DEVELOPMENT DOCUMENT

   '•            FOR FINAL (

  EFFLUENT LIMITATIONS GUIDELINES AND

   NEW, SOURCE PERFORMANCE STANDARDS

      :          FOR THE

        ORE MINING AND DRESSING

         POINT SOURCE CATEGORY
                 A*/1
                 : %P*0^
            Anne E.  Gorsuch
             Administrator
     .Frederic A. Eidsness, Jr.
   Assistant Administrator for Water
           Stephen Schatzow
               Director
    Water Regulations and Standards
             Jeffery Denit
Director, Effluent Guidelines Division
          B. Matthew Jarrett
            Project Officer
             November 1982
     Effluent Guidelines Division
        •    Office of Water..
 U.S. Environmental Protection Agency
        Washington, D.C.  20460

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                        TABLE OF CONTENTS
Sect 1 on

I
II
III
IV
VI
                                                  Page

EXECUTIVE SUMMARY	    1

BEST AVAILABLE TECHNOLOGY ECONOMICALLY
ACHIEVABLE (BAT)	    3

NEW SOURCE PERFORMANCE STANDARDS	    5

BCT EFFLUENT LIMITATIONS	    6


INTRODUCTION	    7

PURPOSE	    7

LEGAL AUTHORITY	    7


INDUSTRY PROFILE	   17

ORE BENEFICIATION PROCESSES	   17

ORE MINING METHODS	   26

INDUSTRY PRACTICE	   34


INDUSTRY SIIBCATEGORIZATON. . .	  Ill

FACTORS INFLUENCING SELECTION OF SUBCATEGORIES..  Ill

SUBCATEGORIZATION	  116

COMPLEXES..	'.	  117


SAMPLING AND ANALYSIS METHODS	  119

SITE SELECTION	  119

SAMPLE COLLECTION,  PRESERVATION, AND
TRANSPORTATION	  125

SAMPLE ANALYSIS	...0	  131

WASTE CHARACTERIZATION.	  155

SAMPLING PROGRAM RESULTS	  155

REAGENT USE  IN FLOATION MILLS		  161

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                                     ABLE OF  CONTENTS (Cpntinued)
                       SPECIAL  PROBLEM  AREAS
                          LECTION OF  POLLUTANT PARAMETERS
                                                                                 194
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                      "™	Tf!tE"n	f o x i c"' P AR AMEfER s.".'...'.	™."";. •;-;.: .•;;••
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        ~^:~~ii^nw	:WTT	RE	C:ATE	GORY... I1."1.'".".1.'.	."".'. ;?.".."...
                                                                                 194
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                                                                                         	iiiiiii	ami
                                                                                       iii	\le	   I
                                                                                 199
                         r]!.N	VEJI	1	0	NAL	^PO.LLUIANJ	P	AR	AM
                                                           !&
                                                       ETERS,
                                                 200
                                                      liiinnjj	pJii	iiliiilijin	i
                                                      mm
                                                                                               !
ONAL
                                          NON- C
                                                          NAL
                       SELECTED..	,.,,,.,,......'.''.'.,.;;.'.;........:.,..'.. , 201
                        S U RROGATE/INDICATOR  RELATIONS HIPS
       !	••	!	!	!	I	f!	!	f
                                                TECH-NOLORY..
                                               	i	i	i	i	i	i	ji	i	ii
                                                                            K. !" ,:„„
                                                 213
iiiH^^^^^^^^^
                        I	M-	P1QCE	SS	C01IRQL	T£CHNOL,OGY	2	13	
                        END-OF-PIPE TREATMENT TECHNIOIJES,
                                                                                224

                                                                                           	!	i	I
                             AN
                                         - SCALE '
'1l!lillil:!lllllll!ll!ilTl1li:i!i;!!i;il!l||il!llll:!!l!lii;iII!!l'>lllllll!! 1
               I!:1,!!!!!111!!,!!!!.111!.!1:]!!!!!!1!!!1!!11!111!!1:1:!!!!1"!"!!,""!!!1'!
                        HISTORICAL DATA SUMMARY,
                                                                                274

                                                                                  294
                                                                                        I	iii	I	iii	ill
                         EVELOPMENT  OF  COST.HATA  BASE...................   399
                    	:	;	;:	ESP	ijAt	frost;;.....'.;...".."..;.".".".	.:.'"::..'.:.."..."   39'9"
                                                                                          1=3	1iM
                                                                                  401
                                                                                             	i
                                                                                  402
                                           ATT(!TR~PTI	on	i	~.	;	:	i	i	i	;	:	:	:	:	:	:	:	:	:	:	:	;	:	¥09	
                                                     ;;*;;^!'.:;^:f':|^ ;••;•!!;•'
                                                                             i.'1
                                                                            ...... ii,;

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                     TABLE OF CONTENTS (Continued)
Section
XI

XII
XIII


XIV


XV
                                                  Page
CHLORINE DIOXIDE TREATMENT		  411

POTASSIUM PERMANGANATE OXIDATION	  411

MODULAR TREATMENT COSTS FOR THE ORE MINING
AND DRESSING INDUSTRY		  412

NON-WATER OUALITY ISSUES	  412


BEST AVAILABLE TECHNOLOGY ECONOMICALLY
AVAILABLE.	  495

SUMMARY OF BEST AVAILABLE TECHNOLOGY	  496

GEN.ERAL PROVISIONS	.	  498

BAT OPTIONS CONSIDERED FOR TOXICS REDUCTION	  507

SELECTION AND DECISION CRITERIA		  511

ADDITIONAL PARAGRAPH 8 EXCLUSIONS	  521

BEST CONVENTIONAL POLLUTANT CONTROL TECHNOLOGY..  523

NEW SOURCE PERFORMANCE STANDARDS (NSPS)	  527

GENERAL PROVISIONS	  527

NSPS OPTIONS CONSIDERED	  530

NSPS SELECTION AND DECISION CRITERIA	  530


PRETREATMENT STANDARDS	  541


ACKNOWLEDGEMENTS			  543


REFERENCES	  545

SECTION III	  545

SECTION V	  ^4fi

SECTION VI	  547

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Secti on
XVI
            TABLE  OF  CONTENTS  (Continued)

                                                 1 '     Page

SECTION  VII	   531

SECTION  "III	 .. ....1... ......... ...."...   548

SECTION  IX. .,	. . .	........ . .	•••••• • • • v   R54
    ,•: ,   .;•.;,  '     	  .  •,•'',••,, '.„>„-,,;/:;,, ...••:., • 	;,  .;] ,  • ,  , ,'.: ; .
SECTION  X...		••,-••   B55
     - "  .  '  '• ;,             •.,.," V  , ji''"'',''!''"! , /" • ;' r ' „ 	'"',',;!;; I <•"  • •,  „,  '„.;. i ...'Vv, '• ]
GLOSSARY.	...'. .'....'. '^. .'.". .... .. .......   557

APPENDIX  A..	 ...... .....•••••••   R77

APPENDIX  R. .	. '.'. . . . ... ...... . . ... ... ..... .. . .   625
                                                                        V!1' .l.li .'!•!« I
                                 vi

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LIST OF TABLES
Number
III-l
III-2
1 1 1- 3
III-4
III-5
IH-6
1 1 1-7
III-R
III-9
III-10
III-ll
111-12
111-13
111-14
111-15
III-lfi
111-17
111-18

111-19

TII-20

ITI-21

III-2?
ITI-23

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 LEAH/ZINC MILLS 	
PROFILE OF GOLD MINES 	 	
PROFILE OF GOLD MILLS 	 	 	 	 	
PROFILE OF MISCELLANEOUS GOLD ANn 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) 	
PROFILE OF MERCURY MINES (1976) 	
PROFILE OF MERCURY MILLS 	 	 	 	 	
Page
55
58
61
67
71
76
79
80
81
82
84
86
87
88
89
90
91

92

93

95

97
98
99
     vii

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                       LIST OF TABLES (Continued)
Number
                                                             Paqe
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 SURCATEGORY..............	   106

111-28    PROFILE OF TITANIUM MINES..		   107

111-29    PROFILE OF TITANIUM DREDGE MILLS.....		   108

111-30    PROFILE OF NICKEL SURCATEGORY....		   10Q

111-31    PROFILE OF VANADIUM SURCATEGORY		   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
                                                       t
V-3       POLLUTANTS ANALYZED AND ANALYSIS TECHNIOUES/
          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 RY TWO DIFFERENT LARORATORIES  USING
          THE BELACK DISTILLATION/PYRIDINE-PYROZOLANE
          METHOD		   151

V-7       ANALTYICAL OUALITY CONTROL PERFORMANCE OF
          COMMERCIAL LABORATORY PERFORMING CYANIDE*
          ANALYSES BY EPA  APPROVED BELACK DISTILLATION
          METHOD		   152

V-8       SUMMARY OF CYANIDE ANALYSIS DATA FOR SAMPLES
          OF ORE MINING AND PROCESSING WASTEWATERS		   153

VI-1      DATA SUMMARY ORE MINING DATA ALL SURCATEGORIES...   166

VI-2      DATA SUMMARY ORE MINING DATA SURCATEGORY IRON
          SUBDIVISION MINE MILL PROCESS MINE DRAINAGE..	   171
                              vi ii

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                       LIST OF TABLES (Continued)
Number
                                                   Page
VI-3
VI-4
VI- 5
VI - 6
VI-7




VI -8



VI - Q •




Vl-lfl


VI-11



VI-12



VI-13
DATA SUMMARY ORE MINING DATA SUBCATEGORY IRON
SMBHIVISON MILL MILL PROCESS PHYSICAL AND/OR
CHEMICAL	   172

DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MINE MILL PROCESS
MINE DRAINAGE....	   173

DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLAT I MUM/
MOLYBDENUM SUBDIVISION MILL MILL PROCESS
CYANIDATION..	   174

DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MILL MILL PROCESS '
FLOTATION (FROTH)			   17B

DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM/
MOLYBDENUM SUBDIVISION MINE/MILL MILL
PROCESS HEAP/VAT/DUMP LEACHING.		   176

DATA SUMMARY ORE MINING DATA SUBCATEGORY
ALUMINUM SUBDIVISION MINE MILL PROCESS
MINE DRAINAGE....	   177

DATA SUMMARY ORE MINING DATA SUBCATEGORY
COPPER/LEAD/ZINC/GOLD/SILVER/PLATINUM
MOLYBDENUM SUBDIVISION MINE/MILL MILL
PROCESS GRAVITY SEPARATION	   178

DATA SUMMARY ORE MINING DATA SUBCATEGORY
TUNGSTEN SUBDIVISION MILL	   179

DATA SUMMARY ORE MINING DATA SUBCATEGORY
MERCURY SUBDIVISION MILL MILL PROCESS
FLOTATION (FROTH).	   180

DATA SUMMARY ORE MINING DATA SUBCATEGORY
URANIUM SUBDIVISION MINE MILL PROCESS MINE
DRAINAGE.		   181

DATA SUMMARY ORE MINING DATA SUBCATEGORY
URANIUM SUBDIVISION MILL MILL PROCESS AND
LOCATIONS	   182
                                ix

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                       LIST OF TABLES (Continued)
Number
Page
VI-14     DATA SUMMARY ORE MINING DATA SIJBCATEGOR.Y
          TITANIUM SUROIVISION MINE MILL PROCESS
          MIME DRAINAGE	  183

VI-1R     DATA SUMMARY ORE MINING DATA SURCATEGORY
          TITANIUM SUBDIVISION MILLS WITH DREDGE
          MINING MILL PROCESS PHYSICAL AND/OR
          CHEMICAL	,	  184

VI-16     DATA SUMMARY ORE MINING DATA SURCATEGORY
          VANADIUM SUBDIVISION MINE MILL PROCESS
          NO MILL PROCESS	  185

VI-17     DATA SUMMARY ORE MINING DATA SUBCATEGORY
          VANADIUM SUBDIVISION MILL MILL PROCESS
          FLOTATION (FROTH)	,...	  186

VI-18     SUMMARY OF REAGENT USE IN ORE FLOATION MILLS	  187

VII-1     DATA SUMMARY ORE MINING DATA ALL SUBCATEGORIES...  204

VII-2     POLLUTANTS TO RE REGULATED	  210

VTI-3     PRIORITY METALS EXCLUSION BY SUBCATEGORY,
          SUBDIVISION, MILL PROCESS	  211

VII-4     TUBING LEACHING ANALYSIS RESULTS	  212

VIII-1    ALTERNATIVES TO SODIUM CYANIDE FOR FLOATION
          CONTROL	  304

VIII-2    RESULTS OF LABORATORY TESTS OF CYANIDE
          DESTRUCTION BY OZONATION AT MILL 6102	  305

VIII-3    RESULTS OF LABORATORY TESTS AT MILL 6102
          DEMONSTRATING EFFECTS OF RESIDENCE TIME, pH,
          AND SODIUM HYPOCHLORITE CONCENTRATIONS ON
          CYANIDE DESTRUCTION WITH SODIUM HYPOCHLORITE	  306

VIII-4    EFFECTIVENESS OF WASTEWATER-TREATMENT
          ALTERNATIVES FOR REMOVAL OF CHRYSOTILE AT
          PILOT PLANTS..	  307

VIII-5    EFFECTIVENESS OF WASTEWATER-TREATMENT
          ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS
          AT ASBESTOS-CEMENT PROCESSING PLANT	  308

VIII-6    EFFECTIVENESS OF WASTEWATER-TREATMENT
          ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS

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                       LIST OF TABLES (Continued)

Number                             .                          Page

          AT ASBESTOS, QUEBEC, ASBESTOS MINE....	   309

VIII-7    EFFECTIVENESS OF WASTEWATER-TREATMENT               t
          ALTERNATIVES FOR REMOVAL OF TOTAL FIBERS
          AT BAIE VERTE, NEWFOUNDLAND ASBESTOS MINE	   310

VIII-8    COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS
          FOR TOTAL FIBERS AND CHRYSOTILE AT SEVERAL
          FACILITIES SURVEYED....	   311

VIII-9    EFFLUENT QUALITY ATTAINED BY USE OF BARIUM
          SALTS FOR REMOVAL OF RADIUM FROM WASTEWATER
          AT VARIOUS URANIUM MINE  AND MILL FACILITIES	   314

VIII-10   RESULTS OF MINE WATER TREATMENT BY LIME
          ADDITION AT COPPER MINE  2120..	   31B

VIII-11   RESULTS OF COMBINED MINE AND MILL WASTEWATER
          TREATMENT BY LIME ADDITION  AT COPPER MINE/
          MILL 2120.		   316

VIII-12   RESULTS OF COMBINED MINE WATER + BARREN  LEACH
          SOLUTIONS TREATMENT BY LIME ADDITION AT  COPPER
          MINE/MILL 2120..	   317

VIII-13   RESULTS OF COMBINED MINE WATER + BARREN  LEACH
          SOLUTIONS MILL TAILINGS  TREATMENT BY LIME
          ADDITION AT COPPER MINE/MILL 2120.		   31R

VIII-14   CHARACTERISTICS OF RAW MINE DRAINAGE TREATED
          DURING PILOT-SCALE EXPERIMENTS IN NEW
          BRUNSWICK, CANADA			   319

VIII-15   EFFLUENT OIIALITY ATTAINED DURING PILOT-SCALE
          MINE-WATER TREATMENT STUDY  IN NEW BRUNSWICK,
          CANADA...		   320

VIII-16   RESULTS OF MINE-WATER TREATMENT BY LIME
          ADDITION AT GOLD MINE 4102....	   321

VIII-17   RESULTS OF LABORATORY-SCALE MINE-WATER
          TREATMENT STUDY AT LEAD/ZINC MINE 3113.	'..   322

VIII-18   EPA-SPONSORED WASTEWATER TREATABILITY STUDIES
          CONDUCTED BY CALSPAN AT  VARIOUS SITES IN ORE
          MINING AND DRESSING INDUSTRY	   323

VIII-19   CHARACTERIZATION OF RAW  WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
                               xi

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                       LIST OF TABLES (Continued)

Number                                                       Page

          ZINC MINE/MILL 3121	  324

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	  32B

VIII-21   OBSERVED VARIATION WITH TIME OF pH AND COPPER
          AND ZINC CONCENTRATIONS OF TAILING-POND DECANT
          AT LEAD/ZINC MINE/MILL 3121	  326

VIII-22   SUMMARY OF TREATED EFFLUENT DUALITY ATTAINED
          WITH PILOT-SCALE UNIT TREATMENT PROCESS AT
          MINE/MILL 3121 DURING AUGUST STUDY	  327

VIII-23   SUMMARY OF TREATED EFFLUENT DUALITY ATTAINED
          WITH PILOT-SCALE UNIT TREATMENT PROCESS AT
          MINE/MILL 3121 DURING MARCH STUDY	•	  32«

VIII-24   RESULTS OF OZONATION FOR DESTRUCTION OF
          CYANIDE	  329

VIII-25   EFFLUENT FROM LEAD/ZINC MINE/MILL/SMELTER/
          REFINERY 3107 PHYSICAL/CHEMICAL-TREATMENT
          PLANT		  330

VIII-26   CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
          ZINC MINE/MILL/SMELTER/REFINERY 3107	  331

VIII-27   SUMMARY OF TREATED EFFLUENT DUALITY ATTAINED
          WITH PILOT-SCALE UNIT TREATMENT PROCESSES AT
          MINE/MILL/SMELTER/REFINERY 3107	  332

VIII-28   CHARACTER OF MINE DRAINAGE FROM LEAD/ZINC
          MINE 3113...	  333

VII1-29   CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER) AT LEAD/
          ZINC MINE 3113...	  334

VIII-30   SUMMARY OF PILOT-SCALE TREAT-ABILITY STUDIES
          AT MINE 3113	  335

VIII-31   CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER) AT ALUMINUM
          MINE 5102	  336

VIII-32   SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
                              xii

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                       LIST OF TABLES (Continued^

Number                                                       Page

          AT MINE 5102		.	   337

VIII-33   CHARACTERIZATION  OF RAW  WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER)  AT MILL.
          9402 (ACID LEACH  MILL WASTEWATER)	   338

VIII-34   CHARACTERIZATION  OF RAW  WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER)  AT MILL
          9402 (ACID LEACH  MILL WASTEWATER).		   339

\7III-35   SUMMARY OF A WASTEWATER  TREATABILITY RESULTS
          USING A LIME ADDITION/BARIUM CHLORIDE
          ADDITION/SETTLE PILOT-SCALE TREATMENT SCHEME	   340

VIII-36   CHARACTERIZATION  OF RAW  WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER)  AT MILL
          9401	„	   341

VIII-37   SUMMARY OF TREATED EFFLUENT OUALITY ATTAINED
          WITH PILOT-SCALE  TREATMENT SYSTEM  AT MILL
          9401		.			'	   342

VIII-38   RESULTS OF BENCH-SCALE ACID/ALKALINE MILL
          WASTEWATER NEUTRALIZATION		   343

VIII-39   CHARACTERIZATION  OF INFLUENT TO WASTEWATER
          TREATMENT PLANT AT COPPER MINE/MILL/SMELTER/
          REFINERY 2122 (SEPTEMBER 5-7 1479)...		   349

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^	   345

VIII-41   SUMMARY OF TREATED EFFLUENT OUALITY FROM
          PILOT-SCALE TREATMENT PROCESSES AT COPPER
          MINE/MILL/SMELTER/REFINERY 2122
          (SEPTEMBER 5-10,  1979)	'	   346

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)....	   348

VIII-43-   SUMMARY OF TREATED EFFLUENT OUALITY FROM
          PILOT-SCALE TREATMENT PROCESSES AT COPPER
          MINE/MILL/SMELTER/REFINERY 2121
          (SEPTEMBER 18-19, 1979)	   349
                              xiii

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                       LIST OF TABLES (Continued)

Number                                                       paqe

VIII-44   HISTORICAL D'ATA SUMMARY FOR IRON ORE MINE/
          MILL 1808 (FINAL DISCHARGE)..		   350

VIII-45  .HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL
          2121 (FINAL 'TAILING-POND DISCHARGE:  TREATMENT
          OF MINE PLUS MILL WATER)...-.		   351

VIII-46   HISTORICAL DATA SUMMARY FOR COPPER.MINE/MILL
          2120 (TAILING-POND OVERFLOW:  TREATMENT OF
          UNDERGROUND MINE, MILL, AND LEACH-CIRCUIT
          WASTEWATER STREAMS)	   352

VIII-47   HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL
          2120 (TREATMENT SYSTEM-BARREL POND-EFFLUENT:
          TREATMENT OF MILL WATER AND OPEN-PIT MINE
          WATER).	-.	   353

VIII-48   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3105 (TREATED MINE EFFLUENT).	.	..	   354

VIII-49   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE
          3130 (UNTREATED MINEWATER)				   355

VIII-50   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE
          3130 (TREATED EFFLUENT)	•>•••	.••••••   356

VIII-51   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3101 (TAILING-POND DECANT TO POLISHING
          PONDS)		.	   357

VIII-52   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3101 (FINAL DISCHARGE FROM POLISHING
          POND)	, .	   358

VIII-53   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3102 (FINAL DISCHARGE)		   359

VIII-54   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3103 (TAILING-POND EFFLUENT TO SECOND
          SETTLING POND).	   360

VIII-55   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3103 (EFFLUENT FROM SECOND SETTLING
          POND)......		   361

VHI-56   HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/
          MILL 3104 (TAILING-LAGOON OVERFLOW)
          (EFFLUENT)			   362
                               xiv

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                       LIST OF TABLES (Continued)

Number                                                       Page

VIII-57   HISTORICAL DATA SUMMARY FOR LEAH/ZINC MILL
          3110 (TAILING-LAGOON OVERFLOW)	   363

VIII-58   HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE
          6103 (TREATEO EFFLUENT)	.	   364

VIII-B9   HISTORICAL DATA SUMMARY FOR MOLYBDENUM MILL
          filOl (TREATED EFFLUENT)	   36B

VIII-60   HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/
          MILL 6102 (CLEAR POND BLEED STREAM-INFLUENT
          TO TREATMENT SYSTEM)	   366

VIII-61   HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/
          MILL 6102 (FINAL DISCHARGE FROM RETENTION
          POND-TREATED EFFLUENT)...	   367

VIII-62   TREATED EFFLUENT DATA FOR  ALUMINUM ORE
          MINE BIO2	    363

VIII-63   HISTORICAL DATA SUMMARY FOR TUNGSTEN MINE
          6104 (TREATED EFFLUENT)..	    369

VIII-64   HISTORICAL DATA SUMMARY FOR URANIUM MINE
          7008 (TREATED MINE WATER)	.......		    370

VIII-65   HISTORICAL DATA SUMMARY FOR NICKEL MINE/MILL
          6106 (TREATED EFFLUENT)..	    371

VIII-66   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	   372

VIII-67   CHARACTERIZATION OF RAW WASTEWATER (INFLUENT
          TO PILOT-SCALE TREATMENT TRAILER) AT BASE
          AND PRECIOUS METALS MILL 2122  DURING PERIOD
          8-19 JANUARY 1979	   373

VIII-68   SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
          PERFORMED AT MILL 2122 DURING  PERIOD OF
          6-14 SEPTEMBER 1978	   374

VIII-69   PERFORMANCE OF A-DUAL MEDIA FILTER WITH TIME-
          FILTRATION OF TAILING POND DECANT AT MILL
          2122		   37 B

VIII-70   RESULTS OF CHLORINATION BUCKET TESTS FOR
          DESTRUCTION PHENOL AND CYANIDE IN MILL 2122
                               xv

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                       LIST OF TABLES (Continued)

Number                                                       Page

          TAILING PONO DECANT	  376

VITI-71   RESULTS OF PILOT-SCALE OZONATION FOR DESTRUCTION
          OF PHENOL AND CYANIDE IN MILL 21?.?. TAILING POND
          DECANT	  377

VIII-72   EFFLUENT QUALITY ATTAINED AT SEVERAL PLACER
          MINING OPERATIONS EMPLOYING SETTLING-POND
          TECHNOLOGY	  37 8

VITI-73   CHARACTERIZATION OF RAW WASTEWATER (TAIL ING-POND
          EFFLUENT AS INFLUENT TO PILOT-SCALE TREATMENT
          TRAILER) AT COPPER MILL 2122 DURING PERIOD OF
          6-14 SEPTEMBER 1978	  370

VIII-74   CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO
          PILOT-SCALE TREATMENT TRAILER) AT BASE AND
          PRECIOUS METALS MILL 2122 DURING PERIOD OF 8-10
          JANUARY, 1979	  380

VIII-7S   SUMMARY OF PILOT-SCALE TREATABILITY STUDIES
          PERFORMED AT MILL 2122 DURING PERIOD OF 6-14
          SEPTEMBER 1978	  381

VIII-76   PERFORMANCE OF A DUAL MEDIA FILTER WITH TIME-
          FILTRATION OF TAILING POND DECANT AT MILL 2122...  382

VIII-77   RESULTS OF ALKALINE CHLORINATION BUCKET TESTS FOR
          DESTRUCTION OF PHENOL AND CYANIDE IN MILL 2122
          TAILING POND DECANT	  383

VIII-78   RESULTS OF PILOT-SCALE OZONATION FOR DESTRUCTION
          OF PHENOL AND CYANIDE IN MILL 2122 TAILING POND
          DECANT	,.	  384

VIII-79   EFFLUENT DUALITY ATTAINED AT SEVERAL PLACER
          MINING OPERATIONS EMPLOYING SETTLING-POND
          TECHNOLOGY		  38 R

IX-1      COST COMPARISONS GENERATED ACCORDING TO
          TREATMENT PROCESS AND ORE CATEGORY	  414

IX-2      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)	  415

IX-3      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ihs)	  424
                              xvi

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Number
                       LIST OF TABLES (Continued)
Page
IX-4      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)	   431

IX-5      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)...	   439

IX-6      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ihs)	,	   443

IX-7      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEQUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ihs)		   444

IX-8      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEHUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)	   452

IX-9      COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEOUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)	   453

IX-10     COST COMPARISON FOR VARIOUS TYPES OF TREATMENT
          TECHNOLOGIES AND SUBSEOUENT COST PER TON OF
          ORE MINED (1 tn = 2000 Ibs)...	   454

IX-11     SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID
          LEACHATE TEST  (mg/1)...	   459
                               xvii

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                         LIST OF FIGURES

Number                                                       Page

VIII-1    POTABLE.WATER TREATMENT FOR ASBESTOS REMOVAL
          AT LAKEWOOD PLANT, DULUTH, MINNESOTA	  386

VIII-2    EXPERIMENTAL MINE-DRAINAGE TREATMENT SYSTEM
          FOR UNIVERSITY OF DENVER STUDY	  387

VIII-3    CALSPAN MOBILE ENVIRONMENTAL TREATMENT PLANT
          CONFIGURATIONS EMPLOYED AT BASE AND PRECIOUS
          METAL MINE AND MILL OPERATIONS	  388

VIII-4    MOBILE PILOT TREATMENT SYSTEM CONFIGURATION
          EMPLOYED AT URANIUM MILL 9402	  389

VITI-5    PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED
          AT URANIUM MILL 9401...	  390

VIII-6    MODE OF OPERATION OF SEDIMENTATION TANK DURING
          PILOT-SCALE TREATABILITY STUDY AT MINE/MILL
          9402	.,	  391

VIII-7    FRONTIER TECHNICAL ASSOCIATES MOBILE TREATMENT
          DLANT CONFIGURATION EMPLOYED AT MINE/MILL/
          SMELTER/REFINERIES #2l?.l and #2122	  392

VIII-8    DIAGRAM OF WATER FLOW AMD WASTEWATER TREATMENT
          AT COPPER MINE/MILL 2120...	  393

VIII-9    DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT
          AT COPPER MINE/MILL 2120	  394

VIII-10   SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT
          FACILITIES AT LEAD/ZINC MINE/MILL 3103	  395

VIII-11   PLOTS OF SELECTED PARAMETERS VERSUS TIME AT
          NICKEL MINE/MILL 6106 (1962-1974)	  396

VIII-12   PLOT OF TSS CONCENTRATIONS VERSUS COPPER
          CONCENTRATIONS IN TAILING-POND DECANT AT
          MINE/MILL/SMELTER/REFINERY 2122	  397

IX-1      SECONDARY SETTLING POND/LAGOON - TYPICAL LAYOUT..  464

IX-2      ORE MINING WASTEWATER TREATMENT  SECONDARY
          SETTLING POND/LAGOON COST CURVES	  465

IX-3      ORE MINE WASTEWATER TREATMENT SETTLING PONDS -
          LINING COST CURVES	  466

IX-4      FLOCCULANT (POLYELECTROLYTE)  PREPARATION AND
                             xviii

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                       LIST OF FIGURES (Continued)

Number                                                       Page

          FEED-FLOW SCHEMATIC...	  467

IX-R      ORE MINE WASTEWATER TREATMENT FLOCCULANT
          (POLYELECTROLYTE) PREPARATION.* FEED SYSTEM
          COST CURVES	. ..	  468

IX-6      OZONE GENERATION AND FEED FLOW SCHEMATIC.	  469

IX-7      ORE MINE WASTEWATER TREATMENT OZONE GENERATION
          K FEED SYSTEM COST CURVES		  470

IX-R      ALKALINE-CHLORINATION FLOW SCHEMATIC..	  471

IX-9      ORE MINE WASTEWATER TREATMENT ALKALINE
          CHLORINATION COST CURVES..	  47?.

IX-in     ION EXCHANGE FLOW SCHEMATIC....		  473

IX-11     OR.E MINE WASTEWATER TREATMENT ION EXCHANGE
          COST CURVES	  474

IX-12     GRANULAR MEDIA FILTRATION PROCESS FLOW
          SCHEMATIC	  475

IX-13     ORE MINE WASTEWATER TREATMENT GRANULAR MEDIA
          FILTRATION PROCESS COST CURVES			  476

IX-14     pH ADJUSTMENT FLOW SCHEMATIC		  477

IX-15     ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT
          CAPITAL COST CURVES.	  478

IX-16     ORE MINE WASTEWATER TREATMENT pH ADJUSTMENT
          ANNUAL COST CURVES	  479

IX-17     WASTEWATER RECYCLE FLOW SCHEMATIC	  480

IX-18     ORE MINE WASTEWATER TREATMENT RECYCLING
          CAPITAL COST CURVES,.	  481

IX-19     ORE MINE WASTEWATER TREATMENT RECYCLING
          ANNUAL COST CURVES		  482

IX-20     ACTIVATED  CARBON ADSORPTION  FLOW SCHEMATIC	  483

IX-21     ACTIVATED  CARBON ADSORPTION  CAPITAL COST
          CURVE FOR  PHENOL REDUCTION  IN BPT  EFFLUENT
          DISCHARGED FROM  BASE  AND  PRECIOUS  METAL ORE
          MILLS	  484
                               xix

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Number
                       LIST OF FIGURES (Continued)
Page
IX-22     ACTIVATED CARBON ADSORPTION ANNUAL COST CURVE
          FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
          FROM BASE AND PRECIOUS METAL ORE MILLS	  48R

IX-23     CHEMICAL OXIDATION - HYDROGEN PEROXIDE FLOW
          SCHEMATIC	  48fi

IX-24     HYDROGEN PERO.XIOE TREATMENT CAPITAL COST CURVE
          FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
          FROM BASE « PRECIOUS METAL ORE MILLS	. .  487

IX-25     HYDROGEN PEROXIDE TREATMENT ANNUAL COST CURVE
          FOR PHENOL REDUCTION IN BPT EFFLUENT DISCHARGED
          FROM BASE ft PRECIOUS METAL ORE MILLS	  488

IX-26     CHEMICAL OXIDATION-CHLORINE DIOXIDE FLOW
          SCHEMATIC	  489

IX-27     CHLORINE DIOXIDE CAPITAL COST CURVE FOR PHENOL
          REDUCTION IN BPT EFFLUENT DISCHARGED FROM
          BASE AND PRECIOUS METAL ORE MILLS...	  490

IX-28     CHLORINE DIOXIDE ANNUAL COST CURVE FOR PHENOL
          REDUCTION IN BPT EFFLUENT DISCHARGED FROM
          BASE AND PRECIOUS METAL ORE MILLS	  491

IX-29     CHEMICAL OXIDATION-POTASSIUM PERMANGANATE
          FLOW SCHEMATIC	  492

IX-30     POTASSIUM PERMANGANATE TREATMENT CAPITAL COST
          CURVE FOR PHENOL REDUCTION IN BPT EFFLUENT
          DISCHARGED FROM BASE AND PRECIOUS METAL ORE
          MILLS		.		  493

IX-31     POTASSIUM PERMANGANATE TREATMENT ANNUAL COST
          CURVE FOR PHENOL REDUCTION FROM BASE AND
          PRECIOUS METAL ORE MILLS.	  494
                              xx

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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  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 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. Costie, 14 ERC 2161  (D.D.C. 1980).  The
settlement agreement outlines a strategy  for  regulation  of  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
from an industrial plant to- a publicly owned treatment works, are
not promulgated.

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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, and silver ore mines
and mills.  This new subcategory is renamed as  the copper,  lead,
zinc,  gold,  silver,  and molybdenum ores subcategory.  Tungsten
ore mines and mills are placed in a new and separate subcategory.
Four new subcategories have been added since the time  the  court
sustained  the  final  BPT  rule.   Each of the new  subcategories
consists of a single facility.  Additionally,   for   clarification
the  BPT  and BAT  subcategorization  schemes  have  been made the
same.

An extensive  sampling and analysis effort was undertaken in  1977
and  extends  to  the  present.   As  part  of  this effort,   20
facilities were visited  for  screening  and  14  facilities  for
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-
lations.    The  three conditions that determine whether a regula-
tion is classified as major are:

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 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    manor   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 m^aor increase in costs,  or significant  adverse  effects
 on  the industry.

 This  regulation  was  submitted   to the Office of Management and
 Budget for review as required  by  Executive  Order  12251.    Any
 comments  from  OMB   and  EPA's  responses   to those  comments are
 available for public inspection at  the  EPA  Public   Information
 Reference Unit,  Room 2922 (EPA Library),  Environmental Protection
 Agency,  401  M Street,  S.W.,  Washington,  B.C.

 BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE (BAT)

 The  presence  or absence of the  129 toxic  pollutants  and several
 conventional  and  nonconventional  pollutants has  been   determined
 as   a  result  of the sampling and analysis program.   One hundred
 twenty-four  of the 129  toxic pollutants have been excluded in all
 subcategories based  upon criteria  contained  in  the   Settlement
 Agreement cited previously:   (1)  they  were  not detected,  (2)  they
 wfre  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 were proposed and are  promulgated.   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
 BPT  facilities which consist of lime precipitation  and  settling.
Of    the   three   alternatives,   no  statistically  significant

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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 sulfide 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 this upset or bypass  (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  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
BAT  review.   Specific technology  for  the destruction of cyanide
was considered,  but is not necessary because in-process  controls

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lets lhlnnoiTma<%  h^SHeWat?^  in tailing ponds reduce cyanide to
less tnan  u.4 mg/1  based  on  the precision  and  acrurarv  r>f  i-h^
analytical  method   for   wastewafer  discharges f^o^mines and
                                           BAT «"d the subpart  is
                                      ""ines  ace located in remote
HEW SOURCE PERFORMANCE STANDARDS (NSPS)


                                      imPlei"ent the best and most
                                  .
                                 ^^
                      ra

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                 m-r
The discharge  is subject to standards for mine drainage.

BCT EFFLUENT LIMITATIONS
(BOD), oil and grease  (OsG), and fecal  coliform.  F??aJ ~-

B^0&^AfTe\^^^^
toxic metals.
          BCT effluent limitations based on BPT were proposed  for
 time.

 in^Pad the Aaency has  proposed  BCT  limitations  for  the  ore
 mining  iSduStS  as part of  its consolidated  BCT rulemakxng.   47
 FR 49176, October 29,  1982.

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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
 Hoi »  ?UrCe cate<3°ry for Possible regulation.  This  development
 ™  T*  P^ents  tne  technical data base compiled by EPA with
 regard to these pollutants and their treatability for  regulation
 ±6nonthe  CJean  Water Act.  The concentrations of conventfona?
 a"d"onc°nventional pollutants were also examined for the  estab-
 lishment of effluent limitations guidelines.  Treatment technolo-
 niS   rr? /JS°u assessed  f°r designation as the best available
 demonstrated technology (BADT) upon which new source  performance^
 standards  (NSPS  are based.   This document outlines the technSl-
 ogy options considered  and  the  rationale  for  selecting  each
                el '   TheSe technol°9y ievels are the basis for the
                      -
LEGAL AUTHORITY


IS?
w*tr
Water
to  he
to the
                           :      Und€r authority 'of Sections  301,
       D   i  <-    '     '      501  of the Clean Water  Act (the Federal
       Pollution Control  Act  Amendments of 1972, 33  USC  1251   et
       Aa^ame£ued   by   the  Clean Water Act of  19?7,  P.L.  95-217)
       SfM    JS! re<3ulations  are  also promulgated in  response
       Settlement Agreement  in Natural Resources  Defense Council
            Trai  8 ERC  2120 (D"n)-c7T976) ,  modifi^— T2
The Clean Water Act
             Water  p°llution  Control  Act  Amendments  of  1972
 h~        v, a comprehensive program to "restore and maintain the
waters-' lecMon 'im? ,biolS9ical  integrity  of  the  HatioS'I
waters,   Section  101 (a).   By  1  July 1977. existinq industrial
dischargers  were  required  to  achieve  "effluent   limitations
requiring   the  application  of  the  best  practicable  cont?Sl
          .cu^"tly available" (BPT),  Section 301 (b) (t ) (A)    By
           '    Se ^scha^ers were required to achieve "effluent
             re(5ufrin?  the  application  of  the  best available
           ecnomcally achievable  .  .  .  which  will  result  in
                     K pr°9ress   toward  the  national  goal  of
                 dl?c5ar?e -of  a11  Poll"tants"  (BAT),  Section
                   indust^ial direct  dischargers were required to
             h" 3°?,n?r S°U^Ce Perfo^ce standards (Slis)?'
             best   available   demonstrated   technology.     Th4
            !  5°r.direCJ discha^ers  were  to be incorporated into
         Pollutant Discharge Elimination System  (NPDES)   permits
        under Section 402  of the Act.   Although Section 402?a?(lT
based
based
issued

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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. y_._  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
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 In general, ore njines and mills are located  in rural
 from a POTW.
areas,   far
 The  1977  Amendments  added  Section  301(b)(2)(E)  to  the  Act
 ?S£2bli?hi'ng "best  conventional  pollutant  control  technology"
  th056 for  BPT.   As  BPT is  the minimal  level
 of control required  by law, no possible application of   the  BCT
 SSUJ5  %  °°Uld   E2fui^ in  BCT  limitations lower than  those
 proposed.  However,  EPA did not  promulgate BCT   as   proposed  and
 reserved  BCT   until  the revised BCT  methodology is proposed.  At
 the time  EPA will again propose  BCT for this  category.

 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-
 •?US4-i,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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(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).

On 14 June 1982,  EPA again proposed BAT and  NSPS  and  requested
comments  from the public.  Over  50  comments were received.  The
Aaencv reviewed the comments and where  data   indicated,   amended
the  proposed  regulations.  The  final regulation was promulgated
3, December  1982."

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
                                   in

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pollutants that Congress subsequently designated
also regulated under BPT.
as  toxic  were
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
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
                                  11

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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.

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 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  aftd  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 promulgating 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 informations

   1 .  Screening and verification sampling and analysis
       programs

   2.  Engineering cost site visits

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    3.  Supporting data from EPA regional offices

    4.  Treatability studies

    5.  Industry self-monitoring sampling

    6.  BPT data base

    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 nonconventional 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
                                  13

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   4.
screening; and

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
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.
                                 14

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 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.

 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.                                                        J

 EPA  obtained the data for  its  economic  analysis primarily from a
 survey conducted under Section  308 of the  Clean Water  Act    The
Agency sent questionnaires  to 138 companies engaged in mining and
milling  of  metal  ores.   The data collected included production
 levels,  employment, revenue, operating   costs,  working  capital,
ore  grade,   and other relevant  information.   The economic, survey
data were supplemented  by  data  from  government  publications
trade journals, and visits to several mine/mills.
                                IS

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                            SECTION III

                         INDUSTRY PROFILE

 ORE BENEFICIATION PROCESSES

 As    mined,   most  ores   contain  the  valuable  metals  (values)
 disseminated  in  a matrix  of less  valuable  rock   (gangue).    The
 purpose   of   ore  beneficiation   is  the  separation of the  metal
 bearing minerals from  the gangue to  yield  a  product   which  is
 higher  in  metal  content.    To  accomplish  this,   the ore must
 generally be  crushed and/or ground  small   enough  so  that   each
 particle  contains mostly  the  mineral  to be recovered or mostly
 gangue.   The  separation of the particles on  the   basis  of   some
 difference  between the ore mineral and  the gangue can  then  yield
 a concentrate high in   metal  value,   as  well   as waste   rock
 (tailings) containing  very little metal.   The separation is  never
 perfect,  and  the degree of success which is attained is generally
 described by two numbers:   (1)  percent  recovery  and (2)  grade of
 the product (usually a concentrate).  Widely varying results  are
 obtained  in  beneficiating  different ores;  recoveries may  range
 from 60 percent  or less to greater than  95  percent.    Similarly,
 concentrates  may  contain  less   than 60  percent or more than 95
 percent of the primary ore mineral.   In  general,  for a  given  ore
 and  process,   concentrate  grade  and  recovery  are   inversely
 related.  Higher recovery  is   achieved   only  by   including   more
 gangue,   thereby yielding  a  lower  grade  concentrate.  The process
 must  be optimized,  trading off recovery  against   the value   (and
 marketability)   of the   concentrate  produced.   Depending on end
 use,  a particular  grade of concentrate is  desired,   and  specific
 gangue components  are  limited  as  undesirable impurities.

 Many  properties   are  used  as the basis  for  separating  valuable
 minerals  from gangue,  including:   specific  gravity,  conductivity,
 magnetic permeability,  affinity for certain   chemicals,   solubil-
 ity,  and the tendency to  form chemical  complexes.   Processes for
 effecting the separation may be generally  considered  as:

 1.  gravity concentration

 2.  magnetic separation

 3.  electrostatic  separation

 4.  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,  CaW04).

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   (FeTiOS.)  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.

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Sink/Float Separation

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

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separators  are  most  often  run  wet  on the slurry produced in
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  peirticle  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)  eind  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,

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etc.)  differ  at  each  operation  where  it  is practiced  and  may
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  (NaS2COC2H5_).   Frothers  are generally organics with a
soluble group and a nonwettable hydrocarbon.   Pine  oil  hydroxyl
(C6HT_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 metals,  or other
specific solutions.  This process is called leaching.
                                 22

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Leaching solutions are categorized as  strong,  general  solvents
(e.g., acids) and weaker, specific solvents (e.g., cyanide).  The
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 o>ne 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 collectton 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.
                                23

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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. + 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
(UO2(S04)3 --- >  R4 UO2 (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  HI   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)  affect  separation
 of the desired product.  Solvent handling ease permits  the con-
 struction of  multistage,  concurrent and countercurrent, solvent
 extraction concentrators which are useful even  when  each  stage
 effects   only  partial  separation of a  value from an  interferent.
 Unfortunately, the solvents are easily  polluted by  slime  so
 complete  liquid/solid  separation is necessary.   Ion  exchange and
 solvent extraction circuits can  be combined to take advantage  of
 the   slime   resistance  of  resin-in-pulp  ion  exchange  and the
 separatory efficiency of solvent extraction (Eluex process).

 Ion  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-pillar, block caving,
timbered stopes, open stopes, shrinkage stopes,  sublevel  stopes
and  others  (Reference  3).  Underground mining usually is inde-
pendent of surface mining, but sometimes preceeds or follows  it.
Waste removal is proportionately much less in underground than in
surface  mining, but still requires surface waste disposal areas.
Underground mines supply substantially  all  the  lead  and  zinc
mined domestically.

In  situ  mining  procedures  include the leaching of uranium and
copper.

Considerations given to the choice of  mining  method  and  brief
descriptions of the methods typically employed are given below.

Surface Mining Operations

Whether  an  ore  body  will  be  mined by surface or underground
methods will be determined by the economics of the operation.  In
general, surface or  open-pit  mining  is  more  economical  than

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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.

Drilling

Drilling  is the basic part of the breaking operation in open-pit
mining;  considerable  effort  has  therefore  been  expended  to
develop equipment for drilling holes at the lowest possible cost.
There  are  several  types  of  drills  which can be used.  These
include churn drills, percussion drills, rotary drills  and  jet-
piercing drills.  All are designed with one objective in mind: to
produce  a hole of the required diameter, depth, and direction in
rock for later insertion of explosives.

Blasting

Basically, explosives are  comprised  of  chemicals  which,  when
combined,  contain  all  the requirements for complete combustion
without  external  oxygen  supply.   Early  explosives  consisted
chiefly of nitroglycerine, carbonaceous material and an oxidizing
agent.    These   mixtures  were  packaged  into  cartridges  for
convenience in handling and loading into holes.  Many  explosives
are still manufactured and packaged to the basic formulas.

In  recent  years,  it  has been discovered that fertilizer-grade
ammonium 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  States, 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  10.0-200  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  arid
 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.

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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  strin
 overburden when required.

 At  a  few sites in Alaska,  bucket-line dredges are still used to
 mine gold from placers.   Prior to dredging,   the  frozen  gravels
 are thawed by circulating water through 3.8  cm (1.5 inches)  pipes
 contained  in  drill  holes  spaced on 4.9 meter (16 foot) centers
 and drilled to bedrock.   Thawed gravels are  dredged with a  chain
 of buckets which dump their  contents into a  hopper on the dredge.

 Titanium  minerals  contained  in sand deposits in New Jersey and
 ancient beach placers in  Florida  are  also  mined  by  dredging
 methods.    In  these  operations,  a pond is  constructed above the
 ore body,  and a dredge is floated  on  the   pond.    The  dredges
 currently used are normally  equipped with suction  head cutters to
 mine the mineral sands.

 Solution Mining

 In  situ  or  solution mining  techniques are  used in some parts of
 Arizona,  Nevada and  New  Mexico  to  recover copper and in Texas  to
 recover   uranium.    In  situ  mining involves leaching the desired
 metal  from mineralized ground in place.   During in situ leaching,
 the ore  body  must  be penetrated and  permeated   by  the  leachina
 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 leachina 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  milling  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  of  the   U.S.  iron  ore
 industry uses  milling  operations which  result  in  no  discharge,  31
percent  discharge  to  surface  waters,  and  the  discharge  practices
 for  15 percent are  unknown.   For pelletizing operations  alone,  56
 percent  of  total production  is represented  by  operations   prac-
 ticing  no  . discharge  of  process wastewater, 35  percent  discharge
 to surface  waters,  and the  discharge  practices of 8   percent   are
 unknown.    A  summary   presented   in  Tables III-l and II1-2 shows
 production  data, processes  and wastewater  technology  employed,
 and discharge  methods  and volumes.

 Unlike  the  milling  segment,  the mining segment of the iron ore
 industry does  discharge,  either directly to  the  environment  or
 into the mill  water circuit, either as the primary source of pro-
 cess  water  or  as  makeup  water.  Water can cause a variety of
 problems if allowed to  collect  in  mine  workings.    Therefore,
 water is collected and 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

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 single  facility  uses  alum  and   a   long-chain   polymer
 flocculation aids for fine-grained suspended solids.
as
 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-
 £ent ?£.iron ore Pellet production will come from  zero-discharge
 facilities,  and  zero-discharge for the Mesabi Range subcategory
 is required for BPT.                                         y^y

 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  bv  the
 fine-grinding process.

 A  selective  flocculation  technique  has  been  developed   that
 reduces the slimes which are so  detrimental.    in  this  process,
 nrLnif°nh-Tin?KalS-^re  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
 reaction      ^  iS   lime-treated   to  create   a  water-softening


 Copper,  Lead,  Zinc,  Gold,  Silver,  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 II1-3  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
XJ!®*   llst?d  in, ^he  tables   are located in Arizona, while the
others  are located in seven other states.   In  addition   to  the
£i=??h  i?^d. i" JabJe   m-3'  a  recent MSHA  (M^e Safety and
Health  Administration)  tabulation   indicates  that  22   smaller
operations  with  an  average  employment  of about 10 people are

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presently engaged in copper mining.   However, production,  Process
and water use information for these small mines is not available.


The tabulation below provides a production cross section  of  the
major copper mining states in 1976 (Reference 6):
      State

       Arizona
       Utah
       New Mexico
       Montana
       Nevada
       Michigan
       Tennessee
       Idaho
1000 Metric Tons

       157,339
        26,817
        22,690
        15,220
         7,092
         3,448
         1,845
           128
  Production
1000 Short Tons

     173,472
      29,567
      25,016
      16,781
       7,820
       3,801
       2,034
         141
   The  total domestic  copper mine production from  1968 to  1979 is
 shown below  (1968-1973 production - Reference 7, 1974-1976
 production - Referenced,  1977-1979 production
                            - Reference 5)
       Year

        1968
        1969
        1970
        1971
        1972
        1973
        1974
        1975
        1976
        1977
        1978
        1979
 1000 Metric 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
  Production
 1000 Short Tons

     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    Learhinn
 operations  currently  employ  sulfuric acid (5 to "o'oercInM  or
 iron sulfate to dissolve copper from the oxide  o'  Sixel  ox de-
 sulfide   ores  in   dumps,   heaps,  vats or in situ.   The coDoer  is
 subsequently recovered  from solution in a highly  pure  ?orm  via
 precipitation,   electrolytic   deposition  (electrowinninaT   or
 solvent   extraction-electrowinningT    Production  of  cement anrf
 per?en?Win roSSf" f  C?"tribut^  •   significant  qLnU?y" (17°6
 mining       6>" °f  th&   recoverable  copper  produced  through
  «nh  *eachin9  circuits  require makeup water,  total recycle of
leach circuit water  is  common  practice.   Therefore   the  BPT
effluent  limitations guidelines for mines and mill s which emnlov
dump, heap, in situ,, or vat leach processes for the extraction of






Copper Milling
                                                  o
                                                      seen in the




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and  mill  wastewater   and   reports  a  discharge    At least one
milling operation is reported  to discharge intermittently.
Lead/Zinc
read and zinc are often found in the  same ore and  as  such,  are
dilculsed  together throughout this document.  The domestic lead/
zncmfntng industry presently consists  of  2^ n^^|to°f ^^ions
(which  may have more than one mine)  and 23  concentrators located
in 12 states   Three of these mines and  four of the concentrators

           ^--^r-ESt »•£  dsssrs -KS-SSS
and six concentrators during the same time  period.

D S   mine  production  of  lead  in.  1975  dropped six percent to
56?; 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  !n  977 totaled 537,499 metric tons (592 500 short tons)
 Reference 5)   Production of lead continued to   decline and  in
 5?r ^Suction totaled 525,569 metric tons (579 300 short tons)
 Reference 5)   Missouri remained the leading producer,  with 89.8
pSrllnlol the nation's total lead production ,  followed  by  daho
 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  1 974
 and 1 979, domestic lead prices  continued to rise ^om a_price  of
 47  4   cents  per  kilogram  (21.5 cents per pound)  in 1975 to $1.16
 r>er kiloSram (52.6 cents per  pound)  in  1979.  The  current  price
 ??! Sep?Sr  1981)  is 93cents per  kilogram (42 cents per  pound)
 (Reference  9) .
 short tons) in 1975.   It then recovered  to about  440,000  metric
 ?ons (4?4?500 short tons)  in 1976  (Reference 5).  Zinc production
 deceased to 407,900 metric tons (449,600 short tons  in  977 and
 further  to  302 700 metric tons (333,700 short tons) in 1978 and
 26^5o5 metric tons (294,600 short tons) .in  1979  (Reference  5).
 Tennessee  was  the  leading zinc  producing  state in  1979 witn si
 SranTSf total production and was followed by Missouri, 23 per-
 Snt* SeW Jersey,P12 percent; and  Idaho,   11   percent  (Reference
 5K   Tennessee  led the nation in product ion  of "nc for 1 5 con-
 secutive vears prior to 1973 and regained  that status   in  1975,
 ?977,  ?978?  and  1979  due to the opening  of two  new  concentra-

 tions.

 in  contrast to climbing lead prices, zinc  prices h™ ^1J^r_a
 downward   trend  over  the  past decade in terms of real Collars.
 Followina  inflationary price increases during  the period 1972  to
 Following  intiatio   y P          ^& average  1975  price of 85.91

 cents per  kilogram  (38.96 cents per pound)   to  81.61  cents  per

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 kilogram (37.01  cents per pound) in 1976 to 75.85 cents  oer kilo-


                                          domestic
              °ref are ?roduced almost   exclusively  from  under-

                                     "  •                     «


Gold
                 ^
                 -ss-
                      gold Producers accounted for  73 percent of



                                                     -;h

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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
have D|en ™j      ^y   ^.^  ig75  ^  lg?^ fQur fco gix new

cvanidation operations  began  full  scale  operation.   However,
despite  thesS  prospects,  domestic  production has continued to
decline during rScent 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
nrices   in   1977  to  1978  led to a decline  in byproduct production
of  qold-  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  II1-9  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  Caching
 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.

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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
 £°"!  
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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   1982, 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

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2e±M.»:r"! ^^irecoyeries 2f over " Percent
 Molybdenum
     Year

     1949
     1953
     1958
     1962
     1968
     1972
     1974
     1977
     1979
                                { i
                                                         The orice
                          Metric Tons

                            10,222
                            25,973
                            18,634
                            23,250
                            42,423
                            46,368
                            51,000
                            55,484
                            657302
Production
Short Tons

  11,265
  28,622
  20,535
  25,622
  46,750
  51,098
  56,000
  61,204
  71,984
in the planning and development stages.                 mills  are


                               43

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Aluminum

In 1979, the major aluminum ore,  bauxite,  was  mined  by  eight
 r,'iQ7Q  (Reference 5)   The only operations  mining  bauxite  for





 "Combination  Process,"  which   is   dified   as  SIC   289   A
 1974  to  1979  (References  5  and
      year

      1974
      1975
      1976
      1977
      1978
      1979
            Metric Tons

               1,980
               1,800
               1,989
               2,013
               1,669
               1,821
Production
Short Tons

   2,181
   1,983
   2,191
   2,217
   1,839
   2,006
 Average annual production for the last 10 years is  approximately
 1,880;000 metric tons (2,070,000 short tons).
 All
   from  the  two
                                  domestic aluminum ore operations
 Silica  (Si02_)
 Tnnhent  (%) of Ore
           Percentage of  Total  Domestic Shipments
           TO-?*    1976   1977    1978    1979
  less  than  8
  8  to  15
  greater  than
15
4
62
34
6
50
44
2
54
44
2
55
43
1
55
44
     pilot  project  proved the economic viability of alu.n^e *|  an.
        domestic  bauxite ore operations require discharge of large

                                   44

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 nfh          a"ributed to a single operation.  Characteristics
 of the two domestic bauxite operations are shown in Table 111-17?

 Tungsten
 Tungsten
Iff|
1969
       97








      1979
                  and  milling  is conducted by numerous  (probably
              _	s, the majority of which  are  very  small
              intermittently.   As  illustrated  below,  tungsten
                             constant  between   1969   and
                                        (Contained Tungsten)
                                             Short Tons -
                                             - 3~~9l)3 —
                                                * "I
                                                3 450
                                                4 075
                    Domestic Production
                           Metric Tons
                             3,543

                             4'369
                             3'132
                             3'699
                            3,015
                                                 3,321
lll-?Sfil?«h?f
111-19.   Table
                         rniriing  is Presented  in Tables  111-18 and
                      describes   the   larger  operations   (annual
                      hSn  5'°??  metric  tons  0?s p~cS2Sed/;S«T)
 than    nnn      • de?cribes smaller  operations  (production  less
 than   5,000 metric tons ore processed/year).  The maioritv of the
 mines  are located in the western  states  of   California  Oreaon
 Idaho   Utah,  and Nevada.  Almost all are underground  mines  and
 many have no discharge of mine water.      "Aground  mines, and
                         Profile tungsten mills.  Processes  used
nn  n,     •  n f?Para^on a^/or fatty acid and sulfide flotation
One mill  in California produces  the  majority  of  the  tunasten
concentrate .   Wastewater treatment methods vary? but may include

the "rt?ve SiJlS9^^^*11^"^016 and/°r evaporation y nSstol
    '                                     ^  ecause they are  in
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.    Historical iS
SiSbe?UtSftJinnhe ^"f^ -States has come from a relatfvelyC?arg4
number  of small production operations.  However,  since 1969

                             mlnes has decllned ^ % " '

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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
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 II1-22 and  II1-23.

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

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                    '
  pollutant rather than as a product.
                                                                 a
                                                                 a
 Primary  deposits  of  uranium  ores  are  widely  distributed in
 granites and pegmatites.  These black ores contain  the  tetrava-
 df?h nminfralS  uraninite  (002) and coffinite 
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employ fewer than five people.  The actual number of actjye mines
at  any  given time will vary, depending on market conditions and
company status.

Most mines ship ore to the  mill  by  truck.   The  economics  of
hauling  unbenSficiated  ore  require  that the distance from the
m?ni tl the mill be no more than a few kilometers  (^°™tely
one  mile).   However, certain high-grade ores (0.6 percent U|0|)
mav be shinned up to  200  kilometers  (120  miles).   The  large
numbe? X Sail Sines often requires individual mills to Purchase
ore  from  several  different  mines,  both company and P"jately
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  to™5  ??J
day  (7,000  short  tons per day).   In addition,  four of these mills
are  practicing vanadium byproduct  recovery,  on* »jjl/*  "J^f£
ing molybdenum concentrate  as byproduct,   and  another  jntermit
tentlv 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 Caching the
 calcine with water  generates  a  vanadic  acid  solution.   This
 SlSSon,   which contains about eight percent V205  i| 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  ^^PP1"?*1^  JJj|h
 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.

  Tn  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
 IS   copplr    The  dissolved  uranium  is  recovered in  a  solvent

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 exchange circuit and then directed to  the main  plant   for   final
 processing and recovery as yellowcake.

 In Situ Recovery

 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
 ^niionS a  U10i  (Reference  25).   Typically,  alkaline   leach
 solutions are  pumped  into  a  series  of  strategically  placed
 ih?™^°?   well£3'   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
           ?^ uncertainties exist about the future use of nuclear
           th;f>.country,  increases in yellowcake requirements  are
           W1  in  fc     6Xt  10  years'   The  ann"al  demand for
                                       '
           n   exPeCJed4.t° grow from approximately  15,400  metric
 tons  (17,000  short  tons)  of U308 in 1976 to 35,400 metric tons
 (39,000 short tons)  in 1985.   As^'result,  the  d4creas!ng  gradS
 S «on nnn  °l* •  "±U  require °-s-  mill  capacity to increase from
 2c ?nA°nSnmetriC.tonS (9'800^00 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

                                -
                                                    "trie
Projected  increases  in U3_08. demand  have  resulted  im

(1) the exploration  and expansion of known  sandstone  deposits;

c™,4-hthen  ®xPloration of new sandstone areas  in Nevada, North and
South   Dakota,  Colorado,   Wyoming,   and    Montana-   and   (3)
              investi9ation  of  "hardrock"   areas  in  Colorado
       rat          nn      .Pa"erns  include  three new uranium
       lit. nnn  31Z'9°° metric  tons  (350,000  short  tons)  per
    Ann   ;°°° 4.metr^c  tons  (700,000  short tons) per year, and
   ,000 metric tons (750,000 short tons) of  ore  pe?  year   and
seven new in situ leach operations.
«SSa?6S <-in _the  deman<*  and  price of yellowcake will provide
added impetus for the extraction of lower grade ores, the ?ef ine-
        co"ventional milling processes, the  development  of  new
        and  milling  techniques (e.g. , hydraulic mining of sand-

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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.

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
T?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)
fPeculation   and  settling  for heavy metal and  suspended solids
removal,  (4)  BaC12  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 tne country.

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 Antimony
 Antimony
 byproduct
 centrated
 operation
 is  mined
 flotation
 wastewater
 of process
           is  recovered  from  antimony  ore  (stibnite) and as a
           of silver and lead concentrates.  This industry is con-
           /2-   /vx?rftes:  *daho and Montana.  Currently, only one
           (Mine/Mill 9901) recovers only antimony ore.   The  ore
            underground and concentrates are obtained by the froth
           process.   There is no -discharge  from  the  mine,  but
             from  the mill flows to an impoundment.   No discharge
            wastewater to surface waters occurs.
 A  second  facility,  Mine/Mill  4403  recovers  antimony  as   a
 byproduct  from  tetrahedrite,  a  complex silver-copper-an?Lony
 sulfide mineral.   The antimony  is  recovered  from  tetrahedrite
 o?n?he Snvl  ^  ^ electr°lytic extraction plant SperSteS by onJ
 Jdaho         mining GomPanies in the Coeur d'Alene  district  of
Antimony   is  also  contained  in  lead  concentrates  and
                                                         recovered

           shrt              Pr?ducti?n °? antimony was 655 metric
         2  short  tons).   Production  at  facility 9901  in  1979  was
 b;.   in  1979,  the  total   domestic   mine   production   of   a
 concentrate was  reported  as  2,990 metric  tons  (3  294.  short tons
 This  concentrate   contained  655 metric  tons   722 short  torn? of
 antimony.  Mine/mill  9901  is profiled  in  Table  III-!?

 Titanium
and
s ^
                              °£ ttn'UB
                                                Hmenlte   (PeTiOS)
                                                      in    estle
produced domestically are from titanium dredging operations   The

    l°
During recent years, domestic production of i linen ite
                               had dropped to m844 n,etri  ton
                                ^
                               Bl

-------
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  JII-30.   As
 shown  below, production has decreased slightly from 1969 to 1980
 (References 5, 18, 20, and  27):
                                  52

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     Year
     1969
     1970
     1971
     1972
     1973
     1974
     1975
     1976
     1977
     1978
     1979
     1980
Metric Tons
  15,483
  14,464
  15,465
  15,309
  16,587
  15,086
  15,421
  14,951
  13,024
  12,263
  13,676
  13,302
Production
Short Tons
  17,056
  15,933
  17,036
  16,864
  18,272
  16,618
  16,987
  16,469
  14,347
  13,509
  15,065
  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  sulflde
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  i's  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
priinai!Y~^r^c©vecy__xof 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.

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At vanadium Mine/Mill 6107, vanadium pentoxide, V2_O5_,  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  (Na6VK)028)  is concentrated by solvent  extraction,
and ammonia is added to precipitate ammonium vanadate.   This  is
dried and calcined to yield a V2_05_ product.

The  most  significant  effluent  streams  are  from leaching and
solvent extraction, wet scrubbers or roasters,  and  ore  dryers.
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Production of vanadium is summarized below (References 5 and 18):
                         Metric Tons
                           3,737
                           4,756
                           4,731
                           7,330
                           6,866
                           4,036
                           5,302
Production
Short Tons
 4,117
 5,240
 5,213
 8,076
 7,565
 4,446
 5,841

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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 givens

"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-
categorization are:

   1.  Designation as a mine or mill
   2.  General geologic setting
   3.  Type of mine (e.g., surface or underground)
   4.  Ore mineralogy
   5.  Type of mill process (beneficiation, extraction
       process)
   6.  Wastes generated
   7.  End product
   8.  Climate, rainfall, and location
                             111

-------
   9.  Reagent use
   10.  Water use or water balance
   11.  Treatment technologies
   12.  Topography
   13.  Facility age

These factors have been examined to determine   if  the  BPT  sub-
categorization should be retained or  if any modification would be
appropriate.

Designation as a Mine 6r 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'us$.  In many  instances,  it  is  advanta-
geous to separate mine water from mill process  wastewater because
of  differing  water quality, flow rate, or treatability.  Levels
of pollutants in mine waters are often lower or less complex than
those in mill process wastewaters.   For  many  mine/mill  opera-
tions,  it is more economical to treat mine water separately from
mill water, especially if the mine water requires minimum  treat-
ability.  Mine water contact with finely divided ores (especially
oxidized  ores)  is  minimal and mine water is  not exposed to the
process reagents  often  added  in  milling.    Wastewater  volume
reduction  from  a  mine  is  seldom  a viable  option whereas the
technology is available to reduce  or  eliminate  discharge  from
many  milling  operations.   Therefore,  development of treatment
alternatives and guidelines may be difficult for mines and mills.

Because many operations follow this approach,   designation  as  a
mine  or  mill  provides  an  appropriate  basis  to classify the
industry within subcategories.

That is not to say that mine water and mill water  might  not  be
advantageously  handled  together.  In some instances, use of the
mine wastewater-as mill process water will result in an  improved
discharge  quality  because of interactions of  the process chemi-
cals and the mine water pollutants.

General Geologic Setting

The general geologic setting (e.g., shape of  deposit,  proximity
to  surface)  determines  the  type  of  mine (i.e., underground,
surface or open-pit, placer, etc.).  Therefore, geologic  setting
is not considered a basis for subcategorization.

Type of_ Mine

The choice of mining method is determined by the general geology;
ore   grade,  size,  configuration,  and  depth;  and  associated
overburden of the ore body.   Because no  significant  differences
resulted  from  application  of  mine water control and treatment
                               II?.

-------
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,
Generated
Type  of  Beneficiation  Process  and  Wastes
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 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  hydrometallurgy.   Physical  processing using water, such as
gravity separation, discharge the suspended solids generated from
washing, dredging, crushing, or grinding.  The exposure of finely
divided ore and gangue to water also leads to  solution  of  some
material.  The dissolved and suspended metals content varies with
the ore being processed, but wastewater treatments are similar.

Froth  flotation  methods  affect  character  of mill effluent in
several ways.  Generally, pH is adjusted  to  increase  flotation
efficiency.   This  and the finer ore grind (generally finer than
for physical processing) may have the secondary  effect  of  sub-
stantially increasing the solubility of ore components.  Reagents
used   in  the  flotation  processes  include  major  pollutants.
                               113

-------
Cyanide and phenol compounds, for example, are  used   in  several
flotation  processes.  Although their usage  is usually  low, their
presence in effluent streams have potentially harmful  effects.

Ore leaching  operations  differ  from  physical  processing,  and
flotation  plants.   The use of large quantities of reagents such
as strong acids and bases and the  deliberate  solubilization  of
ore  components   (resulting  in  higher  percent  soluble  metals
content)  characterizes   these   operations.    Therefore,   the
characteristics   of  the  wastewater  quality  as  well  as  the
treatment and control technologies employed  are  different  than
for physical processing and flotation wastewater.

The  wastes  generated  as part of mining and beneficiating metal
ores are highly dependent upon mineralogy and processes employed.
This was considered in all subcategories.

End Product

The end products are closely allied to the mineralogy of the ores
exploited; therefore, mineralogy and processing were found to  be
more advantageous methods of subcategorization.

Climate, Rainfall, and Location

There  is  a  wide diversity of yearly climatic variations in the
United States.  Unlike many other industries,  mining  and  asso-
ciated  milling operations cannot choose to  locate in areas which
have desirable characteristics.  Some mills  and mines are located
in arid regions of the country and can use evaporation to  reduce
effluent  discharge  quantity.   Other  facilities are located in
areas of net positive precipitation and high  runoff  conditions.
Treatment  of large volumes of water by evaporation in many areas
of the U.S. cannot be used  where  topographic  conditions  limit
space and provide excess surface drainage water.  A climate which
provides  icing  conditions  on  ponds  will also make control of
excess water more difficult than in a semi-arid  area.   Climate,
rainfall, and location were, therefore, considered in determining
whether a particular subcategory can achieve zero discharge.

Reagent Use

Reagent use in many segments of the industry (for example, in the
cyanidation  process for gold) can potentially affect the quality
of wastewater.  However, the types and quantities of reagents are
a function of the mineralogy of the ore and  extraction  processes
employed.    Reagent   use,   therefore,   was  included  in  the
consideration of benefipiation processes (e.g.,  cyanidation  for
gold).
                              114

-------
 Water Use and/or Water Balance

             °r ^ater balance ^ highly dependent on the choice of
 and/or water balance are considered with benef iciation processes

 Treatment Technologies

 Many  mining  and  milling  establishments  use  a  single  tvpe  of
 fndnSS* tr^tment method.   Treatment procedures  vary  within^ the
 industry   but  widespread  adoption of  differing technologies  is
 not prevalent.  Therefore,  it was  determined that ore  mine?alogi
 SiiJ Processed waste characterization  provide a mo?e acceptable
 basis for subcategorization than treatment technology.     PuaDae

 Topography

 Topographical  differences  between  areas  are beyond  the control

            usssss-

                             1SSl.C5ta-
          -
regulations.
                       However, topography is known to  influence
                                                            "a
Facility Age

Many mines and mills have operated for the past 100  years.    For
Tn m?nimS?eSa^°nS' installation of replacement equipment results
in minimal differences in water quality.  Many mill processes for
concentrating ores in the industry hav4 not cLnged PconJfdl?ab?y
(e.g.    froth   flotation,   gravity  separation,   arindina  and
crushing)   but  improvements  in  reagent  use?  monitor ina  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 i5
StrLtlSf 12^°?ithe age 5f the facilities, bu? Is a funcUon of
extractive  metallurgy  and  process  changes.    Virtually  everv
facility  continuously  updates  in-plant  processing  and   ffow
JSSfffS'  eVe? ,.tho^h  basic processing may remain the same.   In
addition, most treatment systems  employ  end-of-pipe  techniaueS
which  can  be installed in either old or new plantl   ThSrSforS

              lt  1S n0t 3 U6           f°r s5bcatwriSS?oS in.
                               115

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SUBCATEGORIZATION

After a review of BPT subcategorization,  and  after  a  36-rnonth
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 ma^or  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  and tungsten ha've been put into separate subcategories 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 and silver.  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
                                116

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 wastewater  generated,  the same wastewater treatment  technologies
 are   applicable in this  subcategory regardless of  the type of ore
 mined and milled.   Because  of  these  factors,  the  subcategory
 formerly  called Base  and Precious  Metals is  expanded and renamed
 the   Copper,   Lead,  Zinc,   Gold,   Silver  and  Molybdenum   Ores
 Subcategory.

 From   the   comments  received   on   the   proposed regulation it is
 apparent that  retention  of  the old  BPT   subcategorization  scheme
 for   the  BPT   limitations   only confused, rather  than  clarified
 matters.    The  commenters   suggested  that,   to eliminate  this
 confusion,  the Agency should  use the identical  subcategorization
 for all the limitations  and standards.   Accordingly,  in  the final
 regulation,    the    Agency    is     eliminating    a    separate
 subcategorization   scheme   for the  BPT  limitations,   and  is
 instead, using the same  scheme for  all  the  BPT,  BAT,   and  NSPS
 limitations.     It  should   be noted  that   the   BPT   effluent
 limitations for any existing mill are the same as those  which are
 currently in force and are  not subject  to additional  review.

 An additional  modification  to  the   BPT   subcategorization  scheme
 has been made.   The Agency  is  establishing  a separate subcategory
 tor  platinum  mines and  mills.   The Agency  received  comments that
 there were  no  platinum mines or  mills in existence and that a new
 platinum mine   and mill  was   being  considered  that   would be
 substantially   different  than  existing  mines and mills upon which
 the Agency  based best demonstrated  technology.   The  Agency  is
 therefore,   establishing  a new  subcategory addressing  platinum ore
 and   is  reserving NSPS  for  this  subcategory.  Antimony ore has
 been  added  in  a separate   subcategory  and  BAT    limitations
 reserved.   BPT  did not include  antimony  ores.

 COMPLEXES
The  subcategorization  scheme  has  subdivisions  for  mines and
mills; complexes are not included.  Because of the  individuality
    complexes,  regulation  of  them  has  been  delegated to the
of
Agency's Regional offices and that practice  will  continue.   As
discussed  in  Section  V, Sampling and Analysis Methods, several
complexes have been sampled during BAT guideline development  and
a separate guidance document has been prepared to aid the Regions
in preparation of permits commensurate with BAT.
                               117

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TABLE IV-1   PROPOSED SUBCATEGORIZATION FOR BAT - ORE MINING AND
            DRESSING
SUBCATEGORY

Iron Ore

Copper, Lead, Zinc, Gold,
Silver,
Molybdenum Ores
Aluminum Ore
Tungsten Ore
Nickel Ore
Vanadium Ore*
Mercury Ore
Uranium Ores
Antimony Ores
Titanium Ores
Platinum Ore
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 Mill
or In-Situ Mines
Mine Drainage
Mills
Mine Drainage
Mills
Mills with Dredge
Mining
...Mine Drainage
Mills
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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     3

     4

     5

     6
- 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

Aluminum Ore Subcategory - Mine 5102

Tungsten Ore Subcategory - Mine/Mill 6104

Mercury Ore Subcategory - Mill 9202

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
                                 19.1

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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
                                122

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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:
                             123

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     1.  Iron Ore SubCategory - Mine/Mill 1132

     2.  Copper, Lead, Zinc, Gold, Silver, Platinum, Molybdenum
         Ore Subcategbry
         - Copper Ore—Mine 2110 and Mine/Mill 2116
         - Zinc Ore-i-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 Mining Study

A  sampling  effort  was conducted to evaluate treatment technol-
ogies at Alaskan placer mines.  BAT regulations for  gold  placer
mining  are  reserved  for  further study.  However, several gold
placer mining operations, all located in Alaska, were sampled  to
determine  performance  capabilities  of  existing settling ponds
used to remove suspended and settleable solids.  A summary report
has been issued and the data are included in  Reference  1.   The
operations visited as part of this effort are:

     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 Ore Subcategory - Mine/Mill 9906, Mine/Mill
         9907 and Mine/Mill 9910.
                                124

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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/Mill/Smelter/Refinery 6106)

      4.   Vanadium Ore (Mine/Mill 6107)

      5.   Mercury Ore (Mine/Mill  9202)

      6.   Antimony Ore (Mine/Mill 9901)

 SAMPLE COLLECTION,  PRESERVATION.  AND TRANSPORTATION

 Collection,   preservation,   and   transportation  of  samples  were
 accomplished  in accordance with  procedures outlined  in   Appendix
 III   of   "Sampling   and  Analysis  Procedures   for   Screening  of
 Industrial Effluents for Priority Pollutants"  (published   by   the
 EPA   Environmental  Monitoring  and Support Laboratory, Cincinnati,
 Ohio,  March 1977,  revised  April  1977) and in  "Sampling  Screening
 Procedure  for  the  Measurement of Priority Pollutants"  (published
 by the EPA Environmental Guidelines  Division,   Washington,  D.C.
 October   1976).   The  procedures  used  are   summarized   in   the
 paragraphs which  follow.

 In general, four  types of  samples were collected:

 Ty_Ee 1.  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
                                125

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ISCO     sampler.      Also,     in     the     case    of    one
mine/mill/smelter/refinery  complex,  two   consecutive   24-hour
composite  samples  were usually collected to better characterize
the several waste streams involved.

Type 2.  A 24-hour  composite  sample,  totaling  1  liter  (33.8
ounce) in volume, was analyzed for the presence of total cyanide.
This  was  a  composite  of  four 250-ml (8.5 ounce) grab samples
collected at 6-hour intervals.

Type 3.  A 24-hour composite  sample,  totaling  0.47  liter  (16
ounce)  in  volume,  was  analyzed for the presence of phenolics.
This was a composite  of  four   118-ml  (4-ounce)  grab  samples,
collected at 6-hour intervals.

Type 4_.  Two 125-ml (4.2 ounce)  grab samples  (one a backup sample
collected  midway  in  the  24-hour sampling period) were analyzed
for the presence of volatile organic compounds by the  "purge  and
trap"  method   (discussed   further under Sample Analysis later in
this section).

All sample containers were  labeled  to  indicate  sample  number,
sample  site,   sampling  point,  individual collecting  the sample,
type of sample  (e.g., composite  or grab, raw  discharge or treated
effluent), sampling dates and times, preservative used (if  any),
etc.

Collection and  Preservation

Screen, Verification, and Additional  Sampling Programs

Whenever  practical,  all samples collected at each  sampling point
were  taken from mid-channel at mid-depth  in   a  turbulent,  well-
mixed  portion of the  waste  stream.  Periodically,  the  temperature
and pH of each  waste  stream sampled were measured  on-site.

Each  large composite  (Type  1) sample  was  collected in  a new  11.4-
liter   (3-gallon),  narrow-mouth  glass   jug  that  had  been washed
with  detergent  and water, rinsed  with   tap   water,   rinsed   with
distilled water,  rinsed  with methylene  chloride,  and air dried  at
room  temperature in a dust-free  environment.

Before  collection of Type  1  samples,  new Tygon tubing was  cut  to
minimum  lengths and  installed on the  inlet  and  outlet  (suction
and   discharge)  fittings   of  the automatic  sampler.   Two  liters
 (2.1  quart)  of  blank  water, known to be free  of organic compounds
and  brought  to  the sampling site from the analytical  laboratory,
were   pumped through  the sampler and its attached tubing into the
glass jug;  the  water  was then distributed to cover  the  interior
of the jug  and  subsequently discarded.

A  blank  was   produced  by  pumping  an additional 3 liters (3.2
quarts)  of  blank water  through  the  sampler,   distributed  inside
                                 126

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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  sampl,e  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
                                   127

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analysis, a 0.95-liter (1-quart) aliquot was placed in  a  sample
container supplied by the representative.

Sample  Types  2  and 3 were stored in new bottles which had been
iced and labeled, one liter (33.8-ounce)  clear  plastic  bottles
for  Type  2,  and  0.47-liter  (16-ounce) amber glass for Type 3.
The bottles had been cleaned by rinsing with distilled water  and
the samples were preserved as described below.

To  each  Type  2 (cyanide) sample, sodium hydroxide was added as
necessary to elevate the pH to  12 or more (as measured  using  pH
paper).  Where the presence of  chlorine was suspected, the sample
was  tested  for chlorine  (which would decompose most of the cya-
nide) by using  potassium  iodide/starch  paper.   If  the  paper
turned  blue,  ascorbic  acid crystals were slowly added and dis-
solved until a drop of the sample produced no change in the color
of the test paper.  An  additional  0.6  gram   (0.021  ounce)  of
ascorbic  acid  was added, and  the sample bottle was sealed (by a
Teflon-lined cap), labeled,  iced  and  shipped  to  Calspan  for
analysis.

To  each  Type 3  (total phenol) sample, phosphoric acid was added
as necessary to reduce the pH to 4 or less  (as measured using  pH
paper).  Then, 0.5 gram  (0.018  ounce) of copper sulfate was added
to  kill bacteria, and the sample bottle was sealed  (by a Teflon-
lined cap),  labeled, iced  and shipped to Calspan for analysis.

Each Type 4  (volatile organics) sample was stored  in a new  125-ml
(4.2-ounce)  glass bottle that had been rinsed with tap water  and
distilled  water, heated to 105 C  (221 F) for 1 hour, and cooled.
This method  was also used  to prepare the septum and  lid for  each
bottle.  Each bottle, -when used was filled  to overflowing,  sealed
with  a  Teflon-faced  silicone septum   (Teflon side down) and a
crimped  aluminum  cap,  labeled,  and iced.    Hermetic  sealing  was
verified by  inverting  and  tapping the sealed  container to confirm
the  absence of  air bubbles.   (If bubbles  were found, the  bottle
was opened,  a few additional drops of sample  were  added,   and  a
new  seal  was   installed.)  Samples were maintained hermetically
sealed and iced until  analyzed  by Gulf  South  Research.

Verification Monitoring  Program

Sampling methods  for the monitoring program were similar  to those
used in  the  screening  and  verification   efforts.    However,   the
monitoring   samples  were  collected  by  industry  personnel,  in
contractor-supplied  containers  and per   contractor  instructions,
over   time   periods  ranging  from   2   to 12  weeks.   Samples were
shipped  to Calspan  for analysis.

Surveillance and  Analysis  Program

As  discussed previously,  the   samples  for   this  program were
collected  and analyzed by  Agency regional  personnel  in  three dif-

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 ferent  regions.  Techniques were very similar to those described
 above and EPA approved protocol was observed.

 Cost Site Visit Sampling Program

 As discussed earlier, this  program  was  primarily  designed  to
 collect  cost  data  and  sampling  was conducted only during the
 short site visit  necessary  to  gather  cost  data.   Therefore,
 single  grab  samples were taken.  For total metals and classical
 pollutant analyses,  a  one-liter  plastic  bottle  and  cap  were
 rinsed  several  times  with  the  stream  to be sampled, and the
 bottle was filled.  For  dissolved  metals,  a  portion  of  this
 sample  was  sucked  by hand pump through a Millipore 0.45 micron
 tilter 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
 shipped to Radian Corporation.

 Sample Transportation
styrofoam  chest  with  ice,  and
 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.                         s»««npxe

 Associated Data Collection
         dn                  "Iatin9  to  P^nt  operations  were
          S«J2?i  SLt?   samPifn<3   visits.    This  additional  data
'r rnni-   ?   *!   information   on  production,  water use,  waste-
water control,  and wastewater  treatment  practices.   Flow diagrams
were obtained or  prepared  to indicate  the course  of  significant
XJ^n^S   Stfeamf '  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.                               *-u*ic

Uranijum Study
 u.          automatici  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.    i                                       *

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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
upbn 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.
                                  130

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  h  mh    "ere 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

             '  da?6'  and  additional  sampling information! This
     TP      a£Val?° recorded in field notes.  For  samples  sent

                  T?A  thS  Sample  WaS  assigned  an EPA sample
    iA      /  and the aPPropciate traffic report was completed.
 Field  data  on  PH,  settleable  solids,  and  temperature  were
 collected to augment the data base.                   "mre  were

 SAMPLE ANALYSIS


 Sampling and Analytical Methods


 fu   C°!?gfesst 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

                      fSW years ago'  and onlv  on  «re  occasions
 mhor  has industry monitored or even developed
 methods to monitor  these pollutants.   Section 304(h)  of the  Act

 estSbfUh f;q^ireS   Jhe Administrator to promulgate guidelines  to
 establish test procedures  for the analysis of  toxic  pollutants.
 As  a  result,  EPA scientists,  including staff of the Environmental

 ^n?arhMLab0ra^°ry in  Athens'  Ge°rgia,  and staff  of  the Environ-
 ™nd^J°ni^!:in9<- and  SuPP°rt   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
         Sampl1?9  a?d analytical Protocols and  were  incorporated
            pr?toc°ls ultimately   adopted  for the study of  other
            Cate9orles   See Sampling  and  Analysis  Procedures for

 reved Apriftl 9^77^^^- Effluents   for   Priority   PollutantsT


 Because Section 304(h) methods  were   available  for  most   toxic

 S?fnK?'-FrtEeS i°   *' °Yanfde  and Phenolics  (4AAP),  the  analytical
 orn^n^  ?  ?   on -developing  methods for  sampling  and analyses of
 organic  toxic pollutants.  The three basic  analytical   approaches
 considered  by  EPA  are infrared spectroscopy  (IS), gas ?hromato-
 graphy  (GC) with multiple detectors, and  gas   chromatography/mass

XJn^°rtry   (GC/MS)'1  Evaluation of these alternatives led ?hJ
Agency  to proposed analytical techniques  for   113   toxic  organic

         S   (S6   4      u69464'  3  December  1979, amended 44 TO
                         based °n:  (1) GC with  selected  detec-
             7Performance liquid chromatography (HPLC), depending
   o-^Ular  P°nutant and (2) GC/MS.  In selecting am^ng
these alternatives,  EPA considered their sensitivity,  laboratory
               c?sts' applicability to diverse waste streams f?om
                               131

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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, ot  65
pollutants   and    classes  of  pollutants  potentially  includes
thousands  of  specific  pollutants,   and  the    expenditure   of
resources   in  government  and  private  laboratories  would  be
overwhelming if analyses were attempted for all these pollutants.
Therefore, to make  the task more  manageable,  EPA selected   129
specific  toxic pollutants for study  in this rulemaking and other
industry rulemakings.  The criteria for selection of these   129
pollutants  included  frequency  of occurrence  in water, chemical
stability  and  structure,   amount  of chemical   produced,    and
availablity of chemical  standards for measurement.

As   discussed  in  Sample Collection,  EPA collected  four  types of
samples  from each sampling point:   (1) a  9.6  liter,  24-hour   com-
posite  sample  used  to  analyzed metals,  pesticides,  PCBs,  asbes-
tos,  organic compounds,  and  the  classical parameters;  (2)   a  l-
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  an'd  verification  of  the
 volatile,   base-neutral, and acid organics was gas chromatograpny
 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-

-------
 ture,   TSS,   VSS,  COD,  TOG,  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  plannning data  generation  for the BAT rulemaking,  EPA con-
 sidered  requiring  dischargers to monitor and analyze toxic pollu-
 tants  under Section 308  of the Act.   The Agency did  not use  this
 authority, however, because  it was reluctant to increase the cost
 to  the   industry   and   because  it desired to keep direct control
 over sample analyses  in  view of  the developmental nature  of  the
 methodology and  the need for close quality control.   In addition,
 EPA believed  that  the slow pace  and limited  laboratory  capability
 for toxic  pollutant analysis would have hampered mandatory sam-
 pling  and analysis.   Although EPA believes  that  available  data
 support   the  BAT   regulations,  it  would have preferred  a larger
 data base for some  of the  toxic  pollutants and will   continue to
 seek additional  data.  EPA will  periodically review  these regula-
 tions,   as  required by  the  Act, and make any  revisions supported
 by  new data.

 Parameters Analyzed

Analyses varied  for  the  different  sampling  programs   and  to
simplify  the  discussion,   they   are   cited by category  whenever
possible.  The categories  are:

     Toxics
        Organics (see Table V-l)
           All
           Total Phenolics (4AAP)
        Metals (see Table V-2)
           Total
          Dissolved
        Cyanide  (total)
        Asbestos
           Total Fiber
                                133

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           Chrysotile
     Conventionals
        Total  Suspended Solids (TSS)
        pH
     Non-Conventionals
        Temperature
        Volatile Suspended Solids (VSS)
        Chemical Oxygen Demand (COD)
        Total  Organic Carbon (TOO
        Radium 226
           Total
           Dissolved
        Total  Phenolics (4AAP)
        Total  Settleable Solids
     Miscellaneous Others
The 114 toxic organics (listed in Table V-l), the 13 toxic metals
(listed   in   Table   V-2),   the   conventionals,    and    the
nonconventionals  were  analyzed  in  separate groups, with a few
exceptions  as  follows:   phenolics,  which  were   occasionally
analyzed  without the other toxic organics; radium 226, which was
only analyzed at uranium facilities; and some exceptions  to  the
toxic  metals  group,  to be noted in the following discussion of
specific programs.

Screen Sampling Program

The screen sampling program was designed to build the  data  base
on  toxics.   Therefore,  all mine/mill samples  (and most complex
samples) were analyzed for the toxics (except dissolved  metals).
The   raw  data  presented  in  Supplement   1  include  only  the
parameters detected.

All screen samples were also analyzed for  the   conventional  and
nonconventional  parameters,  with the exception of uranium mines
and/or mills, which only analyzed for radium  226.

Verification Sampling Program

Sample Set  1.  From the six  facilities  visited  in   the  screen
samplingprogram,  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
 specific parameters, as follows:
was  taken  to  determine
                                134

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        Facility
     Pb/Zn-Mine/Mill 3121
        Ag-Mine/Mill 4401
        Mo-Mill 6101
        Al-Mine 5102
         U-Mine 9408
           Mill 9405

Additional Sampling Program
                                     Parameter
                                          CN
                                          Asbestos
                                          Asbestos
                                          Phenolics,  Asbestos
                                          Asbestos
                                          Asbestos
 sulfate,  uranium, vanadium, and radium 226.

 Verification Monitoring Program






 Surveillance and Analysis Program





 from Mine/Mill  6107 were  analyzed for  tSxf?C1organics   ^Non^ o?
 the samples  were tested for asbestos.         w-y-nics.    wone  of

 All
 a
                      cobalt'
Cost Site Visit Program
   uo         W!r!  analyzed for all toxic metals (total and dis-
solved), but not for toxic organics, or cyanide and asbestos.

The samples were screened for both conventional, and  a  varietv
                                              -ettleable  '
                                13S

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Uranium Study


ffi.urSSS Blni3i!^^
were" analyzed.  The wastewater was analyzed for  total  suspended
solidS  (TSS),  cadmium, zinc, arsenic, copper, uranium, ^olybde-
n£m, vanadium, chemical oxygen demand  (COD), PH, sulfate and both
total and dissolved radium 226.
Gold Placer Mining Study

Titanium  Sand Dredge Mining and Milling Study

qamnles   from   two  facilities  were analyzed for toxic organics,
?oxic  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-
ronventional  parameters  such  as chemical oxygen  demand, total
organic   Sarbon   total   phenolics   (4AAP),  iron,   manganese,
titanium, and oil  and  grease  were also measured.

Solid Waste Program
 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 °f
          industrial Effluents for Priority Pollutants" (published
          bv EPA Environmental Monitoring and Support Laboratory,
          Cincinnati, Ohio, March 1977, revised April 1977).

      •>   "Analytical Methods for the Verification Phase of the
          BSSSviS  (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).
                                136

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           Standard Methods for  the Examination of Water  and Waste
          water   (published by the American Public Health
          Association, Washington, B.C.,  14th edition,  1976).

          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).                '

          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
      r?/M?rch.Institute for analysis of organic compounds by
      GC/MS using the liquid/liquid extraction method.)
      6.
 The choice
 conducted,
 programs.
             of   laboratories  depended  on  the  analysis  to  be
             and  laboratories were changed for different sampling
 of
 Detection  limits  (the  lowest  concentration at which  a  parameter
 can  be  quantified)  vary  between  sampling  programs arid even within
                  detection limit  (DL)  for a  parameter depends on
                instrument  used,  the  range of  standards  each  set
              analyzed,   the   complexity of the sample matrix,  and
              of  the  instrument    response.     Therefore,    the
              nuts given  herein are only indicators.of the  minimum
     i.  ^ L  ,  concentrations.   In  fact,   some .data   points  are
 reported below  the  listed  detection  limit.             points  are

 Screen, Verification and Verification Monitoring  Programs

 The  analysis   methods,  laboratories,  and   detection limits  for
 samples ana yzed during these programs  are given   in   TabTes  V-3
 four programs )        Parameters listed  were  not  analyzed  in  all


 Surveillance and Analysis Program

As discussed, this program was  conducted  by   the  EPA  r
 laboratories.   The methods used were EPA approved and the
tion limits are approximately those shown in Table V-4.

Cost Site Visit Sampling Program
         !it€ Y1! ifc 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 ProgFiHis" - ~ - ~~  -    9  Mining  —
                                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 commie ial
 laboratories evaluated the analytical recovery  of  cyaniSe  from
            n<3  *0"*  decant  samPles  which  had been spiked wi?h
                        deljver?d to  these  laboratories^  immedi-
                           to eliminate the possibility of cyanide
 in  able V-7       °f thiS qUality C°ntro1 Program ar*  P^sSntel

 A  study  of the analysis of cyanide in ore mining and processina
 ^onn™^/38-00"^0^ in '^operation with the American  M?n ing
 Congress  to  investigate  the causes of analytical interference!
 observed and to determine what effect these interferences had  on
            10"  °f •  the  analvtical  method.   Samplefc? five o?e
            processing wastewaters were  obtained  along  with  two
  v    h   i  J[astewater effluents.   Cyanide analyses werV performed
 EMSL g?ihi^?rat°r-eSA.inCluding  Six AMC  reprSsentativeS,   EPA^s
 laboratory      y  ""  Cincinnati and Rad^n Corporation's chemica?


 and  Tmodfli^  ^ih ^V Wf^  tO  evaluate th^  EPA-approved method
 m^?- 2    fu  method for  the determination   of   cyanide.    The
 modified method  employed  a   lead  acetate   scrubber  to
            Krh8  produced d"ri"9 the  reflux-distillatio"0.
          have  been suspected  of providing  an  interference in
 color imetric  determination  of cyanide   concentra :ion2     AI
 several  samples  were  spiked  with th?SJyanat2 to  SSer?!!n if th?4
 compound caused  interference in the cyanide analysis!
                          the resultant data shows no significant
when annp.rf  «  o  i     °^ accuracy of the two methods employed
wnen applied to metal ore mining and milling  wastewaters
prorm
                                                      —lyst  t
cohcetratio.
                                     initial and adjsted cyanide
                               139

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   Sample
       A
       B
       C
       D
       E
       F
       G
    Initial CN          Theoretical CN
 by Approved Method    After Adjustment
 (mq/1)	 (mq/1)	
—	0. 26        ~     0. 26 -(no adjustment)
     0.02
     0.02
       0.54*
     0.02
     84.00
     0.02
    0.210
    2.73
0.54 (no adjustment)
    0.389
 2.5 - 3.5**
    0.273
      * Sample showed strong interferences during color imetric
       procedure.
     **Results were not repeatable.
Sample  D  contained  approximately  180  mg/1  of
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  IhisS  data   the following conclusions have bee J 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/1  of thiocyanate.

 Based upon the relative standard deviations calculated,  it can be
 said that for an  ore ! mining  or  processing  wastewater  sample
 containing  0.2 mg/1 of cyanide, 95 percent of the ana^^/0"^
 be between 0.08 and 0.32  mg/1  using  the  modif led  *eth<£  and
 between  0 09  and  0.31  mg/1 using the approved method.  Over 99
 parcel? 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

-------
allow for analytical measurement  of   up
cyanide.  (See discussion Section X).

PCBs and Pesticides
to  .4  mcj/1  for  total
In  analyses of pesticides and PCBs, extraction efficiencies were
determined.   Known  standards  were  prepared/  extracted    Jnd
                       , finirm extra?ti£n/concentra?ion losses
                    analytical components were carried  out  with
to  err   *:       i*1*-  A calibration mixture was run daily
to  ensure  that  retention time and instrumentation response for
each parameter analyzed did not change due to column anfdeKctor
                               141

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                            Table V-l

                          TOXIC ORGANICS
Compound Name

  1.  *acenaphthene   (B)***
  2.  *acrolein       (y)***
  3.  *acrylonitrile  (V)
  4.  *benzene        (V)
  5.  *benzidene      (B)
  6.  *carbon tetrachloride  (tetrachloromethane)
                                             (V)
   *Chlorinated benzenes  (other  than  dtchlorobenzenes)

  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.
  11.
  12.
  13.
  14.
  15.
  16.
                     (V)
1,2-dichloroethane
1,1,1-trichlorethane
hexachlorethane   (B)
1,1-dichloroethane   (V)
1,1,2-trichloroethane
                       (V)
1,1,2,2-tetrachloroethane
chloroethane   (V)
                        (V)
                             (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)
  24.  *2-chlorophenol    (A)
                                  (V)
                                 142

-------
                       Table V-l  (Continued)

                           TOXIC  ORGANICS
   *Dichlorobenzenes

 25.  1,2-dichlorobenzene
 26.  1,3-dichlorobenzene
 27.  1,4-dichlorobenzene

   *Dichlorobenzidine
                            (B)
                            (B)
                            (B)
 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)

   *pichloropropane and  dichloropropene

 32.   1,2-dichloropropane   (V)
 ??'   i>2;dlchloropropylene (1,3-dichloropropene)
 34.   *2,4-dimenthylphenol   (A)

   *Dinitrotoluene

 35.   2,4-dinitrotoluene    (B)
 36.   2,6,-dinitrotoluene   (B)
 37.   *l,2-dipheriylhydrazine    (B)
 38.   *ethylbenzene    (V)
 39.   *fluoranthene    (B)

  *Haloethers (other  than  those  listed elsewhere)

40.   4-chlorophenyl phenyl ether    (B)
41.  4-bromophnyl phenyl ether    (B)
42.  bis(2-chloroisopropyl) ether   (B)
43.  bis(2-chloroethoxy) methane    (B)

  *Halomethanes (other than those listed elsewhere)
                                                    (V)
44,
45,
46.
47.
48.
     methylene chloride (dichloromethane)
     methyl chloride (chloromethane)   (V)
     methyl bromide (bromomethane)   (V)
     bromoform (tribromomethane)   (V)
     dichlorobromomethane   (V)
(V)
                              143

-------
                     Table V-l (Continued)

                         TOXIC ORGANICS
     trichlorofluoromethane   (V)
     dichlorodifluoromethane   (V)
     chlorodibromomethane   (V)
     *hexachlorobutadiene   (B)
53.  *hexachlorocyclopentadiene   (B)
54.  *isophorone   (B)
55.  ^naphthalene   (B)
56.  ^nitrobenzene    (B)
49
50
51
52
  *Nitrophenols (including 2,4-dinitrophenol  and  dinitrocesol)
57.  2-nitrophenol    (A)
58.  4-nitrophenol    (A)
59.  *2,4-dinitrophenol    (A)
60.  4,6-dinitro-o-cresol    (A)

  *Nitrosamines

61.  N-nitrosodimethylamine    (B)
62.  N-nitrosodiphenylamine    (B)
63.  N-nitrosodi-n-propylamine   (B)
64.  *pentachlorophenol    (A)
65.  *phenol    (A;
   *Phthalate  esters

 66.   bis(2-ethylhexyl)  phthalate
 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)
                                    (B)
   *Polvnuclear aromatic hydrocarbons
 72.
 73.
 74.
 75.
 76.
 77.
 78.
 79.
 80,
 81.
      benzo (a)anthracene  (1,2-benzanthracene)    (B)
      benzo (a)pyrene (3,4-benzopyrene)    (B)
      3,4-benzofluoranthene    (B)
      benzo(k)fluoranthane (11,12-benzofluoranthene)
      chrysene   (B)
      acenaphthylene    (B)
      anthracene    (B)
      benzo(ghi)perylene  (1,12-benzoperylene)    (B)
      fluorene    (B)
      phenathrene    (B)
(B)
                                144

-------
                      Table V-l (Continued)

                          TOXIC ORGANICS


 82.   dibenzo  (ah)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.  4,4l-DDD(p,p'TDE)    (p)

  *endosulfan and metabolites
                          (P)
                          (P)
                          (P)
  95.   a-endosulfan-rAlpha
  96.   b-endosulfan-Beta
  97.   endosulfan sulfate

   *endrin  and  metabolites

  98.   endrin    (p)
  99.   endrin  aldehyde      (p)

   *heptachlor  and metabolites

100.   heptachlor    (p)
101.   heptachlor epoxide    (P)

   *hexachlorocyclohexane (all  isomers)

102.  a-BHC-Alpha   (P)  (B)
103.  b-BHC-Beta   (P)  (y)  •
104.  r-BHC (lindane)-Gamma   (P)
105.  g-BHC-Delta   (P)
                              145

-------
                      Table V-l (Continued)

                          TOXIC ORGANICS
   *polychlorinated biphenyls (PCB's)
106.
107.
108.
109.
110.
111.
112.
PCB-1242
PCB-1254
PCB-1221
PCB-1232
PCB-1248
PCB-1260
PCB-1016
(Arochlor
(Arochlor
(Arochlor
(Arochlor
(Arochlor
(Arochlor
(Arochlor
1242)
1254)
1221)
1232)
1248)
1260)
1016)
                                 (P)
                                 (P)
                                 (P)
                                 (P)
                                 (P)
                                 (P)
                                 (P)
113.  *Toxaphene   (P)
114.  **2,3,7,8-tetrachlorodibenzo-p-dioxo.n
  *Specific compounds  and  chemical  classes as listed in the
***B
   V
 compoun      specifically listed  in  the  consent  degree,
analyzed in the base-neutral extraction fraction
analyzed in the volatile organic  fraction
         in the acid extraction fraction
   A = analyzed
                                 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
toUl lutptndad totidt
volatile sutpt nd«d lolidt
COD
TOC
radium 226 (tout)
radium 226 (dissolved)
antimony ttoUl)
•rMnic (toul)
tayHium {toul)
cadmium (total)
chromium {toul}
copp«r {to til)
ltd (toul)
mercury (tout)
nickel (to til)
it It mum (total)
lilvir (toul)
thallium (total)
zinc {toul)
atbenot (fibrout)
cyanide dotal)
phtnol (toul)
aldrm
ditldrn
ehlordane
(tectinicat mixture and meubolitef)
4.4'.OOT
4.4*.DOE (p^'-DDX)
4.4'.DDD(pj)'.TDei
Cf -«ndoiu1f an
/J-endotulfan
tndoiuUan lulfate
endrm
f ndrm aldehyde
heptichlor
heptichlor epoxxJe
CT-BHC
0-BHC
7 BHC (Ittxlane)
6-BHC
PCS 1242 (Arochlor 1242)
PCS 12&4 (Arochioi 1254)
acenaph thine
acrolem
aciyloniirile
benzene
bentidine
carbon telrachtonde
(tetrachloiomethane)
chlorobeniene
1,2.4 tnchlorobenzene
hexachlorobenzene
1,2
-------
TABLE V-4.
total impended tolidi
COD
foe " ~
radium 226 (total) "~
radium 226 (diuolved)
arwnic (total) : "~
beryllium (total)
cadmium (total)
chromium (total)
copper (total) "
lead (total)
mercury (total)
selenium (total)
thallium (total) :
zinc (total)
cyanide (total)
phenol (total)
aldrin
diefd rin .
(technical mixture and metabolites
4,4'-DDT
4,4'-DDE (pj)'-DDX)
4.4'-DDD tp,p'-TDE)
GT-endoiulfan
endrin
endrin aldehyde
CT-BHC '• "
>-BHC(lindane)
5-BHC ~~
PCS 1254 (Arochlor 1254)
acenaphthene
acrolein
acrylonitrile
benzene
benzidine
carbon tetrachlonde
(tetrachloromethane)
chlorobenzene
1 ,2,4-trichlorobenzene
hexachlo[obenzene
1,2-dichloroethane
hexachloroethane
1,1-dichloroethane
,1.2-trichloroethane
1,1,2,2-tetrat:hloroethane
bis (chloromethyl) ether
bis (chloroethyl) ether
2-chlorophenol
1.2-dichlorobenzene
LIMITS OF DETECT
1mo/l
1 mg/l
2mg/l
	 1 mp/l 	
1 pCi/l
IpCi/l
0.002 mg/l •
0.005 mg/l
0.002 mg/l
0.02 mg/l
0.01 mg/l
0.05 mg/l
0.0005 mg/l ••
0.02 mg/l
0.002 mg/l •
0.01 mg/l
0.1 mg/l
0.005 mg/l
2.2 X 105fibers/l
0.02 mg/l
0.002 mg/l
	 0.1 H9/I 	
	 0-s«a"
	 as/! 	
	 as/! 	
nan
	 as/! 	
	 as/! 	
	 m/' 	
	 as/! 	
O.Sug/l
0.5 ug/l
0.1 ug/l
0.1 ug/l
0.1 ug/l
0.1 ug/l
0.1 ug/l
0,1 ug/l
lug/i
lUS/l
0.03 ug/l
no eitimite
	 no eitimate 	 _,
0.04 ug/l
	 ~&.0U,/I 	
0.35 uo/l
0.15 uq/l
~ 1.0ug/l
	 ~ 1.0mi/l 	
0.08 Wg/l
0.15 |jg/l
~0.10^s/l
0.15 ^g/l
0.10 us/1
0.90 jig/I
~ 0.50 Jig/I
no estimate
no e>t mate
~ 0.50 (ig/l
0.05 0g/l [-
~ 1.0 Wg/l
— 1.0 /ig/l
0.05 «g/l
1.00 Ug/l
0.10O.20^g/l
ON FOR POLLUTA
3,3'-dichlorobenzidine
1,1-drchloroethylene
1 ^-trant-dichloroethylene
(1,3-dichloropropene)
2,4-d in itro toluene
fluorantherw
4-chlorophenyl phenyl ether
bis (2-chloroiiopropyl) ether
bis (2-chloroethoxy) methane
methylene chloride (dichloromethane
methyl chloride (chloromethane)
bromoform (tribromomethane)
drchlorodifluororomethane
hexachlofocy dope n tad iene
J -nitrosodrphenylamine
N-nitrosodi-n-propylamine
pentachlorophenol
bis (2-ethylhexyl) phthalate
i-n-butyl phthalaie
i methyl phthalate
1 ,2-benzanthracene
benzo (a) pyrcne (3,4-benzopyrene)
fluorene
1 ,2.5,6-dibenzanthracene
mdeno (1,2,3-c^l) pyrene
2,3,7 ,8-letrachlorodibenzo-p-dioxin
(TCDD)
tetrachloroethylene
toluene "
tnchloroethylene
vrnyl chloride

NTS ANALYZED
0.100.20 jig/1
0.1OO.20 «o/l
	 ~ 2S.O un/l 	 	
	 O.I50 ug/l 	
	 o-:»s u°/i 	
	 ~ I.OUQ/I 	 	
0.1)2 utt/l
O.JIS «g/l
	 O.«0ug/l 	
o.;n fig/i
0.75 pg/l
o.;.o us/i
0.10 /ig/l
0.20 «ig/l
0.06 /JO/I
~ 0.50 ps/l
no imimite
no intimata
0.3B 
-------
                            Table V-5
      LIMITS OF DETECTION FOR POLLUTANTS ANALYZED FOR COST -
               - SITE VISITS BY RADIAN, CORPORATION
Parameter
Sb
As
Be
Cd
Cr
Cu
Fe
Pb
Hg
Mn
Ni
Se
Ag
Tl
 Zn
Method of Analyses
AAS* - hydride generation
AAS  - hydride generation
AAS - flameless - HGA**
ICPES***
ICPES
ICPES   ;
ICPES  .  . .           	
AAS -  flameless - HGA
AAS -  flameless - cold vapor
ICPES
ICPES
AAS  -  hydride generation
 ICPES
AAS  -  flameless - HGA
 ICPES
Detection
Limit (ppm)
  0.005
  p.002
  0.001
  0.005
  0.005
  0.005
  0.004
  0.002
  0.001
  0.001
  0.020
  0.005
  0.005
  0.002
   0.002
   *Atomic Absorption Spectrophotometry
  **Inductively Coupled Argon Plasma Emission Spectrometry
 ***Heated Graphite Analyzer
                                150

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TABLE V-6. COMPARISON OF SPLIT SAMPLE ANALYSIS* FOR CYANIDE
          BY TWO DIFFERENT LABORATORIES USING THE BELACK
          DISTILLATION/PYRIDINE-PYROZOLONE METHOD
Sample Description
1. Tailing Pond Influent
2. Tailing Pond Influent
3. Tailing Pond Influent
4. Tailing Pond Effluent
5. Tailing Pond Effluent
6. Tailing Pond Effluent
Analytical Result (mg Total CN/I )
Lab#1
0.02
0.05
0.08
0.04
0.04
0.03
Lab #2
0.06
<0.02
0.03
0.50
0.47
0.57
    •All samples collected at the same time at Mine/Mill 3121
                          151

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TABLE V-7. ANALYTICAL QUALITY CONTROL PERFORMANCE OF COMMERCIAL
          LABORATORY PERFORMING CYANIDE* ANALYSES BY EPA APPROVED
          BELACK DISTILLATION METHOD
SAMPLE DESCRIPTION
Distilled H2O -I- 0.1 mg/l CN
as NaCN
Distilled H2O + 0.2 mg/l CN
as NaCN
Distilled H2O + 0.4 mg/l CN
as NaCN
Tailing Pond Decant (no spike)
Tailing Pond Decant Spiked with
0.1 mg/l CN as NaCN
Tailing Pond Decant Spiked with
0.2 mg/l CN as NaCN
Tailing Pond Decant Spiked with
0.4 mg/l CN as NaCN
Tailing Pond Decant Spiked with
1.0mg/ICNasK3Fe(CN)6
Tailing Pond Decant Spiked with
1.0 mg/l CN as K4Fe(CN)6
CYANIDE AS
CN mg/l
0.125
0.250
0.200
0.069
0.144
0.119
0.136
0.028
0.032
0.027
0.079
% ANALYTICAL
RECOVERY OF SPIKE
125
125
50
-
85
44
29
3
3
3
7
*AII cyanide analyses are total cyanide
                                152

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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.

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Values of pH vary, but are often in the alkaline range  (7 to 14).
This  is  because several mill processes operate at 'elevated pHs.
As indicated by discussions in Section III, pH, TSS,  and  metals
values  are closely allied.  The solubility of many metals varies
greatly with pH, and the  status  of  the  metals   (dissolved  v.
solubilized) affects the concentration of TSS.  This relationship
is  used by the industry for ore beneficiation and for  wastewater
treatment.

Nonconventional parameters such as COD, TOC,  volatile  suspended
solids  (VSS), and iron were also analyzed for many samples.  The
concentrations of the organic related parameters, COD,  TOC,  and
VSS,  were  always  low.   Any  organic  compounds  added in mill
processes are not indicated by these tests which are designed  to
measure relatively large masses of organics (in the mg/1 range at
a  minimum).   Iron is common in metal ores and the summary table
reflects this.

The entire BAT  data  set  is  discussed  below  by  subcategory,
subdivision,  and  as  a mill process or mine drainage, and these
discussions more completely characterize  mine/mill  wastewaters.
In general  it can be noted from Table VI-1 that organic compounds
are  not  the  major  concern in this category (a point discussed
thoroughly  in Section III), metals are prevalent, pH  values  are
generally alkaline, and cyanide and asbestos are often  present.

Iron Subcategory, 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
                               1R6

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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/Molybdenum
Drainage Subdivision
   Subcategory,
Mine
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/Molybdenum
Mill Process
Subcategory, Cyanidation
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/Molybdenum
Subdivision, Froth Flotation Mill Process
   Subcategory,
Mill
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

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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/Molybdenum     Subcateqory,
Subdivision, Heap/Vat/Dump/In-Situ Leaching
                  Mill
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/Molybdenum
Operations Recovering Gold
Subcategory,
Placer
A  study  was  conducted  in  1978 to evaluate current wastewater
handling practices at gold placer mines.  Eleven operations,  all
located in Alaska, were sampled to determine performance capabil-
ities  of  existing settling ponds.  Only two of the toxic metals
were monitored during the program, arsenic and mercury.   Settle-
able  solids  were  also  monitored  to  provide an indication of
treatment pond performance.  As can be seen in  Table  VI-8,  the
settleable  solids  concentrations range from not detected to 500
ml/l/hr.  However, many of the different samples  are  discharges
that had not been treated in settling ponds.

Aluminum Subcategory, Mine Drainage Subdivision

As  shown  in  Table  VI-9, aluminum mine drainage is low in most
pollutants.  The toxic metals present in  the .discharge  are  in
relatively  low  concentrations  (compared to Table VI-1) and are
chromium,  copper,  mercury,  nickel,  and  zinc.   Asbestos  was
present  in  moderate concentrations (compared to Table VI-1) and
was not affected by the  existing  treatment  methods.   Acid  pH

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levels were noted in the raw, but these increased to the alkaline
range (7 pH 14) after pH adjustment.

Tungsten Subcategory, Mill Subdivision

As  shown in Table VI-10, 13 of the toxic metals were detected in
the raw wastewater.  However, these are reduced during  treatment
leaving  only  seven  above  20 ug/1 in the discharge.  Of these,
copper,  lead, and zinc have high concentrations (compared to  the
other discharge metals concentrations).

Asbestos  and phenolics were detected in the raw samples; cyanide
was not.  The  values  of  asbestos  are  high  relative  to  the
category as a whole (see Table VI-1).  The effluent phenolics are
low relative to the values in Table VI-1.

Mercury Subcategory, Mill Subdivision

As  seen  in  Table  VI-11,  the  toxic  metals are found in high
concentrations in the raw wastewater in this subdivision, as  are
asbestos  and  phenolics.  That is why the applicable BPT regula-
tion is no  discharge  of  process  wastewater.   The  discharged
sample in Table VI-11 is actually treated recycle water.

Uranium Subcategory, Mine Drainage Subdivision

Uranium  mine drainage, is, relative to mill water less polluted.
As seen in Table VI-12, many of the toxic metals  were  detected,
all  but zinc in concentrations less than 65 ug/1.  Only six were
detected in the treated samples, none greater than 50 ug/1.

Cyanide was not detected, and phenolics were detected  at  a  low
concentration  (10  ug/1).  Asbestos was detected in both raw and
treated samples at moderate concentrations (as compared to  Table
VI-1).

Not  listed  in  Table  VI-12,  but  shown  in  the  support data
(Supplement A), are radium 226 concentrations.   Uranium  ore  is
radioactive  and  radium  226 is a radionuclide always associated
with uranium.  It is one of the uranium decay series  and  has  a
half  life  of  1,620  years.   Raw  mine  water may have several
hundred to a thousand pico-Curies per liter (p Ci/1) of  Ra  226,
but  existing  treatment  is  capable of reducing this to the BPT
guideline of 10 p Ci/1 (total, 30-day average).

Uranium Subcategory, Mill Subdivision

As seen in Table VI-13, several of the toxic metals are found  in
both  raw  and  treated  wastewater.   Treated wastewater in this
table  is  actually  recycle  water.    The  facilities  do   not
discharge.   This  recycle  water is not treated specifically for
metals,  and, therefore, little reduction occurs.
                                  159

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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 Subcategory, Mill Subdivision

As shown in Supplement A (Support Data; Sample Points 1A and  2A,
for  Mill  9905),   seven toxic  metals  were detected in the raw
wastewater; all but selenium and lead at  concentrations  greater
than  200  ug/1.  These  concentrations were reduced by treatment,
leaving only five detected toxic metals ranging in  concentration
from 20 to 100 ug/1.

Asbestos  was  detected  at  moderate concentrations (compared to
Table  VI-1).   Cyanide  was  not  detected  and  phenolics  were
detected at 10 ug/1 in raw and discharged samples.

Titanium Subcategory, Mills with Dredge Mining Subdivision

Table  VI-15 summarizes  the data for the titanium mills employing
dredge mining.   Ten toxic metals were detected in the raw  water,
at  concentrations  less  than or equal to 80 ug/1.  In the treated
effluent, six toxic metals were detected.  Only zinc was detected
in concentrations greater than 10 ug/1.

COD and TOC  concentrations  in  the  raw  water  were  generally
present  in  higher  concentrations than the rest of the category
due to the presence of organic material in some of the ores.  The
treatment processes used substantially reduced the concentrations
of both COD and TOC.  The TSS concentration of the effluents were
less than 10 mg/1.

Vanadium Subcategory, Mine Drainage Subdivision

Table VI-16 illustrates  the character of vanadium mine  drainage.
Several  toxic metals were present both in the raw and discharged
water.   Discharge  concentrations  greater  than  20  ug/1  were
reported  for  chromium, copper, lead, nickel, and zinc.  Cyanide
and total phenolics were not detected.  The asbestos values  were
low relative to the category as a whole.
                               160

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Vanadium Subcateqory, Mill Subdivision

As  seen  in Table VI-17, many toxic metals were detected in both
the raw and discharged waters  from  this  subdivision.   Of  the
metals, only mercury was reduced below the detection limit by the
existing  treatment  system.   Cyanide was also reduced below the
detection limit, and no total phenolics were detected  in  raw  or
discharged water.

Antimony Subcateqory, Mill Subdivision

The  data  for  this  subcategory  are  presented in Table VI-18.
There is no discharge of treated wastewater from the single  mill
in  this subdivision.  Relatively high concentrations  of antimony
and arsenic are  present  in  the  raw  and  treated   wastewater.
Phenolics  were  not  detected  in the raw or treated  wastewater.
Asbestos was detected  in  moderate  concentrations  compared  to
Table VI-1 .  The pH of the impounded water was greater than 12.0.

REAGENT USE IjN 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.
                                16?.

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Chromium

Sodium  dichrornate  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

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Cresylic acid is used as a flotation reagent at Mills 2117, 2121,
and  4403.   Xylenols,  C2_H5_.C6H40H  or   (CH3_)2_.C6_H3_.OH,  are the
dominant constituents of commercial cresylic  acids  and  include
the  toxic  pollutant,  2,4-dimethyphenol,  which  has  not  been
detected in raw or treated wastewater samples a.t Mills  2117  and
2121.    Mill   4403  was  not  sampled   for  the  organic  toxic
substances.   Nitrobenzenes  are  present  in   Aero ... 633,   but
nitrobenzene  was not detected in wastewater during this program.
However,  screening  and  verification    sample   data   strongly
implicate these phenol-based flotation reagents as the sources of
total   phenol  (4AAP)  in  mill  process  wastewaters=.   From  a
practical standpoint, cresylic acid  can  be  considered  as  100
percent  phenolic with the relative phenolic content of the other
phenol-containing reagents  being  considerably  less.   Phenolic
concentrations of 5.2 mg/1 and 5.0 mg/1 have been detected in the
mill  tailing  samples at Mill 2117, and  treated effluent samples
were found to contain 0.30 mg/1 and 0.36  mg/1  on  2  consecutive
days.  The large consumption of cresylic  acid at Mill 2117 (0.035
kilogram/metric  ton  equivalent to 0.070 pound per short ton, of
ore) and the consistency of data substantiate  cresylic  acid  as
being  a  significant  source  of phenolic compounds in flotation
mill process effluents.

Phenolic compounds were found to  be  the  most  prevalent  toxic
organic  species  detected  in the screen samples, but concentra-
tions did not exceed 0.03 mg/1 except  at  operations  which  are
known  to  employ  one  or  more  of  the  phenol based flotation
reagents previously discussed.

SPECIAL PROBLEM AREAS

Backfilling of Mines With Mill Sand Tailings

A review of sample data and historical monitoring  data  supplied
by   the   industry   indicates   the   presence  of  significant
concentrations of  cyanide  in  several   mine  water  discharges.
Further  examination revealed that the facilities with cyanide in
mine water backfilled mined-out stopes using mill  sand  tailings
from  flotation  circuits  which use cyanide compounds as process
reagents.

A variety of undergound mining techniques are used throughout the
mining industry.  Typical mining methods  include room-and-pillar,
vein (or drift) mining, open stoping,  pillar  stoping,  cut-and-
fill,   and   panel-and-fill.   The  selection  of  method(s)  is
dependent on many factors, such as the type and shape of the  ore
deposit, the depth of excavations, and the ground conditions.

Cut-and-fill,  pillar stoping, and panel-and-fill techniques have
found common application in lead, zinc, and silver mines  located
in  Colorado,  Utah,  and  the  Coeur  d'Alene Mining District in
Idaho.   An inherent  feature  of  these  mining  methods  is  the
refilling  of  worked-out and abandoned stopes and other workings
                                164

-------
to prevent subsidence and cave-ins as mining  progresses   through
the  ore  body.   For  many  years,  waste  rock  from  the  mine
exploration crosscuts was used as  fill  material;  however,  the
development  of  hydraulic sandfill procedures has simplified the
backfill operation.   In  current  practice,  the  coarse   (sand)
fraction  of the flotation-mill tailings is often segregated from
the tailings pulp stream by hydro-cyclones and  pumped  into  the
mine for backfilling.

Nine  mines (Mines 3107, 3113, 3120, 3121, 3130, 4104, 4105, 4401
and 4402) are known to practice hydraulic backfilling  with  mill
sand tailings.  Eight of these nine mills use cyanide either as a
flotation reagent (Mills 3107, 3113, 3121, 3130 and 4401)  or as a
leaching  agent  (Mills 4104, 4105, and 4402).  The nature of the
mechanism by which cyanide depresses  pyrite  and  sphalerite  is
such  that  much  of  the  cyanide added to the flotation  circuit
associates with the depressed minerals in- the tailings and ulti-
mately is leached into mine water during hydraulic backfill.

Mine  3130  is  the  only  facility with a separate mine drainage
treatment system that periodically monitors for cyanide.   Efflu-
ent  monitoring data (summarized in Section VIII) include  cyanide
analyses of five 24-hour composite samples collected  during  the
period of June 1!977 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 cyemide 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.
                               IfiB

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-------An error occurred while trying to OCR this image.

-------
      Table  VI-11
        DATA SUMMARY
       ORE MINING DATA
SUBCATEGORY MERCURY
SUBDIVISION MILL
MILL PROCESS FLOTATION (FROTH)
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL) 1 1
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
_ VSS
00 TSS
0 TOC
PH (UNITS)
PHENOLICS (4AAP)
1
1
1
1
1
0
1
1
1
0
1
1
1
1
1
1
1
1
1
ASBESTOS (CHRYSO) (F/L) 1 1
TOTAL FIBERS (F/L) 1 1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
53
1.1
0.09
0.56
0.46
0.85

1
230
1.6

0.01
0.2
2.4
60
4300
139000
21
8
0.92
150OE8
1300E9
53
1.1
0.09
0.56
0.46
0.85

1
230
1.6

0.01
0.2
2.4
60
4300
139000
21
8
0.92
1500E8
1300E9
53
1.1
0.09
O.56
0.46
0.85

.1
230
1.6
;
0.01
0.2
2.4
60
4300
139000
21
8
0.92
1500E8
1300E9
53
1.1
0.09
0.56
0.46
0.85

1
230
1.6

0.01
0.2
2.4
60
4300
139000
21
8
0.92
150OE8
1300E9
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
*
*
*
*
*
*
*
*
*
*
*
*
£


<



0
0
1
0
0
0
0
* 1 1
* 1 1
* 1 0
* 1 1
* 1
* 1
* 1
* 1
* 1
TREATED (MG/L)
DETECTED
MEAN MEDIAN
0.2
0.11

0.008
0.015
0.05


0.05




0.04
22

16
13
8.3
0.22
5700E4
7700E5
0.2
0.11

0.008
0.015
0.05


0.05




O.O4
22

16
13
8.3
O.22
570OE4
7700E5
VALUES ONLY
90% MAX
0.2
0.11

O.OO6
0.015
0.05


0.05




0.04
22

16
13
8.3
O.22
57OOE4
77OOE5
0.2
O.11

0.006
O.015
0.05


0.05




0.04
22

16
13
8.3
0.22
5700E4
7700E5

-------
        Table  VI-12

        DATA SUMMARY
       ORE MINING DATA
SUBCATEGORY URANIUM
SUBDIVISION MINE
MILL PROCESS MINE DRAINAGE
RAW(M6/L)
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MEDIAN 90% MAX
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOL I CS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
3
17
3
16
4
14
3
4
3
4
5
3
3
17
15
2
18
2
13
3
1
3
2
1
16
0
13
3
14
0
3
1
1
3
0
0
17
15
2
18
2
13
1
1
3
2
0.05
0.0195

.00381
.04333
.01673

0.09
0.0038
0.06
.02333


.04306
22 . 504
23.5
144.58
8.5
7.6519
0.01
0.319
1050E5
1950E6
0.05
0.007

0.003
0.045
0.0075

0.05
O.O038
O.OS
0.028


0.02
7
23.5
21
8.5
8.05
0.01
0.319
1 100E5
1950E6
0.05
0.0832

0.0092
0.05
0.075

0.18
O.O038
O.OS
0.037


0.158
104.2
28
415.94
9
8.655
0.01
0.319
1900E5
2300E6
0.05
0.17

0.01
0.05
0.11

0.18
0.0038
0.08
0,037


0.19
140.5
28
1639.5
9
8.825
0.01
0.319
1900E5
230QE6
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
TREATED (MG/L)
NUMBER OF NUMBER DETECTED
SAMPLES DETECTED MEAN MEDIAN
3
13
3
13
3
11
3
3
3
3
5
3
3
13
12
2
13
2
9
3
1
2
2
0
11
0
10
2
8
0
1
1
0
3
0
0
12
12
2
13
1
9
1
1
2
2

.00798

0.0038
O.0425
.00575

0.05
0.0091

. 03633


.01983
10. 169
1.5
33.185
10
7.8833
0.01
0.054
4000E4
5000E5

0.006

0.003
0.0425
0.006

0.05
0.0091

0.048


0.014
8.95
1.5
27
10
7.9
0.01
0.054
4000E4
5000E5
VALUES ONLY
90% MAX

0.0228

0.0069
0.06
0.011

0.05
O.O091

0.051


0.0666
33.5
2
75.8
10
8.5
0.01
0.054
5300E4
5700E5

0.024

0.007
O.O6
0.011

0.05
0.0091

0.051


0.078
38
2
83
1O
8.5
0.01
0.054
5300E4
5700E5

-------
                                                          Table VI-13
00
ro
                                                          DATA SUMMARY
                                                         ORE MINING DATA
                                                  SUBCATEGORY URANIUM
                                                  SUBDIVISION MILL
                                                  MILL PROCESS ARID LOCATIONS
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS) .
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/D
TOTAL FIBERS (F/L)
4
10
6
12
8
12
2
8
4
8
6
6
4
12
5
1
5
1
6
2
7
1
1
2
9
2
11
8
10
1
5
1
8
6
3
2
12
5
1
5
1
6
2
5
1
1
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
0.516
4.2602
0.274
14791
1.738
0.9966
0.046
1 . 9076
0.036
2 . 3422
0. 1705
0.069
1.205
26.176
95 . 206
20
19134
24
6.43
0.0085
1462.1
2300E4
2900E5
0.516
0.243
0.274
0.1
1.575
0.485
0.046
1.3
0.036
2.835
0. 1525
0.056
1.205
22 . 365
26
20
64
24
7.45
0.0085
1660
2300E4
2900E5
1.03
10.6
0.295
0.4068
3.7
3.4
0.046
4.18
0.036
3.68
0.49
0.1
1.24
59.13
386
20
95450
24
8.3
0.01
2040
2300E4
2900E5
1.03
10.6
0.295
0.423
3.7
3.4
0.046
4.18
0.036
3.68
0.49
O.t
1.24
60.9
386
20
95450
24
8.3
0.01
2040
2300E4
2900E5
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*


NUMBER OF NUMBER
SAMPLES DETECTED
5
12
7
13
10
14
3
10
5
10
7
7
5
14
7
2
9
2
9
3
7
2
2
3
11
3
11
5
11
0
5
1
8
6
4
2
13
6
2
9
2
9
2
5
2
2
TREATED
(MG/L)
DETECTED
MEAN MEDIAN
0.299
.11518
0.0072
.03545
0.0406
0.192

0.3888
0.14
.82637
.06383
0.016
0.79
4.729
59 . 505
6
55.611
21.5
6.65
0.01
1.4164
1750E5
1750E6
100E-5
0.029
0.01
0.029
0.028
0.1

0.2
0.14
0.955
0.0215
0.0195
0.79
2.52
10.5
6
26
21.5
7.7
0.01
0.4
1750E5
1750E6


VALUES ONLY
90% MAX
0.895
0.65
0.011
0.0746
0.1
0.84

0.959
0.14
1.28
0.213
0.023
0.84
11.06
279
10
157
27
8.45
0.01
3.87
2000E5
2300E6
0.895
0.75
0.011
0.077
0.1
0.9

0.959
0.14
1.28
0.213
0.023
0.84
11.1
279
10
157
27
8.45
0.01
3.87
2000E5
2300E6

-------
                                                           Table  VI-14
                                                           DATA SUMMARY
                                                          ORE MINING  DATA
                                                   SUBCATEGORY TITANIUM
                                                   SUBDIVISION MINE
                                                   MILL PROCESS MINE  DRAINAGE
RAW(MG/L)
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MEDIAN 90% MAX
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
VSS
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
*
*
* NUMBER OF NUMBER
* SAMPLES DETECTED
t
*
*
*
*
#
*
* -
*
*
*
*
*
*
*
*
*
*
*.
*
0
0
0
0
0
1
0
1
0
0
0
0
0


(I
0



* 1
* 1
TREATED (MG/L)
MEAN





O.02

0.02





0.02
2


8
7.95
0.03
140000
1900E3
DETECTED
MEDIAN





0.02

0.02





0.02
2


8
7.95
0.03
140000
19OOE3
VALUES
90%





0.02

0.02





0.02
2


8
7.95
0.03
140000
1900E3
ONLY
MAX





0.02

O.O2





O.O2
2


8
7.95
O.O3
140000
1900E3
00
CO

-------
20
                                                          Table VI-15
                                                           DATA SUMMARY
                                                          ORE MINING DATA
                                                    SUBCATEGORY TITANIUM
                                                    SUBDIVISION MILLS WITH DREDGE MINING
                                                    MILL PROCESS PHYSICAL AND/OR CHEMICAL
RAW(MG/L)
NUMBER OF NUMBER DETECTED VALUES ONLY
SAMPLES DETECTED MEAN MEDIAN 90% MAX
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
COD
TSS
TOC
PH (UNITS)
PHENOLICS (4AAP)
IRON (TOTAL)
ASBESTOS (CHRYSO) (F/L)
TOTAL FIBERS (F/L)
9
9
9
9
9
9
6
9
9
9
9
9
9
9
6
9
6
9
6
9


1
3
0
0
7
9
0
4
2
3
3
5
0
9
6
9
6
9
5
9


0.002
. 00867
.04743
. 02733

0.0375
0.006
0.023
0.031
O.O07

.03122
1O76.7
341.44
485
5.9
0.0066
3.1924


0.002
0.009
0.03
0.016

0.042
0.006
0.023
0.029
0.009

0.021
1060.5
160
560
5.7
0.007
1.928


0.002
0.01
0.08
0.063

0.058
0.011
0.033
0.036
0.011

0.071
1900
1100
750
6.6
0.007
6.287


0.002
0.01
0.08
0.063

0.058
0.011
0.033
O.036
0.011

0.071
1900
1100
750
6.6
0.007
6.287


*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
TREATED (MG/L)
NUMBER OF NUMBER DETECTED
SAMPLES DETECTED MEAN MEDIAN
9
9
9
9
9
9
9
9
9
9
9
9
9
9
9
9
9
9
8
9
1
1
0
0
0
1
0
5
0
1
1
0
0
2
0
8
9
8
8
9
1
9
1
1
0.002

0.0058

0.005
100E-5


0.003

. 02675
14
3 . 5625
5.1875
5 . 9444
0.01
.20533
3300E3
2700E3
0.002

0.006

0.005
100E-5


0.003

0.008
14
2.9
5.25
6.8
0.01
0.171
3300E3
2700E3
VALUES ONLY
90% MAX
0.002

0.008

0.005
100E-5


0.003

0.071
17
9
6
7.6
0.01
0.5
3300E3
2700E3
0.002

0.008

0.005
100E-5


0.003

0.071
17
9
6
7.6
0.01
0.5
3300E3
2700E3

-------
                                                              Table VI-16

                                                             DATA SUMMARY
                                                             ORE  MINING DATA
                                                           SUBCATEGORY  VANADIUM
                                                             SUBDIVISION MIME
                                                             MILL PROCESS HO MILL PROCESS
                                                     RAW(UG/L)
                                                                                                TREATED (UG/L)

NUMBER OF
SAMPLES

NUMBER
DETECTED

107.

MEAN

MED

907.

*
*

NUMBER OF
SAMPLES

NUMBER
DETECTED

10%

MEAN

MED

901
00
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
PHENOLICS (4AAP)
IRON (TOTAL-MG/L)
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
18
130
16.333
16.833
120
41.833
317.5
1
446.67
6.3333
3
2
1476.7
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
                                                  69.133  69.133  69.133  69.133
1
1
1
1
1
1
0
1
0
1
1
0
1
1
0
1
                                                                                                               2       2
                                                                                                               5       5
                                                                                                               1       1
                                                                                                             8.2     8.2
                                                                                                          29.333  29.333
                                                                                                            20.5    20.5
                    2       2
                    5       5
                    1       1
                  8.2     8.2
               29.333  29.333
                 20.5    20.5
                                                                                                          171.33  171.33   171.33  171.33
59.333   59.333
    12      12

     1       1
159.17   159.17
59.333   59.333
    12       12

     1       1
159.17   159.17
                                                                                                          .86467  .86467   .86467  .86467

-------
                                                         Table  VI-17
                                                          DATA SUMMARY
                                                         ORE MINING DATA
                                                   SUBCATEGORY VANADIUM
                                                   SUBDIVISION MILL
                                                   MILL PROCESS FLOTATION  (FROTH)
00
RAW(MG/L)
NUMBER OF NUMBER
SAMPLES DETECTED
ANTIMONY (TOTAL)
ARSENIC (TOTAL)
BERYLLIUM (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TOTAL)
COPPER (TOTAL)
CYANIDE (TOTAL)
LEAD (TOTAL)
MERCURY (TOTAL)
NICKEL (TOTAL)
SELENIUM (TOTAL)
SILVER (TOTAL)
THALLIUM (TOTAL)
ZINC (TOTAL)
PHENOLICS (4AAP)
IRON (TOTAL)
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
1
3
2
3
3
3
3
3
0
3
DETECTED VALUES ONLY
MEAN MEDIAN 90% MAX
. 03733
. 29889
0.041
. 18456
1.6493
4.2113
0.29
5.7773
O. 1425
.78883
. 86767
.03089
.29211
34.165

135.24
.04167
. 37333
0.038
0.0245
.47117
.06533
O.29
0.3175
O.1425
. 44667
0.14
.02667
0.002
1 . 4767

69.133
.05233
. 39333
.06867
.51233
4.3567
12.527
0.29
16.717
0.284
1 . 8207
2 . 4567
0.063
.87333
100.82
•*
334.9
.05233
. 39333
.06867
.51233
4.3567
12.527
0.29
16.717
0.284
1 . 82O7
2.4567
0.063
.87333
100.82

334.9
*
*


* NUMBER OF NUMBER
* SAMPLES DETECTED
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
O
3
O
3
3
2
3
3
0
3
TREATED
(MG/L)
DETECTED
MEAN MEDIAN
.02067
O.I 08
.05317
.02384
.14189
.03669

. 55928

. 13636
. 08367
0.009
.11667
. 1 1664

.55572
0.014
0.014
0.0.36
O.O25
.06133
.04225

0.3265

.09475
0.079
0.009
0.0015
O.1115

0.5505


VALUES ONLY
90% MAX
0.046
0.305
0.1225
.03833
0.335
.04733

1.18

0.255
0.16
0.016
O.3475
.15917

. 86467
0.046
0.305
0.1225
.03833
0.335
.04733

1.18

0.255
0.16
O.016
0.3475
.15917

. 86467

-------
                      TABLE VI-1 8  SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
REAGENT

Lime
Caustic soda
Soda' Ash
Sulfuric Acid

Copper Su If ate
Copper Ammonium Chloride
Sodium Suif hydrate
DESCRIPTION

CaOorCa(OH)2
Sodium Hydroxide,
NaOH
Sodium Carbonate,
Na2 C03
H2S04

CuSO* or
CuSO]j • 5H2O
CuNH2 Cl
Sodium Hydrosuif ide
Na SH • 2H2O
FUNCTION
MODIFIERS
Alkaline pH regulator and depressant for
galena, metallic gold, iron sulfides, cobalt,
and nickel sulf ide. Has flocculating effect
on ore slimes.
Alkaline pH regulator
Alkaline pH regulator w/slime dispersing
action.
Acidic pH regulator.
NUMBER OF
MILLS
WHERE
USED*»

15
3
3
2
USUAL DOSAGE
kg/metric ton
ore feed

0.054 - 14.2
0.00015 - 0.025
0.54-12.12
0.018-4.3
ACTIVATORS
Universal activator for sphalerite. Also used
for the reactivation of minerals depressed by
cyanide.
Activator for sphalerite. Purchased as a
waste by-product from the manufacture
of electric circuit boards.
Activator for copper suifide minerals.
13
1
"'• -')''
0.06 - 2.32
0.13 f
0.0094
00
           *Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
           t Expressed as soluble copper metal.
          ** Reagent usage data supplied by 22 milling operations.

-------
                        TABLE VI-18.  SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
                                       (Continued)
REAGENT

Cyanide
Sodium Sulf ite
Zinc Sulfate
Sodium Dichromate
Dextrin
Starch
Guar
Hodag-31
Sulfur Dioxide
Noke's Reagent
Hydrogen Peroxide
DESCRIPTION
FUNCTION
NUMBER OF
MILLS
WHERE
USED»*
USUAL DOSAGE
kg/metric ton
ore feed
DEPRESSANTS
Sodium Cyanide, NaCN
or Calcium Cyanide,
Ca(CN)2
NaSO3
ZnSO> • H20 or
ZnSO4 • 7H2O
Na2 Cr2 O7
Corn starch
so2
Phosphorus Penta-
sulfide
P2S5
H2°2
Strong depressants for the iron sulf ides,
arsenopyrite, and sphalerite. Mild
depressant for chalcopyrite, enargite,
bornite and most other sulfide minerals
w/ exception of galena.
Depressant for pyrite and sphalerite
while floating lead and/or copper.
Depressant for sphalerite while floating
lead and/or copper minerals. Often
used in conjunction w/ cyanide.
Depressant for galena in copper-lead
separations. Excess depresses copper
sulfides and iron sulf ides.
Depressant for galena and molybdenite
while floating copper sulfides
Depressant for galena and activator for
copper sulfides. Often used in conjunc-
tion w/ starch.
Depressant for copper and lead while
floating molybdenite.
Depressant for copper sulfides in
copper-molybdenite separations.
13
2
7
1
4
2
4
1
0.003 - 0.065
0.2 - 7.46
0.1 - 1.35
0.022
0.0005 - 0.071
0.156-0.406
0.0001 - 0.47
0.016
00
00
            *Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
           ** Reagent usage data su pi lied by 22 milling operations.

-------
                          TABLE VI-18,  SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
                                         (Continued)
                   REAGENT
                              DESCRIPTION
                                 FUNCTION
                                       NUMBER OF
                                         MILLS
                                         WHERE
                                         USED**
USUAL DOSAGE
  kg/metric ton
    ore feed
                                                                       DEPRESSANTS
            Sodium Silicate
                           Na2O: nSi02
                      Depressant for quartz and other siliceous
                      gangue minerals. Also acts as slime
                      dispersant
                                                                                                                     0.031 - 2.08
            AERO Depressant
            610,633
                           Composition unknown —
                           Contains ~ 1.5%
                           phenolics
                      Depressant for graphitic and talcose
                      gangue. Also acts as gangue dispersants
                      useful in sand-slime separation.
                                                                                                                     0.001 - 0.16
            Jaguar Mud
                           Colloidal material
                      Depressant for gangue materials
                                                                                                                     0.016
oo
•o
Xanthates:
 AERO 301,325,343,355
 DowZ-3,Z-4,Z-6,Z-11,
     Z-14.
Sodium or potassium
salts of xanthic acid.
                                       R-O-C
                                   /
                                   \
                                       or
                                       R-O-C
                                   /

                                   \
           SNa



           S


           SK
                                       where R is an alky!
                                       group of 2-6 carbon
                                       atoms.
     COLLECTORS/PROMOTERS

Strong promoters for all sulfide minerals.
Essentially non-selective in the absence of
modifiers.
                                                                                             17
                                                                             0.0003 - 0.40
            *Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
           ** Reagent usage data supplied by 22 milling operations.

-------
             TABLE VI-18. SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
                           (Continued)
REAGENT
Dow Z-200
Fuel Oil
Vapor Oil
Tar Oil
Minerec
DRESSENATE TX-65W
SOAP

M.I.B.C. (Methyl
Isobutyl
Carbinol)
Methanol
Pine Oil
Cresylic Acid
DESCRIPTION
Isopropyl Ethyl-
Thionocarbamate
Saturated Hydrocarbons
Composition Unknown
Composition Unknown

Synonomous with
Methyl Amyl Alcohol
(CH3)2 CHCH2CHOHCH3
CH3OH
Composed primarily of
terpene hydrocarbons,
terpene ketones, and
terpene alcohols.
Higher homologs of
phenol, CeH5 • OH,
particularly cresols,
CH3 • C6H4 • OH,
and xylenols,
C2H5 • C6H4 • OH,
or
(CH3)2 • C6H4 • OH
FUNCTION
Promoter for copper sulf ides and activated
sphalerite w/ selectivity over iron sulf ides.
Promoters, usually used for readily float-
able minerals, such as molybdenite.
Promoter.
Promoter.
NUMBER OF
MILLS
WHERE
USED**
3
4
1
1
USUAL DOSAGE
kg/metric ton
ore feed
0.004 • 0.10
0.0013 - 0.78
0.01
0.41
FROTHERS
Alcohol type f rothers are used for the
flotation of sulfide minerals where a
selective, fine textured froth is desired.
Frother
Frother, widely used in sulfide flotation.
It exhibits some collecting properties,
especially for such readily floatable
minerals as talc, graphite and molybdenite.
Pine oil produces a tough, persistent froth
and has a tendency to float gangue.
A powerful fr other exhibiting some
collecting properties. Produces froth of
variable texture and persistence, and
tends to be non-selective.
10
1
5
3
0.008-0.17
0.00005
0.015-0.175
0.003 - 0.034
 * Copper, lead, zinc, silver, and molybedenum concentrators which discharge process wastewater (data available).
** Reagent usage data supplied by 22 milling operations.

-------
                       TABLE VI-18 SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
                                     (Continued)

REAGENT

Aliphatic
Dithiophosphates:
Sodium AEROFLOAT
AEROFLOAT 21 1,249,
3477



AEROFLOAT 31 and 242







"Reco"








ARMAC "C"


DESCRIPTION

R-0 S
>^ /
v/^V
R - O ^S Na

where R is an alkyl
group of 2-6 carbon
atoms.
Aryl Dithiophosphoric
Acids

R-0 £
Xp/
R-0' ^SH
where R is an aryl group
(benzene-based).
Sodium Dicresyl
Dithiophosphate

R - O S
V
R - O S Na
where R is the cresy!
group:
CH, - CeH, - OH
O DO
Acetate Salt of
Aliphatic Amines

FUNCTION

Promoters of variable selectivity, and
strength for the flotation of sulf ide materials
Sometimes used in conjunction with
xanthates for improved precious metal
recoveries.



Promoters for copper, lead, zinc and
silver sulf ide minerals. Has frothing
properties.

\



Promoter, selective to copper sulfide
minerals. Very similar to AEROFLOAT
31 and 242.



          'Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
         **Reagent usage data supplied by 22 milling operations.

-------
               TABLE VI-1!8  SUMMARY OF REAGENT USE (BY FUNCTION) IN ORE FLOTATION MILLS*
                             (Continued)
REAGENT
Polyglycols:
DOWFROTH 200, 250
AEROFROTH 65
Diphenyl Guanidine
UCON-R-23
UCON-R-133
SYNTEX

AEROFLOC
AERODRI 100
VALCO 1801
NALCOLYTE 670
SEPARAN NP-10
SUPERF LOC 3302
Flocculants (unspecified)
DESCRIPTION
Polyglycol Methyl
Ethers (i.e. Poly-
propylene glycol methyl
ether)
CH3 • (OC3H6)X • OH
HN-C (NHC6H5)2
Composition unknown
Composition unknown
Composition unknown
FUNCTION
Frothers, for metallic flotation, w/ froth
persistancy and selectivity against non-
metals.
Frother
Frother
Frother
Frother
NUMBER OF
MILLS
WHERE
USED»»
8
1
1
1
1
USUAL DOSAGE
kg/metric ton
ore feed
0.002-0.17
0.00005
0.035
0.015
0.017
FLOCCULANTS
Anionic, Cationic, or
Nonionic Organic
Polymers.
Used as dewatering aids or filtration
aids for thickening or filtering ore
pulps, concentrates, and tailings.
9
0.00015-0.051
 •Copper, lead, zinc, silver, and molybdenum concentrators which discharge process wastewater (data available).
** Reagent usage data supplied by 22 milling operations.

-------
                           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>(l) of the CWA.

EPA conducted sampling  and  analysis  at  facilities  where  BPT
technologies  are  in  place;  therefore,  any  of  the  priority
pollutants present in treated effluent discharges are subject  to
regulation   by   BAT   effluent   limitations  guidelines.   The
Settlement  Agreement  in  Natural  Resources  Defense   Council,
Incorporated, v^ Train, 8 ERC 2120 (D.D.C. 1976), modified 12 ERC
1833  (D.D.C.  1979)  provides  a  number  of  provisions for the
exclusion of particular pollutants, categories and subcategories.
The criteria for exclusion of pollutants are summarized below:

     1.  Equal or more stringent protection is  already  provided
     by an effluent limitation and guideline promulgated pursuant
     to  Section(s) 301, 304, 306, 307(a), or 307(c) of the CWA.

     2.   The  pollutant  is  present  in  the effluent discharge
     solely as a result of its presence in the intake water taken
     from the same body of water into which it is discharged.

     3.  The pollutant is not detectable in the  effluent  within
     the  category  by  approved  analytical  methods  or methods
     representing the state-of-the-art capabilities.

     4.  The pollutant is detected in  only  a  small  number  of
     sources  within the category and is uniquely related to only
     those sources.
     5.  The pollutant is present in only trace  amounts
     neither causing nor likely to cause toxic effects.
and  is
     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.
                                193

-------
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-six (86) priority organic pollutants were
not found in the  treated  effluents  during  sampling.   Of  the
remaining 28 organic pollutants, 17 were found in the effluent of
only  one  or  two  sources  and  always at or below 10 ug/1, ten
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, 28
of the 114 toxic organics were detected during sampling, while 86
toxic organics were not detected  in  treated  wastewater  during
sampling.   Therefore,  the  86  toxic  organics not detected are
excluded by Criterion 3  (the  pollutant  is  not  detectable  by
approved analytical methods).
                                194

-------
Of   the   28  toxic  organics detected,  17  were detected  at  at  least
one  facility and always  at or  below  10 ug/1, which  is   the   limit
of   detection set  by the Agency  for the  toxic organics  in  these
sampling  and analysis programs.
      1 .  Chlorobenzene
      2.  1 ,1,1-Trichloroethane
      3.  Dichlorobromomethane
      4.  Chloroform
      5.  Fluorene
      6.  Ethylbenzene
      7.  Trichlorofluoromethane
      8.  Diethyl Phthalate
   9.   Tetrachloroethylene
  10.   Toluene
  11.    -BHC (Alpha)
  12.    -BHC (Beta)
  13.    -BHC (Delta)
  14.   Aldrin
  15.   Dieldrin
  16.   Endrin
17.   Heptachlor
Thus, it follows that these  17 compounds are subject to exclusion
under Criterion 3.

Pollutants  Detected  But  Present   iri  Amounts  Too  Low ' to  be
Effectively Reduced by Known Technologies

Toxic Organic Pollutants

There  were  10  organic pollutants  detected during the nine 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-Dichloroethyiene
     (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
     (10) Pentachlorophenol        .

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.
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 VII-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
                                 196

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 field has shown that it is difficult to ensure a  tight  seal  at
 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.
       i
 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
                               W7

-------
of only 0.015 mg/1.   No specific treatment data or application of
specific treatment could be found in the ore mining  and  milling
industry.   Pilot-scale  treatment studies have reported removals
ranging from 10 to 84 percent  using  conventional  technologies,
but  only  cation and anion exchange used in combination achieved
high removal efficiencies (Reference 8).  The  removals  obtained
by utilizing conventional technology are not consistent enough to
base  regulations upon them, and cation and anion exchange, while
possibly providing additional treatment, are  judged  too  costly
for  this  industry.   Consequently,  selenium  is  excluded from
regulation since it is found at levels too low to be  effectively
reduced by known technologies (Criterion 6).

Silver.   Most  data  available  on  treatment  of  waste streams
containing silver represent attempts at recovery of this valuable
metal from the photographic and  electroplating  industries.   In
these  industries  there  has been an ample economic incentive to
develop recovery/removal technology because:  (1) the  silver  is
valuable  and  may  be  reused;  and (2) the concentration levels
present favor the economic recovery  of  silver  from  the  waste
stream.   Four  basic  methods for silver removal from wastewater
are  discussed  in  Reference  8:   (1)  precipitation,  (2)  ion
exchange,  (3) reductive exchange, and  (4) electrolytic recovery.
Levels to 0.1 mg/1 have been reported  by  various  investigators
but most of these in bench- or pilot-scale systems.  In addition,
waste  streams in this industry are high in solids, effluent flow
rates are very high, and treated effluent levels are already  low
(mean  of treated effluent samples 0.015 mg/1; maximum 0.04).  It
has been concluded that the concentration levels present are  too
low  to  be  effectively reduced by known technologies (Criterion
6).

Beryllium.  There are little data available in the  literature for
beryllium removal in wastewater from the ore mining and  dressing
industry.   Only  one  domestic  facility mines beryllium ore and
uses water in a beneficiation  process   (Mine/Mill  9902).   This
mill  uses a proprietary leach process with raw wastewater having
beryllium at a concentration of 36 mg/1 at pH 2.6  (Reference  9).
This  facility  has  no  discharge.   However,  when  TSS   in the
impoundment is reduced from  116,000 to  44,000 mg/1  (after a short
settling  period),  beryllium  is  reduced  to  25  mg/1.   Since
beryllium  is relatively insoluble, it  is believed  that reduction
to an effluent level of TSS of 20 mg/1  after  lime  precipitation
and settling would result in a substantial reduction of beryllium
levels.

In  a related  industry, primary beryllium refining, some data are
available which  indicate effluent beryllium levels  of  0.09  mg/1
are  possible  by   lime  precipitation  and multiple pond settling
(Reference  10).  Of  73  effluent  samples  analyzed  during  BAT
screening for  beryllium, only  10 samples had detectable beryllium
concentrations   with   a   mean   of    0.005  mg/1 and  maximum
concentration  of  0.011 mg/1.   These  are  the  levels  which  are
                                 198

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 achieved by known technologies and since the pollutant is present
 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 (FeCr2O«), 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   i.n  Treated  Effluents at  a Small Number  of
 Facilities  and  Uniquely Related to Those Facilities

 The  toxic organic pollutant,  2,4-dimethylphenol,  was  detected  in
 the   effluent  at  only  one  facility  (9202)   during the screen
 sampling program.   AEROFLOATT.,  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
 V A A ™* j •
                                199

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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.

El

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.

     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

Settlea.ble 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.
                                 200

-------
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 VI1-2
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  concentrations)  of  the  regulated  param-
eter (s) are used to predict the concentration of the toxic pollu-
tant.    When  the  concentrations) 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.
                                201

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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 .
     2.
     3.
     4.
     5.
     6.
     7.
chromium
copper
lead
nickel
selenium
zinc
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.
                               202

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Additonal Paragraph £ 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, the only existing mill discharging.
                                203

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   Table VII-1

  DATA SUMMARY
 ORE MINING DATA
ALL SUBCATEGORIES


NUMBER OF
SAMPLES
ACENAPHTHENE
ACROLEIN
ACRYLONITRILE
BENZENE
BENZIDENE
CARBON TETRACHLORIOE
CHLOROBENZENE
1.2. 3-TRICHLOROBENZENE
HEXACHLOROBENZENE
,2-DICHLOROETHANE
. 1 , 1-TRICHLOROETHANE
HEXACHLOROETHANE
,1 -OICHLOROETHANE
^ . 1 , 2-TRICHLOROETHANE
2 .1.2,2 -TETRACHLOROETHAN
CHLOROETHANE
BIS(CHLOROMETHYL) ETHER
BIS(2-CHLOROETHYL) ETHER
2-CHLOROETHYL VINYL ETHE
2 -CHLORONAPHTHALENE
2.4. 6-TRICHLOROPHENOL
PARACHLOROMETA CRESOL
CHLOROFORM
2-CHLOROPHENQL
1 ,2-OICHLOROBENZENE
1 , 3-DICHLOROBENZENE
1 . 4-DICHLOROBENZENE
3 , 3-DICHLORQBENZIDINE
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
32
32
32
32
32
32
32
32
32
32
RAW(UQ/L)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
0
O
0
10 4.8922 4 10 10
0
1 1111
0
0
O
0
9 6.7208 8.5811 10 10
O
O
0
O
O
O
0
O
0
1 11.667 11.667 11.667 11.667
0
8 7.6096 3.1623 12.5 35
0
0
0
O
0
*
*
*
*
*
*
*
*
*
*
*
*
*
*






*
*
*
*









NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
23
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
TREATED (UQ/L)
NUMBER DETECTED
DETECTED MEAN MED
0
O
0
3 8.3333 7
0
0
1 0.005 O.005
0
0
0
5 7.2649 8.5811
0
0
0
0
0
0
0
0
0
0
0
8 5.1281 3.1623
0
0
0
0
0

VALUES ONLY
. 90% MAX



10.7 11


0.005 O.O05



1O 1O











10 1O






-------
Table VII-1 (Continued)

     DATA SUMMARY
    ORE MINING DATA
   ALL SUBCATEGORIES


NUMBER OF
••.••*• SAMPLES
1. 1-DICHLOROETHYLENE
1 . 2-TRANS-DICHLOROETHYLE
2.4-DICHLOROPHENOL
1 , 2-OICHLOROPROPANE
1 . 3-DICHLOROPROPENE
2,4-DIMETHYLPHENOL
2 , 4-DINITROTQLUENE
2 , 6-DINITROTOLUENE
1 , 2-DIPHENYLHYDRAZINE
ETHYLBENZENE
*LUORANTHENE
METHYL CHLORIDE
METHYL BROMIDE
BROMOFORM
DICHLOROBROMOMETHANE
TRICHLOROFLUGROMETHANE
DICHLORODIFLUOROMETHANE
CHLORODI BROMOMETHANE
HEXACHLOROBUTADIENE
HEXACHLOROCYCLOPENTADIEN
ISOPHORONE
NAPHTHALENE
NITROBENZENE
2-NITROPHENOL
4-NITROPHENOL
2.4-BINITROPHENCL
4.6-DINITRO-O-CRESOL
32
32
32
32
32
32
32
32
32
32
32
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
RAW(UGXL)
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 80% MAX
2 6.5811 3.1623 8.6325 10
0
1 10 10 10 10
O
0
1 140 140 140 140
0
0
0
4 6.7167 1 13.48 17.667
0
1 45 45 45 45
O
0
0
S 5.0325 2.0811 10 1O
0
0
0
0
0
1 12.5 12.5 12.5 12.5
0
0
0
0
0
*
- *
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*






*
*
*

NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28

NUMBER
DETECTED
0
1
0
0
0
1
0
0
0
3
0
0
0
0
2
3
0
0
o
0
0
0
0
0
0
0
o
TREATED (UQ/L)

-------
                                           Table VII-1 (Continued)

                                                DATA SUMMARY
                                               ORE MINING DATA
                                              ALL SUBCATEGORIES
f>0
RAM(UQ/L)
NUMBER OF
SAMPLES
N-NITROSODIMETHYLAMINE
N-NITROSOD1PHENYLAMINE
N-NITROSODI -N-PROPYLAMIN
PENTACHLOROPHENOL
PHENOL
BIS(2-ETHYLHEXYL) PHTHAL
BUTYL BENZYL PHTHALATE
OI-N-BUTYL PHTHALATE
OI-N-OCTYL PHTHALATE
OIETHYL PHTHALATE
DIMETHYL PHTHALATE
BENZO ( A } ANTHRACENE
BENZO(A)PYRENE
BENZO ( B ) FLUORANTHENE
BENZO ( K ) F LUOR ANTHENE
CHRYSENE
ACENAPHTHYLENE
ANTHRACENE
BENZO(G,H,I)PERYLENE
FLUORENE
PHENANTHRENE
DIBENZO( A. H) ANTHRACENE
INDENO(1,2,3-C.D)PYRENE
PYRENE
TETRACHLOROETHYLENE
TOLUENE
TRICHLOROETHYLENE
33
33
33
33
33
33
33
33
10
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
NUMBER DETECTED
DETECTED MEAN MED
0
0
0
1
2
15
2
13
3
16
0
O
0
0
0
0
0
0
0
1
0
0
0
0
2
9
0


10
118
20.18
10.75
16.489
10
24.414








10




7. 75
399.28



10
76
13
0.5
10
10
1O








10




4.5
2.0811

VALUES ONLY
90% MAX


10
143.2
39.833
16.9
26.1
10
59.4








10




9.7
388.3



10
160
100
21
56
10
90








10




11
3560

*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
* -
*
*
*
*
*
*
*
if
*
*
*

NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
7
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28

TREATED
(UQ/L)
NUMBER DETECTED
DETECTED MEAN MED
0
O
0
0
3
18
4
12
3
4
3
0
0
0
0
0
0
0
0
1
0
0
0
0
1
0
0


92.3
12.458
27.791
25.864
12.167
7.875
12.2








10




1.1
2.5967



33.45
10
10
10
10
9.6
5.8








10




1.1
1



VALUES ONLY
90% MAX


166.8
26
52.4
39.2
14.55
10
20.35








10




1.1
5.26



210
60
66
140
16.5
10
25







4/\
10




1.1
10


-------
Table VII-1 (Continued)

     DATA SUMMARY
    ORE MINING DATA
   ALL SUBCATEGORIES
RAW(UQ/L)
NUMBER OF
SAMPLES
VINYL CHLORIDE
ALDRIN
DIELDRIN
CHLORDANE
4.4rDDT
4. 4 -DDE
4.4-DOD
ENDOSULFAN-ALPHA
ENDOSULFAN-BETA
ENDOSULFAN SULFATE
ENDRIN
ENDRIN ALDEHYDE
HEPTACHLOR
HEPTACHLOR EPOXIDE
BHC- ALPHA
BHC-BETA
BHC (LINDANE) -GAMMA
BHC-DELTA
PCB-1242 (AROCHLOR
PCS -1254 (AROCHLOR
PCB-1221 (AROCHLOR
PCB-1232 (AROCHLOR
PCB-1248 (AROCHLOR
PCS -1260 (AROCHLOR
PCB-1016 (AROCHLOR
TOXAPHENE

















1242)
1254)
1221)
1232)
1248)
1260)
1016)

2.3,7. 8 -TETRACHLORODIBEN
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
33
9
9
9
9
9
32
33
NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
O
4
0
0
0
1
1
1
0
0
0
0
1
0
5
5
4
2
0
0
0
0
0
0
0
0
0

6.4156



S
6.6667
10




7.5
5.2649
6.1325
6.2072
5










5 9



5 5
6.6667 6.6SS7
10 10




7.5 7.5
4.0811 7.5
5 8.75
5 8.6667
5 5










10



S
6.6667
10




7.5
10
10
10
5









*
*
*
*
*
*
*
*
*
£
*
*
*
*
*
*
*
*
*
*
*
*
*
*
*
* •
*
*
*
*

NUMBER OF
SAMPLES
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
28
6
6
6
6
a
27
28
TREATED (UG/L)

NUMBER DETECTED VALUES ONLY
DETECTED MEAN MED 90% MAX
0
2 6.5811 3.1623 8.6325
2 B.5811 3.1623 8.6325
O
0
0
0
0
o
0
1 555
0
2 6.5811 3.1623 8.6325
0
3 555
1 5 5 5
0
2 5 5 5
o
0
0
0
0
0
0
0
0

1O
10







5

10
5
g

g










-------
                                         Table VII-1 (Continued)

                                              DATA SUMMARY
                                             ORE MINING DATA
                                            ALL SUBCATEGORIES
RAW(UG/L)

CIS
TRA*
NUMBER OF NUMBER
SAMPLES DETECTED
1 -3-DICHLOROPROPYLEN
1 1.3-DICHLOROPROPYLE
DETECTED VALUES ONLY
MEAN MED 90% MAX


*
*
*
*
*
TREATED (UQ/L)
NUMBER OF NUMBER DETECTED
SAMPLES DETECTED MEAN MED


VALUES ONLY
90% MAX

ro p
oo

-------
                                                   Table VII-1  (Continued)
ro
                                                      DATA i-UMKARY
                                                      OFF.  MIN1NC UATA
                                                     ALL isUyCA
RAU([tr,/L) •

ANTIPCt-.Y (KFAL)
ARiLNIC (TuTAL)
SERYLLIur- (TOTAL)
CADMIUM (TOTAL)
CHROMIUM (TO TAX)
COPPEK (TOTAL)
CYANIC-!:' (TCHAD
LEAD 'FOr»L>
HcRCUr.Y (TOTAL)
MICHEL (TOTAL)
3ELI.N1UH (TOTAL)
S1LVFP (T3TAL)
THALLIUM (TOTAL)
ZINC IK'TAL)
COD
TSo
TOC
PH (UMTS)
PHENOL- ICr- (1/iP)
IRON (TOT 41. )
NUMBER OF
SAMPLES
82
114
P*
1T6
85
103
t>F>
86
P7
86
F.I
&<"
P 2
106
£2
17
6
111
72
?2
NUMBER
or fEcrf.u
6
106
1.\
•51
70
100
ri
70
54
70
•58
?.3
J
.116
7.2
17
: fc
'•5
71
IV
Oi.
H-AX
.C4S.J3
2.2219
. 1 3 7 3 ci
0*21*
J.HI 62
6 •> . t ^ 1
,.3 21' 4 2
J4.1.807
. TD'Jqi-J
3.SU6
lt..
2040 *
MUMPER OF NUMPE"
SAMPLES DETECTED
71
ino
7j
92
75
90
l>7
75
80
ff,
73
73
71
92
23
47
9
?6
r-7
25
3
83
10
36
2?i
£3
14
31
37
13
37
q
3
82
23
46
P
36
49
?1
TREATED
(MG/L)
DETFCTED
MEAN MEDIAE
0.031
.15349
0.0051
.01415
.13623
.23464
.13571
.1348)
.01264
.22202
=05957
.015f 7
.52767
.90209
11.69?
?4.9fl9
5.1B75
6.9714
.07378
.62619
lltOE-5
(1.018
0.005
0.005
0.035
0.06
0.08
0.05
ft OOF -6
0.07
0.015
0.01**
0.74
0.062
11
9.5
5.25
7.7
0.032
O.P09-


VALUES ONLY
90* . MAX
0.1
0.6
0.0109
0.06
0.332
0.548
0.435
0.376
0.0306
0.966
0.112
0.04
O.P4
2.244
20.6
69.2
6
8.16
-0.21
2.192
0.1
1.5
0.011
0.077
1.8
4.6
0.6
0.959
0.25
1.28
0.9
0.04
0.84
11.1
53
157
. 6
8.5
0.46
3.87

-------
                       TABLE VII-2.  POLLUTANTS CONSIDERED FOR REGULATION
SubcBtffQorv
Iron Ore
Copper, L«d,
Zinc, Gold,
Silver, Platinum,
Molybdenum
Aluminum
Tungsten
Mercury
Uranium
Antimony
Titanium
Nickel
Vanadium
Sub-
Division
Mines
Mills
Mines
Mills
Mines
Mines
Mills
Mines
Mills
Mines
Mills, In
Situ Leach
Mines
Mills
Mines
Mills
Mills w/
Dredges
Mines
Mills
Mines
Mills
Mill Process

Phys/Chem
Phys (Meiabi)

Cyanidation or Amalgamation
Heap, Vat, Dump, In Situ Leach
Froth Flotation
Gravity Separation
















Toxic Pollutants
Sb
E
E
As
E
E
•
|
U
G
G
R«
E
E
Crl
E
E
Or
E
E
On
E
E
CN
E
E
Ph
E
E
Ha
E
E
Ni
E
E
S«
E
E
Afl
E
E
Tl
E
E
Zn
E
E
(Total Phenolic*
E
E
Convtn-
tionili
TSS
G
G
pH
G
G
Nonconventfonals
COD


1 F» - Total or Din.
G
G
Ra
226


At


U


i
&
1
1


Zero Discharge at BPT
E
E
G
E
G
E
G | G
G
G
G
E
E
E
G
E
G
G






Zero Discharge at BPT
Zero Discharge at BPT
E

E
E
E

G

E
G
G

G

G
G
G

E

E
E
E

G

E
G
G

E

E
E
E

G

E
G
G

G

E.
E
E



E
E
G



E
E
E
G


E
E
E
G


E
E
E



E
E
E



E
E
E

G

E
G
G

G

E
E
E

G

G
G
G
G
G
G
G
G
G
G








G











G










G




Zero Discharge at BPT
E
E
E
E
G
G
E
E
E
E
E
E
E
E
E
E
E
E
E
E
E
G
E
E
E
E
E
E
G
G
E
E
G
G
G
G
G
G


G
G


G



Reserved

E
E
E
E
E
E
G
G
G
E
E
E
E
E
E
E
E
E
E
E

E
E

E
E

E
E

E
G

E
E
E

E
E
E

E
E
E

E
U
b

E
E
E
G
G
G
G
G
G



G

~5~















Reserved
Reserved
E * Excluded from Guideline Development
G- Guidelines to be Considered

-------
TABLE VI1-3.  PRIORITY METALS EXCLUSION CRITERIA BY SUBCATEGORY, SUBDIVISION,
                MILL PROCESS

Subcategory
Iron Ore


Copper, Lead,
Zinc, Gold,
Silver, Platinum,
Molybdenum


Aluminum
Tungsten

Mercury

Uranium

Antimony

Titanium


Nickel

Vanadium


Sub-
Division
Mines
Mills

Minos
Mills



Mines
Mines
Mills
Mines
Mills
Mines
Mills, In
Situ Leach
Mines
Mills
Mines
Mills
Mills w/
Dredges
Mines
Mills
Mines
Mills

Mill Process

Phys/Chem
Phys (Mesabi)

Cyanidation of Amalgamation
Heap. Vat, Dump, In Situ Leach.
Froth Flotation
Gravity Separation

















Sb
3
3

3


6

3
3
6


3
3


3
6
3





As
6
6
'
6




3




6



3
3
3





Chryiotllt

























Be
3
3

6


6

3
3
6


3
3


3
3
3





Cd
3
3






3




3
3


3
3
6





Cr
3
6
Ze
3


6

6




6
6


3
3
3




T
Cu
6
6
roDc

Zero!
Zero


6



Zero
3
6


6
G
6




owe
CN
3
3
Khar

3 ben.
Disci


3
3
3

Disci
3
3
Rrai

3
3
3
Resei

Rt-»

Pofci
Pb
3
3
pat

argea
targe


3
3


"arge
6
6
wrri

6
6
6
ived

•vtrf

tanti
Hi
3
3
BPT

itBPl
atBP


6
6
3

atBP
6
6


3
3
6





Hi
3
3


r
T


3
3
3

T
3



3

3





Se
3
3

3


6

3
-3
3


6
6


3
3
3





AS
3
3

e


6

3
3
6


3
6


3
3
6





Tl
3
3

e


3

3
3
3


3
3


3
3
3





Zn
6
6






3








6

6





Toul Phtneliei
1 	 — i
3
3

6


6

6
3
6 '


6
6


6
6
6




                                                                                           EXCLUSION CRITERIA
                                                                                          1. Equal or i
                                                                                            provided by EPA's
                                                                                          2. The pollutant is present
                                                                                            as a molt of its yn. yt mi
                                                                                            in the intake vmer.

                                                                                          3. The pollutant is not
                                                                                          4. The pollutant rs unique to
                                                                                                     roff s
                                                                                          5. The polliKam is promt in
                                                                                          6. The pollutant is present in
                                                                                            amounts too small to treat.

                                                                                          7. The pollutant a effectively
                                                                                            oonttoUedl oy tiuatMy
                                                                                            other polmants.

-------
                           Table VII-4

                 TUBING LEACHING ANAYSIS RESULTS
Component

Bis (2-ethylhexyl) Phthalate
     Micrograms/Liter

 Origina.1            Tygon
     Acid Extract
     Base-Neutral Extract
Phenol
     Acid Extract
     Base-Neutral Extract
   915
 2,070
19,650
  N.D.
N.D.
885
N.D.
N.D.
N.D. - Not Detected
                                 212

-------
                          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
flotation as a depressant and in cyanidation for leaching.
froth
                                  213

-------
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
                                  214

-------
increase in dosage that can overshoot the minimum amount required
to produce the most efficient separation.

The degree of sophistication of  in-process  control  of  cyanide
varies  widely in the category.  The greatest degree of sophisti-
cation is used at copper/lead/zinc Mill 3103.  A  Courier  online
X-ray  analyzer  performs  analyses at 10-minute intervals of the
mill heads and tails and of the concentrates, heads, and tails of
the individual flotation circuits.  Analytical results  are  com-
puted  by  a Honeywell 316 computer and automatically printed and
charted.  The mill operator may then adjust the rate  of  reagent
addition  based  on  these  analytical results.  For example, the
rate of cyanide addition is decreased when the copper content  of
the  copper  circuit  tailings  increases during a time increment
(usually two hours).  Conversely, the rate of cyanide addition is
increased when the iron content of the lead and zinc concentrates
increases.  In this manner, the mill operator is able to optimize
the reagent use, percentage recovery, and  grade  of  concentrate
produced.  Several mills (3103, 3105, 3122, 3123, 2117, and 2121)
have on-stream analytical capabilities.

Laboratory  analysis  provides adequate control in the milling of
simple, "clean" ores.  However, the greater  the  complexity  and
variability  of  the  ore  being milled, the more advantageous it
becomes  to  a  mill  operator  to  have   on-stream   analytical
capabilities.

The   prevailing  practice  for  in-process  control  of  reagent
addition consists of manual sampling and laboratory  analysis  of
heads, tails, and concentrates.  Typically, samples are collected
at   two-hour  intervals  and  analyses  are  begun  immediately.
Approximately two hours are required  before  analytical  results
are  available.  This method is slower and produces less informa-
tion than the more sophisticated method previously described.

Many small mills have limited  analytical  capabilities  and  the
control of reagent addition depends on the experience of the mill
operator.   According  to  mill  operators  and  site visit data,
cyanide addition in excess of the amount  required  is  generally
used with limited analytical control.

Control  of the rate of reagent addition depends on the attention
given to the analytical results by the mill operator.  The  atti-
tude, conscientiousness, and experience of the mill operator have
a  significant  effect  on  the degree of control maintained over
reagent usage.  The efficiency of reagent usage impacts the over-
all efficiency and economy of the mill, as well as the  character
of  the  wastewater generated, and operators must remain aware of
this.
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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 to Cyanide Iri Flotation

Cyanide  is believed to function primarily as a reducing agent in
the depression of pyrite in xanthate  flotation  operations.   In
1970,  Miller  (Reference  5)  investigated  alternative reducing
agents and found that, in terms of effectiveness and cost, sodium
sulfite  compared  quite  favorably  with  cyanide  as  a  pyrite
depressant.   In -particular,  it  was found that cyanide exerted
some depressant effect on chalcopyrite  and  pyrite,  but  sodium
sulfite  did  not.  The sodium sulfite alternative appeared to be
applicable to copper ore flotation operations.

Some mills use sulfite or sulfides instead of cyanide.  Mill 3101
is an example of a copper/zinc flotation mill which  uses  sodium
sulfite  and  no  cyanide.   There  was  no  measurable effect on
recovery or grade of  concentrate.   At  Mill  6104,  copper  and
molybdenum  minerals are separated in froth flotation with sodium
bisulfide  used  as  a   copper   depressant,   provide   another
alternative  to  the  use of cyanide for the depression of copper
minerals in selective flotation.

An EPA-sponsored study to identify and evaluate  alternatives  to
sodium cyanide was initiated in May 1978 (References 6 and 7) and
alternatives  were  identified  by  a  literature search.  Points
taken  into  consideration  were:   ability  to  depress  pyrite,
selectivity  of  depressant,  theory of performance, inferred and
specific environmental aspects, state of development as a practi-
cal depressant, and cost.  Fourteen alternatives were identified,
three of which were carried into  the  evaluation  phase  of  the
study.   The  compounds selected for bench-scale evaluation were:
sodium monosulfide (Na2S), sodium sulfite  (Na2SO3_),  and  sodium
thiosulfate  (Na2S2O3_).  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
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the  cyanide  at  the  natural pH level, natural meaning the pre-
vailing pH of the ground ore plus water.  At elevated pH  (1 0  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 effectiverat 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):
                                 217

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 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  are methyl isobutyl  carbinol (MIBC)  and
polyglycol  methyl  ethers.   Frothers are generally nonselective,

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generically related compounds, with  fairly  predictcible  charac-
teristics.   As  such, substitution within this class of reagents
(frothers) should not be difficult.  For example, Mill  2121  has
recently  discontinued use of cresylic acid in a silver flotation
circuit by substitution with polyglycol methyl ethers.

Collectors are much  more  selective  than  frothers,  and  their
effectiveness  is  highly dependent upon their compatibility with
associated modifiers,  promoters,  activators,  and  depressants.
Possible  alternatives to phenolic collectors are dithiophosphate
salts and dithiophosphoric acids with alkyl groups  in  place  of
phenol  groups.   Substitution  of these reagents for phenol con-
taining collectors may be feasible without serious  complications
or  economic consequences; however, the consequences of substitu-
tion are site dependent and require extensive experimentation  at
each mill.

In-Process Recycle of Waste Streams

In-process  recycle  of  concentrate  thickener  overflow  and/or
recycle of filtrate produced by concentrate  filtering  is  prac-
ticed  at  a  number  of flotation mills (e.g., 2121, 3101, 3102,
3108, 3115, 3116, 3119, 3123, and 3140).   In  addition,  several
mills (2120, 6101, and 6157) use thickeners to reclaim water from
tailings prior to the final discharge of these tailings to ponds.
Water  reclaimed  in  this  manner is used as makeup water in the
mill.  In-process recycle of waste streams  produced  by  concen-
trate  dewatering  is incorporated primarily as a process control
                                                \4- V-* ^**I

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Recycle  of  spilled  reagent  can also be  an advantage.  At Mill
3101, the occurrence of spills and overflow from flotation  cells
results  from  the milling of a higher grade ore than the reagent
dosage is optimized  for.   A  system  has   been   implemented  to
collect  spills  and  return them to the  flotation circuit.  This
control practice not only improves the quality of  treated  waste-
water, but the percentages of metals recovered as well.

Use of Mine Water as Makeup in the Mill

A  large  number  of  mine/mill  operations use mine drainage as
makeup in the mill (e.g., 4103, 4104,  4105,  3101,  3102,  3103,
3104, 3105, 3106, 3108, 3110, 3113, 3118, 3119, 3122, 3123, 3126,
3127,   3138,  3142,  6102,  6104,  9402,   and  9445).   In  some
instances, the entire process water requirement of  the  mill  is
obtained from mine drainage.

From  a  wastewater  treatment  aspect  for facilities allowed to
discharge, a great advantage is gained by this practice.   First,
this  practice either eliminates the requirement for a mine water
treatment system or greatly  reduces  the   volume  of  wastewater
discharged to a single system.  As discussed previously, reducing
the volume of wastewater flow to an existing treatment system can
be  an effective means of enhancing the capabilities of that sys-
tem.  Second, in situations where mine water contains  relatively
high  concentrations  of soluble metals,  its use in the mill pro-
vides a more effective means for the removal of these metals than
could generally be attained by treatment of the mine water alone.
This is due to reduced metals solubility  in the  alkaline  condi-
tions maintained in flotation and most mill  circuits.  Therefore,
use of mine water as makeup in a mill can be considered a control
practice  which  improves  the  quality  of mine and mill treated
wastewater.

Techniques for Reduction of_ Wastewater Volume

Pollutant discharges from mining and milling sites may be reduced
by limiting the total volume of discharge,  as well as by reducing
pollutant concentrations in the  wastestream.   Volumes  of  mine
discharges are not, in general, amenable to control, except inso-
far as the mine water may be used as input  to the milling process
in  place  of  water from other sources.  Techniques for reducing
discharges of mill wastewater include limiting water use, exclud-
ing incidental water from the waste stream,  recycle  of  process
water,  and impoundment with water lost to  evaporation or trapped
in the interstitial voids in the tailings.

In most of the industry, water  use  should be  reduced  to  the
extent practical, because of the existing incentives for doing so
(i.e.,  the  high  costs  of  pumping  the   high volumes of water
required, limited water  availability,  and the  cost  of  water
treatment  facilities).  Incidental water enters the waste stream
directly through precipitation and through  the  resulting  runoff
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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.  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.  However, this most often entails  only
the  removal  of  solids in a thickener or tailing pond.  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.    In  general,  plants
practicing  bulk flotation on sulfide ores  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 successful 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

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 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.

 Commenters  came forward following proposal of  the  standards   for
 new  mills  with  data demonstrating  that the  buildup of reagents
 and other contaminants can in fact  interfere with the extractive
 process,    causing  severe  loss of   product.    They have  also
 demonstrated  that treatment of the  recycle water  may   not  always
 be  an   economically  viable  option  for dealing  with this inter-
 ference  problem.   Unfortunately,  this interference is a complex
 phenomenon, which appears  to be related to the characteristics of
 the  ore at particular sites,  making  it impossible to carve out a
 subcategory  of   facilities   afflicted   with    this  problem.
 Accordingly,   £o accommodate the problem,  the  final NSPS contains
 a   special   "bleed"  or  "purge"  provision which   will   allow
 facilities  to  discharge   wastewater  subject  to the  NSPS mine
 drainage standards if  they  can  demonstrate   to   the permitting
 authority   that total  recycle would cause a major interference in
 the  extractive  metallurgical   process  and   that    appropriate
 treatment   of   the  recycle  water  is not adequate to remedy this
 interference.   This  provision  will   allow such  facilities   to
 substitute  some  fresh water for recycle water and thereby avoid
 the losses  associated  with buildup  of contaminants in the recycle
 water.   Specifications of  the exact amounts of  water discharged
 and  the approrpiate  treatment of  recycle water  will, of course,
 be  left  to  the permitting  authority.   This is  discussed  in  more
 detail in Section XII  of this document.

 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

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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
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.
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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.

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.
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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
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 cibsorbing fine
metal   hydroxide   precipitates.    The  major  disadvantage  of
                              225

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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
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.
                                226

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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
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
    Effluent Levels (mg/1)
     Attained Prior to
    Use of Polymer	
TSS
Pb
Zn
   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
Effluent Levels (mg/1)
 Attained Subsequent
to Use of Polymer
              Range
             4 to 34
          0.14 to 0.67
          0.06 to 0.69
Similarly, the use of a polymer  at  Mine  3130  reduced  treated
effluent concentrations of total suspended solids, lead, and zinc
over  concentrations  attained  prior  to  use of the polymer, as
shown in the following (company supplied data):
               Effluent Levels  (mg/1)    Effluent Levels  (mg/1)
            Attained Prior to Use     Attained Subsequent to
           of Polymer and Secondary    Use of Polymer and
Parameter      Settling Pond*        Secondary Settling Pond*
TSS
Pb
Zn
Mean

  19
0.34
0.45
    Range     Mean

   4 to 67       2
0.11 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.

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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.

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.

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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).

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
                                229

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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.

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

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 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.

 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

Cd
Cr
Cu
Ni
Pb
Zn
 Concentration (mq/1-total)
 Initial             Final
0.015 to 0.034       less than 0.005
0.01 to 0.125        less than 0.01
0.04 to 0.15         less than 0.01 to 0.02
",045 to 0.16        less than 0.01 to 0.04
0.32 to 1.32
0.382 to 1.49
less than 0.02
0.02 to 0.417
                                 231

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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.

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 1/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
                                 232

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 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
 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-z  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),.-2
 and/or Zn(CN?), and iron-cyanides (Fe(CN)6-z  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.
                                 233 .

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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:

     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:
C1
CN-
H0 •«• 2NaCl
                  2NaOH --- > CNO~
                      and
    3C12  +  2CuCN + SNaOH --- > 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
                                 234

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 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
 metal-   cyanide  complexes   as hydroxides.   Reactions  of the free
 cyanide and  the metal-cyanide complex with hypochlorite are:
    NaOCl + CN-    	         CND- + NaCl
    SNaOCl  +
      3 NaCl
      and
2CuCN + 2NaOH
                    U20 	 2NaCNO + 2Cu(OH)2
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:
     3C1
2 + 2CNO- + 6   NaOH ---- > 2HC03-
                                N
                                                 6NaCL + 2H0
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
                                 235

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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.

The  reaction mechanism for the destruction of cyanide to cyanate
is generally expressed as:
     CN- + 03 ----
                          CNO-
The reaction mechanism for the subsequent  reaction,  destruction
of   cyanate,  has  not  been  positively  identified.   Proposed
mechanisms include  (Reference 24 and 25):
2CNO-
CNO-
             03
             OH-
 H20	:
H H20	

 or
 2HC03  + N2  +  302
» C03  -2  +  NH3
     CNO- + NH3 	>  NH2- CO- NH2
Regardless of the  actual mechanism, destruction of  cyanate can be
accomplished in approximately  30 minutes  (Reference 26).

Hydrogen Peroxide  Theory.  Two processes   for  the   oxidation  of
cyanide with hydrogen  peroxide (H2.O2.)  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-  n
      CNO-
      H20S
      i- 2H,0
               22
       CNO- H
      >  NH4
    H20
   h 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 1 29 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
                               236

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 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 VII1-2 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
 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  VII1-3 show,
 chlorine doses  in excess of stoichiometric amounts  were required,
 and  pH  was  found  to  be  a major   determinant   of  treatment
 effectiveness.     Results   on a  pilot-plant   scale  were  less
 favorable,  but  improved performance in the  full-scale  treatment
 system  is anticipated through use of a retention  basin  in which
 additional  oxidation of cyanide by  residual chlorine  can  take
 place.

 Mill   6102  has   built  a treatment  system employing lime  precipi-
 tation,   electrocoagulation-flotation,   ion   exchange,    alkaline
 chlorination,  and   mixed  media  filtration.  This  is followed  by
 final  pH adjustment.  The alkaline  chlorination   system  includes
 on-site  generation  of  sodium   hypochlorite by electrolysis  of
 sodium chloride.   The hypochlorite  is  injected   into  the  waste-
 water   following   the  electrocoagulation-flotation  process  and
 immediately preceeding  the filtration  unit.   At  this point  in the
 system,  some  cyanide removal  has  been  realized  incidental to  the
 lime   precipitation-electrocoagulation 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 cyanide concentration of approximately 60 to 70 mg/1.
                                 237

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The  initial  design of the alkaline chlorination treatment plant
was based on extensive laboratory and pilot-plant studies.  Final
design modifications, made in 1975, were based on  a  requirement
to comply with an amended discharge permit limitation of 0.5 mg/1
cyanide  (total).   Because  some cyanide is present in the 3,000
cubic meters (700,000 gallons) per day of tailings  not  treated,
the  alkaline  chlorination  treatment plant is operated with the
goal of destroying the  cyanide  contained  in  that  portion  of
wastewater being treated.  The composite waste stream should then
meet the permit limitation.

The treatment plant includes three FRP (fiber-reinforced-plastic)
tanks,  measuring 3 by 3 meters  (10 by 10 feet), which operate in
series, as reaction chambers.  Chlorine is added at a rate of-540
to  680  kilograms  (1,200  to   1,500  pounds)  per  day  from  a
chlorinator having a capacity of 900 kilograms  (2,000 pounds) per
day  of  chlorine.   The  pH  of  the  combined  waste  stream is
maintained between 11 and 12 by  lime  addition  to  the   incoming
waste    stream.    Process   controls   include   pH   and   ORP
(oxidation/reduction-potential)  recorders  and  a  magnetic  flow
meter.

The  average  cyanide concentration of the total tailing  effluent
discharge between July and December 1975 was 0.18 mg/1 of cyanide
(total).  This compares to the average concentration of 4.72 mg/1
of cyanide  (total) in the discharge prior to installation of  the
treatment  plant.  Performance data for the alkaline-chlorination
treatment plant at Mill 3144 are presented in the following table
(industry data):
     Source
      Chlorination-Plant  Feed
      Chlorination-Plant  Discharge
      Mine Drainage  (Overflow  from
       Backfill)
      Total  Combined Tailings
Total Cyanide
             68.3
              0.13

              0.06
              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-
                                 23B

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 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
 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
 complicates the  theoretical prediction of  phenol  concentrations
 in flotation-mill wastewater  streams.

A  more detailed account of reagent use,   including phenolic-based
 compounds  and  their  presence  in  mill   process  wastewater   is
provided in Section VI, Wastewater Characteristics.
                                239

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Several  methods  are  available  for  treating  phenolic wastes,
including:


     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
thearomatic  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 (C6H402.)
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.
                                  240

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 Hydrogen Peroxide.   In the presence of a  metal  catalyst  (e.g.,
 Fe++,   Fe+++,   A1+++,   Cu-n-,  and  Cr++),  hydrogen peroxide (H202)
 effectively oxidizes phenols  over a wide  range  of  temperatures
 and  concentrations.    Investigations  into  the  use of hydrogen
 peroxide show  phenol removal  efficiencies of  98+  percent  at  a
 dosage  rate of one  to two parts  of H2.02 per part phenol  and at an
 optimum  pH between  3 and 5.   Wastewater containing substituted
 phenols,  such  as  cresylic acid,  can increase the required  perox-
 ide  dosage to four parts of  H2O2 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  (KMnO4).  Phenol reductions  to  less  than  1
mg/1 are reported.  The destruction of simple phenol  by potassium
permanganate is expressed as:

   3C6H60 + 28KMnO£ +  5H2O  — 18CO2.  + 28KOH +  28MnO2

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 KMnOi Pgr part  of
                                241

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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).

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
                                  242

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     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:

     Na2SO4 + 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) + Na2S04

     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 + 2O2	Na2SO4

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  H2S,  which could cause odors and pose a
safety   hazard  to  personnel.   This   may  be  accomplished    at
essentially  the  same  point  as  the  sulfide  treatment, or  by
addition of a solution  containing  both  sodium  sulfide  and  a
strong   base  (such as caustic soda).    The sulfides of  many heavy
metals,  such as copper and mercury, are sufficiently insoluble  to
allow essentially complete  removal  with  low  residual  sulfide
levels.  Treatment'for these metals with close control  on sulfide
concentrations   could  be  accomplished  without  the  need  for
additional treatment.   Adequate aeration should  be  provided   to
yield an effluent saturated with oxygen.

Sulfide  precipitaition  is  presently practiced at most mercury-
cell chloralkali plants to control mercury discharges.   In  this
                                 243

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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
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
                                 244

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 of  the EPA and the Army Corps of Engineers  (References  41,   42,
 and  43).   Filtration  processes investigated included filtration,
 pressure   diatomaceous  earth  (DE)   filtration,   and  vacuum  DE
 filtration.                 .

 Forms  of  asbestiform  minerals evaluated in these studies include
 amphibole  and  chrysotile.   A  basic difference between these forms
 which  influences   treatment   is that   the  treatment  for   the
 amphibole    form    usually    carries   important   conclusions  and
 recommendations from   the  pilot-scale    studies   on   granular
 filtration,  including:

      1.  Granular  filtration  is successful in removal of asbesti-
      form  fibers.

      2.    Sedimentation prior to filtration increases filter run
      length  (i.e.,  time between backwashes)  but does not increase
      fiber removal.  (Note  that untreated  water for  these studies
      was Lake  Superior  water,  clean water  relative  to  mine/mill
      wastewater.    Sedimentation does remove some asbestos fiber
      in this industry  and would be necessary  to   prevent  filter
      clogging  and  frequent  filter backwashing.)

      3.    Two-stage flash-mixing followed   by   flocculation  is
      recommended for conditioning raw water  prior to filtration.

      4.  Alum  is a  more effective coagulant  than  ferric   chloride
      for Duluth raw water.

      5.  Nonionic polyelectrolyte is most  effective  in preventing
      turbidity  breakthrough.

      6.  A positive-lead mixed media filter  designed to  operate

      7.    Backwash  water should be discharged to  a sludge lagoon,
      and supernatant should be returned  to the treatment plant.

      8.    For   large  capacity  plants,  granular filtration  is
           recommended   over  DE filtration.   For  small plants, DE
           filtration should also be considered.

Two kinds  of DE filtration processes were  studied in  the  pilot-
scale  studies,  pressure filtration and vacuum filtration.   Flow
through both filtration  systems  ranged   from  0.0006   to  0.001
mVsec  (10  to 20  gal min).   Both kinds of  DE filters were oper-
ated  in various ways to evaluate  conditioning of   DE with  alum,
cationic polymers,  and anionic  polymers.    Single-step and twostep
precoat were studied.  Conditioned DE was  used in  filter  precoat,
as  well   as  for  body feed.    Various grades of  DE,   from fine to
coarse, were evaluated.  Details of pilot-plant   DE   testing  are
reported in References 41,  42,  and 43.
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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  riot  be  effective  on
warmer waters.)  For the conditions experienced during the pilot-
scale  studies  on  potable Lake Superior water, it was concluded
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/ft«).


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 VII1-4  (Reference 44).

Data  on pilot-plant asbestos removal from wastewater at specific
locations are  summarized   in  Tables  VII1-5   and  VII1-6.    The
authors  conclude  that sedimentation  followed  by  m'ixed-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).
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 Based   on the pilot-scale studies  discussed previously (Reference
 42), a  114,000-mVday  (30,000,000-gal/day)  granular  filtration
 plant   was designed  and constructed  for treating the water supply
 for  the city  of  Duluth,  Minnesota.   This plant  became operational
 in November  1976   (Reference   45).    Figure VIII-1   presents  a
 schematic  diagram   of   the  principal portion  of this full-scale
 plant.   The major components of  the  system  are  the  mixed-media
 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/m*  (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.
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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-
centration of greater than 5 x 10*  fibers/liter  in  the  filter
influent   to.  less  than  detectable  (approximately  3  x  10s
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 an'd comparison of pilot-scale results  from  treatment  of
raw  water  with  granular filtration and DE filtration discussed
earlier (Reference 42)  indicate  that  similar  results  can  be
obtained  from  the  two  systems.  Data in Tables VII1-5 through
VIII-7 (from Reference 44), however, indicates that DE filtration
may be more effective.  An economic analysis  of  both  types  of
systems has been conducted  (Reference 42).

Practices  iri 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.
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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 103
to 104 fibers/liters.  Examination  of  these  treatment  systems
indicates  (several  factors  in  common:   high initial suspended
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.
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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
by  the  requirements  for  periodic stripping or regeneration of
ionic exchangers.

Disadvantages of using ion exchange in treatment  of  mining  and
milling  wastewater  are  relatively high costs, somewhat limited
resin capacity,  and  insufficient  specificity  (especially,  in
cationic exchange resins for some applications).  Also, regenera-
tion  produces  a  waste,  and  its  subsequent treatment must be
considered.

For recovery of specific ions or groups of ions  (e.g.,  divalent
heavy-metal cations, or metal anions such as molybdate, vanadate,
and chromate), ion exchange is applicable to a much broader range
of  solutions.   This  use is typified by the recovery of uranium
from ore leaching solutions using strongly basic  anion  exchange
resin.   Additional  examples  are  the commercial reclamation of
chromate plating and anodizing solutions,  and  the  recovery  of
copper  and  zinc  from  rayon-production  wastewaters.  Chromate
plating and anodizing wastes have been purified and reclaimed  by
ion  exchange  on a commercial scale for some time, yielding eco-
nomic as well as  environmental  benefits.   In  tests,  chromate
solutions  containing  levels  in  excess  of  10  mg/1 chromate,
treated by ion echange at practical resin loading values  over  a
large  number  of  loading/elution (regeneration) cycles, consis-
tently produced an effluent containing no more than 0.03 mg/1  of
chromate.

High  concentrations of ions other than those to be recovered may
interfere with practical removal.  Calcium ions, for example, are
generally collected along with the divalent heavy  metal  cations
of  copper,  zinc,  lead,  etc.  High calcium ion concentrations,
therefore, may make ion exchange removal of divalent heavy  metal
ions   impractical   by  causing  rapid  loading  of  resins  and
necessitating unmanageably large resin  inventories  and/or  very
frequent elution steps.

Less  difficulty of this type is experienced with anion exchange.
Available resins have fairly high selectivity against the  common
anions,  such  as  Cl~  and  SO4~2.   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
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growth. Pretreatment 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, C02_, H2S, bacteria and algae,
and  hardness.   Since there is some latitude in selection of the
ions that are exchanged for the contaminants  that  are  removed,
post treatment may or may not be required.

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
                                 251

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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).

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 VII1-9.

Ion  Exchange.   At  uranium  Mill  9452,  a  unique   mine-water
treatment  system  exists  which  uses radium 226 ion exchange in
addition  to  flocculation,  barium   chloride   coprecipitation,
settling, and uranium ion exchange.  The mine water to be treated
is  pumped  from  an  underground  mine  to  a mixing tank, where
flocculant is added.  The water is then settled in two  ponds  in
series,  before  barium chloride is added.  After barium chloride
addition, the water is mixed and flows to two additional settling
ponds (also in series).   The  decant  from  the  final  pond  is
acidified  before it proceeds to the uranium ion exchange system.
The uranium ion exchange column effluent is pumped to the  radium
226  ion  exchange system.  After treatment for removal of radium
226, the final effluent is pumped to a holding  tank  for  either
recycle to the mill or discharge.

The  total  treatment  system  at  Mine/Mill   9452  is capable of
removing radium 226 from levels of 955  picocuries/liter  (total)
and  93.4  picocuries/liter  (dissolved) to 7.18 picocuries/liter
(total) and less than 1 picocurie/liter  (dissolved).   This  per-
formance  represents 99.2 percent removal of total radium 226 and
greater than 99 percent removal of dissolved radium 226.

Ammonia Stripping

High concentrations of ammonia  in  facility  wastewater  can  be
effectively  removed  by  air  stripping processes.  In this mass
transfer process, air and water are contacted  in a packed or  wet
column.   Water is sprayed from the top of the column and allowed
to trickle down the wood or plastic media in packed  columns,  or
fall  as droplets in wet columns.  Air is conducted in a counter-
current mode (from bottom to top of column) or a cross-flow  mode
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(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
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.
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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
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
                                 254

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reductions in molybdenum concentrations, as demonstrated  by  the
summarized results below.
Test Date (1975)
7/24 and 7/25
7/2 S and 7/29
7/29 and 7/30
7/31 and 8/1
8/1 and 8/2
8/5 and 8/6
Average

Feed
20.5
23.0
22.4
24.4
19.5
22.0
22.0
Mo(mq/l)
Effluent Eluate*
1 .18
0.91
1 .38
1 .76
1 .14
1.38
1.29

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 eluate allow
economical recovery of the molybdenum,  defraying  a  substantial
fraction  (or  possibly  all)  of  the  costs of the ion exchange
operation.  On  the  basis  of  pilot-plant  testing  results,  a
decision was made to install a full-scale ion exchange unit.

Laboratory   tests  of  precipitation  technology  at  this  site
indicate that, at the low effluent  temperatures  which  prevail,
conventional  precipitation  technology would not be effective in
removing heavy metals at retention times  considered  economical.
Electrocoagulation flotation was evaluated as an alternative, and
the  pilot-scale  unit was run to define optimum operating condi-
tions and  performance  capabilities.   Performance  achieved  at
various operating pH levels is summarized below:
Parameter
Concentration
(mq/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
                                 255

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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.

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.
under  consideration  will  provide
suspended  solids  concentration  of
reliability (industry report).
Full-scale designs currently
a  maximum  effluent   total
 5  mg/1  with  100  percent
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.
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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
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 VII1-14.   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

           Pb
           Zn
           Cu
           Fe
 Concentrations (mq/1)
Extractable (total)    Dissolved
      0.25                  0.24
      0.37                  0.26
      0.05                  0.04
      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

             Pb
             Zn
             Cu
             Fe
Concentration (mq/1)
   Extractable (total)
              0.12
              0.19
              0.04
              0.17
     3.   The initial acidity and total metal  concentrations  had
     little  effect  on  the  final effluent quality,  but greatly
                                257

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     influenced the volume and density of sludge produced;  these
     factors   affected  the  quantity  of  neutralizing  reagent
     required.

     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
     sol id/1 iqui'd 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

             pH (units)
             Fe
             Cu
             Mn
             Al
             Zn
             Pb
             Ni
             As
             Cd
             Sulfate
Concentration (mg/1)
          2.6
          336
         51 .6
         4.52
         20.8
          122
         0.04
         0.19
         6.01
         0.44
        1,400
to 3.1
to 800
to 128
to 19.0
to 62.5
to 294
to 0.50
to 0.51
to 22.0
to 1 .0
to 3,820
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The study was conducted specifically to evaluate the capabilities
of   the  treatment  scheme  depicted  in  Figure  VII1-2.   This
treatment consisted of a two-stage process of  chemical  addition
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  5.07  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:
     pH (Stage 1)*
     pH (Stage 2)*
     Fe (total)
     Zn
     Mn
     Cu
     Al
     Ni
     Cr
                   Exp. 1   Exp. 2   Exp. 3   Exp. 4   Exp. 5
  5.0
  6.4
  ND
 12.7
  6.4
  ND
  ND
LT 0.13
  ND
5.5
6.4
ND
0.3
0.5
ND
ND
0.13
ND
  5.0
  5.5
  ND
 30.0
  6.8
LT 0.5
  ND
  0.29
  ND
  5.0
  5.6
  ND
 30.0
  7.1
LT 0.3
  ND
  0.19
  ND
5.0
6.5
ND
0.2
0.4
ND
ND
0.13
ND
     *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 VII1-16, this method has been  demonstrated
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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
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
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with   a   theoretical  retention  time  of  approximately 6  to 9  hours
with  sulfuric  acid  neutralization  prior to discharge.

EPA Treatability  Studies

In August 1978, comprehensive   studies   of  the   treatability   of
wastewaterf 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
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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
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).  This 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)1 sent to the pilot-scale treatment system during the
period of study is presented in Tables VI11-19 and  VI11-20.   As
illustrated, pH, TSS, and total metals concentrations varied over
a  wide range during the August study.  This variability appeared
to be related to the schedule of mill operation  (Refer  to  Table
VIII-21).   During  periods when the mill was not operating, only
mine water was being discharged into the tailing pond.  (However,
this facility does not use  lime treatment;  alkaline  mill  water
provides  pH adjustment for mine water).  During the March study,
the concentrations of the parameters of interest  in  the  decant
were generally much higher  and less  variable than during August.

Two basic experimental designs were  employed to  investigate metal
removal  by  effluent polishing.   Initially, direct filtration of
the tailing pond decant was investigated.  Subsequently, a second
set of experiments was conducted to  determine the improvement  in
metals  removal  attained   by  lime  addition, coagulant (polymer)
addition, and settling prior to dual-media filtration.

A summary of the  treated  effluent  concentrations  attained  is
presented  in  Tables  VIII-22   (August study) and VIII-23  (March
study).  Results of experiments to evaluate  cyanide  destruction
by  ozonation  are  presented  in  Table VII1-24.  Conclusions and
observations made on the  basis of  these  results  are  summarized
below:

      1.   During the August study  the metal  removal efficiency of
      filtration was found to be dependent  on the pH maintained in
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     7.   During the March study the cyanide concentration in the
     tailing pond  decant  decreased  to  0.04  mg/1  before  the
     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/Mi11/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 VII1-25).

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 and 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  VII1-26.   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  m3/m2/day,  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-
                                  264

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27).   A  theoretical retention time of 11 hours was employed for
secondary settling experiments.  The effluent quality produced by
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).
 i
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 VII1-29.  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.

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Experimental systems identified as C and E in Table VI11-30  were
used  to investigate the extent to which the treatment efficiency
of the basic lime and settle treatment system could  be  improved
by  incorporating  polymer  addition and flocculation.  Review of
the experimental  results  demonstrates  that  polymer  addition,
followed  by  flocculation,  greatly improved the capabilities of
the basic lime and settle system.  Most notable was the threefold
improvement in removal  efficiency  of  total  suspended  solids,
zinc, and iron.

A  dual-media, granular filtration step was used with all systems
as a final  polishing  step  (systems  D,  F,  H,  and  J).   The
incremental improvements in removal of total suspended solids and
total  metals  resulting  from filtration are also represented in
Table VIII-30.   Results  indicate  filtration  is  an  effective
polishing  treatment  showing significant (14 to 5 mg/1) improve-
ments in TSS in all  cases  and  general  improvement  in  metals
concentrations.

On  the  basis  of the experimental results, the treatment scheme
producing optimum removals of suspended solids and  heavy  metals
from  acid  mine drainage at Mine 3113 consisted of adjustment of
pH to 10.5 with lime, flocculant  addition,  flocculation,  sedi-
mentation, and filtration (experimental system J in Table IX-30).
Other,  less  rigorous  treatment schemes with lower lime dosages
and without filtration would not reliably produce  the  excellent
effluent quality attained with system J.

Aluminum  Mine  5102.  At this site, bauxite is mined by open-pit
methods.  Watewater  (approximately  17,000  m3,  or  4.5  million
gallons,  per  day)  emanates  as runoff and as drainage from the
open-pit mine.  This wastewater  is  generally  characterized  as
acidic  (with  a  pH of 2.2 to 3.0), with total iron and aluminum
concentrations in the range of 50 to 150 mg/1 and 50 to 200 mg/1,
respectively.  The treatment system used for this mine water con-
sists of lime addition  and  sedimentation  in  a  multiple  pond
system.

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
                                 9.66

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treated was  already  of  a  very  high  quality  (low  pollutant
concentrations).

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  VII1-32.   This  is  to  be  expected  from  the  low
     dissolved metal concentrations in the mine water influent to
     the pilot-scale treatment units (Table VII1-31).

     2.  The use  of  a  polymer  improved  settling  performance
     during  lime  addition experiments,  species, i.e., arsenate
     (As04~3),  molybate   (Mo04~2),   selenite   )Se03-2),   and
     vanadates  (HVO^-2,  V0«.-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/m2/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.

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
second  phase  during  the  period of 3 to 12 December 1978.  The
basic  pilot-scale treatment scheme employed during  the  initial
phase  consisted of lime precipitation (pH adjustment), aeration,
flocculant (polymer) addition, barium  chloride  coprecipitation,
flocculation, single-stage or two-stage settling, and mixed-media
filtration.   Reagent  dosages  and operating parameters employed
                              267

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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  VII1-33
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
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
                                  268

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 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:
            +'• 3  (VO3)~i  +  30H-  --- >   Fe(VO3)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.

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 .                                              .

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Reduction of effluent TSS concentrations would also  improve  the
total   metal  concentrations  for  all  the  metals  subject  to
precipitation or coprecipitation removal mechanisms.

Uran i um/Vanad i um Mill 9401.  This mill  is  located  in  an  arid
region  of  New Mexico.  An alkaline leaching process is employed
at this mill to selectively leach uranium and vanadium from  ore.
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,  VO4~3,  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 H2SO4, 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
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.
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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  VII1-36.

The  results summarized   in  Table   VII1-38   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  BaCl2_
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 (SO4), BaCl,2 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 remove from wastewater.

Enhanced   metals  removal  was  observed  under   all  conditions
studied.  Optimum removal  of metals was achieved at  the   highest
ratio  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-
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
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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 VII1-7 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
phenols, and the 13 toxic metals.
solids,  cyanide,
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  VII1-40.   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.

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
                                 272

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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 2inc 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/Refinerv 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
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
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gpm/ft2  and  9.3  gpm/ft2  respectively.  Test 0.3 combined lime
addition to a pH of 8.8 with dual media filtration at a hydraulic
loading of 9.1 gpm/ft2.

The results of sample analyses indicate no significant change  in
the  already  low  concentrations of most pollutants.  TSS levels
dropped from an average of 4.1 mg/1 to an average  of  1.3  mg/1.
In  test  03,  the  pH  dropped to 7.7 through the filter column,
while causing partial clogging of the filter.  Lime  was  visible
in the filter effluent.  This phenomenon presumably is due to the
low  solubility  of lime in the facility wastewater which is high
in sodium and calcium chloride salts.  A summary of  treatability
study effluent sampling data is presented in Table VII1-43.

HISTORICAL DATA SUMMARY

This  subsection presents long-term monitoring data gathered from
individual  facilities  in  several  subcategories.    Facilities
considered  here  include  those  for  which  long-term  data are
available and which are regularly  achieving  or  surpassing  BPT
limitations by optimizing their existing treatment systems.

Iron Ore Subcategory

Mine/Mill 1108 is located in the Marquette Iron Range in northern
Michigan.   The  ore  body "consists  primarily  of  hematite and
magnetite  and  is  mined  by  open-pit  methods.   Approximately
8,800,000  metric  tons  (9,700,000  short tons) of ore are mined
yearly.  The concentration plant produces approximately 2,800,000
metric tons (3,100,000 short tons) of iron ore pellets  annually.
Wastewater  is presettled prior to treatment with alum and a long
chain  polymer  to  promote  flocculation  and  improve  settling
characteristics.   The  treated  water  is polished in additional
small settling ponds prior to discharge.

Table VIII-44 is a summary of industry supplied  monitoring  data
for the period January 1974 through April 1977.  As indicated, pH
is well controlled and always in the range of 6 to 9.  Alum and a
polymer  have  been  used  on  a  continuous  basis since 1975 to
improve TSS removal at  this  facility.   Since  that  time,  TSS
control  has  been  excellent and has exceeded 30 mg/1 only three
times on a daily maximum basis.  On a monthly average basis, this
facility  has  exceeded  20  mg  TSS/1  only  once  since   1975.
Dissolved  iron  concentration  averages approximately 0.36 mg/1.
Since 1975, dissolved iron concentration  has  not  exceeded  2.0
mg/1 on a daily basis or 1.0 mg/1 on a monthly basis.

Copper, Lead, Zinc, Gold, Silver, 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
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1976  was  approximately  3,281,000  metric tons (3,617,000 short
tons) of ore with 125,000 metric tons  (138,000  short  tons)  of
copper  concentrate,  and  185  metric  tons  (204 short tons) of
silver concentrate.  The  primary  mineral  form  is  chalcocite.
Concentration  is  accomplished  by  the froth flotation process.
The smelter and refinery contribute wastewater  to  the  combined
treatment   system.    Wastewater   also  originates  from  power
generation, sewage treatment, and  collection  of  storm  runoff.
Wastewater  from  the above sources is combined in a tailing pond
and decanted to a series of small settling  basins  before  final
discharge.  The alkaline pH of the treatment system is maintained
by  the alkaline nature of the discharge from the mill as well as
by the addition of lime to the slimes fraction of  the  tailings.
The  limed  slimes are combined with all other wastewater sources
in a mixing basin and then pumped into  the  tailing  pond.   The
mine  water  contribution to the total discharge ranges from 0 to
4,500 mVday (0 to 1.2 million  gal/day),  and  this  mine  waste
stream  is  released  into  the tailing pond on a seasonal basis.
The total pond discharge  volume  averages  approximately  79,000
mVday (approximately 21 million gal/day).

Discharge  monitoring data supplied by the company for a 58-month
period between March 1975 and  December  1979  are  presented  in
Table  VII1-45.   This summary presents data derived from monthly
averages for all parameters.  The data presented for pH  and  TSS
represent  almost  continuous  daily  monitoring  throughout  the
reporting period.  For these two parameters, the values shown are
based on approximately 1,500 measurements.  These  data  indicate
that,  for  the  parameters  monitored,  effluent  performance is
consistently far below BPT  effluent  standards,  even  when  the
maximum values reported are considered.

Wastewater treatment practices at Mine/Mill/Smelter/Refinery 2121
which are employed to attain its high quality effluent are:

     1.   Supplemental  lime  addition  for improved coagulation,
     metals removal, and pH control

     2.  Use of a multiple pond system for improved settling con-
     ditions and system control

     3.  Sufficient pond volume  to  provide  adequate  retention
     time for sedimentation of suspended particulates and metals
                                             j                    L
     4.   Provision  of  a tailing pond design resulting in rela-
     tively efficient and undisturbed sedimentation conditions

     5.  Use of a decant configuration which effectively controls
     pond  levels  without  disturbing  settled  solids  in   the
     vicinity of the decant towers
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     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 VII1-8 and VII1-9.  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 VII1-46 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
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.
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Figure VII1-9 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 VII1-46.  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
     conditions.
during  upset
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  VII1-47
do  not  accurately  reflect  the capabilities of the barrel pond
treatment system.  The reason for this, as  indicated  in  Figure
VI11-9,  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
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
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report  that  actions  are presently being taken to eliminate the
necessity for this practice.

Lead/Zinc/Copper Mine/Mill 3105

This underground mine is located  in  Missouri.   Galena,  sphal-
erite, and chalcopyrite (lead, zinc, and copper minerals) are the
primary  minerals  recovered.   Ore production began in 1973, and
reported mine production was  1,032,000  metric  tons  (1,137,700
short tons) in 1976.

Mining  and  milling  wastewater  streams are treated separately.
The mill operates in a closed loop system; tailings  are  treated
in  a  tailing  pond, and the pond decant is recycled back to the
mill.  Some mine water is  used  as  makeup  water  in  the  mill
flotation  process.   Excess  mine  water,  averaging 8,300 cubic
meters (2.1 million gallons) per day is treated by  sedimentation
in a 11.7 hectare (29 acre) settling pond.

Effluent  monitoring  data for the mine water treatment system at
Mine 3105 are presented in Table VII1-48.  This data  summary  is
based on NPDES monitoring reports submitted for this facility for
the period January 1974 through January 1978.

The  mine is an example of low solubilization of heavy metals due
to the mineralization of the ore body.   More  specifically,  the
ore  body  is  low  in pyritic minerals and exists in a dolomitic
host rock.  Mines exhibiting  low  solubilization  potential  are
characterized by mine waters of near neutral to slightly alkaline
pH.

Mine  water  treatment  at  Facility 3105 illustrates that simple
sedimentation at mines exhibiting  low  solubilization  potential
may  be sufficient to achieve water quality superior to BPT  limi-
tations, by effective removal of suspended solids and  associated
particulate metals  (see Table VII1-48).

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
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
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from operations which practice cut and fill techniques
discussed in Section VI.
has  been
Tables  VII1-49  and  VII1-50  summarize  data on raw and treated
waste streams by industry for the period  June  1977  to  October
1977.   A  new treatment system was recently brought online, so a
great deal of data are not available.  Examination of  raw  waste
data  indicates  that the mine water contains high concentrations
of metals.  As shown in Table VII1-50, 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
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
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was  recycled.   Approximately
was obtained in this manner.
56 percent of the mill feed water
Other wastewater control technologies included the segregation of
spills from the copper and zinc flotation circuits and control of
surface drainage with ditches and surface grading.

Effluent data submitted by the company for the period of  January
1974 to August 1977 are summarized in Tables VII1-51 and VII1-52.
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  VII1-51  and  VII1-52   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

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-
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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 VII1-53.  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
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
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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
VII1-54  and  VII1-55.   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  gal'ena.   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
metric tons  (124,400 short tons) of lead  and  zinc  concentrates
that year.

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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
Mine/Mill 3104 are:
the  effluent  quality  attained  at
     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  irs  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
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
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influence of surface drainage, daily discharge  volume  from  the
treatment  pond  system  averages  2,650  cubic  meters  (650,000
gallons) per day.

Table VIII-57 presents a summary of company monitoring  data  for
the period January 1974 to April 1977.  These data represent grab
samples  collected  once  monthly  for 40 months.  All parameters
analyzed  were  well  below  BPT   limitations   throughout   the
monitoring period.

Molybdenum Mine 6103

This  operation,  located  in  Colorado,  is  a  recently  opened
underground  mine  yielding  molybdenum  ore  at  the   rate   of
approximately  2,200,000  metric  tons (2,425,000 short tons) per
year.  A discharge of 9,100 m3 (2.4 million gallons) per  day  is
treated  by  spray cooling, and suspended solids are removed in a
multiple pond  system  with  the  aid  of  flocculants  prior  to
discharge.   The mill which recovers molybdenite by flotation, is
located some distance from the mine and is connected to the  mine
by  a  long  haulage tunnel.  Extensive recycle is practiced, and
there is no wastewater discharge at the mill site.

Table VIII-58 summarizes the limited data provided by the company
for  the  period  July  1976  to  June  1977.   In  general,  TSS
concentrations are well below 20 mg/1.  Effluent metal values are
reduced substantially below BPT limitations.

Molybdenum Mill 6101

This  facility,  which  uses the flotation process to concentrate
molybdenum ore, is located in mountainous terrain in New  Mexico.
Ore  is  obtained  from a large open-pit mine, with production at
5,700,000 metric tons  (6,300,000  short  tons)  per  year.   The
flotation  mill  produces  an  alkaline  tailings discharge which
flows approximately 16.1 kilometers (10 miles)  to  the  tailings
disposal  area,  where  sedimentation  in  primary  and secondary
settling ponds takes place.  The average  discharge  volume  from
this  treatment  system  is  11,000  cubic  meters  (4.6  million
gallons) per day.

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
concentrations from approximately 0.05 mg/1  to  less  than  0.02
mg/1.
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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 VII1-60.

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
problems  with  the  columns (most notably, excessive pressure at
flow exceeding 3.8 cubic meters, or 1,000  gallons,  per  minute)
                                285

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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.

Mixed-media filtration was incorporated into the treatment scheme
to  provide  effluent  polishing for optimum removal of suspended
solids and metals.
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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
(A12_O3_)  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 10,900 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 VII1-62 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 VII1-65 through VII1-67.  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
complex tungsten, molybdenum,  and copper ore is mined at the rate
of 640,00.0 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.
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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 VIII-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
U308_  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_O8_.

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.

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.
                                288

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Nickel Ore Subcateqory

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
VII1-70.   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 VII1-11 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 Subcateqory

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.

A summary of discharge monitoring data  from  Mine  6107  between
July 1978 and December 1980 is presented in Table VIII-71.  Rela-
tively  low  concentrations of suspended solids were consistently
reported.  The average total iron concentration  was  0.65  mg/1.
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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
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-
                                290

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 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  VII1-73 and VII1-74.   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-
 tively  removes the suspended  solids will also effectively remove
 the metals.
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Results of the pilot-scale treatability studies are presented  in
summary  form in Tables VIII-75, VIII-76, VIII-77, and VIII-78. A
review of the results presented in Table VIII-75  indicates  that
secondary  settling at a theoretical retention time of 10.4 hours
was sufficient to produce  effluent  total  metal  concentrations
well  below BPT limitations.  In a full scale system, even longer
times (24 to 72 hours) would be recommended to  reliably  achieve
this  limit.  A larger pond would also provide protection against
surge loads and short-circuiting.

Polymer and lime addition,  followed  by  flocculation  prior  to
settling  (2.8 hour retention time), produced effluent suspended-
solids concentrations comparable to those achieved  by  secondary
settling with a longer retention time (10.4 hours).  The observed
improvement  in efficiency of removal of TSS at shorter retention
time is attributed to the addition of polymer.  This  is  further
evidenced by the fact that treatment schemes employing lime addi-
tion and settling resulted in much higher suspended solids levels
in the effluent when a polymer was not employed.

Experiments employing lime addition were conducted to investigate
its  effect on dissolved metal precipitation and suspended solids
settleability.  However, because dissolved  metal  concentrations
in  the  tailing  pond recycle water were already very low (i.e.,
less than 0.04  mg/1),  this  treatment  provided  little  or  no
benefit.

Three   dual-media,  downflow,  pressure  filters  consisting  of
different filter-media sizes and depths, were evaluated over a  a
range  of  hydraulic  loadings  of  117 to 880 mVm2/day  (2 to 15
gpm/ft2).   All  three  filters  employed  consistently  produced
filtrates  with  suspended  solids concentrations of less than 10
mg/1 throughout the range of 30 to 50 mg/1.   On  two  occasions,
however,  filter  performance  was  adversely  impacted  by shock
loads.  At these times, the suspended  solids  concentrations  of
the  tailing  pond  recycle water being treated ranged upwards to
several percent solids, and the filtrate concentrations  attained
were 13 and 30 mg/1.

Because dual media filtration at hydraulic loadings of 293 to  880
m3/m2/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
frequency of backwash required appears to be  much   greater  than
the frequency required for a loading of  10 gpm/ft2 or less.
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 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   VII1-70.    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 VII1-70 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.

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
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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  ozonation
are also presented in Table VII1-78.  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 and BAT  effluent   limitations  and
standards of performance governing the placer mining of gold were
reserved  because of insufficient economic data and effluent data
from well managed plants.  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
deposits are deeply buried,  bulldozers,   front-end loaders,  and
draglines  are  being  used   for  overburden  stripping,  sluicebox
                                 294

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 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    splids   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
 500.    EPA  Region  X  has  issued  National  Pollution Discharge
Elmination System  (NPDES)  permits to many placer  miners   in  the
                                295

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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 the 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
generally  small  in size relative to  the volume of wastewater  to
be treated.  Therefore, it  is not  unusual that  these  ponds  are
                                 296

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 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
 involved,  geographic  location,  age, disturbed  area, and geology
of  the  mined  areas.   There   is  an  abundance  of   literature
                               297

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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
adequately treated  is  suitable for discharge to   surface  waters,
and this  practice  is also common  to this  industry.
                               298

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 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, and the  geological conditions prevalent  at   each
 site.    Unlined   ponds    are  best  used  under  the following
 circumstances:

     1.  Deep groundwater table and/or soils   exhibiting  permea-
     bilities sufficiently  low to minimize the volume  of  seepage,

     2.  Native  soils with  significant capacity  to remove and  fix
     pollutants  from seepage,

     3.  Arid climates, and

     4.   Geological and hydrological conditions at  the site pre-
     cluding contamination of aquifers or  other  bodies  of  water
     which  are useful as water supplies.


The seepage rates from  unlined uranium mill ponds depend upon  the
characteristics  of the tailings,  soils, underlying geoglogy,  and
hydrologic  conditions prevalent at the site.  Soils  and  tailing
deposits  exhibiting  high permeabilities may permit high seepage
rates,  especially if sandy soils underlie  the pond.  If ponds  are
located on soils containing  high proportions of natural  clay  or
on   impervious rock (such as shale),  seepage rates can be reduced
substantially.   Permeability values as low as  10~« to  10-8 cm/sec
 (down to 0.006 gal/min/acre) can often be  achieved  under  these
circumstances.
                                299

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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
     liners in tailing pond applications  is lacking; and
of
     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
results  in a reduction of the swelling capacity  of the  clay   and
in eventual failure of the seal.
                              300

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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
commonly used machinery,
chemical  constituents.
product is effective in
dissolved solids.  This
liner  for  a  new pond
through use of a slurry
 advantages of being easy to install with
 and they are relatively  inert  to  most
  One supplier of treated clay claims its
sealing ponds containing up to 20 percent
product may be  used  in  constructing  a
or may be used to control lateral seepage
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
     environment or exposure to the elements; and
     3.  They are more prone to puncture and tear during
     lation and may pose difficulties in field handling.
                                  instal-
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.

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
                               301

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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
pond seepage  (Reference  70).   A  system   such  as  this  can  be
effective  only  when  specific   favorable  subsurface conditions
prevail.
                                302

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 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 arroyd,  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 mill  in Eastern Ontario,  Canada,   is controlled by a grout
curtain,  constructed  of   clay,  bentonite, and  cement (Reference
70).  The grout slurry was  injected into  the subsurface alluvium,
down  to an  impervious bedrock layer,   and   forms   an   underground
barrier  to  the  lateral  flow of seepage  to a nearby  recreational
lake.  Monitoring the concentrations of dissolved  radium  226  in
the  groundwater  has  demonstrated the effectiveness  of the cur-
tain.   The  effectiveness  of   this  method  is   attributed   to
increased flow-path length and ion exchange with the montmorillo-
nite clay in the grout.
                              303

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                  TABLE VIIM.  ALTERNATIVES TO SODIUM CYANIDE FOR FLOTATION CONTROL
ORE
copper with
pyrite
copper/iead/zinc
with pyrite
zinc with
pyrite
PH
LEVEL
natural
10 to 12
natural
10 to 12
natural
10 to 12

DEPRESSANT
Sodium cyanide
Sodium sulfite
Sodium cyanide
Sodium sulfite
Sodium
monosulfide
Sodium cyanide
Sodium sulfite
Sodium cyanide
Sodium sulfite
Sodium
monosulfide
Sodium cyanide
Sodium sulfite
Sodium cyanide
(none)
Sodium
monosulfide
Sodium sulfite
DEPRESSANT USAGE'
kg/metric ton
ground ore
0.196
0.504
0.196
0.006
0.312
0.196
0.504
0.196
0.504
0.003
0.196
0.005
0.002
—
0.003**
0.504
Ib/short ton
ground ore
0.392
1.008
0.392
0.013
0.624
0.392
1.008
0.392
1.008
0.006
0.392
0.010
0.004
-
0.006
1.008
PERCENTAGE
- RECOVERY
copper
62.0
56.6
60.9
65.4
67.5
73.6
74.6
58.2
74.2
73.4
NA
NA
NA
NA
NA
NA
lead
NA
NA
NA
NA
NA
78.8
74.9
75.0
80.3
75.9
NA
NA
NA
NA
NA
NA
zinc
NA
NA
NA
NA
NA
NA
NA
NA
NA
NA
7.97
11.6
9.11
11.2
13.1
12.9
PERCEN
DEPREi
iron
84.1%
81.9
82.6
69.5
-66.4
85.7
87.3
87.4
83.9
84.2
93.3
91.2
92.6
91.9
91.4
91.2
TAGE
5SION
zinc
NA
NA
NA
NA
NA
77.8
82.9
77.2
72.7
53.3
NA
NA
NA
NA
NA
NA
DEPRESSANT COST
per metric ton
ground ore
$0.18
$0.14
$0.18
< $0.01
$0.11
$0.18
$0.14
$0.18
$0.14
<$0.01
$0.18


-
<$0.01
$0.14
per short ton
ground ore
$0.16
$0.13
$0.16
<$0.01
$0.09
$0.16
$0.13
$0.16
$0.13
<$0.01
$0.16


-
< $0.01
$0.13
OJ
-p.
         Based on References 37 and 38.
         NA = not applicable
         * Based on laboratory-scale experiments using 500 grams of ground ore.
         **0.312 kg/metric ton ground ore not studied

-------
TABLE VIII-2.
RESULTS OF LABORATORY TESTS OF CYANIDE DESTRUCTION
BY OZONATION AT MILL 6102
PH
5.2
7.4
8.1
9.3
9.4
10.2
11.4
12.7
Final Cyanide Concentration
0.36
0.09
0.06
0.05
0.04
0.03
0.02
0.04
(mg/l)








           Based on ozone dosage =10 times stoichiometric;
           15-minute contact time; and initial cyanide concentration of 0.55 mg/l.
                                    305

-------
TABLE VIII-3. RESULTS OF LABORATORY TESTS AT Ml LL 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/l

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.
                                    306

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TABLE VIII-4.  EFFECTIVENESS OF WASTEWATER-TREATMENT 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
4x1012
4x1012
4x1012
4x1012
Treated Water
5x10ntto1x101'**
ixio9
3x106
1x105
      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
                                          307

-------
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
5 x 109*
5x109*
Treated Water
9.3 x 109t
3.2 x109**
   Source: Reference 44
   Corresponding turbidities are:
     »620 JTU's
     t  1.0 JTU's
    •»  0.38 JTU's
                                     308

-------
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 TO9
1 x 109
1x109
Treated Water
3 x 107
3 x 106
8x104
      Source:  Reference 44
                                    309

-------
TABLE V1II-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)
1x109(1x1011)
1 x 109
1x109
Treated Water
1 x 109 (1 x 1010)
1 x 108 (1 x 109)
2x106
<1x105
    Source: Reference 44

    Parentheses enclose results for a second sample.
                                       3in

-------
       TABLE Vlll-8. COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS FOR TOTAL FIBERS AND
                   CHRYSOTILE AT SEVERAL FACILITIES SURVEYED
FACILITY
4401
(Mine-Water Settling Pond)
4401
(Tailing Pond)
5102 (Mine-Water Treatment System)
5102 (Mine-Water Treatment System)
2122 (Tailing Pond)
2122 (Tailing Pond)
2122 (Tailing Pond)
21 22 (Tailing Pond)
2122 (Tailing Pond)
2121
(Treatment System)
2120
(Tailing Pond)
2120
' (Tailing Pond)
2120
(Mine-Water Treatment System)
2117
(Treatment Plant)
TYPE OF
SAMPLING
S
S
S
V
S
V
S
V
V
S
S
V
S
S
INFLUENT CONC.
(fibers/l'iter)
TOTAL FIBERS
3.8 x 107
7.1 x1011
3.5 x 107
3.6 x 107
2.5 x 1012
ND
6.1 x 1012
ND
ND
3.0 x1011
1.2 x1012
1.3 x1013
4.6 x 107
2.5 x 108
CHRYSOTILE
1.1 xlO7
1.1 x1011
5.5 x106
5.5 x 106
4.3 x1011
ND
6.2 x1011
ND
ND
5.5 x 1010
3.1 x1011
1.7 x1012
1.8x106
5.5 x 106
EFFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
5.7 x 107
2.1 x 109
1.4x109
1.0 x 108
4.3 x 109
6.3 x 10S
3.7 x 107
2.7 x 108
1.3 x107
8.2 x 106
1.2x109
7.8 x 107
7.2 x 107
3.4 x 106
CHRYSOTILE
1.1 x106
1.8 xlO8
2.0 x 108
3.3 x 106
6.7 x 108
<2.2 x 105
8.2 x 106
<2.2 x 105
9.1 x 105
5.5 x 105
3.0 x 108
1.2 x 107
8.2 x 106
<2.2x105
TREATMENT
REDUCTION FACTOR

-
>102
INC
INC
~103
—
~105
—
—
104-105
103
105-106
-
~102

10
~103
INC
—
~103
_
~105
—
—
105
103
~105
-
>10
S = screen sampling
V = verification phase
INC = increase
ND * no data

-------
TOR
SOTILE
3-104

-------
             TABLE Vlll-8. COMPARISON OF TREATMENT-SYSTEM EFFECTIVENESS FOR TOTAL FIBERS AND
                         CHRYSOTILE AT SEVERAL FACILITIES SURVEYED (Continued)
FACILITY
9402
(Mine-Water Treatment System)
9408
(Mine-Water Treatment System)
9408
(Mine-Water Treatment System)
9405
(Mill Settling Pond)
9405
(Mill Settling Pond)
9411
(Mine-Water Treatment System)
6104
(Mine-Water Treatment System)
TYPE OF
SAMPLING
S
S
V
S
V
S
S
INFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
1.2 xlO8
1.6 x 109
1.4x 108
2.9 x 108
1.0 x 108
2.3 x109
7.7 x 106
CHRYSOTILE
5.2 x106
1.9 x 108
1.5 x107
2.3 x 107
<2.2x105
1.1 x 108
<2.1x105
EFFLUENT CONC.
(fibers/liter)
TOTAL FIBERS
4.3 x 108
2.3 x 109
7.3 x 106
1.2x109
6.6 x 107
5.7 x 108
3.3 x 107
CHRYSOTILE
5.3 x 107
2.0 x 108
<2.2 x 105
1.5 x 108
<2.2 x 105
2.7 x TO7
8.2 x 106
TREATMENT
REDUCTION FACTOR


-
10-102
INC
<1°
<10
INC

INC
-
~102
INC
-
<10
INC
CO
(->
CO
       S = screen sampling
       V = verification phase
       INC = increase
       ND = no data

-------
                      TABLE VIII-9. EFFLUENT QUALITY ATTAINED BY USE OF BARIUM SALTS FOR
                                     REMOVAL OF RADIUM FROM WASTEWATER AT VARIOUS URANIUM
                                     MINE AND MILL FACILITIES
oo

OPERATION


94031
Mills 94051
9405*8
9411
941 1*9
Mines 94121'6
9408
9408»
94523
AMOUNT
(mg/l)
OF BaCI2
ADDED
7A2
9.5
9.5*
5
10
10.4
55
55
45
RADIUM CONCENTRATIONS (picocuries/l)


BEFORE BaCI2 TREATMENT
Total
111 (±1.1)
15.9 (±1.6)
39.2 (±3.9)
35.4 (±0.3)4
56.9 (±5.7)
48.9 (±0.2)
123.6 (±1.5)7
142 (±14)
955
Dissolved
__
—
33.3 (±3.3)
15.5 (±2.0)5
60.2 (±6.0)
4.7 (±0.1)
37.7 (±0.3)4
120 (±12)
93.4

AFTER BaCI2 TREATMENT
Total
4.09 (±0.41)
<1.0
5.05 (±0.5)
8.4 (±0.1 )4
<2
10.9 (±0.2)
2.1 (±0.23)7
1.1 2 (±0.11)
7.18
Dissolved
_
—
<2
0.2 (±0.1 )5
—
1.6 (±0.1)
0.6 (±0.1 )4
<0.9
<1
PERCENTAGE
OF RADIUM
REMOVED
Total
96.3
>93.7
87.1
76.3
>96
77.7
98.3
99
99.2
Dissolved
—
-
>93.9
98.7
—
66.0
98.4
>99
>99
               1. Data obtained from single grab sampling and analysis (April, 1976).
               2. Calculated value based on average flow and annual BaCI2 usage.
               3. Includes ion exchange treatment; facility visited August 1978.
               4. Data obtained from analysis of two grab samples (April, 1976).
               5. Company data for February 1975 (Average of 12 grab samples).
               6. Final discharge to dry watercourse.
               7. Colorado Dept. of Health data for period January 1973 through February 1975 (Average of 24 samples analyzed for
                 "extractable" Ra 226).
               8. Data obtained from composite of two grab samples representing two separate influent points (May, 1977).
               9. Note that the dosage has doubled apparently enhancing the treatment system efficiency.
               * Updated data obtained during sampling trips occuring April-May, 1977. All samples, unless otherwise indicated, are 24-hr
                 composites.
               t Dosage rates are assumed to remain the same as previous rates.
               ( ) Parenthetical values indicate analytical accuracy.

-------
TABLE VIIMO. 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.
                                      315

-------
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.
                                             316

-------
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.
                                              317

-------
TABLE VIII-13. RESULTS OF COMBINED MINE WATER + BARREN LEACH SOLUTION +
               MILL TAILINGS TREATMENT* BY LIME ADDITION AT COPPER
               MINE/MILL 2120
SAMPLE
Mine water +
mill tailings +
barren leach water



Tailings (control)
SAMPLE
Mine water +
mill tailings +
barren leach water



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.
                                           318

-------
   TABLE Vlll-14. CHARACTERISTICS OF RAW MINE DRAINAGE TREATED DURING
                PILOT-SCALE EXPERIMENTS IN NEW BRUNSWICK, CANADA
PARAMETER
pH*
Sulfate
Acidity (as CaCO3)
Cu
Fe
Pb
Zn
Suspended Solids
CONCENTRATION (mg/l)
MINE1
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 3,211
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
MINES
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
                                     319

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 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 1 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 0.1 5)
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 0.1 9)
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
(0.1 2 to 2.51)
0.20
(0.04 to 0.41)
0.11
(0.02 to 0.20)
0.47
(0.09 to 1.40)
0.26
(0.01 to 0.61)
0.22
(0.01 to 0.89)
COMPARISON OF EFFLUENT QUALITIES DURING PERIODS OF OPTIMIZED STEADY OPERATION
1


2


3


Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Bucket-Settled
Sand-Filtered
Clarifier Overflow
Basin-Settled
Sand-Filtered
0.18
(0.01 to 0.35)
0.21
(0.16 to 0.25)
0.15
(0.14 to 0.15)
0.44
(0.25 to 0.62)
0.29
(0.01 to 0.50)
0.29
(0.1 7 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 0.1 5)
0.03
(0.02 to 0.04)
0.19
(0.10 to 0.26)
0.18
(0.1 5 to 0.21)
0.20
(0.1 4 to 0.26)
0.42
(0.14 to 0.45)
0.17
(0.04 to 0.29)
0.13
(0.11 to 0.17)
0.26
(0.09 to 0.60)
0.24
(0.10 to 0.30)
0.14
(0.08 to 0.23)
Source: Reference 54
                                     320

-------
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

Total
Dissolved
Dissolved
Dissolved
Dissolved
Pb
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
                              321

-------
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.
 •All metals concentrations are based on "total" analyses.
 tValue in pH units
"Theoretical retention times were variable and not always specified.
N.A. = Not analyzed
                                          322

-------
TABLE Vlll-18. EPArSPONSORED 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)* i
Mine3113(Pb/Zn)*
Mine 5102 (Al)
Mill 9401 (U)»»
Mill 9402 (U)** j
PERIOD OF STUDY
August 3-1 3, 1978
March 19-29, 1979
August 14-19, 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
Chlorination) 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,
Flocculation, Sedimentation, Filtration
Investigated
Polishing Treatments (Filtration, Secondary
Settling) and Lime Precipitation Investigated
Bench-Scale and Pilot-Scale Investigation of
IX, pH Adjustment with H2SO4, Ferrous
Sulfate Coprecipitation, Barium Chloride
Coprecipitation, Settling, and Filtration.
Treatment Scheme Employing Lime Addition,
Aeration, Barium Chloride Addition, Settling,
and Filtration Investigated
    * Operations in base and precious metals subcategory
    •'Operations in uranium ore subcategory
                                            323

-------
TABLE VHI-19.  CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
              TREATMENT TRAILER) AT LEAD/ZINC MINE/MILL 3121
POLLUTANT
PARAMETER
PH
TSS
Sb
As
B«
Cd
Cr
Cu
Pb
Hfl
Ni
St
Afl
T1
Zn




NUMBER OF
OBSERVATIONS
75
13
























1


"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
7.8
4.5
<0.5
0..001
<0.002
0.002
<0.01
0.10
0.21
0.0002
<0.02
0.002
0.01
<0.01
0.74


RANGE
6.7 - 9.1
1-10
<0.5
<0. 001-0. 025
<0.002
0.005-0.011
<0.01
0.02-0.16
0.18-0.25
<0. 0002-0. 0005
<0.02
<0. 002-0. 004
<0. 01-0. 05
<0.01
0.25-1.25


"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
n.














a.













r


RANGE
n.














a.













r


   Based on observations made in period 6 through 10 August 1978.
   n.s. « Not Analyzed.
                        32

-------
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)
(disi.)
Cr
Cu (total)
(di«.)
Pb (total)
(diss.)
Hg
Ni
Se
Ag
T1
Zn (total)
(di».)
Fe (totalj
(diis.)
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
-
-
-
0.1)15-0.0211
<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
                            3 25

-------
TABLE VIII-21.  OBSERVED VARIATION WITH TIME OF pH AND COPPER AND
             ZINC CONCENTRATIONS OF TAILING-POND DECANT AT LEAD/ZINC
             MINE/MILL 3121
DATE
(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/l)
PH»
6.9 to 7.4
6.9 to 7.7
7.3 to 7.7
6.7 to 7.7
7.6 to 7.8
7.4 to 7.6
7.2 to 7.6
7.1 to 7.4
8.2 to 8.3
8.2 to 8.3
8.5 to 8.6
8.5 to 8.9
8.6 to 9.1
Cu
0.004
0.05
0.02
*0.08
0.08
0.09
0.11
0.14
0.14
0.12
0.16
0.14
0.13
Zn
1.1
1.2
1.3
0.94
0.75
0.76
0.79
0.80
0.55
0.44
0.49
0.28
0.25
COMMENTS
Mill not operating
Mill not operating
Mill not operating
Mill startup 4:30 PM, Aug. 7
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
Mill operating
'Value in pH units
                                      326

-------
TABLE Vlll-22. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH
              PILOT-SCALE UNIT TREATMENT PROCESSES AT MINE/MILL 3121
              DURING AUGUST STUDY
UN IT TREATMENT
PROCESS EMPLOYED
Secondary Settling (approx.
1 1- 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 1 1.3, 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)
prf
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**
«1to
2)
<1«*
(0-1)
Cu
0.11
0.05
0.02
0.03
0.02
0.02**
(C0.01 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 analyzed
                                      327

-------
 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
PH*
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)
jng/D*
Zn
1.4
(0.96-1.8)
1.0
(0.52-1.4)
0.26
(0.040-0.48)
  pH Units
* Values given are mean and range, in ( ), concentrations
**All metal concentrations are based on "total" analyses
                                     328

-------
CO
I\J
IO
                        TABLE VII1-24. RESULTS OF OZONATION FOR DESTRUCTION OF CYANIDE
                           Initial CN~ concentration =0.11
0_ Dosage
mg/min Ratio 0,/CN
1.4
0.48
0.36
0.7
0
10:1
10:1
10:1
5:1
-
Retention Time
(min)
10
30
45
10
10
Final CN Concentration
Cmg/1)
0.066
0.080
0.081
0.108
0.115

-------
TABLE VIII-25. EFFLUENT FROM LEAD/ZINC MINE/MILL/SMELTER/
               REFINERY 3107 PHYSICAL/CHEMICAL-TREATMENT PLANT
PARAMETER
pH*»
TSS
Cd
Pb
Zn
Hg
CONCENTRATION (mg/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
                                          330

-------
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






















t


"TOTAL" CONCENTRATION
(mg/0
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.a.



'



t
0.036
n.a.
0.021
0.073
n.



i
a.



>
0.055-


RANGE
n



1
. a.



f
0.025-0.050
n.a.
0.020-0.030
<0. 02-0. 17
n.



i
a.



r
0.030-0.12


  Based on observations made in period 14 through 19 August 1978.
  n.a. = Not analyzed.
                                            331

-------
TABLE Vlll-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**
KD
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)
H9
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
     ••Average concentrations attained
     () Range of concentrations attained
    n.a. Not analyzed
                                                    332

-------
TABLE VIII-28. CHARACTER OF DRAINAGE FROM LEAD/ZINC MINE 3113
PARAMETER
pH**
TSS
TDS
so4 =
Cd
Cu
Fe
Pb
Ag
Zn
Hg
As
CONCENTRATION (mg/1)*
Average
4.2
111
1687
813
0.13
0.60
90
0.070
0.01
44
<0.001
0.021
Range
2.9 to 7.5
86 to 322
214 to 9,958
485 to 3.507
< 0.01 to 0.50
0.18 to 1.6
3.6 to 522
0.01 to 0.35
<0.005to0.20
1.5 to 76
< 0.001
<0.01 to 0.040
                  Based on industry data collected during the
                  period of 1970 through 1978.
                 *AII metals concentrations are based on "total'
                  analyses
                **Value in pH units
                                              333

-------
TABLE VHI-29. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
              TREATMENT TRAILER) AT LEAD/ZINC MINE 3113
POLLUTANT
PARAMETER
• null 1 —^»^^^— »»^— ^—
PH
TSS
Sb
As
Be
Cd
Or
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
Fe
so4-


NUMBER OF
OBSERVATIONS
^
53
7






























t
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
=====
3.2
112
<0.1
0.013
<0.003
0.23
0.011
1.5
0.088
<0.0002
0.074
0.006
n.a.
<0.05
71
69
1063
RANGE
=====
3.0 - 3.3
104 - 124
<0.1
0.005 - 0.030
<0.003
0.22 - 0.24
0.010 - 0.015
1.3 - 1.6
0.033 - 0.12
<0.0002
0.060 - 0.090
<»
0.003 - 0.010
n.a.
<0.05
67 - 74
50 - 80
925 - 1320
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
ssszsr^ssssssssssss
n.a.
n.a.
n. a.
n.a.
n.a.
0.23
n.a.
1.4
n.a.
n.a.
n.a.
n.a.
n.a.
n.a.
57
25
n.a.
RANGE
n.a.
n.a.
n..a.
n.a.
n.a.
0.23 - 0.24
n.a.
1.4 - 1.5
n.a.
n.a.
n.a.
n.a.
n.a.
n.a.
56 - 60
22 - 29
n.a.
    Based on observations made in period 24 through 28 September 1978.
    n.a. = Not Analyzed.
                                               334

-------
                          TABLE Vlll-30. SUMMARY OF PILOT-SCALE TREATABILITY STUDIES AT MINE 3113
oo
co
.EXPERIMENTAL
SYSTEM
A
B
C
D
E
F ,
G
H
1
J
UNIT TREATMENT
PROCESS EMPLOYED
Lime Addition to pH ~ 9.5, Sedimentation
Lime Addition to pH ~ 9.5, Sedimentation,
Filtration
Lime Addition to pH ~ 9.5, Aeration,
Sedimentation
Lime Addition to pH ~ 9.5, Aeration,
Sedimentation, Filtration
Lime Addition to pH ~ 9.5, Polymer
Addition, Flocculation, Sedimentation
Lime Addition to pH ~ 9.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration
Lime Addition to pH ~- 8.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation
Lime Addition to pH ~ 8.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration
Lime Addition to pH ~ 10.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation
Lime Addition to pH ~ 10.5, Aeration,
Polymer Addition, Flocculation,
Sedimentation, Filtration

PHf
9.3*"
(9.1 to
9.7)
8.8*»
(8.4 to
9.0)
9.7"
(9.7 to
9.8)
9.5»"
(9.4 to
9.6)
9.3**
(8.8 to
9.8)
8.9**
(8.1 to
9.5)
8.4**
(7.5 to
8.3)
7.9**
(7.3 to
8.3)
10.5**
(10.2 to
10.7)
10.2**
(9.5 to
10.5)
EFFLUENT CONCENTRATIONS* ATTAINED (mg/l)
TSS
33
<2**
«1to
3)
35
1
10**
(10)
<1»*
(0)
6

-------
TABLE VIII-31  CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
              TREATMENT TRAILER) AT ALUMINUM MINE 5102
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Ba
Cd
Cr
Cu
Pb
Hg
N!
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.












a.












t
<0.01
<0.02
<0.02
n.a.
RANGE
n.













a.












r
<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.
                                             336

-------
TABLE Vlll-32. SUMMARY OF PILOT-SCALE TREATABILITY STUDIES AT MINE 5102
UNIT TREATMENT
PROCESS EMPLOYED
No Treatment (Control)
Lime Addition to pH <%> 8.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition to pH ~ 8.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation,
Filtration
Lime Addition to pH ~ 9.0,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition to pH~ 9,0,
Aeration, Polymer Addition,
Flocculation, Sedimentation,
Filtration
Lime Addition to pH +* 10.4,
Aeration, Flocculation,
Sedimentation
Lime Addition to pH "10.2,
Aeration, Polymer Addition,
Flocculation, Sedimentation
Lime Addition 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**
«1to1)
<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**
«0.2to0.2)
Fe
0.15**
(0.020 to 0.24)
0.15
0.067**
(0.06 to 0.07)
0.06
0.08
0.2
0.23
0.073**
'(0.07 to 0.08)
0.040**
(0.2 to 0.10)
   * All metals concentrations are based on "total" analyses
     Value in pH units
   "Average concentrations attained
   ( ) Range of concentrations attained
                                                  337

-------
TABLE Vlll-33. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
TABLfc VIII A3. !J:R£ATMENT TRAILER) AT MILL 9402 (ACID LEACH MILL WASTEWATER)
POLLUTANT
pH
TSS
Sb
At
B»
Cd
Cr
Cu
Pb
Hg
Ni
Sa
Ag
T1
Zn
V
Mo
Fe
Mn
At
so4
TDS
R«226
U
PERIOD OF
OBSERVATIONS
(DATES)
1/3-8/1978
















































NUMBER OF
OBSERVATIONS
5









































•
3/4
3/4
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
1.7
40
<0.5
2.5
0.03
0.05
0.67
4.0
0.93
<0.0002
1.4
2.0
<0.1
<0.2
6.1
100
10
1900
118
786
21760
--
99.7±i%
17
RANGE
1.6-1.9
24-48
<0.5
1 .6-3.3
0.03
0.04-0.05
0.46-0.82
2.7-6.0
0.65-1.1
CO. 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)*!%
(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
--
31,520
88.3±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
--
28,100-35,300
(58.4-127)±l%
(12±9%) to
(20±12%)
                                        3.38

-------
TABLE VIII-34. CHARACTERIZATION OF RAW WASTEWATER (INFLUENT TO PILOT-SCALE
            TREATMENT TRAILER AT MILL 9402 (ACID LEACH MILL WASTEWATER)
'OLLUTANT
>ARAMETER
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





•







i










*




NUMBER OF
OBSERVATIONS
8
3
3
3
3
3
3
3
3
3
3
3
3
3
3

"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
1.6
168
0.86
3.2
1.3
1.0
109
17
3,670
287
1,840
19,600

154.8±1%
19.8

RANGE
1.4-1.8
142-215
0.73-0.98
2.9-3.6
1.1-1,5
0.95-1.1
97-110
15-20
3200-4100
260-310
1,640-2,220
18,400-20,400
-
(168)±1%
16-24.1

"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
X
—
- - . -•- . -
0.82
3.1
1-3
0.93
93
17
• ;.2,32f
157
1,330
--
32,200 .
129±1%
20±12%

RANGE
" __
, . , , __
0.71-0.93
2.9-3.4
1.2-1.3
0.87-1.0
89-98
15-19
2,100-2,600
150-160
1,170-1,480
--
29,700-34,200
(12lil%)-l35+1
16±12%)-(24il2

                                  339

-------
TABLE VIIl-35. SUMMARY OF WASTEWATER TREATABILITY RESULTS USING A LIME
             ADDITION/BARIUM CHLORIDE ADDITION/SETTLE PILOT-SCALE
             TREATMENTSCHEME
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
96 to > 99
97 to > 99
73
96-97

78-97
0-25
FINAL CONCENTRATION
ATTAINED (mg/l)
Total
-
0.11-0.18
<0.20
0.10
< 0.040
< 0.050
0.80-32
0.88-4.9
0.90-17
0.20-1.3
4.6
3.9-4.0*

0.30-2.5
123-292
Dissolved
5590-9740
0.060-0.15
< 0.014
<0.020-0.030
< 0.040
<0.040
0.10-1.0
0.43-4.2
0.50-5.0
£0.20
2.2
1.0-2.3*

0.20-0.50
-
   * Values in picocuries per liter (pc/l)
                                        340

-------
TABLE VJJJ-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
                                     341

-------
            TABLE VIII-37. SUMMARY OF TREATED EFFLUENT QUALITY ATTAINED WITH PILOT SCALE TREATMENT
                          SYSTEM AT MILL 9401
UNIT TREATMENT
PROCESS EMPLOYED
Fe (SO4) (1000 mg/l), Aeration,
Non-Ionic Polymer, BaCI2
(60 mg/l), Flocculation,
Sedimentation, Filtration
IX, Fe (S04) (330 mg/i),
Aeration, Non-Ionic Polymer,
BaCI2 (15 mg/l), Flocculation,
Sedimentation, Filtration
IX, Fe(S04) (500 mg/l).
Aeration, Non-Ionic Polymer,
BaCI2 (120 mg/l), Flocculation,
Sedimentation, Filtration
EFFLUENT CONCENTRATIONS ATTAINED (mg/l)'

pH**
10.0


9.0



8.0



TSS
1744


2.0



468



TDS
21700


24300



23200



so4-
9380


10500



9830



AS
<5.0
(d)

<5.0
(d)


<5.0
(d)



SE
9.6
(d)

8.5
(d)


8.5
(d)



Mo
90


55



55



V
8.0


9.0



5.0



U
59
±13%

6.0
±9%


0.72
±18%



Ra226T
3.1
±2%

2.5
±6%


44
±7.3%


U)
            * Metals are total metals unless otherwise indicated by a (d) = dissolved

            **pH units

            t pd/l

-------
                      TABLE Vlll-38. RESULTS OF BENCH-SCALE ACID/ALKALINE MILL WASTEWATER NEUTRALIZATION
                 ACID/ALKALINE
                MILL WASTEWATER
                  MIX RATIO
                  BY VOLUME
                                 THEORETICAL PARAMETER CONCENTRATION (mg/l)
                                 T /  D
                         Mo
                                           T  /  D
                                   Se
                                                     T  /  D
                                                               T  /  D
                                                          [(MEASURED PARAMETER CONCENTRATION (mg/ll
                                                                         T  /  O
                                                                Mo
                                                                                             T  /  D
                                                                                               pH
                                                                                                               APPARENT REMOVAL (percent)
                                                                                                                           T  /  D
                                                                                                                   Mo
                                                                                                                             Se
                                                                                                                                               T  /  D
                    500/0
                                1900
                                          10
                                                             100
                                                                       1900
                                                                                                     100
                                    1860
                                                                  97
                                                                                      8.9
                                                                                                          97
                    500/300
                                1190
                                                   8.}
                                                             72
                                                                                 1.3
                                    1160
                                              46
                                                                  71
                                                                                      0.70
                                                                                                         <0.50
                                                                                                                  3.9
                                                                                                                                    97.2
                                                                                                                              55.3
                                                                                                                                        98.5
                    500/400
                                          53
                                                             67
                                                                       480
                                                                                           12
                                    1030
                                              53
                                                                                      2.0
                                                                                                                                        96.2
                    400/500
                                850
                                          63
                                                                       400
                                                                                 8.9
                                                                                                                          52.4
                                    830
                                              64
                                                                            350
                                                                                                         
oo
300/500
                                710
                                          70
                                                             54
                                                                                 30
                700
                                                        270
                                                                  30
                                                                                                                         60.3
                                                                                                                              61.3
                    200/500
                                          79
                                                                       150
                                                                                 59
                                                                                                     2.5
                                    530
                                              80
                                                                  47
                                                                                                                  6.6
                                                                                                                         72.6
                                                                                                                                   24.9
                                                                                                                              73.7
                                                                                                                                                       94.7
                                                                                                                                                            98.9
                    150/500
                               440
                                                             43
                                                                       93
                                                                                                     6.0
                                                                  43
                                                                                                          1.0
                     0/500
                               <0.20
                                                             26
                                                                      <0.20
                                                                                 110
                                                                                                     26
                                   <0.20
                                                        19
                                                                  26
                                                                           <0.20
                                                                                                19
                t (theoretical concentration) - (meaiured concentration)

                          theoretical concentration
                                 (100)

-------
TABLE VIII-39.  CHARACTERIZATION OF INFLUENT TO WASTEWATER TREATMENT
             PLANT AT COPPER MINE/MILL/SMELTER/REFINERY 2122
             (SEPTEMBER 5-7,1979}
POLLUTANT
PARAMETER
PH
TSS
Fe
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
TI
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
-
—
                               344

-------
TABLE VIII-40. CHARACTERIZATION OF EFFLUENT FROM WASTEWATER
            TREATMENT PLANT (INFLUENT TO PILOT-SCALE TREATMENT
            PLANT) AT COPPER MINE/MILL/SMELTER/REFINERY 2122
            (SEPTEMBER 5-10,1979)                ..    ..    .
POLLUTANT
PARAMETER
PH
TSS
Fe
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
Tl
Zn
CM
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 EFFLUENT QUALITY FROM PILOT-SCALE
              TREATMENT PROCESSES AT COPPER MIIME/MILL/SMELTER/REFINERY
              2122 (SEPTEMBER 5-10, 1979)
TEST
NUMBER
01


02


03


04


05


06
07




08



09




TREATMENT APPLIED
Filtration*
0.15 m3/min/m2
(3.6 flpm/ft2)
Filtration*
0.26 m3/min/m2
(6.3 gpm/ft2)
Filtration*
0.37 m3/min/m2
(9.1 gpm/ft2)
Filtration*
0.43 m3/min/m2
(10.6 gpm/ft2)
Filtration*
0.50 m3/min/m2
(12.2 gpm/ft2)
One hour settling test
Settling -45 min
Filtration with Lime
Addition to pH 9.1
0.38 m3/min/m2
(9.4 gpm/ft2)
Filtration with Lime
Addition to pH 9.0
O 9
0.38 m /min/m
(9.3 gpm/ft2)
Filtration with Lime
Addition to pH 9.5
0.37 m3/min/m2
(9.5 gpm/ft2)
EFFLUENT CONCENTRATION (TOTAL, mg

pH*
7.2


7.6


8.1


8.1


8.3


8.5
8.95




8.95



9.05




TSS
4


3


2


2


2


8
<1




1



1




Cu
0.084


0.095


0.094


0.093


0.100


0.250
0.060




0.070



0.064




Pb
0.004


0.015


0.010


0.015


0.009


0.043
0.004




0.005



0.017




Zn
0.310


0.260


0.120


0.140


0.210


0.170
0.031




0.043



0.023



/I)

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
                                   346

-------
TABLE VIII-41.  SUMMARY OF TREATED EFFLUENT QUALITY FROM PILOT-SCALE
              TREATMENT PROCESSES AT COPPER MINE/MILL/SMELTER/REFINERY
              2122 (SEPTEMBER 5-10,1979) (Continued)

TEST
NUMBER
10



11



12



13



16





TREATMENT APPLIED
Filtration
5 mins
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration
6 hours
0.38 m /min/m
(9.3 gpm/ft2)
Filtration
12 hours
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration*
1 8 hours
0.38 m3/min/m2
(9.3 gpm/ft2)
Filtration with Lime
Addition to pH 10.0
0.37 m3/min/m2
(9.1 gpm/ft2)
EFFLUENT CONCENTRATION (TOTAL, mg/l)

pH*
„
-


8.45



8.3



8.6



9.75




TSS
1



1



<1



1



2




Cu
0.058



0.073



0.044



0.110



0.056




Pb
0.011



0.002



0.018



0.011



<0.002




Zn
0.038



0.110



0.097



0.038



0.110




Fe
0.06



0.07



0.04



0.03



0.03



 Dual media filtration (anthrafilt and silica sand)
* Field pH values
                                   347

-------
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. Phenolics
NUMBER OF
OBSERVATIONS
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
3
2
TOTAL CONCENTRATION
(mg/l)
MEAN
7.5
4.1
0.27
< 0.003
< 0.002
< 0.0005
< 0.008
< 0.008
0.023
0.004
< 0.001
0.054
< 0.005
<0.01
< 0.003
0.015
0.01
RANGE
7.4-7.65
3.9-4.2
0.25-0.29
<0.003
<0.002
<0.0005
<0.005-0.013
<0.005-0.015
0.022-0.025
0.003-0.006
< 0.001
0.053-0.055
<0.005
<0.01
<0.003
0.011-0.019
0.007-0.013
DISSOLVED CONCENTRATION
(mg/l)
MEAN
—
-
0.088
<0.003
<0.006
<0.0005
<0.008
<0.008
0.022
0.006
<0.001
0.055
<0.005
<0.01
< 0.003
0.032
-
RANGE
—
-
0.071-0.110
< 0.003
< 0.002-0.01 5
<0.0005
<0.005-0.013
<0.005-0.013
0.021-0.023
0.005-0.009
<0.001
0.053-0.057
< 0.005
<0.01
< 0.003
0.018-0.049
—
* Grab samples
••pH units
                                   348

-------
TABLE VI11-43.  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 gpmtft2)
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

r
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
                                      349

-------
TABLE VIII-44. HISTORICAL DATA SUMMARY FOR IRON ORE MINE/MILL 1108 (FINAL DISCHARGE)
PARAMETER*
pH
pH
TSS
TSS
Fe (dissolved)
Fe (dissolved)
MONITORING
PERIOD
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 -Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
3 - 5/mo
1/mo.
(mo. ave.)
4 - 31/mo.
3 - 5/mo.
(mo. ave.)
3 - 5/mo
NUMBER OF
OBSERVATIONS
35
148
35
804
35
147
MEAN
7.1
(ave. of mo. means)
7.1
(daily ave)
5.9
(ave. of mo. means)
6.1
(daily ave.)
0.36
(ave. of mo. means)
0.35
(daily ave.)
STANDARD
DEVIATION
0.2
0.3
4.8
6.8
0.47
0.58
RANGE
6.7 - 7.7
6.5 - 8.0
2-28
1.0-70
0.05-1.92
0.01 - 3.6

-------
           TABLE VIII-45.  HISTORICAL DATA SUMMARY FOR COPPER/SILVER MINE/MILL/SMELTER/REFINERY 2121
                           (FINAL TAILINGS POND DISCHARGE)
PARAMETER*
•%
Flow (m /day)
pH (Std. Units)
TSS (mg/l)**
Copper (mg/l)
Zinc (mg/l)
Chlorides (mg/l)
MONITORING
PERIOD
March 1975 - Dec. 1979
March 1975 - Dec. 1979
March 1975 - Dec. 1979
March 1975 • Dec. 1979
March 1975 • Dec. 1979
March 1975 • Dec. 1979
FREQUENCY OF
OBSERVATION
Continuous
I/month™
1/rnonthn
1 /month
1 /month
1 /month
NUMBER OF
OBSERVATIONS
_
58
58
59
59
57
MEAN»»*
79.107
7.8
5.5
<0.023
<0.026
866
STANDARD
DEVIATION
36,714
0.4
2.9
0.015
0.019
262
RANGE1"
21,953 • 168,400
7.2-9.0
2-29
0.003 • 0.065
<0.002 - 0.09
354-1,494
CO
yi
            *  All metals expressed as total metals unless otherwise specified.
            ** TSS values measured by turbidity and converted to mg/l.
            t  Range of concentrations based on maximum and minimum data between March 1975 and April 1977, and average concentrations between
               May 1977 and December 1979.
            tt Monthly averages based on almost continuous daily monitoring.
            ***Mean of monthly averages.

-------
      TABLE VIII-46. HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL 2120 (TAILING-POND
                   OVERFLOW: TREATMENT OF UNDERGROUND MINE, MILL, AND LEACH-
                   CIRCUIT WASTEWATER STREAMS)
PARAMETER*
PH
pH
TSS
TSS
Cu
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Sept. 75 - June 77
Sept 75 - June 77
Sept. 75 - June 77
Sept 75 - June 77
Sept 75 - June 77
Sept 75 - June 77
Sept 75 - June 77
Sept 75 - June 77
Sept. 75 - June 77
Sept 75 - June 77
FREQUENCY OF
OBSERVATION
15 - 31/mo.
1/mo.
(mo. ave.)
15 -31/mo.
1/mo.
(mo. ave.)
15 -31/mo.
1/mo.
(mo. ave.)
1 - 5/mo.
1/mo.
(mo. ave.)
15 -31/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
538
22
536
22
536
22
82
22
535
22
MEAN
(mg/l)
9.53
(daily ave.)
9.23
(ave. of mo. means)
9.6
(daily ave.)
10
(ave. of mo. means)
0.065
(daily ave.)
0.07
(ave. of mo. means)
<0.01
(daily ave.)
<0.01
(ave. of mo. means)
0.177
(daily ave.)
0.23
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.02
1.07
10.6
3
0.081
0.05
0
0
0.556
0.43
RANGE
(mg/l)
6.4-12.0
7.2-11.0
1-132
5-19
0.01 - 0.88
0.02 - 0.27
none
none
0.01 - 7.1
0.02-1.71
*AII metals expressed as total metals unless otherwise specified.

-------
                 TABLE VIII-47. HISTORICAL DATA SUMMARY FOR COPPER MINE/MILL 2120 (TREATMENT

                              SYSTEM - BARREL POND - EFFLUENT: TREATMENT OF MILL WATER
                              AND OPEN-PIT MINE WATER
PARAMETER*
pH
pH
TSS
TSS
Cu
Cu
Pb
Hg
Zh
Zn
MONITORING
PERIOD
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
Apr. 75 -Sept. 77
Jan. 75 - Sept. 77
Jan. 75 - Sept. 77
FREQUENCY OF
OBSERVATION
25-31 /mo.
1/mo.
25-31 /mo.
1/mo.
25-31/mo.
1/mo.
3-5/mo.
3-5/mo.
25-31/mo.
1/mo.
NUMBER OF
OBSERVATIONS
953
33
954
33
957
33
33
30
957
33
MEAN
(mg/l)
10.8
(daily ave.)
10.7
(ave. of mo. means)
12
(daily ave.)
12
(ave. of mo. means)
0.09
(daily ave.)
0.10
(ave. of mo. means)
0.015
(ave. of mo. means)
0.0003
(ave. of mo. means)
0.14
(daily ave.)
0.15
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.1
0.6
10
4
0.09
0.05
0.012
0.0001
0.16
0.09
RANGE
(mg/l)
2.9 - 13.1
9.7-12.4
0-120
7-22
< 0.01 - 0.98
0.04 - 0.27
< 0.01 - 0.06
< 0.00005 - 0.0006
< 0.01 - 2.00
0.04 - 0.36
oo
yi
CO
          *AII metals expressed as total metals unless otherwise specified.

-------
                   TABLE VIII-48.  HISTORICAL DATA SUMMARY FOR LEAD MINE/MILL 3105 (TREATED
                                  MINE EFFLUENT)
PARAMETER*
pH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
Jan. 74 - Jan. 78
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
49 .
43
49
49
49
49
MEAN
(mg/l)
8.1
(ave. of mo. means)
3.0
(ave. of mo. means)
<0.02
(ave. of mo. means)
0.006
(ave. of mo. means)
0.043
(ave. of mo. means)
0.026
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
0.23
2.0
0.0
0.001
0.023
0.022
RANGE
(mg/l)
7.4 - 8.5
1 -9
na
< 0.005 -0.010
0.01 -0.12
0.005-0.11
CO
           *AII metals expressed as total metals unless otherwise specified.
           na = not applicable

-------
                           TABLE VI11-49. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE 3130
                                         (UNTREATED MIIMEWATER)
PARAMETER*
pH
TSS
TSS
CN
CN
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Aug. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
July 77 - Oct. 77
Aug. 77 - Oct. 77
FREQUENCY OF
OBSERVATION
unk
6/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.)
6/mo.
1/mo.
(mo. ave.)
6/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
unk
18
3
3
3
20
3
20
3
MEAN
(mg/l)
7.7
42.6
44.2
(ave. of mo. means)
0.11
0.11
(ave. of mo. means)
0.93
1.00
(ave. of mo. means)
1.25
1.35
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
n.a.
20.2
na
na
na
0.357
na
0.52
na
RANGE
(mg/l)
7.2 - 8.3
5.6-90.2
33.2 - 61.9
0.04 - 0.16
0.04 - 0.16
0.30-1.60
0.70-1.40
0.45 - 2.50
1.09-1.88
CO
U1
Jl
            *AII metals expressed as total metals unless otherwise specified.
            unk = unknown
            na = not applicable

-------
                TABLE VIII-50. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE 3130 (TREATED EFFLUENT)
CO
01
PARAMETER*
pH
pH
TSS
TSS
Cu
Cu
CN
CN
Pb
Pb
Zn
Zn
MONITORING
PERIOD
June 77 - Sept. 77
June 77 - Sept. 77
Aug. 77 - Oct. 77
Aug. 77 - Oct 77
Aug. 77 - Oct. 77
Aug. 77 - Oct. 77
June 77 - Oct 77
June 77 - Oct. 77
July 77 - Oct 77
Aug. 77 - Oct. 77
July 77 - Oct 77
Aug. 77 - Oct. 77
FREQUENCY OF
OBSERVATION
unk
unk
8/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.)
1/mo.
1/mo.
(mo. ave.)
8/mo
1/mo.
(mo. ave.)
8/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
unk
4
23
3
3
3
5
5
. 25
3
25
3
MEAN
(mg/l)
8.8
8.8
(ave. of mo. means)
2.12
2.1
(ave. of mo. means)
0.12
0.12
(ave. of mo. means)
0.064
0.064
(ave. of mo. means)
0.079
0.08
(ave. of mo. means)
0.32
0.33
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
na
na
1.34
na
na
na
na
na
0.021
na
0.085
na
RANGE
(mg/l)
8.5 - 8.9
8.6 - 8.9
<0,1 -6.1
1.8 - 2.4
0.05 - 0.25
0.05 - 0.25
0.022-0.150
0.022-0.150
< 0.05 -0.1 3
0.067 - 0.097
0.18 - 0.57
0.24-^0.39 ,
            *AII metals expressed as total metals unless otherwise specified.
            unk = unknown

            na = not applicable

-------
                   TABLE VI11-51. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3101 (TAILING-
                                 POND DECANT TO POLISHING PONDS)
PARAMETER*
pH
PH
Cu|
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
FREQUENCY OF
OBSERVATION
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
197
44
136
44
196
44
197
44
MEAN
(mg/l)
9.3
(daily ave.)
9.2
(ave. of mo. means)
0.054
(daily ave.)
0.055
(ave. of mo. means)
0.025
(daily ave.)
0.040
(ave. of mo. means)
0.193
(daily ave.)
0.294
(ave. of mo. means)
STANDARD
DEVIATION
(mg/1)
1.5
1.3
0.072
0.060
0.049
0.049
0.264
0.401
RANGE
(mg/l)
7.0-11.3
6.7-11.2
0.002 • 0.490
0.008 - 0.323
0.004-0.525
0.004 - 0.303
0.010 • 2.20
0.038 -2.15
tn
          *AII metals expressed as total metals unless otherwise specified.
          na = not applicable

-------
                TABLE VIII-52. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3101 {FINAL

                              DISCHARGE FROM POLISHING POND)
PARAMETER*
pH
pH
TSS
Cu
Cu
Pb
Pb
Zn
Zn
MONITORING
PERIOD
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
Oct. 75 - Aug. 77
Jan. 74 - Aug. 77
FREQUENCY OF
OBSERVATION
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
4 - 5/mo.
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
101
44
14
100
44
98
44
101
44
MEAN
(mg/l)
8.1
(daily ave.)
8.3
(ave. of mo. means)
8.4
(daily ave.)
0.024
(daily ave.)
0.020
(ave. of mo. means)
0.009
(daily ave.)
0.020
(ave. of mo. means)
0.211
(daily ave.)
0.176
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
1.3
1.3
na
0.027
0.017
0.008
0.018
0.123
0.096
RANGE
(mg/l)
6.6-11.2
6.9 - 10.9
0.3 - 26.0
0.002-0.124
0.006 - 0.076
0.004 - 0.052
0.005 - 0.082
0.026 - 0.548
0.028 - 0.390
CO
tn
oo
         *AII metals expressed as total metals unless otherwise specified.

         na = not applicable

-------
            TABLE VIII-53. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3102 (FINAL DISCHARGE)
PARAMETER*
PH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
Dec. 73 - Sept. 74
.7
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
1/mo.
(mo. ave.)
NUMBER OF
OBSERVATIONS
10
10
10
10
10
10
MEAN
(mg/l)
7.9
(ave. of mo. means)
2.6
(ave. of mo. means)
<0.02
(ave. of mo. means)
0.001
(ave. of mo. means)
0.002
(ave. of mo. means)
0.01
(ave. of mo. means)
STANDARD
DEVIATION
(mg/l)
0.3
3.8
0
0.001
0.002
0.02
RANGE
(mg/l)
7.6 - 8.4
0-10
na
< 0.001 - 0.003
< 0.001 - 0.007
< 0.001 - 0.07
CO
in
-o
        * All metals expressed as total metals unless otherwise specified.

        na = not applicable

-------
               TABLE VIII-54. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3103 (TAILING-POND
                              EFFLUENT TO SECOND SETTLING POND)
CO
0}
PARAMETER*
PH
TSS
Cu
Pb
Zn
MONITORING
PERIODf
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
32**
35
40
40
40
MEAN
(mg/l)
7.84
<1.63
0.023
0.154
0.309
STANDARD
DEVIATION
(mg/l)
0.26
0.8
0.013
0.0685
0.174
RANGE
(mg/l)
7.4 - 8.5
<1 .0-4.0
0.005 - 0.058
0.038 - 0.33
0.030 - 0.79
             *AII metals expressed as total metals unless otherwise specified.
             tThe following months were excluded due to missing data: Oct. 1975, June 1975,
              Sept. 1975, Oct. 1976, and Jan. 1977.
            **Several illegible data points excluded.

-------
                 TABLE VIII-55.  HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3103 (EFFLUENT
                                 FROM SECOND SETTLING POND)
PARAMETER*
pH
ISS
Cu
Pb
Zn
MONITORING
PERIOD *
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
Feb. 74 - Nov. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
33**
38
39
40
40
MEAN
(mg/l)
7.93
< 1.55
0.013
0.058
0.110
STANDARD
DEVIATION
(mg/l)
0.22
1.48
0.007
0.028
0.086
RANGE
(mg/l)
7.5 - 8.4
1.0-8.0
0.002-0.034
0.01-0.122
0.018-0.440
00
          *AII metals expressed as total metals unless otherwise specified.

          tThe fo"owing months were excluded du^°L_missiin9 data: Oct- 1974' June 1975, July 1975, Sept. 1975, Oct. 1975, and Jan. 1977
         **Several illegible data points excluded.

-------
                   TABLE VIII-56. HISTORICAL DATA SUMMARY FOR LEAD/ZINC MINE/MILL 3104 (TAILING-
                                 LAGOON OVERFLOW)
PARAMETER*
pH
TSS
CN
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
Apr. 77 - Dec. 77
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
Jan. 74 - Dec. 77
FREQUENCY OF
OBSERVATION
4-5/mo. (mo. ave.)
4-5/mo. (mo. ave.)
1/mo.
4-5/mo. (mo. ave.)
4-5/mo. (mo. ave.)
4-5/mo. (mo. ave.)
NUMBER OF
OBSERVATIONS
48
48
9
48
48
48
MEAN
(mg/l)
7.8
6.8
<0.1
0.06
0.07
0.13
STANDARD
DEVIATION
(mq/l)
0.8
2.9
0
0.04
0.02
0.07
RANGE
(mg/l)

2.9-16
na
0.01 - 0.14
0.03 - 0.12
0.03 - 0.42
00
Jl
iNJ
           *AII metals expressed as total metals unless otherwise specified.

          na = not applicable

-------
                         TABLE VIII-57.  HISTORICAL DATA SUMMARY FOR LEAD/ZING MILL 3110
                                       (TAILING-LAGOON OVERFLOW)
PARAMETER*
pH
TSS
Cu
Pb
Zn
MONITORING
PERIOD
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
Jan. 74 - Apr. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
40
40
40
40
40
MEAN
(mg/l)
7.2
5.5
0.02
0.05
0.20
STANDARD
DEVIATION
(mg/l)
0.4
5.1
0.02
0.02
0.18
RANGE
(mg/l)
6.3 - 8.3
0.4 - 20
0.001 -0.111
0.001 -0.106
0.01 - 0.35
CO
0}
         *AII metals expressed as total metals unless otherwise specified.

-------
             TABLE VIII-58.  HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE 6103 (TREATED EFFLUENT)
PARAMETER*
pH
TSS
As
Cd
Cu
Pb
Mo
Zn
MONITORING
PERIOD
July 76 - June 77
July 76 - June 77
Apr. 77 - June 77
July 76 - June 77
July 76 - June 77
Apr. 77 - June 77
July 76 • June 77
July 76 - June 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
12
12
3
12
12
3
12
12
MEAN
(mg/l)
8.5
8.3
<0.01
<0.01
<0.05
<0.01
0.50
0.24
STANDARD
DEVIATION
(mg/l)
0.24
5.9
—
—
—
-
0.21
0.069
RANGE
(mg/l)
8.1 - 8.9
< 5-25
none
none
none
none
0.21-1.0
0.15-0.4
GO
Jt
-P"
          *AII metals expressed as total metals unless otherwise specified.
          na = not applicable

-------
              TABLE Vlll-59. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE 6101 (TREATED EFFLUENT)
PARAMETER*
pH
TSS
COD
Cd
Cu
CN
Mo
Se
Zn
MONITORING
PERIOD
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
1972
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
1972
Jan. 75 - Dec. 76
FREQUENCY OF
OBSERVATION
0-5/mo.
0-5/mo.
0-5/mo
0-5/mo
0-5/mo.
0-5/mo.
0-5/mo
0-5/mo.
0-5/mo.
NUMBER OF
OBSERVATIONS
47
50
33
44
4
51
43
4
48
MEAN
(mg/l)
7.57
(daily ave.)
7.6
(daily ave.)
28.5
(daily ave.)
<0.02
(daily ave.)
< 0.02
(daily ave.)
<0.02
(daily ave.)
1.84
(daily ave.)
< 0.005
(daily ave.)
0.077
(daily ave.)
STANDARD
DEVIATION
0.38
7.5
12
na
0.01
0.021
0.38
na
0.18
RANGE
(mg/l)
6.5 -5.0
1 -34
8-52
< 0.01 - < 0.02
< 0.02 - 0.03
< 0.02 - 0.083
1.1 -2.9
none
< 0.01 - 0.90
co
o>
01
         *AII metals expressed as total metals unless otherwise specified.

         na = not applicable

-------
              TABLE VIII-60. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/MILL 6102 (CLEAR
                           POND BLEED STREAM - INFLUENT TO TREATMENT SYSTEM)
•00
PARAMETER*
TSS
Cd
Cu
Fe
Mn
Mo
Pb
Zn
Cyanide
MONITORING
PERIOD
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
FREQUENCY OF
OBSERVATION
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
NUMBER OF
OBSERVATIONS
33
34
35
33
35
35
35
35
35
MEAN
(mg/l)
65.4
0.02
0.055
2.2
7.3
5.6
0.01
0.84
0.06
STANDARD
DEVIATION
(mg/l)
180
0.022
0.016
2.4
1.2
1.3
0.002
0.68
0.06
RANGE
(mg/l)
10 - 1070
<0.01-0.15
<0.05-0.10
0.60 - 13.0
5.4-12.2
3.0 - 8.3
<0.01 - 0.02
0.20 - 2.7
<0.01 - 0.2
          *AII metals expressed as total metals unless otherwise specified

-------
                TABLE VIII-61. HISTORICAL DATA SUMMARY FOR MOLYBDENUM MINE/MILL 6102 (FINAL
                             DISCHARGE FROM RETENTION POND - TREATED EFFLUENT)
co
PARAMETER*
TSS
Cd
Cu
Fe
Mn
Mo
Pb
Zn
Cyanide
MONITORING
PERIOD
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
July -Oct. 1978
July - Oct. 1978
July - Oct. 1978
July - Oct. 1978
FREQUENCY OF
OBSERVATION
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
1-5/week
NUMBER OF
OBSERVATIONS
36
37
37
37
36
37
37
37
37
MEAN
(mg/l)
9
0.01
<0.05
0.4
0.26
2.5
<0.01
0.29
<0.01
STANDARD
DEVIATION
(mg/l>
11.4
0.006
0
0.32
0.17
1.8
0
0.16
0
RANGE
(mg/l)
<5-60
<0.01 - 0.03
no range
0.15-2.0
0.1 -0.7
0.5 - 7.0
no range
<0.05 - 0.8
no range
             * All metals expressed as total metals unless otherwise specified.

-------
                        TABLE VIII-62. HISTORICAL DATA SUMMARY; BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 008;

                                        JANUARY 1979- DECEMBER 1980
00
en
oo
PARAMETER**
Flow (m3/day)»
pH (Std. Units)
TSS(mg/l)tt
Iron (ing/I)™
Aluminum Img/Oft
MONITORING PERIOD
============
Jan. 1979 - Dec. 1980
Jan. 1979 -Dec. 1980
Jan. 1979 - Dec. 1980
Jan. 1979 - Dec. 1980
Jan. 1979 - Dae. 1980
FREQUENCY OF
OBSERVATIONS
	 : 	 — 	
Continuous
1/week
1/week
1/week
1/week
NUMBER OF
OBSERVATIONS"*
-
21
21
21
21
MEAN*
10,860
-
3.7
0.52
1.04
STANDARD
DEVIATION
7.470
-
2.5
0.46
0.50
RANGE
0-38,153
4.5 - 8.8
OS • 13.3
0.08 - 2.61
0.22 - 4.02
*  Mean flow and standard deviation include zero discharge months. No discharge during July, August and September 1980.

** All metals expressed as total metals unless otherwise specified.

»**0nly one sample collected during September 1979, December 1979, June 1980 and October 1980.

t  Mean of monthly averages.
tt Mean, standard deviation and range for discharging months only. Number rounded to significant figures.

-------
                      TABLE VIII-63. HISTORICAL DATA SUMMARY; BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 009-
                                     FEBRUARY 1979-DECEMBER 1980
co
o>
•o
PARAMETER*
Flow (m3/day)
pH (Std. Units)
TSS(mg/l)tt
Iron (mg/ljtt
Aluminum (mg/\f*
MONITORING PERIOD
Feb. 1979 - Dec. 1980
Feb. 1979 -Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 - Dec. 1980
Feb. 1979 -Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
I/week
1/week
1/week
1/week
NUMBER OF
OBSERVATIONS**
—
21
21
21
21
MEAN*
14,120
-
2.2
0.19
0.67
STANDARD
DEVIATION
8,930
-
0.92
0.08
0.37
RANGE
5,791 - 43,868
6.2 - 8.8
0.5-11.2
0.02 - 0.6
0.12 - 2.25
* All metals expressed as total metals unless otherwise specified
** No data for April 1980 and July 1980
t Mean of monthly averages
tt Numbers rounded to significant figures

-------
                        TABLE Vlll-64. HISTORICAL DATA SUMMARY: BAUXITE MINE 5102; MINE DRAINAGE; DISCHARGE 010;
                                        FEBRUARY 1979 - DECEMBER 1980
CO
PARAMETER*
Flow (m3/day)tt
pH (Std. Units)
TSS (mg/l)***
Iron (mg/l***
Aluminum 
-------
                  TABLE VIII-65. HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
                                AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT- DISCHARGE 001-
                                JUNE 1978 - DECEMBER 1980
CO
PARAMETER*
Flow (m3/day)tt
TSS(mg/l)tt
pH (Std. Units)
Aluminum (mg/l)tt
Iron (mg/l)tt
MONITORING PERIOD
June 1978 -Dec. 1980
June 1978 • Dec. 1980
June 1978 - Dec. 1980
Oct. 1978 - Dec. 1980
Oct 1978 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
1/week
1-3/month
1-3/ month ' •
NUMBER OF
OBSERVATIONS
_
31
31
27
27
MEAN*
6,540
5.2
-
<0.36
<0.23
STANDARD
DEVIATION
2,900
2.8
-
0.32
0.36
RANGE
0-12,536
0-30.0
5.2 - 8.7
<0.1 - 1.50
< 0.01 -5.2
                  •All metals expressed as total metals unless otherwise specified.
                  t Mean of monthly averages.        ...-
                  ttNumbers rounded to significant figures.

-------
                  TABLE Vlll-66  HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
                               ' AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT; DISCHARGE 007;
                                 JANUARY 1978 - DECEMBER 1980
00
-J
ro
PARAMETER**
Flow (m3/day)»
TSS (mg/l)t
pH (Std. Units)
Aluminum (nrg/l)*
Iron (mg/l)1"
MONITORING PERIOD
Jan. 1978 - Dec. 1980
Jan. 1978 - Dec. 1980
Jan. 1978 - Dec. 1980
Jan. 1978 - Dec. 1980
Jan. 1978 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1-4/month
1/week
1-4/month
1-4/month
NUMBER OF
OBSERVATIONS
-
16
16
9
9
MEANtt
370**
6.8
-
<0.15
<0.76
STANDARD
DEVIATION
534**
5.0
—
0.06
0.60
m
RANGE
0-4,360
0-32
6.0-8.9
< 0.1 -0.29
< 0.01 -2.6
                   * Mean flow and standard deviation include zero discharge months.
                   "There were 20 months during the monitoring period in which no discharge occurred.
                   t Mean, standard deviation and range for discharging months only. Numbers rounded to significant figures.
                   ttMean of monthly averages.

-------
                  TABLE'VNI-67.  HISTORICAL DATA SUMMARY; BAUXITE MINE 5101; FINAL DISCHARGE FROM MINE
                                  AREA RUNOFF AND MINE PIT PUMPAGE - TREATED EFFLUENT; DISCHARGE 009-
                                  JANUARY 1980 - SEPTEMBER 1980
co
PARAMETER**
Flow (m3/day)*
TSS(mg/l)tt
pH (Std. Units)
Aluminum (ing/I)"
Iron (mg/l)tt
MONITORING PERIOD
Jan. 1980 - Sept. 1980
Jan. 1980 -Sept. 1980
Jan. 1980 -Sept. 1980
Jan. 1980 - Sept. 1980
Jan. 1980 - Sept. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
3-4/month
1/week
1-2/month
1-2/month
NUMBER OF
OBSERVATIONS
_
7
7
7
7
MEAN*
1,060
5.9
—
<0.27
0.27
STANDARD
DEVIATION
1,050
2.6
—
0.14
0.14
RANGE
0-8,176
0-26
6.1 - 8.2
<0.1 - 0.64
<0.05 - 0.53
                 * Mean flow and standard deviation include zero discharge months. No discharge occurred during January or September 1980.
                 **AII metals expressed as total metals unless otherwise specified.
                 t Mean of monthly averages.
                 ttMean, standard deviation and range for discharging months only. Numbers rounded to significant figures.

-------
               TABLE VIII-68. HISTORICAL DATA SUMMARY FOR TUNGSTEN MINE 6104 (TREATED EFFLUENT)
PARAMETER*
pH
TSS
Cu
Mo
Zn
MONITORING
PERIOD
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
Jan. 75 - Dec. 76
FREQUENCY OF
OBSERVATION
1/mo.
(mo. ave.)
1-4/mo.
(mo. ave.)
1/mo.
1/mo.
1/mo.
NUMBER OF
OBSERVATIONS
35
24
24
24
24
MEAN
(mg/l)
7.98
(ave. of mo. means)
10.8
(ave. of mo. means)
< 0.02
0.074
< 0.01
STANDARD
DEVIATION
(mg/l)
0.27
4.4
na
0.17
na
RANGE
(mg/l)
7.3 - 8.6
4-20
none
< 0.02 -0.74
none
CO
-•4
         *AII metals expressed as total metals unless otherwise specified.
         na = not applicable

-------
              TABLE Vlll-69. HISTORICAL DATA SUMMARY FOR URANIUM MINE 9408 (TREATED MINE WATER)
PARAMETER*
pH
TSS
Ra226
(dissolved)
U238
Ba
Mo
Pb
Zn
MONITORING
PERIOD
Apr. 75 - Jan. 77
Apr. 75 • Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
Apr. 75 -Jan. 77
Apr. 75 - Jan. 77
Apr. 75 - Jan. 77
FREQUENCY OF
OBSERVATION
1/mo.
1/mo.
1/mo.
1/mo.
1/mo.
4/yr.
1/mo.
4/yr.
NUMBER OF
OBSERVATIONS
21
18
19
21
21
7
19
7 -
MEAN
(mg/l)
8.3
6
0.55
2.0
0.52
0.65
0.06
0.02
STANDARD
DEVIATION
(mg/l)
0.3
4
0.45
1.3
0.38
0.53
0.11
0.02
RANGE
(mg/l)
7.8 - 8.7
1 -10
0-1.2
0.2-4.2
0.06-1.6
0.05 - 1.3
0.01 - 0.50
0.008 - 0.08
00
^4
(ft
         * All metals expressed as total metals unless otherwise specified.

-------
                  TABLE VIII-70.   HISTORICAL DATA SUMMARY FOR  NICKEL ORE MINE/MILL 6106 - TREATED
                                    EFFLUENT; DISCHARGE 001; JANUARY 1976 - DECEMBER 1980
CO
PARAMETER*
Flow (m^/day)*
pH (Std. Units)***
TSS(mg/l)***
Chromium (mg/l)***
Manganese (mg/l) ***
	 T**-
Phosphorous (P) (mg/l)
»#«
Ssttleable Solids (mg/l)
MONITORING PERIOD
Jan. 1976 - Dec. 1980
Jan. 1976 -Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
Jan. 1976 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Daily
1/week
1/week
1/month
1/month
I/month
1/week
NUMBER OF
OBSERVATIONS
59
31
32
32
32
31
30
MEAN™
3,520**
8.5
18.6
0.034
0.023
0.05
<0.05
STANDARD
DEVIATION
5,560**
0.30
15.4
0.017
0.016
0.017
0
RANGE
0-67,700
8.1-9.59
0.4-138
Not available
Not available
Not available
<0.05 - 0.05
                    * Mean flow and standard deviation include zero discharge months.
                    ••There are 28 months during the monitoring period in which no discharge occurred.
                    t All metals expressed as total metals unless otherwise specified.
                    ttMean of monthly averages.
                    •••Mean, standard deviation and range for discharging months only. Number rounded to significant figures.

-------
                       TABLE VIII-71.  HISTORICAL DATA SUMMARY; VANADIUM MINE 6107; MINE AREA RUNOFF, WASTE
                                       PILE RUNOFF; DISCHARGE 005; JULY 1978- DECEMBER 1980
PARAMETER*
Flow (m3/day)«*
TSS(mg/l)*»»
Iron (mg/l)*»«
pH (Std. Units)
MONITORING PERIOD
July 1978 - Dec. 1980
July 1978 - Dec. 1980
July 1978 - Dec. 1980
July 1980 - Dec. 1980
FREQUENCY OF
OBSERVATIONS
Continuous
1/week
1/week
5/week
NUMBER OF
OBSERVATIONS
31
27t
27t
28
MEAN™
6,140
4
0.65
-
STANDARD
DEVIATION
5,760
3
0.23
-
RANGE
0-51,098
<1-12
0.15 - 2.9
5.4 - 9.3
CO
*  All metals expressed as total metals unless otherwise specified.
** Mean flow and standard deviation include zero discharge months. No discharge during September 1979, July 1980 and September 1980.
***Mean, standard deviation and range for discharging months only. Numbers rounded to significant figures.
t  No values recorded for April 1980.
tt Mean of monthly averages.

-------
                   TABLE VI11-72. HISTORICAL DATA SUMMARY FOR TITANIUM MINE/MILL 9906;
                                 TREATED MINE/MILL WATER; OCTOBER 1975 - DECEMBER 1979*
PARAMETER
Flow (m3/day)
pH (Std. Units)
TSS (mg/l)
Settleabla Solids (mg/l)
MONITORING PERIOD
Oct. 1975 -Dec. 1979
Oct. 1975 • Dec. 1979
Oct. 1977 • Dec. 1979
Oct. 1979 - Doc. 1979
NUMBER OF
OBSERVATIONS
54
54
27
3
MEAN"
25,927
7.0
8.1
<0.05
STANDARD
DEVIATION
9,176
-
1.1
-
RANGE
1,060-68,130
4.0 - 10.0
2.4-22
-
00
                   * Summarized from Mine/Mill 9906 Discharge Monitoring Reports from October 1975 - December 1979
                   "Mean of monthly averages

-------
TABLE Vlll-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
I
POLLUTANT
PARAMETER
pH
TSS
Sb
As
Be
Cd
Cr
Cu
Pb
Hg
Ni
Se
Ag
T1
Zn
TOC
phenol
NUMBER OF
OBSERVATIONS
26
27
23











1












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
<0.02
0.015-1.9
8-9
0.028-0.036
"DISSOLVED" CONCENTRATION
(mg/l)
MEAN
' X
--
-- -
. -
—
--
0.012
— —
0.027
0.068
-- -,. • •

—
--
	 .
0.007
--
--
RANGE
--

--
	
__
0.010-0.015
__
0.020-0.035
0.050-0.090
--

--
--
_ _
<0. 005-0. 010
--
--
  Based on observations made in period 6 through 14 September 1978.
                                     379

-------
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
CM
Total
Phcnolics
PERIOD OF
OBSERVATIONS
(DATES)
Jan. 8-19, '79


















	
NUMBER OF
OBSERVATIONS
16
10
—
3
__
—
3
3
3
—
3
3
3
—
3
3
5
13
9
14
"TOTAL" CONCENTRATION
(mg/l)
MEAN
X
-
8.8
213
--
0.03
—
--
0.07
0.44
0.11
—
0.06
0.025
0.04
—
0.03
23
39
17
0.03
0.58
RANGE
8.5-8.9,
25-1200
—
0.006-0.08
--

0.04-0.13
0.04-1.24
0.06-0.19
--
0.02-0.14
0.02-0.03
<0. 03-0. 08
—
0.01-0.08.
0.42-68
32-52
6-24
0.003-0.060
0.23-0.81
"DISSOLVED" CONCENTRATIO
(mg/l)
MEAN
X
—
--
•
--
--
--
<0.02
0.08
0.06
--
0.02
—
0.03
—
0.01
0.04
—
—
—
—
RANGE
—
--
--
—
--
--
<0.02
0.02-0.21
0.05-0.08

0.02
--
0.03
__
<0.01-0.03
0.02-0.07
--
—
--
—
                                     3RD

-------
TABLE VIII-75.  SUMMARY OF PILOT-SCALE TREATABILITY STUDIES PERFORMED AT
                MILL 2122 DURING PERIOD OF 6-14 SEPTEMBER 1978
UN IT 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**
(0)
52
1
7.5**
(Cl 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**
(C0.02 to
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

-------
TABLE VIII-76. PERFORMANCE OF A DUAL MEDIA FILTER WITH TIME-FILTRATION

             OF TAILING POND DECANT AT MILL 2122
                                  3  2
Hydraulic Loading on Filter = 762 m /m /day (13 gpm)

Initial TSS concentration = 33 mg/1
          Time Elapsed
        t  + 15 min
         o


        t  + 2 hr 15 min
         o


        t  + 4 hr 15 min
         o


        t  + 7 hr 15 min
         o


        t  + 10 hr 15 min
         o
Final TSS  Concentration

.	mg/1   	



          12



           7



          11



          23



          31
                                     382

-------
   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
  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
             PH
  Contact Time  (min)

                0
               15
               30
               60
               120

NaOCl Dose 50 mg/1
        ~   pH~
  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
9
CN"


Total
Phenolics

•atf&P

•^



10
CN"
0.750
0.012
0.003
0.001
0.001
Total
Phenolics
0.808
0.680
0.396
-
- -
11
CN"
1.09
0.039
0.012
0.011
0.011
Total
Phenolics
0.608
0.452
0.696
_
-

CN"
0.88
0.05
0.02
0.01
0.004
9
Total
Phenolics
0.336
0.228
0.088
0.216
1
CN"
1.2
0.001
0.008
0.003
< 0.001
0
Total
Phenolics
0.720
O.OS8
0.084
0.080
1
CN"
+«
i
Total
Phenolics

All cyanide  and phenol concentrations are mg/1.
                                 383

-------
                       TABLE VIM-78. RESULTS OF PILOT-SCALE OZONATION FOR DESTRUCTION OF
                                   PHENOL AND CYANIDE IN MILL 2122 TAILING POND DECANT*
CO
CO
CL Dosage
mg/min
2.4
0.8
6
2
1
24
8
4
2
Ratio 0,/CN
2:1
2:1
5:1
5:1
5:1
10:1
10:1
10:1
10:1
Initial pH
10
9
10
, 11
10
10
9
10
8.5
Retention Time
10
30
10
30
60
5
15
30
60
Final Concentration (mg/1)
CN
0.610
0.460
0.035
0.040
0.016
-
0.020
0.065
0.035
Total Phenolics
0.416
0.256
0.312
0.428
0.272
0.052
0.026
0.200
0.474
             Initial Phenol Cone. = 0.232-0.808 mg/1; Cyanide  added as NaCN to provide an initial  cone.
             of  1 mg CN/1 of wastewater.

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               TABLE VIII-79. EFFLUENT QUALITY ATTAINED AT SEVERAL PLACER MINING OPERATIONS
                           EMPLOYING SETTLING-POIMD TECHNOLOGY
MINE
4142
4141
4114
4140
4139
4136
4135
4126
4127
4132
4133
4134
DATE
SAMPLED
July 14, 1977
July 20, 1977
July 14, 1977
July 17, 1977
July 12, 1977
Aug. 26, 1978
Aug. 24, 1978
Aug. 15, 1978
Aug. 15, 1978
Aug. 18, 1978
Aug. 19, 1978
Aug. 21, 1978
INFLUENT
pH*
-
- •
-
7.3
7.4
-
-
6.5
6.7
6.6
7.9
-
TSS
(mg/l)
-
-
24,000
1,130
9,000
64,100
2,890
14,800
39,900
1,540
2,260
• -
Settleable
Solids
(ml/l/hr)
1.5
17
1.8
1.7
13
45
7.5
260
550
1.6-2.0
0.7 - 1.6
1.6
As
(mg/l)
-
- .
- •
- •
1.2
3.9
0.04
1.3
5.0
0.05
1.5
-
Hg
(mg/l)
-
-
-
-
0.004
0.001
0.020
0.0002
0.0014
<0.0002
0.0002
-
EFFLUENT
pH*
—
-
-
8.5
7.4
-
-
6.8
6.4
6.5 .
7.7
-
TSS
(mg/l)
2080
120
<0.1
220
230
150
474
76
5,700
1040
170
1420
Settleable
Solids
(ml/l/hr)
0.3
<0.1
<0.1
<0.1
0.15
<0.1
0.7 - 0.9
<0.1
2.5
0.4 - 0.8
<0.1
0.4
As
(mg/l)
0.27
0.031
-
0.057
0.012
<0.002
0.022
0.25
1.2
0.05
0.06
0.28
Hg
(mg/l)
< 0.0002
<0.0002
-
<0.0002
< 0.0002
< 0.0002
< 0.0002,
0.0002
0.0005
<0.0002
0.0002
<0.0002
CO
oo
         *Value in pH units

-------
                       Figure Vlll-1. POTABLE WATER TREATMENT FOR ASBESTOS REMOVAL AT
                                  LAKEWOOD PLANT, DULUTH, MINNESOTA
                  CHEMICAL ADDITION
                 (ALUM, CAUSTIC SODA,
                     POLYMERS)
 LAKE WATER
     I
                        MIX
                       TANKS
                              FLOCCULANTS
FLOCCULATION
                                                I
                                           SEDIMENTATION
CO
00
     I
                                            MULTI-MEDIA
                                             FILTRATION
                                                         BACKWASH
                                                           WATER
                         STORAGE
                          BASIN
                                          POTABLE WATER TO
                                            STORAGE AND
                                            DISTRIBUTION
                                    SLUDGE TO
                                    SANITARY
                                    LANDFILL

-------
Figure Vlll-2. EXPERIMENTAL MINE-DRAINAGE TREATMENT SYSTEM FOR
           UNIVERSITY OF DENVER STUDY
                         MINE
                       DRAINAGE
              LIME'
                    CONTACT TANK
                    (NEUTRALIZATION
                        STAGE)
                     CONTACT TANK
                  (SULFIDE-TREATMENT
                        STAGE)
                          I
 SULFIDE
SOLUTION
  PUMP
rSULFIDE
 MIXING
  TANK
                  DISTRIBUTION TROUGH
                        TREATED
                       EFFLUENT
                                 3R7

-------
Figure Vlll-3. CALSPAN MOBILE ENVIRONMENTAL TREATMENT PLANT CONFIGURATION
             EMPLOYED AT BASE AND PRECIOUS METAL MINE AND MILL OPERATIONS
                                 -EQUILIZATION BYPASS
                   3,785-liter
                  d.OOO^allon)
                   PRIMARY
              SETTLING/EQUALIZATION
                    TANK
      OVERFLOW
      TO WASTE
                 102-liter
                 (27-gillon)
               FLOCCULATION
        GRAV TY    TANK
        OVERFLOW
                                        SECONDARY
                                        TREATMENT
                                         BYPASS
FLOCCULATOR BYPAS
      SLUDGE _^
      TO WASTE"*
                                                     388

-------
  Figure VIII-4. MOBILE PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT
               URANIUM MILL 9402
        INFLUENT
       (END-OF-PIPE)
»s\* 5 gpm
                 METERING  METERING
p
./
BYPASS
~2_
SUMP
TANK
C

                                                                             OVERFLOW
                                                                                TO
                                                                               WASTE
 WASTE
 SLUDGE
OVER FLOW _
TO WASTE
                                                                             OVERFLOW
                                                                             TO WASTE
                                                                 MULTI-MEDIA
                                                                   PRESSURE
                                                                    FILTER
                                                                   COLUMNS
                                                                             OVERFLOW
                                                                             TO WASTE

                                                                           ^-BACKWASH
                                                                             TO WASTE

-------
 Figure VIII-5.  PILOT TREATMENT SYSTEM CONFIGURATION EMPLOYED AT URANIUM
              MILL 9401
OVERFLOW
 TO WASTE
 GRAVITY
OVERFLOW
•—I
1
B! ...
1
f


                                                                              IX
                                                                            ELUANT
                                                                     FERROUS   yw
                                                                     SULFATE    O

                                                                     (W/ORW/0 A
                                                                     H2S04)   
-------
Figure VIM-6. MODE OF OPERATION OF SEDIMENTATION TANK DURING PILOT-SCALE
           TREATABILITY STUDY AT MINE/MILL 9402
                 OVERFLOW WEIR
                                                          BAFFLE
DISCHARGE
                                                              INFLUENT
                                   391

-------
                  Figure VIII-7. FRONTIER TECHNICAL ASSOCIATES MOBILE TREATMENT PLANT CONFIGURATION
                             EMPLOYED AT MINE/MILL/SMELTER/REFINERIES #2121 & #2122
                         LIME
                         SLURRY
                                                                           PRESSURE
                                                                         ^INDICATOR
                                                                                          BACKWASH
                                                                                          PORTS
         INFLUENT
00
•o
                                                                                                       100 GAL*
                                                                                                       FILTRATE
                                                                                                       HOLDING
                                                                                                       TANK
                DRAIN
                   * 90 gallons ~ 0.34 cubic maters
                   100 gallons % 0.38 cubic meters
DRAIN

-------
                       Figure VIII-8.  DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT AT COPPER
                                     MINE/MILL 2120*
00
-o
                  OVERFLOW
                                        OPEN-PIT
                                      MINE WATER
                                                                              1
r
                                                	MISC. WASTES	 I	
                                                                                                               SERVICE
                                                                                                              RESERVOIR
                                       / SURGE \
                                       I   POND   1
                                                           UNDERGROUND
                                                            MINE WATER
                  DISCHARGE
                                       	INTERMITTENT WATER FLOW
                                       	PROPOSED WATER FLOW
                                                                                                 DISCHARGE
            •Water flow configuration representative of period during which historical monitoring data was generated.

-------
                       Figure VIII-9. DIAGRAM OF WATER FLOW AND WASTEWATER TREATMENT AT BASE
                                   AND PRECIOUS METALS MINE/MILL 2120 (COPPER)**
00
,0
                                     OPEN-PIT
                                   MINE WATER
                                                        UNDERGROUND
                                                         MINE WATER
                DISCHARGE
                                             CONTINUOUS WATER FLOW
                                     	INTERMITTENT WATER FLOW

                                     —	PROPOSED WATER FLOW
                                                                       THICKENERS
                                             	MISC. WASTES
                                                                                                        SERVICE
                                                                                                       RESERVOIR
                                                                                           CONCENTRATOR
                                                                                             TAILINGS
                                                                                       r
                                                                                            LIME


r BOX

I TAILING
I POND
 EXCESS
LEACHATE
                                                                                             DISCHARGE
                                                                                                           OVERFLOW
           ••Water flow configuration represents modifications made to system as of September, 1979

-------
Figure VIII-10.   SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT FACILITIES
                 AT LEAD/ZINC MINE/MILL 3103
                               MINEWATERPUMPAGE
                               94.6 I/sec (1,500 gal/min)
                                   MINE SUMP
                               82.0 I/sec (1,300 gal/min)
     TO SMELTER
         31.5 I/sec:
      (500 gal/min)
    MILL-WATER
STORAGE RESERVOIR
                                   113.5/0 I/sec
                                  (1.800/0 gal/min)
            RECYCLE
            63.1/0 I/sec
          (1,000/0 gal/min)
                                 CONCENTRATOR
   MILL TAILINGS
     94.6/0 I/sec
   (1,500/0 gal/min)
                   RAINWATER
                   est. 17.7 I/sec -
                  (est 280 gal/min)
  12.6/94.6 I/sec
(200/1,500 gal/min)
                                                  CONCENTRATE
                                                   THICKENERS
                                 OVERFLOW
                                 18.9/0 I/sec
                                (300/0 gal/min)
   EVAPORATION
   AND SEEPAGE
     est 13.2 I/sec'
   (est 210 gal/min)
     RAINWATER
      est 43.8 I/sec  _. _ _
   (est 695 gal/min)
   TAILING POND
                                  67.5/99.0 I/sec
                               (1,070/1,570 gal/min)
     SMALL
  STILLING POOL
   DISCHARGE
   11.3 I/sec
   (1,765 gal/min)

xxx/xxx
CONTINUOUS OR
SEMI-CONTINUOUS FLOW
INTERMITTENT FLOW
FLOW DURING MILL
OPERATION/FLOW DURING
MILL SHUTDOWN
                                                 395

-------
;=  80-t-
uj  40 • "
<
S  20
                                  MONTHS
   Fioure VIII-11.  PLOTS OF SELECTED PARAMETERS VERSUS TIME AT NICKEL
     a          MINE/MILL 6106 (NOVEMBER 1977 - DECEMBER 1980)

-------
             0.2£-r
                     Figure VIII-12. PLOT OF TSS CONCENTRATIONS VS. COPPER CONCENTRATIONS IN

                                 TAILING-POND DECANT AT MINE/MILL/SMELTER/REFINERY 2122
                                        D RND
TDTHL COPPER CDNCENTRRTIDN5
                                        •f RND	«= DISSOLVED COPPER CDNCENTRHTIONS
oo
-o
             8.2 •-
         .5   B.I5--

             a. i  ••
         tK
         Ul
                                   —i
                                                               	1_
                                                          IBB
             12?
ISB
ITS
HHB
                                              T55 CDNCETNTRHTinN CMB/L)

-------

-------
                            SECTION IX

            COST,  ENERGY,  AND NON-WATER QUALITY ISSUES

 DEVELOPMENT OF COST DATA BASE

 General

 Generalized  capital  and  annual   costs for wastewater treatment
 processes at ore  mining  and dressing facilities have been  estab-
 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
     Alkaline-Chlorination
     Activated  Carbon  Adsorption
     Hydrogen Peroxide Oxidation
     Chlorine Dioxide  Oxidation
     Potassium  Permanganate Oxidation
     Ion  Exchange
     Granular Media Filtration
     pH Adjustment
     Recycle
     Evaporation  Ponds (total evaporation)
Table   IX-1
subcategory.

CAPITAL COST
indicates  the  processes  studied  for  each  ore
Capital Cost of Facilities

Settling Ponds.   Construction costs for settling ponds were based
upon assumptions (specifically documented later in this  section)
                                399

-------
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.

-------
Capital Cost o£ Contingency

Unless otherwise stated, a contingency cost  of  20  percent  was
added to the total capital costs generated.  This was  intended to
cover taxes, insurance, over-runs and other contingencies.

ANNUAL COST

Annual Cost of Amortization

Initial capital costs were amortized on the basis of a  10 percent
annual interest rate with assumed life expectancy of 30 years for
general   civil   and  structural  equipment  and  10  years  for
mechanical and electrical equipment.   Capital  recovery  factors
were calculated using the formula:

                    n
     CRF = (r) (l+r)
                 n
           (l+r)   - 1
                                                . -   t      • •    •  _
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.
                                401

-------
     Hydrated Lime
     Activated Carbon
     Hydrogen Peroxide (70% cone.)                $
     Sulfuric Acid (66 Be)     '                   $
     Ferrous Sulfate (400 Ib. drum dry powder)    $
     Chlorine Dioxide (5% cone, in 55 gal. drum)  $
     Potassium Permanganate (dry)                 $
          $ 65.00/ton
          $  0.50/lb.
             0.35/lb.
             0.04/lb
             0.52/lb.
             9.10/gal.
             0.59/lb.
Reagent  dosages  are  documented  in
discussions in this section.

Annual Cost of Energy
the   treatment   process
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.
                                402

-------
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., OOP/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,
superstructure)  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 basi,s.

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
                               403

-------
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
dosage of five mg/1 and a contact time of  15 minutes.
an  ozone
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.
                                 404

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Alkaline Ch1orination

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.
         i   .        '•+
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,
assumptions were made, including:
several   system   design
     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
     each chemical
       30-day  supply  of
     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.
                              405

-------
Ion Exchange

Capital Costs.  The flow schematic for the ion exchange system is
exhibited in Figure IX-10.  This is a  combination  cation-anion-
mixed  bed  process  with  a pretreatment (filtration) step.  The
system costed consists of skid-mounted  package  units  including
raw   waste   filters   in  steel  tanks,  cation  exchangers,  a
degasifier, anion exchanger, and mixed bed exchangers.   Acid  is
provided  for  regeneration of cation exchangers and caustic soda
for anion  exchangers.   These  waste  solutions  are  mixed  and
require disposal.  (All units are housed in a structure.)

Total  capital  costs  include  equipment, installation, building
construction, and 20 percent contingency.  The capital cost curve
in Figure IX-11 relates this total capital cost to hydraulic flow
rate.

Supply and installation cost estimates  for  all  equipment  were
obtained  from vendor quotations.  Units were sized for hydraulic
loading according to vendor recommendations.  Building  construc-
tion  costs (including concrete foundations) were estimated based
on vendor space requirement quotes and the costing methodology of
References 2 and 3.

Annual Costs.  Amortization of  initial  capital  investment  was
based  on  a  10  percent  annual interest rate at a 10-year life
expectancy for equipment (CRF  =  0.16275)  and  a  30-year  life
expectancy  for construction (CRF = 0.10608).  Annual maintenance
costs were estimated at three percent of  initial  capital  cost.
Reagents   were  costed  at  $0.1 I/pound  for  caustic  soda  and
$0.03/pound for sulfuric acid.   Electric  power  was  costed  at
$0.03/KWH.   Operator  hours  were estimated at 20 hours per week
for plants treating less than 2.5 MGD, at 30 hours per  week  for
plants  treating  2.5  to  10.0 MGD, and at 40 hours per week for
plants treating 10.0 to 35.0 OMGD.  Operator wages  and  benefits
were estimated to total $13.50/hour.  Additional monitoring costs
of $10,000/year were assumed.  Figure IX-11 displays total annual
costs as a function of daily flow rate.

Granular Media Filtration

Capital  Costs.   Figure  IX-12  depicts the basic granular media
filtration  system  proposed  for  cost  estimates.   Industrial,
gravity flow deep bed, granular media filters were selected.  The
filters  would  be  contained  in  prefabricated,  portable steel
filter units.  Treated effluent would discharge  through  a  con-
crete  backwash  wastewater basin where the filtered solids would
settle.  The supernatant would then be pumped back to the filters
for treatment.

All piping is carbon steel, valves are the  butterfly  type,  and
the  pumping  equipment  consists  of  vertical  turbine pumps of
carbon  steel  construction.   Pump  impellers  are  bronze  with
                             406

-------
stainless  steel shafts.  Filter media consists of plastic filter
bottom, gravel, sand, and anthracite.

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.

pH 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.
                               407

-------
Total  capital  cost  estimates  included  storage  bins,  feeder
equipment,  concrete  and  lining  material for slurry and mixing
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
                              40R

-------
rate is graphed in Figure IX-18.  Cost curves are shown  for  25,
50, 75, and TOO percent recycle.

Annual  Costs.   Annual costs for wastewater recycle systems were
assumed to include the following:  (1) amortization calculated at
10 percent annual interest over 1.0 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 (qpm) 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.
cost curves for 25, 50, 75, and 100 percent recycle
shown in Figure IX-19.

Evaporation Pond
Total annual
systems  are
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
$1,000/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 (O.S.lbs) of phenol.  A total
phenol (4AAP) concentration of 0.4 mg/1 was  assumed  for  system
sizing.
                              40Q

-------
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

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
sulfaterphenol.   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
                               4in

-------
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.
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.
                                411

-------
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-
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
                              412

-------
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
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.

TREATMENT OF RECYCLE WATER

The final standard of performance for new froth  flotation  mills
extracting "  copper,  lead,  zinc,  gold,  silver  or  molybdenum
requires  zero  discharge  of  process  wastewater.   However,  a
discharge or bleed can be allowed if there is interference in the
mill  process  that can not be mitigated by appropriate treatment
of the recycle water.  The discharge allowed is  subject  to  the
standards for mine drainage.

Cost  of  treatment of the recycle water and cost of treatment of
the discharge can be determined from the cost data  presented  in
this  section.   Appropriate  treatment  consists  of  either  pH
adjustment followed by settling or pH adjustment,  settling,  and
mixed media filtration.

For example, if a new mill was to be built exactly as an existing
mill,  the  cost  of zero discharge, appropriate treatment of the
recycle water and treatment of the discharge  can  be  determined
from  tables  IX-2 to IX-10 by adding the cost given for recycle,
secondary settling, pH  adjustment,  and,  if  considered,  mixed
media  filters.  This total cost would represent the maximum cost
that would be incurred by a mill which has  a  tailings  pond  in
place,  treats  the  water, and recycles all back to the mill, or
discharges a bleed treated to meet mine drainage standards.
                              413

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TABLE IX-1.  COST COMPARISONS GENERATED ACCORDING TO TREATMENT PROCESS AND ORE CATEGORY
=====
Iron Ore

Copper Ore
Lead/Zinc Ores
Gold/Silver Ores
Aluminum Ore
Ferroalloy Ores
Mercury Ore
Titanium Ore
Uranium Ore
=====
Mines
Mills
Mines
Mills
Mines
Mills
" Mines
Mills
Mines
Mines
Mills
Mines
Mills
Mine/Mill
Mines
Mills
Secondary
x Settling
X
X
X
X
X
X
X
X
X
X


X
X
X

0"°
X
X
X
X
X
X
-X
X
X
X
X


X
X

| Ozonation

X
X
X
X
X
X


X





I
sl
II
<0
"

X
X
X
X
X
X


X





Ion Exchange


X

X

X


X



X

„ Granular Media
* Filtration
X
X
X
X
X
X
X
X
X
X


X
X

djustment
1
a
===^=

X
X
X
X
X
X

X
X


X
X
X
RECYCLE
25%


X

X

X


X





50%


X

X

X


X





75%


X

X

X


X





100%


X

X

X


X




X
» Activated
Carbon


X

X

X








I Hydrogen
Peroxide


X

X

X








!*
°'i
55


X

X

X








» Potassium
Permanganate


X

X

X








Total
Evaporation
Ponds














X

-------
                  TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                             AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                          p. 1 of 9
          Type of Mine: Iron Ore
Mine
Code -
Location
Type
1101-MN
Mine
1101-MN
Mill
1102-MN
-Mine
1102-MN
Mill
1103-MN
Mine/Mill
1104-MN
Mine
1104-MN
Mill
Ore
Production
(1000 tons;
year)
36,376
36,376
3,072
9,072
4,409
1 ,808
1,808
Hater
Uis-
charqed
(MOD)
21.13
0
0.69
0
0
1.05
5.94
	
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost ($1000)
c. Cost: 4/ton of ore mined
Second.
Settling
a. 340
b. 44.8
c. 0.12
a.
b. -
c.
a. 84
b. 19.0
c. 0.21
a.'
b. -
c.
a.
b. -
c.
a. 97
h. 20.2
c. 1.12
a. 187
b. 29
c. -1.60
Floc-
cula-
tion
110
160
0.44
-
65
30
0.33
-
-
70:
32
1.77
90
58
3.21
Ozon-
ation
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-'
Ion
Exchan.
-
-
-
-
-
••-
-
Mixed
Media
Hltr.
2000
310
0.85
-
150
47
0.52
-
-
206
55
3.04
803
140
7.74
PH
Adjust.
-
-
-
-
-
-
-
R e c y c 1 e
25%
-
-
-
-
-
-
-
50%
-
-
-
-
-
- -
-
75%
- .
-
-
-
-
-
-
100%
-
-
-
-
-
-
-

"rocess
Control
-
-
-
-
-
-
-
REMARKS







en

-------
        TABLE IX-2.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                   AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
                                                                                  p. "2 of 9
Type of Mine: I™" Ore
Mine
Code -
Location
Type
1105-HN
1105-MN
Mill
1106-MN
Mine/Mil
1107-HI
Mine
1107-HI
Hill
1108-MI
Mine
1108-MI
Mill
1109-M1
Mine
Ore
Production
(1000 ton$i
year)
9,149
9,149
44,092
5,842
5,842
9,700
9,700
18,078
Water
Ois-
charqed
(MOD)
12.70
0
0
3.53
2.69
4.17
8.71
N.A.
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (>1000)
c. Cost: t/ton of ore wined
Second.
Settling
a. 250
b. 35
c.0.38
a.
b. -
c.
a.
b. -
c.
a.150
b. 25.2
c. 0.43
a. 136
b. 24
c. 0.41
a.161
b. 26.5
c. 0.27
a.225
b. 33
c. 0.34
a.
b. -
c.
Floc-
cula-
tion
100
99
1.08
-
-
82
45
0.77
80
40
0.68
85
50
0.52
95
75
0.77
-
Ozon-
ation
-
-
-
-
-
-

-
Alkal.
Chlor-
ination
-
-
-
-
-
-

-
Ion

-
-
-
-
-
-


Mixed
Media
Filtr.
1400
220
2.40
-
-
525
100
1.71
430
85
1.45
605
110
1.13
1070
182
1.88
-
PH
A/1 1 tic 1-

-
-
-
-
-
-

-
Recycle
25X
-
-
-
-
-
-

-
50%
-
-
-
-
-
-

-
75S
-
-
-
-
-
. -

-

100%
-
-
- •
-
-
-

-

•rocess
ontrol
-
-
-
-
-
-

-
REMARKS









-------
        TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                  AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
                                                                                p. 3 of 9
Mine
Code-
Locatlon
Type
1109-MI
Hill
1110-PA
Mine
1110-PA
Mill
1111-MN
Mine
1112-MN
Mine
1112-MN
Mill
1113-MN
Mine/mill
1114-MO
Mine
Ore
Production
(1UOO tonv
year)
18.078
2,866
2,866
34,172
9.590
9,590
27,558
2,601
Hater
Dis-
charged
(MGD)
•5.94
N.A.
1.71
14.00
7.00
0
0
1.43
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD cdStS: b. Annual Cost (>1000)
c. Cost: • 31.5
c. 0.33
a.
b. -
c.
a.
b. -
c.
a. 107
l>. 21.1
c. 0.81
Floc-
cula-
tion
90
58
0.32
-
75
37
1.29
100
110
0.32
91
65
0.68
-
-
72
35
1.34
Ozon-
atfon
-
-
-
-
-
-
-
-
Alkal.
Chlor-
1 nation
-
' -
-
-
-
-
-
-
Ion
:xchan.

- •
-
-
-
-
-
-
-
Mixed
Media
Flltr.
800
140
0.77
-
300
70
2.44
1500
230
0.67
900
150
1.56
-
-
253
62
2.38
PH
Adjust
-
-
-
-
-
-
- '
-
Recycle
25X
-
-
-
-
-
-
-
-
SOX

-
- •
-
-
-
-
-
75X

-
-
-
-
-
- •
-
100X

-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
REMARKS









-------
                  TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES

                             AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
                                                                                           p. 4 of 9
          Type of Mine: Iron Ore
Mine
Code -
Location
Type
1114-MO
Mill
1115-MO
Mine
1115-MO
Mill
1116-WI
Mine/Mill
1117-UT
Mine/Mill
1118-CA
Mine/Mill
Ore
Production
(1000 ton$<
year)
2.601
2,425
2.425
2.425
2.645
9.028
Vlater
Dis-
charged
(MGD)
1.71
NA
4.17
0
NA
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (J1000)
c. Cost: (/ton of ore mined
Second.
Settling
a. 115
b. 22
c. 0.85
a.
b. -
c.
a. 161
t>- 26.5
c- 1.09
a.
b. .
c.
a.
b. .
c.
a.
b. -
c.
Floc-
cula-
tion
75
37
1.42
-
85
50
2.06
-
- •
-
Ozon-
ation
-
-
-
-
-
-
Alkal.
Chlor-
Ination
-
-
-
-
-
-
Ion

-
-
-
-
-
w
Mixed
Media
Filtr.
300
70
2.69 ;
-
605
110
4.54
-
-
-
PH
Adjust.
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
50%
-
-
-
-
-
-
75X
-
-
-
-
-
-
100X
-
-
-
-
-
-

'rocess
Control
-
-
-
-
-
-
REMARKS






-p.
t—"
JO

-------
        TABLE IX-2.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                   AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
                                                                            p. 5 of 9
Mine
Code -
Location
Type
1119-HY
Mine
1119-HY
Mill
1120-Ml
Mine
1120-HI
Mill
1121-MN
Mine
1121-MN
Mill
1122-MN
Mine
1122-MN
Mill
Ore
Productlor
(1000 tons
year)
4,850
4,850
4,630
4,630
1,194
1,194
8,157
I
8,157
Hater
Dis-
charged
(MGD)
0.45
Minimal
0.18
Minimal
5.94
1.44
17.80
0
a. Capital Cost ($10UO)
TREATMENT TECHNOLOGIES ANO COSTS: b. Annual Cost (HOOO)
c. Cost: t/ton of ore mined
Second.
Settling
a. 75
b. 18.2
c. 0.38
a.
b. -
c.
a. 56
b. 17
c. 0.37
a.
b. -
c.
a. 186
b. 29.2
c. 2.45
a. 108
b. 21.1
c. 1.77
a. 315
b. 42.3
c. 0.52
a.
>. -
<:.
Floc-
cula-
tion
60
28
0.58
-
54
27
0.58
-
90
58
4.86
72
35
2.93
105
125
1.53
-
Ozon-
ation
-
-
-
-
-
-
- .
-
Alkal.
Chlor-
inatlon
-
-
-
-,
-
-
-
-
Ion
Exchan.
-
-
-
-

-
-
-
Mixed
Media
Filtr.
110
42
0.87
-
50
33
0.71
-
800
140
11.73
254
63
5.28
1750
280
3.43
-
PH
Adjust
-
-
-
-
-
-
-
-
Recycle
25X
«•
-
-
-
-
-

i
50%
-
-
-

-
-
-
-
75X
-
-
-
-
-
-
-
-
1002
-
• -
-
-
-
-
-
-
Process
Control
-
-
-
- .
.
_
_
-
REMARKS









-------
        TABLE IX-2.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                   AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
                                                                              p. 6 of 9
Type of Mine:  Iron Ore
— 	
Mine
Code -
Location
Type
1123-MN
Mine
1123-MN
Mill


1124-MN
Mine
1124-MI
Mill

1125-MN
Mine/Mill
1126-MN
Mine/Mill

1127-UT
Mine/Mill

1128-NM
Mine/Mill

Ore
Productior
(1000 tons;
year)
2 .535

2,535


11,905
11,905

1.543
2,425

1,874


71.65

Haler
01s-
charqed
(MOO)
2.98

0


2.98
0

NA
NA

0

-
0
a. Capital Cost ($1000)
TREATMENT TECIWOLORIES AW) COSTS: b. Annual Cost (>1000)
c. Cost: 
-------
      TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Iron Ore
                                                                                 P. 7 of 9
Mine
Code -
Location
Type
1129-TX
Mine
1129-TX
Mill
1130-NY
Mine/Mill
1131-NY
Mine
1131-NY
Hill
1132-WY
Mine/Mill
1133-MN
Mine
1134-MN
Mine/Mill
Ore
Production
(1000 tons;
year)
2,380
2,380
1,984
3,858
3,858
1,433
0
1 ,433
Hater
1)1 s-
charqed
(MGD)
NA
0
NA
0.44
16.36
NA
NA
NA
TREATMENT TECHNOLOGIES AMD
Second.
Settling
a.
b. .
c.
a.
b. :
c.
a.
b. .
c.
a. 74
b. 18.1
c. 0.47
a.306
b. 38
c. 0-98
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floc-
cula-
tlon
-
-
-
60
28
0.73
100
110
2.85
-
-
-
Ozon-
ation
-
-
-
-
-
- -
-
-
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
a. Capital Cost ($101)0)
COSTS: b. Annual Cost (>1000)
c. Cost: 6/ton of ore mined
Ion
Exchan.
-

-
-
-
: -
-
-
Mixed
Media
Filtr.
-
-
-
108
41
1.06
1650
250
6.48
-
-
-
PH
Adjust
/'
-
-
-
-
-
-
-
-
Recycle
25X
: -
-
-
-
-
. -
-
-
SOX
-
-
-
-
-
-
• -
-
75*

-
-
-
-
-
-
«•
100%
1 •
^
-
-
-

-
-
Process
Control
-
-
-
-
-

-
-
REMARKS






Operation
closed


-------
                 TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES

                           AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
         Type of Mine: Iron Ore
                                                                                       p. 8 of 9
Mine
Code-
Locatlon
Type
1135-MN
Mine/
Mill
1136-MI
4ine
11137-CA
Mine/
Mill
1138-MN
Mine/
Mill
1139-GA
Mine
1140-MN
Mine
1141-MN
Mine
1142-MN
Mine/
Mill
Ore
Production
(1UOO tons
year)
1,212
301
496
9735
0
0
0
NA
Water
Dis-
charged
(MGD)
NA
120.46
0
0
NA
NA
NA
•
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AND COSTS: b. Annual Cost (ilOOO)
c. Cost: i/ton of ore mined
Second.
Settling
a.
b. -
c.
a. 820
b. 96
c.31.89
a.
b. -
c.
a.
b. -
c.
a.
b. - .
c.
a.
b. -.
c.
a.
b. -
c.
a.
b. -
• c.
Floc-
cula-
tlon
-
140
860
285.71
-
-
-
-
-
-
Ozon-
atton
-
-
-
-
-
• -

-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
:xchan.
-
-
-
-
-
-
-
-
Mixed
Media
Flltr.
-
6000
1060
352.16
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-

-
Recycle
25?
-
-
-
-
-
-
-
-
50%
-
-
-
•-
-

-
-
75%
-
-
-
-
-
-
-
-

100X
-
-
-
-
-
-
-
-

'rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS




Operation
assumed
closed
it
it

ro
ro

-------
                    TABLE IX-2. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
         Type of Mine:  Iron Ore
                                                                                   p. 9 of 9
rtine
Code -
Location
Type
114 3- MN
Mine/
Mill
1144-MI
Mine
114S-NV
Mine
1146-MN
Mine
1147-MN
Mine
1148-MN
Mill
1149-MN
Mill


Ore
Production
(1UOO tons/
year)
2,648
1,874
115
661
413
1297
0

Water
Dis-
charged
(MGD)
1.06
3.25
0.63
10.00
2.19
0
NA.
1
TREATMENT TECHNOLOGIES AMD
Second.
Settling
a. 98
b. 20.2
c. 0,76
a. 148
b. 25.2
c. 1.34
a. 83
b. 18.8
C. 16.35
a. 240
b. 34.6
c. 5.23
a. 126
b. 23
c. 5.57
a.
b. -
c.
a.
b. -
c.
a.
i,
c.
Floc-
cula-
tion
70
32
1.21
82
43
2.30
64
28
24.35
98
82
12.41
79
37
8.96
-
-

Ozon-
ation

-
-
-
-
-
-

Alkal.
Chlor-
1 nation

-
-
-
-
-
-

a. Capital Cost ($1000)
COSTS: b. Annual Cost (3,1000)
c. Cost: t/ton of ore mined
Ion
Exchan.

-
-
-
•-
-
-

Mixed
Media
Filtr.
200
55
2.08
500
95
5.07
145
46
40.00
1200
185
27.99
360
79
19.13
-
-

PH
Adjust

-
-
-
-
-
-


25%

-
-
-
-
-
-

Recycle
50%

-
-
-
-
-
-

75%

-
-
-
-
-
-

100%

-
-
-
-
-
-

Process
Control

-
-
-
-
- '
-

REMARKS









00

-------
              TABLE IX-3.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mini:
Bmt and Prociouj
Metah (Copper)
                                                                                            P.1of7
Mine Coda •
2101-NV
Mine/Mill
2102-AZ
Mine/Mill
2103-NM
Mine/Mill
2104-NM
Mine/Mill
2107-AZ
Mine/Mill
2108-AZ
Mine/Mill
2109-AZ
Mine/Mill
2110-AZ
Mine
Ore
Production
(1000
7,932
6,015
15,403
8,101
4,402
3,066
3,729
4,090
Water
Discharged
0
0
0
0.18
0
0
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Co»t ($1000)
b. Annual Cost ($1000)
c. Cost: i /ton or ore mined
Second.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 61
b. 17
c. 0.21
a. 61
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
-
-
-
55
27
0.33
-
-
-
-
Ozona-
tion
-
-
-
32
27
0.33
-
-
-
-
Alkal.
Chlorin-
ation
-
-
-
48
27
0.33
-
-
-
-
Ion
Exchan.
-
-
-
850
250
3.09
-
-
-
-
Mixed
Madia
Filtr.
-
-
-
50
33
0.41
-
-
-
-

PH
Adjust.
-r
-
-
26
23
0.28
-
-
-
-
Recycle
25%
-
-
-
7
16
0.20
-
-
-
-
50%
-
-
-
9.5
16.5
0.20
-
-
-
-
75%
-
-
-
13
17
0.21
-
-
-
-
100%
-
-
-
17
18.5
0.23
-
-
-
-
Activated
Carbon
-
-
-
120
67
0.83
-
-
-
-

Hydrogen
Peroxide
-
-
-
140
91
1.12
-
-
-
-

Chlorine
Dioxide
-
-
-
108
83
1.02
-
-
-
-

Potassium
Permang-
anate
-
-
-
160
99
1.22
-
-
-
-

Process
Control
-
-
-
-
-
-
-
-
Remarks




Operation
inactive


Operation
presently
inactive

-------
TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
          AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Base and Precious
Metals (Copper) .p. 2 of 7
Mine Code
Location Type
2111-AZ
Mine/Mill
2112-AZ
Mine/Mill
2113-AZ
Mine/Mill
2115-AZ
Mine/Mill
2116-AZ
Mine/Mill
2117-TN
Mill
2118-AZ
Mine/Mill
2119-AZ
Mine/Mill
Ore
Production
(1000
tons/year
1,631
670
10,340
1,555
9,804
2,024
18,357
15,013
Water
Dischargee
(MGD)
=
NA
0
0
0
0
8.50
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: i /ton or ore mined
Secon
Settlin
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 220
b. 32
c. 1.58
a.
b. -
c.
a.
b. -
c.
Floccu
lation
-
.-
-
-
95
75
171
-
-
Ozona
tion
-
-
-
-
540
157
7.76
-
-
Alkal.
Chlorin
ation
-
-
-
-
-
230
237
11.71
-
-
Ion
Exchan
-'
-
-
-
-
11000
2760
136.36
-
-
Mixed
Media
Filtr.
-
-
-
-
-
1050
180
8.89
-
- .
pH
Adjus
-
-
-
-
68
70
3.46
-
-
Recycle
25%
-
-
-
-
-
60
31
1.53
-
-
50%
-
-
-
-
90
43
2.12
-
-

75%
-
-
-
-'
130
57
2.81
-
-

ioor
-
-

-
••-
170
70
3.45
-
-


Activated
Carbon
-
-

-
-
1650
805
39.77
-
-

Hydroge
Peroxid
-
-
-
-
-
330
195
9.63
-
-

Chlorine
Dioxide
-
-
-'
-
-
285
185
9.14
-
-

Potassium
Permang-
anate
-
-
-
-
-"
1100
345
17.05

-

Process
Control
-
-
-
-
-

-
-


Remarks
Inactive








-------
               TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                          AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mint:
B«o and Precious
MetaU (Coppar)
                                                                                             p. 3 of 7



Mini Cod* •
=====
2120-MT
Minn

2120-MT
Mill

2121-MI
complex

2122-UT
Mill

2123-AZ
Mine/Mill
2124-AZ
Mine/Mill

2125-AZ
Mine


2126-NV
Mine/Mill



On
Production
(1000
—
17,000


17,000

3,617

35,500

2,047


6,710

0



8,000




Water
Discharged
(MGD)
-' —
0.05


9.50

32

8.50

0


0

0



0


TREATMENT TECHNOLOGIES AND COSTS: «. Capital Cost ($1000)
b. Annual Cost ($1000)

Second.
Settling
i -.-.:
a. 58
b. 17
c. 0.10
a. 235
b. 34
c. 0.20
a. 400
b. 50
c. 1.38
a. 220
b. 32
c. 0.09
a.
b. -
c.
a.
b. -
a.
b. -
..
c.
a.
b. -

c.
:loccu-
ation
==
45
26
0.15
97
80
0.47
120
210
5.81
95
75
0.21

' -


~

_



_


Orona-
lion
===
20
24
0.14
600
177
1.04
1800
470
12.99
540
157
0.44

-




—



_


Alkal.
Chlorin-
ation
- - ^
42
25
0.15
260
257
1.51
610
737
20.38
230
237
0.67

-




-



—


Ion
Exchin.

-

12000
3110
18.29
42000
13010
359.69
11000
2760
7.77

—




-



-


Mixed
Media
Fillr.
==
18
28
0.16
1150
190
1.12
2500
400
11.06
1050
180
0.51

—




-



-


PH
Adjust.
3
2
0.13
70
75
0.44
125
190
5.25
68
70
0.20

—




—



-


Recycle
26X

-

65
33
0.19
165
70
1.94
60
31
0.08

—




-



-


50%
-
—

00
46
0.27
270
125
3.46
90
43
0.12

•~




—



—


75X

—

145
62
0.36
380
170
4.70
130
57
0.16

*~




—



—


00%

—

180
75
0.44
485
230
6.36
170
70
0.20

™




—



—



Activated
Carbon

•~

1800
905
5.32
5200
2705
74.79
1650
805
2.27






—



—



lydrooan
'eroxlde

—-

340
205
1.21
580
400
11.06
330
195
0.55






~



~~



Chlorine
Dioxide

~

300
195
1.15
540
390
10.78
285
185
0.52














Permang-
anate



200
360
2.12
3200
820
22.67
1100
345
0.97














'roceu
Control

























Remark!





Already
meeting
BAT.







Temporarily
inactive







-------
              TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine: Base and Precious
        Metals (Copper)
                                                                                               "p. 4 of 7
Mine Code •
Location Type
2130-NM
Mine/Mill
2131-NV
Mine
2132-NV
Mine/Mill
2133-NV
Mine/Mill
2134-ID
Mine
2134-ID
Mill
2135-AZ
Mine
2136-AZ
Mine
Ore
Production
11000
tons/year]
NA
NA
NA
0
NA
NA
8.27
NA
Water
Dischargee
(MGD)
Minimal
0
0
0
0
NA
0
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /(on or ore mined
Second
Settlin
z.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
3. —
Floccu
lation
-
-
- -
-
-"
-
-
Ozona
tion
-
-
-
- .
'-.
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
-
Ion
Exchan
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
- .
-
-

50%
-
-

-
*-
-
-
-

75%
-
-
-

-
-
-
-

100%
-
-
-

--'
-
-'
-


Activated
Carbon
-
-
-
-
- .
-
-
-

Hydrogen
Peroxide
-
-

-
-
-
-
-

Chlorine
Dioxide

-
-
-
-
•-
-
-

Potassium
Permang-
anate
-
-
-
-
. -
-
-
-

Process
Control
-
-
-
-
-
-
-
-


Remarks
Mine dis-
charges to
mill -vary
small or zero


Closed
permanently

Partial
recycle
i ^— •^••••^
Dperation
nactive


-------
                    TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                               AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn « 2000 Ibs)
       Type of Mini:
B»t «rd Precious
Metals (Copper)
p.5of7


Mine Cod. •

2137-AZ
Mine/Mill

2138-AZ
Mine/Mill

2139-AZ
Mine/Mill


2140-AZ
Mine/Mill


2141-AZ
Mine/Mill


2142-AZ
Mine

2143-AZ
Mine


2144-AZ
Mine



Or*
Production
(1000

NA


5,000


32,494



5,800



5,300



1,820


NA


NA




Water
Discharged

0


0


0



0



0



0


NA


0


TREATMENT TECHNOLOGIES AND COSTS: •. Capital COM ($1000)
b. Annual Cost ($1000)
c. Coil: * /ton or ore mined
Second.
a.
b. -
G.
a.
b. -
c.
a.
b. -
c.
a.

b. —
c.
a.


c.
a.


c.
a.
b. -
c.
a.
b. -

c.
Floccu-

-


-


-



—



~






—


	


Ozona-

-


-


-



~*










—


	


Alkal.
Chlorin-
ation

-


-


-



—










—


	


Ion
Exctun.

-


-


-



™










—


	


Mixed
Media
Fillr.

-


-


-



~










—


_


pH
Adjust.

-


-


-



"™










—


_


Recycle
25%

-


-


-



~










—


_


50%

-


-


-














—


_


76%

-


-


-














—


_


100%

-


-


-














—


_


Activated
Carbon

-


-


-














~


-


•lydrogen
Peroxide

-


—


-














~


-


Chlorine
Dioxide

-


—


—

















-



Potassium
Permang-
anate

—


—


—










—






-



Process
Control

—


—


—










_






-




Remarki





















Suspected
inactive





s

-------
             TABLE IX-3.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                        AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
          Base ind Precious
          Metals (Copper)
                                                                                               p. 6 of 7
Minn Coda •
Location Type
2145-AZ
Mine/Mill
2146-AZ
Mine/Mill
2147-AZ
Mine/Mill
2148-AZ
Mine/Mill
2149-AZ
Mine
2150-UT
Mine/Mill
2151-MI
Mine/Mill
2152-NM
Mill
Ore
Production
(1000
tons/year)
NA
MA
19,600
NA
NA
NA
NA
0
Water
Discharged
(MGD)
0
0
0
0
0
NA
NA
NA
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. - '
c.
a.
b. -
c.
Floccu-
lation
-
-
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
- -
-
Ion
Exchan
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-


-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
.-
-
-
-
-
-
50%
-
-
.
-
-
-
-
-
75%
-
• -
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-

Activated
Carbon
-
-
-
-
'
-

-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-'
-
Potaaium
Permang-
anate
-
-
-
: -
- -
-
-
-
Process
Control
-
-
-
-
-
-
-
- -



Temporarily
inactive
Probably
inactive

Under devel-
opment -
) discharge
ikely
Pilot-scale
iroduction
Temporarily
nactive

-------
             TABLE IX-3. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                        AND SUBSEQUENT COST PER TON OF ORE MINED {1 tn = 2000 Ibs)
Typ* at Mint:
Btw and Prociouj
Metils [Copper)
                                                                                               p.7of7
Mint Cod* •
2154-AZ
Mint

















'
Or*
Production
(1000
NA


















Water
Discharged
(MGD)
NA


















TREATMENT TECHNOLOGIES AND COSTS: «. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Coit: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
Flo ccu-
lation
-


















Ozona-
tion
-


















Alkal.
Chlorin-
ation
-


















Ion
Exchan.
-


















Mixed
Media
Filtr.
-


















PH
Adjust.
-


















Recycle
25%
-


















50%
-


















75%
-


















100%
-


















Activated
Carbon
-



















rlydrogen
Peroxide
-



















Chlorine
Dioxide
-



















Potassium
Permang-
anate
-



















Process
Control
-


















Remarks
Under
develop-
ment or
exploration



















-------
                   TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
      Type of Mine:  Base and Preciout
              Metals (Lead-Zinc)
p. 1 of 8



Mine Code •
Location Type
3101
Mine/Mill

3102-MO
Mine/Mill

3103-MO
Mine/Mill

3104-NY
Mill


3105-MO
Mine

3106-PA
Mine

3106-PA
Mill

3107-1 D
Complex



Ore
Production
(1000
tons/year)

206


1,634

1,072


1,112


1,138


383


383


782




Water
Discharged
(MOD)

0.38


5.94

2.58


1.78


2.19


28.53


1.50


. 5.94

TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)

Second
Settling
a. 70
b. 17.5
c. 8.50
a. 180
b. 29
c. 1.77
a. 135
b. 24
c. 2.24
a. 115
b. 22
c. 1.98
a. 124
b. 23
c. 2.02
a. 390
b. 48
c. 12.53
a. 110
b. 21.4
c. 5.59
a. 180
b. 29
c. 3.71

Floccu-
lation
60
27
13.11
90
60
3.67
80
40
3.73
75
37
3.33
78
38
3.34
20
90
49.61
75
35
9.14
90
60
7.67

Ozona-
tion
50
35
16.99
400
122
7.47
195
74
6.90
145
59
5.31
180
66
5.80
1680
422
110.18
130
54
14.10
400
122
15.60

Alkal.
Chlorin-
ation
54
32
15.53
195
172
10.5
120
85
7.92
92
65
5.85
110
75
6.59
585
667
174.15
88
59
15.40
195
172
21.99
c. Cost: t /ton or ore mined
Ion
Exchan
1200
330
160.19
7500
2210
135.25
4000
1210
112.87
3000
860
77.34

_


— •

2700
760
198.43
7500
2210
282.61
Mixed
Media
Filtr.
90
40
19.42
800
140
8.57
400
83
7.74
300
70
6.29
340
75
6.59
2400
370
96.61
260
65
16.97
800
140
17.90
PH
Adjust
30
25
12.1
60
56
3.43
45
38
3.54
42
34
3.06
43
36
3.16
115
170
44.38
41
33
8.62
60
56
7.16
Recycle
25%
9.5
16
7.77
50
27
1.65
30
22
2.05
23
20
1.80

	


_

21
20
5.22
50
27
3.45
50%
14
16.5
8.01
73
37
2.26
44
27
2.58
32
24
2.16

_


_

30
23
6.00
73
37
4.73
75%
18
17.5
8.50
100
47
2.88
60
31
2.89
44
27
2.43

_


—

40
25
6.53
100
47
6.01
100%
24
19.5
9.47
130
57
3.49
75
36
3.36
56
30
2.70




_

52
28
7.31
130
57
7.29
Activated
Carbon

-

1260
600
36.72
700
305
28.45
540
235
21.13




_

460
205
53.52
1260
600
76.73
Hydrogen
Peroxide

_

290
165
10.10
225
130
12.13
203
120
10.79






195
115
30.03
290
165
21.10
Chlorine

_

250
160
9.79
190
125
11.66
170
115
10.34






168
110
28.72
250
160
20.46
Potassium
Permang-

_

870
270
16.52
520
185
17.26
410
160
14.39






380
150
39.16
870
270
34.53
Process

„


_




_











_






Closed





















CO

-------
                   TABLE 1X4.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
     Typ* of Mint: Baa and Precious
             Mtti!» (L«»d-Zinc)
                                                                                                      p.2of8



Mine Code -
Location Type

3108-TN
Mill


31 09 -MO
Mine/Mill

3110-NY
Mill

3111-TE
Mine

3112-NM
Mine


3113-CO
Mine


3113-CO
Mill


3114-ID
Mine/Mill




Ore
Production
(1000
tons/year)
=
391


1,117

103


100


135



203



203


68




Water
Discharged
(MGD)

0.05


7.50

0.58


0.95


0.66



1.69



1.40 -


0.42

TREATMENT TECHNOLOGIES AND COSTS: «. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Coil: t /ton or or* mlntd
Second.
Settling
a. 55
b. 16.5
c. 4.22
a.210
b. 3?
c. 2.86
a. 80
b. 18.6
c. 18.0E
a. 94
b. 19.9
c. 19.90
a. 84
b. 18.9
c. 14.00
8.112

b. 22.0
c. 10.84
a.105

b. 21.0
c. 10.34
a. 74
b. 18.0
c. 26.47
Floccu-
lation
45
24
6.14
95
67
6.00
68
28
27.18
70
32
32.00
65
30
22.22
75

34
16.75
75

34
16.75
61
27
39.71
Oiona-
tion
20
24
6.14
490
147
13.16
64
36
34.95
92
43
43.0
69
38
28.15
140

57
28.08
120

53
26.11
52
33
48.53
Alkal.
Chlorin-
ation
42
25
6.39
220
217
19.43
62
36
34.95
72
47
47.00
64
39
28.89
92

62
30.54
85

58
28.57
56
33
48.53
Ion
Exchan.
460
150
38.35
9000
2610
233.66
1600
450
436.89

-


-

2700

810
399.01
2500

730
359.61
1400
360
529.41
Mixed
Media
Filtr.
18
28
7.16
930
155
13.88
140
46
44.66
190
54
54.00
145
47
34.81
280

67
33.00
250

61
30.05
95
41
60.29
PH
Adjust.
23
22
5.63
65
65
5.82
33
27
26.21
36
28
28.00
34"
26
19.26
42

33
16.26
40

32
15.76
31
26
38.23
Recycle
25%
3.5
15
3.84
55
30
2.69
13
17
16.50

-


-

23

20
9.85
20

19
9.36
11
16
23.53
50%
5
16
4.10
80
43
3.85
17
18
17.48

-


-

32

23
11.33
28

22
10.84
15
16.5
24.26
75X
7
16.5
4.22
125
55
4.92
24
19
18.4!

-


-

45

27
13.30
39

25
12.32
19
17.5
25.74
100%
8
17
4.35
150
66
5.91
30
22
21.36

-


-

55

30
14.78
49

28
13.79
24
20
29.41
Activated
Carbon
52
47
12.02
1500
725
64.91
250
115
111.65

-


-



—

450

195
96.06
200
97
142.65

Hydrogen
Peroxide
124
85
21.74
315
185
16.56
160
99
96.12

-


-



—

190

115
56.65
155
95
139.71

Chlorine
Dioxide
86
75
19.18
275
180
16.11
134
94
91.26

-


-



—

165

108
53.20
126
92
135.29

Potattlum
Permang-
•nate
25
89
22.76
030
320
28.65
240
120
116.50

-


-



—

360

150
73.89
205
115
169.12

Proem
Control

-




-


-


-



—



—


-




Remarks
—-

























CO
ro

-------
                   TABLE IX-4.
                         COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Type of Mine:
             Base and Precious
             Metals (Lead-Zinc)
                                                                                                   p. 3 of 8



Mine Code -
Location Type
3115
Mine/Mill


3116-CO
Mine

3118-VA
Mine

3118-VA
Mine


3118-VA
Mill

3118-VA
Mill

3118-VA
Mine/Mill

3119-MO
Mine/Mill



Ore
Production
(1000
tons/year!

372



198

596

596


596


596


596


647




Water
Discharge
(MGD)

4.7



0.87

•1.80

2.60


0.01


0.20


14.00


1.78

TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)

Seconi
Settlin
—
a. 160
b. 26
c. 6.99

a. 91
b. 19.6
c. 9.90
a. 118
b. 22.1
c. 3.71
a. 135
b. 23.7
c. 3.98
a. 55
b. 16.5
c. 2.77
a. 60
b. 16.8
c. 2.82
a. 280
b. 38
c. 6.38
a. 115
b. 22.0
c. 3.40

Floccu
lation
90
51
13.71

69
32
16.16
75
35
5.87
80
40
6.71
35 '
25
4.19
55
26
4.36
100
110
18.46
75
35
5.41

Ozona
tion
320
98
26.34

84
42
21.21
146
60
10.07
196
75
12.58
15
21
3.52
33
27
4.53
860
242
40.60
145
59
9.12

Alkal.
Chlorin-
ation
165
147
39.52

70
45
22.73
96
67
11.24
120
86
14.42
42
25
4.19
48
28
4.70
350
357
59.89
92
65
10.04

Ion
Exchan
6000
1610
432.80


-

2900
810
135.91
3900
1110
186.24
330
120
20.13
880
250
41.95
16000
4510
756.71
2900
810
125.19
c. Cost: i /ton or ore mined
Mixed
Media
Filtr.
640
120
32.26

180
51
25.76
310
72
12.08
400
88
14.77
10
17
2.85
54
33
5.54
1500
230
38.60
310
72
11.13
pH
Adjust
53
50
13X14

36
28
14.14
42
34
5.70
46
38
6.38
18
22
3.69
28
23
3.86
85
100
16.78
42
34
5.26
Recycle
25%
43
25
6.72


_

23
20
3.36
30
22
3.69
1.7
15
2.52
6.9
16
2.68
87
40
6.71
23
20
3.09
50%
60
32
8.60


_

32
23
3.86
42
26
4.36
2.5
16
2.68
9.5
16.5
2.77
130
59
9.90
32
23
3.55
75%
84
39
10.48


	

45
26
4.36
58
30
5.03
3.1
16.5
2.77
12
17
2.85
170
80
3.42
45
26
4.02
100%
110
45
12.10


—

60
30
5.03
73
35
5.87
4.2
17
2.85
16
18
3.02
240
100
16.78
60
30
4.64

Activatec
Carbon

_



_





20
40
6.71
125
70
11.74
2500
1255
210.57
540
235
36.32

Hydrogen
Peroxide

_









120
85
14.26
140
90
15.10
400
245
41.11
203
120
18.55

Chlorine
Dioxide

_









70
70
11.74
110
82
13.76
350
240
40.27
170
115
17.78

Potassium
Permang-
anate

_









115
85
14.26
165
100
16.78
1650
465
78.02
410
160
24.73

Process
Control

_

























Remarks
Closed.
Annual pro-
duction based
on 250
working days



Ore produc-
tion includes
Mine 31 17
Ore produc-
tion includes
Mine 31 17













-ft.
oo
CO

-------
                   TABLE 1X4.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
    Type of Mini: Bast and Precloui
            Metals (Lead-Zinc)
                                                                                                    p.4of8


Wins Cod. -
Location Typ«
-
3120-10
Mine/Mill

3121 -ID
Mine

3121-ID
Mill

3122-MO
Mine

3123-MO
Mine


3124-NJ
Mine


3125-NY
Mine

3127-TN
Mine



Or*
Production
(1000
tons/year)
=
174

283


283

1,111


1,774


205


24


721



Witir
Discharged
(MGD)

1.24

1.24


1.58

6.90


9.60


0.25


0.37


1.45

TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: i /ton or or* mined
Second.
Sattling
a. 102
b. 20.7
c. 11.90
a. 102
b. 20.7
c. 7.31
a. 111
b. 21.6
c. 7.63
a. 205
b. 30.9
c. 2.78
a. 230
b. 34.0
c. 1.92
a. 64
b. 17.3
c. 8.44
«. 72
b. 17.8
c.74.17
a. 108
b. 21.2
c. 2.94
Floccu-
lation
71
33
18.97
71
33
11.66
75
34
12.01
95
65
5.85
98
80
4.51
57
27
13.17
60
28
116.67
75
34
4.72
Ozona-
tion
110
49
28.16
110
49
17.31
136
55
19.43
460
139
12.51
600
177
9.98
37
29
14.15
47
32
133.33
125
53
7.35
Alkal.
Chlorin-
ation
80
52
29.88
80
52
18.37
88
60
21.20
210
197
17.73
260
257
14.49
50
30
14.63
54
31
129.17
85
56
7.76
Ion
Exchan.
2300
650
373.56

-

2900
780
275.62

—





-


-


—

Mixed
Madia
Filtr.
230
60
34.48
230
60
21.20
270
66
23.32
900
160
14.40
1165
191
10.77
65
35
17.07
90
38
158.33
255
63
8.74
pH
Adjust.
39
31
17.82
39
31
10.95
41
34
12.01
63
62
5.58
71
76
4.28
28
23
11.22
30
24
100.00
40
31
4.29
Recycla
25%
19
19
10.92

-

22
20
7.07

-


-


-


-


"~

SOX
26
21
2.07

-

30
23
8.13

-


-


-


-


—

75X
37
24
3.79

—

42
26
9.19

-


-


-


-


^

100X
47
27
15.52

-

52
29
10.25

-


-


-


-


~

Activated
Carbon
410
180
103.45

—

500
215
75.97

—


—


-


-





Hydrogen
Peroxide
190
113
64.94

—

200
120
42.40

—


—


—


—





Chlorine
Dioxide
160
104
59.77

~

165
110
38.87

—


—


—


—





Potaolum
'ermanB-
anate
340
145
83.33

~~

390
150
53.00

~


~~


—


—





Proem
Control



«




•~


~™


•~


— •






Rtnwkt
— !






















co

-------
                    TABLE IX-4. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                               AND SUBSEQUENT COST PER TON OF ORE MINED (1tn = 2000 Ibs)
      Type of Mine:
Ban and Precious
Matali(Laad-Zinc)
                                                                                                   p. 5 of 8

•-.

Mint Cod* -
Location Type
3128-TN
Mine

3130-UT
Mine


3131-WI
Mine

3132-WI
Mine

3133-WI
Mine



3133-WI
Mill

3134-WA
Mine

3135-WA
Mine/Mill




Ore
Production
(1000
tons/year)

526


NA


MA

NA


0



0



301

*





Water
Dischargee
(MGDI

1.45


8.50


2.00

1.16


NA



0.76



0.71

0


TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)

Seconc
Settling
a. 108
b. 21.2
c. 4.03
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -

c.
a.
b. -

c.
a. 86
b.19.0
c. 6.31
a.
b. -
c.


Floccu-
lation
75
34
6.46

—


—

_


•"_



_


66
30
9.97




Ozona-
tion
125
53
10.0

-


_

_


„'



_


73
40
1329




Alkal.
Chlorjn-
ation
85
56
10.64

-


_




^_






65
41
13.62



c: Cost: t /ton or ore mined
ton
Exchan

-


-


_




__







-




Mixed
Media
Filtr.
255
63
11.98

-


„











165
50
16.61



PH
Adjust
40
31
5.89

-


	











35
28
9.30



Recycle
25%

—


-


_












_




50%

—


-


_












_




75%

_


-















_




100%

_


-















	




Activated
Carbon

_


-















_




Hydrogen
Peroxide

„


—












.


_




Chlorine
Dioxide

_.


_















_




Potassium
Parmang-

__


_








~



™"







Proem

_


„








^~



•~















Inactive


Presently
inactive
Presently
inactive


Presently
inactive


Presently
inactive




"Production
included
with Mine
3134
00
CJI

-------
             TABLE IX-4.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                        AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn « 2000 Ibs)
Typ« of Mini: B*» wid Precious
        Metals (Lead-Zinc)
p. 6 of 8
Mine Code -
3136-NV
Mine/Mill
3137-AZ
Mine/Mill
3138-CO
Mine
31 39-1 L
Mill
3140-NM
Mill
3141-TN
Mine
3141-TN
Mill
3142-UT
Mine/Mill
Ore
Production
(1000 •
tons/year)
126
93.2
98.2
NA
144
0
0
216

Water
Discharged
(MGD)
0
0
NA
0.87
0
0
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost (S1000I
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Flo ecu-
lation
-
-
-

-
-
-
-
Ozona-
lion
-

-
-
-
-
-
-
Alkat.
Chlorin-
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-

pH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-

-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-

Process
Control
-
-
-
-
-
-
-
-
Renwlci
Pilot scale
operation

Discharges
twice
yearly
Presently
inactive

Operation
closed



-------
                    TABLE 1X4.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                               AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
      Type of Mine: Base and Precious
              Metals (Lead-Zinc)
p. 7 of 8



Mine Code -
Location Type

3143-CO
Mine
t
3143-CO
Mill
4103-CO
Mine/Mill
UKA-WI
Mine



UKB-WI
Mine

UKC-TN
Mine/Mill

UKD-TN
Mine


UKD-TN
Mill



Ore
Production
(1000
tons/year)


60

60

0


NA



NA

NA

NA


NA




Water
Discharged
(MGD)


NA

0

NA


1.00



1.90

2.00

4.49


1.06

TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)

Second
Settling
a.

b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -

c.
a.
b. -
c.
a. 120
b. 22.5
c. —
a. '163
b. 27
c. —

a. 95
b. 20
c. —

Floccu-
lation


-






__



-

79
37.5
—
89
51
—

70
32


Ozona-
tion


-






_



_

155
62
—
310
103
—

98
44

c. Cost: t /ton or ore mined
Alkai.
Chlorin-
ation


-










_

100
71'
-
170
137
—

74
48

Ion
Exchan


-






_



_

3200
910
-
_


2200
610

Mixed
Media
Filtr.


-










	

330
75
-
630
110
_

208
58

PH
Adjust


-










__

43
35
-
55
50
_

38
29

Recycle
25%


-










_

26
21
-



17
18
"
50%


—










:

37
24
-



24
20
"
75%


_












50
28
-



32
22
~~
100%


_












63
32
-



41
26
	
Activated
Carbon


_







—




580
255
-



375
165
—
Hydrogen
Peroxide


	







"~




210
125
-



180
105
—
Chlorine


_







~




176
117
—



155
100
-
Potassium
Permang-


_







—




440
165




310
135
-
Process










—





















Operation
closed

Presently
inactive


Presently
inactive


Under-
ground
Mine
Under-
ground
Mine
00

-------
                   TABLE IX-4.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES

                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn » 2000 Ibs)
       Typ« of Mine:
B«o and Prtctouk

Metals (Lead-Zinc)
                                                                                                        p. 8 of 8
Min« Code -
Location Type
•
? (100)-TN
Mine/Mill
? (1021-CO
Mine/Mill






Or*
Production
(1000
tonj/yoar)
11
NA
NA







Water
Discharged
(MGD)
_______
0
NA






TREATMENT TECHNOLOGIES AND COSTS: a. Capital Co»t (S10QO)
b. Annual Cost ($1000)
c. Cost: ( /ton or or* mined
Second.
Settling
•
-






Ho eol-
ation
-






Ozona-
tlon
.
-






Alkal.
Chtorin-
ation
—
-






Ion
Exchan.
- '






Mixed
Media
Fillf.
=====
-







PH
Adjust.
-






Recyct*
25X
-






SOX
-






75X
-






100%
-
-






Activated
Carbon
-
-







lydrogtn
Pcroxid*
-
-







Jhlorln*
Dioxide
-
-







otmlum
>*rmanB-
anat*
-







Procen
Control
-






Remarks
— =







-p*
LO
y>

-------
                    TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                               AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
       Type of Mine:
Base and Precious
Metals (Gold-Silver)
                                                                                                       p. 1of 4
Mine Cod* -
Location Type
4101-NV
Mine/Mill
4102-CO
Mine
4102-CO
Mill
4104-WA
Mine/Mill
4105-SD
Mine
4115-UT
Mine/Mill
4116-AZ
Mine/Mill
4117-NV
Mine/Mill
Or*
Production
(1000
tons/year)
820
179
287
SB
1,560
145
NA
SO
Water
Discharged
(MGD)
0
1.00
0.36
0
3.04
Minima
NA
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: { /ton or ore mined
Second.
Settling
a.
b. -
c.
a. 96
b. 20
c. 11.17
a. 72
b. 17.8
c. 6.20
a.
b. -
c.
a. 145
b. 24.7
c. 1.58
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
70
32
17.88
60
28
9.76
-
82
43
2.76
-
-
-
Ozona-
tion
-
95
43
24.02
50
31
10.80
-
220
79
5.06
-
-
-
Alkal.
Chlorin-
atton
-
74
47
26.26
53
32
11.15
-
130
97
6.22
-
-
-
Ion
Exchan.
-
2100
600
335.20
1200
330
114.98
-

-
-
-
Mixed
Media
Filtr.
-
200
55
30.73
90
38
13.24
-
480
92
5.90

-
-
PH
Adjust.
-
36
28
15.64
30
24
8.36
-
48
44
2.82
-
-
-
Recycle
25%
-
-
9.5
16
5.57
-
-
-
-
-
50%
-
-
15
17
5.92
-
-
-
-
-
75%
-
- '
18
18
6.27
-
-
-
-
-
100%
-
-'
23
20
6.97
-
-
- -
-
-
Activated
Carbon
-
-
180
88
30.66
-

-
-
-
Hydrogen
Peroxide
-
-
152
95
33.10
-
-
-
-
-
Chlorine
Dioxide
-
-
122
91
31.70
-
-
-
-
-
Potassium
Permang-
anate
'-
-
195
108
37.63
-
-
-
-
-
Proceta
Control
-
-
-
-
-
-
-
-
Remarks


<»





OJ
-JO

-------
               TABLE IX-5.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                          AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn » 2000 Ibs)
Type of Mini: B«JB ind Precious
        Metals (Gold-Silver)
p. 2of 4
Mine Coda •
Location Type
4118-NV
Mine/Mill
4119-NV
Mine/Mill
4120-NV
Mine/Mill
4121-NV
Mine/Mill
4122-NV
Mine/Mill
4123-NV
Mine
4124-NM
Mine
4126-AK
Mine/Mill
Ore
Production
(1000
tons/year)
NA
NA
NA
0
0
NA
0
NA
Water
Discharged
(MGD)
0
Minimal
0
0
0
0
0
0
TREATMENT TECHNOLOGIES AND COSTS: ». Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floccu-
lation
-
--
-
-
-
-
-
-
Ozona-
tion
-
-
-
-
-
-
-
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust.
-
-
-
-
-
-
-
-
Recycle
25%
-
-

--
—
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-
100%
-
-
-
-
-
-
-
-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
- '
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-
-
-
-
-
-
-
-
Process
Control
-
-
-
-
-
-
-
-
Remark!



Presently
inactive
Presently
inactive

Temporarily
inactive


-------
              TABLE IX-5.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
Typ« of Mine: Base and Precious
        Metals (Gold-Silver)
p.3of4
Mine Coda -
Location Type
4127-AK
Mine/Mill
4128-NV
Mine/Mill
4129-CO
Mine
4129-CO
Mill
4130-NV
Mine
4131-NV
Mine/Mill
4401-ID
Mine
4401 and
4406-ID
Mills
Ore
Production
(1000
tons/year)
612 (m3)
730
NA
0.18-0.36
0
2,004
181
181 (4401)
108 (4406)
Water
Discharged
(MGD)
NA
0
NA
NA
0
0
0.21
0
TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. —
c.
a. 63
b. 17.1
c. 9.45
a.
b. -
c.
Floccu-
lation .
-
. -
-
-
-
-
55
28
15.47
-
Ozona-
tion
-
-
-
-
-
-
34
28
15.47
-
Alkal.
Chlorin-
ation
-
-
-
-
-
-
48
28
15.47
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
57
34
18.78
-
pH
Adjust.
-
-
-
-
-
-
27
23
12.71
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
-
-
-
75%
-
"-
-
-
-
-
-
-
100%
-
-
-
-
-
-

-
Activated
Carbon
-
-
-
-
-
-
-
-
Hydrogen
Peroxide
-
-
-
-
-
-
-
-
Chlorine
Dioxide
-
-
-
-
-
-
-
-
Potassium
Permang-
anate
-

-
-

-
-
-
Process
Control
-
-
-
-
-
-
- '
-
Remarks


Under
exploration
Under
exploration
Inactive


Wastewater
in combined
tailings
pond

-------
              TABLE IX-5. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
TVM of Mine: Baft and Precious
        Metali (Gold-Silver)
                                                                                                p.4»f4



Mine Code -

4402-CO
Mine


4403-ID
Mill


4404-CO
Mine


4406-ID
Mine

4407-ID
Mine


4408-MT
Mine

4409-ID
Mine


4410-IO
Mine




Production
(1000

74.4


198


407



108


684


74



NA


84





Water
Discharged
(MGD)

0.78


0.83


1.00



0


NA


0



NA


0


TREATMENT TECHNOLOGIES AND COSTS: a. Capital Cost ($1000)
b. Annual Cost ($1000)
c. Cost: t /ton or ore mined
Second.
Settling
a. 88
b.19.4
c. 26.08
a. 90
b.19.5
c. 9.85
a.96
b.20
c- 4.91
a.


c.
a.
b. -
c.
a.
b. -

c.
a.
b. —
c.
a.
b. -

c.
Floccu-
lation
68
30
40.32
69
30.5
15.40
70
32
7.86





—


_



—


_


Ozona-
tion
78
40
53.76
85
41
20.71
95
44
10.81





—


_



*~


_


Alkal.
Chlorin-
•lion
69
43
53.76

-

74
47
11.55





—


_






_


Ion
Exchan.

-

1900
510
257.58

-






—


_



~~


_


Mixed
Madia
Filtr.
165
50
67.20
175
53
26.77
200
55
13.51





—


_



~~


_


pH
Adjust.
35
27
36.29
34
28
14.14
38
28
6.88





—


_



~~


' —


Recycle
25%

-

16
17
8.59

-






—


_



~


_


50%

-

22
19
9.60

-






—


_



~~


_


75X

-

29
21
10.61

-






—


_



~


_


100%

-

38
24
12.12

-






•~


_



~


_


Activated
Carbon

-

320
133
68.18








~~


_



~


_


Hydrogen
Peroxide

-

180
105
53.03

-






~~


_



~


-



Chlorine
Dioxide

-

145
98
49.50

-






™"


-






-



Potassium
Permang-
anate

-

282
127
64.14

-






™" *


-



~


-



Process
Contra)

-


-


-






__


-






-





Remarks




























-------
                   TABLE IX-6.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES

                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                           p. i pf i
         Type of Mine:  Aluminum Ore
Mine
Code -
Location
Type
5101-AR
Mine
5102-AR
Mine






Ore
Production
(1000 ton
-------
          TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                      P. 1 of 8
Type of Mine:  Ferroalloy
Mine
Code -
Location
Type
6101-NM
Mill
6102-CO
Mill
6103-CO
Mine
6104-CA
Mine
61 05- NY
Mine/Mill
6106-OR
Mine/
..Swelter.
6107-AZ
Mine
6108-NV
Mine/Mill
Ore
Productior
(1UOO tons/
year)
6,283
15,430
2,425
705
11
1,322
361
NA
Vlater
01 s-
charqed
(MGD)
2.90
2.90
2.87
8.71
0
Minimal
.5.54
0
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AH!) COSTS: b. Annual Cost ftlOOO)
c. Cost: i/ton of ore mined
Second.
Settling
a. 140
b. 24
c. 0.38
a. 140
b. 24
c. 0.16
a. 140
b. 24
c. 0.99.
a. 230
b. 32
c. 4.54
a.
b. -
c.
a.
b. -
c.
a. 180
b. 28
jc. 7.76
a.
b. -
c.
Floc-
cula-
tion
82
43
0.68
82
43
0.28
82
43
1.77
97
77
10.92
-
-
90
56
15:51
-
Ozon-
ation
215
77
1.23
215
77
0.50
-
-
-
-
-
-
Alkal.
Chlor-
inatlon
130
94.0
1.5(
130
94.0
0.61
-
-
-
-
-
-
Ion
xchan.
4500
1260
20.0!
4500
1260
8.13
-
-
-
-
-
-
Mixed
Media
Filtr.
480
92
1.46
480
92
0.60
480
90
3.71
1071
182
25.82
-
-
750
140
.38.78
-
PH

47
40
0.64
47
40
0.26
47
40
1.65
69
71
10.07
-
-
60
55
15.24
-
Recycle
25X

32
23
0.15
-
-
-
-
-
-
SOX

47
28
0.18
-
-
-
-
-
-
75X

62
33
0.21
-
-
-
-
-
-

100X
2300
5.39
8..5S
80
38
0.24
-
-
-
-
-
-

'rocess
lontrol
-
-
-
-
-
-
. -
-
REMARKS









-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                     p, 2 of 8
Type of Mine:  Ferroalloy
Hine
Code -
Location
Type
6109-CA
Mine/MIT
6110-ID
Mine
6111-AK
Mine
6112-NC
Mine/Mill
6113-NH
Mine
6114-NV
Mine
61 15- CO
Mine/Mil
6116-SC
Mine/ Mil
Ore
Production
1000 tons/
year)
16
NA •
NA
ca.330
50
0
NA
NA
Hater
Dis-
charged
(MOO)
Minimal
NA
NA
NA
0
NA
NA
0
a. Capital Cost ($10UO)
TREATMENT TECHNOLOGIES AND COSTS: b. Annual Cost (»1000)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. "
c.
a.
b. "
c.
a.
b. "
1C.
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
Ion
xchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust,
-
-
-
-
-
-
-
-
Recycle
25%
-
-
-
-
-
•-
-
-
50%
-
-
-
-
-
• -
-
-
75%
-
-
-
-
-
-
-
-

100%
-
-
-
-
-
-
-
-

rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS

Exploratory
operations
underway
it
Temporarl ly
1nact1ve-under
exploratlor

Exploratory
operations
underway
Inactive


-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn « 2000 Ibs)
                                                                                   p. 3 of 8
Type of Mine:  Ferroalloy
Mine
Code -
Location
Type
6117-NV
Mine/Mill
6118-MN
Mine
6119-CO
Mine
6120-UT
Mine
6121-NV
Mine
6122-CA
Mine
6123-ID
Mine
6124-CA
Mine
Ore
Production
(1UOO ton:*
year) '
NA
NA
NA
0
NA
ca. 11
NA
ca.33
Hater
Dis-
charged
(MGD)
0
NA
NA
0
0
0
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost 01000)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
- ,;
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion

-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-

-
-
-
PH

-
-
-
-
-
-
-
-
R e c y c 1 e
25%
-
-
-
-
-
-
-
-
502
-
-
-
-
-
-
-
-
75%
-
- -
-
-
-
-
•• -
-

100*
-
-
-
-
-
-
-
-

'rocess
Control
-
-

-
-
-
-
-
REMARKS


Exploratory '
operations
Inactive


Inactive


-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                     p. 4 of 8
Type of Mine:  Ferroalloy
Mine
Code -
Location
Type
6125-CA
Mine/Hill
6126-10
Mine
61 27- ID
Mine
6128-CA
Mine
6129-NV
Mine
6130-NV
Mine
6131-CA
Mine/Mil
6132-UT
Mine/Mil
Ore
Production
1000 ton;*
year)
NA
NA
NA
HA
NA
HA
NA
NA
Water
Ois-
charqed
(MGD)
NA
Inter-
mittent
NA
NA
0
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AM!) COSTS: b. Annual Cost (*1000)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
:
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
xchan.
-
-
-
-
-
-
- '
-
Mixed
Media
Filtr.
-
-
-
-
-
'
-
-
pit
Adjust.
-
-
-
-
-
-
-
-
Recycle
25Z
-
-
-
-

-
-
-
50%
-
-
-
-
-
-
-
-
75?
-
-
-
-
-
-
-
-

100*
-
-
-
-
-
- -
-
-
'rocess
ontrol
-

-
-
-
-
-
-
REMARKS

Presently
Inactive
Exploration
underway
Inactive

Inactive
ii
ii '

-------
•£»
CO
                   TABLE IX-7. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                             AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
                                                                                               p. 5 of 8
          Type of Mine: Ferroalloy
Mine
Code -
Location
Type
6133-MT
Mine
6134-10
Nine
6135-CA
Mine
6136-UT
Mine
6137-UT
Mine
6138-CA
Mine
6139-CA
Mine
6140- CA
Mine
Ore
Production
(11)00 tonsy
year)
NA
NA
NA
NA
NA
NA
NA
NA
Hater
Dis-
charqed
(MGD)
NA
0
NA
NA
NA
NA
NA
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AH!) COSTS: b. Annual Cost (>1000)
c. Cost: {/ton of ore mined
Second.
Settling
a.
b. _
c.
a.
b. _
c.
a.
b. -
c.
a.
b. "
c.
a.
b. -
c.
a.
b. .
c.
a.
b. *
c.
a.
b. "
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
inatlon
-
-
-
-
-
"-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH

-
-
-
-
-
-
-
- -
Recycle
25%
-
-
-
-
-
-

-
50%
'.-
-
-
-
-
-
-
-
75%
-
-
-
-
-
-
-
-

100%
-
-
-
-
-
-
-
-

'rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS

Inactive


Inactive




-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                     P. 6 of 8
Type of Mine:  Ferroalloy
Mine
Code -
Location
Type
6141-CA
Mine
6142-CA
Mine
6143-CA
Mine
6144-CA
Mine
6145-CA
Mine/Mill
6146-CA
Mine/Mill
6147-CA
Mine
6148-NV
Mine/Mill
Ore
Production
(1000 ton
-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                    p. 7 of 8
Type of Mine: Ferroalloy
Mine
Code -
Location
Type
6149-ID
Mine/Mill
61 50- ID
Hill
6151-HT
Hill
6152-OR
Mine
6153-NV
Hill
6154-MT
Hill
6155-NV
Hill
6156-UT
Hill
Ore
Production
1000 tons/
year)
NA
NA
0.002
1
NA
NA
0.028
NA
Viator
Ois-
charqed
(MGD)
NA
NA
Minimal
NA
NA
NA
0
NA
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (>1000)
c. Cost: t/ton of ore rained
Second.
Settling
a.
b. -
c.
a.
b. -
c.
3.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
Floc-
cula-
tion
-
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-"
-
Ion
xchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
-
-
-
-
PH
Adjust
-

-
-
-
-
- -
-
Recycle
25%
-
-
-
-
-
-
-
-
SOX
-
**
-
-
-
-
-
-
75X
-
-
-
-
-
-
-
-

100X
-
-
-
-
-
-
-
-
rocess
ontrol
*•
-
-
-
-
-
-
-
REMARKS
Inactive
ii







-------
         TABLE IX-7.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                    AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                     p. 8 of 8
Type of Mine:  Ferroalloy
Mine
Code -
Location
Type
6157-NV
Mill
6158-CA
Mill
6159-NN
Mill
6160-TX
Mill
6161-NM
Mill
6162-SC
Mill
6163-CA
Mill

Ore
Production
(1000 ton;,
year)
0.055
(cone.)
0.496
0.110
(cone.)
NA
8
(cone.)
NA
0.055
(cone.)

Hater
Dis-
charqed
(MGD)
0
0
0
0
0
0
0

a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AM!) COSTS: b. Annual Cost (J.1000)
c. Cost: tf/ton of ore mined
Second.
Settling
a.
b.
c.
a.
b. -
c.
a.
b. -
c.
a. .
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
Floc-
cula-
tion
-
-
-
-
-
-
-

Ozon-
ation
-
-
-
-
-
-
-

Alkal.
Chlor-
ination
-
-
-
-
-
-
-

Ion
Exchan.
-
-
-
.-
-
-
-

Mixed
Media
Filtr.
-
-
-
-
-
-
-

PH
Adjust.
-
-
-
-
-
-
-

R e c y
25%
-
-
-
-
-
-
-

50%
-
-
-
-
-
-
-

c 1 e
75%
-
-
-
-
-
-
-


100%
-
-
-
-
-
-
-


Vocess
Control
-

-
-
•-
-
-

REMARKS

'reduction based
on mill capacity







-------
                    TABLE IX-8. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES

                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn - 2000 Ibs)
                                                                                              P. 1 of 1
           Type of Mine:  Mercury
en
ro
iline
Code -
Location
Type
9201 -CA
Mine/Mill
9202-NV
Mine/Mill






Ore
Production
1000 tons/
year)
30
17.5






Hater
Dis-
charqed
(MGO)
0
0

.




a. Capital Cost ($1000)
TREATMENT TECHNOLQfilES AMD COSTS: b. Annual Cost (>1000)
c. Cost: i/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
a.
b.
c.
Floc-
cula-
tion
-
-






Ozon-
ation
-
-






Alkal.
Chlor-
1 nation
-
-






Ion
xchan.
-
-






Mixed
Media
Filtr.
-
-






PH
Adjust
-
-






Recycle
25X
-
-






50%
-
-






75%
-
-







100X
-
-







"rocess
ontrol
-
-






REMARKS









-------
                   TABLE IX-9.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
           Type of Mine:  Titanium
                                                                                          p. 1 of 1
4J1
00
rline
Cotie-
Location
Type
9905-NY
Mine
99U6-FL
Mill
9907-FL
Mill
9908-FL
Mine/Mill
9909-FL
Hi ne/Mill
9910-HJ
Mill
9911-NJ
Mine/Mill

Ore
Production
(1000 tonV
year)
464 .
7260
.7260
NA
HA
6600
NA

Mater
Dis-
charged
(MGD)
0.70
6.84
1.63
NA
NA
3.77
0

a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AN!) COSTS: b. Annual Cost ()1000)
c. Cost: t/ton of ore mined
Second.
Settling
a. 84
b. 19.0
c. 4.09
a. 200
b. 30.5
c. 0.42
a.H4
b. 21.8
c. 0.30
a.
b. -
c.
a.
b. -
c.
a. 155
b. 26
c. 0.39
a.
b. "
c.
a.
li.
c.
Floc-
cula-
tion
66
30
6.47
90
66
0.91
75
34
0.47
-
-
85
46
0.70
-

Ozon-
ation
72
40
8.62
455
137
1.8<
138
56
0.77
-
-
270
92
1.39
-

Alkal.
Chlor-
ination
-


-
-

-

Ion
Exchan.
-
8100
2210
30.44
2800
800
11.02
-
-
5300
1410
21.36
- -

Mixed
Media
Filtr.
160
47
10.13
900
145
2.00
270
66
0.91
-
-
560
100
1.52
-

PH
Adjust
33
27
5.82
63
63
0.87
40
33
0.45
-
-
50
45
0.68
-


25«
-
50
28
0.39
22
20
0.28
-
- .
36
23
0.35


R e c y
sot

78
40
0.55.
30
23
0.32
-
-
52
30
0.45
-

c 1 e
75X
-
110
50
0.69
40
27
0.37
-
-
70
37
0.56
-


loot
-
140
58
0.80
52
29
0.40
-
-
90
42
0.64
-

'rocess
Control
-
-
-
-
-
-
-

REMARKS



Inactive
ii

Inactive


-------
          TABLE IX-10.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                         AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                                      p. 1 of 5
 Type of Mine:
Uranium, Radium,
Vanadium
Mine
Code-
Location
Type
9401 -NM
Mine
9401 -NH
Hill
9402-NM
Mine 35. 3f
9402-NM
Mine 17,22
24,30,33
9402-NM
Mill
9403-UT
Mill
9404-NM
Mill
9405-CO*
Mill
Ore
Production
(1UOO ton;*
year)
750
1,270
lj!25

2409
274*
2,490
439*
Hater
Uis-
(MGO)
0.85
(1.86)**
3.00
0.00
(1.94)**
(0.41)**
(1.38)**
1.00
a. Capital Cost ($101000)
c. Cost: (/ton of ore mined
Second.
Settling
a. 90
b. 19.5
c. 2.60
a. 120
b. 22
c. 1.72
a.
b. -
c.
a.
b. -
c.
a. 120
b. 22
c. 0.91
a. 74
b. 18
c. 6.57
a. 110
b. 21
c. 0.84
a. 96
b. 20
c. 4.55
Floc-
cula-
tion
68
30
4.00
-
-
-
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
Inatlon
-
-
-"
-
-


-
Ion
xchan.
1900
510
68.00
-
-
-
-
-
-
-
Mixed
Media
Filtr.
180
52
6.93
-

-

-
-
-
pH

35
28
3.73
-
-
-
42
35
1.4S
30
25
9.12
39
31
1.24
37
29
6.61
Recycle
25X
-
-
-
-

,--
- ,
-
BOX
-
-
-
-
-
-
-
-
75X
-
-
-
-
-
-
-
-

loot
-
60
30
2.3«
-
-
60
30
1.2
25
20
7.30
47
27.5
1.10
40
26
5.92

'rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS

No point
discharge

0 Discharge
Ref. p. 111-71
Mill receives
ore from
other mines
No point
discharge
ii .
Single
discharging
uranium mill
     indicates production obtained from mill
(  )** for 0 discharge mills, flow Indicated =
                       capacity - assume 365 working days/year.
                       volume discharged to treatment or recycle system.
See also page 5 for off site evaporation
pond design.

-------
in
                   TABLE IX-10.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                              AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
          Type of Mine: Uranium, Radium,
                   Vanadium
                                                                                                   p. 2 of 5
Mine
Code -
Location
Type
9407-WY
Mill
9409-WY
Mine
9409-WY
Mill
9410-HY
Mine
9411-WY
Mill
9413-WY
Mine
9413-HY
Mill
9419-TX
Mill
Ore
Production
(1UOO tons;
year)
724
500
1,086
0
357
595
603
1,046
ii ill
* Indicates proc
( )** for 0 discharc
Mater
Dis-
charged
(MGD)
(1.11)**
0.50
(1.22)**
2.30
(0.39)**
0.36
(2.04)**
4.65
uction o
e mills,
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (»1000)
c. Cost: 
-------
CTt
                    TABLE IX-10.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                                AND SUBSEQUENT COST PER TON OF ORE MINED {1 tn = 2000 Ibs)
                                                                                                 p. 3 of 5
          Type of Mine: Uranium, Radium,
                    Vanadium
i-llne
Code -
Location
Type
9422-CO
Hill
9423-HA
Mill
9425-HY
Mill
9427-HY
Mill
9430-UT
Mill
9437-NM
Mine
9442-HY
Mill
9443-NM
Mine
Ore
Production
[1UQO tons/
year)
165
183
NA
346*
NA
96
563*
NA
* indicates piro
( )** for 0 dischar
Hater
Uls-
charqed
(MOD)
NA
(0.14)**
0
(0.33)**
NA
5.03
(0.28)*
1.45
{Tuctioh" (
ge mills
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost 01000)
c. Cost: i/ton of ore mined
Second.
Settling
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 70
b. 17.5
c. 5.06
a.
b. -
c.
a. 180
b. 28
c. 29.17
a. 66
b. 17.4
c. 3.09
a.
b. -
c.
)bta1ned~F
flow 1nd
Floc-
cula-
tlon
-
-
-
-
-
90
54
56.2
-
-
rom mill
icated =
Ozon-
atlon
-
-
-
-
-
-
-
capaclt
volume
Alkal.
Chlor-
1 nation
-
-
-
-
-
-
-
-
y - assu
dlscharg
Ion
xchan.
-
-
-
-
-
6500
1710
L781.2
-
me 365~w
ed to tr
Mixed
Media
Filtr.
-
-
-
-
-
700
130
-
-
PH
Adjust.
-
-
-
29
24
6.94
-
58
52
54.17
-
-
Recycle
25X
-
-
-
-
-
-
-
-
SOX
-
-
-
-
-
-
-
-
75X
-
-
-
-
-
-
-
-

100X
-
-
-
23
19
5.49
-
-
-
-

rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS
Lined e vapor.
pond: no
discharge
No point
discharge

No point
discharge


. No point
discharge
Under
development
orklng days/year.
eatment or recycle system.

-------
          TABLE IX-10. COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                      AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                        p. 4 of 5
Type of Mine:  Uranium, Radium
          Vanadium
i-11 ne
Code -
Location
Type
9445-NM
Mill
9446-NM
Mill
9447-UT
Mine
9447-UT
Mill
9449-WY
Mine
9449-WY
Mill
9450-WY
Mill
9452-NM
Mine
Ore
Production
(1UOO tonv
year)
540
607*
262
273*
750
NA
483*
500
VJater
Dis-
charged
(MGD)
0
(0.63)**
minimal
(0.31)**
9.8
0
(0.55)**
1.80
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AMD COSTS: b. Annual Cost (»1000)
c. Cost: t/ton of ore mined
Second.
Settling
a.
b. -
c.
a. 80
b. 18.8
c. 3.10
a.
b. -
c.
a. 70
b. 17.5
c. 6.41
a. 240
b. 34
c. 4.53
a.
b. -
c.
a. 80
b. 18.5
c. 3.83
a.
b. -
c.
Floc-
cula-
tlon
-
-
-
-
98
80
10.67
- •
-
-
Ozon-
ation
-
-
-
-
-
-"
-
-
Alkal.
Chlor-
1 nation
-
-
-
-
-
•-
-
-
Ion
Exchan.
-
-
-
-
-
-
-
-
Mixed
Media
Filtr.
-
-
-
-
1200
190
25.33
-
-
-
PH
Adjust.
-
35
27
4.45
• -
-
71
78
10.4
-
34
26
5.38
-
Recycle
25%
-
-
-
. -
- -
- .
-
-
50%
-
-
-
-
-
-

-
75%
-
-
-
• -
-
-
-
-
100%
-
-
-
-

-
29
22
4.55
-
5rocess
Control
-
-

-
'-
-

-
REMARKS

I No point
' discharge

, No point
discharge
»

I No point
(discharge

*i w* !nd1«aif Producti°n obtained from mill capacity - assume 365 working days/year.
( )** for 0 discharge mills, flow indicated = volume discharged to treatment or recycle system.

-------
                         TABLE IX-10.  COST COMPARISON FOR VARIOUS TYPES OF TREATMENT TECHNOLOGIES
                                      AND SUBSEQUENT COST PER TON OF ORE MINED (1 tn = 2000 Ibs)
                                                                                                             p. 5 of 5
          Type of Mine: Uranium, Radium,
                     Vanadium
Mine
Code -
Location
Type
9452-NM
Mill
9456-WA
Mill
New Mine
"A" - MM
Mine
9460-WY
Mine
9460-WY
Mill
9463-TX
Mill
Union
Carbide-
Rifle Mi
9405-CO
Mill
Ore
Production
;iooo tons/
year)
1,448*
764*
0
300
NA
NA
NA
439
Hater
Dis-
charqed
(MOD)
(1.04)**
(0.63)**
0.87
0.87
0
NA
NA
1.00
a. Capital Cost ($1000)
TREATMENT TECHNOLOGIES AKH COSTS: b. Annual Cost ($1000)
c. Cost: i/ton of ore mined
Second.
Settling
a. 95
b. 20
c. 1.38
a.
b. -
c.
a.
b. -
c.
a. 90
b. 19.5
c. 6.50
a.
b. -
c.
a.
b. -
c.
a.
b. -
c.
a. 23, 900
b. 2,550
c. 580
Floc-
cula-
tion
-
-
-
68
30
10. Of
-
-
-
-
Ozon-
ation
-
-
-
-
-
-
-
-
Alkal.
Chlor-
i nation
-
-
-
-
-
-
-
-
Ion
xchan.
-
-
-
1900
510
170.00
-
-
-
-
Mixed
Media
Filtr.
-
-
-
180
52
17.33
-
-
-
-
PH
Adjust.
37
30
2.07
-
-
36
28
9.33
-
-
-
-
Recycle
25%
-
-
-
-
-
-
-
-
50%
-
-
-
-
-
'-
-
-
75%
-
-
-
-
-
-
-
-

100%
40
26
1.79
-
-
-
-
-
-
-

rocess
ontrol
-
-
-
-
-
-
-
-
REMARKS


No production
at present




Evaporation
pond design
JO
         ( )**  for 0 discharge mills, flow indicated = volume discharged to treatment or recycle system.

-------
           TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l)
Waste Type by Segment
Iron Mining and Milling
(1101,1105,1109,1120)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds - Settled
Solids
Copper Mining and Milling
(2101,2104,2118,2119,
2120,2121,2122,2126,
2139,2147,2164)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds - Settled
Solids
Arsenic ,
Range

<0.0005-
0.161
<0.0005-
0.005
<0.0005-
0.010
<0.0005-
0.550

0.002-
0.050
<0.002-
0.0155
0.006-
0.055
0.0026-
0.065
Barium
Range

0.025-
0.715
0.105-
0.51
0.13-
0.39
0.02-
0.41

0.058-
0.62
0.032-
0.11
0.04-
2.8
<0.001-
2.0
Cadmium
Range

<0.008-
0.016
<0.008
<0.008-
0.021
<0.008

<0.008-
0.17
<0.008-
0.071
<0.008-
0.022
<0.008-
0.039
Chromium
Range

<0.001-
0.003
<0.001-
0.009
<0.001-
0.076
<0.001-
0.013

<0.001-
<0.04
<0.001-
0.052
<0.001-
0.057
<0.001-
0.110
Lead
Range

<0.084
<0.084
<0.084
0.084-
0.112

<0.06-
0.840
<0.08-
0.084
<0.084-
0.840
<0.06-
0.084
Mercury
Range

<0.0005-
0.001
<0.0005-
0.001
<0.0005-
0.001
<0.0005-
0.001

<0.0005-
0.002
<0.0005-
0.001
< 0.0005-
0.001
<0.0005-
0.002
Selenium
Range

0.001-
0.009
0.003-
0.009
0.003-
0.015
0.001-
0.021

0.0005-
0.079
0.006-
0.056
0.006
0.104
<0.001-
0.105
Silver
Range

<0.002-
0.008
<0.002-
<0.003
<0.002-
0.01
<0.002

<0.002-
0.024
<0.002-
0.010
<0.002-
0.012
<0.002-
0.021
-ft.
Ul

-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Lead/Zinc Mining and Milling
(3104,3106,3107,3110,
3113,3122,3123,3126)
Mine Waste Rock
Fresh Tailings
Tailing Ponds - Settled
Solids
Mine Water Ponds •
Settled Solids
Gold/Silver Mining and Milling
(4101,4105,4119,4121,
4402, 4407)
Mine Waste Rock
Low Grade Ore
Fresh Tailings
Tailing Ponds - Settled
Solids
Aluminum Mining (5101)
Mine Water Ponds •
Settled Solids
Mine Waste Rock
Arsenic
Range

<0.0041-
0.047
<0.002-
<0.023
<0.002-
0.043
<0.0044-
0.008

<0.0005-
0.027
<0.002-
0.103
0.004-
0.017
0.007-
0.369

<0.025
<0.025
Barium
Range

0.052-
0.270
0.051-
0:665
0.016-
1.68
0.47-
1.5

0.095-
2.90
0.160-
3.25
0.009-
1.73
<0.001-
1.90

0.15-
1.19
0.34
Cadmium
Range

0.039-
0.650
<0.008-
0.17
<0.008-
0.36
0.013-
0.040

<0.003-
0.098
<0.008-
0.087
<0.008-
0.170

-------
TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Molybdenum Mining and
Milling (6101, 6102, 6103)
Low Grade Ore
Mine Waste Rock
Fresh Tailings
Tailing Ponds - Settled
Solids
Wastewater Treatment
Sludge
Mine Water Pond -
Settled Solids
Tungsten Mining and Milling
(6104, 6105)
Mine Waste Rock
Tailing Pond - Settled
Solids
Dry Tailings
Mine Water Pond Settled
Solids
Arsenic
Range

<0.005-
<0.006
<0.005
<0.0005-
0.019
<0,005-
0.017
0.026
0.048

<0.001-
<0.002
0.0218-
0.075
<0.001-
0.02
<0.002-
<0.003
Barium
Range

0.058-
0.155
0.08-
0.19
0.14-
0.27
0.09-
6.2
0.039
0.74

0.22-
0.4
0.395-
0.59
0.2-
0.4
0.31-
0.38
Cadmium
Range

<0.008
<0.008
<0.008
<0.008
0.064
<0.008

0.015-
0.02
0.017-
0.027
<0.01-
0.01
0.011-
0.015
Chromium
Range

0.001-
<0.002
<0.001-
0.004
<0.001-
0.01
<0.001-
0.018
0.21
0.12

<0.001-
0.07
0.0085-
0.032
<0.04
<0.001-
0.002
Lead
Range

<0.084
<0.084
<0.084
<0.084-
0.19
<0.084
<0.084

<0.05-
<0.084
<0.06-
<0.084
<0.05
<0.084
Mercury
Range

<0.0005
<0.0005
<0.0005
<0.0005-
0.0018
<0.0005
<0.0005

<0.0005
<0.0005-
0.0005
0.0001-
0.0004
<0.0005-
<0.0018
Selenium
Range

0.004-
0.018
0.002-
0.010
0.002-
0.043
0.003-
0.023
0.055
0.006

0.0199-
0.052
0.0448-
0.173
0.041-
0.046
0.0048-
0.0212
Silver
Range

<0.002
<0.002
<0.002-
0.004
<0.002-
0.015
0.03
0.011

<0.002-
<0.01
<0.002-
<0.01
<0.01
< 0.002-
0.002

-------
          TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Vanadium Mining and Milling
(6107)
Mine Waste Rock
Mine Water Pond - Settled
Solids
Tailing Pond - Settled
Solids
Mill Wastewater Ponds -
Settled Solids
Nickel Mining, Milling and
Smelting (6106)
Mine Waste Rock
Low Grade Ore
Mine/Smelter Wastewater
Settled Solids
Mercury Mining and
Milling (9202)
Fresh Tailings
Tailing Ponds - Settled
Solids
Mine Waste Rock
Arsenic
Range

< 0.001
<0.001
<0.001
<0.001-
0.54

0.02
<0.001
0.001

0.17
0.26
0.1
Barium
Range

0.13
0.15
1.31-
1.69
0.03-
0.3

0.1
<0.1
0.2

0.76
0.76
0.76
Cadmium
Range

<0.01
0.04
<0.01
<0.01-
0.37

<0.01
<0.01
<0.01

<0.005
<0.005
0.01
Chromium
Range

<0.04
<0.04
<0.04
<0.04-
1.9

<0.04
<0.04
<0.04

<0.05
<0.05
0.06
Lead
Range

<0.05
<0.05
<0.05
<0.05

<0.05
<0.05
<0.05

0.14
0.11
<0.1
Mercury
Range

<0.0001
<0.0001
<0.0001-
0.0002
<0.0001-
0.0036

<0.0001
<0.0001
0.0001

0.0019
0.041
0.14
Selenium
Range

<0.001
0.008
<0.001-
0.004
0.001-
0.03

0.001
<0.001
0.001

<0.015
X0.015
<0.015
Silver
Range

<0.002
0.004
<0.002-
0.046
0.008-
0.25

<0.002
0.032
0.004

<0.001
<0.001
<0.001
ro

-------
        TABLE IX-11. SUMMARY OF RESULTS FOR RCRA, EP ACETIC ACID LEACHATE TEST (all values as mg/l) (continued)
Waste Type by Segment
Uranium Mining
(9402, 9403, 9404, 9405,
9408,9409,9411,9412,
9423, 9447, 9451, 9455,
9460)
Mine Waste Rock
Low Grade Ore
Mine Water Ponds -
Settled Solids
Titanium Dredge Mining
and Milling (9906)
Fresh Tailings
Mine Water Pond -
Settled Solids
Mill Wastewater Pond -
Settled Solids
Antimony Mining and
Milling (9901)
Fresh Tailings
Tailing Pond - Settled
Solids
Arsenic
Range


-------
            Figure IX-1. SECONDARY SETTLING POND/LAGOON - TYPICAL LAYOUT
WASTEWATER
  INFLUENT
 PLAN
                           2.5:1
                                                        2.5:|
2.5:|
          EFFLUENT
            PIPE
 SECTION
                      WASTEWATER INFLUENT
                                                     •OUTLET PIPE
                                                                         EFFLUENT PIPE
               EXISTING
               GRADE

-------
               COST IN THOUSAND  OF DOLLARS
m



m
I
-O

m
w
 H H
 - m

 o ***
39
a


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o

fO
                                                              a*

                                                              |

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                                                             O30
                                                             z m


                                                            33 m
                                                            <30
                                                            m _,
                                                            W30
                                                              m
                                                              Z
                                                              H
                                                              CO
                                                              rn

                                                             CO
                                                             rn

                                                             H




                                                             O

-------
     Figure IX-3.  ORE MINE WASTEWATER TREATMENT SETTLING PONDS - LINING

              COST CURVES
 1000
tr
<
o
o

o
z

CO

o
X
H
to  10
o
o
                                                      7
      O.I
      1

WASTEWATER  FLOW - MGD
10
                                                                  30
                     NOTE: POND LINER USED

                           FOR MINE 9405 ONLY
                         FIRST ISSUE: 2/29/80

-------
          Figure IX-4. FLOCCULANT (POLYELECTROLYTE) PREPARATION AND FEED-FLOW SCHEMATIC
HOPPER
                FEEDER
                      WETTING CHAMBER
                         MIXER
 WATER    +
 SUPPLY
      MIXING TANK
                                           METERING
                                             PUMP
                                            WASTEWATER
                                               INFLUENT
               MIXER


              EFFLUENT,
                            AGING TANK
MIXING TANK

-------
               COST IN MILLIONS OF DOLLARS

                    -              b
                                                 p
                                                 o
0

m


I
rn
JO
> m

i I
w rn
z ^
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m '  P
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                                                      cog
                                                       m
                                                       r-
                                                       rn
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                                                       3d
                                                       O



                                                       m

-------
                  Figure IX-6. OZONE GENERATION AND FEED FLOW SCHEMATIC
      COMPRESSOR
                 COOLING WATER
                    TO WASTE
AIR
     FILTER
    SILENCER
                           COMPRESSOR
                           AFTER COOLER
                 COOLING
                  WATER
                  SUPPLY
REFRIGERANT
  COOLING
  SYSTEM
                                                                    WASTEWATER
                                                                       INFLUENT
                  ELECTRIC
                POWER SUPPLY-
                COOLING
                WATER
                                      1
1
                                OZONE
                              GENERATOR
T
        COOLING
        WATER

           OZONE
          CONTACTOR
                                                                       TREATED
                                                                       EFFLUENT
                                          DESSICANT
                                        DRYING SYSTEM

-------
COST IN THOUSAND OF DOLLARS
                                           00
m


CO
m
o

ro

ro


I
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                                                 t
                                                 3



                                                 §
                                                 O
                                                 33
                                                 m
                                                 m




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                                                 1
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                                                 33


                                                 33


                                                 s
                                                 2
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                                                 H


                                                 8

                                                 §
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                                                 O
                                                 m
                                                 2
                                                 m
                                                 33

                                                 §


                                                 i

                                                 8P
                                                 •n
                                                 m
                                                 m
                                                 O
                                                 m
                                                 o
                                                 37


                                                 m
                                                 V)

-------
          Figure IX-8. ALKALINE-CHLORINATION FLOW SCHEMATIC
 CAUSTIC
  SODA
STORAGE
  TANK
   RAW
 WASTES
               PUMP
PUMP
                                       MIXER
                      MIXING TANK
 SODIUM
  HYPO-
CHLORITE
STORAGE
  TANK
               TREATED
               EFFLUENT

-------
                  CAPITAL  COST IN THOUSAND OF DOLLARS


                              5              8
                                                  8
                                                  O
m
<
w
m
o

IN3
X.
ro

     f-

     
-------
                                 Figure IX-10. ION EXCHANGE FLOW SCHEMATIC
                                             ACID
                                         STORAGE TANK
  CAUSTIC
STORAGE TANK
CO

£. 	 rn
RAW ^ ' ^r"* ' '
nAg ff\ ^ j
WASTE &L i, v v ! 1
T T TT li ,
v ^
* *
J Wt
j \ ' " '
' CATION l DEGASIFIER ! ANION
EXCHANGERS EXCHANGE
i 	 _.._! i ,
I
r
i
WASTE WASTE
STORAGE STORAGE
TANK TANK
'.'•'. - •«.___ 	 _• 1L .
_J


||! IJL '
i TT TT
i i
! ! . !
1 MIXED BED
RS EXCHANGERS
it i v
r | 	 * TREATED
j EFFLUENT
WASTE
STORAGE
TANK
                                                                                   .TREATED
                                                                                  "* WASTES
                                                           NEUTRALIZATION
                                                                TANK

-------
   Figure IX-11.  ORE MINE WASTEWATER TREATMENT ION EXCHANGE COST CURVES
COST IN MILLIONS OF DOLLARS _
— 0 C
b b t

















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                          WASTEWATER  FLOW -  MGD
100.0
(I) WASTEWATER SAMPLE ANALYSIS  NOT  INCLUDED
                                               REVISED  2/29/80
                                 474

-------
                    Figure IX-12. GRANULAR MEDIA FILTRATION PROCESS FLOW SCHEMATIC
   WASTEWATER
     INFLUENT
01
                   SETTLED
                    WATER
                   RECYCLE
                    PUMP
       GRAVITY  FILTER (GRANULAR MEDIA)
                                                                    TREATED
                             BACKWASH
                           WASTEWATER
                              BASIN
BACKWASH
  PUMP
                            SOLIDS TO
                              DECANT
                                                     CLEAR
                                                      WELL
                                                                    EFFLUENT

-------
Figure IX-13.   ORE MINE WASTEWATER TREATMENT GRANULAR MEDIA
           FILTRATION PROCESS COST CURVES
COST IN MILLIONS OF DOLLARS
b - b












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WASTEWATER FLOW - MGD
IOO.O
                  (I) WASTEWATER  SAMPLE
                     ANALYSIS  NOT INCLUDED
                          REVISED 2/29/80
                                 476

-------
                         Figure IX-14. pH ADJUSTMENT FLOW SCHEMATIC
       HYDRATED
         LIME
  FEEDER LVWA
WATER
SUPPLY"
       FEED
       PUMP
D   „
D
                        A
           I     I    3
              LIME SLURRY
                  TANKS
                                                    MIXER
             WASTEWATER
               INFLUENT
                                           NEUTRALIZED
                                             WASTES
                                            MIXING TANK

-------
                           VLV
                COST IN THOUSAND OF  DOLLARS
     p
     b
                                       o
                                       o
  m —

  1
  m
  o


   i
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;o
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-------
                COST IN THOUSAND OF DOLLARS
                     o
                     o
                                                   o

                                                   §
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                                                         33

                                                         m
                                                         00

-------
                          Figure IX-17. WASTEWATER RECYCLE FLOW SCHEMATIC
00
         RECYCLE  FLOW TO MINING
         AND  MILLING  OPERATION
         WASTEWATER
          INFLUENT
                                           X
                                        D
   NOTE:
X  NUMBER OF PUMPS DEPENDS ON THE
   RECYCLE FLOW
               EFFLUENT TO
             -^RECEIVING
               STREAM
                                PUMP STATION
                                  WET  WELL

-------
Figure IX-18.  ORE MINE WASTEWATER TREATMENT RECYCLING CAPITAL COST
          CURVES
                       I.O                K>.0

                     WASTEWATER FLOW ~ MGD
IOO.O
                               481

-------
COST IN MILLIONS  OF DOLLARS
                                           •n

                                           t
                                           3
                                           P

                                           O
                                           33
                                           m
                                           m
                                           I
                                           m

                                           I
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                                           rn
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                                           r;
                                           2
                                           O
                                           >

                                           2
                                           C
                                           O

                                           S
                                           s
                                           30
                                           m
                                           CO

-------
                     Figure IX-20. ACTIVATED CARBON ADSORPTION FLOW SCHEMATIC
JO
GO
          r~
             WASH WATER DRAIN
WASTEWATER
 INFLUENT
                       PUMP
                      STATION
D
              EXISTING
              BPT POND
                                 . i
                                       BY PASS
                  STAGE
T
          o r»d

        STAGE
                                               WASH
                                              WATER
                                              PUMPS
                                                D
CLEAR
WATER
 WELL
                                                        BY-PASS
                                                                  —**•          \
                                         ACTIVATED CARBON FILTERS

-------
COST IN THOUSANDS OF DOLLARS

_ o
0 0
b
0.1
WASTEWATER
i°
i
s
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6
b
0


























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-------
            COST  IN THOUSAND OF DOLLARS
   p
   b
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                   O
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O
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m
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-------
                     Figure IX-23. CHEMICAL OXIDATION - HYDROGEN PEROXIDE FLOW SCHEMATIC
    AIR
 COMPRESSOR
INFLUENH
oo
           OXIDATION BASINS
                                                  — pH ADJUSTMENT
                                                                TO
                                                              DI5CHAR6.E
SLUDGE TO
DISPOSAL
                  \\\\ sn*. • \\   //

                    SULFURIC
                      ACID
FERROUS SULFATE
   DISSOLVER
          FROM
         (TANK
          TRUCK
                                                                          HYDRO6EN PEROXIDE
                                                                           STORAGE TANK

-------
                       LW?
             COST IN  THOUSAND OF DOLLARS
m
i
m
-n
n
O
^
i

2
o
o

-------
            COST IN  THOUSAND OF DOLLARS
C/5
m
m  _
o
 i
o

-------
                      Figure IX-26.  CHEMICAL OXIDATION-CHLORINE DIOXIDE FLOW SCHEMATIC
  FROM

  TANK
  TRUCK
-P.
oo
                  cio
                STORAGE

                 TANK
                          METERING

                           PUMP
                                                DILUTION
                                      EJECTOR
       R  /
I FLUENT—7
                                                                    IN-LINE

                                                                    FILTER
                             izu
                              CH
                                         -TO DISCHAR6E
                                                    CONTACT TANK

-------
                         Ubl7
           COST  IN  THOUSANDS  OF DOLLARS
m


1
m
;o
o
o

-------
                          Ibfr
            COST IN THOUSANDS OF DOLLARS
m
I
m
o
I

-------
         Figure IX-29. CHEMICAL OXIDATION-POTASSIUM PERMANGANATE FLOW SCHEMATIC
FROM
TANK
TRUCK
HzO
       STORAGE
        TANK
                   DISSOLVER
                                  D
                                       OXIDATION BASINS
                                                                 CLARIFIERS
                                                                                       TO
                                                                                     DISCHARGE
                                                                     SLUDGE
                                                                       TO
                                                                     DISPOSAL

-------
                COST  IN THOUSANDS  OF  DOLLARS
    p
    b
                        O
                        o
          o
          o
          0
rn

1
m
71
5
    p
    o
\
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    33 5
    C3 ?
    m "•
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   33 m
   m _
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   53
                                                                 m Z

                                                                 > r-

-------
            COST IN THOUSANDS OF DOLLARS
H
m


1
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                                                    3


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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  (includinq
      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-
t1 on •

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
21; u     • 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
                                 49 5

-------
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.

The   BAT   effluent  limitations  guidelines  were  proposed  on
June 144,  1982  (47 FR 25682) and comments were requested from the
public.    AFter  reviewing  over  50   individual  submissions  of
comments   and  data,  the  Agency concluded that the BAT effluent
limitations and guidelines should be finalized as proposed.   The
rationale  for the Agency's selection  of BAT effluent limitations
is  summarized below  in this section.

SUMMARY OF BEST AVAILABLE TECHNOLOGY

Zero discharge  limitations  are  established  for   the  following
subcategofies and subparts:

     Iron  Ore
           Mills  in the Mesabi  Range
     Mercury Ore
           Mills
     Cu,  Pb, Zn, Au,  Ag,  and  Mo  Ores
           Mills  using the cyanidation  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:
 Subcateqory and Subpart
Nonconventional
  Pollutants
  Controlled
                                                       Toxics
                                                   Controlled
      Iron Ore
        Mine Drainage
        Mills (physical methods)
Fe (dissolved)
Fe (dissolved)
                               496

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Subcateqory and Subpart
   Subcateqory and Subpart
Nonconventional
  Pollutants
  Controlled
    Toxics
Controlled
  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,
     and  Mo Ores
       Mine Drainage (not
         placer mining)
       Mills (froth
         flotation)
Fe, Al
COD, Ra226 (dis-
solved) Ra226
(total), u .
Fe

Fe
    Zn
                     Hg
    Zn
                     Cd, Cu, Zn
                     Cd, Cu, Zn
                     Cd, Cu, Zn,
                     Pb, Hg,
                     Cd, Cu, Zn,
                     Pb, Hg,
                                    497

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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
    mq/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
     mq/1
 2.0
 2.0(1.0)**
 2.0

 10  (pCi/1)
 30  (pCi/1)
  4
        1.0
        1.0(0.5)**
        1 .0
        500
        3 (pCi/1)
        10 (pCi/1)
         2
      *Limitations  applicable  to mine drainage  from  copper,
       lead,  zinc,  gold,  and silver mines
          **Limitations  applicable to mine  drainage from
             aluminum mines.
       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.


 Upset or. Bypass Conditions (Storm Provision)

 An  issue  of   recurrent  concern has   been    whether   industry
 guidelines   should  include   provisions authorizing noncompliance
 with effluent  limitations during  periods of "upset" or  "bypass.
 An  upset,   sometimes  called  an  "excursion,"  is unintentional
 noncompliance  occurring for  reasons  beyond the reasonable control
 of the permittee.   Some argue that  an  upset  provision  in  EPA's
 effluent  limitations guidelines  is  necessary because such upsets
 will inevitably occur  because  of   the  limitations,  even  with
 properly  operated  control   equipment.  Because technology-based

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 limitations require only what technology can achieve, some  claim
 that  liability for such situations is improper.   When confronted
 with this issue,  courts have disagreed on the question of whether
 an explicit upset or excursion exemption is necessary, or whether
 upset  or  excursion  incidents  may  be  handled  through  EPA's
 exercise of enforcement discretion.

 While  an upset is an unintentional episode during which effluent
 limits  are  exceeded,   a  bypass  is  an  act   of   intentional
 noncompliance   during    which  waste  treatment   facilities  are
 circumvented in emergency situations.   Bypass provisions have  in
 the past been included  in NPDES permits.

 EPA has determined that both explicit upset and bypass provisions
 should  be  included  in  NPDES permits and has promulgated NPDES
 regulations that  include upset and bypass permit  provisions  (see
 45  FR  33448,   122.60(g)   and  (h)  {May  19,  1980)).  The upset
 provision establishes an upset as an affirmative   defense  if  an
 operation is prosecuted for violating a technology-based effluent
 limitation.   The  bypass provision authorizes bypassing to prevent
 loss of life, personal  injury,  or severe  property damage.

 The Agency has  received several inquiries on the  relation between
 the  general upset   and  bypass  provisions  set  forth  in  the
 consolidated permit  regulations and the storm exemption contained
 in the regulations  for  ore  mining  and .dressing.    The  storm
 exemption  contained in  the final BAT regulation supersedes the
 generic upset in  bypass provisions with respect to precipitation
 events;   that  is,   an  operator wishing to obtain relief from BAT
 limitations  during precipitation  events   must  comply  with  the
 prerequisites  of  the rainfall  exemption  provision.   However  the
 upset  and bypass provisions  are available in all  other applicable
 situations.   The Agency recognizes that an excursion  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  or   upset  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-vear    24-
 hour rainfall.                                              '

 This   excursion   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
                                 499

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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),  in  the
clarification of regulations (44 FR 7953), and  in  the  proposed
regulation (49FR25682).

The general relief in the BPT regulation states that:

     "Any  excess  water,  resulting  from rainfall or snow melt,
     discharged  from  facilities   designed,   constructed   and
     maintained  to  contain  or  treat the volume of water which
     would result from  a  10-year,  24-hour  pecipitation  event
     shall  not be subject to the limitations set forth in 40 CFR
     440."  43 FR at 29777-78,  440.81(c)(1978).

     The term "ten-year, 24-hour precipitation event" is defined,
     in turn, as:

     "the maximum 24-hour precipitation event with a probable re-
     occurrence interval of once in 10 years as  defined  by  the
     National   Weather  Service  and  Technical  Paper  No.  40,
     'Rainfall Frequency  Atlas  of  the  U.S.,'  May   1961,  and
     subsequent  amendments,  or  equivalent regional or rainfall
     probability information  developed   therefrom."    43  FR  at
     29778, 440.82(d).

Under BAT, the provision has been clarified somewhat as follows:

     1.  Storm Exemption for facilities permitted to discharge:

If,  as  a  result of precipitation or snowmelt, a source with an
allowable discharge under 40 CFR 440 has  an  overflow   or  excess
discharge  of  effluent which does  not meet the  limitations  of 40
CFR 440, the source  may  qualify   for  an  exemption   from  such
limitations  with  respect  to  such  discharge   if  the following
conditions are met:

     (i)  The facility  is designed, constructed  and  maintained to
     contain the maximum volume  of  wastewater  which  would  be
     generated by the  facility during a 24-hour  period  without an
     increase in volume  from precipitation and  the maximum volume
     of    wastewater    resulting   from   a    10-year,   24-hour
     precipitation event or treat   the  maximum   flow   associated
     with  these  volumes.    In  computing   the  maximum volume of
     wastewater  which   would  result   from  a   10-year,  24-hour
     precipitation  event,  the  facility  must  included  the volume
     which would result  from  all areas  contributing  runoff to  the
     individual  treatment  facility,  i.e.,  all  runoff that  is  not
     diverted from the active mining  area and  runoff which  is  not
     diverted from the mill area.
                                son

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      (ii) The facility takes all  reasonable  steps  to  maintain
      treatment  of  the  wastewater  and  minimize  the amount of
      overflow.
      (iii)     The  facility  complies  with   the   notification
      requirements of S122.60(g) and  (h),.

      The  storm  exemption  is designed to provide an affirmative
      defense to an enforcement action.  Therefore,  the  operator
      has the burden of demonstrating to the appropriate authority
      that the above conditions have been met.

      2.    Storm   Exemption  for  facilities  not  permitted  to
discharge:

If, as a result of precipitation (rainfall or snowmelt), a source
which is not permitted to discharge under  40  CFR  440,  has  an
overflow  or  discharge  which violates the limitations of 40 CFR
440,  the  source  may  qualify  for  an  exemption   from   such
limitations  with  respect  to  such  discharge  if the following
conditions are met.
      (i)  The facility is designed,  constructed,  and  maintained
      to  contain  the  maximum  volume  of  wastewater stored and
      contained by the facility during normal operating conditions
     without an increase in volume  from  precipitation  and  the
     maximum  volume  of  wastewater  resulting  from  a 10-year,
      24-hour  precipitation  event.    In  computing  the  maximum
     volume  of  wastewater  which  would  result from a 10-year,
      24-hour precipitation event,  the facility must  include  the
     volume which would result from all areas contributing runoff
     to  the individual treatment facility,  i.e., all runoff that
      is not diverted from the area or  process  subject  to  zero
     discharge,   and  other  runoff   that is allowed to commingle
     with the influent to the treatment system.

      (ii) The facility takes all reasonable steps to minimize the
     overflow or excess discharge.

      (iii)      The  facility  complies  with   the   notification
     requirements of 8122.60(g)  and  (h).

     The  storm  exemption  is designed to provide an affirmative
     defense to an enforcement action.   Therefore,   the  operator
     has   the   burden   of  demonstrating   to  the  appropriate
     authority that the above conditions have been met.

In general,  the following will apply in granting an excursion:

     1.   The excursion 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

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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,
ore mill discharges.
surface  mine.
and
4.   Relief  is  granted as an exemption to the requirements
for normal operating conditions when there is  an  overflow,
increase in volume of discharge, or discharge from a by-pass
system  caused by precipitation.  The relief only applies to
the  increase  in  flow  caused  by  precipitations  on  the
facility and surface runoff.  It does not apply to surges in
the drainage from underground mines.

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  provision  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 can be granted from all
effluent limitations contained in BAT.

8.  As a practical matter,  relief  will  not  generally  be
available  to  treatment facilities which employ clarifiers,
thickeners, or other mechanically  aided  settling  devices.
The   use   of  mechanically  aided  settling  is  generally
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 "contain" for facilities which are allowed  to
discharge  must  be  considered, in  context  with  the term
"treat" discussed in paragraph 11  below.   The  containment
requirement  for facilities allowed to discharge is intended
                               502

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 to  insure  that   the   facility   has   sufficient   capacity  to
 provide  24 hours settling  time  for  the  volume resulting  from
 the  10-year,   24-hour   storm.    This   is   the  settling  time
 required to "treat"  influent so  that   it   meets  the  daily
 maximum  effluent limitations.   The  theory  is that  a  settling
 facility  with   sufficient volume   to   contain the  10-year,
 24-hour  rainfall plus 24 hours  of discharge can  provide   a
 minimum    24-hour   retention    time for   settling  of  the
 wastewaters even if  the  pond is full at the time  the  storm
 occurs.    The   water entering   the pond  as a  result  of the
 storm  is assumed to  follow a  last-in,   last-out  principle.
 Because  of this the  "contain" and "maintain" requirement for
 facilities  which are allowed  to discharge does not require
 providing  for   draw   down  of   the   pool   level  during  dry
 periods.    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's  being
 emptied  of wastewater  prior   to  the  rainfall  or snow  melt
 upon which the  excursion 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.

 11.    The   term  "treat"   applies   to   facilities  which are
 allowed  to discharge and means  the  wastewater   facility  was
 designed,   constructed,  and  maintained   to  meet the daily
 maximum  effluent  limitations for the maximum flow  volume  in
 a   24  hour period.  The  operator  has  the  option  to "treat"
 the flow volume of water that would result  from  a  10-year,
 24-hour  rainfall  in  order  to qualify   for   the  rainfall
 exemption.    To  compute the maximum  flow volume,  the  operator
 includes   the   maximum   flow  of  wastewater  during  normal
 operating   conditions  without   an   increase  in volume .from
 precipitation plus the maximum  flow that would result  from a
 10-year, 24-hour  rainfall.   The  maximum flow from  a  10-year,
 24-hour  rainfall  can be  determined  from the Water  Shed Storm
 Hydrograph,  Penn  State  Urban  Runoff  Model,  or  similar
 models.

 12.  The term "treat" offers to  the  operator alternatives to
 the  simple  settling provided by tailings ponds and  settling
 ponds  identified as part of BAT.  Examples  of  alternatives
 are:   (1)    clarifiers  designed and operated to "treat" the
maximum  flow volume, but which obviously would not have  the
 actual  volume  to "contain" and provide the actual volume to
provide  an  actual  24-hour   retention   time;   and   (2)
 flocculants  to  aid settling and,  if properly used,  allow a
smaller settling pond to obtain  the same results as  a  larger
settling pond,  e.g. 24-hour retention of the waste water.
                           503

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     13.   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.

     14.   The term "contain" for facilities which are not allowed
     to  discharge  means  the  wastewater facility was designed,
     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, e.g.
     without an  increase  in  volume  from  precipitation.   The
     operator must provide for a freeboard under normal operating
     conditions equivalent to the volume that would result from a
     10-year, 24-hour rainfall.

Should  additional  guidance be necessary, the Agency through the
Effluent  Guidelines  Division  will  provide  guidance  on   the
application  of  the storm provision in a clarification notice to
the final regulation.

Net Precipitation Areas

The general relief  for  the  requirement  of  "no  discharge  of
process  wastewater"  as  promulgated for ore mining and dressing
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.   Comments  from industry following the proposal
     of BAT requested that the Agency give a specific example  of
     determining  the  volume  that  may  be  discharged  from   a
     facility  in  a  net  precipitation  area.   We  offer   the
     following example:
                                 504

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The  Climatic  Atlas  of  the United  States,  a publication of .the
National Oceanic and Atmospheric Administration, U.S.   Department
of  Commerce,  contains  generalized  maps  titled  "Normal Annual
Total Precipitation  (Inches) by  State  Climatic  Divisions"  and
"Mean  Annual Lake Evaporation  (In  Inches)."  From  these maps for
the  area  of  southeastern  Missouri,  the   mean   annual   total
precipitation  is  shown  as 45 inches of rain equivalent and the
mean annual  lake evaporation is shown as 36   inches.    The  first
condition  is  met,  e.g.  precipitation exceeds evaporation  (net
precipitation).  If precipitation does  not   exceed evaporation,
then   the   relief  is  not  available.   If  the   tailings  pond
associated with the mill has a surface area of 450  acres and  the
difference between annual precipitation and lake evaporation  is 9
inches  net  precipitation,  then the facility would be allowed a
discharge of:

450 acres x  9 in (net precipitation.)  x 43560  ft2/acre x 7.5 qal/ft2
                               365 days
                                 or
                             302,000  gal/day
Such discharge is subject to the standards for mine drainage.

The example  above for determining excess precipitation  considered
a discreet tailings  pond  uneffected by  rainfall  or snowmelt
draining into the tailings pond, e.g., only wastewater  discharged
from  the  actual  mill process and precipitation directly on the
tailings pond enters the tailings pond.  Additional consideration
must be given to tailings ponds where runoff  is commingled, i.e.,
where a tailings pond, even with  diversion   ditches,   cannot  be
isolated  from runoff resulting from  precipitation  on the general
mine and mill area.  Assuming a net   precipitation  condition  as
defined  in  the  above example of 45 inches  precipitation and 36
inches evaporation (net precipitation of 9 inches), the following
example describes the relief available  for   a  mountainous  area
where the tailings pond is not discreet, but  receives runoff from
a  surface   area  defined specifically by diversion ditches.  The
total area within the diversion ditches  is   300  acres and  the
surface area of the tailings pond within the  diversion  ditches is
100 acres.    The facility would be allowed a discharge of:

(300 acres x 45 inches - TOO acres x  36 inches)  x 43560 ft2/acre x 7.5 qal/fl
                               365 days'
                                 or
                             738,500  gal/day


Specific precipitation data and evaporation data can be developed
by the mill operator by using instruments measuring precipitation
and  pan evaporation (adjusted to lake evaporation).  Also,  local
weather station data can be used rather than  the Climatic  Atlas.
Regardless, the relief for net precipitation  is  determined for an
annual  volume  of  precipitation  and evaporation not  the excess
that may occur over a few days or  weeks.    Such  excess  can  be
                                     BOB

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handled  by  free  board  of  the  facility  and, if necessary an
increase in daily discharge to  be  latter  adjusted  during  dry
periods.    The   annual  net  precipitation  may  be  discharged
according  to  whatever  schedule  is  most  convenient  for  the
operator,  i.e.,  in the examples offered above it is assumed the
operator would discharge daily a volume that over  a  year  would
equal the total annual volume of excess precipitation.

It  is  recognized that both precipitation and evaporation varies
year to year and if normal precipitation and average  evaporation
is  used  in  determining  net  precipitation at a facility, then
common sense  must  prevail  to  allow  additional  discharge  to
account for "wet years" and even shorter periods of excessive and
frequent  precipitation,  i.e.,  snow  melt.  However, the Agency
feels  that  the  data  from  the  Climatic  Atlas  provides  the
parameters that can be used by the operator to design his holding
and  treatment  facilities.   This  design and construction, when
applied in context with  the  storm  provision  discussed  above,
gives  relief  from  the zero discharge requirement by allowing a
discharge of excess precipitation subject to effluent limitations
where precipitation exceeds evaporation and the  storm  provision
establishes  upset and bypass conditions for when the holding and
treatment  facility  is  overcome  by  excess  precipitation  not
provided  for  in the design, construction and maintenance of the
treatment facility.

Finally, mine and mill areas used in dump  or  heap  leaching  to
recover   copper   are   subject   to   no  discharge  and  could
theoretically qualify 'for  relief  under  the  net  precipitation
provision.   However,  all  of  the  data available to the Agency
shows that such operations are in arid  areas  where  evaporation
exceeds  precipitation  and the net precipitation provision would
not be applicable.

Commingling Provision

The general provision as promulgated for ore mining and  dressing
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  and  concentration  of  each
     pollutant or pollutant property in  the  combined  discharge
     that is subject to effluent limitations shall not exceed the
     quantity  and  concentration  of each pollutant o'r pollutant
     property that would have  been  discharged  had  each  waste
     stream  been treated separately.  In addition, the discharge
     flow from the combined discharge shall not exceed the volume
     that would have been discharged had each waste  stream  been
     treated separately."

The  Agency  received comments requesting that the Agency further
explain the general provision having to  do  with  waste  streams
                              506

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which  are  combined  for  treatment  from  various  subparts and
segments.  We stated in  the  provision  that  the  quantity,  and
quality  of  each pollutant or pollutant property in the combined
discharge shall not exceed  the  quality  and  quantity  of  each
pollutant  or  pollutant property that would have been discharged
had each waste stream been treated separately.  Further, the flow
from the combined discharge shall  not  exceed  the  volume  that
would  have  been  discharged  had  each wastestream been treated
separately.  An example that industry wished clarified is whether
mine drainage commingled with the discharge from a mine  or  mill
process  subject  to  the  zero  discharge  requirements  is also
subject to no discharge.  Such combined  waste  streams  are  not
subject  to  the zero discharge requirement and may be discharged
subject to the limitations for mine drainage but the  volume  can
not  exceed  the  volume  of  mine  drainage that would have been
discharged had the mine drainage been treated separately.  It  is
immaterial  whether  the  mine  drainage  is  introduced  to  the
treatment system simultaneously with the discharge from the mill,
i.e., two separate pipes leading to the tailings pond, or whether
the mine drainage is introduced as part of  the  feed  water  and
intake  to  the  mill itself.  The volume of the discharge cannot
exceed the volume of the  mine  drainage  and  the  discharge  is
subject  to  the numerical concentrations for pollutants included
in the effluent limitations guidelines for mine drainage from ore
mines covered in the subcategory.

The second clarification  requested  has  to  do  with  the  zero
discharge requirements for mine areas where dump, heap, or insitu
leach   processes   are  used  to  recover  copper.   The  Agency
promulgated a clarification of Regulations (44 FR 7953,  February
4,  1979) that address.es these areas and the requirement for zero
discharge.  Simply put,  the  zero  limitation  includes  process
water applied by the operator to the leach area, precipitation on
and  runoff  from the areas used in the leaching process.  Runoff
from active mine areas outside of the areas actually used in  the
leach  process is considered mine drainage.  If on occasion, mine
drainage runoff is drained  into  or  channeled  to  the  holding
facility   for   the  leach  solution,  this  commingled  process
wastewater may be discharged  in  a  volume  equal  to  the  mine
drainage  and  the  discharge  must  meet the limitation for mine
drainage.  However,  the Agency feels that such  a  condition  for
commingling   to   be   extremely  rare  because  it  would  mean
discharging the very copper values that operator wants to recover
as efficiently as possible.  It seems obvious that  the  operator
would   retain   these   copper  values  for  recovery,  even  if
commingled, rather than go to the additional expense of  treating
and discharging the volume of mine drainage.

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
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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  for  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
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
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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.

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-
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/Mill/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.
                                 509

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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  for  treatment
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).

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,  if  mine
drainage and mill discharge are normally commingled; and
                                 Bin

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     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, and Molybdenum Ores Subcategory
presently  achieve  zero  discharge  including  31  copper,  five
lead/zinc,  five
silver mills*
primary  gold,  one molybdenum and four primary
There are two methods of water reclamation that are practiced  in
a  number  of mills.  They are in-process recycle and end-df-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 and 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.  Commingling of
mill discharge and mine drainage and the effluent limitations for
the combined discharge is discussed above in this  section  under
General  Provisions.   The  feasibility of combining the mine and
mill streams will depend on the magnitudes of:

     1.  The flow of mrne drainage;
    ;                              —
     2.   The quality of  the  mine  drainage  (does  it  require
          treatment before use);

     3.  The process water makeup flow required for the mill.

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
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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          -  86  {Not Detected)
     Organics          -17  (Detected at levels below EPA's
                               nominal detection limit)
     Organics          -  10  (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.

Subcateqories  and  Subparts  _in  Which Toxic Pollutants Were Not
Detected  Are  Excluded  Under  Paragraph  £  of  the  Settlement
Agreement

There  were  subcategories and subparts in which all of the toxic
pollutants were excluded from further consideration in regulation
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
     deposits.
                               of   sand
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.
                               51?.

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Subcategories
Under BPT
and Subparts Which Were Not Permitted to Discharge
No discharge of wastewater was specified for  facilities   in  the
following subcategories and subparts under BPT:

     1.  Iron Ore Mills in the Mesabi Range;
          i                                             '    •
     2.   Copper,  Lead, Zinc, Silver, Gold, 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.

Subcateqories 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,  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;

     4.  Uranium mine drainage; and

     5.  Titanium mill process water.
                                                         »
Criterion for Developed BAT Limitations

Recycle was considered as an option for froth flotation mills  in
the  copper,  lead,  zinc,  gold,  silver,  and  molybdenum  ores
subcategory.  This option was rejected for froth flotation  mills
because  of  the  costs  that  would be associated with retrofit,
including the downtime required to retrofit  existing  equipment;
the impact of changing the metallurgical process, e.g. total loss
of  income  while the process was being adjusted to 100% recycle;
and the cost to possibly treat the recycle water  and  adjustment
of the mill process to use recycle water.

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
                               513

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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
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., the data
represents 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 performed 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
                                  514

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that  would  enable  the
placer mine operations.
Agency to perform cash flow analysis of
Limitations for  gold  placer  mines  are  reserved  in  the  BAT
rulemaking  in  the absence of information regarding the economic
impact of regulating gold placer mines and to allow the Agency to
acquire data on the effluent from well  operated  settling  ponds
associated with gold placer mines.

Criterion for BAT Metals Limitations

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
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
                                515

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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:
          Arsenic
          Cadmium
          Copper
          Lead
          Mercury
          Nickel
          Zinc
          TSS*
 30-Day
Average
   0.01
   0.005
   0.05
   0.04
   0.001
   0.10
   0.20
  10
 Daily
Maximum
  0.05
  0.01
  0.20
  0.14
  0.002
  0.40
  0.80
 25
           *TSS  limitations were  computed,  but  TSS
           would be  limited  under BCT.

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
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Water Quality Criteria  (WQC)  for  the protection  of   human   health
and  aquatic  life  published by  the EPA's Criteria  and  Standards
Division  (CSD)  in November  1980.  Because WQC  for the  protection
of  aquatic  life  were  not   developed  for   all  of  the  Section
307(a)(l) toxic pollutants, the average plant  effluent concentra-
tions for pollutants  lacking  these WQC are compared  with   pollu-
tant-specific toxicity data reported in the Ambient  Water  Quality
Criteria  Documents.   The  environmental  significance  of toxic
pollutants in post-BAT discharges is determined  by comparing the
achievable  levels  with WQC  or,  for those pollutants  lacking WQC
for the protection of acquatic life, with EPA  toxicity data.

Based on a review of  the sampling and analysis data  available for
this industry, the only  environmentally  significant  pollutants
after  applying  the  median   dilution from the  average  receiving
stream flow available (to this industry) are cadmium and arsenic.
The concentration of  cadmium  currently being discharged  from this
industry  (BPT) is the lowest  of any industry known   to  discharge
cadmium.   In  addition,  the additional BAT reductions  are small
relative to the levels present in raw (untreated) waste  streams.

In  preparing  the  environmental  assessment,   the  Agency also
compared  raw  waste  mass  loadings  to  those  of  BPT  and those
expected by achievable levels.  It was found that the  industry's
current  discharge  is less than  10 percent of the industry's raw
waste  load.   This   is  due   to  the  installation  and   proper
maintenance of the Best Practicable Technology at most plants.


After   considering   the   environmental   assessment,  the BPT
limitations for ore mining and dressing and economic factors that
are associated with more stringent  limitations  the   Agency has
concluded   that   nationally  applicable  regulations  based  on
secondary settling or any of  the  other candidate BAT technologies
are not warranted in the Ore  Mining  and  Dressing   Point   Source
Category.

The  BAT  limitations  are  promulgated  as they were  proposed on
June 14, 1982.   The comments  received on the proposal, with  a few
exceptions, agreed with the Agency's decision  to  establish  BAT
equal  to  BPT.    However,   a  few  commenters requested BAT less
stringent than BPT and a few  requested BAT  more  stringent  than
BPT.   By  law  BAT  can  not  be  less,  stringent   than   the BPT
limitations in  effect.    The  technology  upon  which   the  more
stringent  BAT  limitations suggested by the commenter were  based
is not considered available technology.    The  technology   is  in
bench  scale  development  and yet to be tested  in large scale or
implemented on the total  discharge  from  an  ore  mining   point
source.
                                  517

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Cyanide Control and Treatment

As discussed in Section VIII, cyanide compounds are used in froth
flotation  process  of  copper, lead, zinc, and molybdenum, ores.
In addition, the cyanidation process is used for leaching of gold
and silver ores.  Consequently, residual cyanide is found in mill
tailings and wastewater streams from  these  mills.   Cyanide  is
also  found  in low concentrations in mine drainage at facilities
which backfill  mine  stopes  with  the  sand  fraction  of  mill
tailings.

Of  the control and treatment technologies available for cyanide,
consideration was given  to  the  following  options:  in-process
control,  chemical  oxidation  (alkaline  chlorination,  hydrogen
peroxide oxidation, ozonation), and  natural  oxidation  and  the
incidental   removal  occurring  in  existing  treatment  systems
(tailing ponds).  These options were judged to be most applicable
to the high flow volume and comparatively low  concentrations  of
cyanide  in  the  wastewater  streams  typical  in this category.
Another alternative which was considered was the substitution  of
other   reagents   for   cyanide  compounds  in  froth  flotation
processes.  Bench-scale testing indicated that this  alternative,
although technically feasible, would require extensive testing in
actual  production  of  circumstances  with  specific  ores.   In
addition, it  would  be  difficult  in  these  cases  to  predict
downtime,  loss  of  recovery  (if any), and costs associated with
process modifications.

Alkaline Chlorination

This method was  described   in  detail  in  Section  VIII,  while
operating   cost   assumptions   are   outlined  in  Section  IX.
Basically, oxidation of cyanide by alkaline chlorination  may  be
accomplished  by   infusion   of  gaseous  chlorine  into the waste
stream at a pH greater than  10, or  by  the  addition  of  sodium
hypochlorite  (NaOCl)  as an oxidcint along with an alkali such as
sodium hydroxide  (NaOH).  The  alkali achieves pH  adjustment  and
precipitation  of  metal  hydroxides formed from the breakdown of
metal-cyanide complexes.

Pilot-scale tests  of alkaline  chlorination treatment at Mill 6102
showed reduction of effluent   cyanide  concentrations  from  0.19
mg/1  to  less  than  0.1 mg/1 at pH values greater  than  8.8.  In
addition,  Mill  3144  achieved  reduction  of  effluent  cyanide
concentrations  to an  average  of  0.18  mg/1  from  4.72  mg/1
following the installation of  a full-scale  alkaline-chlorination
treatment system.

Ozonation

Oxidation  of   cyanide  by ozonation is also accomplished at ele-
vated pH  (9 to  12).  Copper  appears  to act as a catalyst  in  this
process,  which  suggests  that  waste  streams containing  copper

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cyanide complexes may be treated more effectively  by  ozonation.
Pilot-scale testing of ozonation at Mill  6102 showed reduction of
cyanide  concentration from 0.55 mg/1 to  less than 0.1 mg/1 at pH
greater than 7.4.

Hydrogen Peroxide Treatment

Hydrogen peroxide (H2O2) has also been tested on a limited  basis
as  an  oxidant  for  cyanide  treatment   in  milling  wastewater
streams.  This process also requires an alkaline pH  and  can  be
enhanced  by  a copper catalyst.  Mill 6101 has achieved approxi-
mately 40 percent removal of cyanide during periods  of  elevated
effluent  levels  (up  to  approximately   0.09  mg/1) by hydrogen
peroxide oxidation.

Process Control

One characteristic of the froth flotation  process  which  poten-
tially affects effluent wastewater quality is the latitude avail-
able to the mill operator at the upper end of the dosage applica-
tion  spectrum.   That  is,  while  the addition of less than the
necessary quantities of cyanide  reagent  may  lead  to  loss  of
recovery or reduced product purity, the addition of more than the
necessary  quantities  of  cyanide  reagent is not accompanied by
penalties to the same degree, except of course, the cost  of  the
additional reagent.

Close  attention  to  mill  feed  characteristics and careful and
frequent analysis of its mineral content can result in  reduction
of  cyanide  dosage  to that actually required.  In recent years,
on-line analysis techniques and reagent  addition  controls  have
become available to minimize excess additions of reagent.

Few  froth  flotation  process  facilities  in  the industry have
reported treated effluent cyanide concentrations equal to  or  in
excess  of 0.1  mg/1, and these only on an infrequent basis.  Mill
6102, the largest consumer of cyanide in terms of dosage per unit
of ore feed, has been observed in the past to  generate  effluent
cyanide  concentrations as high as 0.2 mg/1 to 0.4 mg/1.  Follow-
ing installation of cyanide treatment, this facility is reporting
cyanide levels less than 0.1 mg/1.   Three other  flotation  mills
have  reported  discharge  concentrations  in  excess of 0.1 mg/1
(2122, 3121, 6101).  In each case,  the cyanide  dosages  used  in
mill  feed appear to be consistent with dosages reported through-
out the industry and are not unusual in that  respect.    Fluctua-
tions  and peaking in cyanide concentrations appear to be related
to short-term overdoses of cyanide in the flotation process.  Few
treated  effluent  measurements  in  the  entire  industry   have
exceeded 0.2 mg/1 and we believe that, with close process control
and  reagent  addition  in  combination  with a well designed and
operated treatment system,  the 0.2  mg/1  measurement  for  total
cyanide  can  be achieved without additional treatment technology
for  cyanide.    For  the  rare  case  where  difficulty  may   be
                                  519

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encountered   or   great   reliability   is  required,  treatment
technology (i.e., chemical oxidation) is available  as  discussed
in Section VIII and costed in Section X.

Many  existing NPDES permits for ore mills contain limitations on
total cyanide.  As example, in EPA Region VIII,  there  are  nine
existing  permits  that limit total cyanide in the discharge from
ore mills and these limitations vary from .02  mg/1  to  .2  mg/1
daily  maximum.  In EPA Region X there are twelve permits for ore
mills that limit total cyanide and these  limitations  vary  from
.01  mg/1  to  .3  mg/1 daily maximum.  Monitoring data for these
permits confirm that these mills are  consistently  within  their
permit  limitations on total cyanide and that the limitations can
be obtained by control of the process and the incidential removal
of cyanide as discussed below.

Incidental Cyanide Removal

Frequently,  specific  cyanide  treatment   technology   is   not
necessary  if  close  process  control  combined  with incidental
removal leads to low concentrations  of  total  cyanide  in  mill
water  treated  effluent.   This incidental removal is thought to
involve several mechanisms,  including  ultraviolet  irradiation,
biochemical  oxidation,  and  natural aeration.  As evidence that
such mechanisms are involved, it has  been  noted  that  effluent
cyanide  concentrations  tend to be somewhat higher during winter
months when biological activity in the tailing pond is lower  and
ultraviolet  exposure  is  much  lower  due to shortened daylight
hours, less intense radiation, and ice cover on the ponds.

In addition,  the  association  of  cyanide  with  the  depressed
minerals   (i.e.,  pyrites) will cause a portion of the cyanide to
be removed together with the suspended solids  and  deposited  in
the tailing ponds.

Precision and Accuracy Study

A  study  of the analysis of cyanide  in ore mining and processing
wastewater was conducted in cooperation with the American  Mining
Congress  to  investigate  the causes of analytical interferences
observed and to determine what effect these interferences had  on
the  precision  of  the  analytical  method.  The purpose of this
study was to evaluate the  EPA-approved  method  and  a  modified
method  for  the  determination  of cyanide.  The modified method
employed a lead acetate  scrubber  to  remove  sulfide  compounds
produced during the reflux-distillation step.  Sulfides have been
suspected  of  providing  an  interference  in  the  colorimetric
determination of cyanide concentrations.  Also,  several  samples
were spiked with thiocyanate to ascertain if this compound caused
interference in the cyanide analysis.

A statistical analysis of the resultant data shows no significant
difference  in  precision or accuracy of the two methods employed

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when applied to ore mining and milling wastewaters  having  cyanide
concentrations in the  0.2 mg/1 to  0.4 mg/1  range.   Based upon  the
statistical analysis,  approximately  50  percent of  the   overall
error  of  either method was attributed to  intralaboratory error.
This highlights the need for an experienced analyst   to   perform
cyanide  analyses.   After  considering the results of this study
and the levels achieved through dose control of reagent addition
in  the  mill, EPA considered proposing an  effluent limitation of
0.2 mg/1.  That limitation is based on a grab sample  for any  one
day,  and would have been subject  to 100 percent error to  account
for the precision and  accuracy of  the  analytical   method.    (See
Section  V above).  Therefore, the Agency would have  had to allow
an analytical measurement of up to 0.4 mg/1.   However,  all   the
data   observations    in   our  sample  were  below  that   level.
Accordingly,  the  Agency  is  excluding  cyanide   from national
regulation in the ore  mining category.

However,  it  has  come  to the attention of the Agency that site
specific  measurements of  cyanide  are    being    performed   at
individual laboratories to quantify removals by various treatment
methods  for  cyanide.  Such other analytical methods  can  be used
to monitor cyanide and limitations on cyanide can be  included  in
individual  permits  when  the  permit  specifies   an  alternative
analytical method.

ADDITIONAL PARAGRAPH £ EXCLUSIONS

Exclusion of_ Cyanide

Total cyanide is not regulated in  BAT because the   Agency   cannot
quantify   a   reduction   in   total   cyanide   from  observed
concentrations being discharged by use of technologies, known  to
the   Administrator,   Paragraph   8(a)  iii  of  the   Settlement
Agreement.

The references to total cyanide levels  of   less  than 0.2  mg/1
throughout  this document are for  informational purposes only  and
are subject to the  precision  and accuracy  of  the   analytical
method as discussed here and in Section V.

Exclusion of Arsenic and Nickel

EPA  reviewed  the  achievable  levels  calculated  based  on  the
capabilities of the three candidate  BAT  treatment  technologies
(secondary  settling,   coagulation and flocculation, and granular
media  filtration).     The  Agency  examined  the    necessity   of
proposing  specific limitations for all seven of the toxic metals
considered for regulation.   Limitations on  copper,   lead, and zinc
are necessary since these are the  metals  recovered  from  mining
operations  and  concentrated  in  mills in  this category.   From a
treatability viewpoint, control of some toxic metals  (arsenic  and
nickel)   may  be  achieved  by  limitations  upon   which   other
pollutants  are  controlled.   As discussed  in this section, since
                                521

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most of the metals are in suspended solids, reduction of  arsenic
and  nickel  occurs  in  conjunction  with the removal of TSS and
other toxic metals (copper, lead, zinc and mercury).

The BAT data base for the Ore Mining and  Dressing  Point  Source
Category  was  searched for instances in which arsenic and nickel
concentrations exceeded BPT limitations when copper,  lead,  zinc
and  mercury  concentrations were also below their respective BPT
limitations.  There was only one instance in over 300 samples  in
which  a  nickel  or  arsenic  concentration  exceeded  their BPT
limitations when BPT  limitations  for  copper,  lead,  zinc  and
mercury  were  met.   The  one  instance was the discharge from a
sedimentation pond at Facility 3103.   The  nickel  concentration
was 0.22 mg/1 as opposed to the 0.20 mg/1 BPT limitation.

The  Agency  concluded that the limitations on copper, lead, zinc
and mercury would ensure adequate control of arsenic and  nickel,
and  under Paragraph 8(a)iii of the,Settlement Agreement, arsenic
and nickel are excluded from regulation.

Exclusion of Asbestos

Chrysotile asbestos was detected in  wastewater  samples  in  all
subcategories  and  subparts  within  the ore mining and dressing
point source category.  It was  detected  in  90  of  91  samples
throughout the entire industrial category.

EPA  believes  that  the most appropriate way to regulate a toxic
pollutant is by a  direct  limitation  on  the  toxic  pollutant.
However,  direct  limitation  of  toxic  pollutants is not always
feasible.  In the case of chrysotile asbestos, there  is  no  EPA
approved  method  of  analysis for  industrial wastewater samples.
The method of analysis presently used was developed for  drinking
water  samples.   In  addition,  there are less than half a dozen
laboratories in the United States that are capable of  performing
the analysis by this method.

Chrysotile  asbestos  is known to be present in many ore deposits
throughout the country  (Reference  6).   As  ore   is  mined  and
subsequently milled, it is subjected to. a variety of crushing and
size  reduction  operations.   As   a  result,  smaller solids are
formed, the chrysotile asbestiform  fibers are  liberated  as  the
small  solids  are  made,  and  end up in mine drainage and mill
process water.

The possibility of the chrysotile asbestos fibers   being  present
in  waste  streams  for the same reasons and in the same relative
proportions as the solids, led to   the  examination  of  the  EGD
sampling  data  in an attempt to establish a relationship between
chrysotile asbestos and TSS.
                                 522

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Review of  analyses  for  asbestos  and  TSS  in  samples   of   untreated
and  treated wastewater  shows  that  as TSS is reduced  by  treatment,
observed asbestos concentrations are also reduced.
 Intake  water
 effluents were
 for chrysotile
 ranged from  3.
 value . of  1 .1
 sample values
 104 (detection
 samples  (from  26 industrial categories) and POTW
 reviewed to get an indication of background levels
  asbestos.    The  values  of  chrysotile  asbestos
5 x 1.0* (detectioni. limit) to 1.63 x 1Q8 with a mean
  x 107 fibers per liter.  The treated waste stream
for chrysotile asbestos in the ore data ranged from
 limit) to 108 fibers per liter.
The Agency has determined that when TSS  is  reduced   to   the  BPT
effluent  limitations  for  ore  mining  and  dressing,   observed
chrysotile asbestos levels are reduced   to  or  below background
levels.   Therefore,  EPA  is  excluding chrysotile asbestos from
regulation since it is  effectively  controlled  by   technologies
upon  which  TSS  limitations are based, Paragraph 8(a)iii of the
Settlement Agreement.
Exclusion of Pollutants Detected in a Single Source and
Related to That Source
                                           Uniquely
There  are 19 operating uranium mills in the United States, 18 of
which now achieve zero discharge of  process  wastewater.   There
are  no uranium mills that commingle process wastewater with mine
drainage and it is anticipated that none of these zero  discharge
mills  would  elect to treat and discharge at the BPT limitations
because of the expense to install technology required, i.e.,  ion
exchange,  ammonia stripping, lime precipitation, barium chloride
coprecipitation, and settling.

EPA is excluding uranium mills from BAT, since there is only  one
discharging  facility  and  it is believed that none of the other
existing facilities will commingle mine drainage and mill process
wastewater.  Uranium mills are not regulated in BAT  because  the
pollutants  found  in  the discharge are uniquely related to this
single source,  Paragraph 8(a)iv of the Settlement Agreement.
                               523

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                           SECTION XI

         BEST CONVENTIONAL POLLUTANT CONTROL TECHNOLOGY

Section 301(b)(2)(E) of The Act requires that  there be  achieved,
not  later  than July 1, 1984, effluent  limitations for categories
and classes of point sources, other than publicly-owned treatment
works, that require the  application  of  the  best  conventional
pollutant   control  technology  (BCT) for control of conventional
pollutants  as identified in Section  304(a)(4).   The  pollutants
that  have  been  defined  as conventional by  the Agency, at this
time, are biochemical  oxygen  demand,  suspended  solids,  fecal
coliform, oil and grease, and pH.

BCT  is not an additional limitation; rather,  it replaces BPT for
the control of conventional pollutants.  BCT   must  be  evaluated
for cost effectiveness and a comparison made between the cost and
level  of reduction of conventional pollutants from the discharge
of publicly owned treatment works (POTW) and the cost  and  level
of  reduction  of  such  pollutants  from  a class or category of
industrial  sources.

On October  29, 1982 EPA proposed the methodology to determine the
cost-reasonableness  of   all   BCT   tchnology   options.    The
methodology  consists  of two parts:  a POTW test and an industry
cost-effectiveness  test.   The  POTW  test  is  passed  if   the
incremental  cost  per pound of conventional pollutant removed in
going from BPT to BCT  is  less  than  $.27  per  pound  in  1976
dollars.   The  industry  test is passed if this same incremental
cost per pound is less than 143%  of  the  incremental  cost  per
pound  associated  with achieving BPT.  Both tests must be passed
for a BCT limitation more stringent than BPT to  be  established.
In  those  subcategories  for  which  BAT or BCT limitations were
never  promulgated,  or  were  being  reevaluated  on   technical
grounds, the Agency considered several candidate technologies for
BCT.    These   candidate  technologies  are   those  that  remove
significant amounts of conventional pollutants  beyond  BPT.   In
evaluating their reasonableness, EPA used BPT  as a starting point
and  determined  the  incremental  costs  and  levels of pollutant
removal from BPT to each  of  the  candidate   technologies.   The
selection  of  the  final  BCT  limitations  is based on the most
stringent  technology  option  which  passes  the  reasonableness
tests, as well as the other.

EPA proposed BCT equal to BPT limitations for  seven subcategories
of the ore mining and dressing industry on June 14, 1982.  (47 FR
25682)   The  proposed  limitations  were  published  erroneously
without applying the proposed BCT cost test.   EPA has now applied
the new test to all  seven  subcategories.    None  pass  and  EPA
proposed revised BCT limitations for them equal to BPT on October
29, 1982.
                                525

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                            SECTION XII

              NEW SOURCE PERFORMANCE STANDARDS (NSPS)

 The  basis  for  new  source  performance  standards (NSPS)  under
 Section 306 of the Act is best available demonstrated technology.
 New facilities have the opportunity to  implement  the  best  and
 most  efficient  ore  mining and milling processes and wastewater
 technologies.   Congress,  therefore, directed EPA to consider  the
 best  demonstrated  process  changes  and  end-of-pipe  treatment
 technologies capable of reducing pollution to the maximum  extent
 feasible.


 GENERAL PROVISIONS

 Several  items  of  discussion  apply to options in more than  one
 subcategory.   To avoid  repetition,  these  items  are  discussed
 here.

 Relief From No Discharge  Requirement

 Facilities  which  are  not  allowed  to discharge under "normal"
 conditions may do so as a result of:

 1.   An overflow or increase in  volume from rainfall or   snowmelt
 if   the  facility  is  designed,   constructed,   and maintained to
 contain a  10-year,  24- hour rainfall   provision   over and   above
 normal  pond levels.

 2.   Location  in' a "net precipitation"  area;  such facilities   can
     discharge  the  difference  between the precipitation falling
     on the facility and  evaporation  from this area.

 3.   Groundwater infiltration

 These provisions are discussed below.

 Storm Provision

 1.   EPA proposed that  a  new source subject  to no discharge could
 be granted  an  excursion and allowed  to  discharge  excess  water
 upon    the  occurrence  of  a  10-year,   24-  hour  storm event.
 Conversely,  existing   sources  subject   to   a    no    discharge
 requirement were granted  an excursion upon  the good faith showing
 of  best  engineering   judgment by  the  operator that  the  facility
was designed,  constructed and  operated  to  contain  the  volume
 resulting   from   a   10-year,  24-hour   rainfall   plus the plant's
 regular process  wastewater discharge.   Both require  the   operator
 to  design  based on engineering  judgment, but a  new source would
also have to   show   the   10-year,   24-hour  event   or  equivalent
occurred.  Determining equivalent snowmelt  is difficult.  Storms,
or  snowmelts,   or combinations of  storms and snowmelts occurring
                              527

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subsequently were not granted relief, but this total volume could
exceed  a  single  10-year,  24-hour  precipitation  event.   The
operator,  therefore, had no design volume upon which to base his
design of holding facilities for areas and processes  subject  to
the no discharge requirement.  This could require the operator to
design  not for a TO-year, 24-hour precipitation event, but for a
TOO year event, or the maximum precipitation  event  which  could
statistically  occur on into perpetuity.  The relief granted upon
the  10-year,  24-hour  precipitation  event  for  standards   of
performance for those subparts subject to no discharge of process
wastewater  is  changed in the final standards of performance and
relief is granted upon the design, construction, and  maintenance
of  the  facility to contain the volume which would result from a
10-year, 24-hour rainfall and the process water  to  the  holding
facility as in the BPT and BAT regulation.

The  proposed  storm provision, which conditioned relief upon the
occurence of a 10-year, 24-hour precipitation event, was done  at
the  request  of some permitting authorities who stated that such
requirement  is  currently   included  in   NPDES   permits,   the
requirement  had  caused  no  problem to date, and better meets the
requirement of  no  discharge  of  process  wastewater.   We  now
recognize  that while no  problem has occurred with this condition
contained in existing permits, there is the potential of  such  a
problem  occurring.   In  a  report,  "Evaluation  of Performance
Capability of Surface Mine Sediment Basins," prepared in  support
of  effluent  limitation  guidelines for  coal mining, conclusions
are made based on statistical probability  as   to  occurrence  of
multiple  storm  events which are also applicable to storm events
which  can occur at ore mines and  mills.   Conclusions  in  this
report  include:  1)  It   is impossible   to  design a pond which
guarantees against the possibility that its capacity will  not  be
exceeded by some multiple storm scenario;  2)  Increasing pond size
to  retain  runoff   from  multiple   storm events  obeys a low of
diminishing returns.  As  the pond  size   increases   in  order  to
reduce  the  possible over flow,  large  incremental cost increases
are anticipated for  decreasing  increments of  protection;   and  3)
Without  a  relief which  recognizes  the probability  of subsequent
events in terms of total  flow to  the  pond,  an  overflow could
always  occur  as  a result of  multiple storm events even  if a
10-year  storm does not occur.   This  makes relief granted upon  the
occurrence of  a  10-year,  24-hour  storm  impractical or   impossible
for facilities which must contain run-off from  large areas.

The   storm  provision applicable  to  existing  sources subject  to a
BAT no discharge  requirement is discussed in  Section X.   In   that
the   storm  provision for BAT  is  identical to the  storm provision
for NSPS, the  considerations discussed  under  BAT are the same  for
NSPS.   Similarly,  the  discussion  of   the  relationship   of   the
general  upset  and  bypass conditions  for BAT are  also applicable
to NSPS.

Net Precipitation Areas
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 2.    Relief from no  discharge  of  process  wastewater  is  also
 granted  to new facilities located in net precipitation areas and
 is  the same  relief  granted  to  BAT  limitations  requiring  no
 discharge.    In  that  the  net  precipitation relief for NSPS is
 identicial  to the relief for  BAT  the  considerations  discussed
 under BAT are the same for NSPS.

 Ground Water Infiltration                .

 In   addition,   for  new  sources  subject to zero discharge,  i.e,
 froth flotation mills and  uranium  mills,   the  Agency  received
 comments  stating that new mills may have to locate tailings ponds
 in   valleys  or  other  locations  that  would receive water  from
 natural   springs,   and  run-off  from   higher   elevation   that
 percolates   into  the  ground   an£  seeps into the tailings pond.
 This would  cause a build-up of excees water and  requires  relief
 not   addressed   by  the  net  precipitation  relief  or the storm
 provision.

 The  Agency  believes that such  a situation is unlikely or will   be
 seldom   encountered.    For existing  sources  subject  to  zero
 discharge,  the  Agency knows of only one example,   but  NSPS  does
 require   zero   discharge  of  two additional subparts,  e.g. froth
 flotation mills and uranium mills.   As discussed  above  in  this
 section,  the   storm  provision  for  new sources  subject to  zero
 discharge has been changed from that proposed and,   as   discussed
 latter in   this  section,  the zero discharge requirement for  new
 froth  flotation mills has  been amended  from  that  proposed   to
 allow  a  bleed  to  the system.   These changes will  ameliorate much
 of the problem  caused by seepage  into a tailing  pond.

 However,  for any new source subject to zero discharge of  process
 wastewater,  if  the  operator can demonstrate  to the  permitting
 authority that  the tailings pond  does receive excess  water from
 infiltration  or  seepage   into the tailings pond  that  can not  be
 diverted, i.e.   diversion  ditches  above and around  the  tailings
 pond   and  sealing  and   grouting   the  springs;   the  permitting
 authority can grant  a limited  discharge equivalent to the  excess
 water  and  subject  to the limitations for  mine  drainage from  the
 applicable  subcategory.   This  provision is  as follows:

 "In  the event a  new  source subject  to a no  discharge  requirement
 can  demonstrate   that   groundwater   infiltration   contributes  a
 substantial  amount   of  water  to   the  tailing   impoundment   or
 wastewater  holding   facility,  the  permitting authority  may allow
 the discharge of a volume  of water  equivalent to   the  amount   of
 groundwater infiltration.   This discharge shall be subject  to the
 limitations  for  mine   drainage  applicable   to   the  new  source
 subcategory."
Relief From Effluent Limitations for Those
to Discharge
Facilities  Permitted
                                 529

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Storm Provision

The  storm  exemption  for  facilities  allowed  to  discharge is
exactly the same as that granted for existing sources under  BAT.
Excess   water  resulting  from  precipitation  from  a  facility
designed, constructed, and maintained to  contain  or  treat  the
maximum  volume  of  process wastewater discharged during any 24-
hour period, including the volume that would result  from  a  10-
year,  24-hour  precipitation  event may qualify for an exemption
from the limitations set forth in 40 CFR 440.

In that  the  storm  provision  for  NSPS  is  identical  to  the
provision for BAT, the considerations discussed under BAT are the
same for NSPS.

Commingling Provisions

For  new  sources  that  combine for treatment waste streams from
various sources, the quantity and quality of  each  pollutant  or
pollutant  property  in the combined discharge that is subject to
effluent limitations shall not exceed the quantity and quality of
each  pollutant  or  pollutant  property  that  would  have  been
discharged  had  each  waste stream been treated separately.  The
discharge flow from a combined discharge  shall  not  exceed  the
volume that would have been discharged had each waste stream been
treated separately.

In  that  the  commingling provision for NSPS is identical to the
provision for BAT, the considerations discussed for BAT  are  the
same for NSPS.

NSPS OPTIONS CONSIDERED

The Agency  considered the following NSPS options:

Option One.  Require achievement of performance standards  in each
subcategory based on the same technology as  BAT  (NSPS = BAT).

Option  Two.  Require standards based on a complete water  recycle
system  (NSPS = zero discharge).

NSPS SELECTION AND DECISION CRITERIA

EPA  has  selected  performance  standards   based  on   the  same
technology' as   BAT   for  all  facilities   in   the ore  iriining and
dressing point source category,  except  those   facilities  using
froth   flotation  in  the  copper,   lead,  zinc,  gold, silver, and
molybdenum  subcategory and mills  in  the  uranium subcategory.

Subcateqories and  Subparts Under Option  T_

Option  1  (NSPS  = BAT) has been selected  for  iron ore mills in the
Mesabi  range; copper,  lead,  zinc,  silver,   gold,   and   molybdenum
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 mills   that   use  leaching  to recover copper and the cyanidation
 process for  the recovery of  gold;  and  mercury  mills  since  BAT
 specifies  zero  discharge.    Option 1  (NSPS = BAT)  has also been
 selected for iron ore-mine drainage,   iron  ore  mills,  aluminum
 mine   drainage,   copper,  lead,  zinc,  gold, silver,  and molybdenum
 mine drainage,  titanium mine drainage,   dredges  and  mills,   and
 mercury mine drainage.   The concentration levels of toxic metals
 found  in new sources  in  these  subcategories  and  subparts  are
 expected  to be  similar to  existing  sources.    Following the
 implementation  of BAT,  toxic metals  will  be  found   at  or  near
 detection  levels or  at concentrations below the practical limits
 of  additional technology.  Further reduction of these  pollutants
 can not be technically or economically justified.

 Subcateqories and Subparts Under Option 2 (NSPS = zero discharge)

 1.   For froth  flotation  mills  in  the  Copper  and   Zinc,   Gold,
 Silver,  and Molybdenum  Subcategory  EPA requires that new source
 froth   flotation  mills  achieve  zero   discharge    of   process
 wastewater.   EPA considered zero discharge based on recycle for
 existing copper,  lead,  zinc,  gold, silver  and,  molybdenum  mills
 using   froth flotation,  but rejected it because of  the effect of
 the retrofit required  at  some existing  facilities,   the  cost of
 retrofitting including   the  loss of   income  while the mill is
 adjusted  to 100 percent  recycle,   and  the  possible  changes
 required  in the process.    This concern does not apply to new
 sources.   New sources  have the   option   to  recycle   because   the
 metallurgical  process  can   be  adjusted and designed to recycle
 process  wastewater before  the actual   construction   of  the   new
 source.  Zero discharge  is a demonstrated technology at 46  of the
 90  froth  flotation mills for which  EPA has data  (see wastewater
 discharges as summarized  in  Tables IX-2  through IX-10).   It meets
 the definition   of  standard of  performance  permitting    zero
 discharge  of  pollutants.    Zero  discharge  does   offer further
 reduction of  pollutants than BAT  and   it  will  not   result in
 adverse  economic  impacts.

 There  are   new   sources anticipated  in  copper,  lead,  zinc, gold,
 silver,  and  molybdenum mining.   Standards applied  to   these   new
 source   waste  streams  should   reflect  the  best  treatment  levels
 achievable by the froth flotation  segment of  the subcategory.

 A study  of existing froth  flotation mills reveals  that   a  large
 percentage   of  these  facilities  are   effectively  achieving  TOO
 percent  recycle of mill water.   Many of  the  facilities  practicing
 100 percent  recycle  are  located  in   arid   regions,   but   some
 facilities   are   located   in  humid  regions.   A summary of  some
 existing facilities follows.
Copper Ore

Of the 35 known  froth  flotation  copper  mills  in  the
States, 31 achieve zero discharge of process wastewater.
United
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Lead/Zinc Ores

Five  of  the  27  active  froth flotation mills in the lead/zinc
subcategory achieve zero discharge.

Gold

Four  of  the  five  primary  gold  facilities  employing   froth
flotation techniques discharge process wastewater.

Silver

Three of the four known primary silver facilities which use froth
flotation   methods  are  achieving  zero  discharge  of  process
wastewater.

Molybdenum

Of the three molybdenum operations employing the froth  flotation
process, one facility achieves zero discharge of recycle.

Of  the  46 existing mills which are achieving zero discharge the
majority are located in arid areas but 15 are  located  in  areas
where net precipitation is generally above net evaporation  (i.e.,
3  in  Idaho, 3 in Colorado, 3 in Missouri, 2 in Washington, 2 in
Tennessee, 1 in Wisconsin, and  1  in  California).   The   Agency
therefore  believes  zero  discharge  to be demonstrated for both
rainy