reagent addition controls have
become available to minimize excess additions of reagent.
Few froth flotation process facilities in the industry have
reported treated effluent cyanide concentrations equal to or in
excess of 0.1 mg/1, and these only on an infrequent basis. Mill
6102, the largest consumer of cyanide in terms of dosage per unit
of ore feed, has been observed in the past to generate effluent
cyanide concentrations as high as 0.2 mg/1 to 0.4 mg/1. Follow-
ing installation of cyanide treatment, this facility is reporting
cyanide levels less than 0.1 mg/1. Three other flotation mills
have reported discharge concentrations in excess of 0.1 mg/1
(2122, 3121, 6101). In each case, the cyanide dosages used in
mill feed appear to be consistent with dosages reported through-
out the industry and are not unusual in that respect. Fluctua-
tions and peaking in cyanide concentrations appear to be related
to short-term overdoses of cyanide in the flotation process. Few
treated effluent measurements in the entire industry have
exceeded 0.2 mg/1 and we believe that, with close process control
and reagent addition in combination with a well designed and
operated treatment system, the 0.2 mg/1 measurement for total
cyanide can be achieved without additional treatment technology
for cyanide. For the rare case where difficulty may be
encountered or great reliability is required, treatment
technology (i.e., chemical oxidation) is available as discussed
in Section VIII and costed in Section X.
Many existing NPDES permits for ore mills contain limitations on
total cyanide. As example, in EPA Region VIII, there are nine
existing permits that limit total cyanide in the discharge from
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ore mills and these limitations vary from .02 mg/1 to .2 mg/1
daily maximum. In EPA Region X there are twelve permits for ore
mills that limit total cyanide and these limitations vary from
.01 mg/1 to .3 mg/1 daily maximum. Monitoring data for these
permits confirm that these mills are consistently within their
permit limitations on total cyanide and that the limitations can
be obtained by control of the process and the incidential removal
of cyanide as discussed below.
Incidental Cyanide Removal
Frequently, specific cyanide treatment technology is not
necessary if close process control combined with incidental
removal leads to low concentrations of total cyanide in mill
water treated effluent. This incidental removal is thought to
involve several mechanisms, including ultraviolet irradiation,
biochemical oxidation, and natural aeration. As evidence that
such mechanisms are involved, it has been noted that effluent
cyanide concentrations tend to be somewhat higher during winter
months when biological activity in the tailing pond is lower and
ultraviolet exposure is much lower due to shortened daylight
hours, less intense radiation, and ice cover on the ponds.
In addition, the association of cyanide with the depressed
minerals (i.e., pyrites) will cause a portion of the cyanide to
be removed together with the suspended solids and deposited in
the tailing ponds.
Precision and Accuracy Study
A study of the analysis of cyanide in ore mining and processing
wastewater was conducted in cooperation with the American Mining
Congress to investigate the causes of analytical interferences
observed and to determine what effect these interferences had on
the precision of the analytical method. The purpose of this
study was to evaluate the EPA-approved method and a modified
method for the determination of cyanide. The modified method
employed a lead acetate scrubber to remove sulfide compounds
produced during the reflux-distillation step. Sulfides have been
suspected of providing an interference in the colorimetric
determination of cyanide concentrations. Also, several samples
were spiked with thiocyanate to ascertain if this compound caused
interference in the cyanide analysis.
A statistical analysis of the resultant data shows no significant
difference in precision or accuracy of the two methods employed
when applied to ore mining and milling wastewaters having cyanide
concentrations in the 0.2 mg/1 to 0.4 mg/1 range. Based upon the
statistical analysis, approximately 50 percent of the overall
error of either method was attributed to intralaboratory error.
This highlights the need for an experienced analyst to perform
cyanide analyses. After considering the results of this study
and the levels achieved through dose control of reagent addition
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in the mill, EPA considered proposing an effluent limitation of
0.2 mg/1. That limitation is based on a grab sample for any one
day, and would have been subject to 100 percent error to account
for the precision and accuracy of the analytical method. (See
Section V above). Therefore, the Agency would have had to allow
an analytical measurement of up to 0.4 mg/1. However, all the
data observations in our sample were below that level.
Accordingly, the Agency is proposing to exclude cyanide from
national regulation in the ore mining category.
ADDITIONAL PARAGRAPH 8. EXCLUSIONS
Exclusion p_f Cyanide
Total cyanide is not regulated in BAT because the Agency cannot
quantify a reduction in total cyanide from observed
concentrations being discharged by use of technologies, known to
the Administrator, Paragraph 8(a) iii of the Settlement
Agreement.
The references to total cyanide levels of less than 0.2 mg/1
throughout this document are for informational purposes only and
are subject to the precision and accuracy of the analytical
method.as discussed here and in Section V.
Exclusion of_ Arsenic and Nickel
EPA reviewed the achievable levels calculated based on the
capabilities of the three candidate BAT treatment technologies
(secondary settling, coagulation and flocculation, and granular
media filtration). The Agency examined the necessity of
proposing specific limitations for all seven of the toxic metals
considered for regulation. Limitations on copper, lead, and zinc
are necessary since these are the metals recovered from mining
operations and concentrated in mills in this category. From a
treatability viewpoint, control of some toxic metals (arsenic and
nickel) may be achieved by limitations upon which other
pollutants are controlled. As discussed in this section, since
most of the metals are in suspended solids, reduction of arsenic
and nickel occurs in conjunction with the removal of TSS and
other toxic metals (copper, lead, and zinc).
The BAT data base for the Ore Mining and Dressing Point Source
Category was searched for instances in which arsenic and nickel
concentrations exceeded BPT limitations when copper, lead, and
zinc concentrations were also below their respective BPT limita-
tions. There was only one instance in over 300 samples in which
a nickel or arsenic concentration exceeded their BPT limitations
when BPT limitations for copper, lead, and zinc were met. The
one instance was the discharge from a sedimentation pond at
Facility 3103. The nickel concentration was 0.22 mg/1 as opposed
to the 0.20 mg/1 BPT limitation.
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The Agency concluded that the limitations on copper, lead, and
zinc would ensure adequate control of arsenic and nickel, and
under Paragraph 8(a)iii of the Settlement Agreement, arsenic and
nickel are excluded from regulation.
Exclusion of Asbestos
Chrysotile asbestos was detected in wastewater samples in all
subcategories and subparts within the ore mining and dressing
point source category. It was detected in 90 of 91 samples
throughout the entire industrial category.
EPA believes that the most appropriate way to regulate a toxic
pollutant is by a direct limitation on the toxic pollutant.
However, direct limitation of toxic pollutants is not always
feasible. In the case of chrysotile asbestos, there is no EPA
approved method of analysis for industrial wastewater samples.
The method of analysis presently used was developed for drinking
water samples. In addition, there are less than half a dozen
laboratories in the United States that are capable of performing
the analysis by this method.
Chrysotile asbestos is known to be present in many ore deposits
throughout the country (Reference 6). As ore is mined and
subsequently milled, it is subjected to a variety of crushing and
size reduction operations. As a result, smaller solids are
formed, the chrysotile asbestiform fibers are liberated as the
small solids are made, and end up in mine drainage and mill
process water.
The possibility of the chrysotile asbestos fibers being present
in waste streams for the same reasons and in the same relative
proportions as the solids, led to the examination of the EGD
sampling data in an attempt to establish a relationship between
chrysotile asbestos and TSS.
Review of analyses for asbestos and TSS in samples of untreated
and treated wastewater shows that as TSS is reduced by treatment,
observed asbestos concentrations are also reduced.
Intake water samples (from 26 industrial categories) and POTW
effluents were reviewed to get an indication of background levels
for chrysotile asbestos. The values of chrysotile asbestos
ranged from 3.5 x 104 (detection limit) to 1.63 x 108 with a mean
value of 1.1 x 107 fibers per liter. The treated waste stream
sample values for chrysotile asbestos in the ore data ranged from
104 (detection limit) to 108 fibers per liter.
The Agency has determined that when TSS is reduced to the BPT
effluent limitations for ore mining and dressing, observed
chrysotile asbestos levels are reduced to or below background
levels. Therefore, EPA is excluding chrysotile asbestos from
regulation since it is effectively controlled by technologies
521
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upon which TSS limitations are based, Paragraph 8(a)iii of the
Settlement Agreement.
Exclusion of_ Pollutants Detected in_ a Single Source and Uniquely
Related to That Source
There are 19 operating uranium mills in the United States, 18 of
which now achieve zero discharge of process wastewater. There
are no uranium mills that commingle process wastewater with mine
drainage and it is anticipated that none of these zero discharge
mills would elect to treat and discharge at the BPT limitations
because of the expense to install technology required, i.e., ion
exchange, ammonia stripping, lime precipitation, barium chloride
coprecipitation, and settling.
EPA is excluding uranium mills from BAT, since there is only one
discharging facility and it is believed that none of the other
existing facilities will commingle mine drainage and mill process
wastewater. Uranium mills are not regulated in BAT because the
pollutants found in the discharge are uniquely related to this
single source, Paragraph 8(a)iii of the Settlement Agreement.
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SECTION XI
BEST CONVENTIONAL POLLUTANT CONTROL TECHNOLOGY
Section 301(b)(2)(E) of The Act requires that there be achieved,
not later than July 1, 1984, effluent limitations for categories
and classes of point sources, other than publicly-owned treatment
works, that require the application of the best conventional
pollutant control technology (BCT) for control of conventional
pollutants as identified in Section 304(a)(4). The pollutants
that have been defined as conventional by the Agency, at this
time, are biochemical oxygen demand, suspended solids, fecal
coliform, oil and grease, and pH.
BCT is not an additional limitation; rather, it replaces BPT for
the control of conventional pollutants. BCT must be evaluated
for cost effectiveness and a comparison made between the cost and
level of reduction of conventional pollutants from the discharge
of publicly owned treatment works (POTW) and the cost and level
of reduction of such pollutants from a class or category of
industrial sources.
The technologies considered for treatment of conventional
pollutants are the same as those considered for treatment of
toxic pollutants. As discussed in Section X, the Agency has
determined that the BAT limitations for toxics are equivalent to
BPT limitations for the ore mining and dressing industry. BCT
limitations for the conventional pollutants, TSS, and pH are also
proposed at BPT levels. Accordingly, by definition, BCT for this
industry meets any BCT cost test because there is no incremental
cost to remove conventional pollutants beyond BPT.
Summary of Best Conventional Technology
Daily 30-Day
Maximum Average
Pollutant (mq/1) (mq/1)
TSS 30 20
pH Within the range of
6 to 9
The Agency is currently developing a new BCT cost methodology.
It is possible, though unlikely, that a treatment technology more
stringent than BPT will provide additional removal of
conventional pollutants and pass the new cost test, when
developed. In that event, the Agency will propose new BCT
limitations.
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524
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SECTION XII
NEW SOURCE PERFORMANCE STANDARDS (NSPS)
The basis for new source performance standards (NSPS) under
Section 306 of the Act is through application of the best avail-
able demonstrated technology. New facilities have the opportu-
nity to implement the best and most efficient ore mining and
milling processes and wastewater technologies. Congress, there-
fore, directed EPA to consider the best demonstrated process
changes and end-of-pipe treatment technologies capable of
reducing pollution to the maximum extent feasible.
GENERAL PROVISIONS
Several items of discussion apply to options in more than one
subcategory. To avoid repetition, these items are discussed
here.
Relief From Np_ Discharge Requirement
For new sources that must achieve no discharge of process
wastewater, the Agency provides relief upon the occurrence of a
10-year, 24-hour precipitation event or snow melt of equivalent
volume. Excess water resulting from the occurrence shall not be
subject to effluent limitations.
Relief from no discharge of process wastewater is also granted
for those facilities located in net precipitation areas and is
the same relief granted to BAT limitations requiring no
discharge.
Relief From Effluent Limitations for Those Facilities Permitted
to Discharge
The relief is exactly the same as that granted for existing
sources under BAT. Excess water resulting from precipitation
from a facility designed, constructed, and maintained to contain
or treat the maximum volume of process wastewater discharged
during any 24-hour period, including the volume that would result
from a 10-year, 24-hour precipitation event or snow melt of
equivalent volume, shall not be subject to the limitations set
forth in 40 CFR 440.
Commingling Provisions
For new sources that combine for treatment waste streams from
various sources, the quantity and quality of each pollutant or
pollutant property in the combined discharge that is subject to
effluent limitations shall not exceed the quantity and quality of
each pollutant or pollutant property that would have been
discharged had each waste stream been treated separately. The
discharge flow from a combined discharge shall not exceed the
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volume that would have been discharged had each waste stream been
treated separately.
NSPS OPTIONS CONSIDERED
The Agency considered the following NSPS options:
Option One. Require achievement of performance standards in each
subcategory based on the same technology as BAT (NSPS = BAT).
Option Two. Require standards based on a complete water recycle
system (NSPS = zero discharge).
NSPS SELECTION AND DECISION CRITERIA
EPA has selected performance standards based on the same
technology as BAT for all facilities in the ore mining and
dressing point source category, except those facilities using
froth flotation in the copper, lead, zinc, gold, silver,
platinum, and molybdenum subcategory and mills in the uranium
subcategory.
Subcategories and Subparts Under Option ]_
Option 1 (NSPS = BAT) has been selected for iron ore mills in the
Mesabi range; copper, lead, zinc, silver, gold, platinum, and
molybdenum mills that use leaching to recover copper and the
cyanidation process for the recovery of gold; and mercury mills
since BAT specifies zero discharge. Option 1 (NSPS = BAT) has
also been selected for iron ore mine drainage, iron ore mills,
aluminum mine drainage, copper, lead, zinc, gold, silver,
platinum, and molybdenum mine drainage, titanium mine drainage,
dredges and mills, and mercury mine drainage. The concentration
levels of toxic metals found in new sources in these
subcategories and subparts are expected to be similar to existing
sources. Following the implementation of BAT, toxic metals will
be found at or near detection levels or at concentrations below
the practical limits of additional technology. Further reduction
of these pollutants can not be technically or economically
justified.
Subcategories and Subparts Under Option 2_ (NSPS = zero discharge)
EPA is proposing that new source froth flotation mills achieve
zero discharge of process wastewater. EPA considered zero
discharge based on recycle for existing copper, lead, zinc, gold,
silver, platinum, and molybdenum mills using froth flotation, but
rejected it because of the extensive retrofit required at some
existing facilities, the cost of retrofitting, and the possible
changes required in the process. This concern does not apply to
new sources. Zero discharge is a demonstrated technology at 46
of the 90 froth flotation mills for which EPA has data (see
wastewater discharges as summarized in Tables IX-2 through IX-10)
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and meets the definition of standard of performance permitting
zero discharge of pollutants. New sources have the option to
recycle because the metallurgical process can be adjusted and
designed to recycle process wastewater before the actual
construction of the new source. While reagent buildup has been
mentioned by industry as a potential problem in extractive
metallurgy, no evidence has been submitted to validate ths
assertion.
There are new sources anticipated in copper, lead, zinc, gold,
silver, and molybdenum mining. Standards applied to these new
source waste streams should reflect the best treatment levels
achievable by the froth flotation segment of the industry.
A study of existing froth flotation mills reveals that a large
percentage of these facilities are effectively achieving TOO
percent recycle of mill water. Many of the facilities practicing
100 percent recycle are located in arid regions, but some
facilities are located in humid regions. A summary of some
existing facilities follows.
Copper Ore
Of the 35 known froth flotation copper mills in the Un.ited
States, 31 achieve zero discharge of process wastewater.
Lead/Zinc Ores
Five of the 27 active froth flotation mills in the lead/zinc
subcategory achieve zero discharge.
Gold
Four of the five primary gold facilities employing froth
flotation techniques discharge process wastewater.
Silver
Three of the four known primary silver facilities which use froth
flotation methods are achieving zero discharge of process
wastewater.
Molybdenum
Of the three molybdenum operations employing the froth flotation
process, one facility achieves zero discharge of recycle.
In Section IX (Table IX-3 through Table IX-10) cost comparisons
for treatment technologies are made for existing sources. It is
believed that new source mills will have similar mill capacity,
water use per ton of ore, and pollutants in the raw wastewater as
existing mills. Therefore, costs of technology for new sources
will approach those costs determined for existing sources. The
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cost of 100% recycle for mills is generally less than the cost
determined for pH adjustment of the wastewater (lime addition).
Assuming the same settling pond, or tailings pond, would be
required for recycle as for pH adjustment and settling, the cost
to recycle at a new source is approximately the same, or less,
than the cost to implement the technology (pH adjustment and
settle) upon which BPT and BAT effluent limitations are based.
EPA is proposing that new source uranium mills achieve zero
discharge of process wastewater. For this subpart, EPA
considered zero discharge for BAT based on total impoundment and
evaporation, or recycle and reuse of the mill process water, or a
combination of these technologies. Because the pollutants
detected in the current discharge from this subpart are uniquely
related to one point source, the single mill discharging, the
uranium mill subpart is excluded from BAT under Paragraph 8
authority of the Settlement Agreement (see Section X).
However, the Agency believes that for new sources a standard of
performance must be proposed. Otherwise, additional discharges
(new sources) could occur that obviously would not be unique to
one source. New source uranium mills are anticipated by the
Agency. New mill capacity is anticipated to replace existing
mills and to maintain the current production of uranium oxide as
lower grade ore must be mined. Also, an increase in demand for
uranium oxide is anticipated that will require new mills.
Uranium mills can achieve zero discharge as indicated by the fact
that 18 of 19 existing mills currently achieve no discharge.
As discussed in Section X, ammonia stripping, lime precipitation,
barrium chloride co-precipitation or ion exchange, and settling
are the identified technologies for uranium mills to meet BPT
limitations on metals, radium 226, ammonia and TSS. The cost to
implement the technologies to meet the BPT limitations for a new
uranium mill is more than the cost to implement recycle or
evaporation ponds (or the combination of the two) to meet a no
discharge requirement.
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SECTION XIII
PRETREATMENT STANDARDS
Section 307(b) of the Act requires EPA to promulgate pretreatment
standards for both existing sources (PSES) and new sources (PSNS)
of pollution which discharge their wastes into publicly owned
treatment works (POTWs). These pretreatment standards are
designed to prevent the discharge of pollutants which pass
through, interfere with, or are otherwise incompatible with the
operation of POTWs. In addition, the Clean Water Act of 1977
adds a new dimension of these standards by requiring pretreatment
of pollutants, such as heavy metals, that limit POTW sludge
management alternatives. The legislative history of the Act
indicates that pretreatment standards are to be technology based
and, with respect to toxic pollutants, analogous to BAT. The
Agency has promulgated general pretreatment regulations which
establish a framework for the implementation of these statutory
requirements (see 43 FR 27736, 16 June 1978).
EPA is not proposing pretreatment standards for existing sources
(PSES) or new sources (PSNS) in the ore mining and dressing point
source category at this time nor does it intend to promulgate
such standards in the future since there are no known or
anticipated discharges to publicly owned treatment works (POTWs).
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tn
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SECTION XIV
ACKNOWLEDGEMENTS
This document was drafted by Radian Corporation, McLean, Virginia
under the direction of Mr. Baldwin M. Jarrett, Project Officer,
Energy and Mining Branch, Effluent Guidelines Division, EPA.
Direction and assistance were also provided by Mr. William A.
Telliard, Chief of the Energy and Mining Branch and Mr. Ron
Kirby, Project Monitor. The insights and review provided by Mr.
Barry Neuman, formerly of the Office of General Counsel, are also
expressly appreciated. Much of the input for this document was
provided by Radian's subcontractors Frontier Technical
Associates, Buffalo, New York, and Hydrotechnic Corporation, New
York, New York.
The following divisions of the Environmental Protection Agency
(EPA) contributed to the development of this document: All
regional offices; Industrial Environmental Research Laboratory
Cincinnati, Ohio; Office of Research and Development; Office of
General Counsel; Office of Planning and Evaluation; Monitoring
and Data Support; Criteria and Standards Division; Office of
Quality Review; and Office of Analysis and Evaluations.
Appreciation is extended to the following trade associations and
mining companies for assistance and cooperation during the course
of this program:
Aluminum Association
American Iron Ore Association
American Mining Congress
Aluminum Company of America
Amax Lead Company of Missouri
American Exploration and Mining Company
American Smelting and Refining Company
Anaconda Copper Company
Atlas Corporation
Bethlehem Mines Corporation
Brush Wellman Incorporated
Bunker Hill Company
Carl in Gold Mining Company
Cities Service Company
Cleveland-Cliffs Iron Company
Climax Molybdenum Company
Cominco American, Inc.
Continental Materials Corporation
Copper Range Company
Curtis Nevada Mines, Inc.
Cyprus-Bagdad Copper Corporation
Eagle Pitcher Industries, Inc.
E. I. DuPont de Nemours and Company, Inc.
Erie Mining Company
Goodnews Bay Mining Company
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Hanna Mining Company
Hecla Mining Company
Homestake Mining Company
Idarado Mining Company
Inspiration Consolidated Copper Company
Jones and Laughlin Steel Corporation
Kennecott Copper Corporation
Kerramerican, Inc.
Kerr McGee Corporation
Knob Hill Mines, Inc.
Lead and Zinc Institute
Magma Copper Company
Marquette Iron Mining Company
Molybdenum Corporation of America
National Lead Industries, Inc.
New Jersey Zinc Company
Oat Hill Mining Company
Oglebay-Norton Company - Eveleth Taconite
Phelps Dodge Corporation
Pickands Mather and Company - Erie Mining Company
Ranchers Exploration and Development Corporation
Rawhide Mining Company
Reynolds Mining Corporation
Standard Metals Corporation
St. Joe Minerals Company
Sunshine Mining Company
Titanium Enterprises
Union Carbide Corporation
United Nuclear Corporation
U.S. Antimony Corporation
U.S. Steel Corporation
White Pine Copper Company
The initial draft of this document containing most of the
information and data on which EPA relies was developed and
written by Calspan Corporation, Buffalo, New York. Copies of
this draft were distributed to Federal and State agencies, trade
associations, conservation organizations, industry, and
interested citizens. We wish to thank the following who
responded with written comments: American Mining Congress;
Bunker Hill Company; Natural Resources Defense Council, Inc.;
Prather, Seeger, Doolittle, and Farmer; St. Joe Minerals
Corporation, Trustees for Alaska; U.S. Department of Interior -
Bureau of Mines; U.S. Department of Labor; USEPA - Environmental
Research Laboratory (Athens, Georgia); Walter C. McCrone
Associates, Inc.; White Pine Copper Company. An effort has been
made in this document to address those comments which pointed to
deficiencies in the record and data contained in the draft.
532
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SECTION XV
REFERENCES
SECTION III
1. Dennis, W. H., Extractive Metallurgy Principles and
Application, Sir Issac Pitman & Sons, Ltd., London, 1965.
2. U.S. Bureau of Mines, A Dictionary of Mining, Mineral and
Related Terms. Compiled and edited by P. W. Thrush and the USBM
Staff. Washington, D.C.: U.S. Department of Interior (1968).
3. Cummins, A. B. and I. A. Given, Mining Engineering 'Handbook,
Volumes I and II. Littleton, Colorado; Society of Mining
Engineers (1973).
4. "Iron Ore 1976," American Iron Ore Association, Cleveland,
Ohio, 1976.
5. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1978/1979.
6. Personal communications from Commodity Specialist, Bureau of
Mines, U.S. Department of the Interior, to M. A. Wilkinson,
Calspan Corporation, January 1978.
7. "Metals Statistics," American Metal Market, Fairchild
Publications, Inc., New York, 1974.
8. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1975.
9. The Wall Street Journal, 16 September 1981, p. 46.
10. "A Study of Waste Generation, Treatment and Disposal in the
Metals Mining Industry," Midwest Research Institute, October
1976.
11. "Metal-Nonmetal Mine File Reference," Mining and Engineering
Safety Administration, Washington, 9 March 1977.
12. 1979/E/MJ International Directory o_f_ Mining and Mineral
Processing Operations, Engineering and Mining Journal, New York,
1979.
13. Engineering and Mining Journal, McGraw-Hill, February 1980,
Vol. 181, No. 2, p. 25.
14. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1974.
533
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15. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 556, 1963.
16. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 650, 1970.
17. "Mineral Facts and Problems," Bureau of Mines, U.S.
Department of the Interior, Washington, Bulletin 667, 1975.
18. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1977.
19. Personal communications from J. Viellenave, Earth Sciences,
Inc., to M. A. Wilkinson, Calspan Corporation, October 1976.
20. Minerals Yearbook, Bureau of Mines, U.S. Department of the
Interior, Washington, 1973.
21. Personal communication from Linda Carrico, commodity
specialist U.S. Department of the Interior, to D. M. Harty,
Frontier Technical Associates, April 1981.
22. Gordon, E., "UraniumRising Prices Continue to Spur
Development," Engineering and Mining Journal, March 1977.
23. White, G., Jr., "Uranium: Prices Steady at High Level in
"77," Engineering and Mining Journal, March 1978.
24. Merrit, R. C., "Extractive Metallurgy of Uranium," Colorado
School of Mines Research Institute, Inc., Colorado, 1971.
25. Davis, J. F., "U.S. Uranium Industry Continues Active
Development Despite Nuclear Uncertainties," Engineering and
Mining Journal, August 1977.
26. U.S. Nuclear Regulatory Commission, "Draft Generic Environ-
mental Impact Statement on Uranium Milling," Report No. NUREG-
0511, April 1979.
27. Personal communication from Scott F. Sibley, commodity
specialist U.S. Department of the Interior, to D. M. Harty,
Frontier Technical Associates, May 1981.
SECTION V
1. Bainbridge, Kent, "Evaluation of Wastewater Treatment
Practices Employed at Alaskan Gold Placer Mining Operations,"
Calspan Report No. 6332-M-2, Prepared for U.S. Environmental
Protection Agency - Effluent Guidelines Division, Washington,
D.C., 17 July 1979.
534
-------
2. Harty, David M. and P. Michael Terlecky, "Titanium Sand
Dredging Wastewater Treatment Practices," Frontier Technical
Associates, Inc., Buffalo, New York, Report No. 1804-1, Prepared
for U.S. Environmental Protection Agency - Effluent Guidelines
Division, Washington, D.C., Revised 20 October 1980.
3. Hayes, B. J., R. M. Mann, and J. I. Steinmetz, "Cyanide
Methods Evaluation: A Modified Method for the Analysis of Total
Cyanide in Ore Processing Effluents" Radian Corporation DCN No.
80-210-002-14-09, Prepared for U.S. Environmental Protection
Agency - Effluent Guidelines Division, Washington, D.C., 11
August 1980.
SECTION VI
1. "Froth Flotation in 1975," Advance Summary of Mineral
Industry Surveys, Bureau of Mines, U.S. Department of the
Interior, Washington, 1976.
2. Hawley, J. R., "The Use, Characteristics and Toxicity of
Mine-Mill Reage/its in the Province of Ontario," Ontario (Canada)
Ministry of the Environment, Ottawa, 1972.
3. Mining Chemicals Handbook, Mineral Dressing Notes No. 26,
American Cyanamid Company, Wayne, NJ, 1976.
4. Sharp, F. H., "Lead-Zinc-Copper Separation and Current
Practices at the Magmont Mill," Engineering and Mining Journal,
July 1973.
5. Personal communication from Chemist and Concentrator General
Foreman, Lead/Zinc Mine/Mill 3103, to P. H. Werthman, Calspan
Corporation, 1978.
SECTION VII
1. "Condensed Chemical Dictionary," P. Hawley, Van Norstrand,
Reinhold, New York, New York, 1971.
2. Rawlings, G. D., and M. Samfield, Environmental Science and
Technology, Vol. 13, No. 2, February 1974.
3. Development Document for BAT Effluent Limitations Guidelines
and New Source Performance Standards for the Textiles Point
Source Category, U.S. EPA, EGD, 1979.
535
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4. "Sampling and Analysis Procedures for Screening of Industrial
Effluents for Priority Pollutants," U.S. Environmental
Protection Agency, Environmental Monitoring and Support
Laboratory, Cincinnati, Ohio, March 1977, Revised April 1977.
5. "Seminary for Analytical Methods for Priority Pollutants,"
U.S. Environmental Protection Agency, Office of Water Programs,
Savannah, Georgia, 23- 24 May 1978.
6. "Antimony Removal Technology for Mining Industry Waste--
waters," Draft Report Prepared for U.S. Environmental Protection
Agency by Hittman Associates, Columbia, MD., under Contract 68-
03-1566, HIT-C185/200-78-739D, July 1978.
7. Calspan Corporation, "Heavy Metal Pollution from Spillage at
Ore Smelters and Mills," Calspan Report No. ND-5187 M-l (Rev),
EPA Contract No. 68-01-0726, Prepared for U.S. Environmental
Protection Agency, Office of Research and Development, Industrial
Environmental Research Laboratory, Cincinnati, Ohio, 15 July
1977.
8. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Michigan, 1977.
9. "Development Document for Effluent Limitations Guidelines and
New Source Performance Standards for the Ore Mining and Dressing
Point Source Category," U.S. Environmental Protection Agency, EPA
440/178/061-e, PB-286 521, July 1978.
10. Terlecky, P. M., editor, "Draft Development Document for the
Miscellaneous Nonferrous Metals Segment of the Nonferrous Metals
Point Source Category," EPA-440/1-76-067, March 1977.
SECTION VIII
1. Mining Chemicals Handbook, Mineral Dressing Notes No. 26,
American Cyanamid Company, Wayne, NJ, 1976.
2. Personal communication from Chemist and Concentrator General
Foreman, Lead/Zinc Mine/Mill 3103, to P. H. Werthman, Calspan
Corporation, 1978.
3. Personal communication from Water Quality Project Engineer,
Copper Mine/Mill 2122, to K. L. Bainbridge, Calspan Corporation,
1978.
4. Personal communication from Mill Superintendent, Lead/Zinc
Mine/Mill 3101, to K. L. Bainbridge, Calspan Corporation, 1978.
536
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5. Miller, J., "Pyrite Depression by Reduction of Solution
Oxidation Potential," Department of Mineral Engineering,
University of Utah, Salt Lake City, 1970.
6. "Alternative for Sodium Cyanide for Flotation Control
(T2008)," Report Prepared for U.S. Environmental Protection
Agency, Cincinnati, Ohio, BattelleColumbus (Ohio) Laboratories,
1979.
7. "Treatability of and Alternatives for Sodium Cyanide for
Flotation Control," Report for U.S. Environmental Protection
Agency, Cincinnati, Ohio, by Battelle - Columbus (Ohio)
Laboratories, 31 January 1980.
8. Gulp, R. L. and G. L. Gulp, Advanced Wastewater Treatment,
Van Nostrand Reinhold Book, Co., New York, 1971.
9. "Process Design Manual for Suspended Solids Removal," U.S.
Environmental Protection Agency, Washington, EPA 625/1-75-003a,
January 1975.
10. Bernardin, F. E., "Cyanide Detoxification Using Adsorption
and Catalytic Oxidation on Granular Activated Carbon," Journal of
Water Pollution Control Federation, Vol. 45, No. 2, February
1973.
11. Bucksteeg, W. and H. Thiele, "Method for the Detoxification
of Wastewater Containing Cyanide," German Patent 1,140,963,
January 1969.
12. Kuhn, R., "Process for Detoxification of Cyanide Containing
Aqueous Solutions," U.S. Patent 3,586,623, June 1971.
13. Willis, G. M. and J. T. Woodcock, "Chemistry of Cyanidation
II: Complex Cyanides of Zinc and Copper," Proc. Austral. Inst.
Min. Metall., No. 158-159, 1950, 00. 465-488.
14. Woodcock, J. T. and M. H. Jones, "Oxygen Concentrations,
Redox Potentials, Xanthate Residuals, and Other Parameters in
Flotation Plant Pulps," Proceedings of_ the Ninth Commonwealth
Mining and Metallurgial Congress, The Institute of Mining and
Metallurgy, London (England), 1969.
15. Letter from Vice President, Metallurgy, Copper Mine/Mill
2138 and Lead/Zinc Mine/Mills 3120 and 3121, to R.B. Schaffer,
Director, Effluent Guidelines Division (WH-552), U.S.
Environmental Protection Agency, 9 May 1977.
16. Written comments from Owner, Copper Mine/Mills 2101, 2104,
2116, and 2122 and Lead/Zinc Mine/Mill 3142, to U.S.
537
-------
Environmental Protection Agency, Washington, relative to interim
final effluent limitations for ore mining and dressing industry,
7 January 1976.
17. Eccles, A. G., "Cyanide Destruction at Western Mines Myra
Falls Operation," Paper presented at CMP Annual Meeting, 25
January 1977.
18. Eiring, L. V., Uchenye Zapiski Universities, Erevan, No. 2,
1967.
19. Eiring, L. V., "Kinetics and Mechanism of Ozone Oxidation of
Cyanide-Containing Wastewater," Soviet Journal of_ Non-Ferrous
Metals, Vol. 55, No. 106, 1969, pp. 81-83.
20. Chamberlin, N. S. and H. B. Snyder, Jr., "Treatment of
Cyanide and Chromium Wastes," Proceedings of Regional Conference
o_n_ Industrial Health, Houston, Texas, 27-29 September 1951.
21. Chamberlin, N. S. and H. B. Snyder, Jr., "Technology of
Treating Plating Wastes," Paper presented at Tenth Industrial
Waste Conference, Purdue University, West Lafayette, Ind., 9-11
May 1955.
22. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Mich., 1977, Chapter 9,
"Treatment Technology for Cyanide."
23. Goldstein, M., "Economics of Treating Cyanide Wastes,"
Pollution Engineering, March 1976.
24. Bird, A. 0., "The Destruction and Detoxification of Cyanide
Wastes," Chemical Engineer, University of Birmingham (England),
1976, pp. 12-21.
25. Bollyky, L. J., "Ozone Treatment of Cyanide Plating Wastes,"
Paper presented at First International Symposium on Ozone Water
and Wastewater Treatment, 1973.
26. Personal communication from R. Mankes, Telecommunications
Industries, Inc., to R. C. Lockemer, Calspan Corporation,
November 1977.
27. Fiedman, I. D. et al., "Removal of Toxic Cyanides from
Wastewaters of Gold Extracting Mills," Sb. Mosk. Inst. Stali
Splavov (USSR), No. 53, 1969, pp. 106-116 and Chemical Abstracts,
Vol. 72, No. 12, 1970, pp. 272-273.
538
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28. Boriesnko, A. P., "Detoxification of Wastewaters from the
Zolotushinskii Beneficiation Mills," Tsvet. Metal. (USSR), Vol.
42, No. 6, 1969, pp. 16-18.
29. Eiring, L. V., "Processing of Wastewaters from Gold
Extraction Plants," Vodosnabzh. i_ Sanit. Tekln (USSR), (2), 4-5,
1965.
30. Garrison, R. L., C. E. Mauk, and W. Prengle, Jr., "Cyanide
Disposal by Ozone Oxidation," Air Force Weapons Laboratory,
Kirtland Air Force Base, NM, AFWL-TR-73-212, February 1974.
31. Lanouette, K. H., "Treatment of Phenolic Wastes," Chemical
Engineering, Deskbook Issue, 17 October 1977.
32. Patterson, J. W., Wastewater Treatment Technology, Ann Arbor
Science Publishers, Inc., Ann Arbor, Mich., 1975.
33. Rosfjord, R. E., R. B. Trattner, and P. N. Cheremisinoff,
"Phenols: A Water Pollution Control Assessment," Water and
Sewage Works, March 1976.
34. Hodgson, E. W., Jr., and P. M. Terlecky, Jr. (Ed.), "Adden-
dum to Development Document for Effluent Limitations Guidelines
and New Source Performance Standards for Major Inorganic Products
Segment of Inorganic Chemicals Manufacturing Point Source
Category," Prepared for Effluent Guidelines Division, U.S.
Environmental Protection Agency, Washington, by Calspan
Corporation, Buffalo, NY, ND-5782-M-72, June 1978.
35. Larsen, H. P., "Chemical Treatment of Metal-Bearing Mine
Drainage," Journal of Water Pollution Control Federation, Vol.
45, No. 8, 1973, pp. 1682-1695.
36. Shimoiizuka, J., "Recovery of Xanthates from Cadmium
Xanthate," Nippon Kogyokaishi, 88, 1972, pp. 539-543.
37. Rosehard, R. and J. Lee, "Effective Methods of Arsenic
Removal from Gold Mine Wastes," Canadian Mining Journal, June
1972, pp. 53-57.
38. Curry, N. A., "Philosophy and Methodology of Metallic Waste
Treatment," Paper presented at 27th Industrial Waste Conference,
Purdue University, West Lafayette, Ind., 1972.
39. Larsen, H. P. and L. W. Ross, "Two Stage Process Chemically
Treats Mine Drainage to Remove Dissolved Metals," Engineering and
Mining Journal, February 1976.
40. Terlecky, P. M., Jr., M. A. Bronstein, and D. W. Goupil,
"Asbestos in the Ore and Dressing Identification Analysis,
Treatment Technology, Health Aspects, and Field Sampling
Results," Prepared for Effluent Guidelines Division, U.S.
539
-------
Environmental Protection Agency, Washington, by Calspan
Corporation, Buffalo, NY, ND-5782-M-121, (1979).
41. Logsdon, G. S. and J. M. Symons, "Removal of Asbestiform
Fibers by Water Filtration, Journal o_f American Water Works
Association, September 1977, pp. 499-506.
42. "Direct Filtration of Lake Superior Water from Asbestiform
Fiber Removal," Prepared for U.S. Environmental Protection Agency
EPA-670/2-75-050, by Black and Veatch Consulting Engineers, 1975.
43. Robinson, J. H. et al., "Direct Filtration of Lake Superior
Water for Asbestiform - Solids Removal," Journal of_ American
Water Works' Association, October 1976, pp. 531-539.
44. Lawrence, J. and H. W. Zimmerman, "Asbestos in Water: Mining
and Processing Effluent Treatment," Journal of_ Water Pollution
Control Federation, January 1977, pp. 156-160.
45. Patton, J. L., "Unusual Water Treatment Plant Licks Asbestos
Fiber Problem," Water and Wastes Engineering, 50, November 1977,
pp. 41-44. 46. Yourt, R. G., "Radiological Control of Uranium
Mine and Mill Wastes," Proceedings of_ the 13th Conference on_
Industrial Waste, Ontario (Canada), 1966, pp. 107-120.
47. Beverly, R. G., "Unique Disposal Methods are Required for
Uranium Mill Waste," Mining Engineering (Transaction, AIME) 20,
1968, pp. 52-56.
48. Felman, M. H., "Removal of Radium from Acid Mill Effluents,"
WIN-125, 1961.
49. Ryan, R. K. and P. G. Alfredson, "Liquid Wastes from Mining
and Milling of Uranium Ores: A Laboratory Study of Treatment
Methods," ISBN 0-642-99752-4, AAEC/E394, Australian Heights,
October 1976.
50. "The Control of Radium and Thorium in the Uranium Milling
Industry," United States Atomic Energy Commission, WIN-112, 1960.
51. Arnold, W. D. and D. J. Grouse, "Radium Removal from Uranium
Mill Effluents with Inorganic Ion Exchangers," Ind. Egn. Chem.
Process Des. Dev., Vol. 4, No. 3, 1965, pp. 333-337.
52. Clark, J. W., W. Viessman, Jr., M. J. Hammer, Water Supply
and Pollution Control, Third Edition, Harper and Row Publishers,
New York, N. Y., 1977.
540
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53. Metcalf and Eddy, Wastewater Engineering: Treatment,
Disposal, Reuse, Second Edition, McGraw-Hill Book Company, New
York, N. Y., 1979.
54. "Mine Wastewater Pilot Treatment ProjectMetal Removal
Phase," Montreal (Quebec, Canada), Engineering Company, May 1975.
55. Campbell, H. and B. P. LeClair, "Dewatering Base Metal Mine
Drainage Sludge," Proceedings of 1Oth Canadian Symposium on Water
Pollution Research, 1975.
56. Huck, P. M. and B. P. LeClair, "Operational Experience with
a Base Metal Mine Drainage Pilot Plant," EPA 4-WP-74-8, September
1974; Also presented as paper at 29th Industrial Waste
Conference, Purdue University, West Lafayette, Ind., May 1974.
57. Huck, P. M. and B. P. LeClair, "Treatment of Base Metal Mine
Drainage," Paper presented at 30th Industrial Waste Conference,
Purdue University, West Lafayette, Ind., 1975.
58. Huck, P. M. and B. P. LeClair, "Polymer Selection and Dosage
Determination Methodology for Acid Mine Drainage and Tailings
Pond Overflows," Paper presented at Symposium on Flocculation and
Stabilization of Solids in Aqueous and Non-Aqueous Media, Toronto
(Ontario, Canada), November 1974.
59. "Pilot-Scale Wastewater Treatability Studies Conducted at
Various Facilities in the Ore Mining and Milling Industry,"
Prepared for Effluent Guidelines Division, U.S. Environmental
Protection Agency, Washington, by Calspan Corporation, Buffalo,
N. Y., 6332-M-2, in publication (1979).
60. Dean, J. G., F. L. Bosque and K. H. Lanouette, "Removing
Heavy Metals from Wastewater," Env. Sci. and Tech., Vol. 6, No.
6, 19721, pp. 518-522.
61. Nilsson, R., "Removal of Metals by Chemical Treatment of
Municipal Wastewater," Water Research, Vol. 5, 1971, pp. 51-60.
62. Lanouette, K. H. and E. G. Paulson, "Treatment of Heavy
Metals in Wastewater," Pollution Engineering, October 1976, pp.
55-57.
63. Hem, J. D., "Reactions of Metal Ions at Surfaces of Hydrous
Iron Oxide," Geochimica et Cosmochimica Acta, Vol. 41, 1977, pp.
527-538.
64. Michalovic, J. G., J. G. Fisher and D. H. Bock, "Suggested
Method for Vanadate Removal from Mill Effluent," J_._ Environ.
Sci. Health, Vol. A12, Nos. 1 and 2, 1977, pp. 21-27.
541
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65. Kunz, R. G., J. F. Giannelli and H. D. Stensel, "Vanadium
Removal from Industrial Wastewaters," Jour. Water Poll. Control
Fed., Vol. 48, No. 6, 1976, pp. 762-770.
66. LeGendre, G. R. and D. D. Runnells, "Removal of Dissolved
Molybdenum from Wastewaters by Precipitates of Ferric Iron," Env.
Sci. and Tech., Vol. 9, No. 8, 1975, pp. 755-759.
67. Gott, R. D., "A Discussion and Review of the Development of-
Wastewater Treatment at the Climax Mine, Climax, Colorado," Paper
presented at 1977 American Mining Congress, San Francisco,
California.
68. Federal Register, Vol. 43, No. 133, 11 July 1978, p. 29773.
69. Bainbridge, K., "Evaluation of Wastewater Treatment
Practices Employed at Alaskan Gold Placer Mining Operations,"
Calspan Report No. 6332-M-2 prepared for U.S. Environmental
Protection Agency Effluent Guidelines Division, Washington, D.C.
70. Jackson, B. et al., "Environmental Study on Uranium Mills
Part I," Draft Final Report Prepared for Effluent Guidelines
Division, U.S. Environmental Protection Agency, Washington, by
TRW, Inc., under Contract 68-03-2560, December 1978.
71. Letter from D. J. Kuhn, Secured Landfill Contractors, Inc.,
Tonawanda, NY, to P. M. Terlecky, Jr., Frontier Technical
Associates, Inc., Buffalo, NY, 12 February 1979.
SECTION IX
1. Engineering News Record, Volume 203, Number 24, 13 December
1979.
2. Robert Snow Means Company, Building Construction Cost Data,
Robert Snow Means Co., Duxbury, Mass., 1980.
3. Dodge Building Cost Services, Construction Systems Costs,
McGraw Hill, 1979.
4. Harty, David M. and P. Michael Terlecky, "Characterization of
Wastewater and Solid Wastes Generated in Selected Ore Mining
Subcategories (Sb, Hg, Al, V, W, Ni, Ti)," Contract No. 68-01-
5163, Frontier Technical Associates Report No. 2804-1, 24 August
1981, Prepared for U.S. Environmental Protection Agency,
Washington, D.C.
5. PEDCo Environmental, Inc., "Evaluation of Best Management
Practices for Mining Solid Waste Storage, Disposal and Treatment
Presurvey Study," February 1981, Contract No. 68-03-2900,
542
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Prepared for U.S. Environmental Protection Agency, Office of
Research and Development, Cincinnati, Ohio.
6. Environmental Protection Agency, "Mining Industry Solid
Waste, an Interim Report," Office of Solid Waste, February 1981.
SECTION X
1. "Development Document for BPT Effluent Limitations Guidelines
and New Source Performance Standards for the Ore Mining and
Dressing Industry," U.S. Environmental Protection Agency, 1975.
2. "Process Design Manual for Suspended Solids Removal," U.S.
Environmental Protection Agency, Washington, EPA 625/1-75-003a,
January 1975.
3. Personal communication from R. Mankes, Telecommunications
Industries, Inc., to R. C. Lockemer, Calspan Corporation,
November 1977.
4. Bainbridge, K., "Evaluation of Wastewater Treatment Practices
Employed at Alaskan Gold Placer Mining Operations," Calspan
Report No. 6332-M-2 prepared for U.S. Environmental Protection
Agency Effluent Guidelines Division, Washington, D.C.
5. "Dana's Manual of Mineralogy," 17th Edition, 1961, John Wiley
and Sons, New York, New York.
6. Terlecky, P. M., M. A. Bronstein, and D. W. Goupil, Asbestos
in the Ore Mining and Dressing Industry: Identification,
Analysis, Treatment Technology, Health Aspects, and Field
Sampling Results, U.S. Environmental Protection Agency, Contract
Number 68-01-3281, 20 July 1979, page 7.
543
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SECTION XVI
GLOSSARY
absorption: The process by which a liquid is drawn into and
tends to fill permeable pores in a porous solid body; also the
increase in weight of a porous solid body resulting from the
penetration of liquid into its permeable pores.
acid copper: Copper electrode deposited from an acid solution of
a copper salt, usually copper sulfate.
acid cure: In uranium extraction, sulfation of moist ore before
leach.
acid leach: (a) Metallurgical process for dissolution of values
by means of acid solution (used on the sandstone ores of low lime
content); (b) In the copper industry, a technology employed to
recover copper from low grade ores and mine dump materials when
oxide (or mixed oxide-sulfide, or low grade sulfide)
mineralization is present, by dissolving the copper minerals with
either sulfuric acid or sulfuric acid containing ferric iron.
Four methods of leaching are employed: dump, heap, in-situ, and
vat (see appropriate definitions).
acid mine water: (a) Mine water which contains free sulfuric
acid, mainly due to the weathering of iron pyrites; (b) Where
sulfide minerals break down under the chemical influence of
oxygen and water, the mine water becomes acidic and can corrode
ironwork.
activator, activating agent: A substance which when added to a
mineral pulp promotes flotation in the presence of a collecting
agent. It may be used to increase the floatability of a mineral
in a froth, or to reflect a depressed (sunk mineral).
adit: (a) A horizontal or nearly horizontal passage driven from
the surface for the working or dewatering of a mine; (b) A
passage driven into a mine from the side of a hill.
adsorption: The adherence of dissolved, colloidal, or finely
divided solids on the surface of solids with which they are
brought into contact.
aeroflocs: Synthetic water-soluble polymers used as flocculating
agents.
all sliming: (a) Crushing all the ore in a mill to so fine a
state that only a small percentage will fail to pass through a
200-mesh screen; (b) Term used for treatment of gold ore which is
ground to a size sufficiently fine for agitation as a cyanide
pulp, as opposed to division into coarse sands for static
leaching and fine slimes for agitation.
545
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alluminothermic process: The reduction of oxides in an
exothermic reaction with finely divided aluminum.
alluvial deposit; placer deposit: Earth, sand, gravel or other
rock or mineral materials transported by and laid down by flowing
water. Alluvial deposits generally take the form of (1) surface
deposits; (2) river deposits; (3) deep leads; and (4) shore
deposits.
alunite: A basic potassium aluminum sulfate, KA13.(OH)6_(S04) 2.
Closely resembles kaolinite and occurs in similar locations.
amalgamation: The process by which mercury is alloyed with some
other metal to produce amalgam. It was used extensively at one
time for the extraction of gold and silver from pulverized ores,
now is largely superseded by the cyanide process.
AN-FO - Ammonium nitrate: Fuel oil blasting agents.
asbestos minerals: Certain minerals which have a fibrous
structure, are heat resistant, chemically inert and possessing
high electrical insulating qualities. The two main groups are
serpentine and amphiboles. Chrysotile (fibrous serpentine, 3MgO
. 2Si02_ . 2H20) is the principal commercial variety. Other
commercial varieties are amosite, crocidolite, actinolite,
anthophyllite, and tremolite.
azurite: A blue carbonate of copper, Cu3_(C03_) 2 (OH) 2_,
crystallizing in the monoclinic system. Found as an alteration
product of chalcopyrite and other sulfide ores of copper in the
upper oxidized zones of mineral veins.
bastnasite; bastnaesite: A greasy, wax-yellow to reddish-brown
weakly radioactive mineral, (Ce,La) (C03_)F, most commonly found
in contact zones, less often in pegmatites.
bauxite: (a) A rock composed of aluminum hydroxides, essentially
A1203_ . 2H20. The principal ore of aluminum; also used
collectively for lateritic aluminous ores. (b) Composed of
aluminum hydroxides and impurities in the form of free silica,
clay, silt, and iron hydroxides. The primary minerals found in
such deposits are boehmite, gibbsite, and diaspore.
Bayer Process: Process in which impure aluminum in bauxite is
dissolved in a hot, strong, alkalai solution (normally NaOH) to
form sodium aluminate. Upon dilution and cooling, the solution
hydrolyzes and forms a precipitate of aluminum hydroxide.
bed: The smallest division of a stratified series and marked by
a more or less well-defined divisional plane from the materials
above and below.
546
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beneficiation: (a) The dressing or processing of ores for the
purpose of (1) regulating the size of a desired product, (2)
removing unwanted constituents, and (3) improving the quality,
purity, assay grade of a desired product; (b) Concentration or
other preparation of ore for smelting by drying, flotation, or
magnetic separation.
Best Available Technology Economically Achievable (BAT): The
level of technology applicable to effluent limitations to be
achieved by 1 July 1983, for industrial discharges to surface
waters as defined by Section 301(b)(l)(A) of the Act.
Best Practicable Control Technology Currently Available (BPT):
The level of technology applicable to effluent limitations to be'
achieved by 1 July 1977, for industrial discharges to surface
waters as defined by Section 301(b)(l)(A) of the Act.
byproduct: A secondary or additional product.
carbon absorption: A process utilizing the efficient absorption
characteristics of activated carbon to remove both dissolved and
suspended substances.
carnotite: A bright yellow uranium mineral, K2_(U02_)^( V04 )2_ .
3H20.
cationic collectors: In flotation, amines and related organic
compounds capable of producing positively charged hydrocarbon-
bearing ions for the purpose of floating miscellaneous minerals,
especially silicates.
cationic reagents: In flotation, surface active substances which
have the active constituent in the positive ion. Used to
flocculate and to collect minerals that are not flocculated by
the reagents, such as oleic acid or soaps, in which the surface-
active ingredient is the negative ion.
cement copper: Copper precipitated by iron from copper sulfate
solutions.
cerium metals: Any of a group of rare-earth metals separable as
a group from other metals occurring with them and in addition to
cerium includes lanthanum, praseodymium, neodymium, promethium,
samarium and sometimes europium.
cerium minerals: Rare earths; the important one is monazite.
chalcocite: Copper sulfide, Cu2S.
chalcopyrite: A sulfide of copper and iron, CuFeS2_.
chert: Cryptocrystalline silica, distinguished from flint by
flat fracture, as opposed to conchoidal fracture.
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chromite: Chrome iron ore, FeCr20i.
chrysocolla: Hydrated copper silicate, CuSiOS^ . 2H20.
chrysotile: A metamorphic mineral, an asbestos, the fibrous
variety of serpentine. A silicate of magnesium, with silica
tetrahedra arranged in sheets.
cinnabar: Mercury sulfide, HgS.
claim: The portion of mining ground held under the Federal and
local laws by one claimant or association, by virtue of one
location and record. A claim is sometimes called a "location."
clarification: (a) The cleaning of dirty or turbid liquids by
the removal of suspended and colloidal matter; (b) The
concentration and removal of solids from circulating water in
order to reduce the suspended solids to a minimum; (c) In the
leaching process, usually from pregnant solution, e.g., gold-rich
cyanide prior to precipitation.
classifier: (a) A machine or device for separating the
constituents of a material according to relative sizes and
densities thus facilitating concentration and treatment.
Classifiers may be hydraulic or surface-current box classifiers.
Classifiers are also used to separate sand from slime, water from
sand, and water from slime; (b) The term classifier is used in
particular where an upward current of water is used to remove
fine particles from coarser material; (c) In mineral dressing,
the classifier is a device that takes the ball-mill discharge and
separates it into two portionsthe finished product which is
ground as fine as desired, and oversize material.
coagulation: The binding of individual particles to form floes
or agglomerates and thus increase their rate of settlement in
water or other liquid (see also flocculate).
coagulator: A soluble substance, such as lime, which when added
to a suspension of very fine solid particles in water causes
these particles to adhere in clusters which will settle easily.
Used to assist in reclaiming water used in flotation.
collector: A heteropolar compound containing a hydrogen-carbon
group and an ionizing group, chosen for the ability to adsorb
selectively in froth flotation processes and render the adsorbing
surface relatively hydrophobic. A promoter.
columbite; tantalite; niobite: A natural oxide of niobium
(columbium), tantalum, ferrous iron, and manganese, found in
granites and pegmatites, (Fe, Mn) Nb, Ta) 206^.
concentrate: (a) In mining, the product of concentration; (b) To
separate ore or metal from its containing rock or earth; (c) The
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enriched ore after removal of waste in a beneficiation mill, the
clean product recovered in froth flotation.
concentration: Separation and accumulation of economic minerals
from gangue.
concentrator: (a) A plant where ore is separated into values
(concentrates) and rejects (tails). An appliance in such a
plant, e.g., flotation cell, jig, electromagnet, shaking table.
Also called mill; (b) An apparatus in which, by the aid of water
or air and specific gravity, mechanical concentration of ores is
performed.
conditioners: Those substances added to the pulp to maintain the
proper pH to protect such salts as NaCN, which would decompose in
an acid circuit, etc. Na2C03^ and CaO are the most common
conditioners.
conditioning: Stage of froth-flotation process in which the
surfaces of the mineral species present in a pulp are treated
with appropriate chemicals to influence their reaction when the
pulp is aerated.
copper minerals: Those of the oxidized zone of copper deposits
(zone of oxidized enrichment) include azurite, chrysocolla,
copper metal, cuprite, and malachite. Those of the underlying
zone (that of secondary sulfide enrichment) include bornite,
chalcocite, chalcopyrite, covellite. The zone of primary
sulfides (relatively low in grade) includes the unaltered
minerals bornite and chalcopyrite.
crusher: A machine for crushing rock or other materials. Among
the various types of crushers are the ball-mill, gyratory
crusher, Hadsel mill, hammer mill, jaw crusher, rod mill, rolls,
stamp mill, and tube mill.
cuprite: A secondary copper mineral, Cu20.
cyanidation: A process of extracting gold and silver as cyanide
slimes from their ores by treatment with dilute solutions of
potassium cyanide and sodium cyanide.
cyanidation vat: A large tank, with a filter bottom, in which
sands are treated with sodium cyanide solution to dissolve out
gold.
cyclone: (a) The conical-shaped apparatus used in dust
collecting operations and fine grinding applications; (b) A
classifying (or concentrating) separator into which pulp is fed,
so as to take a circular path. Coarser and heavier fractions of
solids report at the apex of long cone while finer particles
overflow from central vortex.
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daughter: Decay product formed when another element undergoes
radioactive disintegration.
decant structure: Apparatus for removing clarified water from
the surface layers of tailings or settling ponds. Commonly used
structure include decant towers in which surface waters flow over
a gate (adjustable in height) and down the tower to a conduit
generally buried beneath the tailings, decant weirs over which
water flows to a channel external to the tailings pond, and
floating decant barges which pump surface water out of the pond.
dense-media separation: (a) Heavy media separation, or sink
float. Separation of heavy sinking from light floating mineral
particles in a fluid of intermediate density; (b) Separation of
relatively light (floats) and heavy ore particles (sinks), by
immersion in a bath of intermediate density.
Denver cell: A flotation cell of the subaeration type, in wide
use. Design modifications include receded-disk, conical-disk,
and multibladed impellers, low-pressure air attachments, and
special froth withdrawal arrangements.
Denver jig: Pulsion-suction diaphragm jig for fine material, in
which makeup (hydraulic) water is admitted through a rotary valve
adjustable as to portion of jigging cycle over which controlled
addition is made.
deposit: Mineral or ore deposit is used to designate a natural
occurrence of a useful mineral or an ore, in sufficient extent
and degree of concentration to invite exploitation.
depressing agent; depressor: In the froth floation process, a
substance which reacts with the particle surface to render it
less prone to stay in the froth, thus causing it to wet down as a
tailing product (contrary to activator).
detergents, synthetic: Materials which have a cleansing action
like soap but are not derived directly from fats and oils. Used
in ore flotation.
development work: Work undertaken to open up ore bodies as
distinguished from the work of actual ore extraction or
exploratory work.
dewater: To remove water from a mine usually by pumping,
drainage or evaporation.
differential flotation: Separating a complex ore into two or
more valuable minerals and gangue by flotation; also called
selective flotation. This type of flotation is made possible by
the use of suitable depressors and activators.
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discharge: Outflow from a pump, drill hole, piping system,
channel, weir or other discernible, confined or discrete
conveyance (see also point source).
dispersing agent: Reagent added to flotation circuits to prevent
flocculation, especially of objectionable colloidal slimes.
Sodium silicate is frequently added for this purpose.
dredge; dredging: A large floating contrivance for underwater
excavation of materials using either a chain of buckets, suction
pumps, or other devices to elevate and wash alluvial deposits and
gravel for gold, tin, platinum, heavy minerals, etc.
dressing: Originally referred to the picking, sorting, and
washing of ores preparatory to reduction. The term now includes
more elaborate processes of milling and concentration of ores.
drift mining: A term applied to working alluvial deposits by
underground methods of mining. The paystreak is reached through
an adit or a shallow shaft. Wheelbarrows or small cars may be
used for transporting the gravel to a sluice on the surface.
dump leaching: Term applied to dissolving and recovering
minerals from subore-grade materials from a mine dump. The dump
is irrigated with water, sometimes acidified, which percolates
into and through the dump, and runoff from the bottom of the dump
is collected, and a mineral in solution is recovered by chemical
reaction. Often used to extract copper from low grade, waste
material of mixed oxide and sulfide mineralization produced in
open pit mining.
effluent: The wastewater discharged from a point source to
navigable waters.
electrowinning: Recovery of a metal from an ore by means of
electrochemical processes, i.e., deposition of a metal on an
electrode by passing electric current through an electrolyte.
eluate: Solutions resulting from regeneration (elution) of ion
exchange resins.
eluent: A solution used to extract collected ions from an ion
exchange resin or solvent and return the resin to its active
state.
exploration: Location of the presence of economic deposits and
establishing their nature, shape, and grade and the investigation
may be divided into (1) preliminary, and (2) final.
extraction: (a) The process of mining and removal of ore from a
mine. (b) The separation of a metal or valuable mineral from an
ore or concentrate, (c) Used in relation to all processes that
are used in obtaining metals from their ores. Broadly, these
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processes involve the breaking down of the ore both mechanically
(crushing) and chemically (decomposition), and the separation of
the metal from the associated gangue.
ferruginous: Containing iron.
ferruginous chert: A sedimentary deposit consisting of
chalcedony or of fine-grained quartz and variable amounts of
hematite, magnetite, or limonite.
ferruginous deposit: A sedimentary rock containing enough iron
to justify exploitation as iron ore. The iron is present, in
different cases, in silicate, carbonate, or oxide form, occurring
as the minerals chamosite, thuringite, siderite, hematite,
limonite, etc.
flask: A unit of measurement for mercury; 76 pounds.
flocculant: An agent that induces or promotes flocculation or
produces floccules or other aggregate formation, especially in
clays and soils.
flocculate: To cause to aggregate or to coalesce into small
lumps or loose clusters, e.g., the calcium ion tends to
flocculate clays.
flocculating agent; flocculant: A substance which produces
flocculation.
flotation: The method of mineral separation in which a froth
created in water by a variety of reagents floats some finely
crushed minerals, whereas other minerals sink.
flotation agent: A substance or chemical which alters the
surface tension of water or which makes it froth easily. The
reagents used in the flotation process include pH regulators,
slime dispersants, resurfacing agents, wetting agents,
conditioning agents, collectors, and frothers.
friable: Easy to break, or crumbling naturally.
froth, foam: In the flotation process, a collection of bubbles
resulting from agitation, the bubbles being the agency for
raising (floating) the particles of ore to the surface of the
cell.
frother(s): Substances used in flotation processes to make air
bubbles sufficiently permanent principally by reducing surface
tension. Common frothers are pine oil, creyslic acid, and amyl
alcohol.
gangue: Undesirable minerals associated with ore.
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glory hole: A funnel-shaped excavation, the bottom of which is
connected to a raise driven from an underground haulage level or
is connected through a horizontal tunnel (drift) by which ore may
also be conveyed.
gravity separation: Treatment of mineral particles which
exploits differences between their specific gravities. Their
sizes and shapes also play a minor part in separation. Performed
by means of jigs, classifiers, hydrocyclones, dense media,
shaking tables, Humphreys spirals, sluices, vanners and briddles.
grinding: (a) Size reduction into relatively fine particles, (b)
Arbitrarily divided into dry grinding performed on mineral
containing only moisture as mined, and wet grinding, usually done
in rod, ball or pebble mills with added water.
heap leaching: A process used in the recovery of copper from
weathered ore and material from mine dumps. The liquor seeping
through the beds is led to tanks, where it is treated with scrap
iron to precipitate the copper from solution. This process can
also be applied to the sodium sulfide leaching of mercury ores.
heavy-media separation: See dense-media separation.
hematite: One of the most common ores of iron, Fe203_, which when
pure contains about 70% metallic iron and 30% oxygen. Most of
the iron produced in North America comes from the iron ranges of
the Lake Superior District, especially the Mesabi Range,
Minnesota. The hydrated variety of this ore is called limonite.
Huntington-Heberlein Process: A sink-float process employing a
galena medium and utilizing froth flotation as the means of
medium recovery.
hydraulic mining: (a) Mining by washing sand and soil away with
water which leaves the desired mineral, (b) The process by which
a bank of gold-bearing earth and rock is excavated by a jet of
water, discharged through the converging nozzle of a pipe under
great pressure. The debris is carried away with the same water
and discharged on lower levels into watercourses below.
hydrolysate; hydrolyzate: A sediment consisting partly of
chemically undecomposed, finely ground rock powder and partly of
insoluble matter derived from hydrolytic decomposition during
weathering.
hydrometallurgy: The treatment of ores, concentrates, and other
metal-bearing materials by wet processes, usually involving the
solution of some component, and its subsequent recovery from the
solution.
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ilmenite: An iron-black mineral, FeO . TiO^. Resembles
magnetite in appearance but is readily distinguished by feeble
magnetic character.
in-situ leach: Leaching of broken ore in the subsurface as it
occurs, usually in abandoned underground mines which previously
employed block-caving mining methods.
ion(ic) exchange: The replacement of ions on the surface, or
sometimes within the lattice, of materials such as clay.
iron formation: Sedimentary, low grade, iron ore bodies
consisting mainly of chert and fine-grained quartz and ferric
oxide segregated in bands or sheets irregularly mingled (see also
taconite).
jaw crusher: A primary crusher designed to reduce large rocks or
ores to sizes capable of being handled by any of the secondary
crushers.
jig: A machine in which the feed is stratified in water by means
of a pulsating motion and from which the stratified products are
separately removed, the pulsating motion being usually obtained
by alternate upward and downward currents of the water.
jigging: (a) The separation of the heavy fractions of an ore
from the light fractions by means of a jig. (b) Up and down
motion of a mass of particles in water by means -of pulsion.
laterite: Red residual soil developed in humid, tropical, and
subtropical regions of good drainage. It is leached of silica
and contains concentrations particularly of iron oxides and
hydroxides and aluminum hydroxides. It may be an ore of iron,
aluminum, manganese, or nickel.
launder: (a) A trough, channel, or gutter usually of wood, by
which water is conveyed; specifically in mining, a chute or
trough for conveying powdered ore, or for carrying water to or
from the crushing apparatus, (b) A flume.
leaching: (a) The removal in solution of the more soluble
minerals by percolating waters, (b) Extracting a soluble metallic
compound from an ore by selectively dissolving it in a suitable
solvent, such as water, sulfuric acid, hydrochloric acid, etc.
The solvent is usually recovered by precipitation of the metal or
by other methods.
leach ion-exchange flotation process: A mixed method of
extraction developed for treatment of copper ores not amenable to
direct flotation. The metal is dissolved by leaching, for
example, with sulfuric acid, in the presence of an ion exchange
resin. The resin recaptures the dissolved metal and is then
recovered in a mineralized froth by the flotation process.
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leach precipitation float: A mixed method of chemical reaction
plus flotation developed for such copper ores as chrysocolla and
the oxidized minerals. The value is dissolved by leaching with
acid, and the copper is reprecipitated on finely divided
particles of iron, which are then recovered by flotation,
yielding an impure concentrate in which metallic copper
predominates.
lead minerals: The most important industrial one is galena
(PbS), which is usually argentiferous. In the upper parts of
deposits the mineral may be altered by oxidation to cerussite
(PbC03^) or anglesite (PbSCU). Usually galena occurs in intimate
association with sphalerite (ZnS).
leucoxene: A brown, green, or black variety of sphene or
titanite, CaTiSiO, occurring as monoclinic crystals. An earthy
alteration product consisting in most instances of rutile; used
in the production of titanium tetrachloride.
lime: Quicklime (calcium oxide) obtained by calcining limestone
or other forms of calcium carbonate. Loosely used for hydrated
lime (calcium hydroxide) and incorrectly used for pulverized or
ground calcium carbonate in agricultural lime and for calcium in
such expressions as carbonate of lime, chloride of lime, and lime
feldspar.
lime slurry: A form of calcium hydroxide in aqueous suspension
that contains considerable free water.
limonite: Hydrous ferric oxide FeO(OH) . nH20. An important ore
of iron, occurring in stalactitic, mammillary, or earthy forms of
a dark brown color, and as a yellowish-brown powder. The chief
constituent of bog iron ore.
liquid-liquid extraction, solvent extraction: A process in which
one or more components are removed from a liquid measure by
intimate contact with a second liquid, which is itself nearly
insoluble in the first liquid and dissolves the impurities and
not the substance that is to be purified.
lode: A tabular deposit of valuable mineral between definite
boundaries. Lode, as used by miners, is nearly synonymous with
the term vein as employed by geologists.
magnetic separation: The separation of magnetic materials from
nonmagnetic materials using a magnet. An important process in
the beneficiation of iron ores in which the magnetic mineral is
separated from nonmagnetic material, e.g., magnetite from other
minerals, roasted pyrite from sphalerite.
magnetic separator: A device used to separate magnetic from less
magnetic or nonmagnetic materials. The crushed material is
conveyed on a belt past a magnet.
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magnetite, magnetic iron ore: Natural black oxide of iron,
Fe3_CH. As black sand, magnetite occurs in placer deposits, and
also as lenticular bands. Magnetite is used widely as a
suspension solid in dense-medium washing of coal and ores.
malachite: A green, basic cupric carbonate, Cu2(OH) 2C03_,
crystallizing in the monoclinic system. It is a common ore of
copper and occurs typically in the oxidation zone of copper
deposits.
manganese minerals: Those in principal production are
pyrolusite, some psilomelane, and wad (impure mixture of
manganese and other oxides).
manganese nodules: The concretions, primarily of manganese
salts, covering extensive areas of the ocean floor. They have a
layer configuration and may prove to be an important source of
manganese.
manganese ore: A term used by the Bureau of Mines for ore
containing 35 percent or more manganese and may include
concentrate, nodules, or synthetic ore.
manganiferous iron ore: A term used by the Bureau of Mines for
ores containing 5 to 10 percent manganese.
manganiferous ore: A term used by the Bureau of Mines for any
ore of importance for its manganese content containing less than
35 percent manganese but not less than 5 percent manganese.
mercury minerals: The main source is cinnabar, HgS.
mill: (a) Reducing plant where ore is concentrated and/or metals
recovered, (b) Today the term has been broadened to cover the
whole mineral treatment plant in which crushing, wet grinding,
and further treatment of the ore is conducted. (c) In mineral
processing, one machine, or a group, used in comminution.
minable: (a) Capable of being mined, (b) Material that can be
mined under present day mining technology and economics.
mine: (a) An opening or excavation in the earth for the purpose
of excavating minerals, metal ores or other substances by
digging, (b) A word for the excavation of minerals by means of
pits, shafts, levels, tunnels, etc., as opposed to a quarry,
where the whole excavation is open. In general the existence of
a mine is determined by the mode in which the mineral is
obtained, and not by its chemical or geologic character. (c) An
excavation beneath the surface of the ground from which mineral
matter of value is extracted. Excavations for the extraction of
ore or other economic minerals not requiring work beneath the
surface are designated by a modifying word or phrase as: (1)
opencut mine - an excavation for removing minerals which is open
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to the weather; (2) steam shovel mine - an opencut mine in which
steam shovels or other power shovels are used for loading cars;
(3) strip mine - a stripping, an openpit mine in which the
overburden is removed from the exploited material before the
material is taken out; (4) placer mine - a deposit of sand,
gravel or talus from which some valuable mineral is extracted;
and (4) hydraulic mine - a placer mine worked by means of a
stream of water directed against a bank of sand, gravel, or
talus. Mines are commonly known by the mineral or metal
extracted, e.g., bauxite mines, copper mines, silver mines, etc.
(d) Loosely, the word mine is used to mean any place from which
minerals are extracted, or ground which it is hoped may be
mineral bearing, (e) The Federal and State courts have held that
the word mine, in statutes reserving mineral lands, included only
those containing valuable mineral deposits. Discovery of a mine:
In statutes relating to mines the word discovery is used: (1) In
the sense of uncovering or disclosing to view ore or mineral; (2)
of finding out or bringing to the knowledge the existence of ore,
or mineral, or other useful products which were unknown; and (3)
of exploration, that is, the more exact blocking out or
ascertainment of a deposit that has already been discovered. In
this sense it is practically synonymous with development, and has
been so used in the U.S. Revenue Act of 19 February 1919 (Sec.
214, subdiv. AID, and Sec. 234, subdiv. A9) in allowing
depletion of mines, oil and gas wells. Article 219 of Income and
War Excess Profits Tax Regulations No. 45, construes discovery of
a mine as: (1) The bona fide discovery of a commercially
valuable deposit of ore or mineral, of a value materially in
excess of the cost of discovery in natural exposure or by
drilling or other exploration conducted above or below the
ground; and (2) the development and proving of a mineral or ore
deposit which has been apparently worked out to be a mineable
deposit or ore, or mineral having a value in excess of the cost
of improving or development.
mine drainage: (a) Mine drainage usually implies gravity flow of
water to a point remote from mining operation, (b) The process of
removing surplus ground or surface water by artificial means.
mineral: An inorganic substance occurring in nature, though not
necessarily of inorganic origin, which has (1) a definite
chemical composition, or commonly a characteristic range of
chemical composition, and (2) distinctive physical properties, or
molecular structure. With few exceptions, such as opal
(amorphous) and mercury (liquid), minerals are crystalline
solids.
mineral processing; ore dressing; mineral dressing: The dry and
wet crushing and grinding of ore or other mineral-bearing
products for the purpose of raising concentrate grade; removal of
waste and unwanted or deleterious substances from an otherwise
useful product; separation into distinct species of mixed
minerals; chemical attack and dissolution of selected values.
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modifier(s): (a) In froth flotation, reagents used to control
alkalinity and to eliminate harmful effects of colloidal material
and soluble salts. (b) Chemicals which increase the specific
attraction between collector agents and particle surfaces, or
conversely which increase the wettability of those surfaces.
molybdenite: The most common ore of molybdenum, MoSz^
molybdenite concentrate: Commercial molybdenite ore after the
first processing operations. Contains about 90% MoS^ along with
quartz, feldspar, water, and processing oil.
monazite: A phosphate of the cerium metals and the principal ore
of the rare earths and thorium. Monoclinic. One of the chief
sources of thorium used in the manufacture of gas mantles. It is
a moderately to strongly radioactive mineral, (Ce, La, Y, Th)
P0£. It occurs widely disseminated as an accessory mineral in
granitic igneous rocks and gneissic metamorphic rocks. Detrital
sands in regions of such rocks may contain commercial quantities
of monazite. Thorium-free monazite is rare.
New Source Performance Standard (NSPS): Performance standards
for the industry and applicable new sources as defined by Section
306 of the Act.
niccolite: A copper-red arsenide of nickel which usually
contains a little iron, cobalt, and sulfur. It is one of the
chief ores of metallic nickel.
nickel minerals: The nickel-iron sulfide, pentlandite (Fe, Ni)
is the principal present economic source of nickel, and
garnierite (nickelmagnesium hydrosilicate) is next in economic
importance.
oleic acid: A mono-saturated fatty acid, CH3.(CH2.)_CH:CH(CH2)7
COOH. A common component of almost all naturally occurring fats
as well as tall oil. Most commercial oleic acid is derived from
animal tallow or natural vegetable oils.
open-pit mining, open cut mining: A form of operation designed
to extract minerals that lie near the surface. Waste, or
overburden, is first removed, and the mineral is broken and
loaded. Important chiefly in the mining of ores of iron and
copper.
ore: (a) A natural mineral compound of the elements of which one
at least is a metal. Applied more loosely to all metalliferous
rock, though it contains the metal in a free state, and
occasionally to the compounds of nonmetallic substances, such as
sulfur. (b) A mineral of sufficient value as to quality and
quantity which may be mined with profit.
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ore dressing: The cleaning of ore by essentially physical means
and the removal of certain valueless portion. Synonym for
concentration. The same as mineral dressing.
ore reserve: The term usually restricted to ore of which the
grade and tonnage have been established with reasonable assurance
by drilling and other means.
oxidized ores: The alteration of metalliferous minerals by
weathering and the action of surface waters, and the conversion
of the minerals into oxides, carbonates, or sulfates.
oxidized zone: That portion of an ore body near the surface,
which has been leached by percolating water carrying oxygen,
carbon dioxide or other gases.
pegmatite: An igneous rock of coarse grain size usually found as
a crosscutting structure in a larger igneous mass of finer grain
size.
pelletizing: A method in which finely divided material is rolled
in a drum or on an inclined disk, so that the particles cling
together and roll up into small, spherical pellets.
pH modifiers: Proper functioning of a cationic or anionic
flotation reagent is dependent on the close control of pH.
Modifying agents used are soda ash, sodium hydroxide, sodium
silicate, sodium phosphates, lime, sulfuric acid, and
hydrofluoric acid.
placer mine: (a) A deposit of sand, gravel, or talus from which
some valuable mineral is extracted, (b) To mine gold, platinum,
tin or other valuable minerals by washing the sand, gravel, etc.
placer mining: The extraction of heavy mineral from a placer
deposit by concentration in running water. It includes ground
sluicing, panning, shoveling gravel into a sluice, scraping by
power scraper, excavation by dragline or extraction by means of
various types of dredging activities.
platinum minerals: Platinum, ruthenium rhodium, palladium,
osmium, and iridium are members of a group characterized by high
specific gravity, unusual resistance to oxidizing and acidic
attack, and high melting point.
point source: Any discernible, confined and discrete conveyance,
including but not limited to any pipe, ditch, channel, tunnel,
conduit, well, discrete fissure, container, rolling stock,
concentrated animal feeding operation, or vessel or other
floating craft, from which pollutants are or may be discharged.
pregnant solution: A value bearing solution in a
hydrometallurgical operation.
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pregnant solvent: In solvent extraction, the value-bearing
solvent produced in the solvent extraction circuit.
promoter: A reagent used in froth-flotation process, usually
called the collector.
rare-earth deposits: Sources of cerium, terbium, yttrium, and
related elements of the rare-earth's group, as well as thorium.
raw mine drainage: Untreated or unprocessed water drained,
pumped or siphoned from a mine.
reagent: A chemical or solution used to produce a desired
chemical reaction; a substance used in assaying or in flotation.
reclamation: The procedures by which a disturbed area can be
reworked to make it productive, useful, or aesthetically
pleasing.
recovery: A general term to designate the valuable constituents
of an ore which are obtained by metallurgical treatment.
reduction plant: A mill or a treatment place for the extraction
of values from ore.
roast: To heat to a point somewhat short of fuzing in order to
expel volatile matter or effect oxidation.
rougher cell: Flotation cells in which the bulk of the gangue is
removed from the ore.
roughing: Upgrading of run-of-mill feed either to produce a low
grade preliminary concentrate or to reject valueless tailings at
an early stage. Performed by gravity on roughing tables, or in
flotation in a rougher circuit.
rutile: Titanium dioxide, TiO^.
scintillation counter: An instrument used for the location of
radioactive ore such as uranium. It uses a transparent crystal
which gives off a flash of light when struck by a gamma ray, and
a photomultiplier tube which produces an electrical impulse when
the light from the crystal strikes it.
selective flotation: See differential flotation.
settling pond: A pond, natural or artificial, for recovering
solids from an effluent.
siderite: An iron carbonate, FeCO!3.
slime, slimes: A material of extremely fine particle size
encountered in ore treatment.
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sludge: The precipitant or settled material from a wastewater.
slurry: (a) Any finely divided solid which has settled out as
from thickeners, (b) A thin watery suspension.
solvent extraction: See liquid-liquid extraciton.
sphalerite: Zinc sulfide, ZnS.
stibnite: An antimony sulfide, Sb2S!3. The most important ore of
antimony.
suction dredge: (a) Essentially a centrifugal pump mounted on a
barge. (b) A dredge in which the material is lifted by pumping
through a suction pipe.
sulfide zone: That part of a lode or vein not yet oxidized by
the air or surface water and containing sulfide minerals.
surface active agent: One which modifies physical, electrical,
or chemical characteristics of the surface of solids and also
surface tensions of solids or liquid. Used in froth flotation
(see also depressing agent, flotation agent).
tabling: Separation of two materials of different densities by
passing a dilute suspension over a slightly inclined table having
a reciprocal horizontal motion or shake with a slow forward
mot-ion and a fast return.
taconite: (a) The cherty or jaspery rock that encloses the
Mesabi iron ores in Minnesota. In a somewhat more general sense,
it designates any bedded ferruginous chert of the Lake Superior
District, (b) In Minnesota practice, is any grade of extremely
hard, lean iron ore that has its iron either in banded or well-
desseminated form and which may be hematite or magnetite, or a
combination of the two within the same ore body (Bureau of
Mines).
taconite ore: A type of highly abrasive iron ore now extensively
mined in the United States.
tailing pond: Area closed at lower end by constraining wall or
dam tc which mill effluents are run.
tailings: (a) The parts, or a part, of any incoherent or fluid
material separated as refuse, or separately treated as inferior
in quality or value; leavings; remainders; dregs. (b) The gangue
and other refuse material resulting from the washing,
concentration, or treatment of ground ore. (c) Those portions of
washed ore that are regarded as too poor to be treated further;
used especially of the debris from stamp mills or other ore
dressing machinery, as distinguished from concentrates.
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tall oil: The oily mixture of rosin acids, and other materials
obtained by acid treatment of the alkaline liquors from the
digesting (pulping) of pine wood. Used in drying oils, in
cutting oils, emulsifiers, and in flotation agents.
tantalite: A tantalate of iron and manganese (Fe, Mn) Ta20,
crystallizing in the orthorhombic system.
tetrahedrite: A mineral, the part with Sb greater than As of the
tetrahedrite-tenantite series, Cu3_(Sb, As)S3_. Silver, zinc, iron
and mercury may replace part of the copper. An important ore of
copper and silver.
thickener: A vessel or apparatus for reducing the amount of
water in a pulp.
thickening: (a) The process of concentrating a relatively dilute
slime pulp into a thick pulp, that is, one containing a smaller
percentage of moisture, by rejecting liquid that is essentially
solid free. (b) The concentration of the solids in a suspension
with a view to recovering one fraction with a higher
concentration of solids than in the original suspension.
tin minerals: Virtually all th.e industrial supply comes from
cassiterite (SnO_2), though some has been obtained from the
sulfide minerals stannite, cylindrite, and frankeite. The bulk
of cassiterite comes from alluvial workings.
titanium minerals: The main commercial minerals are rutile
(Ti02_) and ilmenite (FeTi03_).
tyuyamunite: A yellow uranium mineral (Ca(U02^2V04.) 2, . 3H^O. It
is the calcium analogue of carnotite.
uraninite: Essentially U02^ It is a complex uranium mineral
containing also rare earths, radium, lead, helium, nitrogen and
other elements.
uranium minerals: More than 150 uranium bearing minerals are
known to exist, but only a few are common. The five primary
uranium-ore minerals are pitchblende, uraninite, davidite,
coffinite, and brannerite. These were formed by deep-seated hot
solutions and are most commonly found in veins or pegmatites.
The secondary uranium-ore minerals, altered from the primary
minerals by weathering or other natural processes, are carnotite,
tyuyamunite and metatyuyamunite (both are very similar to
carnotite), torbernite and metatorbernite, autunite and meta-
autunite, and uranophane.
vanadium minerals: Those most exploited for industrial use are
patronite (VS4_), roscoelite (vanadium mica), vanadinite
(Pb C1(V04)3), carnotite and chlorovanadinite.
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vat leach: Employs the dissolution of copper oxide minerals by
sulfuric acid from crushed, non-porous ore material placed in
confined tanks. The leach cycle is rapid and measured in days.
weir: An obstruction placed across a stream for the purpose of
diverting the water so as to make it flow through a desired
channel, which may be an opening or notch in the weir itself.
wetting agent: A substance that lowers the surface tension of
water and thus enables it to mix more readily. Also called
surface active agent.
Wilfley table: Widely used for of shaking table. A plane
rectangle is mounted horizontally and can be sloped about its
long axis. It is covered with linoleum (occasionally rubber) and
has longitudinal riffles dying at the discharge end to a smooth
cleaning area, triangular in the upper corner. Gentle and rapid
throwing motion is used on the table longitudinally. Sands,
usually classified for size range are fed continuously and worked
along the table with the aid of feedwater, and across riffles
downslope by gravity tilt adjustment, and added washwater. At
the discharge end, the sands have separated into bands, the
heaviest and smallest uppermost, the lightest and largest lowest.
xanthate: Common specific promoter used in flotation of sulfide
ores. A salt or ester of xanthic acid which is made of an
alcohol, carbon disulfite and an alkalai.
xenotime: A yttrium phosphate, YPCU, often containing small
quantities of cerium, terbium, and thorium, closely resembling
zircon in crystal form and general appearance.
yellow cake: (a) A term applied to certain uranium concentrates
produced by mills. It is the final precipitate formed in the
milling process. It is usually considered to be ammonium
diuranate, (NH£) 2U201_, or sodium diuranate, Na2U207_, but the
composition is variable and depends upon the precipitating
conditions, (b) A common form of triuranium octoxide, U308_, is
yellow cake, which is the powder obtained by evaporating an
ammonia solution of the oxide.
zinc minerals: The main source of zinc is sphalerite (ZnS), but
some smithsonite, hemimorphite, zincite, willemite, and
franklinite are mined.
zircon: A mineral, ZrSi04_. The chief ore of zirconium.
zircon, rutile, ilmenite, monazite: A group of heavy minerals
which are usually considered together because of their occurence
as black sand in natural beach and dune concentration.
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APPENDIX A
INDUSTRY PROCESSES
(Refer to Section III)
The following ore types are included: iron, copper, lead-zinc,
gold, silver, molybdenum, tungsten, vanadium, mercury, uranium,
antimony and titanium.
Iron Ore Milling Processes
Beneficiation of iron ore includes such operations as crushing,
screening, blending, grinding, concentrating, classifying,
briquetting, sintering and agglomerating. Beneficiation is often
done at or near the mine site. Methods selected are based on
physical and chemical properties of the crude ore. General
techniques utilized in the beneficiation of iron ore are
illustrated in Figure A-l. Processes enhance either the chemical
or physical characteristics of the crude ore to make more
desirable feed for the blast furnace. Beneficiation methods have
been developed to upgrade 20 to 30 percent iron 'taconite' ores
into high-grade materials.
Physical concentrating processes, such as washing, remove
unwanted sand, clay, or rock from crushed or screened ore. For
those ores not amenable to simple washing operations, other phys-
ical methods are used such as jigging, heavy-media separation,
flotation, and magnetic separation. Jigging involves stratifica-
tion of ore and gangue by utilizing pulsating water currents.
Heavy-media separation employs water suspension of ferrosilicon
whereby iron ore particles sink while the majority of gangue
(quartz, etc.) floats. The flotation process uses air bubbles
attached to iron particles conditioned with flotation reagents to
separate iron from the gangue. Magnetic separation techniques
are used on ores containing magnetite.
At the present time, there are only three iron ore flotation
plants in the United States. Figure A-2 illustrates a typical
flowsheet used in an iron ore flotation circuit, while Table A-l
lists types and amounts of flotation reagents used per ton of ore
processed. Various flotation methods which utilize these rea-
gents are listed in Table A-2. The most commonly adopted flow-
sheet for the beneficiation of low grade magnetic taconite ores
is illustrated in Figure A-3. Low grade ores containing magne-
tite are very susceptible to concentrating processes, yielding a
high quality blast furnace feed. Higher grade iron ores con-
taining hematite cannot be upgraded much above 55 percent iron.
Figure A-4 illustrates the beneficiation of a fine-grained
hematite ore.
Agglomeration, which follows concentration processes, increases
the particle size of iron ore and reduces "fines" which normally
would be lost in the flue gases. Agglomerating methods include
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sintering, palletizing, briquetting, and nodulizing. Sintering
involves the mixing of small portions of coke and limestone with
the iron ore, followed by combustion. A granular, coarse, porous
product is formed. Pelletizing involves the formation of pellets
or balls composed of iron ore fines, followed by heating (Figure
A-5 illustrates a typical pelletizing operation). Hot ore
briquetting requires no binder, is less sensitive to changes in
feed composition, requires little or no grinding and requires
less fuel than sintering. Small or large lumps of regular shape
are formed. Nodules or lumps (nodulizing) are formed when ores
are charged into a rotary kiln and heated to incipient fusion
temperatures.
Copper Ore Milling Processes
Processing of copper ores may involve hydrometallurgical or
physical-chemical separation from the gangue material. A general
scheme of methods employed for recovery of copper from ores is
shown in Figure A-6. These methods include dump, heap, vat and
in-situ leaching, and froth flotation.
Cement copper is produced from dump, heap and in-situ leaching
and cathode copper is produced by electrowinning the pregnant
solution from a vat leach. Major copper areas employing dump,
heap and in-situ leaching are shown in Figure A-7.
Copper bearing froth from the froth flotation process is
thickened, filtered and sent to a smelter whereby blister copper
(98 percent Cu) is produced. The blister copper is then sent to
a refinery which produces pure copper (99.88 to 99.9 percent Cu)
for market.
One combination of the hydrometallurgical and physical-chemical
processes, termed LPF (leach-precipitation-flotation) has enabled
the copper industry to process oxide and sulfide minerals
efficiently. Also, tailings from the vat leaching process, if
they contain significant sulfide copper, can be sent to the
flotation circuit to float copper sulfide, while the vat leach
solution undergoes iron precipitation or electrowinning to
recover copper dissolved from oxide ores by acid.
Lead-Zinc Ore Milling Processes
Generally, lead-zinc ores are not of high enough grade to be
smelted directly, therefore it is sent through the milling pro-
cess first. In most cases, the only process utilized is froth
flotation, but in some cases, preliminary gravity separation is
practiced prior to flotation. The general milling procedure is
to crush the ore and then grind it, in a closed circuit with rod
mills, ball mills and classifying equipment, to a small enough
size to allow the ore minerals to be freed from the gangue.
Chemical reagents are then added which, in the presence of forced
air bubbles, produce selective flotation and separation of the
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desired ore minerals. In some cases, the reagents used in the
flotation process are added in the mill; in other cases, the fine
material from the mill flows to a conditioner (mixing tank),
where the reagents are added. The particular reagents utilized
are a function of the mineral concentrates to be recovered. The
specific choice of reagents used at a facility is usually the
result of determining empirically which reagents yield the opti-
mum mineral values versus reagent costs. In general, lead and
zinc as well as copper sulfide flotations are run at elevated pH
(8.5 to 11, generally) levels so that frequent pH adjustments
with hydrated lime (CaOH^) are common. Other reagents commonly
used are:
Reagent Purpose
Methyl Isobutyl-carbinol Frother
Propylene Glycol Methyl Ether Frother
Long-Chain Aliphatic Alcohols Frother
Pine Oil Frother
Potassium Amyl Xanthate Collector
Sodium Isopropol Xanthate Collector
Sodium Ethyl Xanthate Collector
Dixanthogen Collector
Isopropyl Ethyl Thionocarbonate Collectors
Sodium Diethyl-dithiophosphate Collectors
Zinc Sulfate Zinc Depressant
Sodium Cyanide Zinc Depressant
Copper Sulfate Zinc Activant
Sodium Dichromate Lead Depressant
Sulfur Dioxide Lead Depressant
Starch Lead Depressant
Lime pH Adjustment
The finely ground ore slurry is introduced into a series of
flotation cells, where the slurry is agitated and air is
introduced. The desired minerals are rendered hydrophobic (non-
water-accepting) by surface coating with appropriate reagents.
Usually, several cells are operated in a countercurrent flow
pattern, with the final concentrate being floated off the last
cell (cleaner) and the tails being removed from the first or
rougher cells.
In many cases, more than one mineral is recovered. In such
cases, differential flotation is practiced. The flow diagram in
Figure A-8 depicts a typical differential flotation process for
recovery of lead and zinc sulfides. Chemicals which induce
hydrophilic (affinity-for-water) behavior by surface interaction
are added to prevent one of the minerals from floating in the
initial separation. The underflow of tailings from this separa-
tion is then treated with a chemical which overcomes the depres-
sing effect and allows the flotation of the other mineral.
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The floated concentrates are dewatered (usually by thickening and
filtration), and the final concentratewhich contains some
residual wateris eventually shipped to a smelter for metal
recovery. The liquid overflow from the concentrate thickeners is
typically recycled in the mill.
After the recovery of the desirable minerals, a large volume of
tailings or gangue material remains as underflow from the last
rougher cell in the flow scheme. These tails are typically
adjusted to a slurry suitable for hydraulic transport to -the
treatment facility, i.e., tailing pond. In some cases, the
coarse tailings are removed by a cyclone separator and then
pumped' in to the mine for backfilling.
The tailings from a lead/zinc flotation mill contain residual
solids from the original ore which has been finely ground to
allow mineral recovery. The tailings also contain dissolved
solids and excess mill reagents. In cases where the mineral
content of the ore varies, excess reagents will undoubtedly be
present when the ore grade drops suddenly, conversely lead and
zinc will escape with the tails if high-grade ore creates a
reagent-starved system. Accidental spilling of the chemical
reagents used are another source of adverse discharges from a
mill.
Figure A-8 depicts a typical lead-zinc ore mining and processing
operation.
Gold Ore Milling Processes
Milling practices applicable to the processing and recovery of
gold and gold-containing ores are cyanidation, amalgamation,
flotation, and gravity concentration. All of these processes
have been used in the beneficiation of ore mined from lode
deposits. Placer operations, however, employ only gravity
methods which in the past were sometimes used in conjunction with
amalgamation.
Prior to 1970, the amalgamation process was used to recover
nearly 1/4 of the gold produced domestically. Since that time,
environmental concerns have caused restricted use of mercury. As
a result, the percent of gold produced which was recovered by the
amalgamation process dropped from 20.3 percent in 1970 to 0.3
percent in 1972. At the same time, the use of cyanidation
processes was increasing. In 1970, 36.7 percent of the gold
produced domestically was recovered by cyanidation, and this
increased to 54.6 percent in 1972.
The amalgamation process as currently practiced (used by a single
mill in Colorado) involves crushing and grinding of the lode ore,
gravity separation of the gold-bearing black sands by jigging,
and final concentration of the gold by batch amalgamation of the
sands in a barrel amalgamator. In the past, amalgamation of lode
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ore has been performed in either the grinding mill, on plates, or
in special amalgamators. Placer gold/silver-bearing gravels are
beneficiated by gravity methods, and, in the past, the precious
metal-bearing sands generally were batch amalgamated in barrel
amalgamators. However, amalgamation in specially designed sluice
boxes were also practiced.
There are basically four methods of cyanidation currently being
used in the United States: heap leaching, vat leaching,
agitation leaching, and the recently developed carbon-in-pulp
process. Heap leaching is a process used primarily for the
recovery of gold from low-grade ores. This is an inexpensive
process and, as a result, has also been used recently to recover
gold from old mine waste dumps. Higher grade ores are often
crushed, ground, and vat leached or agitated/leached to recover
the gold.
In vat leaching, a vat is filled with the ground ore (sands)
slurry, water is allowed to drain off, and the sands are leached
from the top with cyanide, which solubilizes the gold (Figure A-
9). Pregnant cyanide solution is collected from the bottom of
the vat and sent to a holding tank. In agitation leaching, the
cyanide solution is added to a ground ore pulp in thickeners, and
the mixture is agitated until solution of the gold is achieved
(Figure A-10). The cyanide solution is collected by decanting
from the thickeners.
Cyanidation of slimes, generated during wet grinding, is cur-
rently being done by a recently developed process, carbon-in-pulp
(Figure A-9). The slimes are mixed with a cyanide solution in
large tanks, and the solubilized gold cyanide is collected by
adsorption onto activated charcoal. Gold is stripped from the
charcoal using a small volume of hot caustic; an electrowinning
process is used for final recovery of the gold in the mill.
Bullion is subsequently produced at a refinery.
Gold in the pregnant cyanide solutions from heap, vat, or agitate
leaching processes is recovered by precipitation with zinc dust.
The precipitate is collected in a filter press and sent to a
smelter for the production of bullion.
Recovery of gold by flotation processes is limited, and less than
3 percent of the gold produced in 1972 was recovered in this
manner. This method employs a froth flotation process to float
and collect the gold-containing minerals (Figure A-ll). The one
operation that uses this method, further processes tailings from
the flotation circuit by the agitation/cyanidation method to
recover the residual gold values.
Gold has historically been recovered from placer gravels by
purely physical means. Present practice involves gravity separa-
tion, which is normally accomplished in a sluice box. Typically,
a sluice box consists of an open box in which a simple rectangu-
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lar sluice plate is mounted on a downward incline. To effect the
separation of gold from gravel or sand, a perforated metal sheet
is fitted on the bottom of the loading box and riffle structures
are mounted on the bottom of the sluice plate. These riffles may
consist of wooden strips or steel or plastic plates which are
angled away from the direction of flow in a manner designed to
create pockets and eddy currents for the collection and retention
of gold.
During actual sluicing operations, pay gravels (i.e., goldbearing
gravels) are loaded into the upper end of the sluice box and
washed down the sluice plate with water, which enters at right
angles to (or against the direction of) gravel feed. Density
differences allow the particles of gold to settle and become
entrapped in the spaces between the riffle structures, while the
less-dense gravel and sands are washed down the sluice plate.
Eddy currents keep the spaces between riffle structures free of
sand and gravel but are not strong enough to wash out the gold
which collects there.
Other types of equipment which may be employed in physical
separation operations include jigs, tables, and screens.
However, this equipment is typically found only at dredging
operations.
Cleanup of gold recovered by gravity methods is normally
accomplished with small (102 cm (40 in.)) sluices, screens, and
finally, by hand-picking impurities from the gold.
Silver Ore Milling Processes
Present extractive metallurgy for silver was developed over a
period of more than TOO years. Initially, silver, as the major
product, was recovered from rich oxidized ores by relatively
crude methods. As the ores became leaner and more complex, an
improved extractive technology was developed. Today, silver pro-
duction is predominantly as a byproduct, and is largely related
to the production of lead, zinc, and copper from the processing
of sulfide ores by froth flotation and smelting. Free-milling,
easily liberated gold/silver ores, processed by amalgamation and
cyanidation, now contribute only 1 percent of the domestic silver
produced. Primary sulfide ores, processed by flotation and
smelting, account for 99 percent.
Selective froth flotation processing can effectively and effi-
ciently beneficiate almost any type and grade of sulfide ore.
This process employs various well-developed reagent combinations
and conditions to enable the selective recovery of many different
sulfide minerals in separate concentrates of high quality. The
reagents commonly used in the process are generally classified as
collectors, promoters, modifiers, depressants, activators, and
frothing agents. Essentially, these reagents are used in combin-
ation to cause the desired sulfide mineral to float and be
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collected in a froth while the undesired minerals and gangue
sink. Practically all the ores presently milled require fine
grinding to liberate the sulfide minerals from one another and
from the gangue minerals.
A circuit which exemplifies the current practice of froth
flotation for the primary recovery of silver from silver ores or
complex ores is shown in Figure A-12. Primary recovery of silver
occurs mainly from the mineral tetrahedrite, (Cu, Fe, Zn, Ag)
]_2Sb4S]_3. A tetrahedrite concentrate contains approximately 25
to 32 percent copper in addition to the 25.72 to 44.58 kilograms
per metric ton (750 to 1,300 troy ounces per ton) of silver. A
low-grade (3.43 kg per metric ton; 100 troy ounces per ton)
silver/pyrite concentrate is produced at one mill. Antimony may
comprise up to 18 percent of the tetrahedrite concentrate and may
or may not be extracted prior to shipment to a smelter.
Various other silver-containing minerals are recovered as
byproducts of primary copper, lead, and/or zinc operations.
Where this occurs, the usual practice is to ultimately recover
the silver from the base-metal flotation concentrates at the
smelter or refinery.
Molybdenum Ore Milling Processes
The only commercially important ore of molybdenum is molybdenite,
MoS^. It is universally concentrated by flotation. Significant
quantities of molybdenite concentrate are recovered as a
byproduct in the milling of copper and tungsten ores.
Flotation concentration has become a mainstay of the ore milling
industry. Because it is adaptable to very fine particle sizes
(less than 0.01 mm, or 0.0004 inch), it allows high rates of
recovery from slimes which are inevitably generated in crushing
and grinding and are not generally amenable to physical
processing. As a physicochemical surface phenomenon, it can
often be made highly specific, allowing production of high-grade
concentrates from very-low-grade ore. Its specificity also
allows separation of different ore minerals (e.g., CuS and MoS2^
where desired, and operation with minimum reagent consumption
since reagent interaction is typically only with the particular
materials to be floated or depressed.
The major operating plants in the industry recover molybdenite by
flotation. Vapor oil is used as the collector, and pine oil is
used as a frother. Lime is used to control pH of the mill feed
and to maintain an alkaline circuit. In addition, Nokes reagent
and sodium cyanide are used to prevent flotation of galena and
pyrite with the molybdenite. A generalized, simplified flowsheet
for an operation recovering only molybdenite is shown in Figure
A-13. Water use in this operation currently amounts to
approximately 1.8 tons of water per ton of ore processed,
essentially all of which is process water. Reclaimed water from
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thickeners at the mill site (shown on the flowsheet) amounts to
only 10 percent of total use.
Where byproducts are recovered with molybdenite, a somewhat more
complex mill flowsheet results, although the molybdenite recovery
circuits remain quite similar. A very simplified flow diagram
for such an operation is shown in Figure A-14. Pyrite flotation
and monazite flotation are accomplished at acid pH (4.5 and 1.5,
respectively), thereby increasing the likelihood of solubilizing
heavy metals. Flow volumes at those locations in the circuit are
low, however, and neutralization occurs upon combination with the
main mill water flows for delivery to the tailing ponds. Water
flow for this operation amounts to approximately 2.3 tons per ton
of ore processed, nearly all of which is process water in contact
with the ore. Essentially 100 percent recycle of mill water from
the tailing ponds at this mill is prompted by limited water
availability as well as by environmental considerations.
Tungsten Ore Milling Processes
Commercially important tungsten ores include the scheelite
(CaW04) and wolframite series, wolframite ((Fe, MN) W04_),
ferberite (FeW04_), and huebnerite (MnW04_) . Concentration is by a
wide variety of techniques. Gravity concentration, by jigging,
tabling, or sink/float methods, is frequently employed. Because
sliming due to the high friability of scheelite ore (most U.S.
ore is scheelite) reduces recovery by gravity techniques, fatty-
acid flotation may be used to increase recovery. Leaching may
also be employed as a major beneficiation step and is frequently
practiced to lower the phosphorus content of concentrates. Ore
generally contains about 0.6 percent tungsten, and concentrates
containing about 70 percent W03_ are produced. A tungsten
concentrate is also produced as a byproduct of molybdenum milling
at one operation in a process involving gravity separation,
flotation, and magnetic separation.
Figure A-15 depicts a simplified flow diagram for a small
tungsten concentrator.
Vanadium Ore Processes
Eighty-six percent of vanadium oxide production has recently been
used in the preparation of ferrovanadium. Although a fair share
of U.S. vanadium production is derived as a byproduct of the
mining of uranium, there are other sources of vanadium ores. The
environmental considerations at mine/mill operations not
involving radioactive constituents are fundamentally different
from environmental considerations important to uranium
operations, and it seems appropriate to consider the former
operation separately. Vanadium is considered as part of this
industry segment: (a) because of the similarity of
nonradioactive vanadium recovery operations to 'the processes used
for other ferroalloy metals and (b) because, in particular,
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hydrometallurgical processes like those used in vanadium recovery
are becoming more popular in SIC 1061.
Vanadium is chemically similar to columbium (niobium) and
tantalum, and ores of these metals may be beneficiated in the
same type of process used for vanadium. There is also some
similarity to tungsten, molybdenum, and chromium.
Recovery of vanadium phosphate rocks in Idaho, Montana, Wyoming,
and Utahwhich contain about 28 percent P205_, 0.25 percent V205^
and some Cr, Ni, and Moyields vanadium as a byproduct of
phosphate fertilizer production. Ferrophosphate is first
prepared by smelting a charge of phosphate rock, silica, coke,
and iron ore (if not enough iron is present in the ore). The
product when separated from the slag typically contains 60 per-
cent iron, 25 percent phosphorus, 3 to 5 percent chromium, and 1
percent nickel. It is pulverized, mixed with soda ash (Na2C03_)
and salt, and toasted at 750 to 800 degrees Celsius (1382 to 1472
degrees Fahrenheit). Phosphorus, vanadium, and chromium are
converted to water-soluble trisodium phosphate, sodium
metavanadate, and sodium chromate, while the iron remains in
insoluble form and is not extracted in a water leach following
the roast.
Phosphate values are removed from the leach in three stages of
crystallization. Vanadium can be recovered as V2_0!5_ (redcake) by
acidification, and chromium is precipitated as lead chromate. By
this process, 85 percent of the vanadium, 65 percent of the
chromium, and 91 percent of the phosphorus can be extracted.
Another, basically non-radioactive, vanadium ore, with a grade of
1 percent V205_, is found in a vanidif erous, mixed-layer
montmorillonite/illite and geothite/montroseite matrix. This ore
is recovered by salt roasting, following extrusion of pellets, to
yield sodium metavanadate, which is concentrated by solvent
extraction. Slightly soluble ammonium vanadate is precipitated
from the stripping solution and calcined to yield vanadium
pentoxide. A flow chart for this process is shown in Figure A-
16.
Mercury Ore Milling Processes
The principal mineral source of mercury is cinnabar (HgS). The
domestic industry has been centered in California, Nevada, and
Oregon. Mercury has also been recovered from ore in Arizona,
Alaska, Idaho, Texas, and Washington and is recovered as a
byproduct from gold ore in Nevada and zinc ore in New York.
Until recently, the typical practice of the industry has been to
feed mercury ore directly into rotary kilns for recovery of
mercury by roasting. This has been such an efficient method that
extensive beneficiation is precluded. However, with the deple-
tion of high grade ores, concentration of low-grade mercury ores
572
-------
is becoming more important. The ore may be crushed and sometimes
screened to provide a feed suitable for furnacing. Gravity con-
centration is also done in a few cases, but its use is limited
since mercury minerals crush more easily and more finely than
gangue rock.
Flotation is the most efficient method for beneficiating mercury
ores when beneficiation is practiced. An advantage of flotation,
especially for low-grade material, is the high ratio of
concentration that results. This permits proportionate
reductions in the size and costs of the final mercury extraction
process. Only recently has flotation of mercury been practiced
in the United States. During 1975, a single mill, located in
Nevada, began operation to beneficiate mercury ore by this method
(Figure A-17). The concentrate produced is furnaced at the same
facility to recover elemental mercury. The ore, which averages
4.8 kg of mercury per metric ton (9.5 Ib/short ton), is obtained
from a nearby open-pit mine; the major ore minerals present are
cinnabar (HgS) and corderoite (Hg3S2C12_).
Uranium Ore Milling Processes
Blending, Crushing and Roasting. Ore from the mine can be quite
variable in consistency and grade. Procedures have been
developed to weigh and radiometrically assay the ores. This is
done to achieve uniform grade and consistency.
Ore high in vanadium is sometimes roasted with sodium chloride
after crushing. This converts insoluble heavy-metal vanadates
(vanadium complex) and carnotite to more soluble sodium vanadate,
which is then extracted with water. Ores high in organics may be
roasted to carbonize and oxidize the organics and prevent clog-
ging of hydrometallurgical processes. Clay bearing ores attain
improved filtering and settling characteristics by roasting at
300 degrees Celsius (572 degrees Fahrenheit).
Grinding. Ore is ground less than 0.6 mm (28 mesh) (0.024 in.)
for acid leaching, less than 0.7 mm (200 mesh) for alkaline
leaching in rod or ball mills using water (or preferably, leach)
to obtain a pulp density of about two-thirds solids. Screw
classifiers, thickeners, or cyclones are sometimes used to
control size or pulp density.
Acid Leach. Ores with a calcium carbonate (CaC03_) content of
less than 12 percent are preferentially leached in sulfuric acid,
which extracts values quickly (in four hours to a day), and at a
lower capital and energy cost than an alkaline leach. Any
tetravalent uranium must be oxidized to the uranyl form by adding
an oxidizing agent (typically, sodium chlorate or manganese
dioxide), which is believed to facilitate the oxidation of U(4)
to U(6) in conjunction with the reduction of Fe (3) to Fe (2) at
a redox (reduction/oxidation) potential of about minus 450 mV.
Free-acid concentration is held to between 1 and 100 grams per
573
-------
liter. The larger concentrations are suitable when vanadium is
to be extracted. The reactions taking place in acid oxidation
and leaching are:
2U02^ + 02 2U03_
2U03_ + 2H2S04 + 5H20 2(U02S04)
. 7H20
Uranyl sulfate (U02S04_) forms a complex, hydrouranyl trisulfuric
acid (H4U(D2(S04_)3_ in the leach, and the anions of this acid are
extracted for value.
Alkaline Leach. A solution of sodium carbonate (40 to 50 g per
liter) in an oxidizing environment selectively leaches uranium
and vanadium values from their ores. The values may be
precipitated directly from the leach by raising the pH and adding
sodium hydroxide. The supernatant can be recycled after its
exposure to carbon dioxide. A controlled amount of sodium
bicarbonate (10 to 20 g per liter) is added to the leach to lower
pH which prevents spontaneous precipitation.
This leaching process is slower than acid leaching since other
ore components are not attached and these ore components tend to
shield the uranium values. Therefore, alkaline leach is used at
elevated temperatures of 80 to 100 degrees Celsius (176 to 212
degrees Fahrenheit) and is subjected to the hydrostatic pressure
at the bottom of a 15 to 20 m (49.2 to 65.6 ft) tall tank which
contains a central airlift for agitation (Figure A-18). In some
mills, the leach tanks are pressurized with oxygen to increase
the rate of reaction which normally takes one to three days. The
alkaline leach process is characterized by the following
reactions:
2UOI2 + 02 2U03_ (oxidation)
_ _ + U03 + H20 2NaOH +
Na4(U02.) (C03.)3. (leaching)
2NaOH + C02 Na2C03_ + H20
(recarbonization)
2Na4(U02) (C03_)3_ + 6NaOH
Na2U2p;7 + 6Na2C03_ + 3H20 (precipitation)
Alkaline leaching can be applied to a greater variety of ores
than is currently being done/ however, this process, because of
its slowness, apparently involves greater capital expenditures
per unit production. In addition, the purification of yellow-
cake, generated in a loop using sodium as the alkali element,
consumes an increment of chemicals that tend to appear in stored
or discharged wastewater. Purification to remove sodium ion is
necessary both to meet the specifications of American uranium
574
-------
processors and for the preparation of natural uranium dioxide
fuel. The latter process will be used to illustrate the problem
caused by excess sodium. Sodium diuranate may be considered as a
mixture of sodium and uranyl oxidesi.e., Na2U2_0;7 = Na20 + 2U03_.
The process of generating U02^ fuel pellets from a yellowcake feed
involves reduction by gaseous ammonia at a temperature of a few
hundred degrees C. At this temperature, ammonia thermally
decomposes into hydrogen, which reduces the U03_ component to U02_
and nitrogen (which acts as an inert gas and reduces the risk of
explosion in and around the reducing furnace). With sodium
diuranate as a feed, the process results in a mix of U02^ and Na20
that is difficult to purify (by water leaching of NaOH) without
impairing the ceramic qualities of uranium dioxide. When, in
contrast, ammonium diuranate is used as the feed, all byproducts
are gaseous, and pure U02. remains. The structural integrity of
this ceramic is immediately adequate for extended use in the
popular CANDU (Canadium deuterium-uranium) reactors. Sodium ion,
as well as vanadium values, can be removed from raw yellowcake
(sodium diuranate) produced by alkaline leaching. First, the
yellowcake is roasted, and some of the sodium ion forms water-
soluble sodium vanadate, while organics are carbonized and burned
off. The roasted product is water leached, yielding a V205
concentrate as described below. The remaining sodium diuranate
is redissolved in sulfuric acid,
Na2U201 + 3H2S04 Na2S04_ +
3H20 + 2(U02)S04
and the uranium values are precipitated with ammonia and filtered
to yield a yellowcake (ammonium diuranate or U03_) that is low
in sodium.
U02SOi + H20 + 2NH3_
(NH4J2S04 + U03
The byproduct that is formed, sodium sulfate, being classed
approximately in the same pollutant category as sodium chloride,
requires expensive treatment for its removal. Ammonium-ion dis-
charges, which might result from an ammonium carbonate leaching
circuit, are viewed with more concern, even though there is a
demand for ammonium sulfate for fertilization of alkaline south-
western soils. Ammonium sulfate could be generated by neutraliz-
ing the wastes of the ammonium loop with sulfuric acid wastes
from acid leaching wastes. Opponents of a tested ammonium pro-
cess argue that nitrites, an intermediate oxidation product of
accidentally discharged ammonium ion, present a present health
hazard more severe than from sulfate ion.
Vanadium Recovery. Vanadium, found in carnotite (K2_ (U02_) 2 (V04_) 2_
3H20) as well as in heavy metal vanadatese.g., vanadinite
(9PbO . 3V205^ . PbCD is converted to sodium orthovanadate
(Na3_V04_), which is water-soluble, by roasting with sodium
575
-------
chloride or soda ash (Na2_03_). After water leaching, ammonium
chloride is added, and poorly soluble ammonium vanadates are
precipitated:
Na3_V04 + 3NH4C1 + H20 3NaOH +
NH4V03_ + 2NH40H
(ammonium metavanadate)
NaSVOi + 3NH4C1 SNaCl +
(NH4_)3_V04
(ammonium orthovanadate)
The ammonium vanadates are thermally decomposed to yield vanadium
pentoxide:
3(NH4)3_V04 6NH3_ + 3H20 +
V205.
A significant fraction (86 to 87 percent) of V2_05^ is used in the
ferroalloys industry. There, ferrovanadium has been produced in
electric furnaces (the following reaction applies):
V5. + Fe2_03_ + 8C SCO + 2FeV
or by aluminothermic reduction (See Glossary) in the presence of
scrap iron.
Air pollution problems associated with the salt roasting process
have led many operators to utilize a hydrometallurgical process
for vanadium recovery which is quite similar to uranium recovery
by acid leaching and solvent exchange. The remainder of V2_0j[
production is used in the inorganic chemical industry.
Concentration and Precipitation. Approximately one metric ton of
ore with a grade of about 0.2 percent is treated with one metric
ton (or cubic meter) of leach, and the concentrations) of
uranium and/or vanadium in the pregnant solution are also about
0.2 percent. If values were directly precipitated from the
solution, a significant fraction of the values would remain in
solution. Therefore yellowcake is recycled and dissolved in a
pregnant solution to increase precipitation yield. Direct
precipitation by raising the pH is effective only for an alkaline
leach, because it is more selective for uranium and vanadium. If
this technique were applied to the acid leach process, most heavy
metalsparticularly, ironwould be precipitated, thus severely
contaminating the product.
Uranium (or vanadium and molybdenum) in the pregnant leach liquor
can be concentrated through ion exchange or solvent extraction.
Typical concentrations in the eluate of some of the processes are
shown in Table A-3.
576
-------
Precipitation of uranium from the eluates is achievable without
recycling yellowcake, and the selectivity of these processes
under regulated conditions (particularly, pH), improves the
purity of the product.
All concentration processes operate best in the absence of
suspended solids, and considerable effort is made to reduce the
solids content of pregnant leach liquors (Figure A-19-a). A dis-
tinction is made between quickly settling sands that are not
tolerated in any concentration process and slimes that can be
accomodated to some extent in the resin-in-pulp process (Figure
A-19-b-c). Sands are often repulped, by the addition of some
wastewater stream, to facilitate flow to the tailing pond.
Consequently, there is some latitude for the selection of the
wastewater sent to the tailing pond, and mill operators can take
advantage of this fact in selecting environmentally sound waste
disposal procedures.
Ion exchange and solvent extraction (Figure A-19-b-e) are based
on the same principle: Polar organic molecules tend to exchange
a mobile ion in their structuretypically, C1-, N03_-, HS04_-,C03_
(anions), or H+ or Na+ (cations)for an ion with a greater
charge or a smaller ionic radius. For example, let R be the
remainder of the polar molecule (in the case of a solvent) or
polymer (for a resin), and let X be the mobile ion. Then, the
exchange reaction for the uranyltrisulfate complex is:
4RX + (U02(S04)3_)
R4_UO^(S04)3_ + 4X
This reaction proceeds from left to right in the loading process.
Typical resins adsorb about ten percent of their mass in uranium
and increase by about ten percent in density. In a concentrated
solution of the mobile ionfor example, in N-hydrochloric acid
the reaction can be reversed and the uranium values are eluted
in this example, as hydrouranyl trisulfuric acid. In general,
the affinity of cation exchange resins for a metallic cation
increases with increasing valence (Cr+++, Mg++, Na+), and because
of decreasing ionic radius, with increasing atomic number (92U,
42Mo, 23V). The separation of hexavalent 92U cations by IX or SX
should prove to be easier than that of any other naturally
occurring element,
Uranium, vanadium, and molybdenumthe latter being a common ore
consitutentalmost always appear in aqueous solutions as
oxidized ions (uranyl, vanadyl, or molybdate radicals). Uranium
and vanadium also combine with anionic radicals to form trisul-
fates or tricarbonates in the leach. The complexes react anion-
ically, and the affinity of exchange resins and solvents is not
simply related to fundamental properties of the heavy metal
(uranium, vanadium, or molybdenum), as is the case in cationic
exchange reactions. Secondary properties, including pH and redox
potential, of the pregnant solutions influence the adsorption of
577
-------
heavy metals. For example, seven times more vanadium than
uranium is adsorbed on one resin at pH 9 whereas at pH 11, the
ratio is reversed, with 33 times as much uranium as vanadium
being captured. These variations in affinity, multiple columns,
and control of variations in affinity, multiple columns, and con-
trol of leaching time with respect to breakthrough (the time when
the interface between loaded and regenerated resin, e.g., Figure
A-19-d, arrives at the end of the column) are used to make an IX
process specific for the desired product.
In the case of solvent exchange, the type of polar solvent and
its concentration in a typically nonpolar diluent (e.g., kero-
sene) effect separation of the desired product. The ease with
which the solvent is handled (Figure A-19-e) permits the con-
struction of multistage co-current and countercurrent SX concen-
trators that are useful even when each stage effects only partial
separation of a value from an interferent. Unfortunately, the
solvents are easily polluted by slimes, and complete liquid/solid
separation is necessary. IX and SX circuits can be combined to
take advantage of both the slime resistance of resin-in-pulp ion
exchange and the separatory efficiency of solvent exchange (Eluex
process-Figure A-19-f). The uranium values are precipitated with
a base or a combination of base and hydrogen peroxide. Ammonia
is preferred by a plurality of mills because it results in a
superior product, as mentioned in the discussion of alkaline
leaching. Sodium hydroxide, magnesium hydroxide, or partial
neutralization with calcium hydroxide followed by magnesium
hydroxide precipitation, are also used. The product is rinsed
with water that is recycled into the process to preserve values,
then filtered, dried and packed into 200-liter (55-gallon) drums.
The strength of these drums limits their capacity to 450 kg (1000
pounds) of yellowcake which occupies 28 percent of the drum
volume.
Figure A-19-g illustrates the Split Elution Concentration
process. Figure A-20 illustrates a "Generalized Flow Diagram for
Production of Uranium, Vanadium and Radium."
Antimony Ore Milling Processes
Antimony is recovered from antimony ore and as a byproduct from
silver and lead concentrates.
Only a small percentage of antimony (13 percent in 1972) is
recovered from ore being mined primarily for its antimony
content. Nearly all of this production can be attributed to a
single operation which is using a froth flotation process to
concentrate stibnite (Sb2S3_) (Figure A-21).
The bulk of domestic production of antimony is recovered as a
byproduct of silver mining operations in the Coeur d'Alene dis-
trict of Idaho. Antimony is present in the silver-containing
mineral tetrahedrite and is recovered from tetrahedrite concen-
578
-------
trates in an electrolytic antimony extraction plant owned and
operated by one of the silver mining companies in the Coeur
d'Alene district. Mills are usually penalized for the antimony
content in their concentrates. Therefore, the removal of anti-
mony from the tetrahedrite concentrates not only increases their
value, but the antimony itself then becomes a marketableiitem.
Antimony is also contained in lead concentrates and is ultimately.
recovered as a byproduct at lead smeltersusually as antimonial
lead. This source of antimony represents about 30 to 50 percent
of domestic production in recent years.
Titanium Ore Milling Processes
The method of mining and beneficiating titanium minerals depends
upon whether the ore is contained ina sand or rock deposit. Sand
deposits occurring in Florida, Georgia, and New Jerrsey contain 1
to 5 percent Ti02^% and are mined with floating suction or bucket-
line dredges handling up to 1,088 metric tons (1,200 short tons)
of material per hour. The sand is treated by wet graiity methods
using spirals, cones, sluices, or jigs to produce a bulk, mixed,
heavy-mineral concentra As many as five individual marketable
minerals are then separated from the bulk concentrate by a
cbination of dry separation techniques using magnetic and
electrostatic (high-tension) separators, sometimes in conjunction
with dry and wet gravity concentrating equipment.
High-tension (HT) electrostatic separators are employed to
separate the titanium minerals from the silicate minerals. The
minerals are fed onto a high-speed spinning rotor, and a heavy
corona (glow given off by a high voltage charge) discharge is
aimed toward the minerals at the point where they would normally
leave the rotor. The minerals of relatively poor electrical con-
ductance are pinned to the rotor by the high surface charge they
recieve on passing through the high voltage corona. The minerals
of relatively high conductivity do not readily hold this surface
charge and so leave the rotor in their normal trajectory. Titan-
ium minerals are the only ones present of relatively high elec-
trical conductivity and are, therefore, thrown off the rotor.
The silicates are pinned to the rotor and are removed by a fixed
brush.
Titanium minerals undergo final separation in induced-roll
magnetic separators to produce three products: ilmenite,
leucoxine, and rutile. The separation of these minerals is based
on their relative magnetic properties which, in turn, are based
on their relative iron content: ilmenite has 37 to 65 percent
iron, leucoxine has 30 to 40 percent iron, and rutile has 4 to 10
percent iron.
Tailings from the HT separators (nonconductors) may contain
zircon and monazite (a rare-earth mineral). These heavy minerals
are separated from the other nonconductors (silicates) by various
579
-------
wet gravity methods (i.e., spirals or tables). The zircon (non-
magnetic) and monazite (slightly magnetic) are separated from one
another in induced-roll magnetic separators.
Beneficiation of titanium minerals from beach-sand deposits is
illustrated in Figure A-22.
Ilmenite is also currently mined from a rock deposit in New York
by conventional open-pit methods. This ilmenite/magnetite ore,
averaging 18 percent Ti02_, is crushed and ground to a small
particle size. The ilmenite and magnetite fractions are
separated in a magnetic separator, the magnetite being more
magnetic due to its greater iron content. The ilmenite sands are
further upgraded in a flotation circuit. Beneficiation of
titanium from a rock deposit is illustrated in Figure A-23.
580
-------
Table A-1
REAGENTS USED FOR FLOTATION OF IRON ORES
(Reagent quantities represent approximate maximum usages. Exact chemical composition of reagent
may be unknown.)
1. Anionic Flotation of Iron Oxides (from crude ore)
Petroleum sulfonate: 0.5 kg/metric ton (1 Ib/short ton)
Low-rosin, tall oil fatty acid: 0.25 kg/metric ton (0.5 Ib/short ton)
Sulfuric acid: 1.25 kg/metric ton (2.5 Ib/short ton) to pH3
No. 2 fuel oil: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium silicate: 0.5 kg/metric ton (1 Ib/short ton)
2. Anionic Rotation of Iron Oxides (from crude ore)
Low-rosin tall oil fatty acid: 0.5 kg/metric ton (1 Ib/short ton)
3. Cationic Flotation of Hematite (from crude ore)
Rosin amine acetate: 0.2 kg/metric ton (0.4 Ib/short ton)
Sulfuric acid: 0.15 kg/metric ton (0.3 Ib/short ton)
Sodium fluoride: 0.15 kg/metric ton (0.3 Ib/short ton)
(Plant also includes phosphate flotation and pyrite flotation steps. Phosphate flotation employs
sodium hydroxide, tall oil fatty acid, fuel oil, and sodium silicate. Pyrits flotation amployi
xanthata collector.)
4. Cat ionic Flotation of Silica (from crude ore)
Amine: 0.15 kg/metric ton (0.3 Ib/short ton)
Gum or starch (tapioca fluor): 0.5 kg/metric ton (1 Ib/short ton)
Methylisobutyl carbinol: as required
5. Cationic Flotation of Silica (from magnetite concentrate)
Amine: 5 g/metrtc ton (0.01 Ib/short ton)
Mettiylisobutyl carbinol: as required
581
-------
Table A-2
VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
OF HIGH-GRADE IRON-ORE CONCENTRATE
1. Anionic flotation of specular hematite
2. Upgrading of natural magnetite concentrate by cationic flotation
3. Upgrading of artificial magnetite concentrate by cationic flotation
4. Cationic flotation of crude magnetite
5. Anionic flotation of silica from natural hematite
6. Cationic flotation of silica from non-magnetic iron formation
582
-------
Table A-3
URANIUM CONCENTRATION IN IX/SX ELUATES
PROCESS
U3Og CONCENTRATION (%)
Ion exchange
Rwin-in-pulp
Fixed-bed IX:
Chloride alution
Nitrate elution
Moving-bed IX:
Nitrate elution
0.8 to 1.2
05 to 1.0
1.0 to 2.0
1.9
Solvent extraction
Alkyl phosphates, HCI eluent
Anwx process
Oaoex process
Split elution minewater treatment
30.0 to 60.0
3 to 4
5.0 to 6.5
1.2 to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
583
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ORE
CRUSHING AND
SCREENING
BLENDING
I
T
CONCENTRATING PROCESSES:
PHYSICAL
CHEMICAL
1 f 1 J_ .L
r r r 1 i
WASHING JIGC
..,- MAGNETIC
,IIMI, SEPARATION
"
HEAVY-
MEDIA
SEPARATION
AGGLOMERATION
PROCESSES
|
SINTERING
t
PELLETI2ING N(
t
3DULIZING
1
FLOTATION
*
3RIQUETTING
i
T
i
TO STOCK PILE AND/OR SHIPPING
Figure A-1
BENEFICIATION OF IRON ORES
584
-------
DENSIFYING THICKENER
UNDERFLOW
CONDITIONERS
ROUGHER FLOTATION
ROUGHER
TAIL
TO
TAILING
BASIN
ROUGHER
CONCENTRATE
<10 CELLS)
FROTH OF
FIRST 2 CELLS
CLEANER FLOTATION
CLEANER
TAIL
I FROTH OF
CLEANER FIRST 2 CELLS
CONCENTRATE |
(8 CELLS)
i
RECLEANER FLOTATION
RECLEANER
TAIL
RECLEANER
CONCENTRATE
(7 CELLS)
TOTAL
FLOTATION
CONCENTRATE
I
TO AGGLOMERATION
(FIGURE III-*)
Figure A-2
IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
585
-------
CRUSHED CRUDE ORE
i
3 V
I
COBBER MAGNETIC SEPARATION
CONCENTRATE
BALL MILL
I
CLEANER MAGNETIC SEPARATION
CONCENTRATE
HYDROCYCLONE
OVERSIZE UNDERSIZE
L_
HYDROSEPARATOR
CONCENTRATE
f
FINISHER MAGNETIC SEPARATION
CONCENTRATE
THICKENING
I
LTl
T
TO TAILING BASIN
TO PELLETIZING
(FIGURE 111-1)
Figure A-3
MAGNETIC TACONITE BENEFICIATION FLOWSHEET
586
-------
CRUDE IRON ORE
en
00
_ TAILINGS
Y
CRUSHER
T
AUTOGENOUS
GRINDING
t
t
DESLIMING
t
SELECTIVE
FLOCCULATION
t
SILICA
FLOTATION
1
H CAUSTIC
SODIUM SILICATE
65% IRON
TO USE
Figure A-4
SIMPLIFIED FLOWSHEET FOR FINE-GRAINED HEMATITE BENEFICIATION
-------
CONCENTRATE FILTER CAKE
BALLING DRUM
I
SCREEN
UNDERSIZE OVERSIZE
I
AGGLOMERATION FURNACE
PELLETS
i
EXHAUST GASES
t
TO STOCK PILE TO ATMOSPHERE
AND/OR SHIPPING
Figure A-5
AGGLOMERATION FLOWSHEET
588
-------
ORE K0.t» C.I
ORE (0.1-04% Ciil ORE ournc- Df* i«o
ORE WASTE DUMP!
REFINERY
( ' ( ' 1 TO SM
WASTE PRIMARY - .,«,,,.,,. J
DUMP HEAP IN SITU CRUSHER *^ SCREENING
lACloV IACIOI (ACIO. u..fp t
_ , __._., QXtnf S^HflOf SECONDARY CEM
i < QBE CRUSHER COP
ACIO ACIO ACIO ± t
SOLUTION SOL N SOL N J
ACIO ADD ACID r .
RtCVCLEO RECYCLED MICYCIED SCREENING PRECIfl
f
'Hta7r.*TION M"' »l TERTIARY
(^| CRUSHER MIXED OXIOE/
, ACIO i SULFlUt ORE
ACID Ht CYCLED wminu L_T
. . L 1
' SPONGE IRON f
CEMENT rnrCKlANTTIOW COPPER AND BALL ^_JWATER| '
ELTER
ENT
PER
TATIOT
NT
1
t
WASH
WATER
1 -] ' 1 j-^ 1 .^^ -. VAT LEACH
1 ' T [REAGENTS [ IACIOI
lOOJFITfM TAILING ^ tinricrivir lit .* FLOTATION P« TAILI
1 PONO niiCKCNtns C£LLJ ^ TAILS
1 i i L BARREN
P
s
1 '
HfMNERr RECYCLED WATER CONtENTHATE ELECTRO
WINNING
' f FACILITY
I THICKENfRS 1 O'SCAKO
1 StllFIOC
j III ILK | f
TO DUMP
i ' '
BTI'HODUCI j CUri'tHISI CATHOOf
MOLYBDENUM | CONCfNTKAfl COPPER
T IU SMELUH
IU SM(LTt II 1 i
REGNANT
OLUTION
1 1 TOMARKfT
I (OR flFINtHYI
Figure A-6
GENERAL FLOW DIAGRAM DEPICTING METHODS FOR
TYPICAL RECOVERY OF COPPER FROM ORE
589
-------
ui
to
o
LEACHING ZONES
Figure A-7
MAJOR COPPER AREAS EMPLOYING ACID LEACHING
-------
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS
FUELS
LUBRICANTS
TO POND
AND/OR
MILL
WATER FROM MINE.
RECYCLE OR OTHER
REAGENTS
THICKENING
AND FILTRATION
USUALLY
RECYCLED
TO PROCESS
WATER SYSTEM
CONCENTRATE
TO LEAD
SMELTER
TO
RECYCLE
TO SUBSURFACE
DRAINAGE
WATER
DISSOLVED SOLIDS
SUSPENDED SOLIDS
EXCESS REAGENTS
CONCENTRATE
TO ZINC
SMELTER
USUALLY RECYCLED
TO PROCESS
WATER SYSTEM
Figure A-8
TYPICAL LEAD-ZINC MINING AND PROCESSING OPERATION
591
-------
TO Hintttn-r
TO 911LTIH
Figure A-9
CYANIDATION OF GOLD ORE: VAT LEACHING OF SANDS
AND 'CARBON-IN-PULP1 PROCESSING OF SLIMES
592
-------
ORE
i
CRUSHING
GRINDING
CONDITIONING
COUNTERCURRENT
LEACHING IN
THICKENERS
I
PRECIPITATION
OF GOLD FROM
LEACHATE BY
ADDITION OF
ZINC DUST
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
1
PRECIPITATE
FILTERED AND
THICKENED
TO SMELTER
REAGENTS (CN)
BARREN
PULP
TAILING-POND
DECANT
RECYCLED
BARREN SOLUTION
RECYCLED
Figure A-10
CYANIDATION OF GOLD ORE: AGITATION/LEACH PROCESS
593
-------
ORE
CRUSHING
GRINDING
CONDITIONING
i
SELECTIVE i
FROTH >
FLOTATION
CONCENTRATE
FILTERED
AND THICKENED
T
TO SMELTER
FLOTATION CIRCUIT
TAILINGS
LEACHING IN
THICKNERS
PRECIPITATION OF GOLD
FROM LEACHATE BY
ADDITION OF ZINC DUST
COLLECTION OF
PRECIPITATE IN
FILTER PRESS
PRECIPITATE FILTERED
AND THICKENED
T
REAGENTS (CM)
.BARREN
PULP
TO TAJ LING
POND
TO SMELTER
Figure A-1 1
FLOTATION OF GOLD-CONTAINING MINERALS WITH RECOVERY OF
RESIDUAL GOLD VALUES BY CYANIDATION
594
-------
ORE
CLASSIFICATION
NO. 1
FLOTATION CIRCUIT
CONCENTRATE
NO. 2
FLOTATION CIRCUIT
PYRITE
CONCENTRATE
RETREATMENT
CIRCUIT
NO. 3
FLOTATION CIRCUIT
FINAL
TAILINGS
FINAL Ag
CONCENTRATE
FINAL PYRITE
CONCENTRATE
CONTAINS
25.7 TO 44.6 KILOGRAMS PER
METRIC TON
(750-1300 OUNCES PER SHORT TON):
25 TO 32% COPPER
0 TO 18% ANTIMONY
tCONTAINS 3.43 KILOGRAMS PER
METRIC TON (100 TROY OUNCES
PER SHORT TON)
Figure A-12
RECOVERY OF SILVER ORE BY FROTH FLOTATION
595
-------
MINING
ORE
±
CRUSHING:.
WEIGHING. AND
SCREENING
i
r
BALL
MILLS
'
CYCLONES
t
UNDERFLOW
OVERFLOW
-J
-TAILS -
ROUGHER FLOAT
MIDDLINGS
1 MIDDLINGS
SCAVENGER
FLOAT
(4 STAGES WITH
REGRIND AND
INTERNAL RECYCLE!
-CONCENTRATE -
CONCENTRATE -
TAILS
TAILS-
CONCENTRATE
--UNDERFLOW
OVERFLOW
-L
-UNDERFLOW-
OVERFLOW
RECLAIM
WATER
DRYER
MOLYBDENUM
PRODUCT
TO TAILING
POND
Figure A-1 3
SIMPLIFIED MOLYBDENUM MILL FLOWSHEET
SHOWING RECOVERY OF MOLYBDENITE ONLY
596
-------
CONCENTRATE
LIGHT TO TAILS
LIGHT TO TAILS
MONAZITE
CONCENTRATE ~
TO TAILS
CRUSHING
(3 STAGES)
1
28% + 3 MESH
1
GRINDING
BALL MILLS
1
36% + 100 MESH
^1
*
FLOTATION
t
FLOTATION
96% OF MILL FEED
+
GRAVITY
HUMPHREY'S SPIRALS
*
PYRITE
FLOTATION
1
TAILS
1
TABLES
-f
MONAZITE |
FLOTATION
i r
MAGNETIC 1
SEPARATION |
1
NONMAGNETIC
TIN CONCENTRATE i r
CONCENTRATE
1
CLEANER
FLOTATION
(4 STAGES)
1 '
DRYING
1
MOLYBDENUM
CONCENTRATE
(93% + MoS2)
-TAILINGS-
MAGNETIC TUNGSTEN
CONCENTRATE
Figure A-14
SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
SHOWING RECOVERY OF MOLYBDENITE AND BYPRODUCTS
597
-------
ORE
SULFIDE
FLOTATION
CYCLONE
25%.
SLIMES
75% SANDS
i
GRAVITY
TABLES
TAILINGS
OVERFLOW
THICKENER
SCHEELITE
FLOTATION
HC1 LEACH
(15 TO 20% OF
FRACTION)
TUNGSTEN
CONCENTRATE
Figure A-1 5
SIMPLIFIED FLOW DIAGRAM FOR A SMALL
TUNGSTEN CONCENTRATOR
598
-------
6-10%
NaCL
H2S04
TERTIARY
AMINES
1.5 - 2.0% V2O5
t
GRINDING
PELLETIZING
ROASTING
850°C (15G2°F)
NaVO,
LEACHING AND
ACIDIFICATION
pH 2.5 - 3.5
SODIUM DECAVANADATE)
SOLVENT EXTRACTION
NH4OH
PRECIPITATION
I
NH4V03
PRECIPITATE
L
CALCINING
T
V2O5 PRODUCT
Figure A-16
ARKANSAS VANADIUM PROCESS FLOWSHEET
599
-------
ORE
i
CLAY HOPPER
AUTOGENOUS
MILL
CLASSIFIER
ROUGHER
FLOTATION CELLS
CLEANER
FLOTATION CELLS
CLEANER
FLOTATION CELLS
I
THICKENER
FILTER
CONCENTRATE
PRODUCT
OVERFLOW
WATER
TAILINGS
Figure A-17
FLOW DIAGRAM FOR BENEFICIATION OF
MERCURY ORE BY FLOTATION
600
-------
LEACH
AIRLIFT
Figure A-18
PACHUCA TANK FOR ALKALINE LEACHING
601
-------
FROM
LEACH
PREGNANT
LEACH LIQUOR
SLIMES
SLIMY PULP TO
r .
-...^ .......... ---- -j*
SAND
TAILINGS
CLEAR LEACH LIQUOR
TO COLUMN IX OR SX
a) LIQUID/SOLID SEPARATION
SLIMY,
PREGNANT?
PULP
RESIN IN OSCILLATING BASKET
b) RESIN-IN-PULP PROCHSS: LOADING
BARREN
PULP
TO TAILINGS
BARREN
ELUANT
PREGNANT
'ELUATE TO
PRECIPITATION
c) RESIN-IN-PULP PROCESS: ELUTING
Figure A-19
CONCENTRATION PROCESSES AND TERMINOLOGY
602
-------
BARREN ELUANT
ELUTED (OR
REGENERATED)
RESIN
LOADED
RESIN
PREGNANT ELUATE
TO PRECIPITATION
d) FIXED-BED COLUMN ION EXCHANGE/ELUTION
PR EG
LEAC
L1QU
i
0
NANT
.H
OR 1
u
_!_/;{
TV
r ^
1 f LOADED 1
| ' > 1 ORGANIC r4
SOLVENT c
" 1 i 1
° \
M^
BARREN STI
N \ ELUANT SO
_
° o
8 fr 0
7* °
^\
DIPPED
LVENT
SOLVENT
~" ~
(^ ) ((BARREN ^ ) PREGNANT! il
x--x y LIQUOR x x ELUATE II1
PHASE PHASE
LOADING SEPARATION STRIPPING SEPARATION
e! SOLVENT EXTRACTION
LEACH
>
IX
BASHED -$> u
ELUANT
IX
^ PARTIALLY
SX
91 rtirreu
RESIN
g) S
RECYCLE
ELUANT
il
/ \ _,
PREGNANT
ELUATE
STORAGE)
\ . LOADED
X _^ RESIN
PLIT ELUT1ON
PRECIPITATION
f) ELUEX PROCESS
Figure A-19
CONCENTRATION PROCESSES AND TERMINOLOGY (Continued)
603
-------
MINING
I
ORE TREATMENT
LEACHING
'"I
LIQUID/SOLID
SEPARATION
I
ION EXCHANGE
SOLVENT EXTRACTION
PATH I
PATHE
PRECIPITATION
TO
STOCKPILE
URANIUM
CONCENTRATE
VANADIUM
BYPRODUCT
RECOVERY
-3*
TO
STOCKPILE
Figure A-20
GENERALIZED FLOW DIAGRAM FOR PRODUCTION OR URANIUM
VANADIUM, AND RADIUM
604
-------
MINING
1
ORE
_i
CRUSHING
r
GRINDING
I
CLASSIFICATION
1
ROUGHER
FLOTATION
1
FROTH
TAILS-
SCAVENGER
FLOTATION
CLEANER
FLOTATION
1
FROTH
TAILS
I
FROTH
FILTER
I
FILTRATE
THICKENER
i
WASTE
FINAL
CONCENTRATE
TO SHIPPING
TO
WASTE
Figure A-21
BENEFICIATION OF ANTIMONY SULFIUE ORE BY FLOTATION
605
-------
LORE FED
FROM DREDGE
| "_
1 VIBRATING
SCREENS
f
TO
POND
SPIRALS OR LAMINAR Tnlliwr *». TO
FLOWS (ROUGHER* AND CLEANERS) *"" '""-""<-- "" WASTE
WET MILL
| DR> MILL ' '
SCRUBBER PLANT
*
DRIER
V
ELECTROSTATIC
SEPARATORS
V
"* SODIUW
^ HYDROXIDE
\ '
SPIRALS AND/OR MAGNETIC
TABLES SEPARATOR
MAGNETIC
SEPARATOR
i ^^ RU
! "
T
MONAZ1TE ZIRCON
1 I
t """" I """
TO TO T
SHIPPING SHIPPING SHIP
T
riLE ILMENITE I
O TO
PING SHIPPING
Figure A-22
BENEFICIATION OF HEAVY MINERAL BEACH SANDS
606
-------
MINING
I
ORE
t
CRUSHING
GRINDING
1
CLASSIFICATION
i
MAGNETIC
SEPARATION
MAGNETICS-
. NONMAGNETICS'
MAGNETITE
ILMENITE
AND GANGUE
DEWATERER
FLOTATION
CIRCUIT
T
I
THICKENER
TO
WASTE
I
FILTER
DRIER
I
CONCENTRATE
T
J
TO SHIPPING
Figure A-23
BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
607
-------
APPENDIX B
PRELIMINARY INTERIM PROCEDURE
FOR
FIBROUS ASBESTOS
608
-------
JAN
DRAFT
INTERIM METHOD
FOR ASBESTOS
IN WATER
by
Charles H. Anderson and J. MacArthur Long
Revised December 29, 1973
Analytical Chemistry Branch
U.S. Environmental Protect ioi\ Aqency
Environmental Research Laboratory
College Station Road
Athens, Georgia J0605
609
-------
Preface to Revised EPA Interim Method for
Determining Asbestos in Water
In July 1976 the Preliminary Interim Method for Determining
Asbestos in Water was issued by the Athens Environmental
Research Laboratory. That method was perceived as representing
the current state-of-the-art in asbestos analytical method-
ology. The objective of writing the method was to present a
procedure analytical laboratories could follow that would result
in a better agreement of analytical results. In the past two
years, a significant amount of additional experimental work has
generated data that provide the basis for a more definitive
method than was possible previously.
This revised Interim Method reflects the improvements that have
been made in asbestos analytical methodology since the initial
procedure was drafted. The general approach to the analytical
determination/ however, remains the same as previously out-
lined. That is, asbestos fibers are separated from water by
filtration on a sub-micron pore size membrane filter. The
asbestos fibers are then counted, after dissolving the filter
material, by direct observation in ^ transmission electron
microscope.
The major change in the initial procedure is the elimination of
the condensation washer as a means of sample preparation.
Intra- and inter-laboratory precision data for the method are
presented. Also, a suggested statistical evaluation of grid
fiber counts is included.
DRAFT
610
-------
ASBESTOS
(Interim Method)
(Transmission Electron Microscopy Method)
1. Scope and Application
l.i This method is applicable to drinking water and
water supplies.
1.2 The method determines the number of asbestos fibers/
liter, their size (length and width), the size
distribution, and total mass. The method distin-
guishes chrysotile from amphibole asbestos. The
detection limits are variable and depend upon the
amount of total extraneous particulate matter in the
sample as well as the contamination level in the
laboratory environment. Under favorable circum-
stances 0.1 MFL (million fibers per liter) can be
detected. The detection limit for total mass of
asbestos fibers is also variable and depends upon
the fiber size and size distribution in addition to
the factors affecting the total fiber count. The
detection limit under favorable conditions is in the
order of 0.1 ng/1.
1.3 The method is not intended to furnish a complete
characterization of all the fibers in water.
1.4 It is beyond the scope of this method to furnish
detailed instruction in electron microscopy,
electron diffraction or crystallography. It is
assumed that those using this method will be suffi-
ciently knowledgeable in these fields to understand
the methodology involved.
2. Summary of Method
2.1 A variable, known volume of water sample is filtered
through a membrane filter of sufficiently small pore
size to trap asbestos fibers. A small portion of
the filter with deposited fibers is placed on an
electron microscope grid and the filter material
removed by gentle solution in organic solvent. The
material remaining on the electron microscope grid
is examined in a transmission microscope at high
magnification. The asbestos fibers are identified
by their morphology and electron diffraction pattern
and their length and width are measured. The total
area examined in the electron microscope is deter-
mined and the number of asbestos fibers in this area
611
-------
is counted. The concentration in MFL (millions of
fibers/liter) is calculated from the number of
fibers counted, the amount of water filtered, and
the ratio of the total filtered area/sampled filter
area. The mass/liter is calculated from the assumed
density and the volume of the fibers.
3. Definitions
Asbestos - A generic term applied to a variety of
commercially useful fibrous silicate minerals of the
serpentine or amphibole mineral groups,
Fiber - Any particulate that has parallel sides and a
length/width ratio greater than or equal to 3:1.
Aspect Ratio - The ratio of length to width.
Chrysotile - A nearly pure hydrated magnesium silicate, the
fibrous form of the mineral serpentine, possessing a
unique layered structure in which the layers are
wrapped in a helical cylindrical manner about the
fiber axis.
Amphibole Asbestos - A double chain fibrous silicate
mineral consisting of Si40i_]_ units, laterally
linked by various cations such as aluminum, calcium,
iron, magnesium, and sodium. The members of the
amphibole asbestos consist of the following:
crocidolite, cummingtonite-gruenerice, and the
fibrous forms of tremolite, actinoiite and antho-
phyllite. These minerals consist of or contain
fibers formed through natural growth processes.
Mineral fragments that conform to the definition of
a fiber and that are formed through a crushing and
milling process are analytically indistinguishable
from the naturally formed fibers by this method.
Detection Limit - The calculated concentration in MFL,
equivalent to one fiber above the background or
blank count. (Section 8.6).
Statistically Significant - Any concentration based upon a
total fiber count of five or more in 20 grid squares.
4. Sample Handling and Preservation
4.A Sampling
It is beyond the scope of this procedure to furnioh
detailed instructions for field sampling; the general
principles of sampling waters are applicable, There are
some considerations that apply to asoestos fibers, a
special type of particulJte matter. These fibers are
612
-------
small, and in water range in length from .1 um to 20 um or
more. Because of the range of size there may be a verti-
cal distribution of particle sizes. This distribution
will vary with depth depending upon the vertical distri-
bution of temperature as well as the local meteorological
conditions. Sampling should take place according to the
objective of the analysis. If a representative sample of
a water supply is required a carefully designated set of
samples should be taken representing the vertical as well
as the horizontal distribution and these samples compos-
ited for analysis.
4.1 Containment Vessel
The sampling container shall be a clean conventional
polyethylene, screw-capped bottle capable of holding
at least one liter. The bottle should be rinsed at
least two times with the water that is being sampled
prior to sampling.
NOTE: Glass vessels are not suitable as sampling
containers.
4.2 Quantity of Sample
A minimum of approximately one liter of water is
required and the sampling container should not be
filled. It is desirable to obtain two samples from
one location.
4.3 Sample Preservation
No preservatives should be .added during sampling and
the addition of acids should be particularly
avoided. If the sample cannot be filtered in the
laboratory within 48 hours of its arrival, suffi-
cient amounts (1 ml/1 of sample) of a 2.71% solution
of mercuric chloride to give a final concentration
of 20 ppm of Hg may be added to prevent bacterial
growth.
NOTE 1: It has been reported that prevention of
bacterial growth in water samples can be achieved by
storing the samples in the dark.
5. Interferences
5.1 Mi 5idencification
The guidelines 3et forth in this method for counting
fibrous asbestos require a positive identification
by both morphology nnd crystal structure as shewn by
an electron diffraction pattern. Chrysotile
asbestos has a uniaue tubular structure, usuallv
613
-------
showing the presence of a central canal, and
exhibits a unique characteristic electron diffrac-
tion pattern. Although halloysite fibers may show a
streaking similar to chrysotile they do not exhibit
its characteristic triple set of double spots or
5.3A layer line. It is highly improbable that a
non-asbestiform fiber would exhibit the distin-
guishing chrysotile features. Although amphibole
fibers exhibit characteristic morphology and
electron diffraction patterns, they do not have the
unique properties exhibited by chrysotile. It is
therefore possible though not probable for misiden-
tification to take place. Hornblende is an amphi-
bole and, in a fibrous form, will be mistakenly
identified as amphibole asbestos.
It is important to recognize that a significant
variable fraction of both chrysotile and amphibole
asbestos fibers do not exhibit the required confir-
matory electron diffraction pattern. This absence
of diffraction is attributable to unfavorable fiber
orientation and fiber sizes. The results reported
will therefore be low as. compared to the absolute
number of asbestos fibers that are present.
5.2 Obscuration
If there are large amounts of organic or amorphous
inorganic materials present, some small asbestos
fibers may not be observed because of physical over-
lapping or complete obscuration. This will result
in lew values for the reported asbestos content.
5.3 Contamination
Although contamination is not strictly considered an
interference, it is an important source of erroneous
results, particularly for chrysotile. The possi-
bility of contamination should therefore always be a
consideration.
5.4 Freezing
The effect of free-ing on asbestos fibers is not
known cut there is reason to ouspect tnat fibsr
breait down could occur an»j result: in .1 higher fiooi
count than was pr3^ent in the original sample.
Therefore the ja;tiple shoulJ be transported to the
lac-oratory undor conditions that would avoid
free~i n<7.
614
-------
6. Equipment and Apparatus
6.1 Specimen Preparation Laboratory
The ubiquitous nature of asbestos, especially chry-
sotile, demands that all sample preparation steps be
carried out to prevent the contamination of the
sample by air-borne or other source cf asbestos.
The prime requirement of the sample preparation
laboratory is that it be sufficiently free from
asbestos contamination that a specimen blank deter-
mination using 200 ml of asbestos-free water yields
no more than 2 fibers in twenty grid squares of a
conventional 200 mesh electron microscope grid.
In order to achieve this low level of contamination,
the sample preparation area should be a separate
conventional clean room facility. The room should
be operated under positive pressure and have incor-
porated electrostatic precipitators in the air
supply to the room, or alternatively absolute (HEPA)
filters. There should be no asbescos floor or
ceiling tiles, transite heat-resistant boards, nor
asbastos insulation. Work surfaces shoul,; en 3t-ir.-
less steel or Formica or equivalent. A laminar flow
hood should be provided for sample manipulation.
Disposable plastic lab coats and disposable over-
shoes are recommended. Alternatively new shoes for
all operators should be provided and retained for
clean room use only. A mat (Tacky Mat, Liberty
Induotries, 539 Deming Rd., Serlin, Connecticut
06037, or equivalent) should be placed inside the
entrance to the room to trap any gross contamination
inadvertently brought into the room from contani-
nated shoes. Normal electrical and water services,
including a distilled water supply should be
provided. In addition a source of ultra-pure water
from a still or filtration-ion exchange system is
desirable.
6.2 Instrumentation
6.2.1 Trans:.;, jj iO.i -l^^u.'^.i ."l-croscoce. A trans-
mission electron microscope thit operates at
a minimum of liO KV, has a resolution of 1.0
n;n and a maqn i L" iccit ion range of 300 tc
100,000. If tho upper limit is not attain-
able direct!./ it nay bo attained through tho
use oc auxiliary optical viewing. Et is
mandatory tiiat the instrument be capable of
carrying out 5elected area electron diffrac-
tion (5AED) on an area of JOO nm:. Tho
viewing screen shall have either a railii-
615
-------
meter scale, concentric circles of known
radii, or other devices to measure the.
length and width of the fiber. Most modern
transmission microscopes meet the require-
ments for magnification and resolution.
An energy-dispersive X-ray spectrometer is
useful Cor the identification of-suspected
asbestiform minerals; this accessory to the
microscope, however, is not mandatory.
6.2.2 Data Processor. The large number of repeti-
tive calculations ma*e it convenient to use
computer facilities together with relatively
simple computer programs.
6.2.3 Vacuun Evaporator. For depositing a layer
of carbon on the Nuclepore filter, and for
preparing carbon coated grids.
5.2.4 Low Temperature Plasma Asher . To be used
for the removal of organic material
(including the filter) from samples
containing so much organic matter that
asbestos fibers are obscured. The sample
chamber should be at least 10-cm diameter.
6.3 Apparatus, Supplies and Reagents
6.3.1 Jaffe Wick Washer. For dissolving Nuclecore
filter. Assemble as in 3.3.1. It is illus-
trated in Figure 1.
6.3.2 Filtering Apparatus. 47-mm funnel (Cat Mo.
XX1504700, Millipore Corporation, Order
Service Dept., Bedford, MA 01730). Used to
filter water samples. 25-mm funnel
(Millipore Cat No. XX1C02500). Used to
filter dispersed ash samples.
6.3.3 Vacuum Pump. For use in sample filtration.
Should provide vacuum u? to 20 inches cc
mercury.
5.3.4 EM Grids. 200-mesh copper or nickel grids,
covered with formvar film tor use with
Nuclopore-Jaffe sample preparation method.
These grids may be purchased from manufac-
turers of electron microscopic supplies or
prepared by standard electron microscopic
grid preparation procedures. Finder grids
may be substituted and are useful if the
re-examination of a specific grid opening is
Jes i red.
- 616
-------
Screen Supper'
with Grid
lidga
^i
Glass Slides
A.
Petri Dish
Layer of Filter Paoers
Nucleoore Filter
Carbon
Chloroforn
Forravar
Grid
Carbon
/
Grid
B.
A. v:a shiny Acr-.iratus
D. V.'ashiruj Process
617
-------
6.3.5 Membrane Filters.
47-mm diameter Millipore membrane filter,
type KA, 0.45-um pore size. Used as a
Nuclepore filter support on top of glass
frit.
47-mm diameter Nuclepore membrane filter;
0.1-^m pore size. (Nuclepore Corp., 7035
Commerce Circle, Pleasanton, CA 94566) For
filtration of water sample.
25-mm diameter Millipore membrane filter,
type HA; 0.45-um pore size. Used as
Nuclepore filter support on top of glass
frit.
25-mm diameter Nuclepore membrane filter;
0.1-um pore size. To filter dispersed ashed
Nuclepore filter.
6.3.6 Glass Vials. 30-mm diameter x SO-mm long.
For holding filter during ashing.
\
6.3.7 Glass Slides. 5.1-cn x 7.5-cm. For support
of Nuclepore filter during carbon
evaporation.
6.3.3 Scalpels. With disposable blades and
scissors.
6.3.9 Tweezers. Several pairs for the many
handling operations.
6.3.10 "Scotch"'Doublestick tape. To hold filter
section flat on glass slide while carbon
coating.
6.3.11 Disposable Petri dishes, 50-mm diameter, for
storing membrane filters.
6.3.12 Static Eliminator, 500 microcuries ?o-210.
(Nuclepore Cat. Mo. V090POL001C;!} or equiva-
lent. To eliminate static charges from
membrane filters.
6.3.13 Carbon roJ.:, spectrochemicalLy nuro, 1.3"
dia., 3.6 mm x 1.0 mm neck. For carbon
coating .
6.3.14 Carbon rod sharpener. (C.it. Mo. 1204,
Ernest F. Fullam, Inc., ?. 0. »ox 444,
Schenectady, NY 12301) For sharpening
carbon ro^js to a neck of specified length
and diameter.
3 618
-------
6.3.15 Ultrasonic Bath. (50 watts, 55 HKz). For
dispersing ashed sample and for general
cleaning.
6.3.16 Graduated Cylinder, 500 ml.
6.3.17 Spot plate.
6.3.18 10-yl Microsyringe. For administering drop
of solvent to filter section during sample
preparation.
6.3.19 Carbon grating replica, 2160 lines/mm. For
calibration of EM magnification.
6.3.20 Filter paper. S & S S539 Black Ribbon (9-cm
circles) or equivalent absorbent filter
paper. For preparing Jaffa Wick Washer.
6.3.21 Screen supports (copper or stainless steel)
12 nm x 12 mm, 200 mesh. To support
spaci.T.en griJ in JafUe Wick Washer.
6.3.22 Chlorqform, spectro grade, doubly
distilled. For dissolving Nuclepore filters.
6.3.23 Asbestos. Chrysotile (Canadian),
Crocidolite, Amosite. UICC (Union
Internationale Centre le Cancer) Standards.
Available from Duke Standards Company, 445
Sherman Avenue, Palo Alto, CA 94306.
6.3.24 Petri dish, glass {100 mm diameter x 15 ir.ra
high). For modified Jaffa Wick Washar.
6.3.25 Alconox. (Alcor.ox, Inc., New York, MY
10003) For cleaning glassware. Add 7.5 g
Alconox to a liter of distilled water.
6.3.26 Parafilm. (American Can Company, Neena'.-.,
v;i) Use as protective covering for clean
glassware.
5.3.27 Pipecs, disposable, 5 mi and 50 :?.! .
i.3.23 D i :>t i I lod or ..!»: on i ^c?«J w,i*:or. Filtor
through 0.1-'.:m Muclepor'? filter for JM'-C i :v_;
'jp all r >;? a;] e n t-. .; and tor final rin.:>tn<] o -J
glassware, and for preparing blanks.
6.3. 29 Mercuric chloride, 2.71s solution w/v. Used
a3 sample preservative. See 4.3. Add 5.42
g of reagent grade mercuric chloride (HgCl:'
619
-------
to 100 ml distilled water and dissolve by
shaking. Dilute co 200 ml with additional
water. Filter through 0.1-um iJuciepore
filter paper before using.
7. Preparation of Standards
Reference standard samples of asbestos that can be used
for quality control for a quantitative analytical method
are not available. It is, however, necessary for each
laboratory to prepare at least two suspensions, one of
chrysotile and another of a representative amphibole.
These suspensions can then be used for intra-laboratcry
control and furnish standard morphology photographs and
diffraction patterns.
7.1 Chrysotile Stock Solution.
Grind about 0.1 g of UICC chrysotile in an agate
mortar fcr several minutes, or until it appears to
be a powder. Weigh out 10 me and transfer to a
clean 1 liter volumetric flask, add several hundred
ml of filtered distilled water containing one mi of
a stock mercuric chloride solution and then make up
to 1 liter with filtered distilled water. To pre-
pare a working solution, transfer 10 ml of the above
suspension to another 1-liter flask, add 1 ml of a
stock mercuric chloride solution and make up tc 1
liter with filtered distilled water. This suspen-
sion contains 100 ug per liter. Finally transfer 1
ml of this suspension to a 1-liter flask, add 1 ml
of a stock mercuric chloride solution and make up to
volume with filtered distilled water. The final
suspension will contain 5-10 MFL and is suitable for
laboratory testing.
7.2 Amphibole Stock Dispersion.
Prepare arnphibole suspensions from UICC arnphibcle
samples as in Section 7.1.
7.3 IJ o r. z i f i z 3. t i o n S'~. a r. -1 a r ~1: .
?r-?paro electron microscopi: jrids .ror. ra in :. P.C the
UICC asbestos fibers ,.ieco'_.::. p.'; to 3, Procedure, .ir..i
obtain I'C-pr OGen tat i. vo phonographs ot each fiber typ:
ana its diffraction p.ittorr. fcr future r -3 i-n' or.c .
3 . Proc
-------
quent deposition on a membrane filter is a very
critical step in the procedure. The objective of
the filtration is not only to separate, but also to
distribute uniformly the particulate matter such
that discreet particles are deposited with a minimum
of overlap.
The volume filtered will range from 50-500 ml. In
an unknown sample the volume can not be specified in
advance because of the presence of variable amounts
of particulate matter. In general, sufficient
sample is filtered such that a very faint stain can
be observed on the filter medium. The maximum
loading that can be tolerated is 20 ug/cn11 , or about
200 ug on a 47-mm diameter filter; 5 ug/crn2 is near
optimum. If the total solids content is known, an
estimate of the maximum volume tolerable can be
obtained. In a sample of high solids content, where
less than 50 ml is required, the sample should be
diluted with filtered distilled water so that a
minimum total of 50 ml of water is filtered. This
step is necessary to allow the .insoluble material to
deposit uniformly on the,filter. The filtration
funnel assembly must be scrupulously clean and
cleaned before each filtration. The filtration
should be carried out in a laminar flow hood.
NOTE 1: The following cleaning procedure has been
found to be satisfactory:
Wash each piece of glassware three times with
distilled water. Following manufacturer's recommen-
dations use the ultrasonic .bath with an Alccno:;-
water solution to clean all glassware. After the
ultrasonic cleaning rinse each piece of glassware
three times with distilled water. Then rinse eacn
piece three times with deionized water which has
been filtered through 0.1-um Nuclepore filter. Dry
in an asbestos-free oven. After the glassware is
dry, seal openings with parafilm.
8.1.1 Filtration
a. Assemble the vacuum filtration apparatus
incorporating the . l-'.:;r. Nuclepor:? backed
with 0.45-11.71 MLllipore filter. See S.j,2.
b. Vigorously agitate the water sample in
i ts container.
c. I':. c..o L^qu.Leu filtration volume can oe
estimated, either from turbidity estimates
of suspended solids or previous experience,
immediately withdraw the proper volume from
1 l 621
-------
the container and add the entire volume to
the 47-nim diameter funnel. Apply vacuum
sufficient for filtration but gentle enough
to avoid the formation of a vortex. If a
completely unknown sample is being analyzed,
a slightly modified procedure must be
followed. Pour 500 ml of a well-mixed
sample into a 500-ral graduated cylinder and
immediately transfer the entire contents to
the prepared vacuum filtration appa?atus.
Apply vacuum gently and continue suction
until all of the water has passed through
the filter. If the resulting filter appears
obviously coated or discolored, it is
recommended that another filter be prepared
in the same manner, but this time using only
200 or 100 ml of sample.
NOTE 1: Do not add more water after filtra-
tion has started and do not rinse the sides
of the funnel.
NOTE 2: Nuclepore filter is basically a
hydrophobic material. The manufacturer
applies a detergent to the surface of the
filter in order to render it hydrophilic;
this process, however, does not appear to be
entirely satisfactory in some batches. Pre-
treatment of the filter in a low temperature
asher at 10 watts for 10 seconds can be used
to render the surface of the filter hydro-
philic. This process will significantly
decrease the islands of sparse deposit fre-
quently observed.
d. Disassemble the funnel, remove the
filter and dry in a covered petri dish.
8.2 Preparation of Electron Microscope Grids.
The preparation of the grid for examination in the
microscope is a critical step in the analytical
procedure. The objective is to -rar.ove the organic
tilcer material from the asbestos fibers with a
minimum loss and movement and with a minimum break-
age of the grid support film.
I. the sample contains organic matter in such
amounts that interfere with fiber counting and
identification a preliminary ashing step is
required. See 3.5.
i >
622
-------
8.3 Nuclepore Filter, Modified Jaffe Wick Technique,
8.3.1 Preparation of Modified Jaffe Washer
Place three glass microscope slides (75 mm x
22 mm) one on top of the other in a petri
dish (100 mm x 15 mm) along a diameter.
Place 14 S & S 1589 Black Ribbon filter
papers (9cm circles) in the petri dish over
the stack of microscope slides. Place three
copper raesh screen supports (12 mm x 12 mm)
aior.g the ridge formed by the stack of
slides underneath the layer of filter
papers. Place an EM specimen grid on each
of the screen supports. See Fig. 1.
NOTE 1: A stack of 30-40 S & S filters
(7-cm circle) can be substituted for the 14
filters and microscope slides in preparing
the Jaffe washer.
8.3.2 Vacuum Filtration Unit
Assemble the vacuum filtration unit. Place
a 0.45-ua Millipor-3 filter type KA on the
glass frit and then position a 0.1-um
Nuclepore filter, shiny side up, on top of
the Millipore filter. Apply suction to
center the filters flat on the frit. Attach
the filter funnel and shut off the suction.
8.3.3 Sample Filtration
See 8.1.1.
8.3.4 Sample Drying
Remove the filter funnel and place the
Nuclepore filter in a loosely covered petri
dish to dry. The petri dish containing the
filter may be placed in an asbestos-free
oven at 45° C for 30 minutes to shorter.
the drying time.
3.3.5 Selection of Section for Carbon Coating
Using a small pair ot scissors or sharp
scalpel cut out a retanoular section oc the
Nuclepore filter. The minimum approximate
dimensions anould be 15 ir.m long and 3 mo
..j.Jo. ."v.vx.... _ ^^-_-o_.or. r.£.ar the perimeter of
the filtration area.
623
-------
8.3.6 Carbon Coating the Filter
Tape the two ends of the selected filter
section to a glass slide using "Scotch"
tape. Take care not to stretch the filter
section. Identify the filter section using
a china marker on the slide. Place the
glass slide with the filter section into the
vacuum evaporator. Insert the necked carbon
rod and, following manufacturer's instruc-
tions, obtain high vacuum. Evaporate the
neck, with the filter section rotating, at a
distance of approximately 7.5 cm from the
filter section to obtain a 30-50 nm layer of
carbon on the filter paper. Evaporate the
carbon in several short bursts rather than
continuously to prevent overheating the sur-
face of the Nuclepore filter.
NOTE 1: Overheating the surface tends to
crosslink the plastic, rendering the filter
dissolution in chloroform difficult.
'i
NOTE 2: The thickness of the carbon film
can be monitored by placing a drop of oil on
a porcelain chip that is placed at the same
distance from the carbon electrodes as the
specimen. Carbon is not visible in the
region of the oil drop thereby enabling the
visual estimate of the deposit thickness by
the contrast differential.
3.3.7 Grid Transfer
Remove the filter from the vacuum evaporator
and cut out three sections somewhat less
than 3 mm x 3 mm and such that the square cf
Nuclepore fits within the circumference of
the grid. Pass each of the filter sections
over a static eliminator and then place each
of the three sections carbon-sice down on
separate specimen grids previously placed in
the modified Jaffe Washer. Using a micro-
syringe, place a 10-^1 drop of chloroform on
each filter section resting on a grid and
then saturate the filter pad until pooling
of the solvent occurs below the ridge formed
by the glass slides inserted under the layo:
of filter papers. Place the cover on the
oetri dish and allow the grids to remain in
the washer for approximately 24 hour".
not allow the chloroform to completely
evaporate before the grids are removed. To
remove the grids from the washer lift the
14
624
-------
screen support with the grid resting upon it.
and set this in a spot plate depression to
allow evaporation o£ any solvent adhering to
the grid. The grid is now ready for
analysis or storage.
8.4 Electron Microscopic Examination
3.4.1 Microscope Alignment and Magnification
Calibration
Following the manufacturer's recommendations
carry out the necessary alignment procedures
for optimum specimen examination in the
electron microscope. Calibrate the
routinely used magnifications using a carbon
grating replica.
NOTE 1: Screen magnification is not
necessarily equivalent to plate
magn if icat ion .
3.4.2 Grid Preparation Acceptability
After inserting the specimen into the micro-
scope adjust the magnification low enough
(300X-1QOOX) to permit viewing complete grid
squares. Inspect at least 10 grid squares
for fiber loading and distribution, debris
contamination, and carbon film continuity.
Reject the grid for counting if:
1) The grid is too "heavily loaded with
fibers to perform accurate counting and
diffraction operations. A new sample prepa-
ration either from a smaller volume of water
or from a dilution with filtered distilled
water must then be prepared.
2) The fiber distribution is noticeably
uneven. A new sample preparation is
required .
3) The debris contamination 13 too
to perform accurate counting and diffraction
operation a. If the debris 13 Largely
organic the filter must be ashed and :: o ,.: i _-, -
porsinl (.ioe 3,3) . 1 1: inorganic th
-------
preparation from the same initial filtration
must be substituted.
8.4.3 Procedure for Fiber Counting
There are two methods commonly used for
fiber counting. In one method (A) 100
fibers, contained in randomly selected
fields of view, are counted. The number of
fields plus the area of a field of view raust
be known when using this method. In the
other method (B) , all fibers (at least 100)
in several grid squares or 20 grid squares
are counted. The number of grid squares
counted and the average area of one grid
square raust be known when using this method.
NOTE 1: The method to use is dependent upon
the fibar loading on the grid and it is left
to the judgement of the analyst to select
the optimum method. The following guide-
lines can be used: If it is estimated that
a grid square (80 urn x 80 um) contains
50-100 fibers at %a screen magnification of
20000X it is convenient to use the field-of-
view counting method. If the estimate is
less than 50, the grid square method of
counting should be chosen. On the other
hand, if the fiber count is estimated to be
over 300 fibers per grid square, a new grid
containing less fibers must be prepared
(through dilution or filtration of a smaller
volume of water) .
8.4.3A Field-of-View Method
After determining that a fiber count can be
obtained using this method adjust the screen
magnification to 15,-20,OOOX. Select a
number of grid squares that would be as
representative as possible of the entire
analyzable grid surface. From each of thece
squares select a sufficient number of fieldo
of view for fi-or co.:-. ': in:: . The rvjrsbec of
fields of view per -j. i J square £3 dependent
upon the fi^er loading. If more than one
fi-sld of view per ^riJ square is selected,
scan the grid opening orthogonally in an
arbitrary pattern which prevents overlapping
of t'iolds of view. Carry out the analysis
by counting, measuring and identifying (3ee
8.4.4) approximately 50 fibers on each of
two gr ids .
626
-------
The following rules should be followed when
using the field of view method of fiber
counting. Although these rules were derived
for a circular field of view they can be
modified to apply to square or rectangular
designs.
1) Count all fibers contained within the
counting area and not touching the circum-
ference of the circle.
2) Designate the upper right-hand quadrant
as I and number in clockwise order. Count
all fibers touching or intersecting the arc
of quadrants I or IV. Do not count fibers
touching or intersecting the arc of quad-
rants II or III.
3) If a fiber intersects the arc of both
quadrants III and IV or I and II count it
only if the greater length was outside the
arc of quadrants IV and I, respectively.
4) Count fibers intersecting the arc of
both quadrants I and III but not those
intersecting the arc of both II and IV.
These rules are illustrated in Fig. 2.
8.4.3B Grid Square Method
After determining that a fiber count can be
obtained using this, method adjust the screen
magnification to 15,-20,OOOX. Position the
grid square so that scanning can be started
at the left upper corner of the grid
square. While carefully examining the grid,
scan left to right, parallel to the upper
grid bar. When the perimeter of the grid
square is reached adjust the field of view
down one field width and scan in the oppo-
site direction. The tilting section of the
fluorescent screen may be used conveniently
as the Ciela of view. Examine the square
until all the area has be-?n covered. The
analysis should bo carried out by counting,
measur ing and ivient i ty in.q (see 3.4.4)
approximately 50 Libert on each ot tv;o arid.;
or until 10 grid squares on each of two
nri-Ja have been counted. Do not count
fibers intersecting a gric oar.
17
627
-------
IV
III
II
Counted
Counted
2. I lluj tr.LM L Lori of Co
for F i. o Ld -o f -V L ow
tir.g Rules
628
-------
8,4.4 Measurement and Identification
Measure and record the length and width of
each fiber having an aspect ratio greater
than or equal to three. Disregard obvious
biological, bacteriological fibers and
diatom fragments. Examine the morphology of
each fitter using optical viewing if
necessary. Tentatively identify, by refer-
ence to the UICC standards, chrysotile or
possible amphibole asbestos. Attempt to
obtain a diffraction pattern of each fiber
utilizing the shortest camera length
possible. Move the suspected fiber image to
the center of the screen and insert a suit-
able selected area aperture into the
electron beam so that the fiber image, or a
portion of it, is in the illuminated area.
The size of the aperture and the portion of
the fiber should be such that particles
other than the one to be examined are
excluded from the selected area. Observe
the diffraction ^pattern with the 10X binocu-
lars. If an incomplete diffraction pattern
is obtained move the particle image around
in the selected area to get a clearer
diffraction pattern or to eliminate possible
interferences from neighboring particles.
Determine whether, or not the fiber is chry-
sotile or an amphibole by comparing the
diffraction pattern obtained to the diffrac-
tion patterns of known standard asbestos
fibers. Confirm the tentative identifi-
cation of chrysotile and amphibole asbestos
from their electron diffraction patterns.
Classify each fiber as chrysotile, amphi-
bole, non-asbestos, no diffraction or
ambiguous.
MOTE 1: It is convenient to use a tape
recorder during the examination of the
fibers to record all pertinent data. This
information car. then bo summarized on data
sheets or punched cards for subsequent auto-
matic data processing.
NOTE 2: Chrysotile fibers occur as single
fibrils, or in bundles. The fibrils gone-
rally show a tubular structure with a hollow
canal, altnough the absence of the canal
does not rule out its identitication.
Amphibole asbestos fibers usually exhibit a
lath-like structure with irregular ends, but
1'.)
629
-------
occasionally will resemble chrysotile in
appearance.
NOTE 3: The positive identification of
asbestos by electron diffraction requires
some judgment on the part of the analyst
because some fibers give only partial
patterns. Chrysotile shows unique prominent
streaks on the layer lines nearest the
central one and a triple set of double spots
on the second layer line. The streaks and
the set of double spots are the distin-
guishing characteristics of chrysotile
required for identification. Amphibole
asbestos requires a more complete diffrac-
tion pattern to be positively identified.
As a qualititative guideline, layer lines
for amphibole, without the unique streaks
(some streaking may be present) of chryso-
tile, should be present and the arrangement
of diffraction spots along the layer lines
should be consistent with the amphibole
pattern. The pattern should be distinct
enough to establish these criteria.
NOTE 4: Chrysotile and thin amphibole
fibers may undergo degradation in an elec-
tron beam; this is particularly noticeable
in small fibers. It may exhibit a pattern
for a 1-2 seconds and disappear and the
analyst must be alert to note the character-
istic features.
NOTE 5: An ambiguous fiber is a fiber chat
gives a partial electron diffraction pattern
resembling asbestos, but insufficient to
provide positive identification.
8.4.5 Determination of Grid Square Area
Measure the dimensions of several represen-
tative grid squares from each batch of griis
with an optical microscope. Calculate the
average area of a grid square. This should
be done to compensate ror variability in
grid square Jirr.ens ior.3 .
3.5 A3 h i n 5
Some samples contain sufficiently hi'5h levels of
organic material that an ashing step is required
beCore fiber identification and counting can be
carried out.
23
630
-------
Place the dried Nuclepore filter paper containing
the collected sediment into a glass vial (23 mm dia-
meter x 80 nun high) . Position the filter such that
the filtration side touches the glass wall. Place
the vial in an upright position in the low tempera-
ture asher. Operate the asher at 50 watts (13.56
MHz) power and 2 psi oxygen pressure. Ash the
filter until a thin film of white ash remains. The
time required is generally 6 to 8 hours. Allow the
ashing chamber to slowly reach atmospheric pressure
and remove the vial. Add 10 ml of filtered
distilled water to the vial. Place the vial in an
ultrasonic bath for 1/2 hour to disperse the ash.
Dilute the sample if required.
Assemble the 25-mm diameter filtering apparatus.
Center a 25-mm diameter . 1-ym Nuclepore filter (with
the 0.45-um Millipore backing) on the glass frit.
Apply suction and recenter the filter if necessary.
Attach the filter funnel and turn off the suction.
Add the water containing the dispersed ash from the
vial to the filter funnel. Apply suction and filter
the sample. After drying this filter it is ready to
be used in preparing sample grids as in 8.3.
NOTE 1: In specifying a 25-mm diameter filter it is
assumed that the ashing step is necessary mainly
because of the presence of organic material and that
the smaller filtering area is desirable from the
point of view of concentrating the fibers. If the
sample contains mostly inorganic debris such that
the smaller filtering area will result in over-
loading the filter, the 47-mm diameter filter should
be used.
NOTE 2: It will be noted that a 10-ml volume is
filtered in this case instead of the minimum 50-ml
volume specified in 8.1. These volumes are consis-
tent when it is considered that there is approxi-
mately a 5-fold difference in effective filtration
area between the 25-mm diameter and 47-mm diameter
filters.
3.5 Determination of Blank Level
Carry out a blank determination with each batch of
samples prepared, but a minimum of one per week.
Filter .3 urosh supply (500 ml) at distilled,
deionized water through a clean 0.1-'..m membrane
filter. Filter 200 ml of this w.itor through a
0.1-um Nuclepore filter, prepare the electron micro-
scope grid, and count exactly as in the procedures
8.1 - 3.4. Examine 20 grid squares and record this
number of fibers. A maximum oi two fibers in 20
grid squares is acceptable for the blank sample.
631
-------
NOTE 1: The monitoring of the background level of
asbestos is an integral part of the procedure. Upon
initiating asbestos analytical work, blank samples
must be run to establish the initial suitability of
the laboratory environment, cleaning procedures, and
reagents for carrying out asbestos analyses.
Analytical determinations of asbestos can be carried
out only after an acceptably low level of contami-
nation has been established.
Calculations
9.1 Fiber Concentrations
Gr id Square Counting Method - If the Grid Square
Method of counting is employed, use the following
formula to calculate the total asbestos fiber
concentration in MFL.
C = (F x Af)/(Ag x V0 x 1000)
If ashing is involved use the same formula but
substituting the effective filtration area of the
25-mm diameter filcer for Af inscead of that for
the 47-mm diameter filter. If one-half the filter
is ashed, multiple C by two.
C = Fiber concentration (MFL)
F = average number of fibers per grid opening
AJ = Effective filtration area of filter
paper (mnr ) used ia grid preparation for
fiber counting
A_ = Average area of one grid square (mm")
V"o = Original volume of sample filtered (ml)
Field-ot'-View Counting Method - If the Field-of-View
Method of counting is employed use the following
formula to calculate the total asbestos fiber
concentrations (MFL)
C * (F x Af x 1000) ' (Av x V0)
If J3hinq 13 involved use the same formula but
substitutinq the effective filtration area of the
25-mm diameter filter for ."...- i~:::::o of that for the
47-mm diameter filter.
C = Fiber concentration
632
-------
F = Average number of fibers per field of
view
Af = Effective filtration area of filter
paper (mm2) used in grid preparation for
fiber counting
Av = Area of one field of view ( um 2)
VQ = Original volume of sample filtered (ml]
9,2 Estimated Mass Concentration
Calculate the mass (yg) of each fiber counted using
the following formula:
M = L x W2 x D x 10-6
If the fiber content is predominantly chrysotile,
the following formula may be used:
M = I x L x W2 xs D x 10~5
4
where M = mass (ug)
L - length (ym)
W = width (um)
D = density of fibers fc/'cm3 )
Then calculate the mass concentration (ug/1)
employing the following formula.
l\, = C x Mf x 10
where M = mass concentration (ug/1)
C = fiber concentration (MFL)
Mf = mean mass per fiber (ug)
To calculate M use the following formula:
M . ,
/ n
where M. 3 maas of each fiber, respectively
n - number of fibers counted
633
-------
NOTE i: Because many of the amphibole fibers are
lath shaped rather than square in cross section the
computed mass will tend to be high since laths will
in general tend to lie flat rather than on edge.
NOTE 2: Assume the following densities: Chrysotile
2.5, Amphibole 3.25
9.3 Aspect Ratio
The aspect ratio for each fiber is calculated by
dividing the length by the width.
10. Reporting
10.1 Report the following concentration as MFL
a. Total fibers
b. Chrysotile
c. Amphibole
10.2 Use two significant figures for concentrations
greater than 1 MFL, and one significant figure for
concentrations less than 1 MFL.
10.3 Tabulate the size distribution, length and width.
10.4 Tabulate the aspect ratio.distribution.
10.5 Report the calculated mass as ug/1.
10.6 Indicate the detection limit in MFL.
10.7 Indicate if less than five fibers were counted.
10.3 Include remarks concerning pertinent observations,
(clumping, amount of organic matter, debris) amount
of suspected though not identifiable as asbestos
fibers (ambiguous).
11. Precision
11.1 Intra-Laboratory
The precision that is obtained within an individual
laooratory is dependent upon the number of fibers
countod. If LOO fibers are counted and the loading
is at least 3.5 fibers/grid square, computer
modeling of the counting procedure shows a relative
standard deviation of abouc 10s can be expected.
24
634
-------
In actual practice some degradation from this
precision will be observed but should not exceed +
15% if several grids are prepared from the same
filtered sample. The relative standard deviation of
analyses of the same water sample in the same labo-
ratory will increase due to sample preparation
errors and a relative standard deviation of about
about + 25 - 35% will occur. As the number of
fibers counted decreases, the precision -..ill also
decrease approximately proportional to /N where N is
the number of fibers counted.
Based upon the analysis of one laboratory utilizing
a different analyst for each of three water samples,
intra-laboratory precision data is presented in
Table 1.
Table 1. Intra-Laboratory Precision
Sample Number of Mean Fiber Precision,
Type Sample Concentration Relative
Aliquots MFL (millions of Standard
Analyzed asbestos fibers/1) Deviation
Chrysotile 2S 23 37%
(UICC)
Crocidolite 20 8 36%
(UICC)
Taconite 20 15 24%
(raw water)
11.2 Inter-Laboratory
Based upon the analysis by various government and
private industrial laboratories of filters preparao
from nine water samples, inter-laboratory precision
data of the method is oresented in Table 2.
635
-------
Table 2, Inter-Laboratory Precision
Sample
Type
Chrvsotile
Amphibole
Number of
Labs
Reporting
10
9
11
9
9
3
11
4
14
Mean Fiber
Concentration
MFL (millions of
asbestos fibers/1)
377
119
59
31
23
25
139
95
35
Precision,
Relative
Standard
Deviation
35%
43%
41%
65%
32%
35%
50%
52%
66%
12. Accuracy
12.1 Fiber Concentrations
As no standard reference materials are available,
only approximate estimates of the accuracy of the
procedure can be made. At 1 MFL, it is estimated
that the results should be within a factor of 10 of
the actual asbestos fiber content.
This method requires the positive identification of a
fiber to be asbestos as a means for its quantitative
determination. As the state-of-the-art precludes the
positive identification of-all of the asbestos fibers
present, the results of this method, as expressed as
MFL, will be biased on the low side and assuming no
fiber loss represent 0.4 - 0.3 of the total asbestos
fibers present.
12.2 Mass Concentrations
As in the case of the fiber concentrations, no stan-
dard samples of the sizs distribution found in vitsr
are available. The accuracy of the mass determi-
nation should be somawhat better than the fiber
determination if a statistically significant number
of the larger fibers, winch contribute the major
portion of the mass, are identified, meajurou, .ind
counted. This will reduce the bias oc low results
duo to J i f L" icuL c i e:3 in identification. At the same
time, the? assumption that the thickness of the fibor
equals the width will result in a positive error in
determining the volume of the fiber and thus give
high results cor the mass-
636
-------
13. Suggested Statistical Evaluation of Grid Fiber Counts
13.1 Since the fiber distribution on the sample filter,
resulting from the method of filtration, has not been
fully characterized, the fiber distribution obtained
on the electron microscope grids for each sample
should be tested statistically against an assumed
distribution and a measure of the precision of the
analysis should be provided.
13.2 Assume that the fibers are uniformly and randomly
distributed on the sample filter and grids. One
method for confirming this assumption is given below.
13.3 Using the chi-square test, determine whether the
total number of fibers found in individual grid
openings are randomly and uniformly distributed among
the openings, by the following formula:
N , .2
(nv-np,-)
np^
where X: - chi-square statistic
N = number of grid openings examined for the
sample
n.- = total number of fibers found in each
respective grid opening
n = total number of fibers found in M grid
openings
p~; = ratio of the area of each respective
grid opening to the sum of the areas of all
grid openings examined
NOTE 1: If an average area for the grid squares has
been measured as outlined in 8.4.5, the term np-
represents the mean fiber count per grid square.'
If the value for X: exceeds the value listed in
statistical tables for the O.ls significance lev?!
with N-1 degrees of freedom, the fibers are not
considered to i:o uniformly and randomly d i 31 r ibuto J.
among the grid openings. IP, this case, it is
advisable to try to improve the uniformity of fiber
deposition by filtering another aliquot of the sample
and repeating t'-. : .analysis.
13.4 If uniformity and randomness of fiber deposition on
no microscope grids
has been demonstrated as in
13.3, the 95'i confidence interval about the mean
637
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fiber counts for chrysotile, amphibole, and total
asbestos fibers may be determined using the following
formulae:
(1)
where
N
N
X)
N(N-l)
1/2
Sc = standard deviation of the chrysotile
fiber count
N = number of grid openings examined for the
sample
Xv = number of chrysotile fibers in each grid
opening, respectively
Obtain the standard deviations of the fiber counts
for anphibole asbestos fibers and for total asbestos
fibers by substituting the corresponding value of X
into equation (1) .
ts
(2)
3)
X = X
/N
ts
N
where X = upper value of 95% confidence interval
for chrysotile
XL - lower value of 95% confidence interval
for chrysotile
H = average number of fibers per grid opening
t = value listed in t-distribution tables at
the 95% confidence l-»vel for a two tailed
distribution with N-1 degree of freedom
^ = standard deviation of the floor count::
Cor chrysot ile
=> number of griJ openings examined for the
sample
The values of Xu and XL can be converted to concen-
trations in millions of fibers per liter using the
formula in section 9 and substituting either X^ or XL
for the term F.
638
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Obtain the upper and lower values of the 95% confi-
dence interval for amphibole asbestos fibers and
total asbestos fibers by substituting the corres-
ponding values of X and S into equations (2) and (3) .
Report the precision of the analysis, in terras of the
upper and lower limits of the 95% confidence
interval, for chrysotile, araphibole, and total
asbestos fiber content. If a lower limit is found to
be negative, report the value of the limit as zero.
SELECTED BIBLIOGRAPHY
Seaman, D. R. and D. M. File. Quantitative Determination of
Asbestos Fiber Concentrations. Anal. Chem. 48(1): 101-110, 1976
Lishka, R. J. , J. R. Millette, and E. F. McFarren. Asbestos
Analysis by Electron Microscope. Proc. AWWA Water Quality Tech.
Conf. American Water Works Assoc., Denver, Colorado XIV-1 -
XIV-12, 1975.
Millette, J. R. and E. F. McFarren. EDS of Waterborne Asbestos
Fibers in TEM, SEM and STEM. Scanning Electron Microscopy/1975
(Part III) 451-460, 1976.
Cook, P. M. , I. B. Rubin, C. J. Maggiore, and W. J. Nicholson.
X-ray Diffraction and Electron Beam Analysis of Asbestiform
Minerals in Lake Superior Waters. Proc. Inter. Conf. on
Environ. Sensing and Assessment 34(2): 1-9, 1976.
McCrone, W. C. and I. M. Stewart.
10-13, 1974.
Asbestos. Amer . Lab. 6(4)
Mueller, P. K., A. E. Alcocer, R. L. Stanley, and G. R. Smith.
Asbestos Fiber Atlas. U.S. Environmental Protection Agency
Technology Series, EPA 650/2-75-036, 1975.
Glass, R. W. Improved Methodology for Determination of Asbestos
as a Water Pollutant. Ontario Research Foundation Report, April
30, 1976. Mississauga, Ontario, Canada.
Samudra, A. V. Optimum Procedure for Asbestos Fibers Identifi-
cation from Selection Area Electron Diffraction Patterns in a
Modern Analytical Electron Microscope Using Tilted Specimens.
Scanning Electron Microscopy, Vol. I, Proceedings of the Work-
shop on Analytical Electron
I llinois.
Microscopy, March, 1977.
Chatfield, E. J., R. W. Gl--. ;, :." . i M. J. Dillon. Preparation of
Water Samples for Asbestos Fiber Counting by Electron Micros-
copy. EPA Report EPA-600/4-73-0 I I , January 1978.
639
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Asher, I. M. and. P. P. McGrath. Symposium on Electron Micros-
copy of Microfibers. Proceedings of the First FDA Office of
Science Summer Symposium, August 1976. Pennsylvania State
Universi ty.
Chopra, K. S. "Interlaboratory Measurements of Amphibole and
Chrysotile Fiber Concentrations in Water." Journal of Testing
and Evaluation, JTEVA, Vol. 6, No. 4, July 1978, pp. 241-247.
National Bureau of Standards Special Publication 506.
Proceedings of the Workshop on Asbestos: Definitions and
Measurement Methods held at NBS, Gaithersburg, MD, July 18-20,
1977. (Issued March 1978).
*U.S. GOVERNMEKT PRINTING OFFICE: 1982-0-361-085/4458
640
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