vvEPA
            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
           Ore Mining and Dressing
            Point Source Category

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            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 nlstrator
          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
                                             cifon Agency
                          Wl«v,v.

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u,s
Agency

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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)






Secti on                                                     Page






          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
111-16
111-17
JII-18

111-19

111-20

111-21


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), 	 	
Page
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/
          LABORATORIES	   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
                              VI 1

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                     LIST OF TABLES (Continued)


Number                                                      Pa g e

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
          COP PER/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 ATI ON	  175

VI-6      DATA SUMMARY ORE MINING DATA SUBCATEGORY
          COP PER/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
                   ;i
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                                                      Page

          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                                                      Page

          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 ACID/ALKALINE MILL
                              XI 1 1

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                     LIST OF TABLES (Continued)


Number                                                      Page

          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
                               X ! V

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                     LIST OF TABLES (Continued)


Number                                                      Pa g e


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
                               xv

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                     LIST OF  TABLES  (Continued)


Number                                                      Pag e

          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
                               xvi

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                     LIST OF TABLES (Continued)


Number                                                      Page
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                                                      Page

          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

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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

VII1-3    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

VII1-7    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
                               XX7

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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, D.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 108 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.

NM SOURCE PERFORMANCE STANDARDS JNSPSJ

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 (18  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   (BOD5), 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), njo.dif.ied, 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 discharcrars 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.  Althouah 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 the 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. v.
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  best  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   methods
 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   subcategories 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
                                12

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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 sampling 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 wastewaters from ore mining and milling opera-
tions according to what were then considered key parameters-total
suspended  solids,  pH,  lead,  zinc,  copper,  and other metals.
However, little information on  other  environmental  parameters,
such  as  other  toxic  metals  and  organics, was available from
industry or government sources.  To establish the levels  of these
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 co.sts, 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.
                                 15

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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.   eler-rostatic separation

4,   flotation

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  paragraphs
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,

Gravity 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,  CaWCK).

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
                               18

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into layers by density within the jig, the lighter  gangue   being
drawn  off  at  the  top,  with the water overflow and the denser
mineral drawn off at a screen on the  bottom.   Often  a  bed  of
coarser ore or iron shot is used to aid the separation; the  dense
ore  mineral migrates down through the bed under the influence of
the agitation within the jig.  Several jigs  are  often  used  in
series  to  achieve both acceptable recovery and high concentrate
grade.   Jigs are employed in the  processing  of  ores  of   iron,
gold, and ferroalloys.

Tables

Shaking tables of a wide variety of designs have found widespread
use  as  an  effective  means  of achieving gravity separation of
finer ore particles 0.08 mm (.003 inch) to 2.5 mm (0.1  inch)  in
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  has  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  in  the  processing  of  ores  of  gold,   ferroalloys,
titanium, and zirconium.

Spirals

Humphreys  spiral  separators,  a  relatively recent development,
provide an  efficient  means  of  gravity  separation  for   large
volumes  of  material   between  0.1  mm and 2 mm (.004 inch to .08
inch) in diameter.  They have been widely applied,  particularly,
in  the  processing  of  heavy  sands  for  ilmenite (FeTi03_) and
monazite (a rare earth phosphate).   Spirals consist of a  helical
conduit  (usually of five turns) about a vertical axis.  A slurry
of ore is 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  (0.55 to 2.64 short tons) of ore per hour; 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  "Huntington-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
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grinding  mills.   Where  smaller  amounts  of  material  are handled,
wet and crossed-belt  separators are  frequently  employed.

Magnetic  separation  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
                              21

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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 (NaS2COC2H!>).  Frothers are generally  organics  with  a
soluble  group  and a nonwettable hydrocarbon.  Pine oil hydroxyl
(C6H]_20H),  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  nietals,  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
                              22

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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 acid leach)
are recovered by leaching processes.  Leaching is  also  used  in
processing ores of ferroalloys, radium, and vanadium.

Amalgamation

Amalgamation   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 + 2H20	
      2Na2Zn(CN)4_ + 2Au + H2_ f 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
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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  metallic  cation
increases with increasing valence:

     Cr i|| * 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
(uranyl, 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) eiffect  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  exchange  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
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these  procedures  has  been defined  in  a United  States Bureau  of
Mines  (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
quarrying, 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   all  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.

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
mining  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-piliar,   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
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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.

Drill ing

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  nitrate  mixed with about six percent fuel oil could be
detonated by a high explosive primer.  This new  application  has
spread to the point where virtually all open-pit mining companies
use  this  mixture  (called  ANFO) for some or all of the primary
blasting.
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Typicaly, multiple  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
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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 falling, 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.

Stopes

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   and
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 serve 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 fall  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 each
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   sublevels,  using fan blasts to break the ore.  Since
only the sublevels 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.
                             32

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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  ate  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  leachability.
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 nulling category.  A profile  of  the  mining  and
milling practices used in each subcategory follows.

Iron Ore Subcategory

General

American  iron  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  1976 to 77,957,500 metric tons (85,932,552
short tons) (Reference  4).   Iron  ore  shipments  decreased  to
54,918,000  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  88.3 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
industry uses milling operations which result
percent discharge to surface waters, and the
for 15 percent are unknown.   For pelletizing
percent  of  total  production is represented
ticing no discharge of process wastewater,  35
to  surface  waters, and the discharge practi
unknown.  A summary presented in Tables  Ill-
production  data,  processes  and wastewater
and discharge methods and volumes.
of the U.S. iron ore
 in no discharge, 31
discharge  practices
operations alone, 56
 by operations prac-
  percent  discharge
ces of 8 percent are
1   and  111-2  shows
technology employed,
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 pumped out of the mine.
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   uses   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 pellets.   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 of 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 coarse-grained hematite ores
have  proven  unsuccessful because of the  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.

Copper, 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):

                                 Production
      State      1000 Metric Tons 1000 Short Tons

       Arizona
       Utah
       New Mexico
       Montana
       Nevada
       Michigan
       Tennessee
       Idaho

   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):
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
      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   10   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
mining.

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 r.he 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  III-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 producers accounted for 73 percent   of
total  production  during  1975.   Approximately 95 percent of all
production came from 25 mines  or  mine/mill  operations,   10   of
which were operated primarily  for recovery of gold (Reference 8).
Thirty-six  percent of total  production was recovered as a bypro-
duct of other mining (for example, where copper or lead/zinc  are
primary  metals);   the  remainder  was recovered at gold lode and
placer operations.

Gold prices have risen 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  $800.00   (Reference  13)  per 31.1  gram (1  troy
ounce).   It is currently (15  March 1982)  $315 per  31.1   gram   (1
troy  ounce)   (Reference  9).    An  expected response to the  high
                                39

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price of gold is the increased gold prospecting  and  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 III-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 a 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  byproduct  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  39.1  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 90 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 by  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.
                              44

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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

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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  U308_   (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 (PO4_)2.. 1 0-1 2H.2O) .   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 U30£ (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  U3_0£  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.

Mining

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.
                               47

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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  USO!^)
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
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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 U3_08_ 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_0£ 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, 63!,000 metric tons (700,000  short  tons)  per  year,  and
680,00?  metric  tons  (750,000  short tons) of ore per year, and
seveu 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

-------
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) BaCl2_ 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

-------
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 wac 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  ilmcnitc 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.

-------
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, V205_, 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]_002J^) is concentrated by  solvent  extraction,
and  ammonia  is added to precipitate ammonium vanadate.  This is
dried and calcined to yield a V205_ 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              8f076
     1977                  6,866              7,565
     1978                  4,036              4,,446
     1979                  5,302              5,841
                                54

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-------
                           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-
categori^atirvn 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_ Mi.rie 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 instcinces,  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_  Benef iciation  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
separation,  froth flotation using reagents,  chemical extraction,
and hydr.metallurgy.   Physical processing using  water,  such   as
gravity separation,  discharge the suspended solids generated from
wasV.,ncf -.lr ,?dging,  crushing, or grinding.  The exposure of finely
  , ided  one  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

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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
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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  varies 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

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
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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 Ferroalloys—SIC 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
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ruined and milled.  Because  of  these  factors,  the  subcategory
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.
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TABLE IV-1.    PROPOSED SUBCATEGORIZATIOIM 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

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                            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 Ore—Mine/Mill 2120, Mine/Mill/Smelter/
           Refinery 2122, Mine/Mill/Smelter/Refinery 2121,
           and Mine/Mill/Smelter 2117
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         - Lead and Zinc Ore—Mine/Mill  3110, Mine/Mill/Smelter/
           Refinery 3107, and Mine/Mill  3121

         - Gold Ore—Mine/Mill 4105

         - Silver Ore—Mine/Mill 4401

         - Molybdenum Ores—Mill 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  Set  ]_.  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 Ores—Mine/Mill/Smelter/Refinery 2122 (two
           trips),  Mine/Mill Smelter 2121, and Mine/Mill 2120
         - Lead and Zinc Ores—Mine/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 Ores—Mine/Mill 4401 (asbestos only)
         - Molybdenum Ores—Mill 6101 (asbestos only)

     2.  Aluminum Ore Subcategory—Mine 5102 (total phenol and
         asbestos)

     3.  Uranium Ore Subcategory—Mine 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 Ores—Mine/Mill/Smelter/Refinery 2122 and
           Mine/Mill 2120
         - Lead and Zinc Ores—Mine/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 Ore—Mine 2110 and Mine/Mill 2116
         - Zinc Ore—Mine/Mill 3106
         - Lead/Zinc Ore—Mine/Mill 3113 and Mine 3117

     3.  Tungsten Ore—Mine/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:

     Uranium Ore Subcategory - Mine 9401, Mine 9402, Mine 9408,
          Mine 9411, Mine 9412, Mill 9404, Mill 9405, Mill 9414,
          Mill 9415, Mill 9416.

Gold Placer Mini_ng

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:

     1.  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
9907 and
Ore Subcategory
Mine/Mill 9910.
- Mine/Mill 9906, Mine/Mill
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 The  data  collected  have  been  summarized  in  a  compehensive  report
 on    titanium    sand   dredging  wastewater  treatment   practices
 (Reference  2).

 Solid 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/MilI/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  "l_,   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 O-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-irtl
(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 arid 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 points  (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-
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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  sophisiticated 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 plannnirig data generation for the  BAT  rulemaking,   EPA  con-
sidered requiring dischargers to monitor and analyse 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 arid 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
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           Chrysotile
     Conventionals
        Total Suspended Solids (TSS)
        pH
     Non-Convent ionals
        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   conventional,   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:
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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
conventional or nonconventional  parameters occurred.

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 analyzed for all toxic metals  (total  and   dis-
solved),  but not 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
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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 rr.etals 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 coriventionals,  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,  and 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, revisad 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,  B.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
         100 mg/J 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

                          TOXxC ORGANICS
Compound Name

  1.   *acenaphthene
  2.   *acrolein
  3.   *acrylonitrile  (V)
  4.   *benzene        (V)
  5.   *benzidene      (B)
  6.   *carbon tetraohloride (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.  *l,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)

  *Nltrophenols (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(Z-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.   *trichloroethylene   (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 and 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)

   *heptachlor and metabolites

100.   hF« 'achlor   (P)
101.   he"tachlor 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


   *polychlorinated 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 futpcndad tolidi
volatile tuip*nd«d tolidi
COO
TOC
radium 226 lion))
radium 226 (diuolved!
antimony (total)
•rstntc (total)
bery II rum (tout)
cadmium dcul)
chromium (total)
copper (total)
it ad (tola!)
mercury (loiaD
motel (tota!)
selenium (total)
ulver (total!
thallium (totaii
unc (toul)
atbetiot (fibroiii)
cyanide (total i
phenol (total)
aldnn
dwldrii
chlordane
(technical mixture and metabolites)
4. 4 '-DDT
4,4' DDE (pj>'DDX)
44' ODD (p,p' TD£)
ff -endoiulfan
/?-endo»ulf»n
«ndovj'*an tuifate
cndnn
endrtn aldehyde
heptachlor
heptachlor epoxide
a BHC
/3 BHC
y BHC (iindane}
6 BHC
PCB 1242 lArochlor 1242)
PCB 1254 (Arochlor 1254)
acenaphlbene
acrclein
acrylomtnie
ben/cne
benttdme
carbon tetrachlonde
(teuachlotomethane)
chloroben/ene
1^4 trichloroben/ene
hexachloroben^ene
l^^ichloroethane
111 inchloi-oethane
hexachiofoethane
V 1 -dichloroethane
1 1,2 tnchloroelhane
1122 tetrachloroetham-
chlo>oethane
bis (chloromtthyl) eihet
bis (chloioethyll eihet

2^hl0rOndphthalene
246 (richlorophenol
parachloromtia crpso!
chloroform (trichtoromelh^ne)
2e
I,4^tchloro6mzwte
3,3'-dichloroh*nz*d.nt
1,1-dtchlOfocthyleflc
1 ,2 trant-dichlorottnyleiM!
2.4-dichk>rophtncH
1 ,2-dtchioroprop«ne ,
l,3-d»chlofopropytent
(1 .3-dtchlofoprop*frt)
2.4^imethytph«rtol
2,4-dmitrotoluerw
2,6-dmitrototuene
1 ,2-diphenylhydruine
•thylbenzene
fluoranthene
4-dilorophcnyl phenvl tther
4-bromophenyl phenvl ether
bn (2 chtOTOtiopropyl) ether
bit (2-chloroethoMv) methane
methytene chloride (diohltHomethane)
methyl chlortde ( ch 1 or omn thane!
methyl bromide (bromom«thin0)
bromoform (trtbrpmomethane)
dichlorobromomethane
tnchtorofluorometharte
dichlorodi'fuorcromeihand
chlorodibromomethane
hexachlorobutadiene
hex achtorocy dope ntadiene
isophorone
napthalene
nttrobenzene
2 nitrophenol
4-nitrophenol
2.4 dinitrophenol
4,6f)
"
PT
PT
LL
LL
PT
PT
LL

I 3d benzof luw a thene
11 12btfn/ofluofanthene
! rt,y«nc
If dcenaphihylene
|) anthracene
][ 1 12 ben/operylene

I, phenanthrene

PT
PT
PT
LL PT
LL

LL
LL
LL
PT
LL
iiukno 11,2,3 c,d) pyenc
_|^PV"~
• 2378 ifirdctilofodiben^o p-dioxm
. , (TCDDJ
• t.-irdchloro-Ihylem.
tOlUI Ml-
tnchloioethvlfnf
vinyl chlonrir
to.dphf n.
"

1
Ll t ----,'
1 ANALYSIS TECHNIQUE/LAB j
LL Guf South RtMvch intt.
LL
LL
PT
PT
LL
PT
PT
LL
LL
LL
LL
PT
LL
LL
LL
LL
LL
PT














LL







L

















































































1















i




J T



PT
PT
PT
PT
LL










             TEM
             SM
Atomic AtMorpiion »p*cuojcopy (flame or flamcleii)
Gat Chromato!}rapfiy b/ Electron Capture detection method
Gas Chromarography/Maii Spectrometry (GC MS) by Liquid Liquid E
Emmion Spectroicopy utrng inductively coupled argon platma (Plairru
Gat Chromjtagraphy/rVlais Spectrometry iGC-MS) by Purge and Trap
RadioaiUy uung icintilration counter or proportional counter
Trammimon Electron Microscopy
Other itandard (EPA approved) .'.t.fiods
                                                               .iracTion melhod
                                                               Afc method)
                                                              method
                                                                  148

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      TABLE V-4.  LIMITS OF DETECTION FOR POLLUTANTS ANALYZED
SUBSTANCE
pH
total suspended aolidt
•ototile Hispended «ol«h
COO
TOC
radium 226 (total)
radium 226 (dMOhedl
antimony (total)
ertamc (total)
beryllium (tout)
cadmium (total)
chromium (total)
copptr (total )
l«d (total)
mercury (tout)
nickel haul)
selenium (total)
tilver (total)
thallium (total)
zinc (total)
aibettot (fibrous) (total)
cyanide (total)
phenol (total)
aldrin
diefdrm
chtordane
ttechmcal mixture and metabolites)
4.4'.DDT
4.4' DDE (pj)' DDX)
4.4'.DDD(p.p-TDE)
OT -endotulfan
Q -endosulf an
endosulfan sulfate
endrin
endrm aldehyde
heptaehlor
heptachlor epoxide
a BHC
0-BHC
7 BHC (hndane)
S-BHC
PCB 1242 (Arochlor 1242)
PCB 1254 (Atochloi 12541
acenaphthene
acrolem
acrylonnrile
benzene
benzidme
carbon letrechlonde
(tetracMoromethene)
chlorobenzene
1 ,2.4-tnehtorobenzene
hexachlorobr-nzene
1 ,24ichloroethane
1 .1 .1 -trichloroethane
hexachloroethane
1 .1 -dichloroethane
1,1,2-trichloroethane
1,1,2,2 letrachloroethane
ctuoroethane
bis (chtoromethyl) ether
bis (chloroethyl) ether
2-chloroethyl vinyl ether [mixed)
2-chloronaphthalene
2.4.6-trichlorophenol
parachloiometa cresol
chloroform (tnchloromethane)
2-chlorophenol
1 ,2-dicblorobenzene
CONCENTRATION
-
1m«/l
1m»M
2n(/l
Ing/1
1pCi/l
IlKSfl
0-2n>»'l
0.002 mg/l -
0.005 mj/t
O.f»2mj/l
OJOmg/l
0.01 mg/l
0.06 mg/l
0.0005 mg/l *•
0.02 m,/l
0.002m8/l'
0.01 mg/l
0.1 mg/l
O.OOSmj/l
2.2x10*fibert/l
0.02 mg/l
0.002 m»/l
0.1 tig/I
Oj|ig/l

1vg/l
1«H/I
1*8/1
1*8/1
i«g/i
1 Wg/l
iMg/i
0.5*9/1
0.5 fig/1
01*g/l
0 1 *g/l
0.1 *g/l
0.1*8/1
0.1 *g/l
0.1 *g/l
1*8/1
1*,/l
0.03 jig/1
no enimate
no estimate
0.04 iig/l
- 25.0*8/1

0 35 jjg/l
O.tS U9/I
~ 10*a/l
~ 1 0 ua/l
0.08*8/1
0 15 ug/l
~ 010*a/l
0.15 *9/l
0.10 *8/l
0.90 *g/l
~ 0.50 (ig/l
no estimate
no estimate
~0.50*e/l
0 05 *g/l
~ 1.0 *»/!
~ 1 0 *«/!
0 05 fjg/l
1 00 *9/l
0100.20*8/1
SUBSTANCE
1,3-dKhlorobenzm
1,4<)ichlorobaniane
3J'e
I.Uichloroethylene
1 ,2.tran*dKhloroettrylene
2.4^lichlorophenol
1 ^^lichlorapropane
1>dchk>ropropylane
(1 .SJiehloropropernl
2,44)imethylpnenal
2,44initrotoluene
2.6-dmitrotoluene
1 ,24liphenvlnvdraiine
ethylbenzene
fluorantttana
4-chlorophenyl pnenyl ether
4^romophanyi phenyl ether
bis (2-chlorotsopropyf) atfier
b» (2-chloroethoxy) methane
methylene chloride (dichloromethane)
methyl chloride (chloromethane)
methyl bromide (bromomethane)
bromoform (tnbromomathane)
dichlorobromomethane
trichlorofluoromethane
dichloredifluororomethane
chlorodibromomethane
hexachlorobutadiene
hexachlorocyclopcntadiene
isophorone
napthalene
nitrobenzene
2-mtrophenol
4^itrophenol
2,4Hlinitrophenol
4.6-dimtro.o.cresof
N^iitrosodimethylamme
N^iitrosodiphenylamme
N^nitiosodi-n^iroovlarmne
pentachlorophenol
phenol
b» (2«thylhi>xyl) phthalate
butyl benzyl phthalate
dt n-butyl phthalale
diethyl phthalate
dimethyl phthalale
1 .2 benzanthracene
benzo (a) pyiene (3.4-benzopynme)
3,4- benzofluorathene
11.1 2*enzofliK>rantri™«
chryserw
acenaphthylene
anthiaceite
1,12^en;operylene
fluoirne
phenaothrene
1 2 5,6^iiberizanthracene
indeno (1.2.3^,d) pyrene
pyrene
2.3.7.8 tetrachlorodibenzo-pxiioxin
1TCDDI
tetradi loroethylene
toluene
trichioroethylene
vinyl chloride
toxaphene



CONCENTRATION
0.1MJO Wg/l
0.10-0.20 uttt
~».0«g/l
0.50 ug/l
0.36^11/1
~ 1.0/lu/l
0.02 iia/l

OJSntil
0.40 uo/l
0.20*8/1
0.7B ,,g/l
0.20 (ig/l
O.10 ug/l
0.20 ug/l
0.06 ug/l
~ 0 50 ug/l
m estimate
noettimata
0.08 ug/l
0.35 ug/l
0.08*8/1
0.40 *a/l
0.05 ug/l
- 0.10*8/1
~ 0.10*0/1
~0.10*o/l
0.08 ua/l
noaitimata
0.10 ug/l
0.15 uo/l
0.85 atg/l
1.0*0/1
no aitimate
wilt not chromatograph
wiH not chromatograph
not establuhed
not established
not attablithed
~ 50^ ug/l
1.0 ug/l
0.20 jjg/l
0.25 ug/l
0.3-0.4 ug/l
0.20 *«/!
0.3S ug/l
0.05 ug/l
- 0.5 ug/l
- 0.5 in/1
~ 0.5 ug/l
0.20*0/1
~ 0.5 |ig/l
0.05 uo/l
•- 0.5 ua/l
010 ug/l
0 06 U9/I
~ 0.5 ug/l
~ 0.5 ug/l
0.40 u 9/1

no estimate
1.10 ug/l
0.35 ug/l
0.35 K 8/1
~ 0.1 ug/l
not anabhshfld



•As cnetiyzcd by slomic abiorptton sptctrotcopy using gtueout hydrtdt (a f(*me method)
        y »tor"« ataocptioo KMCtroscopy min; cold vapor technique U fUmeltm method)
                                           149

-------
                            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

-------
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
    *AII samples collected at the same time at Mine/Mill 3121
                           151

-------
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/I ON
as NaCN
Distilled H-O + 0.2 mg/l CN
as NaCN
Distilled H,O + 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
1.0 mg/l CN as K3Fe(CN)g
Tailing Pond Decant Spiked with
1.0mg/ICNasK4Fe(CN)g
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
'All cyanide analyses are total cyanide
                                152

-------









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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

-------
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, Mine 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 Mill
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 Mill 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/Zinc/Silver/Gold/Platinum/Molybdenum     Subcategory,
Mill Subdivision,  Froth 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/Zinc/Gold/Silver/Platinum/Molybdenum     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 pcnds.

Aluminum Subcategory, Mine Drainage Subdivision

As shown in Table VI-9, aluminum mine drainage  is  lew  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 Subcateqory,  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,  Mine 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 Subcateqory,  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 wit_h 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

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Vanadium Subcategory, 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.

AjTt_irnp_ny_ Subcategory, Mi 1JI 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 IN 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.   Frother—A 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

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     4.  Activator—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 refloat a depressed mineral.  A
         good example of an activating agent is copper sulfate,
         used in the flotation of sphalerite.

     5.  Depressant—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).  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

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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,  C2H5..C6H40H  or   (CH3_) 2. C6H3_.QH,  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 Mi 11 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  mg/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

-------
































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                                         193
    

    -------
                               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, y_._ 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
    sampling and analysis programs.
                               these
          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  these  17 compounds  are subject  to  exclusion
    under  Criterion 3.
    
    Pollutants  Detected   But  Present   i_n  Amounts   Too   Low   to   be
    Effectively Reduced by_ Known Technologies
    
    Toxic  Organic Pollutants
    
    There  were 10 organic  pollutants detected during   the r\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
    pollutants  are  thus  excluded from regulation because  they were
    present in amounts too low to be  effectively  reduced   by   tech-
    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
    

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    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
    available  which  indicate effluent beryllium levels of
    are possible by lime precipitation  and  multiple  pond
    (Reference  10).   Of  73  effluent  samples  analyzed
    screening for beryllium,  only 10 samples had detectable
    concentrations  with  a  mean   of   0.005   mg/1   and
    concentration " of  0.011   mg/1.    These  are the levels
    achieved by known technologies and since the pollutant
     data  are
     0.09 mg/1
      settling
    during BAT
     beryllium
       maximum
     which are
    is present
                                         200
    

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    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  (FeCr20«),  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  ir\  Treated Effluents at  a  Small   Number   of
    Facilities  and Uniquely Related tp_  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
    

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    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.
    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
    

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    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
    

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    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
    

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                               Table VII-4
    
                     TUBING LEACHING ANAYSIS RESULTS
    Component
    
    Bis (2-ethylhexyl) Phthalate
         Acid Extract
         Base-Neutral Extract
    Phenol
                                  Micrograms/Liter
    
                              Origina.1            Tygon
         Acid Extract
         Base-Neutral Extract
                                915
                              2,070
                             19,650
                               N.D.
    N.D.
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    N.D
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    N.D.
    Not Detected
                                     214
    

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                              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 mills.
    
    Control of_ Cyanide
    
    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 212M
     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
    wastewater.
    
    Alternatives t£ Cyanide Ir\ 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  monosulfi.de  (Na2S),  sodium sulfite (Na2S03_), and sodium
    thiosulfate (Na2S^03_).  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 nonphenolic frothers arc methyl isobutyl carbinol (MIBC) and
    polyglycol methyl ethers.   Frothers are  generally  nonselective,
                                   220
    

    -------
    generically   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 o_f_ 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 oe  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
    

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     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 profitability—are substantially improved.    In
     addition,  costs  of   pretreatment of process water—and,  in some
     instances, reagent use—may 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-
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    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  2~126           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):
    Parameter
        Effluent  Levels  (mg/1)
     Attained  Prior  to Use
     of  Polymer and Secondary
    	Settling  Pond*	
                      Effluent Levels (mg/1)
                   Attained Subsequent to
                    Use of Polymer and
                  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.
                                       231
    

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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  to  involve  both  adsorption  and
    filtration within  the carbon bed,.
                                       233
    

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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(CN2),  and iron-cyanides  (Fe(CN)«-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 + CN- + 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
       SNaOCl  +  2CuCN +  2NaOH + H2_O	2NaCNO  +  2Cu(OH)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- + 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 + Nz + 302
         CNO-  + OH- + H20 	>  Co3 -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 (H202_) 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 stoicniometric 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-electrocoajgulation 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 .-/anide concentration of approximately 60 to 70 mg/1.
    
    Tr^;.  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
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    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
                                         240
    

    -------
    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
    complicate:; the theoretical prediction of   phenol   concentrations
    in flot~\Ion-mill wastewater streams.
    
    A  more del .riled account of reagent  use, including phenolic-based
    LX.i?ipounds and  their  presence  in  mill  process  wastewater,    is
    provided in Section VI,  Wastewater Characteristics.
    
    Several  methods  are  available  for  treating  phenolic wastes,
    including:
                                       241
    

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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   (C102^  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 (C6H40;2)
    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 (H202.)
    effectively oxidizes phenols over a wide  range  of  temperatures
                                        242
    

    -------
    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 H202_ 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 (KMn04_).  Phenol reductions to less than 1
    mg/1 are reported.   The destruction of  simple phenol by potassium
    permanganate is expressed as:
    
       3C6H60 + 28KMn04 + 5H20	1 8C02_  + 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).
                                    243
    

    -------
    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.
    
    Phenol  Treatment Practices.  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:
                                      244
    

    -------
         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  H2.S,   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 cnloralkali 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/1.   Although
    lead is also present in the waste streams treated  with  sulfide,
    its  concentration  is not often measured.   The limited available
                                    245
    

    -------
    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  microgram  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.
                                   246
    

    -------
    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
          form fibers.
    filtration is successful in removal of asbesti-
         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
    scale  studies,  pressure filtration and vacuum f
    through both filtration  systems  ranged  from  0
    m3/sec  (10  to 20 gal min).  Both kinds of DE fi
    ated in various ways to evaluate conditioning of
    cationic polymers, and anionic polymers.  Single-
    precoat were studied.  Conditioned DE was used in
    as  well  as  for  body feed.  Various grades of
    coarse, were evaluated.  Details of pilot-plant
    reported in References 41,  42, and 43.
                                    in  the  pilot-
                                   iltration.   Flow
                                   .0006  to  0.001
                                   Iters were oper-
                                    DE  with  alum,
                                   step and twostep
                                    filter precoat,
                                   DE,  from fine to
                                   DE   testing  are
    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
    

    -------
    that pressure DE filtration is  effective  (i.e.,  reduces  fiber
    content  to  4  x  104 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-m3/day  (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
                                    248
    

    -------
    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/min/ft2).
    
    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 TO9  fibers/liter  in  the  filter
    influent   to   less  than  detectable  (approximately  3  x  TO5
    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 VI11-5 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  in 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  S04~2.   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  10*  to  10s  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 TO3
    to TO4 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
                                   251
    

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    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  ^9/1 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 anion 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., H2_S, 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  (picograms)/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).
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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  f locculati.on,  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/1iter  (total)
    and  93.4  picocuries/1iter  (dissolved) to 7.18 picocuries/1iter
    (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
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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
         7/24 and
         7/28 and
         7/29 and
         7/31 and
         8/1 and
         8/5 and
         Average
     7/25
     7/29
     7/30
     8/1
    8/2
    8/6
    Mo(mq/2 )
    Feed
    20.
    23,
    22.
    24.
    19.
    22.
    22.
    5
    0
    4
    4
    5
    0
    0
    Effluent
    1
    0
    1
    1
    1
    1
    1
    Eluate*
    . 18
    .91
    ,38
    .76
    . 14
    .38
    . 29
    16, 140
    16,045
    16,568
    18,090
    12,930
    17,484
    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
    economical recovery of the molybdenum,   defraying  a
    fraction   (or  possibly  all)  of  the   costs of the
    operation.  On  the   basis   of  pilot-plant  testing
                                            eluate allow
                                             substantial
                                            ion exchange
                                             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)
    Feed Effluent a Effluent b Effluent c
    pH (units)
    Iron
    Manganese
    Zinc
    Copper
    Cadmium
    Cyanide
    25 to 35
    6.3 to 6.6
    1.4 to 1.6
    0.59 to 0.74
    0. 03 to 0.04
    0. 22 to 0.33
    8.5
    1 .9
    1 .6
    0. 1
    0. 15
    0.01
    0. 13
    9. 2
    0.6
    0.5
    0.04
    0. 10
    0.02
    0.07
    10.4
    0.8
    0. 1
    0.04
    0.09
    0.01
    0.06
    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.  Chlorination 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.
                                     257
    

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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 VII1-15.
    
    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                    0.25                  0.24
               Zn                    0.37                  0.26
               Cu                    0.05                  0.04
               Fe                    0.28                  0.22
    
         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.
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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 (mg/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
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     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)*    5.0      5.5       5.0       5.0      5.0
         pH (Stage 2)*    6.4      6.4       5.5       5.6      6.5
         Fe (total)       ND       ND        ND        ND        ND
         Zn               12.7      0.3     30.0     30.0      0.2
         Mn               6.4      0.5       6.8       7.1       0.4
         Cu               ND       ND     LT 0.5   LT 0.3      ND
         Al               ND       ND        ND        ND        ND
         Ni             LT 0.13    0.13      0.29      0.19      0.13
         Cr               ND       ND        ND        ND        ND
    
         *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 VIII-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  Mine/Mill/Smelter/Refinery  3107.   Wastewater streams
    generated from mining, milling, smelting, and refining activities
    at this lead/zinc complex are combined in  a  common  impoundment
    pond,  and the effluent from this pond is subsequently treated in
    a  physical/chemical  treatment  plant  by  lime   precipitation,
    aeration,  flocculation,  and  clarification, 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
                                  267
    

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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  51 Q_2.  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.
    
    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   (MoO*-2),   selenite    )Se03-2),    and
          vanadates  (HV04-2,   VO*-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 m3/in2/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 cwo-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,
    Zn,  Ni,  Cr,  V,  Mo,  Fe, Al, and Mn.  Addition of  lime to form
    metal hydroxide precipitates is known to be an effective  removal
                                     270
    

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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 et al_  (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 (VO,)-» + 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 coprecipitation of radium 226, respectively.  A summary
    of the study results is presented in Table VIII-35.
    
    Generally,  pH values  greater  than  the  range  8.2  to  9  were
    required  for  optimum  removal of TDS and most metals.  However,
    optimum 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.
    
    Uranium/Vanadium 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.t
    This  facility  achieves  no  end-of-pipe  discharge  of  process
    wastewater by:  (1) 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
    (Se03-2),  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
                                      272
    

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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 (S04), Bad2, 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 rfaove from wastewater.
    
    Enhan : :.;   netals  removal  was  observed  under  all  conditions
    studied,  ,,t,\.imum removal of metals was achieved at  the  highest
    racio  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-
                                        273
    

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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/Mi11/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/Mill/Smelter/Refinery 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 Scimpling 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 Subcateqory
    
    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,  Zinc,  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)   of1
    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 VIII-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 VII1-47,  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
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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.  Galenaf 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 VIII-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 VIII-49 and VIII-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-50, 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 arc 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 VI11-53.   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 has an underground mine.
    The minerals of   principal  value  are  galena,  sphalerite,  and
    chalcopyrite.   Zinc,   lead,  and copper concentrates  are produced
    by the flotation  process in the mill.   In 1976,  mine  production
    totaled  972,300  metric tons (1,072,400 short tons),  while 92,400
    metric tons  (102,000 short tons) of lead concentrate,  and  9,800
    metric  tons   (10,800  short  tons)  of  copper  concentrate were
    produced at  the mill.
    
    Mine and mill wastewater streams are combined  for  treatment  at
    this facility,  as indicated in Figure VIII-10.   Wastewater treat-
    ment  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 B.lhectare
    (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 VI11-58 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.
    
    Table  VIII-59 summarizes effluent monitoring data for the period
    January  1975  through  December  1976.    Values   reported   are
    substantially  below  BPT  limitations.   Recently, this operation
    used  hydrogen peroxide  addition  to  the  tailing  pond  decant
    stream  during  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
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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 Ore 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
     (A1203.) 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  VlTI-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 Subcategory
    
    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  oi"  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 Subcategory
    
    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 chlorination 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  m3/m2/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  pond
    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   VI11-70.  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  VII1-78.   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 ozonati'oh
    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 at 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 dredge 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 VII1-79.   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/l/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
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    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
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    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
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     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 or 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,  arid 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
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    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  (montmorillonite),  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
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    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
         en-':'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.
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    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 IP', il in Eastern Ontario,  Canada,  is controlled by a grout
    curtai",  rronstructed of clay, bentonite,   and  cement  (Reference
    70),   "rtK1 grout slurry was injected into the subsurface alluvium,
    do'-fn  to  an  impervious  bedrock layer,  and forms an underground
    tarrier 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
    

    -------
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    -------
    TABLE VI11-2.
    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 SODIUM 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/i
    
    0.08
    0.05
    0.07
    10.6
    10 mg/l
    
    0.04
    0.03
    0.04
    20 mg/l
    
    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-TREATMEISIT ALTERNATIVES
                   FOR REMOVAL OF CHRYSOTILE AT PILOT PLANTS
    TREATMENT METHOD
    Sedimentation
    Sedimentation plus Mixed-Media Filtration*
    Sedimentation plus Uncoated-Diatomaecous-
    Earth Filtration
    Sedimentation plus Alum-Coated-Diatomaceous-
    Earth Filtration
    FIBER CONCENTRATION (fibers/liter)
    Raw Water
    4x 1012
    4x1012
    4x1012
    4x1012
    Treated Water
    5x1011t to1x1011**
    1 x109
    3x106
    1 xlO5
           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
    Q*
    5x 109
    Q*
    5x 10a
    Treated Water
    9.3 x 109t
    Q»*
    3.2x10a
       Source: Reference 44
       Corresponding turbidities are:
         •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
    3x 107
    3x 106
    8x 104
          Source:  Reference 44
                                         311
    

    -------
    TABLE VIII 7.  EFFECTIVENESS OF WASTEWATER-TREATMENT 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(1x1011}
    1 x 109(1x 1011)
    1 x 109
    1 x109
    Treated Water
    1 x109(1x1010)
    1 x 108 (1 x 109}
    2x106
    < 1 x 105
        Source: Reference 44
    
        Parentheses enclose results for a second sample.
                                            312
    

    -------
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                                           316
    

    -------
    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
    10.3
    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
    0.008
    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
    
    SAMPLE
    Mine +
    barren
    leach
    water
    
    
    SAMPLE
    
    Mine +
    barren
    leach
    water
    
    
    TREATMENT
    PH
    6.1
    7.7
    
    9.8
    11.5
    
    TREATMENT
    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 Vlil-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
    
    
    
    Tailings (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 I 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
    (0.14 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 1
    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
    (017 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
    (004 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
    

    -------
    TABLE VIII-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 and not always specified.
    N.A. = Not analyzed
                                                   324
    

    -------
    TABLE VIII-18. EPA-SPONSORED WASTEWATER TREATABILITY STUDIES CONDUCTED
                  BY CALSPAN AT VARIOUS SITES IN ORE MINING AND DRESSING INDUSTRY
    SITE IDENTIFICATION
    Mine/Mill 3121 (Pb/ZN)*
    Mine/Mill/Smelter/
    Refinery 3107 (Pb/Zn)*
    Mill 2122 (Cu)*
    Mine 31 13 (Pb/Zn)'
    Mine 5102 (Al)
    Mill 9401 (U)**
    Mill 9402 (U)**
    	 	 	 ,. ....— .-.
    PERIOD OF STUDY
    August 3-13, 1978
    March 19-29, 1979
    August 14-1 9, 1978
    September 5-15,1978
    January 8-19, 1979
    September 22-29, 1978
    October 10-16, 1978
    October 23-30, 1978
    November 1-10, 1978
    December 4-1 4, 1978
    COMMENTS
    Polishing Treatments (Filtration, Secondary
    Settling), Lime Precipitation and Cyanide
    Destruction Technology (Ozonation, Alkaline
    Chforination) Investigated
    Polishing Treatments (Filtration, Secondary
    Settling) Investigated
    Polishing Treatments (Filtration, Secondary
    Settling), Lime Precipitation and Cyanide
    Destruction Technology (Ozonation, Alkaline
    Chlorination) Investigated
    Lime Precipitation, Aeration, Polymer Addition,
    Floc.culation, 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
    Copiecipitation, 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 Vill-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
    
    
    
    
    
    
    
    
    
    
    
    
    r
    
    
    "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.
    
    
    
    
    
    
    
    
    
    
    
    
    
    T
    
    
    RANGE
    n.
    
    
    
    
    
    
    
    
    
    
    
    
    
    a.
    
    
    
    
    
    
    
    
    
    
    
    
    
    t
    
    !
       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
    <0.0050
    0.016
    0.0070
    0.022
    0.19
    0.036
    0.22
    0.024
    0.0005
    <0.020
    <0.0050
    <0.020
    <0.10
    2.0
    0.16
    0.55
    0.022
    0.079
    RANGE
    8.8-9.1
    5-14
    -
    -
    -
    0.01 5-0. 02(1
    <0. 0050-0. 0010
    <0. 020-0. 030
    0.15-0.23
    0.020-0.050
    0.11-0.30
    <0. 020-0. 030
    <0. 0005-0. 0010
    -
    -
    -
    -
    1.4-2.6
    0.080-0.24
    0.24-0.78
    <0. 020-0. 030
    0.040-0.125
                                      327
    

    -------
    TABLE Vlll-21.  OBSERVED VARIATION WITH TIME OF pH AND COPPER AND
                 ZINC CONCENTRATIONS OF TAILING-POND DECANT AT LEAD/ZINC
                 MINE/MILL 3121
    DATE
    (1978)
    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/U
    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)
    PH*
    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**
    Kl to
    2)
    <1«*
    K1-1)
    Cu
    0.11
    0.05
    0.02
    0.03
    0.02
    0.02**
    K0.01 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 concentrations are based on "total" analyses
          t Value in pH units
         ** Average concentrations attained
          ( 'Range of concentrations attained
         n.a. = Not analysed
                                          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/D*
    PHt
    8.9-9.0
    10.5
    10.3
    TSS
    7
    (3-10)
    24
    (20-29)
    2
    «1-3)
    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
    **A1I metal concentrations are based on "total" analyses
                                        330
    

    -------
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    331
    

    -------
    TABLE VIII-25.  EFFLUENT FROM LEAD/ZINC MINE/MILL/SMELTER/
                   REFINERY 3107 PHYSICAL/CHEMICAL-TREATMENT PLANT
    PARAMETER
    PH«*
    TSS
    Cd
    Pb
    Zn
    Hg
    CONCENTRATION (rng/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.
                       * All 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
    
    
    
    
    
    
    
    
    
    
    
    i
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    "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.
    
    
    
    1
    0.036
    n.a.
    0.021
    0.073
    n.
    
    
    
    '
    a.
    
    
    
    t
    0.055
    
    
    RANGE
    n.
    
    
    
    i
    a.
    
    
    
    r
    0.025-0.050
    n. a.
    0.020-0.030
    <0. 02-0. 17
    n.
    
    
    
    a.
    
    
    
    1
    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*»
    «D
    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
        t Value in pH units
        ** Aver age 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
    < 0.005 to 0.20
    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
    I
    POLLUTANT
    PARAMETER
    pH
    TSS
    Sb
    As
    Be
    Cd
    Cr
    Cu
    Pb
    Hg
    Ni
    Se
    Ag
    T1
    Zn
    Fe
    so4 =
    NUMBER OF
    OBSERVATIONS
    53
    7
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    1
    "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
    
    -------
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                                                    337
    

    -------
    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
    X
    n
    
    
    
    
    
    
    
    
    
    
    
    
    1
    .a.
    
    
    
    
    
    
    
    
    
    
    
    
    f
    <0.01
    <0.02
    <0.02
    n. a.
    RANGE
    n.
    
    
    
    
    
    
    
    
    
    
    
    
    i
    a.
    
    
    
    
    
    
    
    
    
    
    
    
    P
    <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 VIII-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 to pH ~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**
    «1 to!)
    < 1»*
    «1 tol)
    Al
    0.49**
    « 0.2 to 0.70)
    <0.2
    <0.2**
    «0.2)
    <0.2
    0.2
    0.2
    0.3
    <0.27**
    « 0.2 to 0.4)
    <0.2**
    « 0.2 to 0.2)
    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 WASTEWATER)
    POLLUTANT
    PARAMETER
    pH
    TSS
    Sb
    As
    Be
    Cd
    Cr
    Cu
    Pb
    Hg
    Ni
    Se
    Ag
    T1
    Zn
    V
    Mo
    Fe
    Mn
    Al
    so4
    TDS
    Ra226
    U 1
    PERIOD OF
    OBSERVATIONS
    (DATES)
    11/3-8/1978
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    I
    NUMBER OF
    OBSERVATIONS
    5
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    r
    3/4
    , vl
    "TOTAL" CONCENTRATION
    (mg/l)
    MEAN
    X
    1 .7
    40
    <0.3
    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
    21750
    --
    99.7+1%
    17
    RANGE
    i .6-1 .9
    24-48
    <0.5
    1 .0-3. 3
    0.03
    0.04-0.05
    0.46-0.82
    2.7-6.0
    0.65-1.1
    <0.0002
    1.3-1.6
    1.1-2.6
    <0.1
    <0.2
    4.8-7.5
    80-110
    9-12
    1800-2000
    100-130
    640-900
    13,400-34, 300
    --
    (62. 6- 133) -11
    (12±9%) to
    (20*12%-)
    "DISSOLVED" CONCENTRATION
    (mg/l)
    MEAN
    X
    --
    --
    --
    --
    --
    --
    0.66
    3.7
    0.88
    --
    --
    --
    
    --
    5.8
    97
    8.9
    1860
    116
    758
    
    51, 520
    88 . 3i 1 %
    16
    RANGE
    --
    --
    --
    --
    
    !
    0.44-0.76
    2.6-5.8
    0.61-1.0
    --
    --
    --
    --
    --
    4.8-7.1
    77-110
    6.3-11
    1800-2000
    95-125
    640-890
    --
    :.H, 100-35,300
    (SH.4-l27~)-l°t,
    (12±9"«) to
    (20±12%)
                                          343
    

    -------
    TABLE VIII-34. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
                TREATMENT TRAILER AT MILL 9402 (ACID LEACH MILL WASTEWATER)
    POLLUTANT
    PARAMETER
    pH
    TSS
    Cr
    Cu
    Pb
    Ni
    V
    Mo
    Fe
    Mn
    Al
    so4
    TDS
    Ra226
    U
    
    PERIOD OF
    OBSERVATIONS
    (DATES)
    12/5/78-
    12/11/78
    
    
    
    
    
    
    
    
    
    
    
    
    
    !
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    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, 3*f
    157
    1,330
    
    32,200
    129±1%
    2o±m
    
    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
    (12lil%)-135ti
    (16±12%)-(24±12
    
                                     341
    

    -------
    TABLE VIII-35. SUMMARY OF WASTEWATER TREATABILITY RESULTS USING A LIME
                 ADDITION/BARIUM CHLORIDE ADDITION/SETTLE PILOT-SCALE
                 TREATMENT SCHEME
    PARAMETER
    IDS
    Cu
    Pb
    Zn
    Ni
    Cr
    Fe
    Mn
    Al
    V
    Mo
    Ra226
    
    U
    NH3
    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
    96to>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
    123292
    Dissolved
    5590-9740
    0.060-0.15
    < 0.014
    <0.020-0.030
    < 0.040
    < 0.040
    0.10-1.0
    0.43-4.2
    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 VIII-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
                                         343
    

    -------
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    (60 mg/l), Flocculation,
    Sedimentation, Filtration
    
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    IX, Fe (SO4) (330 mg/l).
    Aeration, Non-Ionic Poly
    BaCI2 (15 mg/l), Flocculati
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    3?
    M
    
    
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    IX, Fe(S04) (500 mg/l),
    Aeration, Non-Ionic Poly
    BaCI2 (120 mg/l), Floccula
    Sedimentation, Filtration
    •a
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                                                         344
    

    -------
    N
    LU
    •z.
    cc
    LU
    
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    CO
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    <
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                O 5X>
                                                  345
    

    -------
    TABLE VIII-39. CHARACTERIZATION OF INFLUENT TO WASTEWATER TREATMENT
                PLANT AT COPPER MINE/MILL/SMELTER/REFINERY 2122
                (SEPTEMBERS-?, 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
    <0.03
    0.14
    RANGE
    2.4-2.55
    26-790
    13-26
    0.015-0.016
    4.8-5.40
    < 0.0005
    0.076-0.120
    0.098-0.190
    8.40-10.0
    5.0-6.10
    0.010-0.017
    0.130-0.190
    0.034-0.037
    0.016-0.039
    <0.002-0.006
    1.50-2.10
    < 0.02-0.05
    0.04-0.3
    DISSOLVED CONCENTRATION
    (mg/l)
    MEAN
    —
    -
    11.1
    0.014
    5.10
    < 0.0005
    0.102
    0.134
    3.76
    3.66
    0.002
    0.17
    0.050
    <0.021
    0.005
    1.8
    -
    -
    RANGE
    —
    -
    9.4-13
    0.013-0.016
    4.30-5.80
    < 0.0005
    0.081-0.140
    0.094-0.160
    3.60-3.90
    3.30-4.0
    0.002-0.003
    0.16-0.18
    0.035-0.077
    <0.01-0.027
    0.004-0.006
    1.6-2.10
    —
    -
                                 346
    

    -------
    TABLE VII1-40. 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
    <0.013
    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 VIII-41.  SUMMARY OF TREATED EFFLULNT 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/mm/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 m3/min/m2
    (9.3 gpm/ft2)
    Filtration with Lime
    Addition to pH 9.5
    0.37 m3/min/m2
    (9.5 gpm/ft2)
    EFFLUENT CONCENTRATION (TOTAL, mq
    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
    O
    
    
    
    
    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)
    Fe
    0.09
    
    
    0.08
    
    
    0.06
    
    
    0.07
    
    
    0.06
    
    
    0.54
    0.04
    
    
    
    
    0.04
    
    
    
    0.07
    
    
    
      Dual media filtration (anthrafilt and silica sand)
     *Field pH values
                                       348
    

    -------
    TABLE VHI-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*
    _v_
    
    
    
    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 VIII-42. 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
    < 0.005-0.01 3
    <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
    

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    TABLE 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
    POLLUTANT
    PARAMETER
    pH
    TSS
    Sb
    At
    Be
    Cd
    Cr
    Cu
    Pb
    Hg
    Mi
    Se
    Ag
    T1
    Zn
    TOC
    phenol
    NUMBER OF
    OBSERVATIONS
    26
    27
    23
    
    
    
    
    
    
    
    
    
    
    
    i
    
    
    
    
    
    
    
    
    
    
    
    
    3
    2
    "TOTAL" CONCENTRATION
    (mg/l)
    MEAN
    X
    8.0
    2554
    <0.5
    0.10
    <0.002
    0.014
    0.19
    2.0
    0.16
    0.0007
    0.19
    0.022
    0.014
    <0.02
    0.10
    8
    0.032
    RANGE
    7.8-8.3
    19-44,600
    <0.5
    0.009-1.8
    <0.002
    0.010-0.040
    0.025-3.7
    0.010-43
    0.050-1.9
    <0. 0002-0. 0020
    <0.02-3.6
    0.007-0.20
    <0. 01-0. 11
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    8-9
    0.028-0.036
    "DISSOLVED" CONCENTRATION
    (mg/l>
    MEAN
    X
    --
    --
    __
    —
    --
    0.012
    	
    0.027
    0.068
    
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    0.007
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    --
    --
    --
    —
    --
    0.010-0.015
    	
    0.020-0.035
    0.050-0.090
    --
    
    --
    
    	
    
    -------
    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
    Phenol ics
    PERIOD OF
    OBSERVATIONS
    (DATES)
    Jan. 8- 19, '79
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    
    NUMBER OF
    OBSERVATIONS
    16
    10
    --
    3
    --
    --
    3
    3
    3
    --
    3
    3
    3
    --
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    3
    5
    13
    9
    14
    "TOTAL" CONCENTRATION
    (mg/l)
    MEAN
    X
    8.8
    213
    --
    0.03
    --
    --
    0.07
    0.44
    0.11
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    0.06
    0.025
    0.04
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    0.03
    23
    39
    17
    0.03
    0.58
    RANGE
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    25-1200
    --
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    --
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    0.04-1.24
    0.06-0.19
    --
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    0.02-0.03
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    0.42-68
    32-52
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    0.003-0.060
    0.23-0.81
    "DISSOLVED" CONCENTRATION
    (mg/l)
    MEAN
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    --
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    0.06
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    0.04
    	
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    --
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    —
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    --
    --
                                        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)
    pHt
    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**
    KD
    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**
    «0.02to
    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 VHI-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 rain
    o
    t + 2 hr 15 min
    o
    t + 4 hr 15 min
    o
    t + 7 hr 15 min
    o
    t + 10 hr 15 min
    0
    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 (nun)
    
                    0
                   15
                   30
                   60
                   120
    
    NaOCl Dose 50 mg/1
                  pTT~
    
      Contact Time (min)
    
                    0
                   15
                   30
                   60
                   120
    9
    CN~
    0.189
    0.20
    0.020
    0.095
    0.040
    Total
    Phenolics
    -
    -
    -
    -
    -
    10
    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
    -
    -
    
    CN"
    *.*
    9
    Total
    Phenolics
    ^
    1
    CN"
    0.750
    0.012
    0.003
    0.001
    0.001
    0
    Total
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    0.808
    0.680
    0.396
    1
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    0.039
    0.012
    0.011
    0.011
    1
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    0.608
    0.452
    0.696
    9
    CN"
    0.88
    0.05
    0.02
    0.01
    0.004
    Total
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    0.336
    0.228
    0.088
    -
    0.216
    10
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    0.001
    0.008
    0.003
    <0.001
    Total
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    0.720
    0.058
    0.084
    0.080
    -
    11
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    Total
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    v-Ci^
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    All cyanide and phenol concentrations  are mg/1.
                                    385
    

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    Figure VIII-2. EXPERIMENTAL MINE-DRAINAGE TREATMENT SYSTEM FOR
               UNIVERSITY OF DENVER STUDY
                            MINE
                          DRAINAGE
                 LIME
    
    1
    i
    CONTACT TANK
    (NEUTRALIZATION
    STAGE)
    CONTACT TANK
    (SULFIDE-TREATMENT
    STAGE)
                                           "Q
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      SULFIDE
      SOLUTION
        PUMP
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     MIXING
      TANK
                     DISTRIBUTION TROUGH
                      \    I     I    I
    L
                          TREATED
                          EFFLUENT
                                   389
    

    -------
    Figure VIII-3. CALSPAN MOBILE ENVIRONMENTAL TREATMENT PLANT CONFIGURATION
               EMPLOYED AT BASE AND PRECIOUS METAL MINE AND MILL OPERATIONS
    mm
    /mm)
    "*• 3,785-liter
    (1,000-gallonl
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                                                                        OVERFLOW
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                                                                        OVERFLOW
                                                                           T0
                                                                          WASTE
                   102-liter
                   (27-gallon)
                 FLOCCULATION
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           OVERFLOW
         SLUDGE
         TO WASTE
    
    HOLDING
    TANK
    
    -Q-
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                                                                        OVERFLOW
                                                                           TO
                                                                          WASTE
                                                    390
    

    -------
      Figure VIII-4. MOBILE PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT
                  URANIUM MILL 9402
                                                                            OVERFLOW
                                                                          *-   TO
                                                                             WASTE
                    METERING  METERING
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      RECYCLE SLUDGE
    FILTRATE
    HOLDING
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                                                                            OVERFLOW
                                                                            TO WASTE
    
                                                                            OVERFLOW
                                                                            TO WASTE
                                                                          ^-BACKWASH
                                                                            TO WASTE
    OVERFLOW
    TO WASTE
                                               391
    

    -------
     Figure Vlll-5. PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT URANIUM
                 MILL 9401
                               r—C*J-
    OVERFLOW
     TO WASTE "
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    OVERFLOW
    TO WASTE
                                     392
    

    -------
    Figure VIII-6. MODE OF OPERATION OF SEDIMENTATION TANK DURING PILOT-SCALE
               TREATABILITY STUDY AT MINE/MILL 9402
                     OVERFLOW WEIR
    DISCHARGE
                                                                BAFFLE
                          SLUDGE BLANKET :"::.-'J,
                                   SUPPORT BASE
                                                                    INFLUENT
                                        393
    

    -------
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    -------
    Figure VIII-10.
                 SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT FACILITIES
                 AT LEAD/ZINC MINE/MILL 3103
                               MINEWATER PUMPAGE
                               94.6 I/sec (1,500 gal/mini
                                   MINE SUMP
      TO SMELTER
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       (500 gal/min)
                                   82.0 I/sec (1,300 gal/min)
    
                                           t	
                                      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/tec
    (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
    

    -------
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       Figure VIU-11.  PLOTS OF SELECTED PARAMETERS VERSUS TIME AT NICKEL
                    MINE/MILL 6106 (NOVEMBER 1977 - DECEMBER 1980)
                                       398
    

    -------
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    -------
                               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-
    lished.  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
         Alka1ine-Ch1orination
         Activated Carbon Adsorption
         Hydrogen Peroxide Oxidation
         Chlorine Dioxide Oxidation
         Potassium Permanganate Oxidation
         Ion Exchange
         Granular Media Filtration
         pH Adjustment
         Recy,, '-e
         EVT  -.ration Ponds (total evaporation)
    
    T?M,a  IX- •   indicates  the  processes  studied  for   each    ore
    subcategory.
    
    CAPITAL COST
    
    Capita1 Cost of Facilities
    
    Sett.ling 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
    

    -------
    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
    mechanical  and  electrical  equipment.
    were calculated using the formula:
                 years
    for
    Capital recovery factors
                        n
         CRF = (r)  (1+r)
                     n
                1+r)
    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
         Sodium Hydroxide
         Sodium Hypochlorite
             $   2.00/lb.
             $160.00/ton
             $   0.40/lb.
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         Hydrated Lime                                $ 65.00/ton
         Activated Carbon                             $  0.50/lb.
         Hydrogen Peroxide (70% cone.)                $  0.35/lb.
         Sulfuric Acid  (66 Be)                        $  0.04/lb
         Ferrous Sulfate (400 Ib. drum dry powder)    $  0.52/lb.
         Chlorine Dioxide (5% cone, in 55 gal. drum)  $  9.10/gal.
         Potassium Permanganate (dry)                 $  0.59/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 constructing settling ponds can vary
    widely, depending  on  local  topographic  and  soil  conditions.
    Figure   IX-1   depicts  the  typical  layout  assumed  for  these
    estimates.
    
    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.
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    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
    

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    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
    

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    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
    

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    Capital Cogts.   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  C_o_sts.    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  $0.11/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
    

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    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.
    
    pj! 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
    

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    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 (CRF = 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
    

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    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
    computed at $0.03/KWH  based  on pumping horsepower at  75  percent
    efficiency  and   75   feet  total  dynamic  head,  for   which   the
    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
    

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    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  for  hydrogen  peroxide
    oxidation systems to hydraulic flow 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
    

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    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
    

    -------
    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
    

    -------
    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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    REVISED 2/29/80
                                     478
    

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    Figure IX-18.  ORE MINE WASTEWATER TREATMENT RECYCLING CAPITAL COST
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    -------
                                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  have  demonstrated  both  technological  performance  and
    economic viability 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
    

    -------
    Subcategory and Subpart
    Nonconventional
      Pollutants
      Controlled
        Toxics
    Controlled
       Subcategory and Subpart
      Controlled
    Controlled
         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)
    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
    
                                         0.30
                                         1.0 (1
                                         0.6
                                         0.002
                                         0. 10
            30-Day
            Average
        mg/1
    
            0.15
    5)*     0.5 (0.75)*
            0.3
            0.001
            0.05
       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
                                         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
    conditions may do so as a result of:
                                                             "normal"
    
    
         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.,1   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,
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    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/Mi 1I/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
    129
         Organics          -  87   (Not Detected)
         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 and Subparts in Which  Toxic  Pollutants  Were  Not
    Detected  Are  Excluded  Under  Paragraph   8_  of_  the  Settlement
    Agreement
    
    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 (1) 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
                                     511
    

    -------
    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 with 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
                                    512
    

    -------
    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.
                                       513
    

    -------
    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)(1) 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
    -------
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