U.S. DEPARTMENT OF COMMERCE
                                       National Technical Information Service

                                       PB-286 520
DEVELOPMENT  DOCUMENT FOR EFFLUENT LIMITATIONS GUIDELINES
AND NEW SOURCE PERFORMANCE STANDARDS FOR THE ORE  MINING
AND DRESSING POINT SOURCE CATEGORY,   VOLUME  i
ENVIRONMENTAL PROTECTION AGENCY
JULY 1978

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          DEVELOPMENT DOCUMENT
                  for
    EFFLUENT LIMITATIONS GUIDELINES
                  and
    NEW SOURCE PERFORMANCE STANDARDS

                for the

        ORE MINING AND DRESSING
         POINT SOURCE CATEGORY

       VOLUME I - SECTIONS I - VI
           Douglas M. Costle
             Admin istr ator

           Thomas C. Jorling
   Assistant Administrator for Water
        and Hazardous Materials

               Swep Davis
     Deputy Assistant Administrator
    for Water Planning and Standards
           Robert B. Schaffer
 Director, Effluent Guidelines Division

           Baldwin M. Jarrett
            Project Officer

            Ronald G. Kirby
       Assistant Project Officer

                July, 1978

      Effluent Guidelines Division
Office of Water and Hazardous Materials
  U. S. Environmental Protection Agency
        Washington, D.C.  20460

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 50272 -101
 REPORT DOCUMENTATION
         PAGE
                        1. REPORT NO.
EPA 440/l-78/061~d
3. Recipient's Accession No.

 PB-286 520
 4. Title and Subtitle
   Development Document for Effluent  Limitations Guidelines for
   the Ore Mining and  Dressing Point  Source Category (Vol. I)
                                               5. Report Date
                                                April  1978
 7. Author(s)Baldwin M§ jarrett, Project  Officer
 	Ronald G. Kirby,  Assistant Project Officer
                                               8. Performing Organization Rept. No.
 9. Performing Organization Name and Address
   Effluent Guidelines Division
   Office of Water and Waste Management
   U.S.  Environmental Protection Agency
   401 M Street, S.W.
   Washington, D.C.  20460	
                                               10. Project/Task/Work Unit No.

                                                 2BB  156
                                               11. Contract(C) or Grant(G) No.
                                               (C)

                                               (G)
 12. Sponsoring Organization Name and Address
   Same
                                               13. Type of Report & Period Covered

                                                Final Report
                                                                        14.
                                                                         EPA-EGD
 IS. Supplementary Notes
 16. Abstract (Limit: 200 words)
   This document presents the findings of an extensive study of  the ore mining  and
   dressing industry,  for the purpose  of developing  effluent limitations guidelines for
   existing point sources and standards of performance and pretreatement standards for
   new sources,  to implement Sections  304, 306 and 307 of the Federal Water Pollution
   Control Act,  as amended in 1977 by  the Clean Water Act, P.L.  95-217.

   Effluent limitations  guidelines are set forth  for the degree  of  effluent reduction
   attainable through  the application  of the best practicable control technology  currently
   available (BPCTCA)  and the degree of effluent  reduction attainable through the
   application of the  best available technology economically achievable (BATEA) which
   must be achieved by existing point  sources.  The  standards of performance for  new
   sources are set forth for the degree of effluent  reduction which is achievable
   through the application of the best available  demonstrated control technology,
   processes,  operating  methods, or other alternatives.

   Supporting data and rationale for development  of  the proposed effluent limitation
   guidelines and standards of performance are contained in this report (Volumes  I
   and II).
 17. Document Analysis a. Descriptors
   Ore Mining
   Mineral Dressing
   Water Pollution Control
   Waste Water Treatment
   b. Identifiers/Open-Ended Terms
   Best Available Technology
   Economically Achievable
   Best Practicable Control
   Technology Currently Available
   New Source Performance Standards
   c. COSATI Field/Group
 18. Availability Statement
   Release to Public
                                19. Security Class (This Report)
                                  Unclassified
                                                        20. Security Class (This Page)
                                                           Unclassified
          21. No. of Page*
           429
                                                         22. Price
                                                          Al8-AQ1
(SeeANSI-Z39.18)
                See Instruction* on Reverse
          OPTIONAL FORM 272 (4-77)
          (Formerly NTIS-35)
          Department of Commerce

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                          ABSTRACT
This document presents the findings of an extensive study of
the ore mining and dressing industry,  for  the  purpose  of
developing  effluent  limitations  guidelines  for  existing
point sources and standards of performance and  pretreatment
standards  for  new  sources, to implement Sections 304, 306
and 307 of the  Federal  Water  Pollution  Control  Act,  as
amended in 1977 by the Clean Water Act, P.L. 95-217.

Effluent limitations guidelines are set forth for the degree
of  effluent reduction attainable through the application of
the best practicable control technology currently  available
(BPCTCA)   and  the  degree  of effluent reduction attainable
through the application of  the  best  available  technology
economically  achievable   (BATEA)  which must be achieved by
existing point sources by July 1, 1977, and  July  1,  1984,
respectively.  The standards of performance and pretreatment
standards  for  new  sources are set forth for the degree of
effluent  reduction  which   is   achievable   through   the
application of the best available demonstrated control tech-
nology, processes, operating methods, or other alternatives.

Supporting   data  and  rationale  for  development  of  the
proposed effluent limitation  guidelines  and  standards  of
performance are contained in this report (Volumes I and II).
                            iii

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                    CONTENTS  (VOLUME I)

Section                                                       Page

I        CONCLUSIONS                                            1

II       RECOMMENDATIONS                                        3

III      INTRODUCTION                                          11

              PURPOSE AND AUTHORITY                            11

              SUMMARY OF METHODS USED FOR DEVELOPMENT
              OF EFFLUENT LIMITATION GUIDELINES AND
              STANDARDS OF TECHNOLOGY                          13

              SUMMARY OF ORE-BENEFICIATION PROCESSES           17

              GENERAL DESCRIPTION OF INDUSTRY BY ORE
              CATEGORY                                         29

IV       INDUSTRY CATEGORIZATION                              145

              INTRODUCTION                                    145

              FACTORS INFLUENCING SELECTION OF
              SUBCATEGORIES IN ALL METAL ORE CATEGORIES       147

              DISCUSSION OF PRIMARY FACTORS INFLUENCING
              SUBCATEGORIZATION BY ORE CATEGORY               153

              SUMMARY OF RECOMMENDED SUBCATEGORIZATION        174

              FINAL SUBCATEGORIZATION                         174

V        WASTE CHARACTERIZATION                               179

              INTRODUCTION                                    179

              SPECIFIC WATER USES IN ALL CATEGORIES           181

              PROCESS WASTE CHARACTERISTICS BY ORE
              CATEGORY                                        183

VI       SELECTION OF POLLUTANT PARAMETERS                    379

              INTRODUCTION                                    379

              GUIDELINE PARAMETER-SELECTION CRITERIA          379

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SIGNIFICANCE AND RATIONALE FOR SELECTION
OF POLLUTION PARAMETERS                         380

SIGNIFICANCE AND RATIONALE FOR REJECTION
OF POLLUTION PARAMETERS                         405

SUMMARY OF POLLUTION PARAMETERS SELECTED
BY CATEGORY                                     U07
               vi

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                    CONTENTS (VOLUME II)

Section                                                         Page

VII      CONTROL AND TREATMENT TECHNOLOGY                       409

              INTRODUCTION                                      409

              CONTROL PRACTICES AND TECHNOLOGY                  410

              TREATMENT TECHNOLOGY                              426

              EXEMPLARY TREATMENT OPERATIONS BY ORE
              CATEGORY                                          469

VIII     COST, ENERGY, AND NONWATER-QUALITY ASPECTS             582

              INTRODUCTION                                      582

              SUMMARY OF METHODS USED                           583

              WASTEWATER-TREATMENT COSTS FOR IRON-ORE
              CATEGORY                                          589

              WASTEWATER TREATMENT COSTS FOR COPPER-ORE
              CATEGORY                                          596

              WASTEWATER-TREATMENT COSTS FOR LEAD- AND
              ZINC-ORE CATEGORY                                 603

              WASTEWATER-TREATMENT COSTS FOR GOLD-ORE
              CATEGORY                                          614

              WASTEWATER-TREATMENT COSTS FOR SILVER-ORE
              CATEGORY                                          636

              WASTEWATER-TREATMENT COSTS FOR BAUXITE
              CATEGORY                                          646

              WASTEWATER-TREATMENT COSTS FOR FERROALLOY-
              ORE CATEGORY                                      650

              WASTEWATER TREATMENT COSTS FOR MERCURY-
              ORE CATEGORY                                      674

              WASTEWATER TREATMENT COSTS FOR URANIUM-
              ORE CATEGORY                                      686
                            vii

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              WASTEWATER TREATMENT COSTS FOR METAL
              ORES, NOT ELSEWHERE CLASSIFIED                     704

              NON-WATER QUALITY ASPECTS                          718

IX       BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY
         AVAILABLE, GUIDELINES AND LIMITATIONS                   723

              INTRODUCTION                                       723

              GENERAL WATER GUIDELINES                           725

              BEST PRACTICABLE CONTROL TECHNOLOGY
              CURRENTLY AVAILABLE BY ORE CATEGORY
              AND SUBCATEGORY                                    727

X        BEST AVAILABLE TECHNOLOGY ECONOMICALLY
         ACHIEVABLE, GUIDELINES AND LIMITATIONS                  785

              INTRODUCTION                                       785

              GENERAL WATER GUIDELINES                           786

              BEST AVAILABLE TECHNOLOGY ECONOMICALLY
              ACHIEVABLE BY ORE CATEGORY AND SUBCATEGORY         788


XI       NEW SOURCE PERFORMANCE STANDARDS AND
         PRETREATMENT STANDARDS                                  819

              INTRODUCTION                                       819

              GENERAL WATER GUIDELINES                           820

              NEW SOURCE STANDARDS BY ORE CATEGORY               820

              PRETREATMENT STANDARDS                             828

XII      ACKNOWLEDGMENTS                                         833

XIII     REFERENCES                                              835

XIV      GLOSSARY                                                843

XV       BIBLIOGRAPHY                                            869

              LIST OF CHEMICAL SYMBOLS

              CONVERSION TABLE
                            viii

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                               TABLES

No.                             Title
II-l     Summary of Recommended BPCTCA Effluent Limitations
              By Category and Subcategory—Ores for Which
              Separate Limitations Are Proposed                   4
II-2     Summary of Recommended BATEA Effluent Limitations
              By Category and Subcategory~ores for Which
              Separate Limitations Are Proposed                   6
II-3     Summary of Recommended NSPS Effluent Limitations
              By Category and Subcategory—Ores for Which
              Separate Limitations Are Proposed                   8
III-l    Iron-Ore shipments for United States                    31
III-2    Crude Iron-Ore Production for U.S.                      32
III-3    Peagents Used for Flotation of Iron Ores                37
III-4    Various Flotation Methods Available for Pro-
              duction of High-Grade Iron-Ore Concentrate         38
III-5    Total Copper-Mine Production of Ore by Year             43
III-6    Copper-Ore Production from Mines by State  (1972)        43
III-7    Average Copper Content of Domestic Ore                  44
III-8    Average Concentration of Copper in Domestic Ores
              by Process (1972)                                   44
III-9    Copper Ore Concentrated in the United States
              by Froth Flotation, Including LPF Process
              (1972)                                             45
III-10   Copper-Ore Heap or Vat Leached in the
              United States (1972)                               48
III-ll   Average Price Received from Copper in the
              United states                                      51
111-12   Production of copper from Domestic Ore by
              Smelters                                           52
111-13   Mine Production of Recoverable Lead in the
              United States                                      54
111-14   Mine Production of Recoverable Zinc in the
              United states (Preliminary)                         55
111-15   Domestic silver Production from Different
              Types of Ores                                      68
111-16   Silver Produced at Amalgamation and Cyanidation
              Mills in the U.S.  and Percentage of Silver
              Recoverable from All Sources                       69
III-17   Production of Bauxite in the United States              73
111-18   Production of Ferroalloys by U. S. Mining and
              Milling Industry                                   75
111-19   Observed Usage of some Flotation Reagents               87
111-20   Probable Reagents Used in Flotation of Nickel
              and Cobalt Ores                                    91
                             ix

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                           TABLES  (COnt.)

No.                            Title                            Page

111-21   Domestic Mercury Production Statistics                 100
111-22   Isotopic Abundance of Uranium                          105
111-23   Uranium Milling Activity by State, 1972                109
III-2U   Uranium Concentration in IX/SX Eluates                 115
111-25   Decay Series of Thorium and Uranium                    125
111-26   Uranium Milling Processes                              126
111-27   Uranium Production                                     129
III-28   Vanadium Production                                    129
111-29   Vanadium Use                                           129
111-30   Production of Antimony from Domestic sources           131
111-31   Domestic Platinum-Group Mine Production and Value      135
111-32   Production and Mine Shipments of Titanium
              Concentrates from Domestic Ores in the U.S.       140
IV-1     Summary of Industry Subcategorization Recommended      176
IV-2     Final Subcategorization                                177
V-l      Historical Constituents of Iron-Mine Discharges        186
V-2      Historical Constituents of Wastewater from Iron-
              Ore Processing                                    186
V-3      Chemical Compositions of Sampled Mine Waters           187
V-4      Chemical Compositions of Sampled Mill Waters           187
v-5      Chemical Analysis of Discharge 1 (Mine Water)
              and Discharge 2 (Mine and Mill Water) at
              Mine/Mill 1104, Including Waste Loading
              for Discharge 2                                   196
V-6      Chemical Characteristics of Discharge Water
              from Mine 1108                                    199
V-7      Characteristics of Mill 1108 Discharge Water           201
V-8      Principal Copper Minerals Used in the United States    203
V-9      Mine-Water Production from selected Major Copper-
              Producing Mines and Fate(s) of Effluent           206
V-10     Summary of solid Wastes Produced by Plants
              Surveyed                                          207
V-ll     Raw Waste Load in Water Pumped from Selected
              Copper Mines                                      209
V-12     1973 Water Usage in Dump, Heap, and In-Situ
              Leaching Operations                               221
V-13     Chemical Characteristics of Barren Heap, Dump, or
              In-Situ Acid Leach solutions (Recycled:  No
              Waste Load)                                        223
V-14     Water Usage in Vat Leaching Process as a Function
              of Amount of Product (Precipitate or Cathode
              Copper) Produced                                  225
V-15     Chemical Characteristics of Vat-Leach Barren
              Acid Solution (Recycled:  No Waste Load
              Mill 212U)                                        226

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                            TABLES (COnt.)

NO.                             Title

v-16     Miscellaneous Wastes from Special Handling of
              Ore wash Slimes in Mine 2121 (No Effluent)         228
V-17     Examples of Chemical Agents Which May be Employed
              In Copper Flotation                               232
V-18     Water Usage in Froth Flotation of Copper               234
V-19     Raw Mill Waste Loads Prior to Settling in Tailing
              Ponds                                             235
V-20     Range of Chemical Characteristics of Sampled Raw
              Mine Water from Lead/Zinc Mines 3102, 3103,
              and 3104 Showing Low Solubilization               245
V-21     Range of Chemical Characteristics of Raw Mine
              Water from Four Operations Indicating High
              Solubilization Potential                          246
V-22     Range of Constituents of Wastewater and Raw Waste
              Loads for Mills 3101, 3108, and 3109              252
V-23     Characterization of Gold-Placer Raw Wastewater         257
V-24     Chemical Composition of Raw Mine Water from Mines
              4105, 4104, and 4102                              258
V-25     Process Reagent Use at Various Mills Beneficiating
              Gold Ore                                          261
V-26     Minerals Commonly Associated with Gold Ore             261
V-27     Waste Characteristics and Raw Waste Loads at Four
              Gold Milling Operations                           263
V-28     Raw Waste Characteristics of silver Mining
              Operations                                        268
V-29     Major Minerals Found Associated with Silver Ores       270
V-30     Flotation Reagents Used by Three Mills to Bene-
              ficiate Silver-Containing Mineral Tetrahedrite
              (Mills 4401 and 4403) and Native Silver and
              Argentite  (Mill 4402)                             272
V-31     Waste Characteristics and Raw Waste Loads at Mills
              4401, 4402, 4403, and 4105                        273
V-32     Concentrations of Selected Constituents in Acid
              Raw Mine Drainage from Open-Pit Mine 5101         280
V-33     Concentrations of Selected Constituents in Acid
              Raw Mine Drainage from Open-Pit Mine 5102         280
v-34     Concentrations of selected Constituents in Alkaline
              Raw Mine Drainage from Underground Mine 5101      281
V-35     Wastewater and Raw Waste Load for Open-Pit Mine 5101   283
V-36     Wastewater and Raw Waste Load for Underground
              Mine 5101                                         283
V-37     Types of Operations Visited and Anticipated—
              Ferroalloy-Ore Mining and Dressing Industry       284

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                           TABLES  (COnt.)

No.                            Title

V-38     Chemical Characteristics of Raw Mine Water in
              Ferroalloy Industry                                289
V-39     Reagent Use in Molybdenum Mill 6101                    294
V-40     Raw Waste Characterization and Raw Waste Load
              for Mill 6601                                     294
V-41     Reagent Use for Rougher and Scavenger Flotation
              at Mill 6102                                       298
V-42     Reagent Use for Cleaner Flotation at Mill 6102          298
V-43     Reagent Use at Byproduct Plant of Mill 6102  (Based
              on Total Byproduct Plant Feed)                    299
V-44     Mill 6102 Effluent Chemical Characteristics  (Com-
              bined-Tailings Sample)                            299
V-45     Chemical Characteristics of Acid-Flotation Step        300
V-46     Composite Waste Characteristics for Beneficiation
              at Mill 6104 (Samples 6, 8, 9, and 11)            304
V-47     Waste Characteristics from Copper-Thickener Over-
              flow for Mill 6104  (Sample 5)                     304
V-48     Scheelite-Flotation Tailing Waste Characteristics
              and Loading for Mill 6104  (Sample 7)              305
V-49     50-Foot-Thickener Overflow for Mill 6104 (Sample 10)   305
V-50     Waste Characteristics of Combined-Tailing Discharge
              for Mill 6104 (Samples 15, 16, and 17)            306
V-51     Waste Characteristics and Raw Waste Load at Mill
              6105 (Sample 19)                                  308
V-52     Chemical Composition of Wastewater, Total Waste,
              and Raw Waste Loading from Milling and Smelter
              Effluent for Mill 6106                            308
V-53     Waste Characterization and Raw Waste Load for
              Mill 6107 Leach and solvent-Extraction Effluent
              (Sample 80)                                       311
V-54     Waste Characteristics and Waste Load for Dryer
              Scrubber Bleed at Mill 6107  (Sample 81)           312
V-55     Waste Characteristics and Loading for Salt-Roast
              Scrubber Bleed at Mill 6107  (Sample 77)           313
V-56     Expected Reagent Use at Mercury-Ore Flotation
              Mill 9202                                         319
V-57     Waste Characteristics and Raw Waste Loadings at
              Mills 9201 and 9202                               320
V-58     Waste Constituents Expected                            325
V-59     Chemical and Physical Waste Constituents Observed
              in Representative Operations                      326
                            xii

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                           TABLES  (cont.)

No.                            Title                            Page

V-60     Water Use and Flows at Mine/Mills 9401, 9402, 9403,
              and 9404                                          332
V-61     Water Treatment Involved  in U/Ra/V Operations          332
V-62     Padionuclides in Raw Wastewaters from Uranium/
              Radium/Vanadium Mines and Mills                   340
V-63     Organic Constituents in U/Ra/V Raw Waste Water         340
v-64     Inorganic Anions in U/Ra/V Raw Waste Water             341
V-65     Light-Metal Concentrations Observed in U/Ra/V
              Raw Waste Water                                   341
V-66     Concentrations of Heavy Metals Forming Anionic
              Species in U/Ra/V Raw Waste Water                 341
V-67     Concentrations of Heavy Metals Forming Cationic
              Species in U/Ra/V Raw Waste Water                 343
V-68     Other Constituents Present in Raw Wastewater in
              U/Ra/V Mines and Mills                            343
v-69     Chemical Composition of Wastewater and Raw Waste
              Load for Uranium Mines 9401 and 9402              344
V-70     Chemical Composition of Raw Wastewater and Raw
              Waste Load for Mill  9401  (Alkaline-Mill
              Subcategory)                                      344
V-71     Chemical Composition of Wastewater and Raw Waste
              Load for Mill 9402 (Acid- or Combined Acid/
              Alkaline-Mill Subcategory)                        345
v-72     Chemical Composition of Wastewater and Raw Waste
              Load for Mine 9403 (Alkaline-Mill Subcategory)    346
v-73     Chemical Composition of Wastewater and Raw Waste
              Load for Mill 9404 (Acid- or Combined Acid/
              Alkaline-Mill Subcategory)                        347
V-74     Peagent Use at Antimony-Ore Flotation Mill 9901        354
V-75     Chemical Composition of Raw Wastewater Discharged
              From Antimony Flotation Mill 9901                 356
V-76     Major Waste Constituents  and Raw Waste Load at
              Antimony Mill 9901                                357
V-77     Chemical Composition of Raw Wastewater from
              Beryllium Mill 9902  (No Discharge from
              Treatment)                                        359
V-78     Chemical Composition of Raw Wastewater from
              Rare-Earth Mill 9903                              364
V-79     Results of Chemical Analysis for Rare-Earth
              Metals (Mill 9903—No Discharge)                   365
                            xiii

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                           TABLES  (cont.)

No.                            Title                            Page

V-80     Chemical Composition and Raw Waste Load from
              Rare-Earth Mill 9903                              367
V-81     Chemical Composition and Loading for Principal
              Waste constituents Resulting from Platinum
              Mine/Mill 9904 (Industry Data)                    368
V-82     Chemical Composition of Raw Wastewater from
              Titanium Mine 9905                                370
V-83     Chemical Composition of Raw Wastewater from
              Titanium Mill 9905                                373
V-84     Reagent Use in Flotation Circuit of Mill 9905          373
V-85     Principal Minerals Associated with Ore of Mine 9905    374
V-86     Major Waste Constituents and Raw Waste Load at
              Mill 9905                                         374
V-87     Chemical Composition of Raw Wastewater at Mills
              9906 and 9907                             '        377
V-88     Raw Waste Loads for Principal Wastewater Consti-
              tuents from Sand Placer Mills 9906 and 9907       378
VI-1     Known Toxicity of Some Common Flotation Reagents
              Used in Ore Mining and Milling Industry           404
VI-2     Summary of Parameters Selected for Effluent Limi-
              tation by Metal Category                          408
                            xiv

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                              FIGURES


No.                            Title                            Page

III-l    Beneficiation of Iron Ores                              34
III-2    Iron-Ore Flotation-Circuit Flowsheet                    36
III-3    Magnetic Taconite Beneficiation Flowsheet               39
III-4    Agglomeration Flowsheet                                 40
III-5    Major Copper Mining and Milling Zones of the U.S.        42
III-6    General Outline of Methods for Typical Recovery
              of Copper from Ore                                 47
III-7    Major Copper Areas Employing Acid Leaching in
              Heaps, in Dumps, or In Situ                        49
III-8    Lead/Zinc-Ore Mining and Processing Operations          57
III-9    Cyanidation of Gold Ore:  Vat Leaching of Sands
              and •Carbon-in-Pulp1 Processing of Slimes          63
111-10   Cyanidation of Gold Ore;  Agitation/Leach
              Process                                            64
III-11   Flotation of Gold-containing Minerals with
              Recovery of Residual Gold Values by
              Cyanidation                                        65
111-12   Recovery of Silver Sulfide Ore by Froth
              Flotation                                          70
111-13   Gravity-Plant Flowsheet for Nigerian Columbite          83
III-14   Euxenite/Columbite Beneficiation-Plant Flowsheet        84
111-15   Representative Flow Sheet for Simple Gravity
              Mill                                               85
111-16   Simplified Molybdenum Mill Flowsheet                    88
111-17   Simplified Molybdenum Mill Flow Diagram                 89
111-18   Simplified Flow Diagram for Small Tungsten
              Concentrator                                       93
111-19   Mill Flowsheet for a Canadian Columbium
              Operation                                          94
111-20   Flowsheet of Tristage Crystallization Process
              for Recovery of Vanadium, Phosphorus, and
              Chromium from Western Ferrophosphorus              97
111-21   Arkansas Vanadium Process Flowsheet                     98
III-22   Flowsheet of Dean-Leute Ammonium Carbamate
              Process                                            99
III-23   Flow Diagram for Beneficiation of Mercury Ore
              by Flotation                                      102
111-24   Pachuca Tank for Alkaline Leaching                     111
111-25   Concentration Processes and Terminology                117
111-26   Simplified schematic Diagram of Sulfuric Acid
              Digestion of Monazite Sand for Recovery
              of Thorium, Uranium, and Rare Earths              121
                             xv

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                          FIGURES  (cent.)

No.                             Title

111-27   Simplified Schematic Diagram of Caustic soda
              Digestion of Monazite Sand for Recovery
              of Thorium, Uranium, and Rare Earths              122
111-28   Effect of Acidity on Precipitation of Thorium,
              Rare Earths and Uranium from a Monazite/
              Sulfuric Acid Solution of Idaho and
              Indian Monazite Sands                             123
111-29   Generalized Flow Diagram for Production of
              Uranium, Vanadium, and Radium                     127
111-30   Beneficiation of Antimony Sulfide Ore by
              Flotation                                         132
111-31   Gravity Concentration of Platinum-Group Metals         136
111-32   Beneficiation of Heavy-Mineral Beach Sands             142
111-33   Beneficiation of Ilmenite Mined from a Rock Deposit    143
V-l      Flow Scheme for Treatment of Mine Water                189
V-2      water Flow Scheme in a Typical Milling Operation       189
V-3      Water Balance for Mine/Mill 1105 (September 1971)      191
V-4      concentrator Flowsheet for Mill 1105                   193
V-5      Flowsheet for Mill 1104 (Heavy-Media Plant)            197
V-6      Simplified Concentration Flowsheet for Mine/Mill 1108  200
V-7      Wastewater Flowsheet for Plant 2120-B Pit              205
V-8      Flowsheet of Hydrometallurgical Process Used in
              Acid Leaching at Mine 2122                        214
V-9      Reactions by Which Copper Minerals Are Dissolved in
              Dump, Heap, or In-Situ Leaching                   215
V-10     Typical Design of Gravity Launder/Precipitation
              Plant                                             217
V-ll     Cutaway Diagram of Cone Precipitator                   218
V-12     Diagram of Solvent Extraction Process for Recovery
              of Copper by Leaching of Ore and Waste            220
V-13     Vat Leach Flow Diagram (Mill 2124)                      224
V-14     Flow Diagram for Flotation of Copper (Mill 2120)       229
V-15     Addition of Flotation Agents to Modify Mineral
              Surface                                           230
V-16     Flowsheet for Miscellaneous Handling of Flotation
              Tails (Mill 2124)                                 238
V-17     Dual Processing of Ore (Mill 2124)                      239
V-18     Leach/Precipitation/Flotation Process                  240
V-19     Water Flow Diagram for Mine 3105                       243
V-20     Water Flow Diagram for Mine 3104                       249
V-21     Flow Diagram for Mill 3103                             251
                            xvi

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                           FIGURES  (cont.)

No.                              Title                          Page

V-22     Water Flow in Four Selected Gold Mining and
              Milling Operations                                254
V-23     Water Flow in Silver Mines and Mills                   266
V-24     Process and Wastewater Flow Diagram for Open-Pit
              Bauxite Mine 5101                                 277
V-25     Mill 6601 Flow sheet                                   293
V-26     Simplified Mill Flow Diagram for Mill 6102             296
V-27     Internal Water Flow For Mill 6104 Through
              Molybdenum Separation                             302
V-28     Internal Water Flow for Mill 6104 Following
              Molybdenum Separation                             303
V-29     Water Use and Waste Sources for Vanadium Mill 6107     310
V-30     Water Flow in Mercury Mills 9101 and 9102              315
V-31     Typical Water-Use Patterns                             322
V-32     Alkaline-Leach Water Flow                              328
V-33     Ammonium Carbonate Leaching Process                    329
V-34     Water Flow in Mills 9101,  9402, 9403, and 9404         333
V-35     Flow Chart of Mill 9401                                334
V-36     Flow Chart for Mill 9402                               335
V-37     Flow Chart of Mill 9403                                336
V-38     Flow Chart of Mill 9404                                337
V-39     Water Flows and Usage for  Mine/Mills 9901  (Antimony)
              and 9902  (Beryllium)                              350
V-40     Water Flows and Usage for  Mine/Mills 9903
               (Rare Earths) and 9904  (Platinum)                 351
V-41     Water Flows and Usage for  Titanium Mine/Mills
              9905 and 9906                                     352
V-42     Beneficiation of Bertrandite, Mined from a Lode
              Deposit by Flotation  (Mill  9903)                  361
V-43     Beneficiation of Rare-Earth Flotation Concentrate
              by Solvent Extraction (Mill 9903)                 362
V-44     Beneficiation and waste Water Flow of llmenite
              Mine/Mill 9905  (Rock  Deposit)                     372
V-45     Beneficiation of Heavy-Mineral Beach Sands  (Rutile,
               llmenite. Zircon, and Monazite) at Mill 9906      375
                             xvii

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                         SECTION I

                        CONCLUSIONS
To establish effluent limitation guidelines and standards of
performance,  the  ore  mining  and  dressing  industry  was
divided into 41 separate categories  and  subcategories  for
which  separate  limitations  were recommended.  This report
deals with the entire metal-ore mining and dressing industry
and examines the industry by  ten  major  categories:   iron
ore;  copper  ore; lead and zinc ores; gold ore; silver ore;
bauxite ore; ferroalloy-metal ores; mercury  ores;  uranium,
radium  and  vanadium  ores;  and  metal oresr not elsewhere
classified  (ores  of  antimony,  beryllium,  platinum,  rare
earths,     tin,     titanium,    and    zirconium).     The
subcategorization of the ore categories is  based  primarily
upon  ore  mineralogy  and  processing or extraction methods
employed; however, other factors  (such as size,  climate  or
location, and method of mining) are used in some instances.

Based  upon  the application of the best practicable control
technology currently available, mining or milling facilities
in the 12 of 41 subcategories for which separate limitations
are proposed can be operated with no  discharge  of  process
wastewater.  with the best available technology economically
achievable,  facilities in 21 of the 41 subcategories can be
operated  with  no  discharge  of  process   wastewater   to
navigable  waters.   No  discharge  of process wastewater is
also achievable as a new  source  performance  standard  for
facilities in 21 of the 41 subcategories.

Examination  of the wastewater treatment methods employed in
the ore mining and dressing industry indicates that  tailing
ponds  or  other types of sedimentation impoundments are the
most commonly used methods of suspended-solid  removal,  and
that  these  impoundments  provide the additional benefit of
reduction of dissolved parameters as well.  Tailing impound-
ments also  serve to equalize flow rates  and  concentrations
of wastewater parameters.

It is concluded that, for areas of excess water balance, the
practices of runoff diversion, segregation of waste streams,
and reduction in the use of process water will  assist in the
attainment  of no discharge for the specified subcategories.
Effective chemical-treatment methods which  will  result   in
significant  improvement  in  discharge-water  quality  and
pollutant   waste  loads  beyond   those   attained   by   the
application  of  impoundment   and settling are  identified  in
this report.

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                         SECTION II

                      RECOMMENDATION S
The recommended effluent limitation guidelines based on  the
best  practicable  control  technology  currently  available
(BPCTCA)  are summarized in Table II-l.  Eased on information
contained in Sections III through VIII,  it  is  recommended
that  facilities  in  12  of the 41 subcategories achieve no
discharge of process wastewater.

The recommended effluent limitation  guidelines  based  upon
the   best   available  technology  economically  achievable
(BATEA)   are  summarized  in  Table   II-2.    Of   the   41
subcategories  listed  for  which  separate  limitations are
recommended,  it  is  recommended  that  facilities  in   21
subcategories  achieve no discharge of process wastewater by
1984.

The new source performance standards  (NSPS) recommended  for
operations  begun  after  the proposal of recommended guide-
lines  for  the  ore  mining  and  dressing   industry   are
summarized  in  Table  II-3.   With  the  exception  of five
subcategories,  new   source   performance   standards   are
identical   to   BPCTCA   and   BATEA  recommended  effluent
limitations.

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 TABLE 11-1. SUMMARY OF RECOMMENDED BPCTCA EFFLUENT LIMITATIONS BY
           CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
           LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
\ Physical/Chemical Separation
| Magnetic and Physical Separation (Mesabi Range)

X
IX-1
IX-2
COPPER ORES
Mines
Mills
{Open-Pit, Underground, Stripping
Hydrometallurgical (Leaching)
( Vat Leaching
\ Flotation
X
X
IX-3
IX-4
LEAD AND ZINC ORES
Mines
Mills


IX-5
IX-6
GOLD ORES
Mines
Mills
{Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation

X
IX-7
IX-8
IX-9
IX-10
SILVER ORES
Mines
Mills
{Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation

X
IX-11
IX-1 2
IX-1 3
IX-14
BAUXITE ORE
Mines II
IX-15
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BV EPA.

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   TABLE 11-1. SUMMARY OF RECOMMENDED BPCTCA EFFLUENT LIMITATIONS BY
             CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
             LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
( < 5,000 metric tonsVyesr
) > 5,000 metric tons'/year by
I > 5,000 metric tons^/year by
\ Leaching
Physical Processes
Flotation


IX-16
IX-17
IX-18
IX-19
IX-20
MERCURY ORES
Mines
Mills
{Gravity Separation
Flotation Process


X
X
IX-21

URANIUM, RADIUM, VANADIUM ORES
Mines
Mills
{Acid or Acid/Alkaline Leaching
Alkaline Leaching


IX-22
IX-23
ANTIMONY ORES
Mines
Mills
- Flotation Process


X
IX-24

BERYLLIUM ORES
Mines
Mills


X
X


PLATINUM ORES
Mines or Mine/Mi Us II
IX-25
RARE-EARTH ORES
Mines
Mills
— Flotation or Leaching

X
X


TITANIUM ORES
Mines
Mills
( Electrostatic/Magnetic and Gravity/Flotation Processes
\ Physical Processes with Dredge Mining


IX-26
IX-27
IX-28
   '6,000 metric tons - 5,512 short tons
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.

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 TABLE 11-2. SUMMARY OF RECOMMENDED BATEA EFFLUENT LIMITATIONS BY
           CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
           LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
{Physical/Chemical Separation
Magnetic and Physical Separation (Mesabi Range)

X
X-1
X-2
COPPER ORES
Mines
Mills
{Open-Pit. Underground, Stripping
Hydrometallurgical (Leaching)
J Vat Leaching
| Flotation
X
X
X
X-3

LEAD AND ZINC ORES
Mines
Mills

X
X-4

GOLD ORES
Mines
Mills
SCyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation

X
X
X
X-5
(Same as BPCTCA)
SILVER ORES
Mines
Mills
! Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation

X
X
X
X-6
(Same as BPCTCA)
BAUXITE ORE
Mines
x-7
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.

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 TABLE 11-2. SUMMARY OF RECOMMENDED BATEA EFFLUENT LIMITATIONS BY
           CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
           LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
5< 5,000 metric tons'/year
> 5,000 metric tonjt/year by
> 5,000 metric tonst/year by
Leaching
Physical Processes
Flotation


X-8
(Same as BPCTCAl
X-9
X-10
X-11
MERCURY ORES
Mines
Mills
f Gravity Separation
( Flotation Process


X
X
X-12

URANIUM, RADIUM, VANADIUM ORES
Mines
Mills
( Acid or Acid/Alkaline Leaching
( Alkaline Leaching

X
X
X-13

ANTIMONY ORES
Mines
Mills
— Flotation Process


X
(Same as BPCTCA)

BERYLLIUM ORES
Mines
Mills
X
X


PLATINUM ORES
Mines or
Mine/Mills
1
(Same ai BPCTCA)
RARE-EARTH ORES
Mines
Mills
— Flotation or Leaching

X
X


TITANIUM ORES
Mines
Mills
( Electrostatic/Magnetic and Gravity/Flotation Processes
1 Physical Processes with Dredge Mining

X
(Same as BPCTCA)
(Same at BPCTCA)
  5,000 metric tons • 5,512 short tons
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THJS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.

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  TABLE 11-3. SUMMARY OF RECOMMENDED NSPS EFFLUENT LIMITATIONS BY
            CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
            LIMITATIONS ARE PROPOSED (Sheet 1 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
IRON ORES
Mines
Mills
( Physical/Chemical Separation
^ Megnetic end Physical Separation (Mtsabi Range)

X
(Same as BPCTCA)
(SameasBPCTCA)
COPPER ORES
Mines
Mills
( Open-Pit, Underground, Stripping
| Hydrometallurgical (Leaching)
( Vat Leaching
\ Flotation
X
X
X
(Same as BPCTCA)

LEAD AND ZINC ORES
Mines
Mills

X
(Same es BPCTCA)

GOLD ORES
Mines
Mills
SCyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation

X
X
X
(Same as BPCTCA)
(SameasBPCTCA)
SILVER ORES
Mines
Mills
S Flotation Process
Cyantdation Process
Amalgamation Process
Gravity Separation

X
X
X
(SameasBPCTCA)
(Seme as BPCTCA)
BAUXITE ORE
Mines H
	 U 	
(Sam* as BATEA)
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.
                           8

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   TABLE 11-3. SUMMARY OF RECOMMENDED NSPS EFFLUENT LIMITATIONS BY
             CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
             LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
!< 5,000 metric tonst/year
> 5,000 metric tons''' /year by
> 5,000 metric tons'1' /year by
Leaching
Physical Processes
Flotation


XI-1
(Same as BPCTCA)
XI-2
XI-3
(SameasBATEA)
MERCURY ORES
Mines
Mills
( Gravity Separation
^ Flotation Process


X
X
" (Same es BPCTCA)

URANIUM, RADIUM, VANADIUM ORES
Mines
Mills
( Acid or Acid/Alkaline Leaching
{ Alkaline Leaching

X
X
XI-4

ANTIMONY ORES
Mines
Mills
- Flotation Process


X
(Same as BPCTCA)

BERYLLIUM ORES
Mines
Mills
X
X


PLATINUM ORES
Mines or
Mine/Mills

(Same at BPCTCA)
RARE-EARTH ORES
Mines
Mills
— Flotation or Leaching

X
X


TITANIUM ORES
Mines
Mills
( Electrostatic/Magnetic and Gravity/Flotation Procestes
1 Physical Processes with Dredge Mining

X
(Same as BPCTCA)
(Same as BPCTCA)
   5,000 metric tons - 5,512 short tons

NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA.

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                        SECTION III

                        INTRODUCTION
PURPOSE AND AUTHORITY

The United States Environmental Protection Agency  (EPA)   is
charged  under  the Federal Water Pollution Control Act (the
Act)  with establishing effluent limitations  which  must  be
achieved  by  point  sources of discharge into the waters of
the United states.

Section 301(b) of the Act requires the achievement,  by  not
later  than  July 1, 1977, of effluent limitations for point
sources, other than publicly owned  treatment  works,  which
are based on the application of the best practicable control
technology  currently  available  as defined by the Adminis-
trator pursuant to  Section  304(b)  of  the  Act.   Section
301(b) also requires the achievement, by not later than July
1,  1984,  of  effluent limitations for point sources, other
than publicly owned treatment works, which are based on  the
application  of  the  best available technology economically
achievable which will result in reasonable further  progress
toward the national goal of eliminating the discharge of all
pollutants,  as  determined  in  accordance with regulations
issued by the Administrator pursuant to  Section  304(b)   to
the Act.  Section 306 of the Act requires the achievement by
new  sources  of a Federal standard of performance providing
for  the  control  of  the  discharge  of  pollutants  which
reflects the greatest degree of effluent reduction which the
Administrator   determines  to  be  achievable  through  the
application  of  the  best  available  demonstrated  control
technology,   processes,   operating   methods,   or   other
alternatives,  including,  where  practicable,  a   standard
permitting  no  discharge  of pollutants.  Section 304(b) of
the Act requires the Administrator to  publish,  within  one
year  of  enactment of the Act, regulations providing guide-
lines for effluent limitations setting forth the  degree  of
effluent reduction attainable through the application of the
best  practicable control technology currently available and
the degree of  effluent  reduction  attainable  through  the
application  of  the  best  control  measures  and practices
achievable including treatment techniques, process and  pro-
cedure    innovations,    operating    methods   and   other
alternatives.

The  regulations  proposed   herein   set   forth   effluent
limitations guidelines pursuant to Section 304(b) of the Act
                             11

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for  the  Ore  Mining  and  Dressing  Industry  point source
category.

Section 306 of the Act requires  the  Administrator,  within
one  year  after a category of sources is included in a list
published pursuant to Section 306(b)  (1) (A) of the Act,  to
propose   regulations   establishing  Federal  standards  of
performance for new sources within such categories.  Section
307  of  the  Act  requires  the  Administrator  to  propose
pretreatment  standards  for new sources simultaneously with
the promulgation of standards of performance  under  Section
306.   The  Administrator published, in the Federal Register
of January 16, 1973 (38 F.R. 1624),  a  list  of  27  source
Publication  of  an amended list constitutes announcement of
the Administrator's intention of establishing, under Section
306, standards of  performance  applicable  to  new  sources
within  the  ore  mining  and  dressing  industry, and under
Section 307, pretreatment standards.  The list  was  amended
when  regulations  for  the Ore Mining and Dressing Industry
were published in the Federal Register on November  6,  1975
(40 FR 51722).

The  subgroups of the metal mining industries are identified
as major group 10 in the Standard Industrial  Classification
(SIC)  Manual, 1972, published by the Executive Office of the
President  (Office of Management and Budget) .  This industry
category includes establishments engaged in mining ores  for
the  production of metals, and includes all ore dressing and
beneficiating  operations,  whether   performed   at   mills
operating  in  conjunction with the mines served or at mills
operated  separately.    These  include  mills  which  crush,
grind,  wash,  dry, sinter, or leach ore, or perform gravity
separation or flotation operations.

The industry categories covered by this report  include  the
following:

SIC 1011 - Iron Ores
SIC 1021 - Copper Ores
SIC 1031 - Lead and Zinc Ores
SIC 1041 - Gold Ores
SIC 1044 - Silver Ores
SIC 1051 - Bauxite Ores
SIC 1061 - Ferroalloy Ores
SIC 1092 - Mercury Ores
SIC 1094 - Uranium/Radium/Vanadium Ores
SIC 1099 - Metal Ores, Not Elsewhere Classified
                            12

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The  guidelines  in  this document identify, in terms of the
chemical,  physical,  and  biological   characteristics   of
pollutants,  the  level  of  pollutant  reduction attainable
through  application  of  the   best   practicable   control
technology   currently   available,   and   best   available
technology   economically    achievable.     Standards    of
performance  for  new  sources  and  pretreatment  are  also
presented.  The guidelines also consider a number  of  other
factors,  such  as  the  costs  of  achieving  the  proposed
effluent  limitations  and  nonwater  quality  environmental
impacts    (including   energy  requirements  resulting  from
application of such technologies).

SUMMARY  OF  METHODS  USED  FOR  DEVELOPMENT   OF   EFFLUENT
LIMITATION GUIDELINES AND STANDARDS OF TECHNOLOGY

Scope

The  effluent  limitations  guidelines and standards of per-
formance proposed herein  were  developed  in  a  series  of
systematic  tasks.  The Ore Mining and Dressing Industry was
first studied to determine whether separate limitations  and
standards would be appropriate for different SIC categories.
Development    of   reasonable   industry   categories   and
subcategories and establishment of effluent  guidelines  and
treatment  standards require a sound understanding and know-
ledge of the Ore Mining and Dressing  Industry,  the  mining
techniques and milling processes involved, the mineralogy of.
the  ore  deposits,  water  use,  wastewater  generation and
characteristics, and the capabilities  of  existing  control
and treatment technologies.

Approach

This  report describes the results obtained from application
of the above approach to the  mining  and  beneficiating  of
metals  and  ore  minerals  for  the ore mining and dressing
industry.  The survey and sampling and  analysis  covered  a
wide  range of processes, products, and types of wastes.  In
each sic  category, slightly  different  evaluation  criteria
were  applied  initially,  depending  upon the nature of the
extraction  processes  employed,   locations   where   mining
activities  occur,  mineralogical  differences, treatment and
control technology  employed, and water  usage in the industry
category.  The following discussion illustrates  the  manner
in   which   the    effluent   guidelines  and  standards  of
performance were developed.
                             13

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Data Base

Each SIC category was first examined to determine the  range
of  activities  incorporated by the industry classification.
Information used as a data base for detailed examination  of
each  category  was  obtained from a wide variety of sources
including published data from journals and trade literature,
mining industry directories, general business  publications,
texts  on  mining/milling  technology,  texts  on industrial
wastewater  control,  summaries   of   production   of   the
particular  metals  of interest, U.S. Bureau of Mines annual
summaries,    U.S.    Environmental    Protection     Agency
publications,  U.S.  Geological Survey publications, surveys
performed by industry trade associations, NPDES permits  and
permit   applications,   and   numerous  personal  contacts.
Additional information was supplied by surveys  of  research
performed   in   the   application   of  mining,  extractive
processing, and effluent control technology.  Various mining
company personnel, independent researchers,  and  state  and
federal  environmental  officials  also  supplied  requested
information.   In  addition.  Environment  Canada   provided
information  on current practices within the Canadian Mining
and Dressing Industry.

Cateqori zation and Waste Load Characterization

After assembly of an extensive  data  base,  each  SIC  code
group  or subgroup was examined to determine whether differ-
ent limitations and  standards  would  be  appropriate.   In
several  categories, it was determined that further subdivi-
sion was unnecessary.  In addition, after further study  and
site  visits,  subcategory  designations  were later reduced
within a category in  some  instances.   Where  appropriate,
subcategorization  consideration  was based upon whether the
facility was a mine or a concentrating facility (mill) ,  and
further   based   upon  differences  such  as  raw  material
extracted  or  used,  milling   or   concentration   process
employed,  waste  characteristics,  treatability  of wastes,
reagents used in the process, treatment technology employed,
water use and balance, end products  or  byproducts.   Other
factors  considered  were  the  type  of  mine  (surface  or
underground),  geographic  location,  size,   age   of   the
operation, and climate.

Determination  of  the  wastewater usage and characteristics
for  each  subcategory  as  developed  in  Section  IV   and
discussed  in Section V included:  (1) the source and volume
of water used in the particular  process  employed  and  the
source  of  waste  and wastewaters in the plant, and (2)  the

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constituents  (including  thermal)    of   all   wastewaters,
including pollutants, and other constituents which result in
taste, odor, and color in water or aquatic organisms.   Those
constituents discussed in Section V and Section VI which are
characteristic  of  the  industry  and present in measurable
quantities were selected as pollutants subject  to  effluent
limitation guidelines and standards.

Site Visits and Sampling Program

Based upon information gathered as part of the assembly of a
data  base,  examination  of NPDES permits and permit appli-
cations, surveys by trade associations, and  examination  of
texts,  journals,  and the literature available on treatment
practices in the industry, selection of mining  and  milling
operations  which were thought to embody exemplary treatment
practice was made for the purpose of sampling and  verifica-
tion,  and  to supplement compiled data.  All factors poten-
tially   influencing   industry    subcategorization    were
represented  by  the  sites chosen.  Detailed information on
production,  water  use,  wastewater  control,   and   water
treatment  practices  was  obtained.   As  a  result  of the
visits,  many  subcategories  which  had  been   tentatively
determined were found to be unnecessary.  Flow diagrams were
obtained   indicating  the  course  of  wastewater  streams.
Control and treatment plant design and  detailed  cost  data
were compiled.

Sampling  and  analysis of raw and treated effluent streams,
process source water, and intermediate process or  treatment
steps  were  performed  as part of the site visits.  In-situ
analyses for selected parameters such  as  temperature,  pH,
dissolved  oxygen,  and  specific conductance were performed
whenever possible.   Historical  data  for  the  same  waste
streams was obtained when available.

Raw waste characteristics were then identified for each sub-
category.   This included an analysis of all constituents of
wastewaters which might be expected in effluents from mining
and milling operations.  In addition to examination of  can-
didate control parameters, a reconnaissance investigation of
some  55  chemical  parameters  was  performed  upon raw and
treated  effluent  for  each  site  visited.   Additionally,
limited  sampling of mine waters for radiological parameters
was accomplished at selected sites.  Raw and  treated  waste
characterization during this study was based upon a detailed
chemical  analysis  of  the  samples and historical effluent
water quality data supplied by the  industry and Federal  and
State regulatory agencies.
                             15

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Cos-t Data Base

Cost  information  contained  in  this  report  was obtained
directly from industry during plant visits, from engineering
firms, equipment suppliers, and from  the  literature.    The
information  obtained  from  these  sources has been used to
develop general capital, operating  and  overall  costs  for
each  treatment and control method.  Where data was lacking,
costs  were  developed  parametrically  from  knowledge   of
equipment  required,  processes  employed, construction, and
maintenance requirements.  This generalized cost  data  plus
the specific information obtained from plant visits was then
used  for  cost  effectiveness estimates in Section VIII and
wherever else costs are mentioned in this report.

Treatment and Control Technologies

The full range of control and treatment technologies  exist-
ing  within  each subcategory was identified.  This included
an identification of each control and treatment  technology,
including  both  in-plant  and  end-of-process technologies,
which is existent or capable  of  being  designed  for  each
subcategory.   It  also  included  an  identification of the
amounts and the characteristics of pollutants resulting from
the  application  of  each  of  the  control  and  treatment
technologies.  The problems, limitations, and reliability of
each  control and treatment technology were also identified.
In addition, the nonwater-quality environmental impact—such
as the effects of the application of such technologies  upon
other pollution problems, including air, solid waste, noise,
and radiation—was also identified.  The energy requirements
of  each  of  the  control  and  treatment technologies were
identified, as well as the cost of the application  of  such
technologi es.

Selection of BPCTCA, BATEA, and New Source Standards

All data obtained were evaluated to determine what levels of
treatment  constituted  "best practicable control technology
currently available" (BPCTCA),  "best  available  technology
economically  achievable"   (BATEA) ,  and  "best demonstrated
control technology, processes, operating methods,  or  other
alternatives."   several  factors were considered in identi-
fying such technologies.  These included the application  of
costs  of  the  various  technologies  in  relation  to  the
effluent reduction benefits  to  be  achieved  through  such
application,  engineering  aspects  of  the  application  of
various types of control techniques or process changes,  and
nonwater-quality  environmental  impact.   Efforts were also
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made to determine the feasibility of transfer of  technology
from subcategory to subcategory, other categories, and other
industries  where  similar  effluent  problems  might occur.
Consideration of the technologies was not limited  to  those
presently  employed in the industry, but included also those
processes in pilot-plant or laboratory-research stages.

SUMMARY OF ORE-BENEFICIATION PROCESSES

General Discussion

As mined, most  ores  contain  the  valuable  metals,  whose
recovery  is  sought,  disseminated  in  a  matrix  of  less
valuable  rock,  called  gangue.    The   purpose   of   ore
beneficiation   is   the  separation  of  the  metal-bearing
minerals from the gangue to yield a more useful product--one
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  either  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
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, yielding a lower-grade concentrate.)
The process must be optimized, trading off recovery  against
the  value  (and marketability) of the concentrate produced.
Frequently, depending on end use, a particular minimum grade
of concentrate is required,  and  only  limited  amounts  of
specific gangue components are acceptable without penalty.

Many  properties  are  used  as  the  basis  for  separating
valuable minerals from gangue, including:  specific gravity,
conductivity, magnetic permeability,  affinity  for  certain
chemicals,  solubility,  and  the  tendency to form chemical
complexes.  Processes for effecting the  separation  may  be
generally  considered  as:   gravity concentration, magnetic
separation,   electrostatic   separation,   flotation,   and
leaching.   Amalgamation  and  cyanidation  are  variants of
leaching which bear special mention.  Solvent extraction and
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ion exchange are widely applied techniques for concentrating
metals from leaching solutions, and for separating them from
dissolved  contaminants.   All  of   these   processes   are
discussed in general terms—with examples—in the paragraphs
that  follow.   This  discussion  is  not  meant  to be all-
inclusive; rather, its purpose is  to  discuss  the  primary
processes  in  current  use  in  the  ore mining and milling
industry.  Details of processes used in typical  mining  and
milling  operations  are  provided,  together  with  process
flowcharts, under "General Description of  Industry  By  Ore
Category."

Gravity-Concentration Processes

General.       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  viscosity
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  viscosity  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 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 thus depends on  a  feed  to  the  process  which
contains a small dispersion of particle sizes.  A variety of
classifiers—spiral   and  rake  classifiers,  screens,  and
cyclones—is used to assure a reasonably uniform  feed.   At
some mills, a number of sized fractions of ore are processed
in different gravity-separation units.

Viscosity  forces  on  the  particles  set a lower limit for
effective  gravity  separation  by   any   technique.    For
sufficiently  small  particles, even the smallest turbulence
suspends the particle for long periods of  time,  regardless
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of  density.   Such slimes, once formed, cannot be recovered
by gravity techniques and may cause very low  recoveries  in
gravity processing of highly friable ores, such as scheelite
(calcium tungstate, CaWOJ4).


Jigs.    Jigs  of many different designs are used to achieve
gravity separation of relatively coarse  ore  (generally,  a
secondary  crusher product between 0.5 mm and 25 mm--up to 1
in.—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 into
layers by density within the jig, the lighter  gangue  being
drawn off at the top with the water overflow, and the denser
mineral, 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
most  often  used,  in  series,  to  achieve both acceptable
recovery and high concentrate grade.

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 to 2.5  mm—
up  to  0.1  in.—in diameter).  Fundamentally, they are, as
the name implies,  tables  over  which  water  carrying  ore
particles   flows.    A   series   of   ridges  or  riffles,
approximately perpendicular to the water flow,  traps  heavy
particles,  while  lighter ones are suspended by shaking the
table 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 is generally
some material  (termed "middlings") which has  been  diverted
somewhat  by  the  riffles,  although  less  than the heads.
These are often collected separately  and  returned  to  the
table feed.  Reprocessing of either heads or tails, or both,
and multiple stages of tabling are not uncommon.  Tables may
be  used to separate minerals differing relatively little in
density, but uniformity of feed becomes extremely  important
in such cases.

Spirals.    Humphreys spiral separators provide an efficient
means of gravity separation for large  volumes  of  material
between  0.1  mm  and 2 mm (up to approximately 0.01 in.) in
diameter and have been widely applied—particularly/ in  the
processing of heavy sands for ilmenite  (FeTi03) and monazite
(a rare-earth phosphate).  They consist of a helical conduit
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(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.  Nash 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 2.1 metric tons (0.55 to  2.6U  short
tons)  of  ore per hour; in large plants, as many as several
hundred spirals may be run in parallel.

Sink/Float Separation.  Sink/float (heavy media  separation)
separators differ from most gravity methods in that buoyancy
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 whose  density  is  higher
than  that  of the gangue and less than that of 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—often,  by  means  of  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
particle   size  for  practicable  separation,  since  small
particles 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.

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 processing of ores of iron, columbium  and  tantalum,
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and  tungsten,  to  name  a few.  The separation is based on
differences in magnetic permeability (which, although small,
is measurable for almost all materials)  and 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,
behind which is the pole of a large electromagnet,  and  are
carried by it out of the main stream of ore.  As the surface
leaves  the  high-field  region,  the  particles  drop off--
generally, 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, these drum 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.

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  and  are thrown off and
collected, while the  non-conductive  particles  keep  their
charge  and  adhere  by  electrostatic attraction.  They may
then be removed from the drum separately.

Flotation Processes

Basically, flotation is a process whereby 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, then, 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 compatible with the mineral to be recovered be
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used, and that water conditions in the  cell  not  interfere
with attachment of reagents to mineral or to air bubbles.

Flotation  concentration  has  become  a mainstay of the ore
milling industry.  Because it  is  adaptable  to  very  fine
particle sizes (less than 0.001 cm), it allows high rates of
recovery  from  slimes,  which  are  inevitably generated in
crushing and grinding and which are not  generally  amenable
to  physical  processing.   As  a  physico-chemical  surface
phenomenon, it can often be made highly  specific,  allowing
production  of  high-grade  concentrates from very-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), where desired, and
operation with minimum reagent  consumption,  since  reagent
interaction  is typically only with the particular materials
to be floated or depressed.

Details of the flotation process—exact suite and dosage  of
reagents,  fineness  of grinds, number of regrinds, cleaner-
flotation steps, etc.—differ at each operation where it  is
practiced  and  may often vary with time at a given mill.  A
complex system of reagents is generally used, including five
basic types  of  compounds:   pH  conditioners  (regulators,
modifiers),    collectors,    frothers,    activators    and
depressants.  Collectors serve to attach  ore  particles  to
air   bubbles   formed  in  the  flotation  cell.    Frothers
stabilize  the  bubbles  to  create  a  foam  which  may  be
effectively  recovered  from  the water surface.  Activators
enhance the attachment of the collectors to  specific  kinds
of   particles  and  depressants  prevent  it.   Frequently,
activators are used to allow flotation of ore  depressed  at
an  earlier  stage  of  the  milling process.  In almost all
cases, use of each reagent in the mill  is  low  (generally,
less  than  0.5  kg—approximately  1  Ib—per  ton  of  ore
processed), and the bulk of the reagent adheres to  tailings
or concentrates.

Sulfide  minerals  are  all  readily  recovered by flotation
using similar reagents  in  small  doses,  although  reagent
requirements  and  ease  of flotation do 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  hydroxyl group and a "non-wettable" hydrocarbon.
Sodium cyanide  is  widely  used  as  a  pyrite  depressant.
Activators  useful  in  sulfide-ore  flotation  may  include
cuprous  sulfide   and   sodium   sulfide.    Other   pyrite
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depressants  which  are less damaging to the environment may
be used to replace the sodium cyanide.  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,
manganese,   the   rare   earths,  tungsten,  titanium,  and
columbium and tantalum, for example,  may  be  processed  in
this way.  Flotation of these ores involves a very different
suite  of  reagents  from sulfide flotation and has, in some
cases, required substantially  larger  dosages.   Experience
has  shown  these  flotation  processes  to  be, in general,
somewhat  more  sensitive  to  feed-water  conditions   than
sulfide   floats;   consequently,  oxidized  ores  are  less
frequently run with recycled water.  Reagents  used  include
fatty  acids  (such  as  oleic acid or soap skimmings), fuel
oil, and various amines as collectors; and compounds such as
copper sulfate,  acid  dichromate,  and  sulfur  dioxide  as
conditioners.

Leaching

General.    Ores  can  be  leached by dissolving away either
gangue or values in aqueous acids or bases,  liquid  metals,
or  other  special  solutions.   The  examples  which follow
illustrate various possibilities.

    (1)  water-soluble compounds of sodium,  potassium,  and
         boron  which  are  found  in arid climates or under
         impervious strata can be mined,  concentrated,  and
         separated  by  leaching  with  water and recrystal-
         lizing the resulting brines.

    (2)  Vanadium and some other metals form anionic species
          (e.g., vanadates) which occur  as  insoluble  ores.
         Roasting   of   such  insoluble  ores  with  sodium
         compounds converts the  values  to  soluble  sodium
         salts   (e.g., sodium vanadate).  After cooling, the
         water-soluble sodium salts  are  removed  from  the
         gangue by leaching in water.

    (3)  Uranium ores are only mildly soluble in water,  but
         they   dissolve   quickly   in   acid  or  alkaline
         solutions.

         Native gold which is  found  in  a  finely  divided
         state is soluble in mercury and can be extracted by
         amalgamation   (i.e., leaching with a liquid metal).
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         One  process  of  nickel   concentration   involves
         reduction  of  the nickel by ferrosilicon at a high
         temperature and extraction of the nickel metal into
         molten iron.  This process, called skip-ladling, is
         related to liquid-metal leaching.

    (5)   Certain   solutions   (e.g.,   potassium   cyanide)
         dissolve  specific  metals  (e.g.,  gold)   or their
         compounds,  and  leaching   with   such   solutions
         immediately concentrates the values.

Leaching  solutions  can  be  categorized as strong, general
solvents (e.g., acids) and weaker, specific solvents  (e.g.,
cyanide).   The acids dissolve certain 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 ground  very  fine,  and  separation  of  the
tailings from the value-bearing (pregnant) leach is then not
difficult.    In  the  case  of  sulfuric  acid, the leach is
cheap, but  energy  is  wasted  in  dissolving  unsought-for
gangue constituents.

Specific  solvents  attack only one  (or, at most, a few) ore
constituent(s), including the one being sought.  Ore must be
ground finer to expose the  values.   Heat,  agitation,  and
pressure  are  often  used to speed the action of the leach,
and considerable effort goes  into  separation  of  solids—
often, in the form of slimes—from the pregnant leach.

Countercurrent  leaching,  preneutralization  of lime in the
gangue,  leaching  in  the  grinding  process,   and   other
combinations  of  processes  are often seen in the industry.
The values contained in  the  pregnant  leach  solution  are
recovered by one of several methods, including precipitation
(e.g.,  of  metal hydroxides from acid leach by raising pH),
electrowinning  (which is  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  takes  place  in  the  ore body, with the leaching
solution  applied  either  by  plumbing  or  by  percolation
through  overburden.   The pregnant leach solution is pumped
to the recovery facility and can often be recycled.  In-situ
leaching is most economical when the ore body is  surrounded
by  impervious  strata.   When  water  suffices  as  a leach

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solution and is plentiful, in-situ leaching  is  economical,
even  in  pervious  strata.   Ore  or tailings stored on the
surface can be treated by heap or dump  leaching.   In  this
process,  the  ore is placed on an impervious layer (plastic
sheeting or clay)  that  is  furrowed  to  form  drains  and
launders   (collecting   troughs),  and  leach  solution  is
sprinkled over the resulting heap.  The launder effluent  is
treated  to  recover  values.   Gold  (using cyanide leach),
uranium using  (sulfuric  acid  leach),  and  copper   (using
sulfuric  acid  or acid ferric sulfate leach) , are recovered
in this fashion.

Amalgamation.   Amalgamation is the process by which mercury
is alloyed with some other  metal  to  produce  an  amalgam.
This  process  is  applicable to free milling precious-metal
ores, which are those in which the gold is free,  relatively
coarse,  and has clean surfaces.  Lode or placer gold/silver
that is  partly  or  completely  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 has, in the past, been used extensively
for  the extraction of gold and silver from pulverized ores,
it has, due to environmental  considerations,  largely  been
superseded, in recent years, by the cyanidation process.

The  properties  of  mercury  which make amalgamation  such a
relatively simple and efficient process are:   (a)  its  high
specific  gravity   (13.55  at 20 degrees Celsius, 68 degrees
Fahrenheit);  (b) the fact that mercury is a liquid  at  room
temperature;  and  (c) the  fact that it readily wets  (alloys)
gold and silver in the presence of water.

In the past,  amalgamation  was  frequently  implemented  in
specially  designed boxes containing  plates  (e.g., sheets of
metal such as copper or Muntz  metal   (Cu/Zn  alloy),  etc.)
with  an  adherent  film of mercury.   These boxes, typically,
were located downstream of the  grinding  circuit,  and  the
gold  was seized from the pulp as it  flowed over the amalgam
plates.  In the U.S., this process  has  been  abandoned  to
prevent stream pollution.

The  current practice of amalgamation in the U.S. is limited
to barrel amalgamation of  a  relatively  small   quantity  of
high-grade,   gravity-concentrated    ore.    This    form  of
amalgamation  is  the simplest method cf treating  an  enriched
gold-   or    silver-   bearing   concentrate.    The  gravity
concentrate  is ground for  several hours in an amalgam  barrel
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(e.g., a small cylinder batching mill) with steel  balls  or
rods  before  the  mercury  is  added.  This mixture is then
gently ground to bring the mercury and  gold  into  intimate
contact.   The  resulting  amalgam is collected in a gravity
trap.

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, concentrates, 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	> UNaAu(CN) 2 + INaOH

and  subsequently  sorbed onto activated carbon ("Carbon-in-
Pulp" process)  or precipitated with metallic zinc  according
to the reaction:

    2_NaAu(CN)2! + Zn	>  Na2Zn(CN)jl * 2Au

The  gold  particles  are  recovered  by  filtering, and the
filtrate is returned to the leaching operation.

A recently developed process to recover  gold  from  cyanide
solution  is  the  Carbon-in-Pulp process.  This 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 6 x 16
mesh activated coconut charcoal in a series of  tanks.   The
pulp  and  enriched  carbon are air lifted and discharged on
small vibrating screens between tanks, where the  carbon  is
separated  and  moved  to the next adsorption tank, counter-
current to the pulp flow.  Gold  enriched  carbon  from  the
last  adsorption  tank  is  leached with hot caustic cyanide
solution to desorb the gold.  This hot, high-grade  solution
containing  the  leached  gold  is then sent to electrolytic
cells,  where  the  gold  and  silver  are  deposited   onto
stainless  steel  wool cathodes.  The cathodes are then sent
to the refinery for processing.

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 then is leached in  vats  with
dilute,  well aerated cyanide solution.  The slime fraction,
after  thickening,  is  treated  by  agitation  leaching  in
mechanically   or  air  agitated  tanks,   and  the  pregnant
                            26

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solution is separated from the slime residue  by  thickening
and/or  filtration.  Alternatively, the entire finely ground
ore  pulp  may  be  leached  by  countercurrent  decantation
processing.   Gold  or  silver  is  then  recovered from the
pregnant leach solutions by the methods discussed above.

Different types of gold/silver ore require  modification  of
the  basic  flow  scheme  presented  above.  At one domestic
operation, the ore is carbonaceous  and  contains  graphitic
material,  which  causes  dissolved  gold to adsorb onto the
carbon, thus causing premature precipitation.  To make  this
ore   amenable  to  cyanidation,  the  refractory  graphitic
material is oxidized by  chlorine  treatment  prior  to  the
leaching  step.   Other  schemes  which  have  been employed
include oxidation by roasting and blanking the  carbon  with
kerosene  or  fuel  oil  to  inhibit adsorption of gold from
solution.

Other refractory ores  are  those  which  contain  sulfides.
Roasting to liberate the sulfide-enclosed gold and precondi-
tioning  by  aeration with lime of ore containing pyrrhotite
are two processes which allow  conventional  cyanidation  of
these ores.

The  cyanidation  process  is  comparatively  simple, and is
applicable to many types of gold/silver ore,  but  efficient
low-cost dissolution and recovery of the gold and silver are
possible  only  by  careful  process  control  of  the  unit
operations  involved.   Effective  cyanidation  depends   on
maintaining and achieving several conditions:

    (1)  The gold and silver must  be  adequately  liberated
         from  the encasing gangue minerals by grinding and,
         if necessary, roasting or chemical oxidation.

    (2)  The concentration of "free" cyanide  and  dissolved
         oxygen  in  the leaching solution must be kept at a
         level that will enable reasonably fast  dissolution
         of the gold and silver.

    (3)  The "protective" alkalinity of the  leach  solution
         must  be  maintained  at a level that will minimize
         consumption of cyanide by the dissolution of  other
         metal-bearing minerals.

    (U)  The  leach  residues  must  be  thoroughly   washed
         without   serious  dilution  to  reduce  losses  of
         dissolved values and cyanide to acceptable limits.
                            27

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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 structure — typically, C1-, NO3-,  HSO4-,
or  C03_ — (anions) or H* or Na* (cations) —for an ion with a
greater charge or a smaller ionic radius.  For example,  let
R  be  the remainder of the polar molecule (in the case of a
solvent)  or polymer (for a resin) , and let X be  the  mobile
ion.   Then,  the  exchange  reaction for the example of the
uranyltrisulfate complex is:

    URX + (002(801)3)   ---- >    R4U02(S04)3 * 4X-
                       < ----

This reaction proceeds from left to  right  in  the  loading
process.    Typical  resins adsorb about ten percent of their
mass in  uranium  and  increase  by  about  ten  percent  in
density.    In a concentrated solution of the mobile 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:
                         >  Mg*+  >

and, because of decreasing ionic radius, with atomic number:

                   92U  >   42MO   >   23V

and the separation of hexavalent 92D 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) almost always appear in  aqueous  solutions
as  oxidized  ions (uranyl, vanadyl, or molybdate radicals),
with  uranium  and  vanadium  additionally  complexed   with
anionic radicals tc form trisulfates or tricarbonates in the
leach.  The complexes react anionically, and 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, including pH and reduction/oxidation  potential,
of  the pregnant solutions influence the adsorption of heavy
metals.  For example, seven times more vanadium than uranium
                            28

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was adsorbed on one resin at pH 9; at pH 11, the  ratio  was
reversed,  with  33  times as much uranium as vanadium being
captured.  These variations in affinity,  multiple  columns,
and  control  of  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 the case of solvent extraction, the type of polar solvent
and its concentration in a typically nonpolar diluent  (e.g.,
kerosene) affect separation of  the  desired  product.   The
ease  with  which  the  solvent  is handled permits the con-
struction  of  multistage,  cocurrent  and   countercurrent,
solventextraction  concentrators  that  are useful even when
each stage effects only partial separation of a  value  from
an  interferent.   Unfortunately,  the  solvents  are  easily
polluted by slimes, and complete liquid/solid separation  is
necessary.   lonexchange and solvent-extraction circuits can
be combined to take advantage of  the  slime  resistance  of
resin-in-pulp  ion exchange and of the separatory efficiency
of solvent extraction  (Eluex process).

GENERAL DESCRIPTION OF INDUSTRY BY ORE CATEGORY

The ore groups categorized in SIC groups 1011,  1021,  1031,
1041,    1044,   1051,   1061,  1092,  1094,  and  1099  vary
considerably in terms of their  occurrence,  mineralogy  and
mineralogical   variations,  extraction  methods,  and end-
product  uses.   For  these  reasons,  these  industry  areas
generally are treated separately  except for groups SIC 1061,
Ferroalloys   (members of which are differently occurring ore
minerals but are classed as one group), and SIC 1099,  Metal
Ores,  Not  Elsewhere Classified  (a grouping of ore minerals
whose  mining   and   processing   operations   bear    little
resemblance to one another).

Iron Ore

American iron-ore shipments  increased from 82,718,400  metric
tons   (91,200,000  short  tons) in 1968 to 92,278,180  metric
tons  (101,740,000 short tons) in  1973,  an  increase of  11.56%
 (Reference  1).   in   this   period,   the    shipments   of
agglomerates, most of which were  produced  by processing low-
grade    iron    formations,    increased   by   19.1%.   Total
consumption of  iron ore  in  the United   states   in 1973  was
139,242,640 metric tons  (153,520,000 short tons), with 76.5%
produced domestically.  Domestic agglomerates accounted for
66,256,350  metric tons  (73,050,000  short tons), or 47.6%  of
United  States   consumption.   A  summary   of   U.S.  iron-ore
                             29

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shipments is shown in Table Ill-l.   A  breakdown  of  crude
iron-ore  production in the U.S. is shown in Table III-2.  A
breakdown of U.S. iron-ore shipments by producing company is
given in Supplement B to this document.  Except for  a  very
small tonnage, iron ores are beneficiated before shipping.

Beneficiation of iron ore includes such operations as crush-
ing, screening, blending, grinding, concentrating, classify-
ing,  briquetting,  sintering and agglomerating and is often
carried on at or near the mine site.  Methods  selected  are
based  on physical and chemical properties of the crude ore.
A noticeable trend has been developing in furthering efforts
to  use  lower-grade  ores.   As  with  many  other  natural
resources,  future  availability will largely be a matter of
cost rather than of absolute depletion as these  lower-grade
ores are utilized.  Benefication methods have been developed
to  upgrade  20-30%  iron  «taconite1  ores  into high-grade
materials.

In most cases, open-pit mining is more economical than  con-
ventional  underground methods.  It provides the lowest cost
operation and is employed whenever the ratio  of  overburden
(either  consolidated  or  unconsolidated)   to  ore does not
exceed an economical limit.  The  depth  to  which  open-pit
mining   can  be  carried  depends  en  the  nature  of  the
overburden   and   the   stripping    ratio    (volume    of
overburden/crude   ore).   Economic  stripping  ratios  vary
widely from mine to mine  and  from  district  to  district,
depending  upon  a number of factors.  In the case of direct
shipping ores, it may be as high as 6 or 7 to 1; in the case
of taconite, a stripping ratio of less than  1/2  to  1  may
become  necessary.   Stripping  the  overburden necessitates
continually cutting back the pit walls to  permit  deepening
of  the  mine  to recover ore in the bottom.  Power shovels,
draglines,  power  scrapers,  hydraulicking,  and  hydraulic
dredging  are  used  to  recover ore deposits.  Drilling and
blasting  are  usually  necessary  to  remove   consolidated
overburden  and  to loosen ore banks directly ahead of power
shovels.  Iron ore is loaded into buckets  ranging  in  size
from  0.75  to  7.5 cubic meters  (1 to 10 cubic yards).  The
ore is transported out of the pit by railroad cars,  trucks,
truck   trailers,   belt   conveyors,   skip  hoists,  or  a
combination of these.  It is then transferred to a  crushing
plant  for  size reduction, to a screening plant for sizing,
or to a concentrating plant for treatment  by  washing   (wet
size  classification  and  tailings rejection) or by gravity
separation.
                           30

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         TABLE 111-1. IRON-ORE SHIPMENTS FOR UNITED STATES
                       a. QUANTITIES SHIPPED BY REGION
REGION
real Lakes
ortheastern
outhem
'•stern
OTAL U.S.

REGION
Great Lakes
Northeastern
Southern
Western
TOTAL U.S.
AMOUNT SHIPPED
1968
METRIC TONS
66,093,239
3,602.706
3.474,203
10,566,860
82,736,905
LONG TONS
64.065,185
3,545,806
3,419,333
10,399,972
81,430,195
1969
METRIC TONS
72,534,630
3,453,486
4,733,087
10,454,364
91,175,667
LONG TONS
71,389,050
3,398,943
4.658,335
10,289.252
89.735,580

1970
METRIC TONS
70,180.666
3,043357
5,022.389
10,544.782
88,791.674
LONG TONS
69.072,263
2,996.784
4.943,048
10,378,242
87.389,337

AMOUNT SHIPPED
1971
METRIC TONS
62.766.873
2.859,973
4,240,720
8.253,243
78.120,810
LONG TONS
61,775,561
2,814,804
4.173,744
8.122,895
76.887,004
1972
METRIC TONS
65,759,357
2,362,067
4,032,651
7,397,816
79,637,152
LONG TONS
64,720.783
2,324,762
3,968,961
7,266,471
78,280.977
1973
METRIC TONS
77,504.865
2,406.456
3,923.618
8,462.579
92,296.418
LONG TONS
78,280,787
2.367,465
3.861 ,552
8.328,925
90,838,729
    b. SHIPMENTS FROM GREAT LAKES REGION AS PERCENTAGES OF TOTAL US. SHIPMENTS

YEAR

1968
1969
1970
1971
1972
1973
GREAT LAKES SHIPMENTS
AS PERCENTAGE OF
TOTAL U.S. SHIPMENTS
78.7
79.6
79.0
80.4
82.7
84.0
AGGLOMERATES AS
PERCENTAGE OF
GREAT LAKES SHIPMENTS
61.9
63.6
66.2
70.1
74.8
73.5
GREAT LAKES AGGLOMERATES
AS PERCENTAGE OF TOTAL
U.S. SHIPMENTS
48.7
50.6
52.3
56.3
61.8
61.7
                    c. PERCENTAGES OF TOTAL U.S. SHIPMENTS


Direct Shipping
Coarse Ores
Fine Ores
Screened Ores
Concentrates
Agglomerates

YEAR
1968
8.2


3.2
28.3
60.3
100.0
1969
7.0


3.1
27.5
62.4
100.0
1970
5.0


2.7
28.2
64.1
100.0
1971
4.3


3.1
23.7
68.9
100.0
1972
2.0
12.8
11.9


73.3
100.0
1973
2.4
12.9
12.9


71.8
100.0
SOURCE:  Reference 1
                               31

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           TABLE III 2. CRUDE IRON-ORE PRODUCTION FOR U.S.
                          a. QUANTITIES PRODUCED
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
GREAT LAKES
METRIC TONS
159,349,027
169,328,525
172,799,898
161,947.509
158,183,907
186,627,840
LONG TONS
156,832,339
166,654,225
170,070,772
159,389,781
155,685,620
183,680,322
NORTHEASTERN
METRIC TONS
10.236,712
9,728,661
9,173,800
7.774,210
6.721,672
6,915,338
LONG TONS
10,075,038
9,575,01 1
9,028,913
7,651,428
6,615,513
6,806.120
SOUTHERN
METRIC TONS
7,743.542
9,135,951
10,542,987
9,414,016
9,333.043
8,629,278
LONG TONS
7,621,244
8,991,662
10.376,387
9,265,335
9.185,641
8,492,991
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
WESTERN
METRIC TONS
19,671,003
19.270.778
19,981,771
18.422,861
13,347,447
18.080,995
LONG TONS
19,360,328
18,966,424
19,666,188
18,131,898
13,136,643
17,795.432
TOTAL U.S. PRODUCTION
METRIC TONS
197,000,285
207,463.916
212,498.366
197.558,596
187,586,069
220,253.451
LONG TONS
193,888,949
204,187,322
209,142,260
194,438,442
184,623,417
216,774,865
                b. PERCENTAGE OF U.S. CRUDE IRON-ORE PRODUCTION
RCSSirtM

Great Lakes
Northeastern
Southern
Western

YEAR
1968
80.9
5.1
4.0
10.0
100.0
1969
81.6
4.7
4.4
9.3
100.0
1970
81.3
4.3
5.0
9.4
100.0
1971
82.0
3.9
4.8
9.3
100.0
1972
84.3
3.6
5.0
7.1
100.0
1973
84.7
3.2
3.9
8.2
100.0
SOURCE: Reference 1
                              32

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Special problems are associated with the mining of taconite.
The extreme hardness  of  the  ore  necessitates  additional
drilling/blasting  operations  and  specialized, more rugged
equipment.  The low  iron  content  irakes  it  necessary  to
handle  two or four times as much mined material to obtain a
given quantity of iron  as  compared  to  higher  grade  ore
deposits.

Water  can cause a variety of problems if allowed to collect
in mine workings.  Therefore, means  must  be  developed  to
collect  water  and  pump it out of the mine.  This drainage
water is often used directly to make up for water losses  in
concentration operations.

Underground  methods are utilized only when stripping ratios
become too high  for  economical  open-pit  mining.   Mining
techniques  consist  of  sinking vertical shafts adjacent to
the deposit but far enough away  to  avoid  the  effects  of
surface   subsidence   resulting   from  mining  operations.
Construction of shafts,  tunnels,  underground  haulage  and
development   workings,  and  elaborate  pumping  facilities
usually requires expensive capital investments.   Production
in  terms  of iron ore/day is much lower than in the case of
open-pit production, necessitating the presence of very high
grade  ores  for  economic  recovery.   General   techniques
utilized in the beneficiation of iron ore are illustrated in
Figure  Ill-l.   Processes  enhance  either  the chemical or
physical characteristics of  the  crude  ore  to  make  more
desirable feed for the blast furnace.

Crude  ore  not  requiring further processing may be crushed
and screened in order to eliminate handling problems and  to
increase  heat transfer and, hence, rate of reduction in the
blast furnace.  Blending produces a more uniform product  to
comply with blast furnace requirements.

Physical  concentrating processes such as washing remove un-
wanted sand, clay, or rock from  crushed  or  screened  ore.
For  those  ores  not amenable to simple washing operations,
other physical methods such as jigging, heavy-media  separa-
tion,  flotation, and magnetic separation are used.  Jigging
involves  stratification of ore and gangue by pulsating water
currents.  Heavy-media separation employs a water suspension
of ferrosilicon in which iron ore particles sink  while  the
majority  of  gangue   (quartz,  etc.)  floats.   Air bubbles
attached  to  ores  conditioned  with   flotation   reagents
separate  out  iron  ore during the flotation process, while
magnetic  separation  techniques   are   used   where   ores
containing magnetite are encountered.
                             33

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Figure MM. BENEFICIATION OF IRON ORES
                 ORE
             CRUSHING AND
              SCREENING
CONCENTRATING PROCESSES:
PHYSICAL
| CHEMICAL
               I
            MAGNETIC
           SEPARATION
              I
      HEAVY-
      MEDIA
    SEPARATION
            AGGLOMERATION
              PROCESSES
      PELLETIZING
NODULIZING
                        T
BRIQUETTING
      TO STOCK PILE AND/OR SHIPPING
             34

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At the present time, there are only three iron ore flotation
plants  in  the  United  states.  Figure III-2 illustrates a
typical flowsheet used in an  iron  ore  flotation  circuit,
while  Table  III-3  lists  types  and  amounts of flotation
reagents used per ton of ore processed.   Various  flotation
methods  which  utilize  these  reagents are listed in Table
III-4.   The  most  commonly  adopted  flowsheet   for   the
beneficiation  of low grade magnetic taconite ores is illus-
trated in Figure III-3.  Low grade ores containing magnetite
are very susceptible to concentrating processes, yielding  a
high   quality   blast  furnace  feed.   Higher  grade  ores
containing hematite cannot be upgraded much above 55% iron.

Agglomerating processes follow concentration operations  and
increase  the  particle size of iron ore and reudces "fines"
which normally would be lost in the flue gases.   Sintering,
pelletizing,  brig'uetting,  and  nodulizing are all possible
operations involved in  agglomeration.   Sintering  involves
the  mixing of small portions of coke and limestone with the
iron ore,  followed  by  combustion.   A  granular,  coarse,
porous   product   is   formed.   Pelletizing  involves  the
formation of pellets or balls of iron ore fines, followed by
heating.  (Figure III-H illustrates  a  typical  pelletizing
operation.)    Nodules  or  lumps  are  formed  when ores are
charged into a rotary kiln and heated  to  incipient  fusion
temperatures in the nodulizing process.  Hot ore briquetting
requires  no  binder,  is  less sensitive to changes in feed
composition, requires little or  no  grinding  and  requires
less  fuel  than sintering.  Small or large lumps of regular
shape are formed.

Copper Ore

The copper ore segment of the ore mining and dressing indus-
try includes facilities  mining  copper  from  open-pit  and
underground  mines, and those processing the ores and wastes
by hydrometallurgical  and/or  physical-chemical  processes.
Other  operations  for  processing  concentrate  and  cement
copper, and  for  manufacturing  copper  products   (such  as
smelting,  refining,  rolling,  and  drawing) are classified
under other SIC codes and are covered under limitations  and
guidelines  for those industry classifications.  However, to
present a comprehensive view of the history  and  statistics
of  the  copper  production in the United States, statistics
pertaining to finished copper are included  with  those  for
ore production and beneficiation.

Evidence  of the first mining of copper in North America, in
the  upper  Peninsula  of  Michigan,  has  been   found   by
                             35

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Figure III 2.  IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
              DENSIFYING THICKENER
                   UNDERFLOW
                  CONDITIONERS
                    1
               ROUGHER FLOTATION
             ROUGHER
               TAIL
TO
TAILING
BASIN
          ROUGHER
        CONCENTRATE
          (10 CELLS)
     I
  FROTH OF
FIRST 2 CELLS

I
CLEANER FLOTATION
CLEANER
TAIL CLE/
| CONCEIT
	 ^

I
FROTH OF
USIER FIRST 2 CELLS
JTRATE | _
(8 CELLS)

RECLEANER
I
RECLEANER
TAIL
FLOTATION
I
RECLEANER
CONCENTRATE
(7 CELLS)
I

                                               DISC
                                              FILTER
                                              TOTAL
                                             FLOTATION
                                           CONCENTRATE
                                                J
                                        TO AGGLOMERATION
                                           (FIGURE III-4)
                    36

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         TABLE  111-3. REAGENTS USED FOR FLOTATION OF  IRON ORES
(Reagent quantities represent approximate maximum usages.  Exact chemical composition of reagent
may be unknown.)


1.   Anionic Flotation of Iron Oxides (from crude ore)

     Petroleum sulfonate: 0.5 kg/metric ton (1 Ib/short ton)
     Low-rosin, tall oil fatty acid: 0.25 kg/metric ton (0.5 Ib/short ton)
     Sulfuric acid: 1.25 kg/metric ton (2.5 Ib/short ton) to pH3
     No. 2 fuel oil: 0.15 kg/metric ton (0.3 Ib/short ton)
     Sodium silicate:  0.5 kg/metric ton (1 Ib/short ton)


2.   Anionic Flotation of Iron Oxides (from crude ore)

     Low-rosin tall oil fatty acid:  0.5 kg/metric ton (1  Ib/short ton)


3.   Cationic  Flotation of Hematite (from crude ore)

      Rosin amine acetate: 0.2 kg/metric ton (0.4 Ib/short ton)
     Sulfuric acid: 0.15 kg/metric ton (0.3 Ib/short ton)
     Sodium fluoride: 0.15 kg/metric ton (0.3 Ib/short  ton)
     (Plant also includes phosphate flotation and pyrite flotation steps. Phosphate flotation employs
     sodium hydroxide, tall oil fatty acid, fuel oil, and sodium silicate. Pyrite flotation employs
     xanthate  collector.)


4.   Cationic  Flotation of Silica (from crude ore)

     Amine: 0.15 kg/metric ton (0.3 Ib/short ton)
     Gum  or starch (tapioca fluor):  0.5 kg/metric ton (1 Ib/short ton)
     Methylisobutyl carbinol:  as required


5.   Cationic  Flotation of Silica (from magnetite concentrate)

     Amine:  5 g/metric ton  (0.01 Ib/short ton)
     Methylisobutyl carbinol:  as required
                                          37

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TABLE 111-4. VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
             OF HIGH-GRADE IRON-ORE CONCENTRATE
             1.   Anionic flotation of specular hematite
            2.   Upgrading of natural magnetite concentrate by cationic flotation
            3.   Upgrading of artificial magnetite concentrate by cationic flotation
            4.   Cationic flotation of crude magnetite
            5.   Anionic flotation of silica from natural hematite
            6.   Cationic flotation of silica from non-magnetic iron formation
                                    38

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  Figure 111-3. MAGNETIC TACONITE BENEFICIATION FLOWSHEET
                    CRUSHED CRUDE ORE

                           i
                          I
               COBBER MAGNETIC SEPARATION
            CONCENTRATE
                 i
           I BALLMILL|
     CLEANER MAGNETIC SEPARATION
            CONCENTRATE


                 I
           HYDROCYCLONE
        OVERSIZE    UNDERSIZE

           I   	L_
                HYDROSEPARATOR
            CONCENTRATE


                 *
     FINISHER MAGNETIC SEPARATION
CONCENTRATE
 THICKENING
     1
     LTI

     T
TO PELLETIZING

 (FIGURE 111-1)
                                                TO TAILING BASIN
                          39

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  Figure 111-4. AGGLOMERATION FLOWSHEET
CONCENTRATE FILTER CAKE
                                BENTONITE
                 BALLING DRUM
                     ±
                    SCREEN
                 1         I
             UNDERSIZE  OVERSIZE
                                 1 FUEL I
                 AGGLOMERATION FURNACE
             AND/OR SHIPPING
                 PELLETS        EXHAUST GASES


                   t                t
              TO STOCK PILE      TO ATMOSPHERE
                 40

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archeologists.  Copper was first produced in the colonies at
Simsbury,  Connecticut, in 1709.  In 1820, a copper ore body
was found in Orange County, Vermont.  In the  early  1840ts,
ore  deposits  located  in  Northern  Michigan accounted for
extensive copper production in  the  United  States.   Other
discoveries  followed in Montana (1860), Arizona (1880), and
Bingham Canyon, Utah  (1906).  Since 1883, the United  states
has led copper production in the world.  As indicated by the
tabulation  which  follows,  seven  states presently produce
essentially all of the copper mined in the  U. S.  (See  also
Figure III-5.)

              Arizona
              Utah
              New Mexico
              Montana
              Nevada
              Michigan
              Tennessee


A series of tables follow which give statistics for the U.S.
copper  industry.   Table  III-5  lists  total  copper  mine
production of ore by year, and Table III-6 gives copper  ore
production by state for 1972.  The average copper content of
domestic   ores  is  given  by  Table  III-7.   The  average
concentration  of  copper  recovered  from  domestic   ores,
classified  by extraction process, is listed in Table III-8.
Copper concentrate production by froth flotation is given in
Table III-9, while production of copper concentrate by major
producers in 1972 is given as part of Supplement B.

Twenty-five mines account for 95% of the U.S. copper output,
with  more  than  50%  of  this  output  produced  by  three
companies  at  five  mines.   Approximately  90%  of present
reserves (77.5 million metric tons, 85.5 million short tons,
of copper  metal  as  ore)  average  0.86%  copper  and  are
contained  in  five  states:  Arizona,  Montana,  Utah,  New
Mexico, and Michigan.  Mining produced  154  million  metric
tons  (170 million short tons)  of copper ore and 444 million
metric tons (490 million short tons) of waste in 1968.

Open-pit mines produce 83% of the total copper  output  with
the   remainder   of   U.S.   production   from  underground
operations.  Ten percent of mined  material  is  treated  by
dump  (heap)  and  in-situ leaching producing 229,171 metric
tons (253,000 short tons) of  copper.   Recovery  of  copper
from  leach  solutions  by  iron precipitation accounted for
                            41

-------
                         Figure IM-5.  MAJOR COPPER MINING AND MILLING ZONES OF THE U.S.
•u
to
                                                                     LA. 1 M'SS. '  ALA.  j   GA

                    J"
            '~S
                             ^£v
                               %
     MINING AND MILLING COPPER AS A PRIMARY METAL -



E§:g| MINING AND MILLING COPPER AS A COPRODUCT

-------
  TABLE 111-5. TOTAL COPPER-MINE PRODUCTION OF ORE BY YEAR
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION
1000 METRIC TONS
154,239
202,943
233.760
220,089
242,016
263,088
1000 SHORT TONS
170,054
223,752
257,729
242,656
266,831
290,000
        SOURCE: REFERENCE 2
TABLE III-6. COPPER-ORE PRODUCTION FROM MINES BY STATE [1972]

STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
TENNESSEE
ALL OTHER
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
150,394
32,250
18,077+
15,531+
12.052*
7,483
1,598
< 4,631
242,016
1000 SHORT TONS
165,815
35.557
19,930+
17,126*
13,288+
8,250
1,762
< 5,106
266,831
        SOURCE: REFERENCE 2
                      43

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     TABLE 111-7. AVERAGE COPPER CONTENT OF DOMESTIC ORE
YEAR
1968
1969
1970
1971
1972
1973
PERCENT COPPER
0.60
0.60
0.59
0.55
0.55
0.53
                 SOURCE: REFERENCE 2
TABLE 1118. AVERAGE CONCENTRATION OF COPPER IN DOMESTIC ORES
          BY PROCESS (1972)
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
IDAHO
TENNESSEE**
COLORADO
ALL OTHER
TOTAL U.S.
CONCENTRATION (%)
FLOTATION*
0.51
0.58
0.70
0.55
0.54
0.82
—
0.64
—
1.35
0.55
DUMP/HEAP
LEACH
0.47
1.10
-
-
0.38
N/A
_
N/A
—
—
0.47
DIRECT SMELTER
FEED
1.94
—
0.07f
4.06
0.68
—
2.65
—
10.24
2.30
1.68
     • INCLUDES FROTH FLOTATION AND LEACH-REDUCTION/FLOTATION
    *• FROM COPPER/ZINC ORE
     t JUST AS A FLUXING MATERIAL

      SOURCE: REFERENCE 2
                          44

-------
TABLE 111-9. COPPER ORE CONCENTRATED IN THE UNITED STATES BY
          FROTH FLOTATION, INCLUDING LPF PROCESS (1972)
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
TENNESSEE*
ALL OTHER
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
138,998
31,702
18,019
15,508
12,003
7,483
1,598
228
225,537
1000 SHORT TONS
153,250
34,952
19,866
17,098
13,234
8,250
1,762
251
248,663
         FROM COPPER/ZINC ORE
         SOURCE:  REFERENCE 2
                         45

-------
87.5% of the leaching  production;  recovery  of  copper  by
electromining amounted to 12.5%.

Approximately   98%   of   the   copper   ore  was  sent  to
concentrators  for  beneficiation  by  froth  flotation,   a
process  at  least  60 years old.  Copper concentrate ranges
from 11% to 38% copper as  a  result  of  approximately  83%
average recovery from ore.

Secondary  or coproduction of other associated metals occurs
with copper mining and processing.  For instance,  in  1971,
41%  of  U.S.  gold production was as base-metal byproducts.
Fourteen copper plants in 1971 produced molybdenum as  well.
From  63.5  million  metric  tons (70 million short tons) of
molybdenum byproduct ore, 18,824 metric tons  (20,750  short
tons) of byproduct molybdenum were produced.

Processes  Employed to Extract copper from Ore.    The mining
methods employed by the  copper  industry  are  open-pit  or
underground  operations.  Open-pit mining produces step-like
benched tiers of mined areas.  Underground  mining  practice
is usually by block-caving methods.

Processing  of  copper  ores  may  be  hydrometallurgical or
physical-chemical separation from the  gangue  material.   A
general  scheme  of  methods employed for recovery of copper
from ores is given as Figure III-6.   Hydrometallurgical pro-
cesses  currently  employ  sulfuric  acid   (5-10%)  or  iron
sulfate  to  dissolve  copper  from the oxide or mixed oxide
sulfide ores in dumps, heaps, vats or insitu (Table III-10).
Major  copper  areas  employing  heap,  dump,  and   in-situ
leaching  are  shown  in  Figure  III 7.  The copper is then
recovered from solution in a highly pure form  by  the  iron
precipitation,  electrolytic deposition (electrowinning), or
solvent extraction-electrowinning process.

Ore may also be concentrated by froth flotation,  a  process
designed for extraction of copper from sulfide ores.  Ore is
crushed  and  ground  to  a  suitable  mesh size and is sent
through flotation  cells.   Copper  sulfide  concentrate  is
lifted in the froth from the crushed material and collected,
thickened,  and filtered.  The final concentrate, containing
15-30% copper, is sent to  the  smelter  for  production  of
blister  copper (98% Cu).  The refinery produces pure copper
(99.88-99.9% Cu) from  the  blister  copper,  which  retains
impurities  such  as  gold, silver,  antimony, lead, arsenic,
molybdenum,  selenium,  tellurium,  and  iron.   These   are
removed in the refinery.

-------
Figure 111-6. GENERAL OUTLINE OF METHODS FOR TYPICAL RECOVERY OF
         COPPER FROM ORE
ORE «0.1%Cu)
ORE <0 10.4* Cut «« 	 _ nwcDRimriFN nwn
ORE """••"• WASTE DUMPS



y$£ «EAP .M.BTU CRUSHER " SCREENING
LEACH LEACH LEACH ' 	 	 ±
. _ OXinF/SHI FIDE SFmiunARV
i , i i ORE CRUSHER
ACID ACID ACID X
ACID ACID ACID croetMHur
RECYCLED RECYCLED HfCYClED SCREENING
r ' ' ' ' i
PRECIPITATION HEAP TERTIARY
pLANIi |»- CRUSHER MIXED C
, ACID + ""-rlui
ACID HICYCLED ^>ui.«i«i™ | 	 r cycLO(gES
' SPONGE IRON Y
CEMENT lnrXIDE/
ORE
i



WASH
WATER
	 1 VAT LEACH
ITS (ACID)
TAILS
BARREN
P
S
ELECTRO
WINNING
FACILITY
OW
IDE
JMP
<

CATHODE
COPPER
'
TO MA
(OR RE
REGNANT
OLUTION



RKET
FINEHYI
                                        HE FINE HV
                                  MARKET
                         47

-------
TABLE 111-10. COPPER ORE HEAP OR VAT LEACHED IN THE UNITED STATES (1972)

STATE
ARIZONA
UTAH
NEW MEXICO/NEVADA
MONTANA
TOTAL U.S.
PRODUCTION
1000 METRIC TONS
11.071
549
4.400
N/A
16,039
1000 SHORT TONS
12.228
605
4,851
N/A
17,684
          SOURCE: REFERENCE 2
                            48

-------
Figure 111-7.  MAJOR COPPER AREAS EMPLOYING ACID LEACHING IN HEAPS,
            IN DUMPS, OR IN SITU
                            :j:j:j::3 LEACHING ZONES

-------
One  combination  of  the  hydrometallurgical  and physical-
chemical   processes,   termed   LPF   (leach-precipitation-
flotation)   has enabled the copper industry to process oxide
and sulfide minerals efficiently.  Also, tailings  from  the
vat  leaching  process,  if they contain significant sulfide
copper, can be sent to the flotation circuit to float copper
sulfide,  while  the  vat  leach  solution  undergoes   iron
precipitation  or electrowinning to recover copper dissolved
from oxide ores by acid.

A major factor affecting domestic copper production  is  the
market  price  of the material.  Historically, copper prices
have fluctuated but have generally increased over  the  long
term  (Table  III-ll).   Smelter  production  of copper from
domestic ores has continuously risen and  has  increased  in
excess  of  a  factor of three over the last 68 years (Table
111-12).

Lead and Zinc Ores

Lead and zinc mines and mills in the U.S. range in age  from
over  one  hundred  years  to  essentially new.  The size of
these operations ranges from several hundred metric tons  of
ore  per  day  to  complexes  capable  of  moving  about six
thousand metric tons of ore per day.  Lead and zinc ores are
produced almost exclusively from underground  mines.   There
are some deposits which are amenable to open-pit operations;
a  number  of  mines  during  their  early opening stages of
operation are started as open-pit mines and  then  developed
into underground mines.  At present, only one small open-pit
mine  is  in  operation, and its useful life is estimated in
months.  Therefore, for all practical purposes,  all  mining
can be considered to be underground.

In general, the ores are not rich enough in lead and zinc to
be  smelted  directly.   Normally,  the  first  step  in the
conversion of ore into metal is  the  milling  process.    In
some  cases,  preliminary  gravity  separation  is practiced
prior to the actual recovery of the  minerals  of  value  by
froth flotation, but, in most cases, only froth flotation is
utilized.   The  general procedure is to initially crush the
ore and then grind it, in a closed circuit with  classifying
equipment,  to  a  size  at which the ore minerals are freed
from the gangue.  Chemical reagents are then added which, in
the presence of bubbled air, produce selective flotation and
separation of the desired minerals.  The  flotation  milling
process  can  be  rather complex depending upon the ore, its
state of oxidation, the  mineral,  parent  rock,  etc.   The
                            50

-------
           TABLE 111-11. AVERAGE PRICE RECEIVED FROM COPPER
                        IN THE UNITED STATES
   YEAR
                            PRICE IN CENTS PER KILOGRAM (CENTS PER POUND)
   LAKE COPPER*
              ELECTROLYTIC COPPERt
1865-1874
   1907
   1910
   1915
   1917
   1920
   1925
   1930
   1932
   1935
   1940
   1945
   1950
   1955
   1960
   1965
   1970
   1972
   1973
     60.94 (27.70)
     46.86 (21.30)
     28.86(13.12)
     38.81 (17.64)
     64.20 (29.18)
     39.62(18.01)
     31.77 (14.44)
     29.48 (13.40)
     13.00 (  5.91)
     19.62 (  8.92)
     25.65(11.66)
     26.40(12.00)
 40.96- 54.16(18.62
 66.00- 94.60(30.00
 66.00- 72.60(30.00
 74.80- 83.60(34.00
116.6 -132.0(53.00-
109.7 -114.7(49.88-
110.3 -159.2(50.13-
- 24.62)
•43.00)
- 33.00)
• 38.00)
60.00)
52.13)
72.38)
 42.90- 53.90(19.50-24.50)
 69.30 - 94.60 (31.50 - 43.00)
 66.00 -      (30.00)
 77.00- 81.40(35.00-37.00)
116.9  -132.3(53.12-60.12)
111.4  -115.8(50.63-52.63)
116.9  -151.1 (53.13-68.70)
  * COPPER FROM NATIVE COPPER MINES OF LAKE SUPERIOR DISTRICT: MINIMUM 99.90%
    PURITY. INCLUDING SILVER.

  t ELECTROLYTIC COPPER RESULTS FROM ELECTROLYTIC REFINING PROCESSES:
    MINIMUM 99.90% PURITY, SILVER COUNTED AS COPPER

  SOURCE:  REFERENCE 3
                                 51

-------
 TABLE 111-12. PRODUCTION OF COPPER FROM DOMESTIC ORE
            BY SMELTERS
YEAR
1905
1910
1915
1916
1919
1921
1925
1929
1930
1932
1935
1937
1940
1943
1946
1950
1955
1960
1965
1970
1971
1972
1973
I
ANNUAL PRODUCTION
METRIC TONS
403,064
489,853
629,463
874,280
583,391
229,283
759,554
908,299
632,356
246,709
345,834
757,038
824.539
991,296
543,888
826.596
913,631
1,036,563
1.272,345
1,455.973
1,334,029
1,513.710
1,569.110*
SHORT TONS
444,392
540,080
694,005
963,925
643,210
252.793
837,435
1,001,432
697,195
272,005
381,294
834,661
909,084
1,092,939
599,656
911,352
1.007,311
1,142,848
1,402,806
1,605,262
1.470,815
1,668,920
1.730.000»
•PRELIMINARY BUREAU OF MINES DATA
SOURCE: REFERENCE 3
                    52

-------
recovered  minerals  are shipped in the form of concentrates
for reduction to the respective metals recovered.

The most common lead mineral mined in  the  U.S.  is  galena
(lead sulfide).  This mineral is often associated with zinc,
silver, gold, and iron minerals.

The principal zinc ore mineral is zinc sulfide  (sphalerite).
There  are,  however,  numerous other minerals which contain
zinc.   The  more  common  include  zincite  (zinc   oxide),
willemite  (zinc  silicate), and franklinite (an iron, zinc,
manganese oxide complex).   Sphalerite  is  often  found  in
association  with sulfides of iron and lead.  Other elements
often found in association with sphalerite  include  copper,
gold, silver, and cadmium.

Mine  production  of lead increased during 1973 and 1974, as
illustrated in Table 111-13, which has  been  modified  from
the   Mineral  Industry  surveys,  U.S.  Department  of  the
Interior,  Bureau  of   Mines,   Mineral   Supply   Bulletin
(Reference 4).

Missouri  was  the  foremost  state with 80.78% of the total
United states production, followed  by  Idaho  with  10.24%,
Colorado  with 4.66%, Utah with 2.28%, and other states with
the remaining 2.04%.  This same  trend  continues  with  the
preliminary  figures  for  1974  for  the  period of January
through June.  Based on this information and  the  estimated
60-year  life  for  the lead ores in the "Viburnum Trend" of
the "New Lead Belt" of southeast Missouri, it is likely that
this area will be the predominant lead source for many years
to come.

Mine  production  of  zinc  during  1973   and   preliminary
production figures for December and January 1974 and January
through  May  1974  are presented in Table III-14, which has
been  modified  from  the  Mineral  Industry  Surveys,  U. S.
Department  of  Interior,  Bureau  of  Mines, Mineral Supply
Bulletins.

The mine production figures by state for zinc in 1973,  how-
ever, are misleading, because Tennessee was ranked third due
to  prolonged  strikes,  the replacement of some older mine-
mills,  and  the  development  and   construction   of   new
production  facilities.   Therefore, note that Tennessee led
the nation in the production  of  zinc  for  15  consecutive
years   (until 1973) and should regain the number one ranking
back from Missouri  (1973), based on the preliminary  produc-
tion figures given for the first half of 1973.
                             53

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        TABLE 111-13. MINE PRODUCTION OF RECOVERABLE LEAD
                    IN THE UNITED STATES


STATE
Alaska
Arizona
California
Colorado
Idaho
Illinois
Maine
Missouri
Montana
New Mexico
New York
Utah
Virginia
Washington
Wisconsin
Other States


1973

RANK



3
2


1



4




%



4.66
12.24


80.78



2.28




Total
Daily average*
1973
JAN .-DEC.
METRIC TONS
5
692
40
25.497
56.002
491
185
441,839
160
2.318
2,090
12,456
2,392
2.011
765
—
546,943
1.498
SHORT TONS
6
763
44
28.112
61.744
541
204
487.143
176
2,556
2,304
13,733
2,637
2.217
844
—
603.024
1.652
1974 (PRELIMINARY)
JAN.-JUNE
METRIC TONS
._
357
11
11.317
25,667
122
98
251.571
51
1,078
1,331
5,674
1.359
443
596
486
300,163
1,658
SHORT TONS
...
394
12
12,478
28,299
135
108
277,366
56
1,189
1,467
6.256
1,499
489
657
536
330,941
1.828
'Based on number of days in month without adjustment for Sundays or holidays.
                                  54

-------
         TABLE 111-14. MINE PRODUCTION OF RECOVERABLE ZINC
                    IN THE UNITED STATES (PRELIMINARY)


STATE
Arizona
California
Colorado
Idaho
Illinois
Kentucky
Maine
Missouri
Montana
New Jersay
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Washington
Wisconsin


1973

RANK


4
5


7
1

6

2

3
9
8


Total
%


11.94
9.55


4.13
17.27

6.94

17.4

13.32
3.48
3.51



Daily average*
1973
JAN.-DEC. TOTALS
METRIC TONS
7,638
16
51,533
41,216
4.823
245
17,843
74,576
379
29.955
11,147
73,861
17,104
57,474
15.023
15.131
5.768
7.866
431,599
1,183
SHORT TONS
8,421
18
56317
45.442
5,318
270
19,672
82,223
418
33,027
12,290
81,435
18,858
63,367
16,564
16,682
6,359
8,672
475,853
1,304
1974
JAN. TOTALS
METRIC TONS
600
_.
3,961
3,279
224
—
1,238
6,589
82
2.361
863
6.961
1,575
7,239
1,130
1,281
528
733
38,644
1746
SHORT TONS
662
_.
4,367
3,615
247
.„
1,365
7,266
90
2,603
951
7,675
1,737
7,981
1.246
1.412
582
808
42,606
1,374
'Based on number of days in month without adjustment for Sundays or holidays.
                                    55

-------
Description  of Lead/Zinc Mining and Milling Processes.  The
recovery of useful lead/zinc minerals involves  the  removal
of  ores  containing  these minerals from the earth  (mining)
and the subsequent separation of the useful mineral from the
gangue material (concentration).  A generalized  flow  sheet
for such a mine/mill operation is presented in Figure III-8.

Mine  Operations.   The mining of lead- and zinc-bearing ores
is  generally  accomplished  in  underground   mines.    The
mineralcontaining  formation  is usually fractured utilizing
explosives such as  ammonium  nitrate-fuel  oil  (AN-FO)  at
slurry  gels,  placed  in  holes  drilled  in the formation*
After blasting, the rock fragments are  transported  to  the
mine  shaft  where  they  are  lifted up the shaft in skips*
Primary  or  rough  crushing  equipment  is  often  operated
underground.   The drilling and transportation equipment is/
of course, highly  mechanized  and  usually  employs  diesel
power.   At  some  locations, the equipment is maintained in
underground shops, constructed in  mined-out  areas  of  the
workings.

Water enters a mine naturally when aquifers are intercepted;
in  highly fractured and fissured formations, water from the
surface may seep into the mine.  Minor amounts of water  are
introduced  from the surface by evaporation of cooling water
and through water expired by workers.   At  some  locations/
water  enters  with  sand  or  tailings  used  in  hydraulic
backfill operations.

The water is pumped from the mine at  a  rate  necessary  to
maintain operations in the mine.  The amount of water pumped
does  not  bear  any necessary relationship to the output of
ore or mineral.  The amount pumped may vary  from  thousands
of  liters  per  day  to 120 to 160 million liters (30 to «0
million gallons) per day.  In many cases, there  is  a  sub'
stantial  seasonal  variation  in  the amount of water which
which must be pumped.

The water pumped from a mine  may  contain  fuel,  oil,  and
hydraulic   fluid  from  spills  and  leaks,  and,  perhaps/
blasting agents and partially oxidized blasting agents.  The
water, most certainly, will  contain  dissolved  solids  and
suspended  solids  generated  by the mining operations.  The
dissolved and suspended solids may consist  of  lead,  zinc/
and associated minerals.

Milling  Operations.   The  valuable  lead/zinc minerals are
recovered from the  ore  brought  from  the  mine  by  froth
flotation.   In some cases, the ore is preconcentrated using
                             56

-------
     Figure 111-8. LEAD/ZINC-ORE MINING AND PROCESSING OPERATIONS
            ORE MINING |	DRAINAGE —
                                         WATER
                                         DISSOLVED SOLIDS
                                         SUSPENDED SOLIDS •
                                         FUELS
                                         LUBRICANTS
                              TO POND
                              AND/OR
                              MILL
                                                WATER FROM MINE.
                                                RECYCLE OR OTHER
                                                REAGENTS
 GRINDING AND
CLASSIFICATION
                               LEAD ROUGHER
                              LEAD FLOTATION
 CONCENTRATE
                               ZINC ROUGHER
  FINAL LEAD
 CONCENTRATE
                                        ZINC ROUGHER
                                        CONCENTRATE
  THICKENING
AND FILTRATION
                                        ZINC CLEANER
         USUALLY
         RECYCLED
        TO PROCESS
       WATER SYSTEM
         CONCENTRATE
 CONCENTRATE
                                       AND FILTRATION
                           WATER
                           DISSOLVED SOLIDS
                           SUSPENDED SOLIDS
                           EXCESS REAGENTS
                    TO SUBSURFACE
                      DRAINAGE
          CONCENTRATE
            TO ZINC
            SMELTER
     "1
       I
USUALLY RECYCLED
   TO PROCESS
  WATER SYSTEM
                                      57

-------
mechanical devices based  on  specific  gravity  principles.
The  ore  is  initially  crushed  to  a  size  suitable  for
introduction into fine grinding equipment, such as rod mills
and ball mills.  These mills run wet and are usually run  in
circuit  with  rake  or cyclone classifers to recycle to the
mill material which is coarser than the  level  required  to
liberate  the  mineral  particles.  The fineness of grind is
dependent on the degree of dissemination of the  mineral  in
the  host  rock.  The ore is ground to a size which provides
an economic balance  between  the  additional  metal  values
recovered versus the cost of grinding.

In  some  cases,  the reagents used in the flotation process
are added in the mill; in other  cases,  the  fine  material
from  the  mill  flows to a conditioner (mixing tank), where
the reagents are added.  The  particular  reagents  utilized
are  a function of the mineral concentrates to be recovered.
The specific choice of reagents at a facility is usually the
result of determining empirically which reagents  result  in
an  economic  optimum  of  recovered  mineral  values  which
reagents result in an economic optimum of recovered  mineral
values  versus  reagent costs.  In general, lead and zinc as
well as copper sulfide flotations are  run  at  elevated  pH
(8.5   to   11,   generally)   levels  so  that  frequent  pS
adjustments with hydrated lime (CaOH2)  are  common.   other
reagents commonly used and their purposes are:

         Reagent                       Purpose

Methyl Isobutyl-carbinol               Frother
Propylene Glycol Methyl Ether          Frother
Long-Chain Aliphatic Alcohols          Frother
Pine Oil                               Frother
Potassium Amyl Xanthate                Collector
Sodium Isopropol Xanthate              collector
Sodium Ethyl Xanthate                  Collector
Dixanthogen                            Collector
Isopropyl Ethyl Thionocarbonate        Collectors
Sodium Diethyl-dithiophosphate         Collectors
Zinc sulfate                           Zinc Depressant
Sodium Cyanide                         Zinc Depressant
Copper Sulfate                         Zinc Activant
Sodium Dichromate                      Lead Depressant
Sulfur Dioxide                         Lead Depressant
Starch                                 Lead Depressant
Lime                                   pH Adjustment

The  finely ground ore slurry is introduced into a series of
flotation cells, where the slurry is  agitated  and  air  is
                            58

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introduced.   The  minerals  which  are to be recovered have
been rendered hydrophobic (non-water-accepting)   by  surface
coating  with  appropriate reagents.  Usually, several cells
are operated in a  countercurrent  flow  pattern,  with  the
final  concentrate being floated off the last cell (cleaner)
and the tails taken over the first  or  rougher  cells.   In
some  cases,  regrinding  is  used  on the underflow for the
cleaner cells to improve recovery.

In many cases, more than one mineral is recovered.  In  such
cases,  differential flotation is practiced.  The flow shown
in Figure III-8 is typical of such a differential  flotation
process  for  recovery of lead and zinc sulfides.  Chemicals
which induce hydrophilic  (affinity-for-water)  behavior  by
surface interaction are added to prevent one of the minerals
from  floating  in the initial separation.  The underflow of
tailings  from  this  separation  is  then  treated  with  a
chemical  which  overcomes  the depressing effect and allows
the flotation of the other mineral.

After the recovery of the desirable minerals, a large volume
of tailings or gangue material remains as the underflow from
the last rougher cell in the flow scheme.  These  tails  are
typically adjusted to a slurry suitable for hydraulic trans-
port  to  the treatment facility, termed a tailing pond.  In
some cases,  the  coarse  tailings  are  separated  using  a
cyclone separator and pumped to the mine for backfilling.

The   floated   concentrates   are   dewatered   (usually  by
thickening and filtration), and the final concentrate—which
contains some residual water—is  eventually  shipped  to  a
smelter  for  metal  recovery.  The liquid overflow from the
concentrate thickeners is typically recycled in the mill.

The tailings from a lead/zinc flotation  mill  contains  the
residual solids from the original ore which have been finely
ground   to  allow  mineral  recovery.   The  tailings  also
contains dissolved solids  and  excess  mill  reagents.   In
cases  where  the  mineral content of the ore varies, excess
reagents will undoubtedly be  present  when  the  ore  grade
drops suddenly, and lead and zinc will escape with the tails
if  high-grade ore creates a reagent-starved system.  Spills
of  the  chemical  used  are  another  source   of   adverse
discharges from a mill.

Gold Ore

The gold ore mining and milling industry is defined for this
document  as  that  segment  of the industry involved in the
                            59

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mining and/or milling of ore for the primary  or  byproduct/
coproduct  recovery  of  gold.   in  the United States, this
industry is concentrated in eight states:  Alaska,  Montana,
New  Mexico,  Arizona,  Utah,  Colorado,  Nevada,  and South
Dakota.  Reported domestic production of gold for  1972  was
45.1 million grams  (1.45 million troy ounces) (Reference 2).
Of  this,  approximately  76%  was accounted for by the four
leading  producers,  one  of  which  recovered  gold  as   a
byproduct of a copper mining operation.  Approximately 1% of
the  United  states production was attributed to recovery at
placer mining operations,  most  of  which  are  located  in
Alaska.  The remainder of this production occurred primarily
as byproduct recovery at operations engaged in the mining of
copper, lead, and/or zinc ores.

The domestic production of gold has been on a downward trend
for  the  last 20 years, largely due to the reduction in the
average grade of ore being mined, depletion of ore  at  some
mines, and a labor strike at the major producer during 1972.
However,  large  increases  in the free-market price of gold
during recent years (from approximately $70 per  troy  ounce
in  1972 to nearly $200 per troy ounce in 1974)  stimulated a
widespread increase in prospecting and exploration activity.
As a result, the recovery of gold  from  low-grade  ore  has
become  economically  feasible  for  the  present.  One such
instance  is  an  operation,  located  in   the   State   of
Washington,   which  recovers  gold  from  gold  ore.   This
operation, one of the top ten producers during recent years,
had become ecnomically submarginal and was due to  close  at
the  end  of  1972.   However, the higher gold prices allowed
sustained production from low-grade ore  at  this  mine/mill
through  1975.    The  higher gold prices have also caused an
apparent revival of  gold  placer  mining  in  Alaska.   One
source  (Reference 5)  reports that, although only 40 placers
were active during 1971, 71 placers were active during 1973.
Results of a more recent survey reveal  that,  during  1975,
196 placer operations were active in 34 Alaskan quadrangles.

Mining Practices.  Gold is mined from two types of deposits:
placers  and  lode or vein deposits.  Placer mining consists
of excavating waterborne or glacial deposits of gold-bearing
gravel and sands which can be separated by  physical  means.
Since  most  such  placer deposits are deeply buried, modern
earthmoving equipment is necessary for profitable operation.
In  most  instances,  bulldozers,  front-end  loaders,   and
draglines  are  being  employed  for  overburden  stripping,
sluicebox loading, and tailing-removal operations.   However,
where  water  availability  and   physical   characteristics
permit,  dredging  or hydraulic methods are often favored on
                            60

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an economic  basis.   Lode  deposits  are  mined  by  either
underground  or  open-pit  methods,  the  particular  method
chosen depending on such factors as size and  shape  of  the
deposit,  ore grade, physical and mineralogical character of
the ore and surrounding rock, and depth of the deposit.

Milling Practices.  Milling practices for the processing and
recovery of gold and gold-containing ores  are  cyanidation,
amalgamation,  flotation,  and  gravity  concentration.  All
these processes have been employed in the  beneficiation  of
ore  mined  from lode deposits.  Placer operations, however,
employ only gravity methods which in the past were sometimes
used in conjunction with amalgamation.

Prior to 1970, amalgamation was the process used to  recover
nearly  1/H  of  the gold produced domestically.  Since that
time, environmental concerns have caused restricted  use  of
mercury.   As  a  result, the percent of gold produced which
was recovered by the amalgamation process dropped from 20.3%
in 1970 to 0.3E in 1972.  At the same time that the  use  of
amalgamation   was   decreasing,   the  use  of  cyanidation
processes was  increasing.   In  1970,  36.7%  of  the  gold
produced domestically was recovered by cyanidation, and this
increased to 54.6% in 1972.

The  amalgamation  process as currently practiced  (used by a
single mill in Colorado) involves crushing and  grinding  of
the  lode  ore, gravity separation of the gold-bearing black
sands by jigging, and final concentration  of  the  gold  by
batch amalgamation of the sands in a barrel amalgamator.  In
the  past,  amalgamation  of  lode ore has been performed in
either the grinding mill, on plates, or in  special  amalga-
mators.  Placer gold/silver-bearing gravels are beneficiated
by  gravity  methods,  and, in the past, the precious metal-
bearing sands generally were  batch  amalgamated  in  barrel
amalgamators.   However,  amalgamation in specially designed
sluice boxes was also practiced.

There are basically four methods  of  cyanidation  currently
being  used  in  the  United  States:   heap  leaching,  vat
leaching, agitation leaching,  and  the  recently  developed
carbon-in-pulp  process.   Heap  leaching  is a process used
primarily for the recovery  of  gold  from  low-grade  ores.
This  is  an  inexpensive process and, as a result, has also
been used recently to  recover  gold  from  old  mine  waste
dumps.  Higher grade ores are often crushed, ground, and vat
leached or agitated/leached to recover the gold.
                            61

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In vat leaching, a vat is filled with the ground ore (sands)
slurry,  water  is  allowed  to drain off, and the sands are
leached from the top with  cyanide,  which  solubilizes  the
gold (Figure III-9).  Pregnant cyanide solution is collected
from  the  bottom of the vat and sent to a holding tank.  In
agitation leaching, the  cyanide  solution  is  added  to  a
ground  ore  pulp in thickeners, and the mixture is agitated
until solution of the gold is achieved (Figure III-10).  The
cyanide  solution  is  collected  by  decanting   from   the
thickeners.

Cyanidation  of  slimes  generated  in  the  course  of  wet
grinding is currently being done  by  a  recently  developed
process,  carbon-in-pulp  (Figure  III-9).   The  slimes are
mixed with a  cyanide  solution  in  large  tanks,  and  the
solubilized  gold  cyanide  is  collected by adsorption onto
activated charcoal.  Gold  is  stripped  from  the  charcoal
using  a  small  volume  of  hot  caustic; an electrowinning
process is used for final recovery of the gold in the  mill.
Bullion is subsequently produced at a refinery.

Gold  in  the  pregnant cyanide solutions from heap, vat, or
agitate leaching processes  is  recovered  by  precipitation
with  zinc  dust.   The precipitate is collected in a filter
press and sent to a smelter for the production of bullion.

Recovery of gold by flotation processes is limited, and less
than 3% of the gold produced in 1972 was recovered  in  this
manner.   This  method  employs a froth flotation process to
float and collect the gold-containing minerals (Figure  III-
11).   The  single  operation  currently  using  this method
further processes the tailings from the flotation circuit by
the agitation/cyanidation method  to  recover  the  residual
gold values.

Gold  has historically been recovered from placer gravels by
purely physical means.  Present  practice  involves  gravity
separation,  which  is normally accomplished in a sluicebox.
Typically, a sluicebox consists of an open box  to  which  a
simple  rectangular  sluiceplate  is  mounted  on a downward
incline.  To effect the separation of gold  from  gravel  or
sand,  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 plates  which  are  angled
away  from  the  direction  of  flow in a manner designed to
create pockets and eddy  currents  for  the  collection  and
retention of gold.
                            62

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Figure 111-9. CYANIDATION OF GOLD ORE: VAT LEACHING OF SANDS
         AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
                     ORE
                    CRUSHING



\
f
SAND
FRACTION
'

VAT
FILLING


GRINDING
1
CYCLONE
CLASSIFICATION





SLIME
FRACTION


t- 	 DRAIN WATER » TO WASTE



BARREN
SOLUTION
RECVCLED



ALKALINE
WASH
\
CVAI
LEA
•
HIDE
CH


	 ' RESIDUE SANDS "~l""
-WATER— *• SLUICED OUT OF '
VAT
PRECIPITATION
OF GOLD FROM
LEACHATE BY
ADDITION OF
ZINC DUST
* — \

I REACTIVATION
SAMOS AND RECYCLING ^
4 OF STRIPPED ^^
CARBON
TO BACKFILL
COLLECTION OF
PRECIPITATE
IN FILTER
PRESS
i

PRECII-ITATE
FILTERED AND
THICKENED
RECYCLED *

'1

* L!
CONDITIONING
.
- I
•" L

EAOENTS (C*IOHI2) 1

REAGENT* (CNI I
DISSOLUTION 1
AGITATORS f
ADSORPTION «, 1- 	
AGITATORS 1*
1 	 TAILS—
DESORPTION OF 1^
LOADED CARBON |"

\

ELECTROLYSIS OF
PREGNANT SOLUTION
* 	 ;

GOLD SPONGE
— j AIR |
-*- TO WATTE
^HOTNiOH
*NtCN
SOLUTION

TO REFINERY
    TO SMELTER
                       63

-------
Figure 111-10. CYAN ID ATI ON OF GOLD ORE:  AGITATION/LEACH PROCESS
      ORE
   CRUSHING

r
WATER
   GRINDING
      I
 CONDITIONING
      I
COUNTERCURRENT
  LEACHING IN
  THICKENERS
 PRECIPITATION
 OF GOLD FROM
 LEACHATE BY
 ADDITION OF
 ZINC DUST
      *
COLLECTION OF
PRECIPITATE IN
 FILTER PRESS
     I
  PRECIPITATE
 FILTERED AND
  THICKENED
      t
 TO SMELTER
                        REAGENTS (CN)
                    BARREN
                      PULP
TAILING
 POND
                                                TA'"ND
                           BARREN SOLUTION
                              RECYCLED

-------
Figure 111-11. FLOTATION OF GOLD-CONTAINING MINERALS WITH RECOVERY OF
          RESIDUAL GOLD VALUES BY CYANIDATION

       ORE
     CRUSHING
                          WATER
     GRINDING
                        REAGENTS
   CONDITIONING
       i
    SELECTIVE
      FROTH
    FLOTATION
   CONCENTRATE
     FILTERED
  AND THICKENED
                         FLOTATION CIRCUIT
                             TAILINGS
   TO SMELTER
   REAGENTS (CN)
                           LEACHING IN
                            THICKNERS
.BARREN.
  PULP
.TO TAILING
 POND
                       PRECIPITATION OF GOLD
                         FROM LEACHATE BY
                       ADDITION OF ZINC DUST
           BARREN
          SOLUTION
          RECYCLED
                          COLLECTION OF
                          PRECIPITATE IN
                           FILTER PRESS
                        PRECIPITATE FILTERED
                          AND THICKENED
                            TO SMELTER
                            65

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During  actual sluicing operations, pay gravels (i.e., gold-
bearing 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  which  collects
there.

Other  types  of equipment which may be employed in physical
separation operations include  jigs,  tables,  and  screens.
This   equipment   is   typically  found  only  at  dredging
operations, however.

Cleanup of gold recovered by  gravity  methods  is  normally
accomplished  with small (102 cm (40 in.))  sluices, screens,
and, finally, by hand-picking impurities from the gold.

Silver Ores

The silver ore mining and milling industry  is  defined  for
this  document  as  that segment of industry involved in the
mining and/or milling of ore for the primary  or  byproduct/
coproduct recovery of silver.  Domestic production of silver
for  1972  was  1.158  million  kilograms  (37,232,922  troy
ounces).  Over 38% of this production came from  Idaho,  and
most  of  this,  from the rich Coeur d'Alene district in the
Idaho panhandle.   The remaining production was  attributable
to  eleven  states:   Alaska, Arizona, California, Colorado,
Michigan,  Missouri,  Montana,  Nevada,  New  Mexico,  South
Dakota,  and Utah.  The 25 leading producers contributed 85%
of this total  production,  and  nine  of  these  operations
produced over one million troy ounces each.  During the past
ten  years,  the annual production of silver has varied from
approximately 1 to 1.4 million kilograms (32 to  45  million
troy  ounces).    Prices  have  also varied and, during 1972,
ranged from a low of 4.41 cents per gram  (137.2  cents  per
troy  ounce)   to  a high of 6.54 cents per gram (203.3 cents
per troy ounce).   Average price for 1972 was 5.39 cents  per
gram (167.7 cents per troy ounce).

Current  domestic production of new silver is derived almost
entirely from exploitation of low-grade and complex  sulfide
ores.   About  one-fourth of this production is derived from
ores wherein silver is  the  chief  value  and  lead,  zinc,
and/or  copper are valuable byproducts.  About three-fourths
                            66

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of this production is from ores in  which  lead,  zinc,  and
copper  constitute  the  principal  values,  and silver is a
minor but important byproduct.  The types, grade, and  rela-
tive  importance  of  the  metal  sulfide  ores  from  which
domestic silver is produced are listed in Table 111-15.

Present extractive metallurgy of silver was developed over a
period of more than 100 years.  Initially,  silver,  as  the
major  product,  was  recovered  from  rich oxidized ores by
relatively crude methods.  As the  ores  became  leaner  and
more   complex,   an   improved  extractive  technology  was
developed.   Today, silver production is predominantly  as  a
byproduct,  and is largely related to the production of lead,
zinc,  and  copper  from  the  processing of sulfide ores by
froth flotation and smelting.  Free-milling—simple,  easily
liberated—gold/silver  ores,  processed by amalgamation and
cyanidation, now contribute only 1 percent of  the  domestic
silver   produced.    Primary  sulfide  ores,  processed  by
flotation and smelting, account for 99 percent  (Table  III-
16).

Selective  froth  flotation  processing  can effectively and
efficiently beneficiate almost any type and grade of sulfide
ore.  This process employs  various  well-developed  reagent
combinations and conditions to enable the selective recovery
of  many different sulfide minerals in separate concentrates
of high quality.  The reagents commonly used in the  process
are   generally   classified   as   collectors,   promoters,
modifiers,  depressants,  activators,  and  frothing  agents.
Essentially, these reagents are used in combination to cause
the  desired  sulfide mineral to float and be collected in a
froth  while  the  undesired  minerals  and   gangue   sink.
Practically  all  the  ores  presently  milled  require fine
grinding to liberate the sulfide minerals from  one  another
and from the gangue minerals.

A  circuit  which  exemplifies the current practice of froth
flotation for the primary recovery of silver from silver and
complex ores is shown in Figure 111-12.  Primary recovery of
silver is largely from  the  mineral  tetrahedrite,  (Cu,Fe,
Zn,Ag) 12sb4S13_.    A   tetrahedrite   concentrate   contains
approximately 25 to 32% copper in addition to the  25.72  to
44.58  kilograms  per metric ton (750 to 1300 troy ounce per
ton) of silver.  A low-grade  (3.43 kg per  metric  ton;  100
troy  oz  per  ton)  silver/pyrite concentrate is produced at
one  mill.    Antimony  may  comprise  up  to  18%   of   the
tetrahedrite  concentrate  and  may  or may not be extracted
prior to shipment to a smelter.
                            67

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   TABLE 111-15. DOMESTIC SILVER PRODUCTION FROM DIFFERENT TYPES OF ORES
TYPE
SILVER
COPPER
LEAD/ZINC/
COPPER
LEAD
ZINC
OTHERS*
SILVER ORE PRODUCTION
1000 METRIC TONS
405.43
187,960.33
35,641.47
7,929.90
1,104.73
1,599.04
1000 SHORT TONS
447
207,233
39,296
8.743
1,218
1,763
GRADE OF SILVER
GRAMS PER
METRIC TON
679.0
2.06
10.29
20.57
3.53
6.86
OUNCES PER
SHORT TON
19.8
0.06
0.3
0.6
0.1
0.2
DOMESTIC
PRODUCTION
24
32
28
14
< 0.5
1.5
DERIVED FROM GOLD AND GOLD/SILVER ORE



SOURCE: REFERENCE 2
                                 68

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    TABLE 111-16. SILVER PRODUCED AT AMALGAMATION AND
               CYANIDATION MILLS IN THE U.S. AND
               PERCENTAGE OF SILVER RECOVERABLE
               FROM ALL SOURCES

YFAR


1968
1969
1970
1971
1972
SILVER BULLION AND PRECIPITATES RECOVERABLE BY

AMALGAMATION
KILOGRAMS
2862.2
2605.7
2963.8
30.9
77.4
TROY OUNCES
92,021
83,775
95,287
993
2.490
CYANIDATION
KILOGRAMS
1669.2
1533.8
774.2
3321.4
3110.1
TROY OUNCES
53,666
49,312
24.892
106,785
99,992
YEAR
1968
1969
1970
1971
1972
SILVER RECOVERABLE FROM ALL SOURCES (%}
AMALGAMATION
0.28
0.20
0.21
t
0.01
CYANIDATION
0.16
0.11
0.05
0.26
0.27
SMELTING*
99.55
99.68
99.73
99.74
99.72
PLACERS
0.01
0.01
0.01
t
t
'Crude ores and concentrates
fLess than 1/2 unit
SOURCE: REFERENCE 2
                         69

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Figure 111-12. RECOVERY OF SILVER SULFIDE ORE BY FROTH FLOTATION
                               ORE
              REAGENTS
     CONCENTRATE
      NO. 1
FLOTATION CIRCUIT
                               NO. 2
                         FLOTATION CIRCUIT
     RETREATMENT
       CIRCUIT
                          PYRITE
                        CONCENTRATE
                               NO. 3
                         FLOTATION CIRCUIT
       FINAL Ag
    CONCENTRATE*
     FINAL
    TAILINGS
•CONTAINS
 25.7 TO 44.6 KILOGRAMS PER
 METRIC TON
 (750-1300 OUNCES PER SHORT TON):
 25 TO 32% COPPER
 0 TO 18% ANTIMONY
                 'CONTAINS 3.43 KILOGRAMS PER
                 METRIC TON (100 TROY OUNCES
                 PER SHORT TON)
                             70

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Various other silver-containing minerals  are  recovered  as
byproducts  of primary copper, lead, and/or zinc operations.
Where this occurs,  the  usual  practice  is  to  ultimately
recover   the   silver   from   the   base-metal   flotation
concentrates at the smelter or refinery.

Bauxite

Bauxite mining for the eventual production of  metallurgical
grade  alumina occurs near Bauxite, Arkansas, where two pro-
ducers mined approximately 1,855,127 metric tons  (2,0«*5,344
tons)   of  ore in 1973.  Both operations are associated with
bauxite  refineries   (SIC  2819),  where  purified   alumina
(A12IO3J   is produced.  Characteristically, only a portion of
the bauxite  mined  is  refined  for  use  in  metallurgical
smelting, and one operation reports only about 10 percent of
its  alumina is smelted, while the remainder is destined for
use as chemical and refractory  grade  alumina.   A  gallium
byproduct  recovery operation occurs in association with one
bauxite mining and refining complex.

The domestic bauxite resource began to be tapped  about  the
turn  of  the century, and one operation has been mining for
about 75 years.  However, the  aluminum  industry  began  to
burgeon  during  World  War  II,   and,  almost overnight the
demands for this lightweight metal for aircraft created  the
large  industry  of  today.  Concurrent with the increase in
demand for aluminum was the startup  of  large-scale  mining
operations by both bauxite producers.

Most   bauxite   is  mined  by  open-pit  methods  utilizing
draglines, shovels, and haulers.    Stripping  ratios  of  as
much  as 10 feet of overburden to 1 foot of ore are minable,
and a 15-to-l ratio is considered  feasible.   Pits  of  100
feet  in  depth are common, and 200 feet is considered to be
the economic limit for large ore  bodies.   The  pits  stand
quite  well  for  unconsolidated  sands  and clays, but some
slumping does occur.

Underground mining occurs at one Arkansas facility, and this
operation provides the low-silica ore essential to the  com-
bination  process of refining.  Although this type of mining
is relatively costly, it is a viable alternative to the pur-
chase of foreign ores at elevated prices.  However,  one  of
the operations utilizes imported bauxite for blending of ore
grades.   Milling  of the bauxite ore involves crushing, ore
blending, and grinding  in  preparation  for  refining.   in
1972,  less  than 10 percent of the bauxite used for primary
aluminum  production  was  of  domestic  origin.   With  the
                            71

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increasing  demand for aluminum, it is expected that the use
of imported alumina and aluminum, as well as  bauxite,  will
increase.   Therefore,  the  domestic  supply  of bauxite is
insufficient to meet present  needs  of  the  nine  domestic
refineries.   Recent  price  increases  in  foreign  bauxite
supplies aid in assuring  the  future  of  domestic  bauxite
operations, regardless of the limited national reserves.

The  search  for potential economic sources of aluminum per-
sists, and many pilot projects have been designed to produce
aluminum.  Currently, the most notable attempt to utilize an
alternative source of aluminum is a 9 metric ton  (10  short
ton)    per   day   pilot   plant   which  converts  alunite,
K2Al£(OH)l_2(SOj*) j*,  to  alumina  through  a  modified  Bayer
process,  preceded  by  roasting  and  water  leaching.  The
process yields byproduct sulfuric acid and potassium sulfate
as cost credits.  Additionally, the  processing  of  alunite
creates no significant "red mud" (leach residue) .  Currently
alunite   mining  is  in  the  exploratory  stages,  with  a
commercial scale refinery slated for construction  in  1975.
Full-scale   mining  will  entail  drilling,  blasting,  and
hauling   using   bench   mining   techniques.    From   all
indications, alunite may provide an economical new source of
aluminum.

Bauxite   production  in  the  United  States  has  declined
recently  from  a  peak  year  in  1970,   and   preliminary
production  figures  for 197H indicate a continuation of the
trend.  Production figures in Table  111-17  indicate  total
U.S.  production  of bauxite, which includes that from mines
in Alabama, Georgia, and Arkansas.  These mines also produce
bauxite for purposes other than metallurgical smelting.

Ferroalloy. Ores

The ferroalloy ore mining and milling category embraces  the
mining and beneficiation of ores of cobalt, chromium, colum-
bium  and tantalum, manganese, molybdenum, nickel, and tung-
sten including crushing, grinding, washing, gravity  concen-
tration, flotation, roasting, and leaching.  The grouping of
these  operations  is based on the use of a portion of their
end product in the production of ferroalloys  (e.g.,  ferro-
manganese,  ferromolybdenum,  etc.) and does not reflect any
special similarities among the ores cr among  the  processes
for  their  recovery  and beneficiation.  SIC 1061, although
presently including  few  operations  and  relatively  small
total  production,  covers a wide spectrum of the mining and
milling industry as a whole.  Sulfide, oxide, silicate, car-
bonate, and anionic ores all are or have been recovered  for
                            72

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TABLE 111-17. PRODUCTION OF BAUXITE IN THE UNITED STATES
YEAR
1964
1965
1966
1967
1968
1969
1970
1971
1972
1973
1000 METRIC TONS*
1626
1680
1825
1680
1692
1872
2115
2020
1930
1908
1000 SHORT TONS»
1793
1852
2012
1852
1865
2064
2332
2227
2128
2104
      'Production, given in dry equivalent weight, includes bauxite mined for
       purposes other than metallurgical smelting
                         73

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the  included  metals.   Open-pit  and underground mines are
currently worked, and placer deposits have been mined in the
past and are included in  present  reserves.   Beneficiation
techniques  include  numerous  gravity  processes,  jigging,
tabling,  sink-float,  Humphreys  spirals;  flotation,  both
basic-sulfide  and fatty-acid; and a variety of ore leaching
techniques.  Operations vary  widely  in  scale,  from  very
small  mines  and  mills  intermittently  worked  with total
annual volume measured in hundreds of tons, to  two  of  the
largest   mining  and  milling  operations  in  the  country
(Reference 2 ).  Geographically, mines  and  mills  in  this
category are widely scattered, being found in the southeast,
southwest,  northwest,  north  central,  and  Rocky Mountain
regions and operate under a wide  variety  of  climatic  and
topographic conditions.

Historically, the ferroalloy mining and milling industry has
undergone  sharp  fluctuation  in  response to the prices of
foreign ores, government policies, and production  rates  of
other  metals  with  which some of the ferroalloy metals are
recovered as byproducts  (for example, tin and copper, Refer-
ence 6 ) .  Many deposits of ferroalloy metals  in  the  U. S.
are  of  lower grade  (or more difficult to concentrate) than
foreign ores and  so  are  only  marginally  recoverable  or
uneconomic at prevailing prices.  Large numbers of mines and
mills  were  worked  during  World Wars I and II, and during
government stockpiling programs  after  the  war,  but  have
since  been  closed.   At  present,  ferroalloy  mining  and
milling is at a very low level.  Increased competition  from
foreign  ores, the depletion of many of the richer deposits,
and  a  shift  in  government  policies   from   stockpiling
materials  to  selling  concentrates  from  stockpiles  have
resulted in the closure of  most  of  the  mines  and  mills
active in the late 1950»s.  For some of the metals, there is
little  likelihood  of  further  mining  and  milling in the
foreseeable future; for others, increased production in  the
next  few  years  is  probable.   Production figures for the
ferroalloy  mining  and  milling  industry  since  1945  are
summarized in Table III-18.

As   Table  III-18  shows,  molybdenum  mining  and  milling
constitute the  largest  and  most  stable  segment  of  the
ferroalloy  ore  mining  and  milling industry in the United
States.   The  U.S.   produces  over  85%  of  the   worlds
molybdenum  supply,  with two mines dominating the industry.
These two mines are among the 25 largest  mining  operations
in  the U.S.  Production is expected to increase in the near
future with expanded output from existing facilities, and at
                            74

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            TABLE 111-18. PRODUCTION OF FERROALLOYS BY
                        U.S. MINING AND MILLING INDUSTRY
COMMODITY
Chromium
Columbium and
Tantalum
Cobalt
Manganese
Molybdenum
Nickel
Tungsten
(60% WO3)
Vanadium*
ANNUAL PRODUCTION IN METRIC TONS (SHORT TONS)
1949*
394
(433)
0.5
(0.5)
237
(261)
103,835
(114,427)
10,222
(11,265)
0
1,314
(1,448)
N.A.
1953*
53,470
(58,817)
6.8
(7.4)
572
(629)
129,686
(142,914)
25,973
(28,622)
0
4,207
(4,636)
N.A.
1958t
—
194.7tt
(214.2)
2,202
(2.422)
-
18,634
(20,535)
-
3,437
(3,788)
2,750
(3,030)
1962t
0
—
_
-
23,250
(25,622)
-
7,649
(8,429)
4,749
(5,233)
1968**
0
0
550
(605)
- 43,557
(48,000)
42,423
(46,750)
13,750
(15,150)
8.908
(9,817)
5.580
(6,149)
1972t
0
0
0
16,996
(18,730)
46,368
(51,098)
15,303
(16,864)
6,716
(7,401)
4,435
(4.887)
 •Reference 7
  Reference 3
"•Reference 8
 'Reference 6
                              75

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least one major new operation in Colorado is expected to  be
in operation soon.

The   only  commercially  important  ore  of  molybdenum  is
molybdenite, Mos£.  It is mined by both open-pit and  under-
ground methods and is universally concentrated by flotation.
Commercially  exploited ore currently ranges from 0.1 to 0.3
percent  molybdenum  content  (Reference  8).    Significant
quantities  of  molybdenite  concentrate  are recovered as a
byproduct in the milling of copper and tungsten ores.

Tungsten ores are mined and milled at many locations in  the
U.S.,  but most of the production is from one operation.  In
1971, for example, the Bureau of Mines  reported  66  active
tungsten  mines, but total annual production from 59 of them
was less than 1000 metric tons  (1102 short tons)   each  and,
from five others, less than 10,000 metric tons (11,023 short
tons)  (Reference  2).   These  small  mines  and  mills are
operated intermittently, so it is quite difficult to  locate
and  contact  active  plants  at  any  given time.  Tungsten
production  has  been  strongly  influenced  by   government
policies.    During  stockpiling  in  1955,  750  operations
produced tungsten ore at $63 per unit in 1970 (unit  =  9.07
kg   (20  Ib)  of  70%  w concentrate); with the sale of some
stockpiled material, only about 50  mines  operated  with  a
price  of  $43 per unit (Reference 8).  Projected demand for
tungsten will exceed supply before the year 2000 at  present
prices,  and production from currently inactive deposits may
be anticipated (Reference 8) .

Commercially  important  ores  for  tungsten  are  scheelite
(CaWOj»)  and the wolframite series, wolframite ((Fe, Mn) WO<1) ,
ferberite   (FeWOf»),  and  huebnerite  (MnW04.   Underground
mining predominates, and concentration is by a wide  variety
of  techniques.  Gravity concentration, by jigging, tabling,
or sink/float  methods,  is  frequently  employed.   Because
sliming  due  to  the high friability of scheelite ore  (most
U.S.  ore  is  scheelite)   reduces   recovery   by   gravity
techniques,  fatty-acid  flotation  may  be used to increase
recovery.   Leaching  may  also  be  employed  as  a   major
beneficiation  step and is frequently practiced to lower the
phosphorus content of concentrates.  Ore generally  contains
about  0.6  percent  tungsten,  and  concentrates containing
about 70 percent W03_ are produced.  A  tungsten  concentrate
is also produced as a byproduct of molybdenum milling at one
operation   in   a  process  involving  gravity  separation,
flotation, and magnetic separation.
                            76

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 Manganese and nickel ores are each  recovered  at  only  one
 active  operation  in  the O.S. at this time.  The manganese
 operation is completely dry, having no mine-water  discharge
 and  no mill.  At the nickel mine, small amounts of conveyor
 wash water and scrubber water from  ere  milling  are  mixed
 with  effluents  from  an  on-site smelter and with seasonal
 mine-site runoff.  Water-quality impact from the mining  and
 willing  of  these  two  metals  is  thus presently minimal.
        production of manganese and nickel, however,  may  be
 e*pected to involve considerable water use.

 Manganese is essential to the modern steel industry,  both as
 an  alloying  agent  and  as  a  deoxidizer,   and these uses
 ^omfnate  the  world  manganese  industry   (Reference   9).
 Additional  uses include material for battery electrodes and
 agents for impurity removal in glassmaking.   Domestic  pro-
 auction  of  manganese  ores  and concentrates has generally
 accounted for a very small fraction of U.S. consumption, the
 Majority being supplied from foreign concentrates (Reference
 °) •  A number of  significant  plants  have,   however,   been
 operated   for   manganese   recovery  using   a  variety  of
 Processing methods,  and known ore reserves exist  which  are
 economically recoverable.
    U-S. Bureau of  Mines  divides  manganese-bearing  ores  into
 three classes  (Reference  8) :

    (!)   manganese  ores  (at  least   35  percent  manganese
         content)

    (2)   ferruginous manganese  ore   (10  to  35   percent
         manganese  content)

    (3)   manganiferous   iron  ore   (less  than  10  percent
         manganese  content)

The latter two classes are often  grouped  as  manganiferous
ores  and,  in  recent  years, have accounted for nearly all
domestic production.  In  1971, for example, only  5  percent
°*  the total production  of 43,536 metric tons (48,000 short
tons)  was in the form of  true manganese ores (Reference  8).
future  domestic production is likely on a significant scale
rrom manganiferous ores — particularly, on the Cuyuna Range
*n Minnesota,  where  preparations  for  the  resumption  of
Production  are  currently  underway.   This  area,  although
currently quiescent,  accounted for 85  percent  of  domestic
Production in 1971  (Reference 8) .
                            77

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Manganese  ores  have  been  processed  by a wide variety of
techniques, ranging from  dry  screening  to  ore  leaching.
Notable  concentrating  procedures  in  the recent past have
included   sink-float   separation,   fatty-acid   flotation
(References 10, 11, 12, 13), and ammonium carbamate leaching
(Reference   14).    It  is  most  likely  that  heavy-media
separation will be practiced in the immediate future.

Nickel ores are not currently being exploited  in  the  U.S.
One nickel lateritic deposit is currently being mined.  Some
sulfide  nickel  ore  deposits with commercial possibilities
have been found in Alaska (Reference 2) .  If they are devel-
oped, processes entirely different from those in use at  the
present operation will be employed.  Most likely, processing
will  involve  selective  flotation  with  reagent and water
usage and pollution  problems  quite  similar  to  those  of
Canadian nickel operations  (Reference 15).

There  are  no  mines  or mills currently active in the U.S.
producing ores or concentrates of chromium,  cobalt,  colunr
bium,   and  tantalum.   Further,  no  operations  could  be
identified  where  they  are  recovered  as  a   significant
byproduct,  although  the metals and their compounds are re-
covered at a number of  domestic  smelters  and  refineries.
This  production  is primarily from foreign ores and concen-
trates but includes some recovery from domestic concentrates
of other metals.

Chromium ore production in the U.S. has occurred only  undef
the  impetus  of  government efforts to stimulate a domestic
industry.  Production of chromite ore  from  the  Stillwater
Complex  during  world  War  II, and from 1953 through 1961*
involved gravity concentration by tabling, and this mode  of
operation  is  likely  in  the  event  of future production-
Leaching of foreign concentrates,  as  currently  practiced/
might   provide   an  alternative  method  of  concentrating
chromium values in domestic ores.   Domestic  production  by
any  means is unlikely, however, for the next several years*
Production  costs  for  chromium  from  domestic  ores   are
estimated  to  be  $110 per metric ton ($100 per short ton) i
and no shortage is expected in the near future.

Cobalt has been recovered in significant quantities  at  two
locations in the U.S., neither of which is currently active.
One  of  these,  in the Blackbird district at Cobalt, Idaho/
has some probability  of  further  production  in  the  nest,
future.   At these sites, as at essentially all sites around
the world, cobalt is  a  coproduct  or  byproduct  of  otheJf
metals,  and  the  production rates and world price of these,
                            78

-------
 other metals,  particularly copper  and  nickel, exert  primary
 influence   on  the  cobalt  market   (Reference  6).   Known
 domestic  ore from which  cobalt  might be  recovered  is  a
 complex   copper  cobalt  sulfide   ore  which is likely to be
 processed by selective  flotation and roasting  and  leaching
 of the cobalt-bearing float product  (Reference 6).

 Columbium  and  tantalum   concentrates have in the past been
 produced  at as many as  six sites in the U.S. (Reference 16),
 and   several  potentially  workable  deposits  of  the   ore
 minerals    pyrochlore  and  euxenite   are  known.   Economic
 recovery  would require  a  twofold increase in price  for  the
 metals,   however,  and is  considered unlikely before the year
 2000  (Reference 6).   Production,  should  it  occur,  would
 involve placer mining at  one of the known deposits, with the
 water quality  impact and  treatment problems peculiar to that
 activity.    concentration  techniques  varying  widely  from
 fairly simple  gravity and hand  picking  techniques  through
 magnetic   and   electrostatic separation  and flotation have
 been  used in the past.  Accurate prediction of  the  process
 which would  be  used  in future  domestic production is not
 feasible.

 Vanadium.    Eighty-six  percent of  vanadium oxide  production
 has   recently  been used in the preparation of ferrovanadium.
 Although  a  fair share of  U.S. vanadium production is derived
 as a  byproduct of  the mining of  uranium,  there  are  other
 sources  of  vanadium ores.   The environmental considerations
 at    mine/mill  operations   not   involving    radioactive
 constituents are fundamentally different from those that are
 important  at uranium operations, and it seems appropriate to
 consider   the   former  operation   separately.   Vanadium  is
 considered as  part of this  industry segment:  (a)  because of
 the   similarity   of   non-radioactive   vanadium   recovery
 operations to  the  processes  used for other ferroalloy metals
 and (b) because, in  particular,  hydrometallurgical processes
 like  those  used  in  vanadium  recovery  are becoming more
 popular in SIC  1061.  These  arguments  are also presented  in
 the  discussion of   the  SIC  1094  (uranium,  vanadium,  and
 radium mining  and  ore dressing)  categories.    other  aspects
 of effluent  from uranium/vanadium byproduct  operations under
Nuclear  Regulatory  Commission  (formerly  AEC)  license are
 treated further under that heading.

Vanadium is  chemically  similar to  columbium  (niobium)   and
 tantalum,  and  ores  of these metals may be beneficiated in
the same type of process used for vanadium.   There  is  also
some similarity to tungsten, molybdenum,  and chromium.
                            79

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Ferroalloy Ore Beneficiation Processes

Ore  processing  in  the ferroalloys industry varies widely,
and  even  ores  bearing  the  same  ore  mineral   may   be
concentrated  by widely differing techniques.  There is thus
no scheelite recovery process  or  pyrolusite  concentration
technique  per  se.  On the other hand, the same fundamental
processes may be used to concentrate ores of  a  variety  of
metals  with  differences  only  in  details  of  flow rate,
reagent dosage etc., and some functions  (such  as  crushing
and  grinding  ore)  that  are  common  to  nearly  all  ore
concentration  procedures.   Fundamental  ore  beneficiation
processes  which  require  water  may  be grouped into three
basic classes:

         1.   Purely physical separation (most commonly,  by
              gravity)

         2.   Flotation

         3.   Ore Leaching.

Prior to using any of these processes, ore must, in general,
be crushed and ground; in  their  implementation,  accessory
techniques such as cycloning, classification, and thickening
may be of great importance.

Physical  Ore  Processing  Techniques.   Purely physical ore
beneficiation relies on physical differences between the ore
and  accessory  mineralization  to  allow  concentration  of
values.  No reagents are used, and pollutants are limited to
mill  feed  components  soluble in relatively pure water, as
well as to wear products of milling machinery.  Physical ore
properties  often  exploited   include   gravity,   magnetic
permeability, and conductivity.  In addition, friability (or
its  opposite) may be exploited to allow rejection of gangue
on the basis of particle size.

Gravity  concentration  is  effected   by   a   variety   of
techniques,  ranging  from  the  very  simple  to the highly
sophisticated, including  jigging,  Humphreys  spirals,  and
tabling.   Jigging  is  applicable  to  fairly  coarse  ore,
ranging in size from 1 mm to 13 mm  (approximately  O.OU  to
0.50  inch),  generally  the  product  of secondary crushing
(Reference 6).  Ore is fed as a slurry to the jig,  where  a
plunger  operating  at 150 to 250 cycles per minute provides
agitation.  The relatively dense ore sinks  to  the  screen,
while  the lighter gangue is kept suspended by the agitation
and is removed with the overflow.  Often, a bed  of  coarser
                            80

-------
ore  or  iron  shot is used in the jig to aid in separation.
Sink-float methods rely on the buoyancy forces  in  a  dense
fluid to float the gangue away from denser ore minerals.  It
is   also   a  coarse  ore  separation  technique  generally
applicable  to  particles  which  are   2   mm   to   5   mm
(approximately  0.08 to 0.2 inch in diameter) (Reference 6).
Most commonly, the separation medium is a suspension of very
fine particles of dense materials  (ferrosilicon in the heavy
media separation, and  galena  in  the  Huntington-Heberlein
process).  Light gangue overflows the separation tank, while
ore  is  withdrawn  from  the  bottom.   Both  are generally
dewatered on screens and washed, the separation medium being
reclaimed and returned to the circuit (Reference 17).

Shaking tables and spiral separators are  useful  for  finer
particle   sizes;  generally,  ore  must  be  ground  before
application  of  these  techniques.   A  shaking  table   is
generally  fed  at  one  end and slopes towards the opposite
corner.  Water flows over a  series  of  riffles  or  ridges
which  trap the heavy ore particles and direct them at right
angles to the water flow toward the side of the table.   The
table  vibrates,  keeping the lighter particles of gangue in
suspension, and the particles follow the feed  water  across
the  riffles.   The  separation  is  never  perfect, and the
concentrate grades into gangue at  the  edge  of  the  table
through a mixed product called middlings, which is generally
collected  separately  from  concentrate and gangue and then
retabled.  Frequently, several sequential stages of  tabling
are  required to produce a concentrate of the desired grade.
Particle size, as well as density, affects the  behavior  of
particles  on  a shaking table, and the table feed generally
must be well classified to ensure both high ore recovery and
a good concentration ratio.  Humphreys spiral separators are
useful  for  ore  ground  to  between  0.1  mm  and   2   mm
(approximately  0.004  to  0.08  inch)   (Reference 6).  They
consist of a helical conduit about a vertical axis which  is
fed  at the top with flow down the spiral by gravity.  Heavy
minerals concentrate at the inner edge and may be drawn  off
at ports along the length of the spiral; wash water may also
be  added  there  to  improve separation.  The capacity of a
single spiral is generally 0.45  to  2.27  metric  tons/hour
(0.5 to 2.5 short tons/hour)  (Reference 18).

Magnetic  and  electrostatic  separation are frequently used
for the separation of concentrates of different metals  from
complex  ores — for example,  the separation of cassiterite,
columbite,  and monazite (Reference 6)  or the  separation  of
cassiterite  and  wolframite  (Reference 19).  Although they
are both most frequently implemented as dry processes,  wet-
                            81

-------
belt  magnetic separators are used.  Since ore particles are
charged to 20,000 to 40,000 volts for electrostatic

separation, no wet process exists.  In magnetic  separation,
particles  of high magnetic permeability are lifted and held
to a moving belt by a strong magnetic field, while low  per-
meability  particles  proceed with the original stream  (wet-
belt  separator)  or  belt  (crossed-belt  separator).    In
electrostatic  separation,  charged  nonconductive particles
adhere to a rotating conductive drum, while conductive part-
icles discharge rapidly and fall or are thrown off.

These processes may be combined with each  other,  and  with
various  grinding  mills, classifiers, thickeners, cyclones,
etc., in an almost endless variety of mill flow sheets, each
particularly suited  to  the  ore  for  which  it  has  been
developed.   These  flow  sheets  may  become quite complex,
involving multiple recirculating  loops  and  a  variety  of
processes  as  the  examples from the columbium and tantalum
industry shown in Figures 111-13 and III-1U illustrate.   It
is  believed  that  domestic  mills currently employing only
physical separation will  have  fairly  simple  flow  sheets
since  they  are  all  small  processors.  Such an operation
might be represented by the flow sheet of Figure 111-15.

Water use in physical beneficiation plants may  vary  widely
from  zero  to three or more times the ore milled by weight.
However, there are no technical obstacles  inherent  in  the
process  to  total  reuse  of water  (except for the 20 to 30
percent by weight retained by tails) by recycle  within  the
process or from the tailings pond.

Flotation  Processes.   Flotation concentration has become a
mainstay  of  the  ore  milling  industry.   Because  it  is
adaptable to very fine particle sizes (less than 0.01 mm, or
0.0004  inch),  it allows high rates of recovery from slimes
which are inevitably generated in crushing and grinding  and
are  not  generally  amenable  to physical processing.  As a
physico-chemical surface phenomenon, it can  often  be  made
highly   specific,   allowing   production   of   high-grade
concentrates from very-low-grade ore  (e.g., 95+ percent MoS2_
concentrate  from  0.3   percent)    (Reference   19).    Its
specificity also allows separation of different ore minerals
(e.g.,  CuS  and  MoSJ)  where  desired,  and operation with
minimum reagent consumption  since  reagent  interaction  is
typically  only  with the particular materials to be floated
or depressed.
                             82

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Figure  111-13. GRAVITY-PLANT FLOWSHEET FOR NIGERIAN COLUMBITE
                                                          - OVERFLOW-
                                                                              TO
                                                                             WASTE
                                                    -OVERFLOW-





15-cm (6 in.l
PUMP


CYCLONES
(OPTIONAL)
*
4
SHAKING TABLES
                                                              -OVERFLOW-
                                                               -TAILS -
                                                       •—r
                                                   HEADS
                                                                      -TAILS-
                                                            - MIDDLINGS-

                                                       -HEADS-

                                                         -OVtRSIZE-
                           IBLED PERIODICALLY!
                                    HfADS-
                            SOURCE: REFERENCE 20
                               83

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   Figure 111-14. EUXENITE/COLUMBITE BENEFICIATION-PLANT FLOWSHEET
TO

r
DREDGE 1

HEAVY MINERAL
CONCENTRATE
.t
TO WASTED— SLIME

-^- OUARTZ >
RAGE

RAGE"*" °*"NfT|-
,„ -^—TAILINGS ^
i_

c

i •
-4 	 1 CLASSIFIER -«* ATTRITIONS* «J,




, r
t L
INDUCED-ROLL
MAGNETIC SEPARATOR
*



STORAGE J

1
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MR TABLE
CONCENTRATE
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FcROSSBELT MAGNETIC "1 	 NONMAONETICS
1 	 . iiMRATOB. | AND MIDDLINGS '*J
| MONAZITE ]

EUXENITE 1 COLUMBITE 1
	 •*•
— J
                                                                   STORAGE
                                                                  TO
                                                                  STORAGC
                                    TO STORAGE
                               84

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    Figure 111-15. REPRESENTATIVE FLOW SHEET FOR SIMPLE GRAVITY MILL
                            MINING
                             ORE
   TO
WASTE'
TAILS-
                         GRINDING AND
                           CRUSHING
                          SCREENING
                             FINE
                            1
                           SHAKING
                            TABLE
                      MIDDLINGS	'
                           TAILS -
                                           COARSE
                         - HEADS-*
                                                  MIDD
                                         SHAKING
                                         TABLES
                                           T
                                            .INGS
                                              CONCENTRATE
                                85

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Details of the flotation process ~ exact suite  and  dosage
of  reagents,  fineness  of  grinding,  number  of regrinds,
cleaner-flotation steps etc.t — will differ at each  opera-
tion  where  practiced;  and  may  often vary with time at a
given mill.  The complex system of reagents  generally  used
includes   four   basic   types  of  compounds:  collectors,
frothers,   activators,   and   depressants-     Frequently,
activators  are  used to allow flotation of ore depressed at
an earlier stage of the  milling  process.   In  almost  all
cases,  use  of  each reagent in the mill is low—generally,
less than 0.5 kg per metric ton of ore (1.00  Ib  per  short
ton) —and  the  bulk  of  the reagent adheres to tailings or
concentrates.  Reagents commonly used  and  observed  dosage
rates are shown in Table 111-19.

Sulfide  minerals  are  all  readily  recovered by flotation
using similar reagents  in  small  doses,  although  reagent
requirements  and  ease  of  flotation  do  vary through the
class.  Flotation is generally carried out  at  an  alkaline
pH,  typically 8.5 for molybdenite (Reference 19).  Collect-
ors are most often alkali xanthates with two to five  carbon
atoms — for example, sodium ethyl xanthate (C2H5_0 . NaCS,2) •
Frothers  are  generally  organics  with  a soluble hydroxyl
group and a "non-wettable" hydrocarbon (Reference 18).  Pine
oil (C6.H12OH), for example,  is  widely  used.   Depressants
vary but are widely used to allow separate recovery of metal
values  from  mixed  sulfide ores.  Sodium cyanide is widely
used as a pyrite depressant — particularly, in  molybdenite
recovery.   Activators  useful  in sulfide ore flotation may
include cuprous sulfide and sodium sulfide.

The  major  operating  plants  in  the  ferroalloy  industry
recover  molybdenite by flotation.  Vapor oil is used as the
collector, and pine oil is used as a frother.   Lime is  used
to  control  pH of the mill feed and to maintain an alkaline
circuit.  In addition, Nokes reagent and sodium cyanide  are
used  to  prevent  flotation  of  galena and pyrite with the
molybdenite.  A generalized,  simplified  flowsheet  for  an
operation  recovering  only  molybdenite  is shown in Figure
111-16.  Water use in this operation  currently  amounts  to
approximately  1.8  tons  of water per ton of ore processed,
essentially all of which is process water.  Reclaimed  water
from  thickeners  at  the mill site (shown on the flowsheet)
amounts to only 10 percent of total use.

Where byproducts are recovered with molybdenite, a  somewhat
more   complex   mill   flowsheet   results,   although  the
molybdenite  recovery  circuits  themselves   remain   quite
similar.   A very simplified flow diagram for such an opera-
                            86

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TABLE 111-19. OBSERVED USAGE OF SOME FLOTATION REAGENTS

REAGENT

OBSERVED USAGE
IN KILOGRAMS
PER METRIC TON
IN POUNDS
PER SHORT TON
SULFIDE FLOTATION
Vapor oil
Pine oil
Mokes reagent
MIBC (methylisobutyl carbinol)
Sodium cyanide
Sodium silicate
Starch
Butyl alcohol
Creosote
Miscellaneous xanthates
Commercial frothers
0.1 to 0.4
0.02 to 0.2
0.04
0.02
0.005 to 0.02
0.25 to 0.35
0.0005
0.08
0.45
0.0005 to 0.2
0.002 to 0.2
0.2 to 0.8
0.04 to 0.4
0.08
0.04
0.01 to 0.04
0.50 to 0.70
0.001
0.16
0.90
0.001 to 0.4
0.004 to 0.4
OTHER FLOTATION
Copper sulfate
Sodium silicate
CHeic acid
Sodium oleate
Acid dichromate
Sodium carbonate
Fuel oil
Soap skimmings
Sulfur dioxide
Long-chain aliphatic amines
Alkylaryl sulfonate
Misc. Tradenamed Products
0.4
0.3 to 3
0.06 to 6.5
0.05 to 0.2
0.1 to 0.4
4 to 6
60 to 95*
20 to 50*
6*
—
—
0.02 to 0.4
0.8
0.6 to 6
0.12 to 13
0.1 to 0.4
0.2 to 0.8
8 to 12
120 to 190*
40 to 100*
12*
L
1_mi_
0.04 to 0.8
*IN USE AT ONLY ONE KNOWN OPERATION, NOT NOW ACTIVE
                     87

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             Figure  111-16. SIMPLIFIED  MOLYBDENUM MILL FLOWSHEET
                                          MINING
                                           ORE
CRUSHINGS,
WEIGHING, AND
SCREENING
i


BALL
MILLS
\



CYCLONES 1 	 UNDERFLOW 	 '
, 1
                                            I
                                         OVERFLOW
                                        -W
                        -TAILS
                                      ROUGHER FLOAT
                                        MIDDLINGS
                                                          • CONCENTRATE -
                       — MIDDLINGS —
                        SCAVENGER
                          FLOAT
                       (4 STAGES WITH
                        REGRIND AND
                     INTERNAL RECYCLE)
                                                     —CONCENTRATE-
                                                         • MIDDLINGS •
    CLEANER
     FLOAT
  (6 STAGES WITH
  REGRIND AND
INTERNAL RECYCLE)
                                          TAILS
                                          •i
                                                             i
                                                        CONCENTRATE
                                                                                     — TAILS-
    -*-UNDERFLOW
-L
                        «— OVERFLOW—j     CYCLONES
                                      -UNDERFLOW-
                    CYCLONE
                   OVERFLOW

                      I
                            41   wcnrtwiw
                   THICKENER I— RECLAIM •
                  ..i i  i      I    W/ATFR
                                                           DRYER
                                                                        MOLYBDENUM
                                                                          PRODUCT
TO TAILING
   POND
                                           88

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 Figure 111-17. SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
                      CRUSHING
                      (3 STAGES)
                     28% + 3 MESH
                      GRINDING
                     BALL MILLS
                          1
                    36% + 100 MESH
                     FLOTATION
 CONCENTRATE
                      FLOTATION
                          1
                  96% OF MILL FEED
LIGHT TO TAILS-
                         i
                      GRAVITY
                  HUMPHREY'S SPIRALS
                         i
                       PYRITE
                     FLOTATION
                         1
                       TAILS
                         *
LIGHT TO TAILS
  MONAZITE
CONCENTRATE
   TO TAILS
                       TABLES
                     MONAZITE
                     FLOTATION
                     MAGNETIC
                     SEPARATION
                   1
              NONMAGNETIC
            TIN CONCENTRATE
                                            I
                                       CONCENTRATE
                                            1
                                         CLEANER
                                        FLOTATION
                                        (4 STAGES)
                                           I
-TAILINGS
                                          DRYING
                                            1
                                       MOLYBDENUM
                                       CONCENTRATE
                                        (93% + MoS2)
                        MAGNETIC TUNGSTEN
                           CONCENTRATE
                         89

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tion is  shown  in  Figure  111-17.   Pyrite  flotation  and
monazite flotation are accomplished at acid pH (4.5 and 1.5,
respectively),  somewhat  increasing the likelihood of solu-
bilizing heavy metals.   Volumes  at  those  points  in  the
circuit  are  low,  however,  and neutralization occurs upon
combination with the main mill water flows for  delivery  to
the tailing ponds.  Water flow for this operation amounts to
approximately  2.3 tons per ton of ore processed, nearly all
of which is process water in contact with ore.   Essentially
100  percent recycle of mill water from the tailing ponds at
this mill is prompted by limited water availability as  well
as  by  environmental  considerations  and  demonstrates its
technical  and   economic   feasibility,   even   with   the
complications  induced  by  multiple  flotation circuits for
byproduct recovery.

Other sulfide ores in the ferroalloy cateogry which  may  be
recovered  by  flotation  are  those  of  cobalt and nickel,
although no examples of these practices are currently active
in the U.S.   It  is  to  be  expected  that  they  will  be
recovered  as  coproducts  or  byproducts of other metals by
selective  flotattion  from  complex   ores   in   processes
involving multiple flotation steps.  Some of the most likely
reagents  to  be  used  in these operations are presented in
Table III-20, although  the  process  cannot  be  accurately
predicted  at  this  point.   It  is  expected  that,  as is
generally the case, in  sulfide  flotation,  a  small  total
amount of reagents will be used.

Many  minerals  in addition to sulfides may be and often are
recovered by flotation.  Among the  ferroalloys,  manganese,
tungsten,  columbium, and tantalum minerals are or have been
recovered by flotation.  Flotation of these ores involves  a
very different suite of reagents from sulfide flotation and,
in  some  cases,  has  required substantially larger reagent
dosages.  Experience has indicated these flotation processes
to be, in general,  somewhat  more  sensitive  to  feedwater
conditions  than sulfide floats; consequently, they are less
frequently run with recycled water.

In  current  U.S.  operations,  scheelite  is  recovered  by
flotation  using fatty acids as collectors.  A typical suite
of reagents includes sodium silicate (1.0 kg/metric  ton  or
2.0  lb/short  ton)  oleic  acid   (0,5 kg/metric ton, or 1.0
Ib/short ton), and sodium oleate (0.1 to 0.2 kg/metric  ton,
or 0.2 to 0.4 Ib/short ton).  In addition, materials such as
copper  sulfate  or  acid dichromate may be used in small to
moderate amounts as  conditioners  and  gangue  depressants.
Scheelite  flotation  circuits  may  run  alkaline  or acid.
                            90

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TABLE 111-20. PROBABLE REAGENTS USED IN FLOTATION OF
             NICKEL AND COBALT ORES
                   Lime
                   Amyl Xanthate
                   Isopropyl Xanthate
                   Pine Oil
                   Methyl Isobutyl Carbinol
                   Triethoxybutane
                   Dextrin
                   Sodium Cyanide
                   Copper Sulfate
                  Sodium Silicate
                              91

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depending primarily on the accessory mineralization  in  the
ore.   Flotation of sulfides which occurs with the scheelite
is also common practice.   Sulfide  float  products  may  be
recovered   for   sale  or  simply  removed  as  undesirable
contaminants for delivery  to  tails.   Frequently,  only  a
portion  of  the ore  (generally, the slimes)  is processed by
flotation,  the  coarser  material  being  concentrated   by
gravity  techniques  such  as  tabling.   A  simplified flow
diagram for a small tungsten concentrator illustrating these
features is shown in Figure  III-18.   Note  that,  in  this
operation,  an acid leach is also performed on a part of the
tungsten  concentrate.   This  is  common  practice  in  the
tungsten  industry as a means of reducing phosphorus content
in the concentrates.  Approximately four tons of  water  are
used per ton of ore processed in this operation.

The basic flotation operations for manganese ores and colum-
bium and tantalum ores are not much different from scheelite
flotation; in general, they differ in specific reagents used
and,  sometimes,  in reagent dosage.  One past process for a
manganese ore, however, bears special mention because of its
unusually high reagent usage — which could, obviously, have
a strong effect on effluent character and treatment.

Reagents used include:

Diesel oil                   80 kg/metric ton
                                  (160 Ib/short ton)

Soap skimmings               HQ kg/metric ton
                                  (80 lb/short ton)

Oronite S (wetting agent)      5 kg/metric ton
                                   (10 Ib/short ton)

S02                           5 kg/metric ton
                                  (10 Ib/short ton)

With the exception of reagent consumption,  the  plant  flow
sheet  is  typical  of  a straight flotation operation (like
that shown in Figure 111-16),  involving  multiple  cleaning
floats with recycle of tailings.

While  the  flotation  processes  are similar, columbium and
tantalum flotation plants are likely to possess  an  unusual
degree  of  complexity  due  to  the complex nature of their
ores, which necessitates multiple processes  to  effectively
separate  the  desired concentrates.   This is illustrated in
                            92

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Figure 111-18. SIMPLIFIED FLOW DIAGRAM FOR SMALL TUNGSTEN CONCENTRATOR
                 ORE
                               SULFIOE
                              FLOTATION
                               CYCLONE
                                  I
                               75% SANDS
                               GRAVITY
                               TABLES
                          JJ
                            TAILINGS
                                           25%
                                          SLIMES
                                                         OVERFLOW
    THICKENER
    SCHEELITE
    FLOTATION
                                                    HCI LEACH
                                                   (15 TO 20% OF
                                                     FRACTION)
  TUNGSTEN
CONCENTRATE
                             93

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the flowsheet for a Canadian pyrochlore (NaCaCb206F) mill in
Figure 111-19.                                   "

Ore Leaching  Processes.   While not a predominant  practice
in  the ferroalloys industry, ore leaching has played a part
in a number of operations and is likely to increase as  seg-
ments  of  the industry process ores of lower grade or which
are  less  easily  beneficiated.   A  number   of   leaching
processes  have  been  developed  for  manganese ores in the
search  for  methods  of  exploiting  plentiful,  low-grade,
difficult-toconcentrate  domestic ore (that from most of the
state of Maine, for example) (Reference  7),  and  one  such
process   has  been  commercially  employed.   As  mentioned
previously, leaching of concentrates for phosphorus  removal
is common practice in the tungsten industry, and the largest
domestic  tungsten  producer  leaches scheelite concentrates
with soda ash  and  steam  to  produce  a  refined  ammonium
paratungstate   product.   Leaching  is  also  practiced  on
chromite  concentrates  (although  not  as  a  part  of  the
domestic  mining and milling industry).  Vanadium production
by leaching nonradioactive  ores  will  also  be  considered
here, because of vanadium1s use as a ferroalloy, and because
it   provides   a   welldocumented   example  of  ferroalloy
beneficiation  processes  not  well-represented  in  current
practice, but likely to assume importance in the future.

Leaching  processes  for  the  various  ores  clearly differ
significantly in many details, but all have  in  common  (1)
the  deliberate solubilization of significant ore components
and (2)  the use of large amounts of  reagents  (compared  to
flotation,  for  example).   These processes share pollution
problems not generally encountered elsewhere,  such  as  ex-
tremely  high levels of dissolved solids and the possibility
of establishing density gradients in  receiving  waters  and
destroying  benthic  communities despite apparently adequate
dilution.

The processes for the recovery of vanadium in  the  presence
of  uranium  are  discussed  in  the  subsection on uranium.
Recovery from phosphate rocks in  Idaho,  Montana,  Wyoming,
and  Utah  —  which contain about 28% P^OI), 0.25% V205_, and
some Cr, Ni, and Mo — yields vanadium  as"" a  byproduct  of
phosphate  fertilizer  production.   Ferrophosphate is first
prepared by smelting a charge  of  phosphate  rock,  silica,
coke,  and  iron  ore   (if not enough iron is present in the
ore) .   The  product  separated  from  the  slag   typically
contains  60  percent  iron,  25  percent phosphorus, 3 to 5
percent chromium, and 1 percent nickel,  it  is  pulverized,
mixed with soda ash (Na_2C0.3) and salt, and roasted at 750 to
                              94

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Figure 111-19. MILL FLOWSHEET  FOR A CANADIAN COLUMBIUM OPERATION
                           CONCENTRATE
                                    MAGNETIC
                                   SEPARATION
                                                                   MAGNETITE
                                    DESLIMING
                                   UNDERFLOW
                                  BULK FLOTATION
                                   CONCENTRATE
                                    OESLIMING
                                   UNDERFLOW
                                      i	
                                    MAGNETIC
                                   SEPARATION
                                 PRIMARY CLEANING
                                  (FIRST STAGE I
                                  CONCENTRATE
                                      I-	1
                                 PRIMARV CLEANING
                                  (SECOND STAGE)
                                  CONCENTRATE
                                          	|
                               SECONDARY CLEANING
                                 (THREE STAGES)
                                  CONCENTRATE
                                                         ™l
                                    TABLING
                                               I	TAILS -*J
                                    REVERSE
                                   FLOTATION
                                                                                      TO
                                                                                   STOCKPILE
—CONCENTRATE-
                    SPIRALS
                                                                    TAILS
                                                                   TABLING
                                                                 CONCENTRATE
                                                                                CONCENTRATE
                                 SOURCE: REFERENCE 5
                                         95

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800  degrees  Celsius  (1382  to  1472  degrees Fahrenheit).
Phosphorus, vanadium, and chromium are converted  to  water-
soluble trisodium phosphate, sodium metavanadate, and sodium
chromate,  while  the  iron remains in insoluble form and is
not extracted in a water leach following the roast.

Phosphate values are removed from the leach in three  stages
of   crystallization   (Figure  111-20).   Vanadium  can  be
recovered as V2QS (redcake)  by acidification,  and  chromium
is  precipitated" as  lead  chromate.   Ey  this process, 85
percent of vanadium, 65 percent of chromium, and 91  percent
phosphorus can be extracted.

Another,  basically  non-radioactive,  vanadium  ore, with a
grade of 1 percent V205, is found in a vanidiferous,  mixed-
layer    montmorillonite/illite   and   goethite/montroseite
matrix.  This ore is opened up by salt  roasting,  following
extrusion of pellets, to yield sodium metavanadate, which is
concentrated   by   solvent  extraction.   Slightly  soluble
ammonium  vanadate  is  precipitated  from   the   stripping
solution  and  calcined to yield vanadium pentoxide.  A flow
chart for this process is shown in Figure 111-21.

The Dean Leute ammonium  carbamate  process  has  been  used
commercially  for  the  recovery  of  high-purity  manganese
carbonate  from  low-grade  ore  on  the  Cuyuna  Range   in
Minnesota  and  could  be  employed again (Reference 14).  A
flow sheet is shown in Figure 111-22.

Mercury Ores

The mercury mining and milling industry is defined for  this
document  as  that segment of industry engaged in the mining
and/or milling of ore for the primary or byproduct/coproduct
recovery  of  mercury.   The  principal  mineral  source  of
mercury  is  cinnabar (HgS).  The domestic industry has been
centered in California, Nevada,  and  Oregon.   Mercury  has
also  been  recovered  from  ore  in Arizona, Alaska, Idaho,
Texas, and Washington and is recovered as a  byproduct  from
gold ore in Nevada and zinc ore in New York.

Due to low prices and slackened demand, the domestic primary
mercury  industry  has been in a decline during recent years
(Table   III-21).    During   this   time,   the   potential
environmental  problem and toxic nature of mercury have come
under public scrutiny.  One result has been the cancellation
in March 1972 of all biocidal  uses  of  mercury  under  the
terms  of  the Federal Insectide, Fungicide, and Rodenticide
Act.  In  addition,  registration  has  been  suspended  for
                            96

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Figure 111-20. FLOWSHEET OF TRISTAGE CRYSTALLIZATION PROCESS FOR
              RECOVERY  OF VANADIUM, PHOSPHORUS, AND CHROMIUM
              FROM WESTERN FERROPHOSPHORUS
        I FERROPHOtPHOBUS |
             T_
                  I NiCI j
                              { ROASTING |
                               LEACHING \*-
                              "T~^
                          SETTLING AND DECANTATION

                    SOLUTION
              SOLIDS

              J_
                                                      WATER WASH
                                                 FILTRATION
             PRIMARY CRYSTALLIZATION

                                             RESIDUE
                                                        WASH
                                                        LIQUOR
       REGNANT
       SOLUTION
                               C"V»T*L«
          1
                i
                                        DISSOLUTION
       [ TERTIARY CRYSTALLIZATION
              ™
   | SECONDARY CRYSTALLIZATION]
              TERTIARY CENTRIFUCING
       N»jHP04
      " CRYSTALS
            SECONDARY CENTRIFUOINO |




        CRYSTALS
                                                        SECONDARV
                                                        SOLUTION
                                                                   *. TO WASTE
                                                                     TO
                                                                    • STOCKPILE
         RED CAKE  MOTHER LIQUOR


                     *	
                         FILTRATION
r— 	 T
•LACK-CAKE
VANADIUM PRODUCT
Cill

CHRC
NO, WIU

MATE
TION

I LEAD 1
| NITRATE |
s»-TO

                                        FILTRATION
                     TO
                    WASTE
CHROMIUM
PRODUCT
                                                         SOLUTION
                                                         TO WASTE
                                     .TO
                                      STOCKPILE
                                 97

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   Figure 111-21. ARKANSAS VANADIUM PROCESS FLOWSHEET
                        1.5 - 2.0% V2O5
                             *
       6-10%
       NaCL
                         GRINDING
                        PELLETIZING
                         ROASTING
                 850°C (1562°F)
                           NaV03
H2S04


LEACHING AND
ACIDIFICATION
TERTIARY
 AMINES
                            I
                                     pH 2.5 • 3.5
                         (Na6V10028
                   SODIUM DECAVANADATE)
                            I
SOLVENT EXTRACTION
                       PRECIPITATION
                            I
                         NH4V03
                        PRECIPITATE
                                                 NH4OH
                        CALCINING
                            t
                       V2O5 PRODUCT
                     98

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  Figure 111-22. FLOWSHEET OF DEAN-LEUTE AMMONIUM CARBAMATE PROCESS

                        RAW SIZED ORE - 1.9 cm (0.75 in.)
             REDUCING FURNACE
                                                     -CO-
                                    ±
                            GRINDING (30 MESH)
 WEAK Mn
 SOLUTION
                  LEACHING
            (TWO 11,356-4 (3000-GAL) )
               REACTION TANKS
                  IN SERIES
                           9.14-m (30-ft) CLARIFYING
                                THICKENER
7.6-m (25-ft) COUNTERCURRENT
    WASHING THICKENERS
              LIVE
              STEAM
 SLURRY
  STILL
NH4NH2C02
TAILINGS
                                       NEW LEACH LIQUOR-
                                           LEACH LIQUOR
                                           REGENERATION
                 TWO 11,356-j£
                  (3000-GAL)
                PRECIPITATION
              TANKS (IN SERIES)

• NH,
                                  MnCO,
                               CLARIFYING
                               THICKENER
                                   MOTHER
                                  ' LIQUOR "
                 70% SOLIDS
                              ROTARY DRYER
                                   NH4NH2C02
                                PRODUCT
                                99

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  TABLE 111-21. DOMESTIC MERCURY PRODUCTION STATISTICS
CATEGORY
No. of producing mines
Production in metric tons
(flasks)
Dollar value (thousands)
YEAR
1969
109
1,029
(29.640)
$14,969
1970
79
948
(27,296)
$11,130
1971
56
621
(17,883)
$ 5,229
1972
21
253
(7.286)
$1,590
1973
6


SOURCE: REFERENCE 2
                       100

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mercury  alkyl  compounds and nonalkyl uses on rice seed, in
laundry products,  and  in  marine  antifouling  paint.   An
immediate effect of this has been a substantial reduction in
the   demand   for   mercury  for  paints  and  agricultural
applications.  However, future growth in the consumption  of
mercury    is    anticipated   for   electrical   apparatus,
instruments, and dental  supplies.   From  consideration  of
these  factors, it is anticipated that demand for mercury in
1985 will remain at the 1972 level.  Given such variables as
market prices and effects of emission standards  promulgated
in  April  1973,  it  has  been predicted that production of
primary mercury will range from a high of 20,000 flasks  (695
metric tons, or 765 short tons) to a  lew  of  3,000  flasks
(104 metric tons, or 115 short tons by 1985 (Reference 21).

Mercury  ore  is  mined  by  both  open-pit  and underground
methods.   In  recent  years,   underground   methods   have
accounted   for   about  two-thirds  of  the  total  mercury
production.  Ore grade has varied greatly, ranging from 2. 25
to 100 kg of mercury per metric ton  (5  to  200  pounds  of
mercury  per  short  ton).  The grade of ore currently mined
averages 3.25 kg of mercury per metric ton  (6.5  pounds  of
mercury per short ton).

Until  recently,  the  typical  practice of the industry has
been to feed the mined  mercury  ore  directly  into  rotary
kilns  for  recovery of  mercury by roasting.  This has been
such an efficient method  that  extensive  beneficiation  is
precluded.   However, with the depletion of high grade ores,
concentration of low-grade mercury  ores  is  becoming  more
important.   The  ore  may  be  crushed  —  and, sometimes,
screened —  to  provide  a  feed  suitable  for  furnacing.
Gravity  concentration  is also done in a few cases, but its
use is limited since mercury minerals crush more easily  and
more finely than gangue rock.

Flotation  is  the  most  efficient  method of beneficiating
mercury ores when beneficiation is practiced.  An  advantage
of flotation, especially for low-grade material, is the high
ratio    of    concentration    resulting.    This   permits
proportionate reductions  in  the  size  and  costs  of  the
subsequent  mercury  extraction installation.  Only recently
has flotation  of  mercury  been  practiced  in  the  United
States.   During  1975,  a  single  mill, located in Nevada,
began operation to beneficiate mercury ore  by  this  method
(Figure  111-23).    The  concentrate produced is furnaced at
the same facility to recover elemental  mercury.   The  ore,
which  averages  4.8  kg  of  mercury  per  metric  ton (9.5
Ib/short ton), is obtained from a nearby open-pit mine;   the
                            101

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Figure 111-23. FLOW DIAGRAM FOR BENEFICIATION OF
           MERCURY ORE BY FLOTATION
      ORE
  CLAY HOPPER
       i
  AUTOGENOUS
      MILL
   CLASSIFIER
    ROUGHER
 FLOTATION CELLS
    CLEANER
FLOTATION CELLS
      I
    CLEANER
 FLOTATION CELLS
      I
WATER
   THICKENER
                 OVERFLOW
    FILTER
 CONCENTRATE
   PRODUCT
       TAILINGS
                 102

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major ore minerals present are cinnabar  (HgS) and corderoite
 (Hg^S2C12) .   Production  of  concentrate  at  this mill was
initially anticipated to be 984  metric  tons  (1,085  short
tons)  per  year,  from which 20,000 flasks  (34.5 kg (76 Ib)
per  flask)  would  be  recovered  by  furnacing.   However,
following  startup,  the  rate  of  production has been only
1,200 flasks per month, approximately 450 flasks  per  month
below  capacity.   The  primary  cause  for this low rate of
production  has  been  the   recent   drastic   market-price
reduction  for  mercury  ($132/flask during 1975, as compared
to $286 during  1973).  However, this production accounts for
nearly all the present domestic production of mercury.

Uranium, Radium, and Vanadium Ores

The mining and milling  of  uranium,  vanadium,  and  radium
constitute   one   industry  segment,  because  uranium  and
vanadium are sometimes found in the  same  ore  and  because
radium, resulting from the radioactive decay of uranium, has
always  been  obtained  from  uranium  ores.  In the past 20
years, the demand for radium has diminished  as  radioactive
isotopes  (e.g., Co 60, Pu 239) with tailored characteristics
as  sources  of  radiation have become available.  Radium is
now treated as a pollutant in the wastes.  Uranium is  mined
primarily  for  its use in generating energy and isotopes in
nuclear  reactors.   In  the  U.S.,  vanadium  is  primarily
generated  as  a  byproduct  of  uranium mining for use as a
ferroalloying metal and, in the form  of  its  oxide,  as  a
catalyst.   Vanadium  used  as  a  ferroalloy metal has been
discussed in the Ferroalloys Section.

The ores of uranium, vanadium, and radium are found both  in
the oxidized and reduced states.  The uranium (IV)  oxidation
state  is easily oxidized and the resulting uranium  (VI), or
uranyl, compounds are soluble in various  bases  and  acids.
In  arid  regions of the western United states, the ores are
found in  permeable  formations  (e.g.,  sandstones),  while
uranium  deposits  in  humid regions are normally associated
with more impervious  rocks.   Uranium  is  often  found  in
association  with  carbonaceous  fossils,  i.e., lignite and
asphalts.  Ores with a grade in excess of a  fraction  of  a
percent  uranium are rare (80% of the industry operates with
ores below 0.2%).

Because it would  be  uneconomical  to  transport  low-grade
uranium  ores  very  far,  mines are closely associated with
mills that yield a concentrate containing about  90  percent
uranium  oxide.   This concentrate is shipped to plants that
produce  compounds  of  natural  and  isotopically  enriched
                            103

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uranium for the nuclear industry.  The processes of crushing
and  grinding,  conventionally  associated  with a mill, are
intimately connected with the  hydrometallurgical  processes
that  yield  the  concentrate,  and  both processes normally
share  a  wastewater  disposal  system.   Mine  water,  when
present,  is  often treated separately and is sometimes used
as a source of mill process water.   Mine  water  frequently
contains  a  significant  amount  of uranium values, and the
process of cleaning up mine water not only yields as much as
one percent of the product of some mines but is  also  quite
profitable.

The uranium oxide concentrate, whose grade is usually quoted
in  percent  of  U30JJ  (although  that  oxide figures in the
assay, rather than in the product), is generated by  one  of
several   hydrometallurgical  processes.   For  purposes  of
wastewater categorization,  they  may  be  distinguished  as
follows:

    (1)  The ore is leached either in sulfuric acid, or in a
         hot  solution  of  sodium  carbonate   and   sodium
         bicarbonate,  depending  on  the  content  of acid-
         wasting limestone in the gangue.

    (2)  Values in the leachate are usually concentrated  by
         ion  exchange   (IX)  or by solvent extraction  (SX).
         They are  then  precipitated  as  the  concentrate,
         yellowcake.

Some  vanadium  finds  are  not  associated with significant
uranium   concentrations.    Some   byproduct    concentrate
solutions  are  sold to vanadium mills for purification, and
not all uranium mills separate vanadium, which appears to be
in  adequate  supply  and  could  be  recovered  later  from
tailings.

ores  and Mining.  Consideration of thermonuclear equilibria
suggests an initial abundance of uranium in the solar system
of  0.11  ppm   (parts  per  million).   Since   uranium   is
radioactive,  its concentration decreases with time, and its
present abundance is  estimated  as  0.054  ppm.   The  four
longest-lived  isotopes are found in the relative abundances
shown in Table III-22.

Primary deposits of uranium ore contain uraninite, the U(IV)
compound U02, and are widely  distributed  in  granites  and
pegmatites.   Pure speciments of this compound, with density
ranging to 11, are rare, but its fibrous form,  pitchblende.
                             104

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TABLE 111-22. ISOTOPIC ABUNDANCE OF URANIUM
ISOTOPE
U238
U 235
U234
U236

HALF-LIFE (YEARS)
4.51 x 109
7.13 x 108
2.48 x 105
2.39 x 107

ABUNDANCE
99.27%
0.72%
0.0057%
Traces Identified
(Moon-1972; Earth-1974)
                105

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has  been   exploited   in  Saxony   since   the  recognition of
uranium  in  1789.

Secondary,  tertiary, and higher-order  deposits  of  uranium
ores  are   formed  by  transport   of   slightly water-soluble
uranyl  (D(VI)) compounds, notably  carbonates.  Typically   a
primary  deposit  is   weathered  by  oxidized water, forming
hydrated oxides of uranium  with   compositions  intermediate
between  W2 and U03,   The composition  U308 — i.e., U02 2U03
--  is   particularly stable.  The  process~occasionally~stops
at gummite  (U02.H20),  an orange or red,  waxy  mineral,  but
-?£vL  invSlveslvfu^her  oxidation and  reactions  with
alkaline   and   alkaline-earth    oxides,   silicates    and
phosphates.  The transport leads to the surface uranium ores
or  arid  lands,  including carnotite  (K2(U02) 2 (V04) 2 3H201
uranophane       (CaU2Si2011. 7H20)  ,     " and       lutunite
(Ca(U02,2(POi,2.10-12H20)  and," if  reducing conditions a"re
encountered,  to  the  redeposition  of   0(IV)    compounds.
Vanadium  is seen to follow a similar  route.  Radium, with a
halflife of only  1600  years,  is  generated  from  uranium
deposits in historical times.                        uj.aiu.um

A   reducing  environment  is  often   provided  by  decaying
biological materials; uranium is found in  association  with
lignite,  asphalt,   and dinosaur bones,  one drift at a
                                       ,       r    a  a m
in New Mexico passes  lengthwise through  the  ribcage  of  a
fossil  dinosaur.   since the requisite conditions are often
encountered in the sediments of lakes  or streams, stratiform
uranium  deposits  are   common,  constituting  95%  of  o q
reserves.   stratiform deposits comprise sandstone, conglom-
erate, and limestone  with uranium values in pores or on  the
surface  of  sand  grains or as a replacement for fossilized
organic tissue.  A small fraction of   steeply  sloping  vein
deposits,  similar  to those in Saxony, is found in associa-
tion with other minerals.  Some sedimentary deposits  extend
over  many  kilometers  with  a  slight  dip with resoect
modern grade that  makes  it  profitable  to  mine  f a
Exploration is conducted initially with airborne and surface
radiation sensors that delineate promising  regions  a^d  ?2
followed  by  exploratory drilling, on a 60-m ^SS^ft)  grid
and development drilling, on  a  15-m  (50-ft)   grid    Test
holes  are probed with scintillation counters,  and cores are
chemically analyzed   Reserves have usually  been  Deified
in  terms  of  ore that can yield uranium at $18 pe?kS 12^
Ib) , a price paid by the government for stockpiling!  Recent
increases in price and the possibility of increased  uranISm
                            106

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demand  due to the current energy situation have resulted in
the mining, for storage, of ore below this threshold and may
effect an increase in  reserves.   currently,  reserves  are
concentrated  in  New  Mexico  and  Wyoming, as shown in the
tabulation below.

  DISTRIBUTION OF U.S. URANIUM ORE RESERVES (JAN. 1, 1975)

                   U3O8      No- of Known        % of total
                  (Short Tons)	Deposits

New Mexico         137,108             j>6             £9
Wyoming             28,300             14             14
Utah and
Colorado           Ilr400              .99             5
Texas              14.400              45             7_
Others              8,800              60             5
The number of separate known deposits in the western  United
States  is 284, but half of the reserves lie in 15 deposits.
Four of these, in central Wyoming,  on  the  border  between
Colorado  and  Utah,  in northwestern New Mexico, and on the
Texas gulf coast,  dominate  the  industry.   In  1974,  New
Mexico provided 43 percent and Wyoming 32 percent of uranium
production.   In  1974, the U. S. production was 7.1 million
tons of ore with a U3O8 equivalent of 12,600 tons.

In  the  eastern  United  States,  uranium   is   found   in
conjunction  with  phosphate  recovery in Florida, in states
throughout the Appalachian Mountains, and in Vermont and New
Hampshire  granites.   The  grade  of  these   deposits   is
currently too low for economic recovery of uranium, which is
recovered as a byproduct only in Florida.  Vanadium, in ores
that do not contain uranium values, is mined in Arkansas and
Idaho.   The  humid  environment  of current and prospective
eastern  deposits  presents  special   problems   of   water
management.   Ocean water contains 0.002 ppm of uranium, and
its recovery with a  process  akin  to  ion  exchange  using
titanium  compounds  as  a  "resin" has been explored in the
United Kingdom.  Uranium can be recovered in this fashion at
a cost of $150 to $300 per kg (2.2 Ib) .

Mining practice is conventional.  There are 122  underground
mines  as  of  1 January 1974, with a typical depth of 200 m
(656 ft) .  Special precautions for the ventilation of under-
ground mines reduce the  exposure  of  miners  to  radon,  a
shortlived, gaseous decay product of radium that could leave
deposits  of  its daughters in miners' lungs.  Mine water is
                            107

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occasionally recycled through the mine to recover values  by
leaching and ion exchange.

Because  of  the  small  size  of pockets of high-grade ore,
openpit mines are  characterized  by  extensive  development
activity.   At  present,  low-grade  ore  is  stockpiled for
future use.  stockpiles  on  polyethylene  sheets  are  heap
leached  at several locations by percolation of dilute H2S04
through the ore stockpiles.  On January  1974,  33  open-pit
mines  were being worked, and 20 other (e.g., heap-leaching)
sources were in operation.

Most mines ship ore to the mill by truck.  In at  least  one
instance,  a  short   (100-km,  or  62-mi.)  railroad  run is
involved.   Most  mining  areas  share  at  least  two  mill
processes,  one  using acid leaching and the other, for high
limestone content, using alkaline leaching.

Milling.   Mills range in ore processing capacity  from  450
metric  tons  (495  short  tons) per day to 6500 metric tons
(7,150 short tons) per day, and 15 to 25 mills have been  in
operation  at  any  one time during the last 15 years.  Mill
activities, listed by state, are given in Table  111-23  and
are tabulated by company in Supplement B.

Blending,  Crushing,  and Roasting,  ore from the mine tends
to be quite variable in consistency and grade and  may  come
from  mines  owned  by  different companies.  Fairly complex
procedures have been developed for weighing and  radiometric
assay of ores, to give credit for value to the proper source
and  to  achieve  uniform  grade, and for blending to assure
uniform  consistency.    Sometimes,   coarse   material   is
separated  from  fines  before  being  fed  to crushers that
reduce it to the 5 to 20 mm (0.2 to 0.8  in.)   range.    This
material is added to the fines.

Ore  high  in  vanadium  is  sometimes  roasted  with sodium
chloride at this  stage  to  convert  insoluble  heavy-metal
vanadates  (vanadium  complex)  and carnotite to more soluble
sodium vanadate, which is  then extracted with water.    Ores
high  in  organics  may  be roasted to carbonize and oxidize
these and prevent clogging of hydrometallurgical  processes.
Clayey   ores   attain   improved   filtering  and  settling
characteristics by roasting  at  300  degrees  Celsius  (572
degrees Fahrenheit) .

Grinding.    Ore  is  ground  to  less than 0.6 mm (28 mesh)
(0.024 in.) for acid leaching and to less than .07  mm  (200
mesh)   for alkaline leaching in rod or ball mills with water
                            108

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TABLE 111-23. URANIUM MILLING ACTIVITY BY STATE, 1972
CTT ATE
STATE
New Mexico
Wyoming
Colorado
Utah
Texas
South Dakota
Washington
TOTAL
TOTAL MILL HANDLING CAPACITY
METRIC TONS PER DAY
12,300
8,250
4,000
1,850
3,400
600
450
30,850
SHORT TONS PER DAY
13,600
9,100
4,400
2,000
3,750
660
500
34,010

NO. OF MILLS
3
7
3
2
3
1
1
20
                   109

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(or, preferably, leach) added to obtain a  pulp  density  of
about  two-thirds solids.  Screw classifiers, thickeners, or
cyclones are sometimes used to control size or pulp density.

Acid Leach.   Ores with a calcium carbonate (CaCO3J  content
of  less  than 12 percent are preferentially leached in sul-
furic acid, which extracts values quickly (in four hours  to
a day), and at a lower capital and energy cost than alkaline
leach   for   grinding,   heating,  and  pressurizing.   Any
tetravalent uranium must be oxidized to the uranyl  form  by
the  addition  of  an  oxidizing  agent  (typically,  sodium
chlorate  or  manganese  dioxide),  which  is  believed   to
facilitate  the  oxidation  of U (IV) to U(VI)  in conjunction
with the reduction of  Fe  (III)  to  Fe  (II)   at  a  redox
(reduction/oxidation)  potential  of  about  minus  450  mV.
Free-acid concentration is held to between 1 and  100  grams
per  liter.   The  larger  concentrations  are suitable when
vanadium is to be extracted.  The reactions taking place  in
acid oxidation and leaching are:

                   2U02 * 02  	>  2003

         2UO3 + 2H2SCW + 5H2O	> 2 (UO2SO.4)  . 7H2O

Uranyl  sulfate   (002,504.)  forms a complex, hydrouranyl tri-
sulfuric acid (H4JJO2 (SOj*) 3) , in the leach, and the anions of
this acid are extracted for value.

Alkaline Leach.   A solution of sodium carbonate  (40 to 50 g
per liter) in an oxidizing environment  selectively  leaches
uranium  and vandium values from their ores.  The values may
be precipitated directly from the leach by  raising  the  pH
with  the addition of sodium hydroxide.  The supernatant can
be recycled by exposure to  carbon  dioxide.   A  controlled
amount of sodium bicarbonate (10 to 20 g per liter) is added
to  the  leach  to  lower pH during leaching to a value that
prevents spontaneous precipitation.

This leaching process is slower  than  acid  leaching  since
other ore components are not attacked and shield the uranium
values.   Alkaline  leach  is,  therefore,  used at elevated
temperatures of 80  to  100  degrees  Celsius  (176  to  212
degrees  Fahrenheit)  under  the hydrostatic pressure at the
bottom of a 15 to 20 m  (49.2 to 65.6 ft) tall tank, agitated
by a central airlift  (Figure 111-24).  in  some  mills,  the
leach tanks are pressurized with oxygen to increase the rate
of  reaction, which takes on the order of one to three days.
The alkaline leach process is characterized by the following
reactions:
                              no

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Figure 111-24. PACHUCATANK FOR ALKALINE LEACHING
                  VEIST
               ;**.'«.*.* D
               *&#£••

               r;#;£fe:
               #?i$ti£.
               :*&*.*&.:.:
               •.!*• • '  .  .  .
                                             AIRLIFT
                                              LEACH
               111

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                     2UO2 + 02	> 2UO.3
                       (oxidation)

      3Na2(COl)  + U03 + H20	> 2NaOH + Na£(U02>) (C03J 3_
                    ""    (leaching)

               2NaOH * CC2	> Na2C03_ + E20
                    (recarbonization)

        UO£) (CO3) 3 + 6NaOH	> Na2U2O7   + 6Na2CO3^ +
                     (precipitation)

The efficient utilization of water  in  the  alkaline  leach
circuit  hats  led  to the trend of recommending its expanded
application in the uranium industry.  Alkaline leaching  can
be  applied  to  a  greater  variety of ores than in current
practice; however, the process,  because  of  its  slowness,
appears  to  involve  greater  capital expenditures per unit
production.  In addition, the purification of  yellow  cake,
generated  in  a  loop  using  sodium as the alkali element,
consumes an increment of chemicals that tend  to  appear  in
stored  or  discharged  wastewater  tut  are  often ignored.
Purification to remove sodium ion is necessary both to  meet
the  specifications  of  American uranium processors and for
the preparation of natural uranium dioxide fuel.  The latter
process will be used to  illustrate  the  problem  caused  by
excess sodium.  Sodium diuranate may be considered as a mix-
ture  of  sodium  and  uranyl oxides—i.e., Na2U.2O2 = Na2O +
2003.

The process of generating U
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product  is  water  leached f  yielding a V20.5 concentrate as
described  below.   The  remaining   sodium   diuranate   is
redissolved in sulfuric acid,
    Na£U207 * 3H2SCW --- > Na^SOU. -f 3H2_0 + 2 (U02J S0£

and  the  uranium  values  are precipitated with ammonia and
filtered, to yield a yellow cake  (ammonium diuranate or U03_)
that is low in or free of sodium.                         ~~
         U02SOU * E20 + 2m3 --- >  (NH4)2SO 3NaOH +NH4VO3. * 2NHj*OH
                  (ammonium metavanadate)     ~~      ~"

                  + 3NH4C1 --- > 3NaCl +  (NH4)3VO<»
                  (ammonium orthovanadate)

The  ammonium  vanadates  are  thermally decomposed to yield
vanadium pentoxide;

               3VOI --- >  6NH3   + 3H£O   + V2_0
                            113

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A significant fraction (86 to 87%)  of V2O5 is  used  in  the
ferroalloys   industry.    There,   ferrovanadium  has  been
prepared in electric furnaces by the reaction:

         V2O5 + FeOJ * 8C --- > SCO   + 2FeV
or  by  aluminothermic  reduction  (See  Glossary)   in   the
presence of scrap iron.

Air  pollution  problems  associated  with the salt roasting
process have led  many  operators  to  a  hydrometallurgical
process  of  vanadium  recovery  that  is  quite  similar to
uranium recovery by acid leaching and solvent exchange.   The
remainder of  V2O5  production  is  used  in  the  inorganic
chemical  industry,  and  its  processing  is not within the
scope  of  these   guidelines.    Since   the   mining   and
beneficiation of vanadium ores not containing uranium values
present an excellent example of hydrometallurgical processes
in  the  mining and ore dressing of ferroalloy metals (under
SIC 1061) , it will be explored further under  that  heading.
Because of the chemical similarity of vanadium to columbium,
tantalum,  and  other  ferroalloy metals, recovery processes
for  vanadium  are   likely   to   be   quite   similar   to
hydrometallurgical  processes  that  will  be  used  in  the
ferroalloys mining industry  when  it  becomes  more  active
again.

Concentration and Precipitation.   Tc a rough approximation,
a  metric  ton  of ore with a grade of about 0.2* is treated
with a metric ton  (or cubic meter) of leach, and the concen-
tration (s)  of  uranium  and/or  vanadium  in  the  pregnant
solution  are  also  of  the  order of 0.2X.  if values were
directly precipitated  from  this  solution,  a  significant
fraction   would   remain  in  solution.   Yellow  cake  is,
therefore, recycled and dissolved in  pregnant  solution  to
increase precipitation yield.  Typically, five times as much
yellow  cake  is  recycled  as  is  present  in the pregnant
solution.  Direct precipitation by raising pH  is  effective
only  with  alkaline  leach, which is somewhat selective for
uranium and vanadium.  If it were applied to the acid  leach
 process,  most heavy metals — particularly, iron — would be
precipitated and would severly contaminate the product.

Uranium  (or vanadium and molybdenum) in the  pregnant  leach
liquor  can  be  concentrated  by a factor of more than five
through  ion  exchange  or  solvent   extraction.    Typical
concentrations  in the eluate of some of these processes are
shown in Table 111-24.
                             114

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TABLE 111-24. URANIUM CONCENTRATION IN IX/SX ELUATES
PROCESS
U3O8 CONCENTRATION (%)
Ion exchange
Resin-in-pulp
Fixed-bed IX:
Chloride edition
Nitrate edition
Moving-bed IX:
Nitrate elution
0.8 to 1.2
0.5 to 1.0
1.0 to 2.0
1.9
Solvent extraction
Alkyl phosphates, HCI eluent
Amex process
Dapex process
Split elution minewater treatment
30.0 to 60.0
3 to 4
5.0 to 6.5
1.2 to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
                   115

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Precipitation of  uranium  from  the  eluates  is  practical
without  recycling yellow cake, and the selectivity of these
processes  under  regulated  conditions  (particulary,   pH)
improves the purity of the product.

All  concentration  processes operate best in the absence of
suspended solids, and considerable effort is made to  reduce
the  solids  content  of pregnant leach liquors (Figure III-
25a) .  A somewhat  arbitrary  distinction  is  made  between
quickly  settling  sands  that  are  not  tolerated  in  any
concentration process and slimes that can be accommodated to
some extent in the resin-in-pulp process (Figure III-25b-c).
Sands are often repulped, by the addition of some wastewater
stream or another, to facilitate flow to the tailing pond as
much as a few kilometers away.  Consequently, there is  some
latitude  for  the  selection  of the wastewater sent to the
tailing pond, and mill operators can take advantage of  this
fact   in  selecting  environmentally  sound  waste-disposal
procedures.

Ion exchange and solvent extraction (Figure  IIl-25b-e)  are
based  on  the same principle:  Polar organic molecules tend
to exchange a mobile ion in their  structure  —  typically,
C1-,  N03_-,  HS04.-, C03— (anions) , or H* or Na+ (cations)  —
for an ion with a greater charge or a smaller ionic  radius.
For  example,  let  R be the remainder of the polar molecule
(in the case of a solvent) or polymer (for a resin), and let
X be the mobile ion.  Then, the exchange  reaction  for  the
uranyltrisulfate complex is

         HRX + (U02(SOjl)3)	>   RUU02(SOJ*)3 * 4X
                           <	

This  reaction  proceeds  from  left to right in the loading
process.  Typical resins adsorb about ten percent  of  their
mass  in  uranium  and  increase  by  about  ten  percent in
density.  In a concentrated solution of the  mobile  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-n-*   Mg*+   Na+) and,
because of decreasing ionic radius, with atomic number  (92U
H2  Mo    23V) .  The separation of hexavalent 92U cations by
IX or SX should prove to be easier than that  of  any  other
naturally occurring element.

Uranium,  vanadium,  and  molybdenum  —  the latter being a
common ore constituent — almost always  appear  in  aqueous
                            116

-------
  Figure 111-25. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 1 of 2)
        FROM  •
        LEACH
                           »
                           *L/
       PREGNANT
       LEACH LIQUOR
                ... v.,. r. v.,v.., ..,.  ,,.v/'Xy
                ;•-.•- : ;-/•'.v.Tr\,V'Y V-'" ••:
                '• .  • A.; -.,,«•» «\>  -'  ••%•_
                                           SLIMES
                  SANDS
                              SLIMY PULP TO
                              RESIN-IN-PULP IX
WATER
.
V° •  o  • • • o o  •
                         REPULPING
            «w
               CLEAR LEACH LIQUOR
               TO COLUMN IX OR SX

                    a) LIQUID/SOLID SEPARATION
                                                       SAND

                                                    TAILINGS
SLIMY.
PREGNANT-
PULP
                     RESIN IN OSCILLATING BASKET

                   b) RESIN-IN-PULP PROCESS: LOADING
                                 BARREN
                                 PULP
                                 TO TAILINGS
   BARREN
   ELUANT
                                  PREGNANT
                                  ELUATE TO
                                  PRECIPITATION
                     c)  RESIN-IN-PULP PROCESS: ELUTING
                               117

-------
solutions  as  oxidized  ions (uranyl, vanadyl, or molybdate
radicals), with uranium and vanadium additionally  complexed
with  anionic  radicals to form trisulfates or tricarbonates
in the leach.  The  complexes  react  anionically,  and  the
affinity  of  exchange  resins  and  solvents  is not simply
related  to  fundamental  properties  of  the  heavy   metal
(uranium,  vanadium,  or  molybdenum),  as  is  the  case in
cationic   exchange   reactions.    Secondary    properties,
including  pH and redox potential, cf 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  as  much  uranium  as vanadium being captured.  These
variations in affinity, multiple  columns,  and  control  of
leaching  time  with  respect to breakthrough  (the time when
the interface between loaded and regenerated  resin.  Figure
III-25d,  arrives at the end of the column) are used to make
an IX process specific for the desired product.

In the case of solvent exchange, the type of  polar  solvent
and its  concentration in a typically nonpolar diluent (e.g.,
kerosene)  effect  separation  of  the desired product.   The
ease with which the  solvent  is  handled   (Figure  III-25e)
permits   the  construction  of  multistage  co-current  and
countercurrent SX concentrators that are  useful  even  when
each  stage  effects only partial separation of a value from
an interferent.   Unfortunately,  the  solvents  are  easily
polluted by slimes, and complete liquid/solid separation is
necessary.  IX and SX  circuits  can  be  combined  to  take
advantage  of both the slime resistance of  resin-in-pulp ion
exchange and the separatory efficiency of   solvent  exchange
 (Eluex process).  The uranium values are precipitated with a
base  or a  combination  of  base  and  hydrogen  peroxide.
Ammonia  is preferred by a  plurality  of  mills  because  it
results   in   a  superior  product,  as  mentioned  in  the
discussion  of   alkaline   leaching.    Sodium   hydroxide,
magnesium  hydroxide, or partial neutralization with calcium
hydroxide, followed by  magnesium  hydroxide  precipitation,
are  also  used.   The  product is rinsed with water that is
recycled into the  process  to  preserve  values,  filtered,
dried    and  packed  into  200-liter   (55-gal)  drums.   The
strength of these drums limits  their  capacity  to  450  kg
 (1000  Ib)  of  yellow  cake, which occupies 28% of the drum
volume.

Thorium.   Thorium is often combined with the  rare  earths,
with  which  it is found associated in monazite sands.  It is
actually an  actinide  (rather  than   lanthanide)  and  chemi-
cally, as well as by nuclear structure, is  closely allied to
                             118

-------
Figure 111-25. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 2 of 2)
                               BARREN ELUANT
                                      ELUTED (OR
                                      REGENERATED)
                                      RESIN
                                      LOADED
                                      RESIN
                              PREGNANT ELUATE
                              TO PRECIPITATION

           d) FIXED-BED COLUMN ION EXCHANGE/ELUTION
PR EC
LEAC
LIQU
0
iNANT
;H
OR
ft

1
0 o°o
~H/°
o


^^___ "-^


BARREN
±PIIIAMT

f LOADED 1 1
i— J 1 	 1 ORGANIC r-U 	 1 U
SOLVENT °o o
	 	 ..^- j— , ° ft 0

^ - N
^) 1, (BARREN ^
^-^ J LIQUOR V-'
PHASE
LOADING SEPARATION STRIPPING
e) SOLVENT EXTRACTION

LEACH
IX
t
sx




RECYCLE
BARREN to. ELUANT
ELUANT I |
IX ' '
STRIPPED
SOLVENT
SOLVENT
.

) PREGNANT! i[
ELUATE II1
PHASE
SEPARATION
PREGNANT
ELUATE
IX
PARTIALLY STORAGE)
STRIPPED \ . LOADED
RESIN \ _ ' RESIN
g) SPLIT ELUTION

        PRECIPITATION

f) ELUEX PROCESS
                             119

-------
uranium.   Although  it  finds  some use in the chemical and
electronics industry, thorium is primarily  of  value  as  a
fertile  material  for  the  breeding of fissionable reactor
fuel.  In this process, thorium 232,  used  in  a  "blanket"
around  the  core of a nuclear reactor, captures neutrons to
form thorium  233,  which  decays  to  uranium  233  by  the
emission  of two beta particles with haIflives of 22 minutes
and 27 days.  Uranium 233 is fissile and can be  used  as  a
fuel.  The cycle is very attractive since it may be operated
in  thermal-neutron,  as  well as fast-neutron, reactors.  A
pseudo-breeding reactor  (burning uranium  235  or  plutonium
239  in  the core and producing uranium 233 in the blanket),
with net breeding gain  (quantity of fissile  material  bred/
quantity  burned)  less  than  one  is already in commercial
operation.

Thorium is about three  times  as  abundant  as  uranium  in
rocks,  but  rich  deposits are rare.  Typical monazite sand
ores contain from 1 to 10 percent thoria  (ThO2).   American
ores  from the North and South Carolinas, Florida, and Idaho
contain 1.2 to 7 percent Th
-------
Figure 111-26. SIMPLIFIED SCHEMATIC DIAGRAM OF SULFURIC ACID DIGESTION
             OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
             AND RARE EARTHS
                       • MAIN STREAM
            TO
      STOCKPILE'
        RESIDUE
(UNDIGESTED MONAZITE SAND.
SILICA. ZIRCON. AND RUTILE)
                   FILTRATE
                (R.E., U, AND P2O5»
                   SELECTIVE
                  PRECIPITATION
                   AT pH 2.3
                                                  MONAZITE
                                                   SAND
                                                  GRINDING
                                                 OPERATION
                                                 DIGESTION
                                                    I
                                                 DISSOLUTION
                                              Th. R.E.. U. AND P2Og
                                                    I
                                                  SELECTIVE
                                                PRECIPITATION
                                                  AT pH 1.06
                                                 PRECIPITATE
                                               (Th. R.E.. AND P20S)
                                       PURIFICATION BY SOLVENT EXTRACTION,
                                         SELECTIVE PRECIPITATION. OR FRAC-
                                            TIONAL CRYSTALLIZATION
                                                CONCENTRATES
                                                TO SHIPPING
     TO WASTE
                  TO SHIPPING
                             SOURCE: REFERENCE 22
                                121

-------
   Figure 111-27. SIMPLIFIED SCHEMATIC DIAGRAM OF CAUSTIC SODA DIGESTION
             OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
             AND RARE EARTHS
— MAIN STREAM


1
FILTRATE
(NaOH AND Na3PO4)
*
CRYSTALLIZATION
*
i <
MONAZITE
SAND
t
0GPRERAToGN
1 f
*
^_ __ DIGESTION 138°C
Inn k. (7Rn°F»


•
HYDROUS METAL-OXIDE CAKE
(Th, U, AND R.E.)



| MnnH (NariP


1 >
i
b"4) 1
DUCT) 1
1
^ SELECTIVE
PRECIPITATION

'
FILTRATE
(RARE EARTHS)
,
PRECIPITATE
r (Th AND U)
fc. SELECTIVE
"^ PRECIPITATION
1
1
FILTRATE
*
TO WASTE
1
1
PRECIPITATE OF PURIFICATION BY
RARE EARTHS SOLVENT EXTRACTION

t »
TO STOCKPILE CONCENTRATES -*- STOCKPILE
SOURCE: REFERENCE 22
                              122

-------
 Figure 111-28. EFFECT OF ACIDITY ON PRECIPITATION OF THORIUM. RARE
            EARTHS AND URANIUM FROM A MONAZITE/SULFURIC ACID
            SOLUTION OF IDAHO AND INDIAN MONAZITE SANDS
     100 r-
Q
ill
O
ill
cc
0.
UJ

5
D
                                         IDAHO MONAZITE SAND
                                       a INDIAN MONAZITE SAND
                                       A
                                              i     i
      20 -
                                ACIDITY (pH)
                 AGITATION TIME:
                 DILUTION RATIO:
                 DIGESTION RATIO:
                 NEUTRALIZING AGENT: 3.1% NH4OH
                                    5 MINUTES
                                    H20: SAND = 45:1 TO 50:1
                                    93% H2SO4: SAND = 1.77
                       SOURCE: REFERENCE 22
                          123

-------
Recycle of leachant should  be  possible  with  an  alkaline
leach  process that has been evaluated in pilot-plant scale.
The process consumes caustic soda in the formation  of  tri-
sodium  phosphate,  which can be separated to some extent by
cooling the hot  (110 to 137 degrees  Celsius)   (230  to  279
degrees  Fahrenheit)  leach to about 60 degrees Celsius (140
degrees Fahrenheit) and filtering.  Uranium is  precipitated
with  the  phosphate if NaOH concentration is too low during
the crystallization step, and NaOH concentration  should  be
raised  to more than ION before cooling.  The cyclic cooling
and heating of leach to separate phosphate values represents
an energy expenditure  that  must  be  weighed  against  the
environmental benefits of the process.

The  alkaline  leach process is unusual in that the leaching
action removes the gangue in the solute, as sodium silicate,
and leaves the values as  rare-earth  oxides,  thorium,  and
uranium  diuranate  in the residue.  They are preserved as a
slurry or filter cake, which is then dissolved in  sulfuric/
nitric acid and subjected to fractional precipitation, as in
the acid leach process.

The methods for recovering thorium and uranium from monazite
sands  are  almost  identical  to those used in the acid and
alkaline leach processes for  recovering  uranium  from  its
primary  ores.   Thorium production in the U.S. is currently
not  sufficient  to   characterize   exemplary   operations.
Guidelines developed for the uranium mining and ore dressing
industry  and other subcategories related to thorium ore may
generally apply.

Radiation parameters of thorium and  uranium  daughters  are
somewhat  different.   The  two decay series are compared in
Table III-25.  The uranium series is  dominated  by  radium,
which—with   a   halflife   of   1620  years  and  chemical
characteristics that are distinctly different from those  of
the    actinides    and   lanthanides—can   be   separately
concentrated in minerals  and  mining  processes.   it  then
forms  a  noteworthy  pollutant  entity  that  is  discussed
further in Section V.  Thorium, by contrast,  decays  via  a
series  of daughters with short halflives; the longest beina
Ra228 at 6.7 years.                                        y

Industry Flow Charts.   of the sixteen  mills  operating  in
1967  (Table III-26), no two used identical leaching concen-
tration and precipitation steps.  The same was probably true
of the 15  mills  operating  in  1974   (Table  Hi-23   also
Supplement  B).   A  general  flow chart, to be used in con-
junction with Table 111-26, is presented in  Figure  Hi-29.
                            124

-------
TABLE 111-25. DECAY SERIES OF THORIUM AND URANIUM

ELEMENT OR
NAME


SYMBOL(S)


HALF-LIFE
ENERGY OF RADIATION
(MeV)
OC
P
r
Thorium Series
Thorium
Mesothorium 1
Masothorium 2
Radiothorium
Thorium X
Thoron
Thorium A
Thorium B
Thorium C
Thorium C'
Thorium C"
Thorium D
Th232
90Tn
88Ra228
84Po212(ThC<)
81TI208 (ThC")
82Pb208 (ThD)
1.34 x 1010 years
6.7 years
6. 13 hours
1.90 years
3.64 days
54.5 seconds
0.1 58 seconds
10.6 hours
60.5 min
3 x 10 second
3.1 minutes
Stable
4.20
-
4.5
5.42
5.68
6.28
6.77
-
6.05
8.77
-
-
Uranium Series
Uranium
Thorium
Protactinium
Uranium
Thorium
Radium
Radon
Polonium
Lead
Bismuth
Polonium
Thallium
Lead
Bismuth
Polonium
Lead
92u238 (un
90Th234(UX1)
91Pa234 (UX2)
7 Id
«*<*•• 41 ill \
\\Jlll
90Tha30.«o)
0.226
88Ra
Rn222
86Bn
MPo218 (RaA)
82Pb214 (RaB)
83BI214 (R.C)
MPo214 (RaC'»
81TI210 (RaC")
82Pb210 (RaD)
83Bi210(RaE)
84Po210 (RaFl
gjPb206 (RaG)
4.55 x 109 years
24.1 days
1.14 minutes
2.69 x 105 yean
8.22 x104 years
1600 years
3.825 days
3.05 minutes
26.8 minutes
19.7 minutes
1.5 xlO"4 second
1.32 minutes
22.2 years
4.97 days
139 days
Stable
4.21
-
-
4.75
4.66
4.79
GAB
5.99
-
5.50
7.68
-
-
-
6.30
—
-
0.053
1.55
-
-
-
/3
0.36
2.20
-
1.82
-
-
-
-
r
-
-
-
-
T
-
2.62
-

-
0.13
2.32
-
-
-
-
ft
0.65
3.1 B
-
1.80
0.025
1.17
-
—
-
0.09
0.80
-
If
0.19
-
-
r
1.8
-
-
0.047
-
r
-
               125

-------
TABLE 111-26. URANIUM MILLING PROCESSES
         (a) 1967 Uranium Mills by Process
MILL
American Metal Climax
Anaconda
Atlas (Acid)
Atlas (Alkaline)
Cotter
Federal/American
Foote Mineral
United Nuclear/Homestake
Kerr-McGee
Mines Development
Petrotomics
Susquehanna Western
UCC Uravan
UCC Gas Hills
Utah Construction & Mining
Western Nuclear
LEACH
Acid
Acid
Acid
Alkaline
Alkaline
Acid
Acid
Alkaline
Acid
Acid
Acid
Acid
Acid
Acid
Acid
Acid
CONCENTRATION
SX
RIP, IX
SX
RIP. IX
-
RIP, IX&SX
SX
-
SX
RIP, IX & SX
SX
SX
IX
RIP, IX
IX&SX
RIP, IX & SX
PRECIPITATION
H2°2
Lime/MgO
Ammonia
Ammonia
NaOH
Ammonia
MgO
NaOH
Ammonia
Ammonia
MgO
NaOH
Ammonia
Ammonia
Ammonia
Ammonia
VANADIUM
Salt roast
-
SX
-
-
-
SX
-
-
Na2 CO3 roast
—
-
IX
-
-
-
       (b) Process by Number of Operations (1967)
ORE TREATMENT
Salt Roasting
Flotation
Pre-leach Density Control
LEACHING
Acid
Alkaline
2-Stage

LIQUID-SOLID SEPARATION
Countercurrent Decantation
Staged Filtration
Sand/Slime Separation
RESIN ION EXCHANGE (IX)
Basket Resin In
Pulp (Acid)
Basket RIP (Alkaline)
Continuous RIP
Fix Bed IX
Moving Bed IX

1
2
3

3
3
4


9
3
7


2
1
3
1
1
SOLVENT EXTRACTION (SX)
Amine
Alkyl Phosphoric
Eluex
PRECIPITATION
Lime/MgO
MgO
Caustic Soda (NaOH)
Ammonia (NH4OH)
Peroxide (H2O2>
VANADIUM RECOVERY









7
3
4

1
3
3
8
1
6








            SOURCE: REFERENCE 23
                  126

-------
Figure 111-29 GENERALIZED FLOW DIAGRAM FOR PRODUCTION OF URANIUM,
          VANADIUM, AND RADIUM
                  MINING
                   I
              ORE TREATMENT
                   I
                 LEACHING
                   I
                LIQUID/SOLID
                SEPARATION
                    I
            I
  ION EXCHANGE
       I
    PATH I
       I
           SOLVENT EXTRACTION
                    I
                    I
                PATH IE
                _J
               PRECIPITATION
      TO
STOCKPILE
  URANIUM
CONCENTRATE
                           •—j

                              I
VANADIUM
BYPRODUCT
RECOVERY
TO
STOCKPILE
                          127

-------
Detailed  flow  charts  of  exemplary mills are presented  in
Section VII.

Production Data.    Recent  uranium  production  data   (U. S.
Atomic Energy Commission, 1974) show that uranium  production
has  been  relatively stable  (between 12,600-14,000 ton U3O8
per year) since 1968.

Table III-27 shows uranium production for  the  period  1968
through  1972,  expressed  in terms of both ore movement and
U3O8. production and reserves.  The reserves are estimated  to
be recoverable at the traditional AEC stockpiling  price   of
$18/kg   ($8/lb); with inflation, this price figure should  be
revised upward.  Reserves were seen to  be  increasing  even
before  this adjustment.  They are presumably expanding even
faster when measured in terms of the energy to be  extracted
from  uranium.  Additional uranium  (and its derivative, plu-
tonium) will become  available  if  and  when  environmental
problems  of  fuel recycling are resolved—particuarly, when
breeder reactors become practical.  The  latter  step  alone
should increase the economic  ($18/kg) reserves, estimated  to
last for about  20 years, to about 500 years.

Vanadium  production.  Table  111-28,  is  treated somewhat
differently, since vandium is often an unwanted byproduct  of
uranium mining and is  only  concentrated   (recovered)  when
needed.  Value  of the product fluctuates with demand, unlike
uranium,  as  indicated  in  the table.  World production  is
also shown, to  indicate that U.S. production presents a fair
fraction of the world supply.  The applications of vanadium
are illustrated in Table 111-29.

Radium  is  traded  from  foreign sources, but not mined,  in
quantities of about 40 grams  (or curies)  (1.4 ounce),  at   a
price of about  $20,000/gram  ($567,000/ounce) each  year.  The
high  price  is  set  by the historically determined cost  of
refining and not by current demand.  Reserves of   radium   in
uranium  tailings  are plentiful at this price,  it has been
estimated  that  concentration  of  radium  to  prevent  its
discharge to uranium tailings would approximately  double the
cost of uranium concentrate  (Reference 24).

Thorium  production  in  the U.S. during  1968 was  100 metric
tons  (110 short tons) as was demand, mostly for the chemical
ai-i<3 A! en~4--rnni f  11R#»r ^***^»tmb^^ Wt A
and electronic uses.  The U.S. imported  210 metric tons  (231
short tons)  to increase privately held stocks  from  560  to
770  metric  tons   (616  to  847  short  tons) .  The General
Services Administration also held a stockpile of 1465 metric
tons (1612 short tons) which was intended to contain only 32
                             128

-------
                     TABLE 111-27. URANIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
1973
ORE MOVEMENT
1000
METRIC TONS
5.861
5,367
5,749
5,708
5,834
6,152
1000
SHORT TONS
6,461
5,916
6,337
6,292
6,431
6,781
U3O8 PRODUCTION
1000
METRIC TONS
11.244
10.554
11.732
11.157
11.727
12.032
1000
SHORT TONS
12.394
11.634
12.932
12.298
12.927
13.263
U3Og RESERVES*
1000
METRIC TONS
146
185
224
248
248
251
1000
SHORT TONS
161
204
247
273
273
277
•At $18,000 per metric ton ($16,340 per short ton).
                    TABLE 111-28. VANADIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
U.S. V2Og
PRODUCTION
1000
METRIC
TONS
5,590
5,369
5,085
4,812
4,771
1000
SHORT
TONS
6,192
5,918
5,605
5,304
5.259
%OF
WORLD
46
31
27
28
26
WORLD V2O5
PRODUCTION
1000
METRIC
TONS
12.119
16,892
18.337
16.883
18,135
1000
SHORT
TONS
13,359
18,620
20,213
18.610
19,990
V20g VALUE
PER
METRIC
TON
$3,910
$5,190
$7,216
$7,887
$6,941
PER
SHORT
TON
$3,547
$4,708
$6,546
$7,155
$6,297
                        TABLE III-29. VANADIUM USE
CATEGORY
Ferrowanadium
Vanadium Oxide
Ammonium Metavanadate
Vanadium Metal/alloys
1971
METRIC
TONS
3,792
130
32
412
SHORT
TONS
4,180
143
35
454
%
87
3
1
9
1972
METRIC
TONS
4,084
172
43
453
SHORT
TONS
4,502
190
47
499
%
86
4
1
9
                                  129

-------
metric tons (35 short tons)—i.e., was in  surplus  by  1433
metric tons (1577 short tons).

Metal Ores, Not Elsewhere Classified

This  category  includes ores of metals which vary widely in
their mode of occurrence, extraction methods, and nature  of
associated  effluents.   The discussion of metals ores under
this category  which  follows  treats  antimony,  beryllium,
platinum,  tin,  titanium,  rare-earth,  and zirconium ores.
Thorium ores  (monazite) have been previously discussed under
the  Uranium,  Radium,  Vanadium  category  because  of  the
similarity of their extractive methods and radioactivity.

Antimony Ores

The  antimony ore mining and milling industry is defined for
this document as that segment of industry  involved  in  the
mining  and/or  milling of ore for the primary or byproduct/
coproduct recovery of antimony.  In the United States,  this
industry  is concentrated in two states:  Idaho and Montana.
A small amount of antimony also comes from a mine in Nevada.
Table 111-30 summarizes the sources and amounts of  antimony
production  for 1968 through 1972.  The decrease in domestic
production during 1972 indicated in Table 111-30 was largely
due to a fire which forced the major byproduct  producer  of
antimony to close in May of that year.

Antimony  is  recovered from antimony ore and as a byproduct
from silver and lead concentrates.

Only slightly more than 13 percent of the antimony  produced
in 1972 was recovered from ore being mined primarily for its
antimony  content.   Nearly  all  of  this production can be
attributed to a single operation  which  is  using  a  froth
flotation  process  to  concentrate stibnite (Sb2S3) (Fioure
111-30) .                                        	

The bulk of domestic production of antimony is recovered  as
a byproduct of silver mining operations in the Coeur d'Alene
district  of  Idaho.  Antimony is present in the silver-con-
taining mineral tetrahedrite and is  recovered  from  tetra-
hedrite  concentrates in an electrolytic antimony extraction
plant owned  and  operated  by  one  of  the  silver  mining
companies  in the Coeur d'Alene district.  Mills are usually
penalized for the antimony content  in  their  concentrates.
Therefore,  the  removal  of  antimony from the tetrahedrite
concentrates  not  only  increases  their  value,  but   the
antimony  itself  then  becomes a marketable item.  In 1972,
                            130

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       TABLE 111-30. PRODUCTION OF ANTIMONY FROM DOMESTIC SOURCES
YEAR
1968
1969
1970
1971
1972
ANTIMONY CONCENTRATE
METRIC TONS
4,774
5,176
6.060
4,282
1.879
SHORT TONS
5,263
5,707
6.681
4,721
2,072
ANTIMONY'
METRIC TONS
776
851
1,025
930
444
SHORT TONS
856
938
1.130
1,025
489
ANTIMONIAL LEADt
(ANTIMONY CONTENT)
METRIC TONS
1,179
1,065
542
751
468
SHORT TONS
1,300
1,174
598
828
516
•Includes production from antimony ores and concentrates and byproduct recovery from silver concentrates.
tByproduct produced at lead refineries in the United States.
                                     131

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Figure 111-30. BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
                     MINING
                      ORE
                      i
                    CRUSHING
                   GRINDING
                      I
                 CLASSIFICATION
       i
    ROUGHER
    FLOTATION
 •TAILS-
        I
SCAVENGER
FLOTATION
     FROTH
        I
 CLEANER
FLOTATION
  FROTH
    I
    •TAILS
                         FROTH
TO
WASTE
                    FILTER
                                      FILTRATE
                   THICKENER
                      I
                                       WASTE
                     FINAL
                 CONCENTRATE
                      F
                  TO SHIPPING
                             132

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the price for antimony was $1.25  per  kilogram   ($0.57  per
pound) .

Antimony is also contained in lead concentrates and is ulti-
mately recovered as a byproduct at lead smelters — usually,
as  antimonial  lead.   This  source  of antimony represents
about 30 to 50 percent  of  domestic  production  in  recent
years.

Beryllium Ores

The beryllium ore mining and milling industry is defined for
this  document  as  that segment of industry involved in the
mining and/or milling of ore for the primary  or  byproduct/
coproduct recovery of beryllium.  Domestic beryllium produc-
tion  data  are  withheld  to  avoid  disclosing  individual
company  confidential  data.   During   1972,   some   beryl
(66^12(51.60.18))  was produced in Colorado and South Dakota.
The  largest  domestic  source  of  beryllium   ore   is   a
bertrandite  (Bej*Si207   (OH) 2)  mine  in  the  Spor Mountain
district of Utah.  Domestic  beryl  prices  were  negotiated
between  producers  and  buyers  and  were not quoted in the
trade press.

Mining and milling techniques for beryl are unsophisticated.
Some pegmatite deposits are mined on a small scale--usually,
by crude opencut methods.  Mining is begun  on  an  outcrop,
where  the  minerals  of value can readily be seen, and cuts
are made or pits are sunk by drilling and blasting the rock.
The blasted rock is hand-cobbed, by which procedure as  much
barren  rock  as practicable is broken off with hand hammers
to recover the beryl.  Beryl and the minerals it is commonly
associated with have densities so nearly the same that it is
difficult   to   separate   beryl   by   mechanical   means.
Consequently, beryl is recovered by hand cobbing.

A  sulfuric  acid  leach  process  is  employed  to  recover
beryllium from the Spor Mountain bertrandrite.   This  is  a
proprietary   process,  however,  and  further  details  are
withheld.  No effluent results from this operation.

Platinum-Group Metal Ores

The platinum-group metal ore mining and milling industry  is
defined  for  this  document  as  those operations which are
involved in the mining and/or milling of ore for the primary
or  byproduct/coproduct  recovery  of  platinum,  palladium,
iridium,  osmium,  rhodium, and ruthenium.  These metals are
characterized by their superior resistance to corrosion  and
                            133

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oxidation.   The  industrial  applications  for platinum and
palladium are diverse,  and  the  metals  are  used  in  the
production  of  high-octane  fuels,  catalysts, vitamins and
drugs, and electrical components.   Domestic  production  of
platinum-group  metals  is  principally  as  a  byproduct of
copper smelting, with production also from platinum placers.
Table III-31 lists annual U.S. mine production and value for
the period 1968 through 1972.

The geologic occurrence  of  the  platinum-group  metals  as
lodes or placers dictates that copper, nickel, gold, silver,
and  chromium will be either byproducts or coproducts in the
recovery of platinum  metals,  and  that  platinum  will  be
largely  a  byproduct.  With the exception of occurrences in
the  Stillwater  Complex,  Montana,  and  production  as   a
byproduct  of  copper  smelting,  virtually  all  the  known
platinum-group minerals  in  the  United  States  come  from
placers.    Platinum   placers   consist  of  unconsolidated
alluvial deposts  in  present  or  ancient  stream  valleys,
terraces,  beaches,  deltas, and glaciofluvial outwash.  The
other domestic source of  platinum  is  as  a  byproduct  of
refining  copper from porphyry and other copper deposits and
from lode and placer gold deposits, although  the  grade  is
extremely low.

Platinum-group  metals  occur  in  many  placers  within the
United States.  Minor amounts have been recovered from  gold
placers  in  California, Oregon, Washington, Montana, Idaho,
and Alaska, but significant amounts have been produced  only
from  the  placers  of  the  Goodnews  Bay District, Alaska.
Production over the past several years  from  this  district
has   remained   fairly  constant,  although  domestic  mine
production declined 5 percent in quantity and 7  percent  in
value in 1972  (Reference 2).

Beneficiation of Ores.

The  mining  and  processing techniques for recovering crude
platinum from placers in the U.S. are similar to those  used
for  recovering gold.  The bulk of the crude placer platinum
is recovered by large-scale bucket-line dredging, but small-
scale hand methods are also used in Columbia, Ethiopia,  and
(probably)  the  U.S.S.R.   A  flow  diagram  for  a typical
dredging operation is presented as Figure 111-31.

In the Republic of South Africa, milling  and  beneficiation
of  platinum-bearing  nickel  ores  consist  essentially  of
gravity concentration, flotation, and smelting to produce  a
high-grade  table  concentrate  called "metallic" for direct
                             134

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TABLE 111-31. DOMESTIC PLATINUM-GROUP MINE PRODUCTION AND VALUE
YEAR
1968
1969
1970
1971
1972
MINE PRODUCTION
KILOGRAMS
460.1
671.4
538.6
560.8
532.2
TROY OUNCES
14,793
21,586
17,316
18,029
17,112
VALUE
$1,500,603
$2,094.607
$1,429,521
$1,359,675
$1,267,298
       SOURCE: REFERENCE 2
                           135

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Figure 111-31. GRAVITY CONCENTRATION OF PLATINUM-GROUP METALS
                DREDGE
              (SCREENING.
              JIGGING, AND
               TABLING)
               TABLING
              MAGNETIC
              SEPARATION
CHROMITE/
MAGNETITE
               DRYING
              SCREENING
                SIZING
             Ilil
               BLOWER
           90% CONCENTRATE
       (PLATINUM GROUP AND GOLD)

             TO SHIPPING
               TO
               WASTE
TO
SHIPPING
                                             WASTE
                        136

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 chemical refining  and a nickel-copper matte  for  subseauent
 smelting and refining.

 Byproduct platinum-group metals from gold or copper ores are
 sometimes  refined  ty  electrolysis and chemical means,  in
 the sudbury District of Canada, sulfide ore is processed  bv
 magnetic  flotation  techniques  to  yield  concentrates  of
 copper  and   nickel   sulfides.    The   nickel   flotation
 concentrate  is roasted with a flux and melted into a matte,
 which is cast into anodes for  electrolytic  refining,  from
 which the precious metal concentrate is recovered.

 In  the  U.S.,  the  major  part  of  output  of platinum is
 recovered as a byproduct of copper refining in Maryland, New
 arnnn7' ^"f8' utan< ?nd  Washington.   Byproduct  platinum-
 group  metals from gold or copper ores are sometimes refined
 by electrolysis and by chemical means.   Metal  recovery  in
 refining is over 99 percent.

 Rare-Earth ores

 The  rare-earth  minerals  mining  and  milling  industry is
 defined for  this  document  as  that  segment  of  industry
 engaged  in the mining and/or milling of rare-earth  minerals
 for their  primary  or  byproduct/coproduct  recovery.    The
 rare-earth  elements,   sometimes  known  as the lanthanides,
 consist of  the series  of 15 chemically similar  elements  with
 atomic  numbers 57 through 71.   Yttrium,  with  atomic  number
 J9, is  often included  in the group,  because its  properties
 are similar,   and  it  more  often  than  not   occurs    in
 association with  the  lanthanides.    The  principal mineral
 sources of  rare-earth  metals are  bastnaesite  (CeFC03)   and
 monazite  (Ce,   La,  Th,   Y)POH.    The  bulk of the  domestic
 production  of   rare-earth  metals  is  from a bastnaesite
 deposit  in  Southern  California   which is  also the world's
 largest known   single   commercial  source   of    rare-earth
 elements.   In 1972, approximately  10,703 metric tons (11,800
 short   tons) of  rare-earth oxides  were obtained  in flotation
 concentrate from 207,239 metric  tons  (approximately  228,488
 short tons) of bastnaesite ore mined and milled  (Reference 2
 ).   Monazite  is  domestically  recovered as a byproduct of
 titanium mining and  milling  operations  in  Georgia  and
 Florida.   A  company  which  recently began a heavy-mineral
 (principally,  titanium)  sand  operation  in   Florida   is
 expected to produce over 118 metric tons  (130 short tons* of
 byproduct monazite annually.                            '

At  the  Southern California operation, bastnaesite is mined
by open-pit methods.  The ore, containing 7  to  10  percent
                            137

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rare-earth oxides (REO), is upgraded by flotation techniques
to a mineral concentrate containing 63 percent REO.  Calcite
is removed by leaching with 10 percent hydrochloric acid and
countercurrent   decantation.    The   bastnaesite   is  not
dissolved by this treatment, and the concentrate is  further
upgraded to 72 percent REO.  Finally, the leached product is
usually  roasted  to  remove  the  carbon  dioxide  from the
carbonate, resulting in a product with over 90 percent REO.

Monzazite  is  recovered  from  heavy-mineral  sands   mined
primarily  for  their  titanium  content.   Beneficiation of
monazite is by the wet-gravity, electrostatic, and  magnetic
techniques   discussed  in  the  titanium  portion  of  this
document.  Monazite, an important source of thorium, is also
discussed under SIC 1094  (Uranium,  Radium,  and  Vanadium).
Extraction of the thorium is largely by chemical techniques.

Tin Ores

The  tin  mining  and  milling  industry is defined for this
document as that segment of industry engaged in  the  mining
and/or  milling  of ore for the byproduct/coproduct recovery
of tin.

There are presently no known  exploitable  tin  deposits  of
economic  grade  or  size in the United States.  Most of the
domestic tin production in 1972, less than 102  metric  tons
(112  short  tons),  came  from  Colorado  as a byproduct of
molybdenum mining.  In addition, some  tin  concentrate  was
produced at dredging operations and as a byproduct of placer
gold  mining operations in Alaska.  A small placer operation
began production in New Mexico in  June  1973.   Feasability
studies  continue  for  mining  and milling facilities for a
4,065-metric-ton-per-day  (4,472-short-ton-per-day)  open-pit
fluorite  tin/tungsten/beryllium  mine  in  Alaska's  Seward
Peninsula which  is  to  open  by  1976.   Reserves  at  the
prospect area represent at least a 20-year supply.  As tech-
nological improvements in beneficiation are made and demands
for tin increase, large deposits considered only submarginal
resources,  in  which  tin  in  only one of several valuable
commodities, are expected to be brought into production.

In general, crude cassiterite  concentrate from placer mining
is   upgraded   by   washing,   tabling,and   magnetic    or
electrostatic  separation.   Tin  ore  from lode deposits is
concentrated  by  gravity   methods   involving    screening,
classification,  jigging,  and  tabling.  The concentrate is
usually a lower grade  than  placer  concentrate,  owing  to
associated  sulfide  minerals.   The  sulfide  minerals  are
                             138

-------
removed by flotation or magnetic separation, with or without
magnetic roasting.  The majority of tin  production  in  the
United states is the result of beneficiation as a byproduct.
Cassiterite concentrate recovery takes place after flotation
of  molybdenum  ore  by magnetic separation of the dewatered
and  dried   tailings.    Despite   considerable   research,
successful  flotation  of  tin ore has never been completely
achieved.

Titanium Ores

The titanium ore mining and milling industry is defined  for
this  document  as  that  segment of industry engaged in the
mining and/or milling of titanium ore  for  its  primary  or
byproduct/   coproduct   recovery.   The  principal  mineral
sources of titanium are ilmenite (FeTiO^) and rutile (Ti02).
The United states is a major source of ilmenite but  not  of
rutile.  Since 1972, however, a new operation in Florida has
been  producing  (5,964 metric tons, or 6,575 short tons, in
1974)  rutile.  About 85 percent of the ilmenite produced  in
the  United  States  during  1972 came from two mines in New
York and Florida.  The remainder of the production came from
New Jersey, Georgia, and a second operation in  Florida.   A
plant  with  a  planned  production  of  168,000 metric tons
(185,000 short tons) per year opened in  New  Jersey  during
1973.    This  plant  and another which opened during 1972 in
Florida are not yet at full production  capability  but  are
expected   to   contribute  significantly  to  the  domestic
production of titanium in the future.   Domestic  production
data are presented in Table 111-32.

Two  types of deposits contain titanium minerals of economic
importance:  rock and sand deposits.  The ilmenite from rock
deposits and some sand deposits commonly contains 35  to  55
percent  Ti02;  however,  some  sand  deposits yield altered
ilmenite (leucoxene) containing 60 percent or more Ti02_,  as
well as rutile containing 90 percent or more Ti02..

The  method  of  mining  and beneficiating titanium minerals
depends upon whether the ore to be mined is a sand  or  rock
deposit.   Sand  deposits occurring in Florida, Georgia, and
New Jersey contain 1 to 5 percent TiO£ and  are  mined  with
floating suction or bucket-line dredges handling up to 1,088
metric  tons  (1,200  short tons)  of material per hour.   The
sand is treated by wet gravity methods using spirals, cones,
sluices, or jigs to produce  a  bulk,  mixed,  heavy-mineral
concentrate.  As many as five individual marketable minerals
are   then   separated   from  the  bulk  concentrate  by  a
combination of dry separation techniques using magnetic  and
                            139

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  TABLE 111-32. PRODUCTION AND MINE SHIPMENTS OF TITANIUM
             CONCENTRATES FROM DOMESTIC ORES IN THE U.S.
YEAR
1968
1969
1970
1971
1972
PRODUCTION*
METRIC TONS
887,508
884,641
787,235
619,549
618,251
SHORT TONS
978,509
931,247
867,955
683,075
681.644
SHIPMENTS*
METRIC TONS
870,827
809,981
835,314
647,244
661,591
SHORT TONS
960,118
893,034
920,964
713,610
729,428
•Includes a mixed product containing rutile, leucoxene, and altered ilmenite.

SOURCE: REFERENCE 2
                        140

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electrostatic    (high-tension)   separators,   sometimes  in
conjunction  with  dry   and   wet   gravity   concentrating
equipment.

High-tension   (HT)  electrostatic separators are employed to
separate the titanium minerals from the  silicate  minerals.
In  this  type  of  separation,  the minerals are fed onto a
high- speed spinning rotor, and a heavy corona  (glow  given
off  by  high- voltage charge) discharge is aimed toward the
minerals at the point where they would  normally  leave  the
rotor.    The   minerals   of   relatively  poor  electrical
conductance are pinned to the  rotor  by  the  high  surface
charge  they  receive  on  passing through the high- voltage
corona.  The minerals of relatively high conductivity do not
as readily hold this surface charge and so leave  the  rotor
in  their normal trajectory.  Titanium minerals are the only
ones present of relatively high electrical conductivity  and
are,  therefore,  thrown  off  the rotor.  The silicates are
pinned to the rotor and are removed by a fixed brush.

Titanium minerals undergo final separation  in  induced-roll
magnetic  separators  to  produce three products:  ilmenite,
leucoxine, and rutile.  The separation of these minerals  is
based  on  their  relative  magnetic  propertities which, in
turn, are based on their relative  iron  content:   ilmenite
has  37  to  65 percent iron, leucoxine has 30 to HO percent
iron, and rutile has H to 10 percent iron.

Tailings from the HT separators (nonconductors)  may  contain
zircon  and  monazite  (a  rare-earth mineral) .   These heavy
minerals  are  separated  from   the   other   nonconductors
(silicates)  by various wet gravity methods (i.e., spirals or
tables).   The  zircon  (nonmagnetic) and monazite (slightly
magnetic)  are separated from  one  another  in  induced-roll
magnetic separators.

Beneficiation  of titanium minerals from beach-sand deposits
is illustrated in Figure 111-32.

Ilmenite is also currently mined from a rock deposit in  New
York  by  conventional  open-pit  methods.   This  ilmenite/
magnetite ore, averaging 18 percent  TiO^r  is  crushed  and
ground to a small particle size.  The ilmenite and magnetite
fractions   are  separated  in  a  magnetic  separator,  the
magnetite being  more  magnetic  due  to  its  greater  iron
content.   The  ilmenite  sands  are  further  upgraded in a
flotation circuit.  Beneficiation of titanium  from  a  rock
deposit is illustrated in Figure 111-33.

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Figure 111-32. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS
ORE FED
FROM DREDGE
1 '
VIBRATING
SCREENS
*
SPIRALS OR LAMINA
FLOWS IROUGHERS AND CL
TO
POND


bANbHSI ^^ lAlLlNliS — ^ WASTE
WET MILL
DRY MILL 1
SCRUBBER PLANT
i
DRIER
i
ELECTROSTATIC
SEPARATORS

^ SODIUM
	 , .^ Hvonoxine


SPIR*LS AND/OR MAGNETIC
TABLES SEPARATOR
t
MAGNETIC .
SEPARATOR _J
r 	 NONMAGNETICS 	 * 	 MAGNETICS 	 1
' 	 1 •'lJ RUTILE | j ILMENITE
• 	 MAGNETICS 	 * 	 NONMAGNETICS 	 1
MONAZITE | ( ZIRCON ]
1 1
\ t 1
TO TO T
SHIPPING SHIPPING SHIP

\>
0 TO
PING SHIPPING
                     142

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Figure 111-33. BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
                              MINING
                               ORE
                              JL
                            CRUSHING
                            GRINDING
                               I
                          CLASSIFICATION
                               I
                            MAGNETIC
                           SEPARATION
          ±
                 •MAGNETICS-
. NONMAGNETICS
       MAGNETITE
               ILMENITE
             AND GANGUE
      DEWATERER
                                                    i
              FLOTATION
               CIRCUIT
                                 TAILINGS
              THICKENER
                                   TO
                                  WASTE
                                                    i
               FILTER
                                                    ±
                                                   DRIER
                                                CONCENTRATE
                                                TO SHIPPING
                            143

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Zirconium Ore

The zirconium ore mining and milling industry is defined for
this  document  as  that  segment of industry engaged in the
mining and/or milling of zirconium or  for  its  primary  or
byproduct/coproduct recovery.

The   principal   mineral  source  of  zirconium  is  zircon
(ZrSi04), which is recovered as a byproduct in the mining of
titanium minerals from ancient  beach-sand  deposits,  which
are  mined  by floating suction or bucket-line dredges.  The
sand is treated by wet gravity methods to produce  a  heavy-
mineral  concentrate.  This concentrate contains a number of
minerals (zircon, ilmenite, rutiler and monazite) which  are
separated from one another by a combination of electrostatic
and   magnetic  separation  techniques,  sometimes  used  in
conjunction  with  wet  gravity  methods.   (Refer  to   the
titanium  section of this document.)  Domestic production of
zircon is currently from three operations:  two  in  Florida
and  one  in Georgia.  The combined zircon capacity of these
three plants is estimated to be about  113,400  metric  tons
(125,000  short  tons).   The  price  of  zircon in 1972 was
$59.50 to $60.50 per metric ton ($54.00 to $55.00 per  short
ton) .
                             144

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                         SECTION IV
                  INDUSTRY CATEGORIZATION

INTRODUCTION
In  the  development of effluent limitations and recommended
standards of performance for new  sources  in  a  particular
industry,  consideration  should  be  given  to  whether the
industry can be treated as a whole in the  establishment  of
uniform  and equitable guidelines for the entire industry or
whether there are sufficient differences within the industry
to justify its division into categories.  For the ore mining
and  dressing  industry,  which  contains  nine  major   ore
categories  by SIC code (many of which contain more than one
metal  ore),  many  factors  were  considered  as   possible
justification     for     industry     categorization    and
subcategorization as follows:
         (1)   Designation as a mine or mill;
         (2)   Type of mine;
         (3)   Type of processing  (beneficiation, extraction
              process);
         (4)   Mineralogy of the ore;
         (5)   End product  (type of product produced);
         (6)   Climate, rainfall, and location;
         (7)   Production and size;
         (8)   Reagent use;
         (9)   Wastes or treatability of wastes generated;
         (10) Water use or water balance;
         (11) Treatment technologies employed;
         (12) General geologic setting;
         (13) Topography;
              Facility age;
                             145

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         (15)  Land availability.


Because of their frequent use in this document, the  defini-
tions  of  a mine and mill are included here for purposes of
recommending  subcategorization  and  effluent   limitations
guidelines and standards:

Mine

"A  mine  is  an  area  of  land  upon  which or under which
minerals or metal ores are extracted from  natural  deposits
in  the  earth by any means or methods.  A mine includes the
total area upon which such activities occur  or  where  such
activities  disturb  the natural land surface.  A mine shall
also include land  affected  by  such  ancillary  operations
which  disturb  the  natural  land surface, and any adjacent
land the use of which is incidental to any such  activities;
all  lands  affected by the construction of new roads or the
improvement or use or existing roads to gain access  to  the
site  of  such  activities  and for haulage and excavations,
workings, impoundments, dams, ventilation  shafts,  drainage
tunnels,   entryways,   refuse   banks,  dumps,  stockpiles,
overburden piles, spoil banks, culm banks,  tailings,  holes
or depressions, repair areas, storage areas, and other areas
upon  which  are  sited  structures,  facilities,  or  other
property or materials on  the  surface,  resulting  from  or
incident to such activities."

Mill

"A  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.   This  includes  such operations as crushing,
grinding, washing, drying, sintering, briquetting, pelletiz-
ing, nodulizing, leaching, and/or concentration  by  gravity
separation,  magnetic  separation, flotation or other means.
A mill includes  all  ancillary  operations  and  structures
necessary  for the cleaning, concentrating or other process-
ing of the mineral or metal  ore  such  as  ore  and  gangue
storage areas, and loading facilities."

Examination  of  the  metal  ore  categories covered in this
document indicates that ores of 23 separate metals (counting
the rare earths as a single  metal)  are  represented.   Two
materials  are  treated in two places in this document:  (1)
vanadium ore is considered as a source of ferroalloy  metals
 (SIC  1061)  and  also  in conjunction with uranium/vanadium
                             146

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extraction under NRC  licensing  surveillance  (SIC 109U);  and
 (2)  monazite,  listed as a SIC 1099 mineral because  it is a
source of rare-earth  elements,  also  serves as an  ore of  a
radioactive  material   (thorium)  and,  therefore,  is  also
treated in SIC 1094.

The discussion that follows is   organized  into  five major
areas which illustrate the procedures and final selection of
subcategories   which  have  been  made  as  part  of these
recommendations:

    (1)  The  factors  considered    in   general   for   all
         categories.    (Rationale for selection or rejection
         of each as a pertinent criterion  for  the  entire
         industry is  included.)

    (2)  The factors which determined the  subcategorization
         within each  specific ore category.

    (3)  The procedures which   led   to  the  designation  of
         tentative  and,  then,  final  subcategories within
         each SIC code group.

    (4)  The final recommended  subcategories  for  each  ore
         category.

    (5)  Important factors and  particular problems pertinent
         to subcategorization in each major category.

FACTORS INFLUENCING SELECTION OF SUECATEGORIES  IN  ALL  ORE
CATEGORIES

The  first  categorization  step  was  to  examine  the  ore
categories   and   determine    the    factors    influencing
subcategorization   for  the  industry  as  a  whole.   This
examination  evolved  a  list   of  15   factors   considered
important  in subcategorization of the industry segments (as
tabulated above).   The discussion  which  follows  describes
the  factors  considered  in  general for all categories and
subcategories.

Designation as a Mine or Mill

It is often  desirable  to  consider  mine  water  and  mill
process  water separately.  There are many mining operations
which do not have an associated mill or in which many  mines
deliver ore to a single mill located some distance away.  In
many  instances,   it  is advantageous to separate mine water
from mill process  wastewater   because  of  differing  water
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quality, flow rate or treatability.  Levels of pollutants in
mine  waters  are generally lower or less complex than those
in mill process wastewaters.  Mine water contact with finely
divided ores, (especially oxidized ores) is minimal and mine
water is not exposed to the suite of process water  reagents
often  added in milling,  wastewater volume reduction from a
mine is seldom a viable option  whereas  the  technology  is
available  to  eliminate  all  discharge  from  many milling
operations.

While  it  is  generally  more  efficient  to   treat   mine
wastewater  and  mill  wastewater separately, there are some
situations in which combining the mine wastewater  and  mill
process  wastewater  cause  a co-precipitation of pollutants
with their resultant discharge being of higher quality  than
either  of  the  individual  treated  discharges.   In  some
instances, use of the mine wastewater as mill process  water
will also result in an improved quality of discharge because
of  the  interactions  of the chemicals added to the process
water with the pollutants in the mine water.

Type of Mine

The choice of mining method is determined by the ore  grade,
size, configuration, depth, and associated overburden of the
orebody   to  be  exploited  rather  than  by  the  chemical
characteristics or mineralogy of the deposit.   Because  the
general  geology  is  the determining factor in selection of
the mining method, and because  no  significant  differences
resulted   from   application   of   control  and  treatment
technologies  for  mine  waters  from  either  open  pit  or
underground  mines,  designation of the type of mine was not
selected as a suitable basis for  general  subcategorization
in the industry.

In  addition,  it  is  recognized  that, in deep underground
mines, it is common practice to  backfill  stopes  with  the
coarse   sand   fraction  of  the  mill  tailings,  or  with
cement/sand mixtures.  This practice is designed to  provide
the  ground support needed to reduce the possibility of rock
bursts or caving during subsequent stoping  operations.   In
those  situations where this practice is employed, the sands
are normally  transported  underground  by  sluicing.   Mill
wastewater  is  a  convenient  source of sluice water and is
usually employed for this  purpose.   This  mill  wastewater
undoubtedly  mingles with any mine seepage which is present,
thereby  reducing  the  quality  of  the   mine   discharge.
However,  this  factor  alone  does  not enable further sub-
categorization,  since  it   introduces   no   significantly
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different    treatability   considerations.    control   and
treatment technologies which are effective in this situation
are essentially identical  to  the  technology  employed  in
treating mine waters which result from seepage alone.

Type of Processing  (Beneficiation, Extraction Process)

The  processing  or  beneficiation of ores in the ore mining
and dressing industry varies  from  crude  hand  methods  to
gravity  separation  methods, froth flotation with extensive
reagent  use,  chemical  extraction,  and   hydrometallurgy.
Purely  physical processing using water provides the minimal
pollution potential consistent with recovery of values  from
an  ore.   All  mills  falling in this group are expected to
share the same major  pollution  problem—namely,  suspended
solids generated either from washing, dredging, crushing, or
grinding.   The  exposure to water of finely divided ore and
gangue also leads to  solution  of  soire  material  but,  in
general,  treatment  required  is  relatively  simple.   The
dissolved material will vary with the ore  being  processed,
but  treatment  is  expected to be essentially similar, with
resultant effluent levels  for  important  parameters  being
nearly identical for many subcategories.

The   practice   of   flotation  significantly  changes  the
character of mill effluent in several ways.  Generally, mill
water pH is altered  or  controlled  to  increase  flotation
efficiency.   This, together with the fact that ore grind is
generally finer than for physical processing, may  have  the
secondary  effect of substantially increasing the solubility
of ore components.  Reagents added to effect  the  flotation
may include major pollutants.  Cyanide, for example, is used
in  several  subcategories.   Although usage is usually low,
its presence in effluent  streams  has  potentially  harmful
effects.   The  added reagents may have secondary effects on
the wastewater as well, such as in the formation of  cyanide
complexes.  The result may be to increase solubility of some
metals and decrease treatment effectiveness.  Some flotation
operations  may  also differ from physical processors in the
extent to which water may be recycled without major  process
changes or serious recovery losses.

Ore  leaching  operations differ substantially from physical
processing and flotation plants in wastewater character  and
treatment  requirements.   The  use  of large quantities (in
relation to ore handled) of  reagents,  and  the  deliberate
solubilization  of ore components characterizes these opera-
tions.  Wide diversity of leaching and  chemical  extraction
processes,  therefore,  affects the character and quantities
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of water quality parameters, as well as  the  treatment  and
control technologies employed.

To  a  large extent, mineralogy and extractive processes are
inextricable,   because   mineralogy    and    mineralogical
variations  are responsible for the variations in processing
technologies.  Both factors influence  the  treatability  of
wastes  and efficiency of removal of pollutants by treatment
and control  technologies.   Therefore,  processing  methods
were  a  major  factor  in  subcategorizing  each  major ore
category.

Mineralogy of the Ore

The mineralogy and host rock present greatly  determine  the
beneficiation  of  ores.   Ore  mineralogy and variations in
mineralogy affect the components present in effluent streams
and thus the treatability of the wastes  and  treatment  and
control technology used.  Some metal ores contain byproducts
and  other  associated  materials,  and  some  do  not.  The
specific beneficiation process adopted  is  based  upon  the
mineralogical  characteristics  of  the  ore; therefore, the
waste characteristics of the mine or mill reflect  both  the
ores  mined  and  the  extraction  process  used.  For these
reasons, ore mineralogy  was  determined  to  be  a  primary
factor affecting subcategorization in all categories.

End Product

The  end product shipped is closely allied to the mineralogy
of the ores exploited; therefore, mineralogy and  processing
were    found   to   be   more   advantageous   methods   of
subcategorization.  Two ores,  vanadium  ores  and  monazite
ores,  are the exceptions treated here which were based upon
considerations of end product or end use.

Climate, Rainfall, and Location

These factors directly influenced subcategorization  consid-
eration  because  of  the  wide diversity of yearly climatic
variations prevalent  in  the  United  states.   Mining  and
associated  milling  operations cannot locate in areas which
have desirable characteristics unlike  many  other  industry
segments.   Therefore,  climate and rainfall variations must
be accommodated or designed for.  Some mills and  mines  are
located  in arid regions of the country, allowing the use of
evaporation  to  aid  in  reduction  of  effluent  discharge
quantity  or attainment of zero discharge,  other facilities
are located in areas of net positive precipitation and  high
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runcff  conditions.   Treatment of large volumes of water by
evaporation in many areas of the  United  States  cannot  be
utilized   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.
Although climate, rainfall, and location were  not  used  as
primary   subcategorization   factors,   they   were   given
consideration  when  determining  treatment  technology  and
effluent limitations (i.e., copper ore industries).

Production and Size

The  variation  of  size and production of operations in the
industry ranges from small hand cobbing operations to  those
mining and processing millions of tons of ore per year.  The
size  or  production of a facility has little to do with the
quality of the water or treatment technology  employed,  but
have  considerable  influence  on the water volume and costs
incurred in attainment of  a  treatment  level  in  specific
cases.   Mines  and  mills processing less than 5,000 metric
tons  (5,512 short tons) of ore per year in  the  ferroalloys
industry (most notably, tungsten) are typically intermittent
in   operation,   have  little  or  no  discharge,  and  are
economically  marginal.   Pollution   potential   for   such
operations  is  relatively  low  due  to the small volume of
material handled if  deliberate  solution  of  ores  is  not
attempted.   Few  of  the  operations  are  covered by NPDES
permits.  Accordingly, size or  production  was  used  in  a
limited  sense  for  subcategorization  in  the  ferroalloys
categories but was not found to be suitable for the industry
as a whole.

Reagent Use

The use of reagents in many segments of the  industry,  such
as  different types of froth flotation separation processes,
can potentially affect the quality of wastewater.   However,
the  types and quantities of reagents used are a function of
the mineralogy of the ore and extraction processes employed.
Reagent  use,  therefore,  was  not  a  suitable  basis  for
subcategorization of any of the metals ores examined in this
program.

Wastes or Treatability of wastes Generated

The  wastes  generated  as  part of mining and beneficiating
metals ores are highly dependent upon  mineralogy  and  pro-
cesses  employed.  This characteristic was not found to be a
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basis  for  general  subcategorizaticn,  however,   it   was
considered in all subcategories.

Water Use and/or Water Balance

Water  use  or water balance is highly dependent upon choice
of process employed or process requirements, routing of mine
waters  to  a  mill  treatment  system  or  discharge,   and
potential for utilization of water for recycle in a process.
Processes  employed  play  a  determining role in mill water
balance  and,  thus,  are  a   more   suitable   basis   for
subcategorization.

Treatment Technologies Employed

Many  mining  and  milling  establishments  currently  use a
single type  of  effluent  treatment  method  today.   While
treatment procedures do vary within the industry, widespread
adoption  of  these  technologies  is  not prevalent.  Since
process and mineralogy control treatability of  wastes  and,
therefore,    treatment   technology   employed,   treatment
technology was not used as a basis for subcategorization.

General Geologic Setting

The general geologic setting determines the type  of  rnine--
i.e.,   underground,   surface  or  open-pit,  placer,  etc.
Significant differences which could be used for subcategori-
zation with respect to geology could not be determined.

Topography

Topographic differences between areas are beyond the control
of mine or mill operators and largely place  constraints  on
treatment  technologies employed, such as tailing pond loca-
tion.  Topographic variations  can  cause  serious  problems
with  respect to rainfall accumulation and runoff from steep
slopes.  Topographic differences were  not  found  to  be  a
practical  basis  on which subcategorization could be based,
but topography is  known  to  influence  the  treatment  and
control  technologies employed and the water flow within the
mine/mill complex.  While not  used  for  subcategorization,
topography  has  been  considered  in  the  determination of
effluent limits for each subcategory.

Facility Age

Many mines and mills  are  currently  operating  which  have
operated  for  the  past  100  years.   in  virtually  every
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operation  involving   extractive   processing,   continuous
modification   of  the  plant  by  installation  of  new  or
replacement equipment results in minimal differences for use
in subcategorization within  a  metal  ore  category.   Many
basic  processes for concentrating ores in the industry have
not changed considerably  (e.g.,  froth  flotation,  gravity
separation,  grinding  and  crushing),  but  improvements in
reagent use  and  continuous  monitoring  and  control  have
resulted  in  improved  recovery or the extraction of values
from lower grade ores.  New and innovative technologies have
resulted in changes of the character of the wastes, but this
is not a function of age of the facilities,  but  rather  of
extractive  metallurgy and process changes.  Virtually every
facility continuously updates in-plant processing  and  flow
schemes,  even  though basic processing may remain the same.
Age of the facility, therefore, is not a useful  factor  for
subcategorization in the industry.

DISCUSSION  OF PRIMARY FACTORS INFLUENCING SUBCATEGORIZATION
BY ORE CATEGORY

The purpose of the effluent  limitation  guidelines  can  be
realized  only by categorizing the industry into the minimum
number of groups  for  which  separate  effluent  limitation
guidelines  and  new  source  performance  standards must be
developed.

This section outlines  and  discusses  briefly  the  factors
which  were  used to determine the subcategories within each
ore category.  A presentation of the procedures  leading  to
the  tentative and then final subcategories, together with a
listing of the final recommended subcategories, is included.
The treatment by ore category also  includes  a  brief  dis-
cussion,   where   applicable,   of  important  factors  and
pertinent problems which affect each category.

Iron Ore

In developing a categorization of the iron ore industry, the
following factors  were  considered  to  be  significant  in
providing a basis for categorization.  These factors include
characteristics  of individual mines, processing plants, and
water uses.

    1.   Type of Mining
         a.   Open-Pit
         b.   Underground

    2.   Type of Processing
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         a.   Physical
         b.   Physical - Chemical

    3.   Mineralogy of the Ore

    4.   General Geologic Setting, Topography, and Climate
         (also Rainfall and Location)

Information for the characterization was developed from pub-
lished  literature,  operating  company  data,   and   other
information sources discussed in Section III.

As a result of the above, the first categorization developed
for  the iron mining and beneficiation industry was based on
whether or not a mine or mill produces  an  effluent.   This
initial   categorization  considered  both  the  mining  and
milling water circuits separately, as  well  as  a  category
where  mines  and  mills were in a closed water system.  The
resulting  tentative  subcategories   which   resulted   are
presented in the listing given below:

    I.   Mine producing effluent - processing plant  with  a
         closed water circuit.

  Ila.   Mine  producing   effluent   -   processing   plant
         producing an effluent - physical processing.

  lib.   Mine  producing   effluent   -   processing   plant
         producing  an  effluent  -  physical  and  chemical
         processing.

  III.   Mine and  processing  plant  with  a  closed  water
         circuit.

Examination of the preliminary subcategorization and further
compilation  of  information  relative  to  iron  mining and
processing methods resulted in a classification of the mines
and mills into the following order by production:

    Open-Pit Mining, Iron Formation, Physical Processing
    Open-Pit Mining, Iron Formation, Physical and Chemical
         Processing
    Open-Pit Mining, Natural Ores, Physical Processing
    Underground Mining, Iron Formation, Physical Processing
    Underground Mining, Iron Formation, Physical and Chemical
         Processing
    Underground Mining, Natural Ores, Physical Processing
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 In preparation for selection of  sites  for  visitation  and
 sampling,  the  operations  were  further  classified on the
 basis of size, relative ager and  whether  they  had  closed
 water systems or produced an effluent from either the mining
 or processing operation:                               -"«.»y

 Operation A
     High tonnage        Older plant (1957)
     Open-pit            Mine produces effluent
     Iron formation      Processing plant has closed water
                              system
     Physical processing

 Operation B
     Medium tonnage      Medium age plant (1965)
     Open-pit            Mine produces effluent
     Iron formation      Processing plant has closed water
                              system
     Physical processing
 Operation C
     Medium tonnage           Older plant (1948)
     Open-pit                 NO effluent
     Natural ore
     Physical processing

 Operation  D
     Low  tonnage              Older  plant (1953)
     Open-pit                 Mine  produces  effluent
     Natural ore              Processing  plant produces effluent
     Physical  processing

 Operation  E
     High tonnage              Medium age  plant (1967)
    Open-pit
    Iron formation            Mine produces  effluent
    Physical  processing      Processing  plant has closed
                                  water  system

Operation F
    High tonnage             Medium age  plant  (1967)
    Open-pit                 No effluent
    Iron formation
    Physical processing

Operation G
    Low tonnage              Older plant (1959)
    Open-pit                 Mine produces effluent
    Iron formation           Processing plant produces
                                  effluent
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    Physical and chemical
         processing

Operation H
    Medium tonnage           Older plant (1956)
    Open-pit                 Mine produces effluent
    Iron formation           Processing plant produces effluent
    Physical and chemical
         processing

Operation I
    Medium tonnage           Medium age plant (1964)
    Open-pit                 Mine produces effluent
    Iron formation           Processing plant produces effluent
    Physical and chemical
         processing

Operation J
    Low tonnage              Older plant (1958)
    Underground              Mine produces effluent
    Iron formation           Processing plant produces effluent
    Physical and chemical
         processing
The  mines  visited  and  sampled  had  a 1973 production of
approximately  43,853,450  metric  tons  (48,350,000   short
tons), or 47.5 percent of the total United States production
of iron ore.

One  of the initial goals of this study was determination of
the validity of the  initial  categorization.   The  primary
source   of  the  data  utilized  for  this  evaluation  was
information obtained during this study,  plant  visits,  and
sampling  program.   This  information was supplemented with
data obtained through  personal  interviews  and  literature
review  and with historical effluent quality data from NPDES
permit applications and monitoring data supplied by the iron
mining and beneficiating industry.

Based on this exhaustive review, the preliminary  industrial
categorization was substantially altered.

The  data  review  revealed  two distinct effluents from the
mining and milling of iron.  The first  (I) coming  from  the
mines  and  second   (II) coming from the mills.  It was also
determined that all mills in general could  not  be  classed
together.   This  is  primarily  because  a  large number of
milling operations  achieve  zero  discharge  without  major
upset to presently used concentrating technology.
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The milling categorized into three distinct classes based on
the type of ore and the type of processing.

    Category Ila.  Mills    using    physical     separation
                   techniques,    exclusive    of   magnetic
                   separation (washing,  jigging,  cyclones,
                   spirals, heavy media).

    Category lib.  Mills using flotation processes and using
                   the addition of chemical reagents.

    Category lie.  Mills using magnetic separation  for  the
                   benefication of iron formations.

Final  Iron-Ore  Subcateqorization.    Based on the types of
discharges found from all mills, the first two subcategories
can be grouped  into  a  single  segment.   Mills  employing
magnetic    separation   (no   chemical   separation)   have
demonstrated that a distinct subcategory can be made because
of the type of ore, and the mode of beneficiation.

I.  Mines Open-pit or underground, removing natural ores  or
         iron formations.

II. Iron  ore  mills   employing   physical   and   chemical
    separation  and  iron  ore mills employing only physical
    separation (not magnetic)

III.   Iron  ore  mills  employing  magnetic  and   physical
         separation

Copper Ores

The  copper-ore  subcategorization  consideration began with
the approach that mineralization and  ore  beneficiating  or
process method were intimately related to one another.  This
relationship  together  with  a  basic division into mining,
milling and  hydrometallurgical  processing  resulted  in  a
preliminary  subcategorization  scheme  based  primarily  on
division  into  mine  or  concentrating  facility  and  then
further  based  on the method of concentrating or extraction
of values from the ore.  Examination of water  quality  data
supplied  by  the  industry and other sources indicated that
division of mills  into  further  subcategories  based  upon
process  resulted  in grouping operations with similar water
quality characteristics.  Other factors such as climate  and
rainfall    presented    problems    of    subcategorization
particularly with respect to conditions prevalent in certain
areas during approximately two months of the year.
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Final Copper-Ore Subcategorization

Based on data collected from existing sources in addition to
visits  and  sampling  of  copper   mines   and   extraction
facilities,  the  following  final  subcategories  have been
established based primarily on designation as a mine or con-
centrating or chemical extraction facility:

    I.   Mines - Open-pit or underground, removing  sulfide,
              oxide,  mixed  sulfide/oxide  ores,  or native
              copper.
    II.  Copper mines employing hydrcmetallurgical processes

    III. Copper mills employing the vat-leaching process

    IV.  Copper mills employing froth flotation

Problems in Subcateqorizinq the Copper Industry.  Copper  is
produced  in  many  areas of the United States which vary in
mineralization,  climate,  topography,   and   process-water
source.  The processes are outlined in Section V.  The froth
flotation  of  copper  sulfide  is adjusted to conditions at
each plant and will also vary from day to day with the  mill
feed.

Excess  runoff from rainfall and snow melt do alter the sub-
categorization, but they can be controlled by enlargement of
tailing ponds and construction of diversion ditching.   Pre-
sently  a  few  mines"  send the drainage to the mill tailing
lagoon or use the water in the leach circuits.   A  decrease
in  excess  water  problems can be realized in many cases if
mine water is treated separately from mill process water.

some  industry  personnel  have   indicated   concern   that
dissolved  salt  buildup may cause problems in the recycling
of mill process waters when the makeup water  source  and/or
ore body contain a high content of dissolved salts; however,
data   has  not  been  provided  to  support  this  concern.
Molybdenum mills in Canada indicate that the  mill  tailings
include  a built-in blowdown in the form of water trapped in
the interstitial voids of  the  tailings  and  the  product.
This  blowdown  removes  part  of the dissolved salts from a
recyle operation with  the  result  that  the  circuits  can
operate  on  a  zero discharge.  Additional treatment of the
process water for removal of some of the waste  constituents
may  be  necessary  for  recycle  of  process  water and may
produce a zero effluent from many plants  where  buildup  of
materials may adversely affect recovery.
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Lead and Zinc Ores

As a result of an initial review of the lead/zinc mining and
milling industry which considered such factors as mineralogy
of ore, type of processing, size and age of facility, wastes
and treatability of waste, water balance associated with the
facilities, land availability, and topography, a preliminary
scheme  for  subcategorization of the lead/zinc industry was
developed.  The preliminary analysis disclosed that size and
age of  a  facility  should  have  little  to  do  with  the
characteristics  of the wastes from these operations in that
the basic flotation cells have not changed significantly  in
a  decade.   The reagents used, even in very old facilities,
can be  essentially  the  same  as  in  the  newest.   These
factors,  in  addition  to  life  of  an  ore body, and such
factors as  land  availability,  topography,  and,  perhaps,
volume  of  water  which  must  be  removed from a mine have
little to do with  technology  of  treatment  but  can  have
considerable  effect  on  the cost of a treatment technology
employed in a specific case.

The  preliminary  subcategorization  scheme   utilized   was
selected to provide subcategorization on basic technological
factors  where  possible.   The  factors  considered  in the
preliminary scheme were:

    I.   End Product Recovered:
          (a)  Lead/zinc
          (b)  Zinc
          (c)  Lead
          (c)  others with lead/zinc byproducts

   II.   Designation as a Mine or Mill:
          (a)  Mine
          (b)  Mill
          (c)  Mine/mill complex

  III.   Type of Processing:
          (a)  Gravity separation (no reagents)
          (b)  Flotation

   IV.   Wastes or Treatability of Wastes Generated:
          (a)  Potential for development of conditions
              with soluble undesirable metals or salts
          (b)  No potential for solubilization

    V.   Water Balance:
          (a)  Total recycle possible
          (b)  Total recycle not possible
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The plant visits and  subsequent  compilation  of  data  and
literature  review  were aimed at establishing which factors
were really significant in determining what effluent quality
could be achieved with respect to the tentative subcategori-
zation.

An analysis of the data compiled indicated that subcategori-
zation within the lead/zinc  industry  could  be  simplified
considerably.   No  basic  differences  in treatability were
found  to  be  associated  with  the  type  of  concentrates
obtained from a facility.

The  proposed  subcategorization  based  on what facility is
discharging—that is, a mine or a mill—is justified because
effluents from a mine dewatering operation and those from  a
milling  operation,  into  which  various  chemicals  may be
introduced, are different.  In  the  case  of  a  mine  dis-
charging  only  into  the water supply of the mill, the only
applicable guideline would be that of the mill.

No evidence of current practice of strictly physical concen-
tration by gravity separation was found.   The  recovery  of
desirable  minerals  from known deposits utilizing only such
physical separations is likely to be so poor as to result in
discharge of significant quantities of  heavy-metal  sulfide
to  the  tailing retention area.  The only ore concentration
process currently practiced in  the  lead/zinc  industry  is
froth flotation.  Subcategorization based on milling process
is, therefore, not necessary.

The   treatability   of  mine  wastewater  is  significantly
affected by the occurrence of  local  geological  conditions
which  cause  solubilization of undesirable metals or salts.
A common, and well-understood, example is acid mine drainage
caused by the oxidation of pyrite (FeS_2) to ferrous  sulfate
and  sulfuric  acid.   This oxidation requires both moisture
and air (oxygen source) to occur.  The acid  generated  then
leaches  heavy  metals  from  the  exposed  rock on particle
surfaces.  Heavy metals may also enter solution as a  result
of  oxidation  over  a  period  of time through fissured ore
bodies to form more soluble oxides of heavy metals  (such  as
zinc)  in  mines which do not exhibit acidic mine drainages.
Another route which may result in solubilized  heavy  metals
involves  the  formation  of acid and subsequent leaching in
very local areas in an ore body.  The resultant acid may  be
neutralized  by  later  contact  with limestone or dolomitic
limestone, but the pH level attained may not be high  enough
to  cause  precipitation  of  the  solubilized  metals.  The
important aspect of all of these situations is that the mine
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water encountered is much more difficult to treat than those
where solubilization conditions do not occur.   The  treated
effluents  from mines in this situation often exhibit higher
levels of heavy  metals  in  solution  than  untreated  mine
waters  from  mines  where  solubilization conditions do not
occur.

It has been determined that subcategorization on  the  basis
of  solubilization  potential is not justified; however, the
effluent limits recommended have  taken  into  consideration
this factor.

The  water-balance  parameter,  of course, does not apply to
mine only operations.  In the case  of  milling  operations,
system  design  and  alteration  of  process  flows can have
considerable effect  on  the  water  balance  of  a  milling
operation.  No justification was found for substantiation of
subcategorization on this basis.

The  final  recommended  subcategorization for the lead/zinc
mining and milling industry is, therefore, condensed to:

I.  Lead and/or zinc mines

II.  Lead and/or zinc mills

Gold Ores

The most important factors considered in determining whether
subcategorization was necessary for the  gold  ore  category
were  ore  mineralogy,  general  geologic  setting,  type of
processing, wastes and waste  treatability,  water  balance,
and   final   product.    Upon   intensive  background  data
compilation (as discussed in Section III), mill inspections,
and communications with the industry, most  of  the  factors
were  found  to  reduce to mineralogy of the ore (and, thus,
product)   and  milling  process   emfloyeS.    The   initial
subcategorization  was  found  to  differ  little from final
subcategorization  selection  after  site   visitation   and
sampling data were obtained.

The  most  effective means of categorizing the gold industry
is based upon relative differences among existing sources of
discharge   (mine  or  mill/mine-mill   complexes)    and   on
characteristics of the beneficiation process.  The rationale
for this is based on several considerations:

    (1)   Apart from milling processing,  the  characteristic
         difference  between  mine  effluents and mill/mine-
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         mill   effluents   is   their   quantitative    and
         qualitative  pollutant  loadings.    This difference
         between  mines  and  mills  makes    necessary   the
         application   of  differing  waste-treatment  tech-
         nologies and/or  the  segregation   of   sources  for
         purposes  of  treatment.   A mill  effluent normally
         contains a greater quantity of total solids—up  to
         UO  to  50 percent more than a mine effluent.  Much
         of these solids are suspended solids,  and treatment
         involves removal  by  settling.   This  is  usually
         treated in tailing ponds.  Where mines occur alone,
         or  where  their  effluents  are treated separately
         from the mill, these effluents may be  treated on  a
         smaller scale by a different technology.

    (2)   The specific beneficiation process adapted is based
         on  the  geology  and  mineralogy   of  the ore.  The
         waste characteristics and treatability of the  mill
         effluent   are   a   function   of  the  particular
         beneficiation process employed.   This  takes  into
         account   the   reagents   used   and   the  general
         mineralization  of  the  ore  by  each   particular
         process  as  these  factors  affect differing waste
         characteristics.  The waste characteristics  affect
         treatability; for example, cyanide removal requires
         different  technology  than  that   used  for  metal
         removal.

Consideration was also given to the regional availability of
water, as this factor is relevant to water  conservation  and
"no  discharge"  and waste-control feasibility.  Since it is
common engineering  practice  to  design  tailing  ponds  to
accommodate  excesses  of  water, and also since pond design
can include systems to divert surface runoff away  from  the
pond,  regional availability of water was judged not to be a
limiting factor with respect to the  feasibility  of  a  no-
discharge system.

Final Gold-Ore Subcategorization

On the basis of the rationale developed above and previously
discussed  in  the introductory portion of this section, six
subcategories  were  identified  for  the  gold  mining  and
milling industry:

    I.   Mine(s) alone.

    II.  Mill(s) or mine/mill complex(es) using the process
              of cyanidation for primary or byproduct
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              recovery of gold.

    III. Mill(s) or mine/mill coroplex(es) using process of
              amalgamation  (includes dredging operations,
              if amalgamation is used).

    IV.  Mill(s) or mine/mill complex(es) using the process
              of flotation.

    V.   Mill(s) or mine/mill complex(es) using gravity
              separation  (includes dredging or hydraulic
              mining operation).

Silver Ores

The  development of subcategorization in the silver industry
was essentially identical  to  that  of  the  gold  industry
previously  discussed.   The primary basis for division into
subcategories was mineralogy of the ore and type of process-
ing.  Since mineralogy and type of extraction processing are
intimately related, these factors served,  just  as  in  the
gold  industry,  to  divide  the industry into mine and mill
categories, and then further into milling  categories  based
upon  type  of  processing.  Also note that, in many places,
gold and silver are exploited as coproducts or, together, as
byproducts of other base metals (such as copper) .

Final Silver-Ore Subcateqorization

Eased upon the previous rationale developed  in  the  intro-
ductory  portion of this section (and also discussed in con-
nection with gold  ores),  tentative  subcategorization  was
developed  and  then  verified  by  field  sampling and site
visits.   Based  upon  field  confirmation,  the   tentative
subcategories, found to be unchanged, are:

    I.   Mine(s) alone

    II.  Mill(s) or mine/mill complex(es) using flotation
              for primary or byproduct recovery of silver.

    III. Mill(s) or mine/mill complex(es) using cyanidation
              for primary or byproduct recovery of silver.

    IV.  Mill(s) using amalgamation process for primary
              or byproduct recovery of silver.

    V.   Mill(s) using gravity separation process for primary
              or byproduct recovery cf silver.
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Bauxite Ores

The bauxite mining industry is classified as SIC 1051, which
includes   establishments  engaged  in  mining  and  milling
bauxite and other aluminum ores.  Since there are  no  other
aluminum  ores  being commercially exploited on a full-scale
basis at present, the bauxite mining industry serves as  the
sole   representative   of   SIC   1051.   Subcategorization
rationale was preceded by this fact, and future  development
of  other  aluminun ores will likely necessitate revision of
present categorization.

In the bauxite mining industry, most criteria  for  subcate-
gorization  bear  directly  or  indirectly  upon  two  basic
factors: (1)  nature  of  raw  mine  drainage,  which  is  a
function  of  the  mineralogy and general geological setting
related to percolating waters; and (2)  treatability of waste
generated,  based  upon  the   quality   of   the   effluent
concentrations.   Initially,  general  factors,  such as end
products,  type  of  processing,  climate,   rainfall,   and
location,  proved  to be of minor importance as criteria for
subcategorization.   The   two   existing   bauxite   mining
operations  are  located adjacent to one another in Arkansas
and share similar rainfall and evaporation rates, 122 cm (48
in.)  and 109 cm  (43 in.).  Both operations produce bauxite,
though slightly different in grade, which  is  milled  by  a
process emitting no wastewater.

After  the  site  visits  to  both  operating  mines, it was
evident that the mining technique is closely associated with
the  characteristics  of  the  mine   drainage,   and   that
mineralization  is directly responsible for mining-technique
and raw mine drainage characteristics.  However,  subsequent
discussion   with   industry  personnel  suggests  that  the
characteristics of the raw mine  drainage  are  more  simply
correlated  with geochemical factors associated with surface
and subsurface drainage rather than with  mining  technique.
Regardless  of  the  factors dictating the nature of the raw
mine  drainage,  an  evaluation  of   wastewater   treatment
efficiency  for a treatment process common to both members of
the  industry  became the prime consideration in determining
attainable  treated effluent concentrations.

Final Bauxite-Ore Subcateqorization

Based  on   the  results   of    intensive   study,   facility
inspections,  NPDES  permit  applications, and communication
with the industry, it was concluded that the bauxite  mining
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and  milling  industry  should  not be subcategorized beyond
that presented below.

         Bauxite mining and associated milling operations
              (essentially grinding and crushing)

Ferroalloy Ores

In development of subcategories for  the  ferroalloy  mining
and  milling category, the following factors were considered
initially:   type  of  process,  and  product,   mineralogy,
climate,  topography,  land  availability,  size,  age,  and
wastes or treatability of wastes generated.

A tentative subcategorization of the industry was  developed
after collection and review of initial data, based primarily
on  end  product  (e.g.,  tungsten,  molybdenum,  manganese,
etc.), with further division on the  basis  of  process,  in
some  cases.   Further  data,  particularly chemical data on
effluents  and  more  complete   process   data   for   past
operations,  indicated  that process was the dominant factor
influencing    waste-stream    character    and    treatment
effectiveness.   Examination  of  the  industry additionally
showed that  size  of  operation  could  also  be  of  great
importance.   Other factors, except as they are reflected in
or derived from the  above,  are  not  believed  to  warrant
industry subcategorization.

Final Ferroalloy-Ore Subcateqorization

It  has  been  determined  that  the  ferroalloy  mining and
milling category should be divided into  five  subcategories
for the purpose of establishing effluent limitations and new
source performance standards:
    I.   Mines

    II,  Mines and Mills processing less than 5,000
              metric tons (5,512 short tons)  per year
              of ore by methods other than ore leaching.

   III.  Mills processing more than 5,000 metric tons per
              year of ore by purely physical methods (e.g.,
              crushing, ore washing, gravity separation,
              and magnetic and electrostatic separation).

    IV.  Mills processing more than 5,000 metric tons per
              year of ore and employing flotation.
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V.
          Mills practicing ore leaching and associated
               chemical benef iciation techniques.
 The  subcategory  including mills processing less than 5,000

 mv^ru°   ns Of ore per year is representative of  operations
 which  are typically both intermittent in operation and eco^
 nomically  marginal.   This  subcategory  is   believed   to

 tungsten ores. preS6nt'  alm°st  exclusively  processors  of
       •Pl?YSiCal.prOCesS:Lng Provides  the  minimum  pollution
 potential  consistent  with  recovery  of values from an ore
 using water.  All mills falling into* this  subcategSry  Ire
 expected  to share the same major pollution problem--namelv

 arSd?ned "St* generated *»  th*  need  for  crushing  and
 grinding.   The  exposure of finely divided ore (and gangue)
 to water may also lead to solution of some material, bu?  ±n
 >
 is   not   r^ 6X?ep^n  the  case  of  nickel,  where water  use
 Jf   ??    f JallS  ln the Process.   Information has  been  drawn
 heavily,  therefore,  from  past  data  and related  mill ?«£
                                       .;
reasonable.  These milling processes   are  fully  compatible
with recycle of all mill water.                Y  comPat:Li3le

The   practice   of   flotation  significantly
character of mill effluent in several  ways
water PH is altered  or  controlled  t^iAcre
efficiency    This, together with the  rart tSt ore  rn   s
generally finer than for physical processing, may  h£e"tte
secondary  effect  of substantially increasing Solubility of
ore components.  Reagents added to effect th            y
include major pollutants.  Cyanide, for exam
used  and   though  usage is^ow/kfLcee
The  added  reagents  may  have  secondary  effects
treatment effectiveness,  some flotaSon SpeStion= deore"e
differ from physical processors in the
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may be recycled without process changes or serious  recovery
losses.

Ore  leaching  operations differ substantially from physical
processors and flotation plants in  effluent  character  and
treatment  requirements.   The  use  of large quantities (in
relation to ore handled) of  reagents,  and  the  deliberate
solubilization of ore components, characterizes these opera-
tions.   The  solutilization  process  is  not,  in general,
entirely specific, and the recovery of desired  material  is
less  than  100 percent.  Large amounts of dissolved ore may
be expected, therefore, to  appear  in  the  mill  effluent,
necessitating  extensive  treatment prior to discharge.  For
these operations, even commonly occurring ions  (i.e.,  Na+,
SO**,  etc.)  may be present in sufficient quantities to cause
ma^or  environmental  effects,  and  total   dissolved-solid
levels  can  become  a  real  (although somewhat intractable)
problem.   Wide  variations   in  leaching  processes   might
justify  further division of this subcategory  (into acid and
alkaline leaching, as in the uranium industry, for example),
but the limited current activity and data available at  this
time do not support such a division.

Other Considerations.   Climate, topography, and land avail-
ability are extremely important factors influencing effluent
volume,  character,  and treatment in the mining and milling
industry—particularly, the  attainment  of  zero  pollutant
discharge by means of discharge elimination.  Zero discharge
may be attainable, for example, despite a net positive water
balance  for  a  region  because rainfall input to a tailing
impoundment balances part of the process water loss, includ-
ing evaporative losses in the  mill  and  retention  in  the
tails  and  product.   It  is  anticipated  that,  under the
impetus of effluent limitations established under PL 92-500,
and the resultant pollution control costs, many mills in the
defined subcategories will choose the often  less  expensive
option of discharge elimination.

Mercury Ores

The  mercury industry in the United states currently is at a
reduced level of activity due to  depressed  market  prices.
Two  facilities  were  found  to  be  operating  at present,
although it is thought that  activity  will  again  increase
with  increasing  demand  and  rising  market  prices.   The
decreased use of mercury due  to  stringent  air  and  water
pollution regulations in the industrial sector may be offset
in  the future by increased demand in dental, electrical and
other uses.   Historically, little beneficiating  of  mercury
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ores  has  been  known in the industry.  Common practice for
most   producers    (since    relatively    low    production
characterizes most operators) has been to feed the cinnabar-
rich   ore   directly   to   a   kiln   or  furnace  without
beneficiation.  Water use in most of the operations is at  a
minimum,  although  a rather large  (20,000-flask-peryear, or
695-metric-ton-per-year      or      765-short-ton-per-year)
production capacity flotation operation with a process water
requirement  of 5.3 cubic meters per minute (1,400 gpm).  In
the year 1985, the industry  could  be  producing  3,000  to
20,000  flasks  (104 to 695 metric tons, or 115 to 765 short
tons)  per year, depending on market price,  technology,  and
ore grade (Reference 8).

Final Mercury Ore Subcateqorization

Since  historically most mercury operations have been direct
furnacing  facilities,   the   resulting   subcategorization
represents that fact.  Little or no beneficiation is done in
the   industry,  with  few  exceptions.   There  are  a  few
operations from which mercury is recovered as a byproduct at
a smelter or refinery.  A single known  flotation  operation
has  recently  begun  operation  and  is  reflected  in  the
subcategorization scheme below based on processing.

    I.   Mine(s) alone or mine(s) with crushing and/or
              grinding prior to furnacing (no additional
              beneficiation).

   II.   Mill(s) or mine/mill complex(es)  using the process
              of gravity separation for primary or byproduct
              recovery of mercury.

  III.   Mill(s) or mine/mill complex(es)  using flotation
              for primary or byproduct recovery of mercury.

Uranium, Radium, and Vanadium Ores

The factors evaluated in consideration of  subcategorization
of the uranium, radium, and vanadium mining and ore dressing
industry   are:     end  product,  type  of  processing,  ore
mineralogy,   waste   characteristics,    treatability    of
wastewater, and climate, rainfall, and location.  Based upon
an  intensive  literature  search,  plant inspections, NPDES
permits, and communications with the industry, this category
is categorized by milling process and mineralogy (and, thus,
product).  A discussion of each of the  primary  factors  as
they   affect   the   uranium/radium/vanadium  ore  category
follows.
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The milling  processes  of  this  industry  involve  complex
hydrometallurgy.   such  point  discharges as might occur in
milling processes  (i.e., the production cf concentrate)  are
expected  to contain a variety of pollutants that need to be
limited.  Mining, for the ores, is expected  to  lead  to  a
smaller  set  of  contaminants.   while mining or milling of
ores for uranium or  radium  produces  particularly  noxious
radioactive  pollutants,  these  are  largely  absent  in an
operation recovering vanadium only.  On the basis  of  these
considerations,   the  SIC  1094  industry  was  tentatively
subcategorized into:   (1) The mining of uranium/radium ores;
(2) The processing of the ores of the first  subcategory  to
yield    uranium   concentrate   and,   possibly,   vanadium
concentrate; (3)  The  mining  of  non-radioactive  vanadium
ores;  and   (4)  The  processing  of  the  ores of the third
subcategory to yield vanadium concentrate.

A careful distinction will be drawn between the  radioactive
processes  and  the  vanadium  industry  by including in the
former all operations within SIC 1094 that are  licensed  by
the  U.S.  Nuclear Regulatory Commission  (NRC, formerly AEC,
Atomic Energy  Commission)  or  by  agreement  states.   The
agreement  states, including the uranium producing states of
Colorado,  Texas,  New  Mexico  and  Washington,  have  been
delegated  all  licensing,  record  keeping,  and inspection
responsibilities for radioactive materials regulated by  the
NRC  upon  establishing  regulations  regarding  radioactive
materials that are compatible with those  of  the  NRC(AEC).
The  licensing  requirements,  as  set  forth in the code of
Federal Regulations, Title 10, Parts 20 and  40,  constitute
present  restrictions  on  the  discharge  of radionuclides.
Uranium mines are regulated by some states for discharge  of
radioactive  materials  but  this regulation is not based on
"agreement state" authority since the NRC does not  regulate
the uranium mines.

To  further  emphasize  the  distinction  between  the  NRC-
licensed  uranium  subcategories  and  the   pure   vanadium
subcategories,  the  latter,  whose products are used in the
inorganic chemical industry and,  to  a  large  extent,  the
ferroalloy  smelting  industry,  are  discussed  further  in
connection with ferroalloymetal ore mining and dressing,  in
another   portion   of   these   guidelines.   The  vanadium
subcategories are summarized there as members of the  mining
and hydronetallurgical process subcategories.

The  variety  of  ores  and  milling  processes discussed in
Section  III  might  lead  to  the  generation  of  as  many
subcategories based on the major characteristics of the mill
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process  as  there  are  ores  and  mills.   It is possible,
however, to group  mills  into  fewer  subcategories.   This
simplification is based on the observations discussed below.

Raw  wastewaters  from mills using acid leaching remain acid
at the process discharge (not to be confused  with  a  point
discharge),  retain  various heavy metals, and are generally
not  suitable   for   recycling   without   additional   and
specialized   treatment.   Those  from  the  alkaline  leach
process can be recycled in part, since the leach process  is
somewhat  selective  for  uranium  and  vanadium,  and other
metals remain in the solid tailings.  At one  time,  it  was
expected  that  mills  using  solvent  exchange would have a
radically different raw waste character due to the discharge
of organic compounds.  The fact that mills not using solvent
exchange often process ore that is  rich  in  organics  make
this distinction less important.  As a result, a distinction
can  be made between mills using acid leaching (or both acid
and alkaline leaching)   of  ore  and  mills  using  alkaline
leaching  of  ore  only.  However, such subcategorization is
not required when addressing effluent limitations for mills.

While other differences between ores and processes, in addi-
tion to  those  mentioned  above,  can  have  an  effect  on
wastewater characteristics, they are not believed to justify
further  subcategorization.   For  example,  there  are some
uranium/radium ores that contain molybdenum and others  that
do  not.  Effluent limitations which may restrict molybdenum
content must be applied at  all  times  and  should  not  be
restricted  to  those  operations which happen to run on ore
containing molybdenum.    The  two  subcategories   (acid  and
alkaline)   considered   reflect  not  only  differences  in
wastewater characteristics but also (a) differences  in  the
volume of wastewater that must be stored and managed and (b)
differences  in  the  ultimate  disposition  of  wastes upon
shutdown of an operation.

Climatic conditions  (such  as  rainfall  versus  evaporation
factors  for  a  region),  although  subject to questions of
measurement, have an important influence on the existence of
present-day point discharges and, thus, have been considered
relative to  present  and  future  exploitation  of  uranium
reserves  in  the  United  states.   All exploitable uranium
reserves presently economical to develop are found  in  arid
climates.   Therefore,  no  point  discharges  are needed to
manage the raw wastewater from  most  current  ore  dressing
operations  in  the  uranium industry.  However, one milling
operation now discharges wastewater.
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Ore  characteristics   were   considered   and,   within   a
subcategory,   cause   short-term   effect   on   wastewater
characteristics    that    does    not    justify    further
subcategorization.  Waste characteristics were, as described
above, considered extensively, and it was found difficult to
distinguish  whether  the acid/alkaline leach distinction is
based on  process,  mineralogy,  waste  characteristics,  or
treatability  of  wastewater,  since  all  are interrelated.
Vanadium operations which are not extracting radioactive ore
or covered under government licensing  regulations  (NRC  or
agreement  states),  are  subcategorized  in the ferroalloys
section.

Final Subcateqorization of  Uranium,  Radium,  and  Vanadium
Category

The  uranium, radium, and vanadium segment of the mining and
ore dressing industry considered  here  has  been  separated
into   the   following  subcategories  for  the  purpose  of
establishing  effluent  guidelines  and  standards.    These
subcategories are defined as:

    I.   Mines which extract  (but do not  concentrate)   ores
              of uranium, radium, or vanadium.

   II.   Mills which process uranium,  radium,  or  vanadium
              ores   to   yield   uranium  concentrate  and,
              possibly, vanadium concentrate by either acid,
              alkaline   or    combined    acid-and-alkaline
              leaching.

Metal Ores. Not Elsewhere Classified

This  group of metal ores was considered on a metal-by-metal
basis  because  of  the  wide  diversity  of   mineralogies,
processes  of  extraction,  etc.   Most of the metal ores in
this group do not have high production figures and represent
relatively few  operations.   For  this  entire  group,  ore
mineralogies  and  type  of  process  formed  the  basis  of
subcategorization.  The  metals  ores  examined  under  this
category  are  ores  of  antimony, beryllium, platinum, tin,
titanium, rare earths  (including monazite) , and zirconium.

Antimony Ores

Mining and milling of ore for primary recovery  of  antimony
is   paracticed  at  one  location  in  the  United  States.
Although antimony is often found  as  a  byproduct  of  lead
                            171

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extract!on,  producers  are  often  penalized  for  antimony
content at a smelter.

Final Antimony-Ore subcategorization

The antimony ore  mining  and  dressing  industry  has  been
separated   into   two  subcategories  for  the  purpose  of
establishing  effluent  guidelines  and  standards.    These
subcategories are defined as:

    I.   Mine(s) alone operating for the extraction of  ores
              to obtain primary or byproduct antimony ores.

   II.   Mill(s) or mine/mill complex(es) using a  flotation
              process  for the primary or byproduct recovery
              of antimony ore.

Beryllium Ores

Beryllium mining and milling in the United States are repre-
sented    by    one    operating    facility.     Therefore,
subcategorization consists simply of division into mines and
mills:

    I.   Mine(s) operated for  the  extraction  of  ores  of
              beryllium.

   II.   Mill(s)  or  mine/mill  complex(es)   using  solvent
              extraction  (sulfuric-acid leach) .

Platinum Ores

As  discussed previously, most production of platinum in the
United States is as byproduct  recovery  of  platinum  at  a
smelter  or refinery from base- or ether precious-metal con-
centrates.  A single operating location mines and  benefici-
ates  ore by use of dredging, followed by gravity separation
methods.  A single category, thus, is  listed  for  platinum
ores:

    I.   Mine/mill complex(es)  obtaining  platinum  concen-
              trates   by   dredging,  followed  by  gravity
              separation and beneficiation.

Rare-Earth Ores

Rare-earth ores currently are obtained  from  two  types  of
mineralogies:  bastnaesite and monazite.  Monazite is an ore
both  of thorium and of rare-earth elements, such as cerium.
                            172

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The subcategorization which follows is based primarily  upon
division  into  mines  and  mills, as well as on the type of
processing  employed  for  extraction  of  the  rare-  earth
elements.

    I.   Mine(s) operated for the extraction of  primary  or
              byproduct ores of rare-earth elements.

   II.   Mill(s) or mine/mill  complex(es)  using  flotation
              process   and/or  leaching  of  the  flotation
              concentrate  for  the  primary  or   byproduct
              recovery of rare-earth minerals.

  III.   Mill(s) or mine/mill complex(e3) operated  in  con-
              junction  with  dredging  or  hydraulic mining
              methods;  wet  gravity  methods  are  used  in
              conjunction with electrostatic and/or magnetic
              methods  for the recovery and concentration of
              rare-earth minerals  (usually, monazite).

Tin Ores

Some tin concentrate was produced at dredging operations  in
Alaska  and  placer  operations  in  New  Mexico.   A single
operating facility currently produces tin as a byproduct  of
molybdenum  mining and beneficiation.  Other placer deposits
of tin may be discovered and could be exploited.  Therefore,
a single subcategory for  mining  and  one  subcategory  for
milling are listed:

    I.   Mine(s) operating  for  the  primary  or  byproduct
              recovery of tin ores.

   II.   Mill(s)  or  mine/mill  complex(es)   using  gravity
              methods.

Titanium Ores

Titanium  ores  exploited  in the United States occur in two
modes and mineralogical associations:  as  placer  or  heavy
sand  deposits  of rutile, ilmenite, and leucoxene, and as a
titaniferous magnetite in a hard-rock deposit.  The titanium
ore industry, therefore, is subcategorized as:

    I.   Mine(s)  obtaining  titanium  ore  by  lode  mining
              alone.

   II.   Mill(s) or  mine/mill  complex(es)   using  electro-
              static  and/or magnetic methods in conjunction
                            173

-------
              with  gravity  and/or  flotation  methods  for
              primary  or  byproduct  recovery  of  titanium
              minerals.

  III.   Mill(s) or  mine/mill  complex(es)  in  conjunction
              with  dredge  mining  operation;  wet  gravity
              methods used in conjunction with electrostatic
              and/or magnetic methods  for  the  primary  or
              byproduct recovery of titanium minerals.

Zirconium Ores

Zirconium is obtained from the mineral zircon in conjunction
with  dredging  operations.  No additional subcategorization
is required.

    I.   Mill(s) or mine/mill complex(es) operated  in  con-
              junction   with   dredging   operations.   Wet
              gravity methods are used in  conjunction  with
              electrostatic  and/or magnetic methods for the
              primary or  byproduct  recovery  of  zirconium
              minerals.
SUMMARY OF RECOMMENDED SUBCATEGORIZATION

Based  upon  the  preceding  discussion  and choice of final
subcategories, a summary  of  categories  and  subcategories
recommended  for  the  ore  mining  and dressing industry is
presented here  in  Table  IV-1.   The  discussions  in  the
following   sections,  including  the  recommended  effluent
limitations in Sections IX, X,  and  XI,  will  address  the
categories and subcategories presented in Table IV-1.

FINAL SUBCATEGORIZATION

After  an  analysis  of available treatment technologies and
the  effluent  quality  that  could  be  achieved   by   the
application of the available treatment technologies, and the
fact  that  many  metals  occur  in  conjunction  with other
metals, it  was  determined  that  the  final  subcategories
previously   discussed   could   be   combined   into  seven
subcategories based on the product or products.   The  seven
subcategories   can   then   be   further  divided  into  22
subdivisions for which separate  limitations  will  be  set,
based  on  considerations  of type of process and wastewater
characteristics  and  treatability.    The   other   factors
recognized  as  causing differences in the wastes discharged
do not significantly effect the treatability of  the  wastes
                            174

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within   a   subcategory.    Table   IV-2  shows  the  final
subcategorization and the components of each subcategory  as
they  will  be presented in the regulations derived from the
development document.
                            175

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TABLE IV-1. SUMMARY OF INDUSTRY SUBCATEGORIZATION RECOMMENDED
CATEGORY
IRON ORES
COPPER ORES
LEAD AND ZINC ORES
GOLD ORES
SILVER ORES
BAUXITE ORE
FERROALLOY ORES
MERCURY ORES
Uf
&
METAL ORES. NOT ELSEWHERE CLASSIFIED
1ANIUM RADIUM
VANADIUM ORES
ANTIMONY ORES
BERYLLIUM ORES

PLATINUM ORES

RARE-EARTH ORES
TIN ORES
TITANIUM ORES
ZIRCONIUM ORES
SUBCATEQORIES
MINES
MILLS
MINES
MILLS
Phyikaj and Chemical Separation,
Phytlcal Separation Only
Magnetic and Phyiical Separation (Metabl Rang*)
Open-Pit, Underground, Stripping
Hydromatallurgfcal (Leaching)
Vat Leaching
Flotation Proceu
MINES
MILLS
MINES
MILLS
Cyanldatlon Proceu
Amalgametion Procete
Flotation ProeaM
Gravity Separation
Byproduct of Beta-Metal Operation
MINES
MILLS
Flotation Procatt
Cyanidation Procett
Amalgamation Proem
Gravity Separation
Byproduct of Bate-Metal Operation
MINES
MINES
MILLS
< 5,000 metric tons (5,512 thort tonil/year
> 5,000 metric toni/yaar by Phytleal Procetan
> 5,000 metric tont/year by Flotation
Leaching
MINES
MILLS
Gravity Separation
Flotation Procett
Byproduct of Bata/Preciout-Metal Operation
MINES
MILLS
MINES
MILLS
Acid or Acid/Alkaline Leaching
Alkaline Leaching

Flotation Procen
Byproduct of Bata/Preciout-Metal Operation
MINES
MILLS
MINES OR MINE/MILLS
MINES
MILLS


Flotation or Leaching
Dredging or Hydraulic Method!
MINES
MILLS
MINES
MILLS


Electrottatie/Magnatic and Gravity/Flotation Proceteet
Physical Proceuei with Dredga Mining
MILLS OR MINE/MILLS
                      176

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TABLE IV-2. FINAL RECOMMENDED INDUSTRY SUBCATEGORIES
SUBCATEGORY
Iron Orel


w
1
a.
•B
s
j


Bauxite Or*
Ferroalloy Metal Orel
fOU pn f»_ j.Mn ftlj W U|l




Mercury Orel

Uranium, Radium,
& Vanadium Ores
Antimony Orel

Beryllium Orel

Rare-Earth Orel

Titanium Orel



Mines
Mills
Mines
 5,000 metric tons/year by
Physical processes
> 5,000 metric tons/year by
Flotation
Leaching

Gravity Separation
Flotation Process

Acid or Acid/Alkaline Leaching
Alkaline Leaching

Flotation Proceis





Electrostatic/Magnetic and Gravity/
Flotation Processes
Physical Processes with Dredge Mining
Zirconium Ores
Dredging or Hydraulic Methods
(Monazite)
                  177

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                         SECTION V

                   WASTE CHARACTERIZATION
INTRODUCTION

This section discusses the specific water uses  in  the  ore
mining  and  dressing  industry,  as  well as the amounts of
process waste materials  contained  in  these  waters.    The
process  wastes are characterized as raw waste loads emanat-
ing from  specific  processes  used  in  the  extraction  of
materials  involved in this study and are specified in  terms
of kilograms per metric ton (and as pounds per short ton) of
product produced in ore processed.  The specific water  uses
and amounts are given in terms of cubic meters (and gallons)
or  liters  per  metric  ton  (and gallons per short ton) of
concentrate produced or ore mined.  Many  mining  operations
are  characterized  by high water inflow and low production,
or by production rates that bear little relationship to mine
water effluent due to infiltration cr precipitation.   Where
this occurs, waste characteristics are expressed in units of
concentration  (mg/1 = ppm),  The discussion of the necessity
for  reporting the data in this fashion in some instances is
discussed below under the heading "Mine Water."

The introductory portions of this  section  briefly  discuss
the  principal  water  uses  found  in  all  categories  and
subcategories in the industry.  A discussion of each  mining
and  milling sukcategory, with the waste characteristics and
loads identified for each, concludes this section.

Because of widely varying wastewater characteristics, it was
necessary to accumulate data from the widest possible  base.
Effluent  data  presented  for  each  industry category were
derived from historical effluent data supplied by the indus-
try and various regulatory and  research  bodies,  and  from
current  data  for  effluent  samples collected and analyzed
during  this   study.   The   wastewater   sampling   program
conducted during this study had two purposes.  First, it was
designed  to   confirm  and supplement the existing data.  In
general, only  limited characterization  of  raw  wastes  has
been  previously  undertaken by industry.  Second, the scope
of the water-quality analysis was expanded  to  include  not
only   previously   monitored  parameters,  but  also  waste
parameters which could be present in mine drainage  or  mill
effluents.
                             179

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Mine Water

The  wastewater  situation  evident in the mining segment of
the  ore  mining  and  dressing  industry  is  unlike   that
encountered   in   most  other  industries.   Usually,  most
industries (such as the milling segment  of  this  industry)
utilize  water  in the specific processes they employ.  This
water frequently becomes contaminated during the process and
must be treated prior to discharge.  In the mining  segment,
process  water is not normally utilized in the actual mining
of ores except where it is used in placer mining  operations
(hydraulic  mining and dredging) and in dust control.  Water
is a natural feature that interferes with mining activities.
It enters mines by  ground-water  infiltration  and  surface
runoff  and  comes  into  contact with materials in the host
rock, ore, and overburden.  An additional source of water in
deep underground mines is that water which results from  the
backfilling  of  stopes with the coarse fraction of the mill
tailings.  Transportation  of  these  sands  underground  is
typically  accomplished  by  sluicing.   Mill  wastewater is
usually the source of the sluice water.  The mine water then
requires treatment depending on its quality before it can be
safely  discharged  into  the  surface   drainage   network.
Generally,  mining operations control surface runoff through
the use of diversion ditching, and grading  to  prevent,  as
much  as  possible,  excess  water from entering the working
area.  The quantity of  water  from  an  ore  mine  thus  is
unrelated,   or   only  indirectly  related,  to  production
quantities.  Therefore, raw waste loadings are expressed  in
terms  of  concentration  rather than units of production in
the ore categories discussed in Section IV.

In addition to handling and treating often  massive  volumes
of  mine drainage during active mining operations, metal ore
mine operators are  faced  with  the  same  problems  during
startup,   idle   periods,  and  shutdown.   Water  handling
problems are generally minor during initial startup of a new
underground mining operation.  These problems  may  increase
as the mine is expanded and developed and may continue after
all  mining  operations have ceased.  The long-term drainage
from tailing  disposal  also  presents  long-term  potential
problems.   Surface  mines,  on the other hand, are somewhat
more predictable and less permanent in their  production  of
mine  drainage period.  Water handling within a surface mine
is fairly uniform throughout the life of the  mine.   It  is
highly    dependent    upon   precipitation   patterns   and
precautionary methods employed, such as the use of diversion
ditches, burial of toxic materials, and concurrent regrading
and revegetation.
                            180

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Because mine drainage does not necessarily cease  with  mine
closure,  a decision must be made as to the point at which a
mine   operator   has   fulfilled   his   obligations    and
responsibilities  for  a  particular  mine site.  This point
will be further  discussed  in  Section  VII,  "Control  and
Treatment Technology."

SPECIFIC WATER USES IN AIL CATEGORIES

Water  is  used  in the ore mining and dressing industry for
ten principal uses falling  under  three  major  categories.
The principal water uses are:

         (1)  Noncontact cooling water

         (2)  Process water - wash water
                              transport water
                              scrubber water
                              process and product consumed
                                water

         (3)  Miscellaneous water -
                              dust control
                              domestic/sanitary uses
                              washing and cleaning
                              drilling fluids

Noncontact  cooling  water  is defined as that cooling water
which does  not  come  into  direct  contact  with  any  raw
material,  intermediate  product, byproduct, or product used
in or resulting from the process.

Process water is defined as that  water  which,  during  the
beneficiation  process,  comes  into direct contact with any
raw material, intermediate product,  byproduct,  or  product
used in or resulting from the process.

Noncontact Cooling Water

The  largest  use  of  noncontact  cooling  water in the ore
mining  and  dressing  industry  is  for  the   cooling   of
equipment,    such   as  crusher  bearings,  pumps,  and  air
compressors.

Wash Water

Wash water comes into direct contact  with  either  the  raw
material,  reactants,  or products.  An example of this type
of water usage is ore washing to remove fines.   Waste efflu-
                            181

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ents can arise from these washing sources because the resul-
tant solution or suspension  may  contain  dissolved  salts,
metals, or suspended solids.

Transport Water

Water is widely used in the ore mining and dressing industry
to  transport  ore  to  and  between  various process steps.
Water is often used to move crude ore from mine to mill,  to
move  ore  from crushers to grinding mills, and to transport
tailings to final retention ponds.

Scrubber Water

Wet scrubbers are often used  for  air  pollution  control—
primarily, in association with grinding mills, crushers, and
screens.

Process and Product Consumed Water

Process water is primarily used in the ore mining and dress-
ing  industry in wet screening, gravity separation processes
(tabling, jigging), heavy-media separation,  flotation  unit
processes  (as carrier water), and leaching solutions; it is
also used as mining water for dredging and hydraulic mining.
Mine water is often pumped from a mine and discharged,  but,
at many operations, mine water is used as part of processing
water  at a nearby mill.  Water is consumed by being trapped
in the intersitual voids of the product and tailings and  by
evaporation.

Miscellaneous Water

These   water   uses  include  dust  control  (primarily  at
crushers), truck and vehicle washing, drilling fluids, floor
washing and cleanup, and domestic and  sanitary  uses.   The
resultant  streams  are  either  not  contaminated  or  only
slightly contaminated with wastes.  The general practice  is
to  discharge  such  streams  without  treatment  or through
leaching fields or septic systems.  Often, these streams are
combined with process water prior to treatment or discharged
directly to tailing ponds.  Water used at crushers for  dust
control is usually of low volume and is either evaporated or
adsorbed on the ore.
                            182

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PROCESS WASTE CHARACTERISTICS BY ORE CATEGORY

Iron Ore

The  quality  and quantity of water discharged from open-pit
and underground iron  mining  operations  and  beneficiation
facilities  vary  from  operation to operation.  In general,
the quality of the water in mines is highly dependent on the
deposit mined and the  substrata  through  which  the  water
flows prior to entry into the mine.

Sources of Waste.   The main sources of waste in iron mining
and ore processing are:

    (1)  Wastewater from the mine itself.  This may  consist
         of  ground  water which seeps into the mine, under-
         ground aquifers intersected by the  mine,  or  pre-
         cipitation and runoff which enter from the surface.

    (2)  Process water, including spillage from  thickeners,
         lubricants, and flotation agents.

    (3)  Water used in the transport of tailings,  slurries,
         etc.,  which,  because  of the volume or impurities
         involved, cannot be reused in processing or  trans-
         port without additional treatment.

In  most  cases,  the last category constitutes the greatest
amount of waste.

Waste Loads and Variability.   Waste loads  from  mines  and
processing  operations  are often quite different, and there
is variability on a  day-to-day  and  seasonal  basis,  both
within  an operation and between operations.  At times, mine
water is used as process feed water, and variability in  its
quality is reflected in the process water discharge.

Nature  of Iron Mining Wastes.   Mine water can generally be
classified as a "clear water," even though  it  may  contain
large  amounts of suspended solids.  The water may, however,
contain significant quantities of dissolved  materials.   if
the  substrata  are  high in soluble material (such as iron,
manganese, chloride, sulfate, or carbonate), the water  will
most likely be high in these components.  Because rain water
and  ground water are usually slightly acidic, there will be
a tendency to dissolve metals  unless  carbonates  or  other
buffers are present.
                            183

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Some   turbidity   may  result  from  fine  rock  particles,
generated in blasting, crushing, loading, and hauling.  This
"rock flour" will depend on the methods used in a particular
mine and on the nature of the ore.

Nitrogen-based blasting agents have  been  implicated  as  a
source  of  nitrogen  in mine water.  The occurrence of this
element (as ammonia, nitrite, or nitrate) would be  expected
to  be  highly  variable and its concentration a function of
both the residual blasting material and the volume of  dilu-
tion water present.

These  effluents in the iron mining operations are generally
unrelated  to  production  quantities  from  the  operation.
Therefore,  waste  loadings  are  expressed in concentration
rather than units of production.  Constituents which may  be
present in the mine water are:

    (1)   Suspended solids resulting from blasting, crushing,
         and transporting ore;  finely  pulverized  minerals
         may be a constituent of these suspended solids.

    (2)   Oils and greases resulting from spills and leakages
         from material handling equipment.

    (3)   Natural hardness and alkalinity associated with the
         host rock or overburden.

    (U)   Natural levels of salts and nutrients in the intru-
         sive water.

    (5)   Residual quantities of unburned or partially burned
         explosives.

Processing Wastes.   The processing cf ore from the mine may
result in the presence of a number of waste materials in the
wastewater.  Some of these are derived from the ore  itself,
and  others  are  added during processing.  Still others are
not intentionally added but  are  inadvertent  and  inherent
contributions.

Dissolved and suspended solids are contributed by the ore to
water  used  in  transport and processing.  Included in this
are metals.  The nature and quantity of these are  dependent
on  the  nature  of  the  water,  the ore, and the length of
contact.

During processing, various flotation agents,  acids,  clays,
and other substances may be added and thereby become consti-
                            184

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tuents  of  wastewater.   Oil  and grease from machinery and
equipment may also contaminate the water.

Inadvertent additions include metals  (such  as  zinc)   from
buildings and machinery, runoff from the plant area and from
stockpiles which may contain dissolved and suspended solids,
and spills of various substances.

Sanitary  sewage  from  employees  and  domestic sewage from
washrooms, lunchrooms, and other areas is  usually  disposed
of  separately  from  process  and  transport wastes through
municipal or drainfield systems.  Even when not, it would be
expected to constitute a minor part of the load.

The principal characteristics of the waste stream  from  the
mill operations are:

    (1)   Loadings of 10 to 50 percent solids (tailings).

    (2)   Unseparated minerals associated with the tailings.

    (3)    Fine  particles  of minerals (particularly, if the
         thickener overflow is not recirculated).

    (4)   Excess flotation reagents which are not  associated
         with the iron concentrate.

    (5)   Any spills of reagents which occur in the mill.

One aspect of mill waste which has been poorly characterized
from  an  environmental-effect  standpoint  is the excess of
flotation reagents.  Unfortunately, it is very difficult  to
detect   analytically   the  presence  of  these  reagents—
particularly, the organics.  COD, TOC, and surfactant  tests
may   give  some  indication  of  the  presence  of  organic
reagents, but no definitive information is related by  these
parameters.

The substances present in mine-water discharges are given in
Table  v-1;  those  present  in process-water discharges are
given in Table V-2.  These values are historically represen-
tative of what is present before and after discharge to  the
receiving  water.   When  mine  water  is used as processing
water, its characteristics often cannot  be  separated  from
those of the processing water.

As  part of this study, a number of mining and beneficiation
operations were visited and sampled.   The  results  of  the
sample  analyses  show  certain potential problem areas with
                            185

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    TABLE V-1. HISTORICAL CONSTITUENTS OF IRON-MINE DISCHARGES
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cr
Cu
Fe
Pb
Mg
HO
Ni
Na
Mn
Zn
Chloride
Cyanide
CONCENTRATION (mg/£)
BEFORE TREATMENT
MIN
1.000
140.0
0.200
5.00"
1.800
0.003
0.003
0.001
0.001
0.060
0.001
0.020
0.002
0.003
0.023
0.001
0.001
1.000
0.010
MAX
5,000.0
1,880.0
36.0
8.40"
9.000
0.350
256.0
0.010
1.000
178.0
0.100
118.0
2.000
0.100
15.0
18.0
8.0
120.0
0.02
AVG
371.51
436.18
6.470
7.45*
4.511
0.066
85.39
0.007
0.167
13.3
0.018
39.35
1.001
0.024
7.511
2.462
1.869
27.143
0.013
NO.
19
17
10
18
9
7
3
9
12
14
9
3
2
6
2
14
9
14
4
AFTER TREATMENT
MIN
1.000
100.0
0.026
6.800*
0.400
0.007
0.002
0.010
0.005
0.008
0.008
0.008
-
0.010
.
0.001
0.010
0.900
0.005
MAX
30.0
1,090.0
42.0
8.500*
20.400
0.350
0.158
0.010
0.370
2.100
0.100
0.029
•
0.075
.
6.900
0.340
180.00
0.020
AVG
10.693
390.10
12.116
7.652*
4.313
0.131
0.045
0.010
0.120
0.446
0.023
0.017
•
0.023
.
1.720
0.185
33.225
0.011
NO.
27
20
14
21
16
9
4
6
10
11
8
3
-
5
.
11
5
20
4
•Value in pH units
     TABLE V-2. HISTORICAL CONSTITUENTS OF WASTEWATER FROM
               IRON-ORE PROCESSING
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cu
Fe
Pb
Ni
Mn
Zn
Chloride
Cyanide
CONCENTRATION (mg/%)
BEFORE TREATMENT
MIN
1.20
0.500
0.200
5.000*
0.030
0.030
55.0
•
0.200
0.100
0.010
0.007
0.006
1.000

MAX
9,999.0
356.0
36.0
8.300*
40.400
5.000
2SO.O
•
10.0
5.0
0.060
20.0
10.0
110.0

AVG
1,894.8
207.1
16.986
7.187*
14.229
0.994
120.0
•
2.568
3.367
0.023
2.772
3.013
22.145

NO.
11
10
7
12
8
6
3
•
9
3
3
9
5
11

AFTER TREATMENT
MIN
0.400
0.300
0.200
6.000*
0.100
0.009
82.0
0.010
0.050
0.045
0.010
0.016
0.010
0.350
0.008
MAX
200.0
1,090.0
90.0
8.300*
90.0
0.270
181.0
0.450
1.610
0.250
0.200
2.100
0.115
180.0
0.020
AVG
25.133
393.27
19.518
7.259*
12.0
0.107
131.5
0.230
0.453
0.111
0.087
0.529
0.056
42.875
0.013
NO.
15
16
12
16
13
8
2
2
10
4
3
10
4
15
4
* Value in pH units
                           186

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       TABLE V-3. CHEMICAL COMPOSITIONS OF SAMPLED MINE WATERS

PARAMETER


PH
Alkalinity
COD
TSS
TDS
Conductlfity
Total Fa
Diuolvad ft
Mn
Sulfatt
CONCENTRATION (mg/t I inWASTEWATER FROM MINE


§
S
7.3"
204
27.4
2
455
440*
0.04
<0.02
0.21
as


i
u
Z
I
7.2'
-
48.2
2
SOS
400*
<0.02
<0.02
<0.02
175


S
UJ
i
7.5*
_
9.2
6
609
700*
<0.02
<002
0.40
215
f
|
fi
8
7.2*
176
4.5
30
246
310*
<0.02
<0.02
<0.02
45
-
1
£
8
7.4*
_
1.0
<1
281
320*
<0.02
<0.02
<0.02
28
}
1
2
£
8
7.4*
—
18.3
21
169
216*
<0.02
<0.02
0.059
21

|
i
8
7.6*
211
22.8
20
271
340*
0.18
<0.02
<0.02
26


|
Ul
Z
I
8.4*
218
18
6
1,302
1,950*
4.60
<0.02
3.20
162
]
m
I
8
7.1*
37.4
4.5
10
118
110*
2.80
<0.02
0.026
11.2
ft
1
I
S
7.2*
118
9.0
2
440
560*
1.30
0.04
0.054
33.2


8
in
i
8J*
181
27.5
12
308
342*
0.30
0.02
0.68
36.7


e
*•
i
7.9*
•8.0
<10
48
1.290
1.126*
1.10
0.08
<0.02
780
1
1
§
1
7.64*
161
18.7
13.26
499.6
666.8*
OJ6
0.027
OJ9
134
Vtlutin pH uniti
V/alu* in micromhoi/cnrt
     TABLE V-4. CHEMICAL COMPOSITIONS OF SAMPLED MILL WATERS



PARAMETER



PH
Alkalinity
COD
TSS*«
IDS
Conductance
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION 
-------
respect to the discharge of pollutants.   Summaries  of  the
major  chemical  parameters in raw wastes from mine and mill
water, measured as part of site visits, are given in  Tables
V-3  and  V-U.   The  basic  waste  characteristics,  on the
average,  are  very  similar  for  both  mines  and   mills.
Elevated  concentrations  of  particular  parameters tend to
associate with a particular mining area or  ore  body.   For
example,  the  dissolved  iron and manganese tend to be much
higher in Michigan ores than in ores from the  mining  areas
of the Mesabi Range in Minnesota.

In the beneficiation of iron-containing minerals, as much as
27.2 cubic meters of water per metric ton (7,300 gallons per
long  ton)  and  as  little as 3.4 cubic meters of water per
metric ton (900 gallons per long ton) of concentrate may  be
used.   The  average  amount  of water per metric ton of ore
produced is approximately 11.8 cubic meters  (3,200  gallons
per  long  ton),   Most  processing  water  in beneficiation
operations is  recycled  to  some  extent.   The  amount  of
recycle  is  dependent  on  the  type  of processing and the
amount of water that is  included  in  the  overall  recycle
system in the mill.

Mills that employ flotation techniques currently discharge a
percentage  of  their  water  to  keep  the concentration of
soluble salts from increasing to excessive levels.   Soluble
salts—especially,   those   of  the  multivalent  ions—are
deleterious to  the  flotation  process,  causing  excessive
reagent  use  and  loss  of  recoverable  iron.   Even these
operations currently recycle at least 80  percent  of  their
water.

Mills   using  physical  methods  of  separation  (magnetic,
washing, jigging, heavy media, spirals,  and  cyclones)  can
and  do recycle greater than 80 percent of their water.  The
amount of water discharged from these operations  is  solely
dependent  on  how  much  water  drains and accumulates into
their impoundment systems.

Typical mining operations take the water that accumulates in
the mine and pump it either to discharge  or  to  a  tailing
basin, where a portion is recycled in the processing operat-
tion.   Mine  water is generally settled to remove suspended
matter prior to discharge or before use in plant  processes.
A  typical  flow  scheme  for the treatment of mine water is
given in Figure V-l.

Process operations generally  recycle  high  percentages  of
their water.  Water in the plant process is used to wash and
                            188

-------
      Figure V-1. FLOW SCHEME FOR TREATMENT OF MINE WATER
                          MINE WATER
                      SEDIMENTATION BASIN
SETTLED
SOLIDS
i p
TO
WASTE
CLAR
EFFL
i
i
TO REC
WAI
IFIED
KENT
1
r
:EIVING
FERS

TO PROCESS
WATER
  Figure V-2. WATER FLOW SCHEME IN A TYPICAL MILLING OPERATION
      TO
STOCKPILE
-*


WATER
i
PROCESS
PRODUCT


p
PROCESS PLANT

COAGULANT








PROCESS
TAILING
1 RECYCLE
_ /""SEDIMENTATION ^^
— ^"x.^^ BASIN ^^^s

SETT
SOL
TO
WASTE
LED
IDS
CLAF
EFFL
1
IFIED
UENT
1 .
(80-97%)
\ '
TO RECEIVING
WATERS
                        189

-------
transport   the   ore  through  grinding  processes.   After
separation of the concentrate, the tailings  are  discharged
to  a  tailing  pond,  where  the coarse and fine waste rock
particles settle  (Figure V-2).  Clarified water is  returned
to   be  used  in  further  processing,  and  a  portion  is
discharged to receiving waters.

Plants or mines that have zero discharge have not been  dis-
cussed  in  this  section  because  they  discharge no waste
materials.  It should be pointed out,  however,  that  every
plant  operation  loses water to some degree and has to make
up this water loss to maintain a water  balance.   The  main
sources  of water loss are losses to within the concentrated
product,  evaporation  and  percolation  of  water   through
impoundment  structures,  loss of water to the tailings, and
evaporation or water loss during processing.

Process Descriptions

The  following  subsections  discuss  particular  processing
operations  to  demonstrate  how  water  is  utilized during
different ore processing, the water flow within each system,
and the waste loads generated.

Mine and Mill 1105.    Mine  and  mill  1105  is  a  typical
taconite  operation.   Open-pit  mines  associated  with the
operation produce an effluent, and the mill operates with  a
closed water system.

Crude  magnetic  taconite  is  mined,  mainly from the lower
cherty member of the Minnesota Biwabik  formation,  by  con-
ventional open-pit methods and then milled to produce a fine
magnetite.  The fine magnetite from the mill is agglomerated
in a grate-kiln system to produce approximately 2.61 million
metric  tons  (2.6  million  long  tons)   of  oxide  pellets
annually for blast-furnace feed.

The mine, mill, and pelletizing plant are located on a large
site controlled by the operating company, with 8094 hectares
(20,000 acres)  utilized at present.  An initial tailing pond
of 405 hectares  (1000 acres)   has  been  filled.   A  second
1,619-hectare (U,000-acre) pond is now being used.

An  open system is used in mine dewatering.  A sketch of the
system with flow rates is shown  in  Figure  V-3.   Settling
basins are used to contain the water before it is discharged
to two lakes.
                            190

-------
          Figure V-3. WATER BALANCE FOR MINE/MILL 1105  (SEPTEMBER  1974)
                 3.4 m/min (900 gpm)
                  (INTERMITTENT)
          1
(   SETTLING BASINSN
    2 PUMPS
 ei1.4m3/min
 (3,000 gpm) EA.
(INTERMITTENT)
                                       MINE OEWATERING
                                        17 to 32 m3/min (4.500 to 8.500 gpml
                             MAX. 8.23 m3/min (2.200 gpm)
                             (INTERMITTENT)
                                               15.1 m3/min (4,000 gpm)
                                               204 m3/min (54,000 gpm)
                                                             189 m3/min
                                                            (50.000 gpm)
                                               1S.1 m3/min 14,000 gpm)
                                                       15.1 m3/mln (4,000 gpm)
                                                                 1
                                                             SETTLING BASINS
NS>v
                                                                                     17 to32m3/min
                                                                                     (4,500 to 8,500 gpm)
                                           191

-------
The mill water system is a closed loop.  Plant processes use
20»*  cubic meters per minute (78 mgd), with 189 cubic meters
per minute (72 mgd) returned from the 91.1-meter  (300-foot)
diameter  tailing  thickener  overflew and 15.1 cubic meters
per minute (5.7 mgd)  returned  from  the  tailing  pond  or
basin.   The tailing thickener receives waste or tailings in
a slurry from the concentrate  pellet  plant.   A  nontoxic,
anionic  polyacrylamide flocculant is added to the thickener
to assist in settling out solids.  Tailing thickener  under-
flow is pumped to the tailing basin.

Rotary  drilling  machines  are  used in the mine to prepare
blast holes for the ammonium  nitrate-fuel  oil  (ANFO)   and
metallized  slurry  blasting  agents.   Electric shovels are
used  to  load  the   broken   ore   into   100-ton-capacity
diesel/electric trucks for haulage to the primary crusher.

The  1.52-meter   (60-inch)  primary crusher is housed in the
pit and reduces the ore to a size of less than 0.15 meter (6
inches).  From the crusherr coarse  ore  is  conveyed  to  a
storage building.

Figure  V-4  is  a flowsheet showing the physical processing
used in the mill.  Coarse ore assaying 22  percent  magnetic
iron  is  reclaimed  from the storage building and ground to
lU-mesh size in the primary, air-swept dry grinding  system.
Broken  ore  is removed from the mill by a heated air stream
and is air classified and  screened.   The  coarse  fraction
goes to a vertical classifier, and the fine fraction goes to
two  cyclone classifiers.  From the cyclone classifiers, the
fine product goes to a wet cobber to recover  the  magnetics
for  the  secondary grinding circuit.  Coarse product of the
air classifiers is screened, and the oversize is returned to
the primary mill for further grinding.  Undersize  from  the
classifiers  is  separated  magnetically  to  produce  a dry
cobber concentrate, a dry tailing,  and  a  weakly  magnetic
material  which is recycled for further grinding and concen-
tration.  About 37 percent of the crude weight  is  rejected
in the primary circuit.

Dust collected in sweeping the dry mill is pulped with water
and fed to a double-drum wet magnetic separator to produce a
final  tailing  and  a  wet  concentrate for grinding in the
secondary mills.

Ball mills are used in the secondary wet grinding section to
reduce the size of the dry cobber and wet dust concentrates.
Slurry from the ball mills is sized in wet cyclones.   Over-
size from the cyclones is returned to the ball mill.  Under-
                            192

-------
Figure V-4. CONCENTRATOR FLOWSHEET FOR MILL 1105
        DRY SEMIAUTOGENOUS GRINDING MILLS
             VERTICAL DRY CLASSIFIER
         OVERSIZE
UNDERSIZE

    t
                           CYCLONE CLASSIFIER
                        OVERSIZE
         SCREEN
        UNDERSIZE
           t
                                         WET MAGNETIC COBBING
   OVERSIZE UNDERSIZE

                t
            I
       CONCENTRATE
                                                         I
   TAIL
            DRY MAGNETIC ROUGHER
            TAIL
                        CONCENTRATE
    DRY MAGNETIC SCREENING
 MIDDLING
                                     WET SECONDARY
                                     GRINDING MILLS
                                          ±
                                     HYDROCYCLONES
                                    UNDERSIZE  OVERSIZE
            TAIL
        i
                                   HYDROSEPARATION
                                   I
                              CONCENTRATE
                TAIL
                  I	
                      WET FINISHER MAGNETIC SEPARATION
                      I
                 CONCENTRATE
                  TAIL
                     TO
                   PELLET
                    PLANT
            TAILING THICKENER
                                      UNDERFLOW
            TO
          WASTE
                                          I
                                                      OVERFLOW
           TO
      TAILING POND
                          f
     TO
REUSE WATER
                       193

-------
size  ore from the cyclones is pumped to hydroseparators.  A
rising current of water is used  in  the  hydroseparator  to
overflow a fine silica tailing.  Hydroseparator underflow is
sent   to   finisher   magnetic  separators.   The  finisher
separators upgrade the hydroseparator underflow and  produce
a  fine  tailing  or discard.  Finisher magnetic concentrate
can be further upgraded, if necessary, by fine screening and
regrinding  and  then  reconcentrating  the  screen-oversize
material.

The final concentrate is thickened and dewatered to about 10
percent, moisture  prior  to  the formation of fgreen balls'
from this material.  A bentonite binder is blended with  the
concentrate  before balling in drums.  The balling drums are
in closed circuit with screens to return undersize  material
to the drum and to control the green ball size.

Fines  are  again  removed  from the green balls on a roller
feeder before they enter a traveling grate.  These fines are
recirculated to a balling drum or to the pellet plant feed.

Green balls are dried in an updraft and downdraft section of
the grate.  Dried balls then pass through a preheat  section
on  the  grate.   The  magnetite  begins to oxidize, and the
balls strengthen while passing through the preheat section.

Balls go directly from the grate to a kiln, where  they  are
baked  at  1315  degrees  Celsius  (2400 degrees Fahrenheit)
before they are discharged to a cooler, where  oxidation  of
the  pellets is completed and pellet temperature is reduced.
The finished pellets contain 67 percent iron and  5  percent
silica  and  are  transported for lake shipment to the steel
industry.

Mine and Mill 1104.   This mine/mill complex  is  a  typical
natural   ore   (one   not   requiring   fine  grinding  for
concentration)   operation,  with  the  mine  and  mill  both
producing  effluents.   Physical  processes  are used in the
mill to remove waste material  from  the  iron.   The  plant
processes a hematite/1imonite/goethite ore and was placed in
operation  at  the  start  of the 1962 shipping season.  The
operation is seasonal for 175 days per year, from  the  last
week in April to about the middle of October.

Mine  water  from one of the two active pits is pumped to an
abandoned mine (settling basin) and overflows to a river  at
a maximum rate of 7,086 cubic meters per day (1,872,000 gpd)
and  at  an  average  rate  of  5,826  cubic  meters per day
(1,539,000 gpd) per day at Discharge  No  1.   Mill  process
                            194

-------
water,  mine  drainage from the other pit, and fine tailings
from the mill are pumped to a 105-hectare (260-acre) tailing
basin.  Process water is recycled from the basin at  a  rate
of  45  cubic  meters  (12,000  gallons)  per minute.  Excess
water  from  the  tailing  basin  is  siphoned  to  a   lake
intermittently  at  an  average  rate  of 3,717 cubic meters
(981,900 gallons) per day at Discharge No. 2.  Table V-5  is
a  compilation  of  the  chemical  characteristics and waste
loads present in mine water (Discharge No.  1—concentration
only) and combined mine and mill process effluent.

Mining  is  carried  out  by  conventional open-pit methods.
Ammonium nitrate explosives are used in  blasting.   Shovels
load the ore into trucks for transport to the plant.

At the mill, the ore, averaging 37 percent iron, is fed to a
preparation  section for screening, crushing, and scrubbing.
A plant flowsheet is shown in Figure V-5.

Reversible conveyors permit rock coarser  than  10.2  centi-
meters  (4  inches)  from the first stage of screening to be
removed as a reject  and  stockpiled  or  processed  further
depending  on  the  quality of the oversize material.  Plant
feed is processed in a  crusher/screen  circuit  to  produce
fractions  which  are 3.2 cm by 0.64 cm  (1.25 inches by 0.25
inch) and less than 0.64 cm (0.25 inch).  The material which
is 3.2 cm by 0.64 cm  (1.25 inches by 0.25 inch)  goes  to  a
heavy-media  surge  pile.   The  fraction which is less than
0.64 cm (0.25 inch) after classification to remove  tailings
which are less than 48 mesh is sent to a jig surge pile.

Material  from  the  heavy-media  surge  pile  is split into
fractions which are 3.2 cm by 1.6 cm  (1.25  inches  x  0.63
inch)  and  1.6 cm x 0.64 cm  (0.63 inch by 0.25 inch).  Both
fractions  go  to  identical  sink/float  treatment   in   a
ferrosilicon suspension.  Float rejects or tailings from the
heavy  suspension  treatment  are  trucked  to  a stockpile.
Concentrates go directly to a railroad loading pocket.   The
ferrosilicon  medium  is  recovered  by magnetic separation.
The magnetic medium is recycled to the process.  Nonmagnetic
slimes go to the tailing pond.  The material which  is  less
than  0.64 cm  (0.25 inch) but greater than 48 mesh goes from
the surge pile to jigs, where pulsating  water  is  used  to
separate  the  concentrate  and  tailing.   Concentrates are
dewatered before shipment, and water from this operation  is
recycled   in  the  plant.   Jig  tailings  are  sent  to  a
dewatering  classifier.   Sands  from  the  classifier   are
trucked  to  a reject pile.  Overflow from the classifier is
pumped to the tailing basin.
                             195

-------
    TABLE V-5. CHEMICAL ANALYSIS OF DISCHARGE 1 (MINE WATER) AND
             DISCHARGE 2 (MINE AND MILL WATER) AT MINE/MILL 1104,
             INCLUDING WASTE LOADING FOR DISCHARGE 2
PARAMETER
PH
TSS
TDS
Total Fe
Dissolved Fe
Mn
CONCENTRATION (mg/£ ) IN WASTEWATER
DISCHARGE 1
6.7 •
6
263
<0.02
<0.02
<0.02
DISCHARGE 2
7.3*
6
210
<0.02
<0.02
<0.02
RAW WASTE LOAD
g/metric ton
—
3.8
132
< 0.01 3
< 0.01 3
<0.013
Ib/short ton
—
0.0074
0.26
<0.00003
<0.00003
<0.00003
*Value in pH units
                            196

-------
                Figure V-5.  FLOWSHEET FOR  MILL 1104 (HEAVY-MEDIA PLANT)
                                            I  MINING  I


                                            CRUDE ORE
                                             (37% IRON)
            DOUBLE-DECK
              SCREEN
      > 16.2 cm
       (>• in.)
 10.2 to 16.2 c
  14 to 6 in.)

_	I
                 |<24*F.


           TO ROCK-REJECTS
             STOCKPILE
1

HEAVY-MEDIA
SURGE PILE
                      O.M to 3.2 cm (0 25 to 1.25 in.)
.26 in.) 	

I
^DOUBLEDECK
SCREEN
1 3»% ft 42% ft
CONE
CRUSHER
t
SCREEN
> 0.64 cm <
!>0,25 in.l |<
1
<0,64cml<0



1
0.64cm
0.26 in.)
\

<0
«o

\
54 cm
25 in.)
                                                                    RAKE CLASSIFIER
                                                             48m«h to
                                                          10.64 cm 10.26 in.)
                                                                 44* Ft
                                                            OVERFLOW
                                                                                 TO
                                                                             TAILING POND
HEAVY-
MEDIA
SEPARATOR
(SP.QR. • 3.10)
14% F«
\

HEAVY-
MEDIA
SEPARATOR
ISP.OR. - 2.901
H%Ft

1«*F.
\

60% ft

                                                           STAR FEEDERS
                                                             SPLITTER
                                                                                               WATER
FLOAT REJECTS   FLOAT REJECTS   CONCENTRATE

     I      -         I
                                                         PULP DISTRIBUTORS
                                                                                         12 JIGS
                                                                          CONCENTRATES


                                                                                I SI* Ft
                                                                                 REJECTS


                                                                                     36% Ft
                                                                           DEWATERING
                                                                            CLASSIFIER
                                                                                DEWATERING
                                                                                CLASSIFIER
 TO FLOAT-REJECTS
    STOCKPILE
                TO
              TAILING
               POND
                             64.76* Ft
                                                                    CONCENTRATE
                                                                       FINE
                                                                    CONCENTRATE
                                                                          68.6* Ft
                                                                                   OVERFLOW OVERFLOW
                         1"
         RECYCLE TO
        HEAVY-MEDIA
         SEPARATORS
CONCENTRATES TO TRANSPORTATION
      I- 32* OF CRUDE ORE)
 WATER
  TO
RECYCLE
  TO
TAILING
 POND
                                                                                                          REJECTS
   \

TO FLOAT-
 REJECTS
STOCKPILE
                                             197

-------
Concentrates produced in the plant are shipped by  rail  and
boat  to  the lower Great Lakes.  The 58-percent-iron heavy-
media concentrate serves as  blast-furnace  feed.   The  58-
percent-iron  jig concentrate is later sintered at the steel
plant before entering the blast furnace.

Mine and Mill 1108.   This mine/mill complex is  located  in
Northern Michigan.  The ore body consists of hematite (major
economic material), magnetite, martite, quartz, jasper, iron
silicates,  and  minor  secondary  carbonates.   All  of the
constituents   appear   in   the   tailing   deposit.    The
concentration  plant  processes  approximately 21,000 metric
tons (20,700 long tons) per day  of  low-grade  hematite  at
35.5 percent iron to produce approximately 9,850 metric tons
(9,700  long  tons)  per  day  of  concentrated  ore at 65.5
percent iron.  The  remaining  11,200  metric  tons  (12,346
short  tons),  at  approximately  10 percent total iron, are
discharged to the tailing basin.

Mine water is currently pumped from the actively  mined  pit
and  discharged  directly.  The chemical constituents of the
discharged water are given in Table V-6.

Water in the concentration process is utilized at a rate  of
11U  cubic meters  (30,000 gallons)  per minute.  Ore is first
ground to a fine state (80 percent less than 325  mesh)   and
the  slime materials removed by wet cycloning.  A simplified
flow scheme is included in Figure  V-6.   Subsequently,  the
concentrated  ore  is  floated  using tall oil - fatty acid.
The flotation underflows are discharged to a tailing stream,
which is discharged directly  to  a  385-hectare  (950-acre)
tailing  basin.   Approximately 80 percent of the water from
the tailing pond is returned to the concentrating  plant  as
reuse  water   (untreated).   The  remaining  20  percent  is
discharged, after treatment, to a local  creek.   This  dis-
charged  wastewater  is first treated with alum, then with a
long-chain polymer to promote flocculation.  It then  passes
to  a  8.5-hectare  (21-acre)   pond,  where  the flocculated
particles settle.  The concentration of chemical  parameters
and  the  waste loading in this discharge are given in Table
V-7.

Copper Ore

Frequently, discharged wastes encountered in the copper  ore
mining  and  dressing  industry  include  waste streams from
mining,  leaching,  and  milling  processes.   These   waste
streams  are  often  combined  for  use  as process water or
treated together for discharge,  other wastes encountered in
                              198

-------
TABLE V-6. CHEMICAL CHARACTERISTICS OF DISCHARGE WATER
          FROM MINE 1108
PARAMETER
PH
Alkalinity
COD
TSS
IDS
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION
(mg/je )
7.2*
118
9.0
2
440
1.3
0.04
0.054
33.2
       •Value in pH units
                     199

-------
Figure V-6. SIMPLIFIED CONCENTRATION  FLOWSHEET FOR MINE/MILL 1108
     FATTY-ACID
    CONDITIONER
      WATER
     BINDING
    MATERIAL
                                  MINING
                                 CRUDE ORE
                                                             9,3 m3/min (5.5 eft)
                                      24.500 metric toni
                                      (20,700 long tons)
                                      per day
                                 CRUSHING
                                    AND
                                 GRINDING
                                    I
                              HYDROSCILLATORS
                                 CYCLONES
                                    i
CONDITIONING
                                 FLOTATION
                                    i
                                THICKENER
    i
                                 VACUUM
                                FILTRATION
                                 PRIMARY
                               CONCENTRATE
                                    i
                                REGRINDING.
                           FLOTATION, THICKENING,
                              AND FILTRATION
                                    I
                                SECONDARY
                               CONCENTRATE
                                   i
PELLETIZING
 OPERATION
                                 PELLETS
                                    t
                               TO STOCKPILE
                                                         17 m3/min (10 cfs)
                                                        41.6 m3/min (24.5 cfi)
                                                        14.4 m3/min (8.5 eft)
                          17 m3/min (10 cf»)
0.85 m3/min (0.5 cf»)
                                       8.1% SOLIDS

                             100 m3/min (59 cfs) I


                                      TO TAILINGS
                                    200

-------
    TABLE V-7. CHARACTERISTICS OF MILL 1108 DISCHARGE WATER
PARAMETER
pH
Alkalinity
COO
TSS
TDS
Tottl Ft
Di»olv.d ft
Mn
SuHili
PROGRAM SAMPLE
CONCENTRATION
(meai IN
WASTE WATER
7.1*
82.0
22.6
10
180
2.06
0.93
0.06
6
WASTE LOAD
In ft/fn*tric ton
lib/inert ton)
PRODUCT
-
213 10.42)
77.4 10. IE)
3.4 10.007)
580 (1.08)
7.06 10.013)
3.2 (0.008)
0.17 (0.0003)
17.2 (0.034)
10-MONTH AVERAGES
AVERAGE
CONCENTRATION
(mo/It)
7.0'
-
-
8.8
-
-
0.78
0.66
-
WASTE LOAD
in g/mttrlc ton
(Ih/ihon ton)
PRODUCT
_
-
-
20.7 (0.040)
-
-
1 .83 (0.0036)
1.S9 (0.0031 1
-
HIGH
CONCENTRATION
(mg/4)
7.9-
-
-
63
-
-
3.60
6.80
-
LOW
CONCENTRATION
Img/D
6.6*
-
-
1
-
-
0.01
0.01
-
•Vilw In pH unlti
                           201

-------
this segment are discharge wastes from copper  smelting  and
refining  facilities, treated sewage effluent, storm drains,
and filter backwash.  The uses of water in copper mining and
milling are summarized below.
           I. Mining:
              a. Cooling
              b. Dust control
              c. Truck washing
              d. Sanitary facilities
              e. Drilling

          II. Hydrometallurgical processes associated with
              mining:  Dump, heap, and in situ leaching
              solutions.

         III. Milling Processes:
              a. Vat leach
                   1. crusher dust control
                   2. Vat leach solution
                   3. Wash solutions
              b. Flotation
                   1. Crusher dust control
                   2. Carrier water for flotation

Copper Ore Mining.   Most of the domestic copper is mined in
low-grade ore bodies in  the  western  United  States.   All
mining  and  milling activities adjust to the type of copper
mineralization which is encountered.  The principal minerals
exploited may be grouped  as  oxides  or  sulfides  and  are
listed  in  Table V-8.  Porphyry copper deposits account for
90 percent of the domestic copper  ore  production  and  are
mined by either blockcaving or open-pit methods.  The choice
of  method  is  determined  by  the size, configuration, and
depth of the ore body.

Open-pit (undercut) mining accounted for 83 percent  of  the
copper  produced  in  the United States in 1968.  The mining
sequence   includes   drilling,   blasting,   loading,   and
transportation.   Primary drilling involves sinking vertical
or near-vertical blast holes behind the face of an  unbroken
bank.   Secondary drilling is required to break boulders too
large for shovels to handle, or to blast unbroken points  of
rock that project above the digging grade in the shovel pit.
Ore and overburden are loaded by revolving power shovels and
hauled by large trucks  (75 to 175 ton capacity)  or by train.
Ore and waste may be moved by tractor-drawn scrapers or belt
conveyors.    Some  mines  have primary crushers installed in
                            202

-------
 TABLE V-8. PRINCIPAL COPPER MINERALS USED
           IN THE UNITED STATES
MINERAL

Chalcocite
Chalcopyrite
Bornite
Covellite
Enargite
COMPOSITION
OCCURRENCE*
SULFIDES
Cu2S
CuFeS2
Cu5FeS4
CuS
Cu3AsS4
SW, NW, NC,**
SW, NW, **
NW, SW
NW, SW
NW
OXIDES
Chrysocolla
Malachite
Azurite
Cuprite
Tenorite
CuSiO3-H2O
Cu2(OH)2-CO3
Cu3(OH)2-(C03)2
Cu2O
CuO
SW**
SW, NW»*
SW, NW**
SW
SW
NATIVE ELEMENTS
Copper
Cu
NC, SW**
 *SW = Southwest U.S.
 NW= Northwest U.S.
 NC = Northcentral U.S.

''Major minerals
                  203

-------
the pit which  send  crushed  and  semi-sorted  material  by
conveyor to the mill.

In 1968, 445 million metric tons (490 million short tons) of
waste   material   were   discarded  (mostly  from  open-pit
operations) after production of 154 million metric tons  (170
million short tons)  of copper ore.  The cutoff grade of ore,
which designates it as  waste,  is  usually  less  than  0.4
percent copper.  However, oxide mineralization of 0.1 to 0.4
percent  copper  in waste is separated and placed in special
dump areas for leaching of copper by means of sulfuric acid.

Underground mining methods provided 17 percent of  the  U. s.
copper  in  1968.   Deep  deposits have been mined by either
caving or supported stopes.  Caving  methods  include  block
caving  and  sublevel  caving.   For supported stope mining,
installation of systematic ground supports  is  a  necessary
part  of  the  mining  cycle.   In underground mining, solid
waste may be left behind.   More  than  60  percent  of  the
material  produced is discarded as tco low in copper content
or as oxide ore, which does not concentrate economically  by
flotation.

Water  sources  and  Usage.    in the mining of copper ores,
water collected from the mines may originate from subsurface
drainage or infiltration from surface runoff, or from  water
pumped  to the mine when its own resources are insufficient.
A minimal amount of water in mining is needed  for  cooling,
drilling,  dust  control,  truck  washing,  and/or  sanitary
facilities (Figure V-7).  For safety, excess mine water  not
consumed  by  evaporation  must  be  pumped  from the mines.
Table V-9  lists  the  amount  of  mine  water  pumped   from
selected  mines  and the ultimate fate of this wastewater at
surveyed mines.  Open-pit mines pumped 0 to 0.27 cubic meter
per metric ton  (0 to 64.7 gallons  per  short  ton)  of  ore
produced,  while  underground  mines  pumped  0.008 to 3.636
cubic meters per metric ton  (1.91 to 871 gallons  per  short
ton) of ore produced.

solid wastes produced are summarized in Table V-10 as metric
tons   (or  short  tons)  of  waste (actually, overburden and
wastes) per metric ton  (short ten) cf ore produced.   Under-
ground  operations  rarely have waste.  Those mines which do
produce wastes yield relatively small amounts in  comparison
to open-pit mining operations.

                       Wthin  <>Pen-pit  mines  consists  of
                            204

-------
Figure V-7. WASTEWATER FLOWSHEET FOR PLANT 2120-B PIT
 NATURAL DRAINAGE,
    SEEPAGE, AND
       RUNOFF
         1
       MINING
      (DRILLING,
      BLASTING,
        AND
      LOADING)
         ORE-
 16,560,000 metric tons/year
(18,250,000 short tons/year)
TO
MILL
       EXCESS
     MINEWATER
          0.06 m /metric ton
          (14.4 gal/short ton)
                         LIME PRECIPITATION
          0.06 m3/metric ton
          (14.4 gal/short ton)
    DISCHARGED
                    205

-------
TABLE V-9. MINE-WATER PRODUCTION FROM SELECTED MAJOR COPPER-PRODUCING
            MINES AND FATE(S) OF EFFLUENT
MINE
2101
2102
2103
2104
2107
2108
2109
2110
2111
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
TYPE*
OP
UG
OP
OP
UG
OP
OP
OP
OP
OP
OP
UG
OP
UG
OP
UG
UG.OP
UG
OP
OP
OP
MINE-WATER PRODUCTION
m3/metric ton
ore produced
0.270
0.008
N.E.
0.086
N/A
N.E.
N.E.
N.E.
N.E.
0.015
40.5 (avg)t
1.769
0.030
0.886
0.014
0.654
0.486
0.170
0.034
0.075
N.E.
gal/short ton
ore produced
64.7
1.85
N.E.
20.6
N/A
N.E.
N.E.
N.E.
N.E.
3.5
9,715.0(avg)t
424.0
7.1
212.3
3.4
156.7
116.4
40.85
8.1
18.0
N.E.
EFFLUENT FATE(S)
Reuse in Dump Leach
Reuse in Mill and Leach
Mine above Water Table
Reuse in Dump Leach
Reuse in Mill
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Evaporation and Seepage in Mine
Reuse in Mill
Discharged
Reuse in Mill
Reuse in Leaching
Discharged
Reuse in Dump Leach
Reuse in Mill
Discharged
Discharged
Reuse in Dump Leach
Reuse in Mill
Evaporation and Seepage in Mine
      *   OP = open pit; UG = underground.
      t   0 to 81.1 m3/metric ton (0 to 19,432 gal/short ton) ore produced; variable due to seasonal rainfall and
          open-pit operations; average calculated assuming six dry (0) and six wet (81.1-m^/19,432-gal) months.
      N/A - not available
      N.E. - no effluent
                                      206

-------
 TABLE V-10. SUMMARY OF SOLID WASTES PRODUCED BY PLANTS SURVEYED
MILL
2101
2102
2103
2104
2107
2108
2109
2110
2111
2112
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
HAULED WASTE (1973)
metric tons
34,765,038*
19,534,193*
51,903,633
20,075,681*
0(UG)
11,400.238*
24,222,246
104,328
8,545,824
45,360 (UG)
17,938,604
10,886,400*
18,144 (UG)
32 .257 .310*
0(UG)
33.623.553*
82,737 (UG)
33,112300*
0(UG)
88,452,000*
10,886,400 *
15,339.844
short tons
38,321,250*
21,532.400*
57,213,000
22,129,279*
0(UG)
12,566,400*
26,700,000
115,000
9,420,000
50,000 (UG)
19,773,594
12,000,000*
20,000 IUG)
35,557,000*
0(UG)
37,063,000*
91.200IUG)
36.500,000*
0(UG)
97,500,000*
12.000,000*
16,909,000
MILL ORE (1973)
metric tons
7,198,015
7,967.575
13,977,230
7,349,938
N/A
3,662.574
1,567,460
3,712,262
1,480,550
635.040
9,383.475
130,386
471.376
11,466.193
1,211.680
16,656,192
19,935,266
23,342.256
8,059,688
34,745.760
1,970.438
7.912,698
short tons
7,934,320
8,782,600
15,407,000
8,101,784
N/A
3.927.000
1.727,800
4,092.000
1,632.000
700,000
10,343.337
143,723
519,593
12,638.000
1.335.626
18,360,000
21,974,500
25,730,000
8,884,136
38,300,000
2,172,000
8.722,000
RATIO
(WASTE/ORE)
4.83
2.45
3.71
2.73
_
3.20
15.45
0.03
5.77
0.07
1.91
83.6T
0.04
2.81
-
2.02
0.004
1.42
-
2.55
5.53
1.94
* All or • portion leached
* Stripping operation
N/A - Not available
UG • Underground
                               207

-------
particulates.   The sludge is commonly evaporated or settled
in holding ponds.

Waste Water characterization.   The  volume  of  mine  water
pumped  from  mines  was previously summarized in Table V-9.
The chemical characteristics of these waters are  summarized
in   Table   V-ll,   which   includes   the  flow  per  day,
concentration of constituents, and raw-waste load per day.

A portion of the copper industry  (less than 5 percent)   must
contend  with acid mine water produced by the percolation of
natural  water   through   copper   sulfide   mineralization
associated  with deposits of pyrite (FeS^) -  This results in
acid water containing high concentrations of  iron  sulfate.
Acid   iron  sulfate  oxidizes  metal  sulfides  to  release
unusually high concentrations of trace elements in the  mine
water.   The  pH of mine water most often is in the range of
1.0 to  8.5.   In  the  southwestern  U.S.,  mine  water  is
obtained from underground shafts, either in use or abandoned
on  the  property.   This source of water is valuable and is
used for other  copper-producing  processes.   In  contrast,
mine  water  in  Utah,  Montana,  Colorado, Idaho, Oklahoma,
Michigan, Maine, and Tennessee—especially,  in  underground
mines—is  often  unwanted excess, which must be disposed of
if reuse in other processes (such as leaching and flotation)
is not possible.

The primary chemical characteristics of mine waters are:  (1)
occasional presence of pH of 2.0 to 9.5; (2) high  dissolved
solids;  (3)  oils  and  greases;  and (4)  dissolved metals.
Often, mine water is characterized by high sulfate  content,
which  may  be  the  result  of  sulfide-ore oxidation or of
gypsum deposits.  Mine water—particularly, acid mine water-
-may cause the  dissolution  of  metals  such  as  aluminum,
cadmium,  copper, iron, nickel, zinc, and cobalt.  Selenium,
lead,  strontium,  titanium,  and  manganese  appear  to  be
indicators  of  local  mineralogy  and  are  not solubilized
additionally by acid mine water.

Handling of Mine Water.   As shown in Table V-9, mine waters
are pumped to leach and mill operations as  a  water  source
for those processes whenever possible.  However, four of the
operations  surveyed  discharge  all  of their mine water to
surface waters.  Half of these treat the water first by lime
precipitation and settling.  These are discussed in  greater
detail in Section VII.
                            208

-------
                                 TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                            SELECTED COPPER MINES (Sheet 1 of 4)
to
o
10
PARAMETER
Flow
PH I!
MINE 2119
CONCENTRATION
(mg/ii1
42.01 3.5m3/ day
9.64*
TDS 544
TSS 1 8
II
Oil *nd Gruse 11 1
TOC 5
COD I <10
B I 0.2
Cu I 0.5
Co < 0.05
Se I < °-003
Te I < 0.50
As I < 0.07
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
< 0.05
<0.2
3.80
< 0.05
< 0.05
< 0.10
< 0.2
0.13
0.0008
< 0.05
RAW WASTE LOAD PER UNIT ORE MINED
kg/1 000 metric tons
752 m3/1000 metric tons
9.64'
418.5
6.2
0.77
3.85
<7.69
0.154
0.385
< 0.038
< 0.002
< 0.385
< 0.054
< 0.038
< 0.154
2.923
< 00385
< 0.0385
< 0.077
< 0.154
0.10
0.00062
< 0.038
lfa/1000 short tons
180.332 gal/1000 short tons
9.64'
837.0
12.4
1.54
7.70
< 15.38
0.308
0.770
< 0.076
< 0.004
< 0.770
< 0.108
< 0.076
< 0.308
5346
< 0.0770
< 0.0770
< 0.154
< 0.308
030
0.00124
< 0.076
MINE 2120-K
CONCENTRATION
(mg/I.)
27.524.5m3/d8y
3.49*
4.590
4
< 1.0
31
20
0.10
92.0
0.32
N/A
< 0.02
< 0.07
172.0
< 0.5
2.000.0
100
0.33
024
< 0.5
1.35
0.0784
<0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
15.173 m3/1000 metric tons
3.49*
69,630.3
60.7
< 15.17
7.13
303.4
1.52
1,395.6
4.85
N/A
< 3.03
< 1.06
2.609.2
< 7.58
30.340
1.517
5.01
3.64
< 759
20.48
1.19
< 1.52
lb/1000 short tons
3.635.997 gal/1000 short tons
3.49*
139.260.6
121.4
< 30.34
14.26
6062
3.04
2.791 .2
9.7
N/A
< 6.06
< 2.12
5,218.4
< 15.17
60,680
3.034
10.02
7.28
< 15.17
40.96
2.38
< 3.04
            *V»lue in pH units

-------
                                   TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                              SELECTED COPPER MINES (Sheet 2 of 4)
to
M
O
PARAMETER
Flow
PH
TOS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hfl
Pb
MINE 2120-B
CONCENTRATION
(ma/Hi
2.725.2m3/day
6.1*
2,152
40
< 1.0
3.2
<10
0.04
5.30
0.1
I 0.007
<0.2
< 0.07
31.25
< 0.5
6.00
26.5
1.3
0.13
<0.5
1.55
0.0005
<0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
60.08 m3/1000 metric tons
6.1*
129.3
2.4
< 0.060
0.192
< 0.601
0.002
0.318
0.006
0.0004
< 0.01 2
< 0.004
1.88
<0.03
0.361
1.592
0.781
0.008
<0.03
0.093
0.00003
< 0.006
lb/1000 short tons
14,400 gal/1000 short tons
6.1 *
258 b
4.8
<0.12
0.384
< 1.202
0.004
0.636
0.012
0.0008
< 0.024
< 0.008
3.76
< 0.06
0.722
3.184
1.562
0.016
< 0.06
0.186
0.00006
< 0.012
MINE 2120-CE
CONCENTRATION
(mg/£)
27252m3/day
4.7*
454
34
17.0
2.3
<10
0.01
6.2
0.06
0.042
< 0.2
< 0.07
6.17
< 0.5
8.6
1.42
0.034
< 0.05
< 0.5
0.09
0.0005
< 0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric torn
17.685 m3/1000 metric tons
4.7*
8.03
0.60
0.30
0.041
< 0.177
0.0002
0.11
0.0011
0.00074
< 0.0035
< 0.0012
0.109
< 0.009
0.152
0.025
0.0006
< 0.0009
< 0.009
0.002
0.000009
< 0.002
lb/1000 short tons
4.239 gal/1000 short tons
4.7*
16.06
1.2
0.6
0.082
< 0.354
0.0004
0.22
0.0022
0.00148
< 0.007
< 0.0024
0.218
< 0.018
0.304
0.05
0.0012
< 0.0018
< 0X118
0.004
0.000018
< 0.004
             "Value in pH units

-------
                                 TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                            SELECTED COPPER MINES (Sheet 3 of 4)
to
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
HB
Pb
MINE 2121
CONCENTRATION
(mg/il
3,815.3m3/day
7.37*
29.250
69
<1.0
<4.5
819
2.19
0.87
<0.04
< 0.077
0.60
<0.07
2.8
<0.5
<0.1
2.22
<0.02
<0.05
<0.5
119
< 0.0001
<0.1
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
17.28m /1000 metric tons
7.37'
5.053.9
11.9
< 0.173
< 0.778
141.5
0.378
0.150
< 0.007
< 0.013
0.104
< 0.012
0.484
< 0.086
< 0.01 7
0.384
< 0.003
< 0.009
< 0.086
20.6
< 0.00002
< 0.01 7
Ib/ 1000 short tons
4.141 gal/1 000 short tons
7.37-
10,107.8
23.8
< 0.346
< 1.556
283
0.756
0.3
< 0.014
< 0.026
0.208
< 0.024
0.968
< 0.1 72
< 0.034
0.768
< 0.006
< 0.018
< 0.172
41.2
< 0.00004
< 0.034
MINE 21 22
CONCENTRATION
(mg/£)
3.274m3/dav
7.61 •
2.288
2
3
21
38.9
0.11
1.90
130
< 0.003
0.2
<0.07
1.33
<0.2
9.5
0.83
<0.05
0.13
<0.2
0.83
< 0.0001
<05
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
34 m3/10OO metric tons
7.61'
78.69
0.069
0.103
0.722
1.34
0.004
0.065
0.065
< 0.0001
0.007
< 0.002
0.046
< 0.007
0.327
0.029
< 0.002
0.004
< 0.007
0.029
< 0.000003
< 0.017
lb/1000 short tons
8.053 pi/1000 short tons
7.61*
157.38
0.138
0206
1.444
2.68
0.008
0.130
0.130
< 0.0002
0.014
<04O4
0.092
< 0.014
0.654
0.058
< 0.004
0.008
< 0.01 4
0.058
< 0.000006
< 0.034
             •Value in pH units

-------
                                 TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM

                                           SELECTED COPPER MINES (Sheet 4 of 4)
to
M
to
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
H«
Pb
MINE 2123
CONCENTRATION
(mg/8,1
409m3/dav
6.96*
1.350
2
7
10
4
0.07
1.05
< 0.06
0.096
< 0.2
< 0.01
0.1
< 0.5
< 0.1
0.9
< 0.03
< 0.05
< 0.2
0.8
< 0.0001
< 0.5
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric torn
75 m3/1OOO metric tons
6.96*
101
T>.2
0.5
0.75
0.3
0.005
0.08
< 0.005
0.007
< 0.02
< 0.0008
0.008
< 0.04
< 0.008
0.07
< 0.002
< 0.004
< 0.02
0.06
< 0.000008
< 0.04
Ib/ 1000 short tons
18.000 gal/1000 short tons
6.96'
202
0.4
1.0
1.5
0.6
0.01
0.16
< 0.01
0.014
< 0.04
< 0.0016
0.016
< 0.08
< 0.016
0.14
< 0.004
< 0.008
< 0.04
0.12
< 0.000016
< 0.08
                                   •Value in pH units

-------
Process  Description-Hydrometallurqical Extraction Process es
(Mining)

The use of acid leaching processes on low-grade  oxide  ores
and  wastes  produces  a significant amount of cement copper
each year.  All leaching is  performed  west  of  the  Rocky
Mountains.   Figure  V-8 is a flow diagram of the process of
acid leaching.

Leaching of oxide mineralization with dilute  sulfuric  acid
or  acid ferric sulfate may be applied to four situations of
ore.  Dump leaching extracts copper from low-grade  (0.1  to
0.4 percent Cu) waste material derived from open-pit mining.
The cycle of dissolution of oxide mineralization covers many
years.

Most  leach  dumps  are  deposited upon existing topography.
The location of the dumps is selected to assure  impermeable
surfaces  and  to  utilize  the  natural slope of ridges and
valleys for the recovery and collection of pregnant liquors.
In some cases, dumps have been placed on specially  prepared
surfaces.   The  leach  material is generally less than 0.61
meter (2 feet)  in  diameter,  with  many  finer  particles.
However, it may include large boulders.  Billions of tons of
material  are  placed  in dumps that are shaped as truncated
cones.

The leach solution is recycled  from  the  precipitation  or
other  recovery  operation,  along  with  makeup  water  and
sulfuric acid additions (to pH 1.5 to 3.0).  It is pumped to
the top of dumps  and  delivered  by  sprays,  flooding,  or
vertical pipes.  Factors such as climate, surface area, dump
height,  mineralogy,  scale  of operation, and size of leach
material affect the choice of delivery method.   Figure  V-9
summarizes  the  reactions  by  which  copper  minerals  are
dissolved in leaching.

Heap leaching of wastes approaching a better  grade  ore  is
usually done on specially prepared surfaces.  The time cycle
is  measured  in  months.   Copper  is dissolved from porous
oxide  ore.   Very  little  differentiates  heap  from  dump
leaching.   In  the  strictest  sense,  the  pad  is  better
prepared, the volume of material is less, the  concentration
of  acid  is  greater,  acid  is  not regenerated due to the
absence of pyrite in the ore,  and  the  ore  is  of  better
copper grade in heap leaching, compared to dump leaching.

In-situ  leaching techniques are used to recover copper from
shattered or broken ore bodies in place on the surface or in
                            213

-------
   Figure V-8. FLOWSHEET OF HYDROMETALLURGICAL PROCESSES USED IN
              ACID LEACHING AT MINE 2122
TO ATMOSPHERE
                                                TO ATMOSPHERE
     43 fli /nMtric ton
     (10.212 fjIMiort ton)
241 m3/mttric ton
157,834 (d/ihort ton)
EVAPORATION
                                                EVAPORATION


SEWAGE.
SEEPAGE,
RUNOFF.
MINE WATER,
AND WELL WATER
MAI
WA
{
296m3
(70,879 ,d
EUP
FEH
\
— ^"RESERVOIR

27 m3/m«tric ton _
(6.426 pi/short ton)

2.294 m'/iMtric to
rnratric ion (549,873 griAhort ton
/•nort ton)
* **
360 m'/rwtric ton SOL
198.303 9.1/^on ton) MAK£up
I WATER '
107 m3/nwtric ton ,—
k. (26.704 gri/ihort ton)


1
DUMP
LEACH
i i
« PREG
1 SOLI
RREN
UTION
±_
WANT
TION
2,386 m3/m*tric ton
(571,914 ori/Uion ton)
PRECIPITATION ^ (3.727 9rt/*ort tonl roTABLE
PLANT
" WATER
1
CEMENT
COPPER
                                                     0.3 m'/nutrie ton
                                                     (64 gtl/ihort ton)
                                                    TO
                                                 STOCKPILE
                                   214

-------
Figure V-9. REACTIONS BY WHICH COPPER MINERALS ARE DISSOLVED IN
          DUMP, HEAP, OR IN-SITU LEACHING
                                AZURITE

            Cu3(OH)2-(C03>2 + 3H2S04 "** 3CuS04 + 2CO2 + 4H2O

                               MALACHITE

              Cu2(OH)2-CO3 + 2H2SO4 ^     2CuSO4 + C02 + 3H.,O

                             CHRYSOCOLLA

                CuSi03-2H20 + H2S04 ~^ CuSO4 + SiO2 + 3H2O

                                CUPRITE

                      Cu2O + H2SO4 ~^ CuSO4 + Cu + H2O

            Cu2O + H2SO4 + Fe2(SO4)3 7;^   ^ 2CuS04 + H20 + 2FeS04

                             NATIVE COPPER

                     Cu + Fe2(S04)3 ^Z^ CuSO4 + 2FeSO4

                               TENORITE

                      CuO + H2SO4 ^1^! CuS04 + H20

            3CuO + Fe2(S04)3 + 3H2O ^± 3 CuSO4 + 2Fe(OH)3

          4CuO + 4FeSO4 + 6H2O + O2 ^^ 4CuSO4 + 4Fe(OH)3

                              CHALCOCITE

                  Cu2S + Fe2(SO4)3 ^H^ CuS + CuS04 + 2FeS04

                  Cu2S + 2Fe2(S04)3 ~   ^ 2CuS04 + 4F«S04 + S

                               COVELLITE

                    CuS + Fe2(S04)3 ~   ^ CuSO4 + 2FeS04 + S


          Chalcopyrite will slowly dissolve in acid ferric sulfate solutions and also
       will oxidize according to:

                      CuFeS2 + 2O2 ^     CuS + FeSO4;

                         CuS + 202 ^	 CuS04.

       Pyrite oxidizes according to:

                 2FeS2 + 2H20 + 702 ^~^ 2FeSO4 + 2H2SO4-
                               215

-------
old underground workings.  Oxide and sulfide ores of  copper
may  be  recovered over a period of years.  The principle is
•the same as in dump or heap  leaching.   Usually,  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  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 the
precipitation plant at the surface.  The leaching of surface
ore bodies is similar to a heap or dump leach.

Recovery of Copper From Leach Solutions.   Copper  dissolved
in  leach  solutions may be recovered by iron precipitation,
electrowinning, or solvent extraction (liquid ion exchange).
Hydrogen reduction has been employed experimentally.

Copper is often recovered by iron  precipitation  as  cement
copper.   Burned and shredded scrap cans are most often used
as the source of iron, although other iron scrap and  sponge
iron  may also be used.  In 1968, 12 percent of the domestic
mine copper production was in the form of cement copper  re-
covered  by  iron  precipitation.  Examples of iron launders
and cone precipitators are shown in Figures V-10 and V-ll.

The pregnant copper solution (0.5 to 2.2 g/1)  is passed over
shredded or burned iron scrap  and  precipitates  copper  by
replacement according to the reaction:

                    + Fe --- > Cu + FeSOl
Scrap iron of other forms and sponge iron may be employed.

Gravity  iron  launders employ gravity to allow solutions to
trickle over and through iron  scrap.   Spray  water  washes
remove    copper   frequently   from   the   can   surfaces.
Occasionally, solution is introduced from  below  and  flows
upward  through  the  iron  to produce a coarser, but highly
pure, cement copper.  (See Figure V-10.)

Cone precipitators may  be  employed  for  copper  recovery.
Solution  is  injected, through nozzles at the bottom of the
cone, into the shredded iron scrap.  This  injection,  under
pressure,  both  precipitates  copper rapidly and removes it
from the iron surface by the turbulent action.  (See  Figure
V-ll.)
                            216

-------
Figure V-10. TYPICAL DESIGN OF GRAVITY LAUNDER/PRECIPITATION PLANT
            DRAINS
DRAINS
        SIDE VIEW

r

E:
r'

r- -

c::
c::
c:




L_
.
T '
,-j- CELL *

SOLUTION
it" CELL FLOW"*
— i
_:J*

' ** 1
_ _ J "~"

7:t |
2
,-4- 5 *
<•-} 30
_I J* -
1:3*

(il t" *

/ '
TOP VIEW

/

/
\



/
/
\
>

J

)

                                      CANS
                               FLOW
                           END VIEW
                      SOURCE: REFERENCE 25
                           211

-------
       Figure V-11. CUTAWAY DIAGRAM OF CONE PRECIPITATOR
  BARREN
  SOLUTION
     COPPER
SETTLING AND
  COLLECTION
       ZONE
    COPPER DISCHARGE
                                                        SCRAP IRON
DYNAMIC
ACTION
ZONE
                                              COPPER-BEARING
                                              SOLUTION
                       SOURCE: REFERENCE 25
                             218

-------
Precipitated  copper  is  recovered by draining and scooping
out the solids.  Recovery from pregnant solution may  be  60
percent.   The  resulting  cement copper is 85 to 99 percent
pure and is sent to the smelter for further purification.

The barren solution from a precipitation plant  is  recycled
from  a  holding  pond  to  the  top  of the ore body, after
sulfuric acid and makeup water are added, if necessary.

Leach solutions containing greater than 25 to 30  grams  per
liter  of  copper are usually sent to electrowinning facili-
ties.  The cathode copper produced is highly pure  and  does
not require smelting.

Solvent  extraction  of  copper from acid leach solutions by
organic reagents is rapidly becoming an important method  of
recovery.   When pregnant liquors contain less than 30 grams
of copper per liter, the process is most  applicable.   (See
Figure v-12.)

In  solvent  extraction,  a  reagent  with high affinity for
copper and  iron  in  weak  acid  solutions,  and  with  low
affinity  for  other  ions, is carried in an organic medium.
It  is  placed  in  intimate  contact  with   copper   leach
solutions,  where  H+  ions  are  exchanged for Cu(++) ions.
This regenerates the acid, which is recycled  to  the  dump.
The  organic  medium,  together  with  copper,  is sent to a
stripping  cell,  where  acidic  copper  sulfate   solutions
exchange H+ ion for Cu(++).  This regenerates the organic/H*
media  and  passes  copper  to the electrolytic cells, where
impurity-free  copper   (99.98  to  99.99  percent   Cu)   is
electrolytically  deposited  on  cathodes  (electrowinning).
Typically, 3.18 kg (7 Ib)  of acid is used per  O.U54  kg  (1
lb) of copper produced.

Acid  Leach  Solution  Characterization.   Water sources for
heap, dump, and in  situ  leaching  are  often  mine  water,
wells,  springs, or reservoirs.  All acid water is recycled.
Makeup water needs result only from evaporation and seepage;
therefore, the water consumption depends largely on climate.
Table v-12 lists the amount of water  utilized  for  various
operations.

The  buildup  of  iron salts in leach solutions is the worst
Problem encountered in leaching operations.  The pH must  be
maintained below 2.4 to prevent the formation of iron salts,
which  can precipitate in pipelines, on the dump surface, or
within the dump, causing uneven  distribution  of  solution.
This  may also be controlled by the use of settling or hold-
                            219

-------
Figure V-12. DIAGRAM OF SOLVENT EXTRACTION PROCESS FOR RECOVERY OF
          COPPER BY LEACHING OF ORE AND WASTE
                                      MINE
                                      DUMP
                                        I
                                    WEAK CuS04
                                  LEACH SOLUTION
                                       i
                       RAFFINATE
                       RECYCLED "
 SOLVENT
EXTRACTION
  PLANT
                                     Cu++ ON
                                 ORGANIC CARRIER
              RECYCLED
              ORGANIC
            CARRIER 
-------
TABLE V-12.1973 WATER USAGE IN DUMP, HEAP. AND IN-SITU
             LEACHING OPERATIONS
MILL
2101
2103
2104
2107
2108
2110
2116
2118
2120
2122
2123
2124
2125
WATER USAGE (1973)
m3/metric ton
precipitate produced
4,848.6
1,600.0*
1, 335.1 1
967 .8*
1,096.5
1,308.7
N/A
1,185.3
4,264.0
1,973.6
922.2
746.3
626.0
gallons/short ton
precipitate produced
1,162,131
383,490*
320,000t
231,967*
262,800
313,683
N/A
284,108
1,022,000
473,040
221,026
178,876
150,048
       •Estimated from 1972 copper-in-precipitate production and
        assuming precipitates are 85% copper (Source: Copoer • A
        Position Survey, 1973, Reference 26)
      t Production taken from NPDES permit application
      N/A - Production not available; only flow available
                        221

-------
ing ponds, where  the  iron  salts  may  precipitate  before
recycling.

Table  V-13  lists  the  chemical  characteristics of barren
leach solutions at selected plants.  This solution is always
recycled and is almost always totally contained.

Other metals, such  as  iron,  cadmium,  nickel,  manganese,
zinc,  and cobalt, are often found in high concentrations in
leach solutions.  Total and dissolved solids often build  up
so  that  a  bleed is necessary.  A small amount of solution
may be sent to a holding or evaporation pond  to  accomplish
the control of dissolved solids.

Handling   and   Treatment   of  Water.    No  discharge  of
pollutants usually occurs from leaching  operations,  except
for  a  bleed,  which  may  be evaporated in a small, nearby
lagoon.

Process Description - Mill Processing

Vat Leaching.   Vat leaching techniques require crushing and
grinding of high-grade oxide ore  (greater than  0.4  percent
Cu).    (See Figure V-13.)  The crushed ore, either dry or as
a slurry, is placed in lead-lined tanks, where it is leached
with sulfuric acid for approximately four days.  This method
is applicable to nonporous oxide ores and  is  employed  for
better recovery of copper in shorter time periods.

The  pregnant  copper  solution,  as  drawn  off  the tanks,
contains very high concentrations of copper, as well as some
other  metals.   The  copper  may  be  recovered   by   iron
precipitation or by electrowinning.

Water  is utilized in the crusher for dust control, as leach
solution, and as wash water.   The  wash  water  is  low  in
copper  content and must go to iron precipitation for copper
recovery.  Table V-1U summarizes water usage  at  vat  leach
plants.   The  vat  ores are washed and discarded in a dump.
If the sulfide concentration is significant, these ores  may
be floated in the concentrator to recover CuS.

Vat  Leach Water characterization.  Table V-15 summarizes the
chemical  characteristics  of  vat  leach  solutions.  These
solutions are recycled directly.  Makeup  water  is  usually
required  when  there  are evaporative losses from the tanks
and  recovery plants.
                             222

-------
       TABLE V-13. CHEMICAL CHARACTERISTICS OF BARREN HEAP.
                 DUMP, OR IN-SITU ACID LEACH SOLUTIONS
                 (RECYCLED:  NO WASTE LOAD)
PARAMETER
PH
TS
TSS
COD
TOC
Oil and Graau
S
As
B
Cd
Cu
Fe
Pb
Mn
Hfl
Ni
Tl
Sa
Ag
T«
Zn
Sb
Au
Co
Mo
So
Cyanid*
CONCENTRATION (mfl/£> IN LEACH SOLUTION FROM MINE
2120
3.56*
28,148
14
515.8
1.3
<1.0
<0.5
<0.07
0.11
7.74
36.0
2.880.0
0.1
260.0
0.0009
2.40
<1.0
< 0.003
<0.1
1.0
940.0
<0.5
<0.06
3.30
.
.
<0.01
2124
2.82*
47,764
186
1.172
28.0
6.0
<0.5
0.23
0.31
0.092
145.0
6,300.0
<0.1
94.0
0.0012
7.20
<0.1
< 0.040
<0.1
1.0
28.5

-------
                   Figure V-13.  VAT  LEACH  FLOW DIAGRAM  (MILL 2124)
                                           TO ATMOSPHERE
ORE WASH WATER
         166 m3/nwtrie ton
         (39.453 gri/ihort ton)
                                                                                     TO ATMOSPHERE
                                                   34 m3/nwtric ton
                                                   18.166 gd/ihort ton)
                                            EVAPORATION
    CRUSHER
     SLIMES
                             1.1 fn3/nwtric ton
                             (26* gd/ihort ton)
                                                                                      EVAPORATION
    ORE MAKEUP
	  WATER
  290 m3/mtrk ton
 169,506 gri/ihort ton)
                                          TAILS
                                              BARREN SOLUTION-
                                              176 m3/nwtrlc ton
                                              (42,181 gri/ihort ton)
_ COPPER-RICH __
  ELECTROLYTE
 207 m3/m*tric ton
(49,678 gil/thoit ton)
                              31 m /metric ton
                              (7.397 gri/riiort ton)
                                                                                    ELECTROWINNINQ
        166 m3/mrttie ton
        (39,483 grt/riiort ton)
                          28 m3/mttrlc ton
                          (5,620 pl/ihort ton)
                               10 m3/mttric ton
                             (2.411 vl/thort ton)
                                                      WASH
                                                     WATER
    TO MILL
                      TO MILL
                                        TO WASTE
                                        189 m3/imtric ton
                                        (45,176 gal/thort ton)
                                                                          TO
                                                                       STOCKPILE
                                                 TO LEACH DUMPS
                                                AND PRECIPITATION
                                                      PLANT
                                           224

-------
TABLE V-14. WATER USAGE IN VAT LEACHING PROCESS AS A FUNCTION OF
            AMOUNT OF PRODUCT (PRECIPITATE OR CATHODE COPPER)
            PRODUCED
MILL
2102
2116
2124
WATER USAGE (1973)
m3/metric ton
product
133.7
52.4
206.85
gallons/short ton
product
32,040
1 2,568 t
49,578
METHOD OF RECOVERY
Solvent Extraction/Iron
Precipitation*
Electrowinning**
Electrowinning**
     *  Product is cement copper or copper precipitate
     t  No 1973 data were received through surveys. 1972 data from Reference 26
       were used to calculate a value which may be a low estimate of water use.
     "•Product is cathode copper
                             225

-------
TABLE V-15. CHEMICAL CHARACTERISTICS OF VAT-LEACH BARREN ACID
          SOLUTION (RECYCLED: NO WASTE LOAD, MILL 2124)
PARAMETER
PH
TDS
TSS
COD
TOC
Oil and Grease
Al
Cd
Pb
Cr
Cu
Fe
Mn
Ni
V
Tl
Se
Ag
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/P )
1.1*
169,000
515
331
96
1.0
1,540.0
0.42
2.0
17.0
27,800
4,800.0
47.3
1.70
2.50
< 0.03
< 0.003
0.17
11.5
51.0
2.0
< 0.01
         "Value in pH units
                        226

-------
Of the three vat leach   facilities  surveyed,  one  recycles
directly.   Another  employs holding  (evaporative) ponds for
dissolved-iron control.  Still another reuses all the  leach
solution  in  a  smelter process  and  requires new process
water.  Therefore, no discharge results.

Variation Within the Vat Leach  Process.   Ores  which  are
crushed  prior  to  the  vat leach process may be washed in a
spiral classifier for control of particulates (slimes)  unde-
sirable for vat leaching.  These slimes may be floated in  a
section  of  the  concentrator to recover copper sulfide and
then leached in a thickener for recovery  of  oxide  copper.
The  waste  tails  (slimes) are deposited in special evapora-
ting ponds.  The leach solution undergoes iron precipitation
to recover cement copper, and the barren solution is sent to
the evaporation pond as  well.  These wastes  are  character-
ized  in Table V-16.  No effluent results, as the wastes are
evaporated to dryness in the special impoundment.

The process has application when mined ores contain signifi-
cant amounts of both oxide and sulfide copper.

Process Description - Froth Flotation

Approximately  98%  of   ore  received   at   the   mill   is
beneficiated  by  froth  flotation at the concentrator.  The
process   includes   crushing,   grinding,   classification,
flotation, thickening, and filtration.  (See Figure V-14.)

Typically,  coarse  ore  is delivered to the mill for two- or
three-stage reduction by truck, rail or conveyor and is then
fed to a vibrating grizzly feeder, which passes its oversize
material to a jaw crusher.  The ore then travels by conveyor
to a screen for further  removal of fines ahead of  the  next
reduction  stage.   Screen oversize material is crushed by a
cone crusher.   When  ore  mineralogy  is  chalcopyrite,  or
contains   pyrite,   an  electromagnet  is  inserted  before
secondary crushing to remove tramp iron.  Crushing to  about
65 mesh is required for  flotation of porphyry copper.

The  crushed  material   is  fed  to  the  mill  for  further
reduction  in  a  ball   mill  and/or  rod  mill.    A  spiral
classifier  or  screen   passes  properly  sized  pulp to the
flotation cells.  Ahead  of the flotation cells,  conditioners
are employed to properly mix  flotation  reagents  into  the
pulp.   (see Figure V-15.)

Reagents  employed  for  this  process  might  include,  for
instance:
                            227

-------
TABLE V-16. MISCELLANEOUS WASTES FROM SPECIAL HANDLING OF
          ORE WASH SLIMES IN MINE 2124 (NO EFFLUENT)
PARAMETER
pH
TDS
TSS
COD
TOC
Oil and Grease
Al
Cd
Cu
Fe
Pb
Mn
Hg
Ni
Se
Ag
Ti
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/£)
SLIME LEACH-THICKENER
UNDERFLOW
2.4»
19,600
292,000
515
21
4.0
320.0
0.27
4,800
5,500
0.22
2.7
0.0026
1.5
< 0.003
0.057
3.8
8.9
1.0
0.5
<0.01
SLIME PRECIPITATION-
PLANT BARREN SOLUTION
1.8*
23,000
277
226
8
1.0
305.0
0.40
4,800
4,500
0.59
3.0
0.0560
1.75
< 0.003
0.054
4.2
35.0
1.0
3.75
<0.01
 'Value in pH units
                        228

-------
        Figure V-14. FLOW DIAGRAM  FOR FLOTATION OF COPPER (MILL 2120)
                                    I  MINING  I
                                       ORE
                                     CRUSHERS
                           195 nr/mttric ton
                         (46.736 gri/ihort ton)
                    RECYCLE
           REAGENTS

1
1

195
(46.73!
J
*|

                     f
                  PROCESS	
                  WATER
                                     BALL MILL
                           195 mj/nwtric ton
            CONCENTRATION



        CuS                T
       FROTH               [

                           t
                        RECYCLE
                                                             195 mj/m«tric ton
                                                             (46.735 gil/ihort ton)
                                                                                    TO ATMOSPHERE
                                                                 -TAILINGS-
                  27 m3/nwtric ton
                 (6.491 gd/ihort ton)
0.05 m3/mttric ton
(11.1 Bil/ihort ton)
                                                                • RECYCLE
                                                  -L
                                                                             THICKENER
RECYCLED
 WATER
            108 m3/mttrle ton
          (25,964 gil/ihort ton)

    77 m3/imtrle ton
    (18,545 gal/ihort ton)
                                                                             THICKENED
                                                                               TAILS
                                                                     15 m3/imtrte ton
                                                                     (3,709 ««l/*ort ton)
                                                               EVAPORATION
                                                                                  (3,709 gil/ihort ton)
RECYCLES (A) * (B
118 m3/nwtrlc ton
128,189 gil/thort ton)
                                             OVERFLOW
                                              (IF ANY)
                                                       DISCHARGE
                                            229

-------
Figure V-15. ADDITION OF FLOTATION AGENTS TO MODIFY MINERAL SURFACE
       REAGENTS TO
        ADJUST pH
                      PULP FROM GRINDING CIRCUIT
                            (25-45% SOLIDS)
                                i
CONDITIONER
                                                 WETTING AGENT
                                                   DISPERSANT
        COLLECTOR
                            CONDITIONER
                       ACTIVATOR
                    (OR DEPRESSANT)
CONDITIONER
                •FROTH-
          TO
      ADDITIONAL
      PROCESSING
 FLOTATION
   CELLS
-TAILINGS-
                         TO
                        WASTE
                            230

-------
 Reagent       Example of          Ib/short ton    kg/metric ton
 _type          Reagent             mill feed       mill  feed

 PH control     lime                    10.0           5.0
 collector     Xanthate                 0.01          0.005
 collector     Minerec                  0.03          0.015
               compounds
 frother       MIBC                     0.04          0.02

 The specific types of reagents  employed and amounts  needed
 vary  considerably  from  plant  to  plant, although  one may
 classify  themr  as  in Table V-17,  as precipitating agents,  pH
 regulators,  dispersants,   depressants,   activators,   collec-
 tors,  and frothers.

 Rougher-cell  concentrate  is   cleaned   in cleaner flotation
 cells.  The  overflow is  thickened,  filtered, and  sent to the
 smelter.   Tailings   (sands)  from   the cleaner   cells  are
 returned  to  the  mill  for  regrinding.   Tailings from the
 rougher cells are  sent to  the tailing pond for  settling   of
 solids.   Scavenger   cells, in  the  last cells of  the  rougher
 unit,  return their concentrate   (overflow)   to  one   of  the
 first  rougher cells.

 In   flotation,  copper sulfide  minerals are recovered in the
 froth  overflow.  The underflow  retains  the sands  and  slimes
 (tailings).    The  final,  thickened and filtered  concentrate
 contains  15  to  35  percent  copper  (typically,   25  to   30
 percent)  as  copper  sulfide.  Copper  recoveries  average  83
 percent,  so  a significant  portion of  the copper is discarded
 to  tailing ponds.  Tailings contains  15  to 50 percent solids
 (typically,  30 percent) and 0.05 to 0.3  percent copper.

 Selective or differential  flotation is  practiced   in  copper
 concentrators,  which (for example)  may separate molybdenum
 from copper  concentrate, copper  sulfide   from  pyrite,  and
 copper  sulfide  from copper/lead/zinc  ore.  Silver may  be
 floated from copper  flotation feed; gold and silver  may   be
 leached   by cyanide from  the  copper   concentrate,  with
 Precipitation by zinc dust.

Water Usage  in Flotation.   The major usage of water in  the
 flotation  process   is  as  carrier water  for the pulp.  The
carrier water added  in the crushing circuit also  serves   as
 contact  cooling water,  sometimes, water  sprays are used to
control dust in the  crusher.   Process  water  for  flotation
comes  from  mine-water  excess,  surface  and  well  water,
recycled tailing thickener, and lagoon water.  The  majority
of  the copper industry recycles and reuses as much water as
                            231

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TABLE V-17. EXAMPLES OF CHEMICAL AGENTS WHICH MAY BE
          EMPLOYED IN COPPER FLOTATION
MINERAL
Bornite
Chalcocite
Chalcopyrite
Native Copper
Azurite
Cuprite
Malachite
PRECIPITATION AGENT
—
—
—
—
Sodium monosulfide
Sodium monosulfide
Sodium monotulfide
PH
REGULATION
Lime
Lime
Lime
Lime
Sodium carbonate
Sodium carbonate
Sodium carbonate
DISPERSANT
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
DEPRESSANT
Sodium cyanide
Sodium cyanide
Sodium cyanide
Sodium cyanide
Quebracho
Quebracho
Tannie acid
ACTIVATOR
—
—
—
—
Polysulfide
Polysulfide
Polysulfide
COLLECTOR
Xanthate
Aeroftoats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats,
Fatty acids
and salts
Fatty acids
and salts,
Xanthates
Fatty acids
and salts,
Xanthates
FROTHER
Pine oil
Pine oil
Pine oil
Pine oil
Pine oil.
Vapor oil ,
Cresylic
•cid
Pine oil.
Vapor oil.
Cresylic
•cid
Pine oil,
Vapor oil,
Cresylic
•cid
                 SOURCE: REFERENCE 27
                    232

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 is available because  the  industries are located in  an  arid
 climate   (i.e., Arizona,  New Mexico, and Nevada).  There are
 plants in areas of  higher  rainfall  and  less  evaporation
 which  have reached 70, 95, and 100 percent recycle (or zero
 discharge) and are researching process changes and treatment
 technology in order to attain zero  discharge  of  all  mill
 water.  Three major copper mills discharge all process water
 from the  tailings lagoon  at this time.

 Table  V-18  outlines  the amount of water used in flotation
 per ton of concentrate produced.

 Noncontact cooling water  in the crushers, if not entirely in
 a closed  circuit, may be  reused in the flotation circuit and
 either  settled  in   holding  ponds  prior  to  recycle   or
 evaporated.  The use of noncontact cooling water in crushing
 appears   to  be  rare,  since  pulp  carrier water serves as
 contact cooling water.

 Waste Characterization.   The  chemical  characteristics  of
 raw  mill wastewater  (mill tailings) are summarized in Table
 V-19.  Residual flotation agents or their  degradation  pro-
 ducts  may  be  harmful   to  aquatic  biota,  although their
 constituents and toxicity have not  been  fully  determined.
 Their  presence (if any), however, does not appear to hamper
 the recycling of tailing  decant water to the  mill  process.
 Water  is  characterized  by  1  to  4  grams  per  liter of
 dissolved solids and by the presence of alkalinity, sulfate,
 surfactant, and fluoride.  Dissolved metals in decant  water
 are  usually  low, except for calcium (from lime employed in
 flotation process), magnesium, potassium, selenium,  sodium,
 and  strontium—which  do not respond to precipitation with
 lime.  On occasion, cyanide, phenol,  iron,  lead,  mercury,
 titanium, and cobalt are  detectable in the decant.  However,
 in  these  cases,  the  water  is  either  recycled fully or
 partially discharged.

 Handling or Treatment of  Decanted Water  From  Mill  Tailing
 Ponds.    The majority of the industry recycles all mill pro-
cess"" water  from the thickeners and the tailing pond due to
 the  need  for  water  in  the  areas  of  major  copper-ore
 production.    of the balance of the industry,  which includes
 approximately six major copper producing facilities  and  an
 undetermined  number  of  operations  producing  copper as a
byproduct, at least half  (50 percent)  are currently  working
toward  attaining  recycle  of mill Frocess water.   Also,  of
 the  six,  three  have  sophisticated  lime   and   settling
 treatment,  or  are installing it, to protect the quality of
                            233

-------
TABLE V-18. WATER USAGE IN FROTH FLOTATION OF COPPER
MILL
2101
2102
2103
2104
2106
2108
2109
2111
2112
2113
2114
2115
2116
2117
2118
2119
2120
2121
2122
2123
2124
WATER USAGE (1973)
m^/metric ton
concentrate
produced
95.8
188.7
77.6
474.3
36.0
141.9
N.P.^
280.4*
78.6
68.3
85.5
366.7
51.8
145.0
112.0
161.6
234.7
149.4
160.9
370.9
110.3
gal /short ton
concentrate produced
22,967
45,233
18,610
113,674
8,625
34,009
N.P.
67,201 »
18,847
16.377
20,503
87,888
12,417
34,763
26,846
38,738
56,257
35,801
38,570
88.905
26,440
    •Concentrate production estimated from known copper content and
    assuming concentrate contains 20.43% copper, as in 1972
    N.P.- No (1973) production
    SOURCE: Reference 26
                           234

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                          TABLE V-19.
RAW MILL WASTE LOADS PRIOR TO SETTLING IN TAILING PONDS
(Sheet 1 of 2)

PARAMETER

Flow
pH
•res
TDS
OilondGnioi
A>
Cd
C«
Ft
H|
Pb
Z«
C^-id.
Opmini
Amuri
Production
ofConcomrMi

CONCENTRATION
tmt/M
105380 m*la*t
(28.000.000 fol/diyl
8.1-
389.000
2.652
<0.05
<0.02
3.0
400
18300
0.003
21
310
0.01
MILL 2120*
RAW WASTE LOAD PER UNIT PRODUCT
ko'tOOO motric torn
1092 m3/natrk ton
81-
42,500.000
289.700

87.400
4.110.000
0.66
4.600
67,700
"
-
349,272 mrlric nra (38S.OOO riiort tan)


CONCENTRATION
lm»/ll
103.462 m3/dm
127 .335.000 oM/day)
9^-
114.000
395
 at nm 24-hour c
                                    May UTS leaMractof dral

                                    bath land nih 4nd din* aid 
-------
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
           TAILING PONDS (Sheet 2 of 2)
PARAMETER
Flow
pH
SES'
IDS
TSS
Oil and Gnaw
802
Al
Ai
Cd
Cu
ft
Pta
Mn
Hg
Ni
So
A«
8r
Zn
Sb
Co
Au
Mo
Phoiphata
Cymida
Oparatlnt
diyi/ynr
Annual
Production
of ConeantraM
MILL 21 24
CONCENTRATION
(ma/Hi
19,322 m3/diy
(5,104,800 gal/day)
10.05*
50%
2.846
466.000
1
46.75
<0.5
<0.07
O.OS
912.5
1.982
0.35
3t
0.0006
2.8
< 0.003
<0.1
1.2
5.6
<0.5
1.68
< 0.06
29.29
20.8
< 0.01
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
100.8 m3/nwtrfc ton
10.06*
-
286,996
47.000,000
100.8
4.714.34
<50.4
<7.06
6.04
92.017.8
199,867.7
35.29
3.126.1
0.0605
282.36
< 0.303
< 10.08
121 .01
564.71
<50.4
169.41
<5.04
2,953.66
2.097.6
<1.01
lb/1000 short tons
24,170 Bll/shon ton
10.06*
-
573,989
04.000,000
201.7
9.428.67
< 100.8
< 14.12
10.08
184,035.6
399,735.5
70.59
6,252.2
0.1210
564.71
< 0.606
< 20.17
242.02
1.129.42
< 100.8
338.83
< 10.08
6.907.29
4.195.0
< 2.02
362
69.362 matric ton) 176.457 diort torn)
       •Valua in pH units
       fSntlMbli solids
                       236

-------
 the  discharge.   These are discussed   in greater   detail   in
 Section  VII.

 As    discussed   previously,   noncontact cooling   water,   if
 present,  remains either  in a   closed   system  or   joins  the
 carrier  water to the flotation cells.

 Sewage   from  the mill is either handled in a treatment plant
 or,  in one case, is  sent to an acid leach holding  reservoir.
 Overflow from the treatment plants is either  discharged   or
 sent to  the tailing  pond.

 Variations in Flotation  Process.   Flotation tailings may  be
 separated at  the  concentrator into slimes and sands.  The
 sands usually are transferred directly to the tailing  pond.
 However,  in  one case,  the  slimes  (fines) are leached in a
 thickener prior to rejoining  the  thickener  underflow  with
 the   sand tails.    Sand  and  slimes   are  then sent to the
 tailing  pond.   Thickener overflow is  sent to a precipitation
 plant for recovery of oxide   copper   (Figure  V-16).   This
 variation is  employed  when  mined ores contain a  mixture  of
 sulfide  and oxide copper.

 Variations in Mill Processes

 Dual  Process.    Ores  which contain mixed sulfide   and  oxide
 mineralization   in  equal  ratios  (greater than 0.4 percent
 copper sulfides or oxides) may  be treated with vat leaching,
 as well as with froth flotation, in a  dual  process  (Figure
 V-17).

 Ore  is crushed  and placed  in  vats for  leaching with sulfuric
 acid,  as described   under "vat leaching."  The leachate  is
 sent  to iron  precipitation   or  electrowinning  plants  for
 recovery  of copper.   The residue, or tails, remaining in the
 vats  contains nonleachable copper sulfides and is  treated  by
 froth  flotation   to   recover the copper, as described under
 "Froth Flotation."

 Water usage and tailing-water quality   are  similar  to  the
 processes  of vat  leaching and  froth flotation.   No discrete
discharge differences result  from this  variation compared to
vat leaching and  froth flotation.

Leach/Freeipitation/Flotation  (LPF)  Process.   Mixed  sulfide
and   oxide  mineralization  may  also  be  handled  by  the
leach/precipitation/flotation process.  Crushing may  be  in
two  or three stages  (Figure V-18).   Both rod and ball mills
may be employed to produce a pulp of  less than 65  mesh  and
                            237

-------
Figure V-16. FLOWSHEET FOR MISCELLANEOUS HANDLING OF FLOTATION
          TAILS (MILL 2124)
                    MINING
                     ORE
                     i
                   SULFIDE
               FLOTATION CELLS
                      I
                   TAILINGS
                     i
               HYDROSEPARATOR
                      I
                    SLIMES
                     i
                    ACID
                    LEACH
                 (THICKENER)
    CONCENTRATE
    TO
    STOCKPILE
  •SANDS-
UNDERFLOW
  (SLIMES)
                       PREGNANT
                       SOLUTION
               BARREN
              SOLUTION
                PRECIPITATION
                    PLANT
                     I
                   CEMENT
                   COPPER
    TO
•>- TAILING
    POND
                      TO
                      STOCKPILE
                            238

-------
        Figure V-17. DUAL PROCESSING OF ORE (MILL 2124)
                   MINING
                     I
                    ORE
                    i
                  CRUSHERS
                    i
                   LEACH
       RECYCLED
        WATER
                 ACID
               ' SOLUTION
ELECTROWINNING
 DIRECT
SULFIDE
 MILL
 FEED
     ORE
    LEACH
    TAILS
                              RECYCLED ACID-
CONCENTRATOR
                    RECYCLED
                   'OVERFLOW
                   RECYCLED
                    DECANT
                      TAILING
                    THICKENERS
                                       CATHODE
                                        COPPER
                                                        TO
                                                     STOCKPILE
                          239

-------
Figure V-18. LEACH/PRECIPITATION/FLOTATION PROCESS
                        MINING
                         ORE
                        J_
                         FIRST
                       CRUSHER
                         I
                       SCREENING
                        SECOND
                        CRUSHER
                       ROD MILL
                       BALL MILL
                        LEACH
                       AGITATORS
                     PRECIPITATORS
                          I
                     SPONGE COPPER
                         AND
                      SPONGE IRON
                      CONDITIONER
                         I
                     FROTH (pH 4.0(
                      FLOTATION
          RECYCLE
           WATER
                      CONCENTRATE
                       THICKENER
                         DISC
                        FILTER
                         I
               COPPER SULFIDE CONCENTRATE
                         AND
                     SPONGE COPPER
                         T
                      TO STOCKPILE
                         240

-------
25  percent  solids.   The  pulp  flows  to acid-proof leach
agitators.  sulfuric acid (to a pH of 1.5 or 2.0)   is  added
to the feed.  The leaching cycle continues for approximately
45  minutes.   The  acid  pulp  then is fed to precipitation
cells, where burned and  shredded  cans  or  finely  divided
sponge  iron  (less than 35 mesh) may be used to precipitate
copper by means of an  oxidation/reduction  reaction,  which
increases the pH of the pulp to 3.5 to 4.0:

              CuS04 * Fe   	>   Cu * FeS04_
                  "(excess)

Copper  precipitates  as  a  sponge,  and  the entire copper
sponge, together with pulp-sponge iron feed, is  carried  to
flotation  cells.  Flotation recovers both sponge copper and
copper  sulfide  in  the  froth  by  means  of  the   proper
conditioning  reagents, such as Minerec A as a collector and
pine oil as a frother.  Flotation is accomplished at a pH of
4.0  to  6.0  (±0.5).   The  concentrate  is  thickened  and
filtered  before  it  is  shipped  to  the  smelter.  Copper
recovery may be as  high  as  91  percent.   An  example  of
reagent consumption for this process is:

    Reagent             kg/metric ton       lb/short ton
      type              of mill feed        of mill feed

    Sulfuric acid            12.5                25
    Sponge iron              18                  36
    Minerec A                 0.09                0.18
    Pine oil                  0.04                0.08

Lead and Zinc Ores

The chemical characteristics of raw mine drainage are deter-
mined  by  the  ore  mineralization  and  by  the  local and
regional geology encountered.  Pumping  rates  for  required
mine dewatering in the lead and zinc ore mining industry are
known  to  range from hundreds of cubic meters per day to as
much as 200,000 cubic meters per day (52 million gallons per
day) .

Raw  wastewater  from  miling  operations  appears   to   be
considerably  less  variable  from facility to facility than
mine wastewater.  The volume of mill discharge  varies  from
as  little as 1000 cubic meters per day (264,200 gallons per
day) to as much as 16,000 cubic meters per  day  (4  million
gallons  per  day).   When  expressed as the amount of water
utilized per unit of ore processed, quantities varying  from
330  cubic  meters  per metric ton  (79,070 gal/short/ton)  to
                            241

-------
1,100 cubic meters per metric  ton   (263,566  gal/short/ton)
are  encountered.  The sources and characteristics of wastes
in each recommended subcategory are discussed below.

Sources  of_  Wastes  -   Mine   Water    (No   Solubilization
Potential) .

The main sources of mine water are:

     (1)  Ground-water infiltration.

     (2)  Water  pumped  into  the  mine  for  machines   and
drinking.

     (3)  Water resulting from hydraulic backfill operations.

     (H)  Surface-water infiltration.

The geologic conditions which prevail in the mines  in  this
subcategory consist of limestone or dolomitic limestone with
little or no fracturing present.  Pyrite may be present, but
the  limestone is so prevalent that, even if acid is formed,
it is almost certainly neutralized in situ before any metals
are solubilized.  Therefore, the extent of heavy  metals  in
solution  is  minimal.   The  principal contaminants of such
mine waters are:

     (1)  Suspended  solids  resulting  from  the   blasting,
         crushing,  and  transporting  of  the ore.  (Finely
         pulverized minerals may be  constituents  of  these
         suspended solids.)

     (2)  Oils and greases resulting from spills and leakages
         from  material-handling  eguipment  utilized   (and,
         often, maintained) underground.

     (3)  Hardness and alkalinity associated  with  the  host
         rock and ore.

     (U)  Natural nutrient level of the subterranean water.

     (5)  Dissolved salts not present in surface water.

     (6)  Small quantities of unburned  or  partially  burned
         explosive substances.

A simplified diagram illustrating mining operations and mine
wastewater   flow  for  a   mining  operation  exhibiting  no
solubilization  potential   is   shown   in   Figure   V-19.
                             242

-------
Figure V-19. WATER FLOW DIAGRAM FOR MINE 3105
ctEEnA/ir? *--
DRILL
WATER 270m<5/day

MINE
-~


   (72,000 gpd)
          T
                   PUMPING
                        7,600 m3/day
                        (2,000,000 gpd)
                                      FUEL AND LUBRICANT
                                      SPILLAGE AND
                                      LEAKAGE
                                      EXPLOSIVE
                                      WASTE PRODUCTS
(MI
^
                 MILL FEED-WATER
                  RESERVOIR
                   243

-------
Typically,  mine water may be treated and discharged or used
in a nearby mill as flotation-process water.

The range of chemical constituents measured for three  mines
sampled as part of this program is given in Table V-20.  The
data,  although limited to 4-hour composite samples obtained
during three site visits, generally confirm other data  with
a  narrower  range of parameters.  Generally, raw mine water
from this class of mine is of good quality, and any  problem
parameters  appear  to  be  readily  remedied by the current
treatment practice of sedimentation-pond systems.

Sources of Wastes - Mine Water (Solubilization Potential)

The  sources  of  water  from  mines   with   solubilization
potential   are   the  same  as  those  for  mines  with  no
solubilization  potential.   The  key  difference  in   this
situation  is  the local geologic conditions that prevail at
the  mine.   These  conditions  lead  to  either  gross   or
localized   solubilization  caused  by  acid  generation  or
solubilization  of   oxidized   minerals.    The   resultant
wastewater  pumped  from  the  mine  contains the same waste
parameters as that from the preceding subcategory  but  also
contains  substantial  soluble metals.  Table V-21 shows the
range of chemical constituents from  four  mines  exhibiting
solubilization potential.

The  following  reactions  are  the basic chemical reactions
that describe an acid mine-drainage situation:

Reaction 1—Oxidation of Sulfide to Sulfate

When natural sulfuritic material in the form  of  a  sulfide
(and,  usually,  in combination with iron)  is exposed to the
atmosphere (oxygen), it may  theoretically  oxidize  in  two
ways  with  the  presence  of  water  (or water vapor) as the
determining factor:

(A) Assuming  that  the  process  takes  place  in   a   dry
    environment,  an  equal amount of sulfur dioxide will be
    generated with the formation of  (water-soluble)  ferrous
    sulfate:

         FeS2 * 302    	>   Fes04 + S0.2

(B) If, however, the oxidation proceeds in the presence of a
    sufficient quantity  of  water   (or  water  vapor) ,  the
    direct  formation  of sulfuric acid and ferrous sulfate,
    in equal parts, results:
                            244

-------
TABLE V-20. RANGE OF CHEMICAL CHARACTERISTICS OF SAMPLED
          RAW MINE WATER FROM LEAD/ZINC MINES 3102,3103,
          AND 3104 SHOWING LOW SOLUBILIZATION
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NHg
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£ )
7.4 to 8.1*
180 to 196
200 to 330
2 to 138
326 to 510
< 10 to 631
<1 to 4
3 to 29
0.03 to 0.15
< 0.05 to 1.0
< 0.0001 to 0.0001
< 0.2 to 4.9 1
0.03 to 0.69
<0.02
<0.002 to 0.015
<0.02
< 0.02 to 0.06
<0.02 to 0.90
37 to 63
3 to 57
0.3 to 1.2
* Value in pH units
' f\tt+M MA*** V«k£tA4»+ iMiMllAnf*A **# a^lfi •+al*ll lVa+1 *«rk S%M »n«JlM*AM+
                       245

-------
TABLE V-21. RANGE OF CHEMICAL CHARACTERISTICS OF RAW MINE WATERS
          FROM FOUR OPERATIONS INDICATING HIGH SOLUBILIZATION
          POTENTIAL
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/A )
IN RAW MINE WATER
3.0 to 8.0*
14.6 to 167
178 to 967
< 2 to 1.047
260 to 1,722
15.9 to 95.3
Itoll
Oto3
0.020 to 0.075
< 0.05 to 4.0
0.0001 to 0.001 3
0.1 to 1.9
1.38 to 38.0
< 0.02 to 2.1
0.01 6 to 0.08
0.17 to 0.42
< 0.02 to 57.2
0.1 2 to 22.0
48 to 775
< 0.01 to 220
0.06 to 0.80
             •Value in pH unit*
                         246

-------
          2FeS,2 + 2E20 * 102   	>  2FeSO£ +

 In   most   mining   environments   in   this    subcategory
 (underground,   as well as in the tailing area) ,  reaction (B)
 is favored.

 Reaction 2.—Oxidation of Iron (Ferrous to Ferric)

 Ferrous  sulfate, in the presence of quantities  of   sulfuric
 acid   and oxygen,  oxidizes  to  the  ferric state to form
 (water-soluble)  ferric sulfate:

 iFeSOO. + 2H2S04  4 02    	>   2Fe2(SOj*)3 * 2H2.0

 Here,  water  is not limiting since it is  not  a   requirement
 for  the reaction but, rather,  is a product of the  reaction.
 Most  evidence  seems to indicate that bacteria  (Thiobacillus
 ferrobacillus.  Thiobacillus  sulfooxidans)   are involved in
 the above reaction  and,   at  least,   are  responsible  for
 accelerating  the  oxidation  of  ferrous iron to the ferric
 state.

 Reaction 3_—Precipitation  of Iron

 The ferric iron  associated with  the  sulfate ion   commonly
 combines  with  the  hydroxyl  ion  of  water to form ferric
 hydroxide.   In an  acid environment,   ferric hydroxide  is
 largely  insoluble and precipitates:

    Fe2(SQijL)3  +  6H20    	>   2Fe(OH)J * 3H2SQ±

 Note   that  the   ferric ion can,   and  does, enter into  an
 oxidation/ reduction  reaction with iron sulfide  whereby the
 ferric   ion  "backtriggers"  the  oxidation of  further amounts
 of  sulfuritic  materials  (iron sulfides,  etc.)  to the sulfate
 form, thereby  accelerating the acid-forming process:

             _  *  FeS2  + H2O  	>   SFeSO]* * 2S

               S  * 30  + H20    	>    H2S04

The fact that  very little  "free" sulfuric acid is   found   in
mine waste drainage is  probably  due to  the reactions  between
other soluble  mineral  species and  sulfuric acid.

In some ore bodies, such reactions—and  subsequent  solubili-
zation of metals—may  occur  in local regions in  which little
or  no limestone  or dolomite  is  available for neutralization
before the harmful solubilization occurs.  Once  a metal such

-------
as copper, lead, or zinc  is  in  solution,  the  subsequent
mixing  and neutralization of that water may not precipitate
the  appropriate  hydroxide  unless  a  rather  high  pH  is
obtained.   Even  if  some of the metal is precipitated, the
particles may be less than 0.45 micrometer  (0.000018  inch)
in  size  and,  thus, appear as soluble metals under current
analytical practice.

Conditions compatible with solubilization of certain metals-
particularly, zinc—are associated with heavily fissured ore
bodies.  Although the minerals being recovered are sulfides,
fissuring of the ore body allows the slight oxidation of the
ore to oxides,  which  are  more  soluble  then  the  parent
minerals.

When   conditions   exist  which  provide  a  potential  for
solubilization, the mine water resulting  is  of  a  quality
which  requires treatment beyond conventional sedimentation.
The best current practice suggests  that  the  treated  mine
water is likely to be of a quality inferior to raw discharge
from  mines where the potential for such solubilization does
not exist.

A flow diagram illustrating flows encountered in a  mine  of
the type described in this subcategory is shown as Figure V-
20.    The   characteristics   of   mine  waters  from  this
subcategory are illustrated by Table V-21,  which  amplifies
the above observations.

These  data suggest that particular problems are encountered
in achieving zinc and cadmium levels approaching the  levels
of  raw mine water from the class of mines with no solubili-
zation potential.

Process Description - Mill Flows and Waste Loading

The raw wastewater from a lead/zinc flotation mill  consists
principally  of  the water utilized in the flotation circuit
itself, along with any housecleaning water used.  The  waste
streams  consist of the tailing streams (usually, the under-
flow of the zinc rougher flotation cell), the overflow  from
the   concentrate   thickeners,   and   the   filtrate  from
concentrate  dewatering.   The  water  separated  from   the
concentrates is often recycled in the mill but may be pumped
with the tails to the tailing pond, where primary separation
of  solids  occurs.  Usually, surface drainage from the area
of the mill is also collected and sent to  the  tailing-pond
system for treatment.
                            248

-------
       Figure V-20. WATER FLOW DIAGRAM FOR MINE 3104
SEEPAGE*
       MINE
(ALL WATER REQUIRED
 FOR DRILLING FROM
     SEEPAGE)
                         I
                     PUMPING
                         3,460 m3/day
                         (915,000 gpd)
                /^SEDIMENTATION^
                V^^ BASIN ^^X
                     nr
                    DISCHARGE
                       T
                                            FUEL AND LUBRICANT
                                            SPILLS AND LEAKAGE
                                            EXPLOSIVE WASTE
                                            PRODUCTS
                         249

-------
The  principal characteristics of the waste stream from mill
operations are:

    (1)  Solid loadings of 25 to 50 percent (tailings).
    (2)  Unseparated minerals associated with the tails.
    (3)  Fine particles of minerals—particularly, if the
         thickener overflow is not recirculated.
    (4)  Excess flotation reagents which are not associated
         with the mineral concentrates.
    (5)  Any spills of reagents which occur in the mill.

Figure V-21 illustrates the sources, flow rates,  and  fates
of  water used for the flotation process in beneficiation of
lead and zinc ores.

One aspect of mill waste which has  been  relatively  poorly
characterized from an environmental-effect standpoint is the
excess flotation reagents.  Unfortunately, it is very diffi-
cult to analytically detect the presence of these reagents—
particularly,  those  which  are  organic.  The TOC and MBAS
surfactant  parameters  may  give  some  indication  of  the
presence   of   the  organic  reagents,  but  no  definitive
information is implied by these parameters.

Significant   characteristics   of   raw   wastewater   from
lead/zinc/ore  flotation  mills are illustrated by data from
three mills visited during this investigation, as summarized
in Table V-22.  Note that some of the lower values shown  in
this  table  result from admixture of excess mine water with
the tailing stream.  Information  for  a  mill  using  total
recycle  and  one  at which mill wastes are mixed with metal
refining wastes in the tailing pond are not included in this
summary.  Feed water for the mills  is  usually  drawn  from
available  mine  waters; however, one mill uses water from a
nearby lake.  These  data  illustrate  the  wide  variations
caused  by  the  ore  mineralogy,  grinding  practices,  and
reagents utilized in the industry.

Gold Ores

Water flow and the sources,  nature,  and  quantity  of  the
wastes  dissolved in the water during the processes of gold-
ore mining and beneficiation are described in this section.

Water Uses

The major use of water in this industry is in  beneficiation
processes, where it is required for the operating conditions
of  the individual process.  Water is normally introduced at
                            250

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                       Figure V-21. FLOW DIAGRAM FOR MILL 3103
                MINE

3.786 m3Atoy
11. 000,000 gpd)
16,150-m3
(4.000,000-pl)
MILL FEED
RESERVOIR

9,500m3
12.500.001

1.890 m3/day
1500,000 gpd)
/diy
>ltprfl
REAC
                                                                WATER
                                                               FROM MILL
                                                             FEED RESERVOIR
                                                                   8.600 m3««y
                                                                   12.500.000 gpd)
        (a) WATER BALANCE
TO TAILING-
"•ONO SYSTEM
                               TO STOCKPILES


                                    (b) MILL PROCESS
                                      251

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 TABLE V-22.
RANGE OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE
LOADS FOR MILLS 3101, 3108, AND 3109.
PARAMETER
PH
TSS
Cd
Cr
Cu
Fe
Mn
Pfa
Zn
RANGE OF
CONCENTRATION (mg/1 )
IN WASTE WATER
lower limit
7.9*
20,500
1.2
9.8
4.8
2,900
295
76
160
upper limit
10.7*
269,000
16.4
40.0
496
35,000
572
560
3.000
RANGE OF RAW WASTE LOAD PER UNIT ORE MILLED
kg/1 000 metric tons
lower limit
—
140,000
8
79
32
19,500
1,500
395
1,400
upper limit
_
1.000,000
79
210
2,600
180,000
2,950
4,800
15.500
lb/1000 short tons
lower limit
_
280,000
16
168
64
39,000
3,000
790
2.800
upper limit
_
2,000.000
168
420
5,200
360,000
5,900
9.600
31,000
'Value in pH units.
                            252

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the grinding  stage  of   lode  ores   (shown  in  the  process
diagrams  of  section  III)  to  produce  a  slurry which is
amenable to pumping, sluicing, or classification  into  sand
and slime fractions for  further processing.  In slurry form,
the  ground   ore  is  most  amenable to beneficiation by the
technology currently used to process the predominantly  low-
grade   and   sulfide    gold   ores—i.e.,  cyanidation  and
flotation.  The gravity  separation process commonly used  to
beneficiate   placer  gravels also requires water as a medium
for separation of the fine and heavy particles.

Other uses of water in gold mills include washing of  floors
and machinery and domestic applications.  Nash water is nor-
mally   combined   with   the  process  waste  effluent  but
constitutes only a small fraction  cf  the  total  effluent.
Some fresh water is also required for pump sealing.  A large
quantity  of  water  is  required in the vat leach process to
wash the leached sands and residual cyanide from  the  vats.
Because  the  sands  must be slurried for pumping twice, the
vat leach process requires approximately twice the  quantity
of water necessary for the milling of gold ore by any of the
other leaching processes.

With  the  exception  of  hydraulic  mining  and dredging of
placers, water is  not  normally  directly  used  in  mining
operations  but, rather, is discharged as an indirect result
of  a  mining operation.   Cooling  is  required  in   some
underground   mines,  and  water  is  used to this end in air
conditioning  systems.   This water does not come into  direct
contact  with the mine and is normally discharged separately
from the mine seepage.

Water flows of  four  gold  mining  and  milling  operations
visited during this study are presented in Figure V-22.

Sources of Wastes

There   are   two  basic  sources  of  effluents  containing
pollutants:   (1)   mines  and  (2)   beneficiation  processes.
Mines  may be either open-pit or underground operations.   in
the case of an open pit, the source of the pit discharge, if
any, is precipitation, runoff,  and ground-water infiltration
into the pit.   Ground-water  infiltration  is  the  primary
source  of  water  in  underground  mines.  However,  in some
cases, sands removed from mill tailings are used to backfill
stopes.   These sands may initially contain 30 to 60  percent
moisture,  and  this water may constitute a major portion of
the  mine  effluent.    The  particular  waste   constituents
present  in  a  mine or mill discharge are a function of the
                            253

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Figure V-22.   WATER FLOW IN FOUR SELECTED GOLD
                 MINING  AND MILLING OPERATIONS
                       1.227 m-Vdtv
                       (321,600 |pd)
                            (a) MINE/MILL 4101
             3.117 m3/ctay
             (1,000,000 gpd)
                         DISCHARGE TO STREAM
      2.280 mj/d.y
      (600,000 gpd)
                                                      DISCHARGE TO STREAM
       •AMALGAMATION OF GRAVITY-SEPARATED SANDS; FINES AND GOLD-EXTRACTED SANDS
        ARE FLOATED FOR RECOVERY OF BASE METALS.
                           (b) MINE/MILL 4102
                                                         RAIN
              2.5S7 TO 17.175 n3/d>y
            1*70,000 TO 4,600,000 gpd)
              311 r»3/d.y
              (94,GOOgpdl
                                       4.947 m'/diy
                                       (17M.OOO gpd)
     'GOLD VALUES PRESENT IN BASE-METAL CONCENTRATES, RECOVERED AT SMELTER OR REFINERY.

                           (c)  MINE/MILL 4103
UNDERGROUND
MINE


DISCHARGE
IMPOUNDED
                                        964 m3/diy
                                        (2SO.OOO gpd)
      'tit m3/diy
       1218,000 gpd)
1,908m3/d«y
(500.000 gpd)
DURING APRIL AND MAY
                                                        INTERMITTENT DISCHARGE
                           (d) MINE/MILL 4104
                                  254

-------
mineralogy  and  geology of the ore body  and  the  particular
milling   process  employed.  The rate and extent to which the
minerals  in an  ore   body  become  solubilized  are  normally
increased  by   a  mining  operation,  due to the exposure of
sulfide   minerals  and  their  subsequent   oxidization   to
sulfuric  acid.  At  acid pHr the potential for solubilization
of  most  heavy  metals  is greatly increased.  Not all mine
discharges  are  acid, however; in those cases where they  are
alkaline,   soluble   arsenic, selenium, and/or molybdenum may
present problems.

In  the   beneficiation  of  placer  gravels,  no   crushing,
grinding, or use  of chemical reagents is necessary.  Gold is
separated  from  the gravels  and  sand by physical methods
alone.  As  a result, the waste parameters of concern are the
suspended  and/or   settleable   solids   generated   during
dredging,   hydraulic   stripping,   or  washing   (sluicing)
activities.

Wastewater  emanating from mills consists almost entirely  of
process  water.   High suspended-solid loadings are the most
characteristic  waste constituent of  a  mill  waste  stream.
This  is  primarily  due to the necessity for fine grinding of
the ore to  make it  amenable to  a  particular  beneficiation
process.    In   addition,  the  increased surface area of the
ground ore  enhances the possibility  for  solubilization  of
the  ore minerals and gangue.  Although the total dissolved-
solid loading may   not  be  extremely  high,  the  dissolved
heavy-metal concentration may be relatively high as a result
of  the highly  mineralized ore being processed.   These heavy
metals, the suspended solids, and process  reagents  present
are the principal waste constituents of a mill waste stream.
Depending  on   the  process conditions, the waste stream may
also have a high  or low pH.  The pH is of concern, not  only
because  of  its potential toxicity, but also because of the
resulting  effect   on   the   solubility   of   the   waste
constituents.

Process Description - Mining

Gold  is mined  from two types of deposits:   placers and lode
(vein)  deposits.  Placer mining consists of excavating gold-
bearing gravel  and sands.   This is currently done  primarily
by  dredging,   hydraulic  mining or mechanical excavation of
deeply buried placers.   Lode deposits are  mined  either  by
either  underground  (mines 4102,  4104,  and  4105)  or open-pit
(mine 4101)  methods, the particular method  chosen  depending
on such factors as size and shape of the deposit,  ore grade,
                            255

-------
physical   and   mineralogical  character  of  the  ore  and
surrounding rock, and depth of the deposit.

Characterization  of  wastewater  emanating  from   selected
placer operations is presented in Table V-23.

Both mines 4104 and 4105 are underground mines, and, at each
of  these,  the  coarse fracton of the mill tailings is used
for backfilling of  stopes.   Mill  wastewater  is  used  to
sluice   the  tailing  sands  underground,  and  this  water
undoubtedly contributes to the pollutant loading of the mine
discharge.  This practice also accounts for the presence  of
mill reagents such as cyanide in the mine-water discharge.

The  chemical  composition  of raw mine effluent measured at
three of the lode-ore mines visited is listed in Table V-24.
Although incomplete chemical data for mine 4102 are  listed,
considerable   variability  was  observed  with  respect  to
several key components (TS, TDS, SO4—, Fe, Mn, and Zn).

Both mines 4104 and 4105 are underground mines, and at  each
of  these,  the coarse fraction of the mill tailings is used
for backfilling of  stopes.   Mill  wastewater  is  used  to
sluice   the  tailing  sands  underground,  and  this  water
undoubtedly contributes to the pollutant loading of the mine
discharge.  This practice also accounts for the presence  of
mill reagents such as cyanide in the mine-water discharge.

Process Descript.ions - Milling

The  gold  milling  processes  requiring  water  usage  with
subsequent  waste  loading  of  this  water,  as   discussed
previously, are:

    (1)  cyanidation,

    (2)  amalgamation, and

    (3)  flotation.

There  are  four  variations  of  the  cyanidation   process
currently being practiced in the U.S.:

    (1)  agitation-leaching,

    (2)  vat leaching,

    (3)  carbon-in-pulp, and

    (4)  heap leaching.
                             256

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                            TABLE V-23.  CHARACTERIZATION OF GOLD-PLACER RAW WASTEWATER
to
in
PARAMETER
PH
CONDUCTIVITY2
DISSOLVED OXYGEN
TURBIDITY (MTU)
TSS
SETTLEABLE SOLIDS
TYPE OF OPERA-
TION
CONCENTRATION imgf t )
MINE 4106
75»
45*
9.0
180
1.720
3.0* •
HYDRAULIC
STRIPPING
AND SLUICE-
BOX
LOADING
MINE 4107
72«
75t
82
1.600
13.000
28"
MECHANICAL
EXCAVATION/
CAT-LOADED
SLUICEBOX
MINE 4108
7.0-
86*
7.7
2,000
17,200
18"
MECHANICAL
EXCAVATION/
CAT-LOADED
SLUICEBOXES
(2)
MINE 4110
72*
74f
5.2
3200
47,100
60"
HYDRAULIC
STRIPPING/
DRAGLINE-LOADED
SLUICEBOX
MINE 4112
6.9"
,88'
-
900
9.670
20"
HYDRAULIC
STRIPPING/
DRAGLINE-LOADED
SLUICEBOX AND
BUCKET LINE
DREDGE
MINE 4113
7.3*
1,320f
<\
31200
535.000
104"
HYDRAULIC
STRIPPING
AND SLUICE-
BOX
LOADING
MINE 41 14
7.2*
1.300f
8.0
32300
94.600
186"
MECHANICAL
EXCAVATION/
FRONT-END-LOADED
SLUICEBOX/
DRAGLINE
             •Value in pH units.
             * Value in micromhos/cm at 25°C (77°F)
            "Value in m I/i .

-------
TABLE V-24. CHEMICAL COMPOSITION OF RAW MINE
              WATER  FROM MINES 4105,4104 AND 4102
PARAMETER
PH
Alkalinity
Color
Turbidity MTU)
TOS
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Al
As
Bt
Ba
B
Cd
Ca
Cr
Cu
Total Fa
Pb
CONCENTRATION (mgll
MINE 4106
.
276
34f
2.40
1.190
1,176
14
733
35.01
12.0
1.0
0.095
<0.2
0.03
< 0.002
< 0.6
0.18
< 0.02
87.0
< 0.02
< 0.02
1.2
< 0.1
MINE 4104**
3.3*(3.2*>
-
-
-
-
4,747
81(20)
-
3.780
•
0.1
-
-
-
•
•
-
(0.04)
-
-
0.7 (1.7)
169(205)
-
MINE 4102**
6.16

-

536
530
5
-
27
•
< 0.1
•
0.143
0.084
-
•
-
0.025
•
-
0.056
25.11
0.82
PARAMETER
Mg
Mn
Hg
Ni
Tl
V
K
Ag
Na
Sr
Ta
Ti
Zn
Sb
Mo
Sulfata
Nltrata
Phoiphata
Cyanida
Phenol
Chloride
Fluorida
CONCENTRATION (mat I
MINE 4105
80,0
0.14
< 0.0001
0.10
< 0.06
<0
44.0
< 0.02
80.0
0.78
0.10

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In general, the cyanidation process involves  solubilization
of  gold with cyanide solution, followed by precipitation of
gold from solution with zinc dust.  (See Figure III-9.)

The agitation-leach process employed by mill  4401  requires
water  to  slurry the ground ore.  Cyanide solution is added
to this pulp in tanks, and this mixture is agitated to main-
tain maximum contact of the cyanide with the ore.   Pregnant
solution  is  separated from the leached pulp in thickeners,
and gold is precipitated from this solution with zinc  dust.
(See Figure 111-10.)

The  vat leaching process is employed by mill 4105.  In this
process, vats are filled with ground  ore  slurry,  and  the
water  is  allowed  to  drain off.  Cyanide solution is then
sprayed into the vats, and gold is  solubilized  by  cyanide
percolating   through   the  sands.   Pregnant  solution  is
collected  at  the  bottom  of  the  vats,   and   gold   is
precipitated with zinc dust.

The  carbon-in-pulp process is also used by mill 4105.  This
process was designed to recover gold from  slimes  generated
in  the  ore grinding circuit.  Water is added to the ore to
produce  a  slurry  in  the  grinding   circuit   which   is
subsequently   cycloned.   Cyclone  underflows  (sands)   are
treated by vat leaching, while cyclone overflow  is  treated
by  the carbon in-pulp process.  In this process, the slimes
are mixed with cyanide solution in large tanks, and  contact
is  maintained by agitation of the mixture (much the same as
for agitation leach).  This mixture is then caused to  batch
flow through a series of vats, where the solubilized gold is
collected  by  adsorption  onto activated charcoal, which is
held in  screens  and  moved  through  the  series  of  vats
countercurrent  to  the flow of the slime mixtures.  Gold is
stripped from this charcoal using  a  small  volume  of  hot
caustic.   An  electrowinning process is used to recover the
gold from this solution.  (See Figure III-9.)

Heap leaching has had only  limited  application  in  recent
years.   This inexpensive process has been used primarily to
recover gold from low-grade ores.  As the price of gold  has
risen  dramatically  since  1970,  the principal use of heap
leaching during this time has been in the recovery  of  gold
from   old  mine  waste  dumps.   This  process  essentially
consists of percolating cyanide solution down through piled-
up waste rock.  The leachate is usually collected by gravity
in a sump; in  some  cases,  use  is  made  of  a  specially
                            259

-------
constructed   pad  to  support  the  rock  and  collect  the
leachate.

Amalgamation can be done in a number of ways.   The  process
employed  by mill 4102 is termed "barrel amalgamation." This
essentially consists of adding  mercury  to  goId-containing
sands in a barrel.  The barrel is then rotated to facilitate
maximum  contact  of  mercury  with the ore.  The amalgam is
collected by gravity, and the gold and mercury are separated
by pressing in a hand-operated press.

Water is used by mill 4104 to slurry ground ore,  making  it
amenable  to  a  flotation  process.   The  slurried  ore is
transported to conditioner tanks,  where  specific  reagents
are added; essentially, this causes gold-containing minerals
to  float  and be collected in a froth, while other minerals
sink and are discarded.   This  separation  is  achieved  in
flotation  cells in which the mixture is agitated to achieve
the frothing.  The froth is collected off  the  top  of  the
slurry  and is further upgraded by filtering and thickening.
Tailings from the flotation process of mill 4104 are further
processed  by  the  cyanidation/agitation-leach  process  to
recover residual gold values.

In  addition  to suspended solids, particulate and dissolved
metals, reagents used in the mill beneficiation process also
add to the pollutant  loading  of  the  waste  stream.   The
particular  reagents  used  are  a  function  of the process
employed to  concentrate  the  ore.   In  the  gold  milling
industry,   cyanide  and  mercury,  clearly,  are  the  most
prominent  reagents  of  the  cyanidation  and  amalgamation
processes.   These  reagents are also of primary concern due
to their potential toxicities.   Table  V-25  indicates  the
quantity  of  each of these reagents consumed per ton of ore
milled.  The bulk of these reagents which are  used  in  the
process are present in the waste stream.

Because  there  is  both  a potential for solubilization and
suspension of the ore minerals present,  heavy  metals  from
these minerals may exist in the mill waste stream.  Table V-
26  lists  the  minerals  most commonly associated with gold
ore.  Since settleable solids  and  most  of  the  suspended
solids  are  collected  and  retained  in tailing ponds, the
dissolved and dispersed particulate heavy metals present  in
the final discharge are of ultimate concern.  Depending upon
the  extent  to which they occur in the ore body, particular
heavy metals may be present in a mill waste  stream  in  the
range  of  from  below  detectable  limits  to  3 to 4 mg/1.
                             260

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     TABLE V-25. PROCESS REAGENT USE AT VARIOUS MILLS BENEFICIATING
                   GOLD ORE


""ILL

4105
4105
4105
4101
4102
— -


MILL PROCESS

Cyanidation/Laach
Cyanidation/Char-in-pulp
Milt 4105 Total*
Cyanidation/Agitation Leach
. . . **tt
Amalgamation
REAGENT CONSUMPTION
CYAN ID ATI ON
kg/metric ton
ore milled
0.13
0.58
0.30
0.18
-
lb/»hort ton
ore milled
0.26
1.16
0.60
0.35
-
AMALGAMATION
kg/metric ton
ore milled
_
_
„
_
0.001
Ib/short ton

_
•••
_
^^
0.002
  Reagent consumption based on 1975 data supplied by the company.
 Total reagent consumption for Mill 4105 is the weighted average based on relative percentages of ore milled in the two
^circuits noted. This figure is based on 1975 data supplied by the company.
  Reagent consumption is based on 1973 data.
  Primary operation at this  mill is flotation for the recovery of base-metal concentrates.
               TABLE V-26. MINERALS COMMONLY ASSOCIATED
                            WITH GOLD ORE
MINERAL
Arsenopyrite
Pyrite
Chalcopyrite
Galena
Sphalerite
Greenockite
Cinnabar
Pentlandite
Calverite
Sylvanite
Native Gold
Selenium
COMPOSITION
FeAsS
FeS
CuFeS
PbS
ZnS
CdS
HgS
(Fe, Ni)g S8
Au Te£
(Au, Ag) Te2
Au
Se»
                          * Accompanies sulfur in sulfide minerals
                                        261

-------
Calcium, sodium, potassium, and magnesium are found at  con-
centrations of less than 100 mg/1 to over 1000 mg/1.

High  levels  of  soluble  metals  usually  result  from the
leaching processes, and  this  is  well-illustrated  by  the
cyanide  leach  process  in  the  gold industry.  Table V-27
summarizes the chemical  composition  and  raw  waste  loads
resulting  from four gold milling operations.  The processes
represented  include  amalgamation,   cyanidation/agitation-
leach,  cyanidation/vat  leach, and the cyanidation/llcarbon-
in-pulp" process.

Silver Ores

Water flow and the sources,  nature,  and  quantity  of  the
wastes  dissolved  in  the  water  during  the  processes of
silver-ore mining and beneficiation are  described  in  this
section.   Coproduct recovery of silver with gold is common,
and similar methods of extraction are employed.

Water Uses

The major use of water in the silver-ore milling industry is
in the beneficiation process, where it is required  for  the
operating  conditions of the process.  It is normally intro-
duced at the ore grinding stage of lode  ores   (see  process
diagrams. Section III) to produce a slurry which is amenable
to  pumping, sluicing, or classification for sizing and feed
into the concentration process.  In slurry form, the  ground
ore  is  most  amenable  to  beneficiation by the technology
currently  used  to  process  the  predominantly   low-grade
sulfide  silver ores—i.e., froth flotation.  A small amount
of silver  is  recovered  from  placer  gravels  by  gravity
methods, which also require water as a medium for separation
of the  fine and heavy particles.

Other  miscellaneous  uses  of water in silver mills are for
washing floors and  machinery  and  for  domestic  purposes.
Wash  water  is  normally  combined  with  the process waste
effluent but constitutes only a small fraction of the  total
effluent.  Some fresh water is also required for pump seals.

With the exception of hydralic mining and dredging, water is
not  normally directly used in mining operations; rather, it
is usually discharged  where  it  collects  as  an  indirect
result  of  a mining operation.  Cooling is required in some
underground mines  for the air  conditioning  systems.   This
water does not come into direct contact with the mine and is
normally discharged separately from the mine effluent.
                             262

-------
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
          MILLING OPERATIONS (Sheet 1 of 2)
MINE/MI LI
4102
(Arnelgernetion)
4101
(AlitMion LMch)
4106
(VM Leech)
4106 (Cerbon-
in-Pulp)
MINE/MILL
4102
(Amelgametron)
4101
(AgMMion LMch)
4105
IVit Leech)
41 OB (Cerbon-
in-Pulp)

MINE/MILL
4102
Umelgemetion)
4101
(Agitation Leech)
4106
(Vn Leach)
4106 (Carbon-
irtPulp)

MINE/MILL
4102
(A mi Igi ration)
4101
(Agitation Leech)
4106
IVit LMChJ
4106 (Carbon-
In-Pulp)
TSS
CONCEN-
TRATION
(mg/ll
496,000
646,000
2
486,000
WASTE LOAD
in kg/1000 mettle toiw
lib/1000 thott ton)
of concentrate produced
61.696,316.000
(123.3S0.630I
11,641,466.000
123.082,930,000)
860.000
(1.720.000)
4.7 x 10] 3
9.4 xlO11
in kg/1000 metric tent
(b/1000 short torn)
of ore milled
2.871.000
(5,742,000)
436,000
(872.000)
4
18)
4.171.000
(8.342,000)
TOC
CONCEN-
TRATION •
Inn/ 1 1
34.3
60.0
11.6
07.0
WASTE LOAD
in kg/1000 metric torn
lib/1000 ehort toiu)
of oonctntriM produced
4,275.000
(8,660,000)
1,069.000
(2.118,0001
4,940,000
(9,890.000)
94,100,000
(188.200,0001
in kg/1000 metric torn
lfe/1000 chart tonl
of ore milled
199
(398)
40
1801
46
(92)
830
11,6601
Cu
CONCEN-
TRATION •
I mo/ 41
0.03
0.17
0.7S
2.0
WASTE LOAD
in kg/1000 metric tent
lib/1000 chart tom)
ol eoncentrete produced
3,740
17,480)
3,600
(7.2001
335.000
(670,000)
1,941.000
13382,000)
in kg/1000 metric tone
Ik/1 000 ehort torn)
of ore milled
0.2
(0.4)
0.1
10.2)
3.1
(6.21
17
(341
Fe
CONCEN-
TRATION
Img/t)
1.6

-------
  TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
             MILLING OPERATIONS (Sheet 2 of 2)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
4105
(Vat LMch)
4106 (Carfaon-
in-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
4105
(Vat Leach)
4105 (Cartaon-
in-Pulpl
MINE/MILL
4102
(Amalgamation)
4101
(Agnation Leech)
4105
(Vat Laaeh)
4106 ICarbon-
In-Pulp)
Pb
CONCEN-
TRATION •
(mg/U
<0.1
<0.1
0.06
<0.1
WASTE LOAD
in kg/1000 metric font
(lb/1000 chart tom)
of eoncantrata produced
<12.500
k 25,0001
<2,100
K 4,200)
25.800
(51,600)
< 97.000
(<194,000)
in kg/1000 metric tent
(fc/IOOOthorttoni)
of ore milled
<0.6
(0.2)
<0.08
K0.16)
0.12
(0.24)
<0.9
«1.8)
Ha
CONCEN-
TRATION
(ma/ It
0.0011
—
0.004*
0.0042
WASTE LOAD
in kg/1 000 metric tom
(Ib/1000ihorttoni1
ol concentrate produced
137
(274)
_
1,700
(3.400)
4,070
(8,140)
In kg/1000 metric tom
lfc/1000 ehort torn)
of ore milled
0.0064
(0.0128)
—
0.016
(0.032)
0.036
(0.072)
SULFIDE
CONCEN-
TRATION
Img/lt)
<0.5
<0.5
0.2 •
1.7
WASTE LOAD
in kg/1000 metric tom
lib/1000 diort torn)
of concentrate produced
< 62,000
K124.000)
< 10,600
K21.200)
86,000
(172,0001
1,660,000
(3.300,000)
in kg/1 000 metric torn
(lb/1000 *hort torn)
of ore milled
<2.9
K5.8I
^0.4
K0.8)
0.8
(1.6)
15
(30)
Cd
CONCEN-
TRATION
tmg/U
<0.02
0.10
<0.01"
<0.02
WASTE LOAD
tn kg/1000 metric tom
(lb/1000 ihort ton*)
of concentrate produced
< 2.500
(cs.000)
2,100
(4,200)
< 4,300
(< 8.600)
< 19,400
K38.800)
in kg/1000 metric tom
(lb/1000 ihort torn)
of ore milled
<0.1
K0.2)
0.08
(0.16)
<0.04
K0.08I
<0.17
K0.34I
CYANIDE
CONCEN-
TRATION
(mg/U
<0.01
5.06
0.09
81.3
WASTE LOAD
in kg/ 1000 metric torn
(lb/1000 ihort tom)
of concentrate produced
<1,260
K2.500)
107,000
[214,0001
38,700
(77,400)
78.800,000
(158,000.000)
In kg/ 1000 metric tom
(lb/1000 ihort tom)
of ore milled
<0.06
kO.121
4
(8)
0.36
(0.72)
890
(1380)

•COMPANY DATA
                                  264

-------
Water  flows  of  some  silver mining and milling operations
visited during this program are presented in Figure V-23.

Sources of Wastes

There are two basic sources  of  effluents:  mines  and  the
beneficiation  process.   Mines  may  be  either open-pit or
underground operations.  In the case of  an  open  pit,  the
source  of  the  pit  discharge,  if  any, is precipitation,
runoff and ground-water infiltration into the pit.   Ground-
water  infiltration  is  the  primary  source  of  water  in
underground mines.  However, in some  cases,  sands  removed
from mill tailings are used to backfill stopes.  These sands
may  initially  contain  30 to 60 percent moisture, and this
water may constitute a major portion of the mine effluent.

The particular waste constituents present in a mine or  mill
discharge  are  a  function of the mineralogy and geology of
the ore body and the particular  milling  process  employed.
The  rate  and  extent  to which the minerals in an ore body
become  solubilized  are  normally  increased  by  a  mining
operation, due to the exposure of sulfide minerals and their
subsequent  oxidization  to  sulfuric acid.  At acid pH, the
potential for solubilization of most heavy metals is greatly
increased.  Not all mine discharges are  acid,  however;  in
those  cases  where  they  are  alkaline,  soluble  arsenic,
selenium, and/or molybdenum  may  present  problems  in  the
silver-ore mining and dressing industry.

Very  minor  production  of  silver  is obtained from placer
deposits as a byproduct of gold recovery.   Wastewater  from
placer  operations  is primarily the water which was used in
the  gravity  separation  processing  of  the   ore   and/or
hydraulic  mining  of  a  deposit.   The  process  water  is
generally discharged either directly to a  stream  or  to  a
settling  pond.   The  principal wastewater constituent from
any  placer  operation,  whether  silver,  gold,  or   other
materials,  is  high  loadings  of  suspended and settleable
solids.

Wastewater  emanating  from  silver  mills  consists  almost
entirely  of  process  water.  High suspended-solid loadings
are the most characteristic waste constituent of silver-mill
waste streams.  This is caused by fine grinding of the  ore,
making  it  amenable  to a particular beneficiation process.
In addition, the increased surface area of  the  ground  ore
enhances  the  possibility for solubilization and suspension
of  the  ore  minerals  and  gangue.   Although  the   total
dissolved  or  suspended  solid loading may not be extremely
                             265

-------
Figure V-23. WATER FLOW IN SILVER MINES AND MILLS
             549 m3/day
             (145,000 gpd)
         FLOTATION
                                                         •it 254 m3/day
                                                         (•it 67.000 gpd)
CHI
1

=tK t 	 ^
J 1,109m3/diy
-^^ f9Q1 flfMtanril


MILL
t

3,161 m3/d.y V P°ND J X. "»
(83I> 700 gprl) ^*^- _J-**^ ^^—

1,636 m3 (432.000 gall /day



                        1.132 m3 (299,000 «*l)/d«y

                       (a) MINE/MILL 4401
UNDERGROUND
MINE
1
DISCH
C CHI
1
kRGE
(775.000 gpd)
jf S4S m3/
_— -**^ (144,0
-------
high, the total heavy-metal concentration may be  relatively
high as a result of the mineralization of the ore being pro-
cessed.   These  heavy  metals,  the  suspended  solids, and
process   reagents   present   are   the   principal   waste
constituents of a mill waste stream.  In addition, depending
on  the process conditions, the waste stream may also have a
high or low pH.  The primary method of ore beneficiation  in
the  silver-ore milling industry is flotation.  As a result,
mill waste  streams  can  be  expected  to  contain  process
reagents.

Process Description - Mining

As   discussed   previously,   very   little  water  use  is
encountered in silver-ore  mining,  with  the  exception  of
dredging for recovery of silver from gold mining operations.
As a result of sampling and site visits to mining operations
in the silver mining industry, the waste constituents of raw
silver-mine  water were determined and are presented here in
Table V-28.  suspended-solid concentrations are  low,  while
dissolved-solid   concentrations   constitute  the  measured
total-solid load,  chlorides and sulfates are the  principal
dissolved-solid    constituents    observed.     Heavy-metal
concentrations observed are not notable, with the  exception
of total iron and total manganese.

Process Description - Milling

Milling  processes  of  silver  ore  which require water and
result in the waste loads present in mill water are:

    (1)   flotation,

    (2)   cyanidation, and

    (3)   amalgamation.

The selective froth flotation process  can  effectively  and
efficiently beneficiate almost any type and grade of sulfide
ore.   This  process  is  employed by mills 4401 and 4403 to
concentrate   the    silver-containing    sulfide    mineral
tetrahedrite and by mill 4402 to concentrate free silver and
the  silver  sulfide  mineral  argentite.  In this flotation
process, water is added  in  the  ore  grinding  circuit  to
produce  a  slurry  for  transporting  the  ore  through the
flotation circuit.   This slurry first  flows  through  tanks
(conditioners),   where   various   reagents  are  added  to
essentially cause the desired mineral to be more amenable to
flotation and the undesired minerals and gangue to  be  less
                            267

-------
          TABLE V-28. RAW WASTE CHARACTERISTICS OF SILVER
                      MINING OPERATIONS
PARAMETER
pH
Acidity
Alkalinity
Color
Turbidity (JTU)
TOS
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
MINE 4401
8.0*
10.2
85.0
47*
2.0
504
504
<2
240.8
11.9
17
4
0.085
<0.2
<0.07
< 0.002
<0.6
0.11
<0.02
46.0
<0.1
<0.02
0.33
<0.1
27.5
MINE 4403
.
4.2
76.2
<5*
2.2
622
622
<2
424.8
19.8
16
2
0.030
<0.2
<0.07
< 0.002
<0.5
0.09
<0.02
44.5
<0.1
<0.02
2.05
0.18
32.0
PARAMETER
Mn
Hfl
Ni
Tl
V
K
Se
Ag
Na
Sr
Ta
Ti
Zn
Sb
Mo
Chloride
Sulfate
Nitrate
Phosphate
Cyanide
Phenol
Fluoride
Kjeldahl N
Sulfide
SiO2
CONCENTRATION (mg/£)
MINE 4401
0.43
0.0020
0.09
<0.1
<0.2
8.0
0.126
<0.02
7.0
0.15
<0.3
<0.5
<0.02
<0.2
<0.2
4.2
175
2.45
0.3
<0.01
<0.01
0.26
<0.2
<0.5
9.75
MINE 4403
63
0.0004
0.06
<0.1

-------
 amenable.     These  reagents  are  generally  classified   as
 collectors, depressants,  and activators,  according to  their
 effect  on  the  ore minerals and gangue.  Also,  pH modifers
 are added   as  needed  to  control  the  conditions  of  the
 reaction.     Following  conditioning,   frothing   agents  are
 added,  and the slurry  is  transported  into  the  flotation
 cells,   where  it  is  mixed and  agitated by aerators at  the
 bottom of  the cells.   The collector and   activating  agents
 cause  the  desired  mineral  to   adhere   to  the rising  air
 bubbles and  collect  in   the  froth,   while  the  undesired
 minerals or gangue are either not collected or are caused to
 sink  by depressing agents.   The  froth  containing the silver
 mineral(s)  is collected by skimming from  the top  of  the
 flotation   cells   and  is  further upgraded by filtering  and
 thickening (Flow  sheets-Section III).

 Recovery of silver is also accomplished  by cyanidation   at
 mill  4105.    This process has been discussed  in  the  part of
 Section V  covering gold ores.

 Currently,  amalgamation is rarely used  for the  recovery   of
 silver   because most  of the ores  containing easily liberated
 silver  have been   depleted.    The  amalgamation   process   is
 discussed  in  sections III and V under gold-ore beneficiation
 methods.

 Quantity of Wastes
Discharge  of  water  seldom  exists  from  open-pit  mines.
However, most underground mines must  discharge  water,  and
the  average  volume  of  this water from the crossection of
mines visited ranges from less than 199 cubic meters per day
(50,000 gallons per day) to more than  13,248  cubic  meters
per   day    (3.5  million  gallons  per  day).   Where  mine
discharges occur, the  particular  metals  present  and  the
extent of their dissolution depend on the particular geology
and  mineralogy  of  the  ore  body  and  on  the  oxidation
potential and pH prevailing within the mine.  Concentrations
of metals in mine effluents are, therefore, quite  variable,
and  a  particular  metal  may range from below the limit of
detectability upwards to 2 ppm.  Calcium, sodium, potassium,
and magnesium may be present in quantities of  less  than  5
ppm  to  about  50  ppm  for each metal.  However, the heavy
metals are of primary concern, due to their  toxic  effects.
Minerals  known  to  be  found in association with silver in
nature are listed in Table V-29.

For the facilities visited, the volumes of the waste streams
discharging from mills  processing  silver  ore  range  from
                            269

-------
TABLE V-29. MAJOR MINERALS FOUND ASSOCIATED
          WITH SILVER ORES
MINERAL
Tetrahedrite
Tennantite
Galena
Sphalerite
Chalcopyrite
Pyrite
Naumannite
Greenockite/
Xanthochroite
Garnierite
Pentlandite
Native Bismuth
Argenite
Stephanite
Stibnite
COMPOSITION
(Cu. Fe. Ag)-|2 Sb4
(Cu, Fe, Ag)i2A$4
PbS
S13
S13
ZnS
Cu Fe $2
FeS
CdS
(Mg, Ni) O- Si 02 • x H20
(Fe, Ni)g S8
Bi
Ag2S
Ag5 Sb 84
Sb2S3
                 270

-------
 1,499  to  3,161  cubic  meters  per day (396,000  to 835,200
 gallons per day).   These waste streams carry solids loads  of
 272 to 1,542 metric tons per day (300 to  1,700  short   tons
 per  day)   from  a  mill,  depending  on  the mill.    Where
 underground mines  are present, the   coarser   solids  may  be
 removed  and used  for backfilling stopes in  the mine.   While
 the coarser  material  is  easily  settled,   the   very   fine
 particles  of  ground ore (slimes)  are normally suspended  to
 some extent in the  wastewater  and  often  present  removal
 problems.    The quantity  of  suspended solids present in a
 particular waste stream is  a function of the ore   type and
 mill  process  because  these  factors  determine  how finely
 ground the ore is.

 Heavy metals present in the minerals listed   in  Table   V-29
 may also be present in dissolved or dispersed colloidal form
 in   the  mill waste stream.   Since  the settlable solids, and
 most suspended solids,  are  collected and retained  in  tailing
 ponds,  the dissolved and dispersed  heavy metals  present   in
 the final  discharge are of  concern.   Depending on  the extent
 to  which they occur in the  ore body,  particular heavy metals
 may  be present in a mill  waste stream in the range  of from
 below detectable limits to  2 to  3   ppm.   Calcium,   sodium,
 potassium,     and     magnesium   normally   are    found   at
 concentrations of  10 to 250  ppm each.   In addition  to the
 suspended   solids  and dissolved metals,  reagents used in the
 mill beneficiation  process  also add  to the pollutant  loading
 of  the  waste stream.  The particular  reagents  used  are  a
 function of  the process  employed to  concentrate the ore.   in
 the  silver  milling industry,  the various flotation reagents
 (frothers,  collectors,  pH modifiers, activating agents,  and
 depressants)    are    the most   prominent reagents  of  the
 flotation  process.   Table V-30   indicates the  quantity   of
 these reagents consumed  per  ton  of ore  milled.  A portion  of
 these   reagents which are consumed in  the  process is present
 in  the  waste  stream.  Note that  a large  number of   compounds
 fall    under    the   more  general  categories  of   frothers,
 collectors,   etc.    At   any   one   mill,   the   particular
 combination  of reagents  used  is  normally chosen on  the basis
 of   research   conducted   to   determine  the conditions under
 which recovery is optimized.  While flotation processes  are
 generally  similar,  they differ specifically with regard to
 the particular reagent combinations.  This is  attributable,
 in  part,  to  the highly variable mineralization of the ore
bodies  exploited.  Waste  characterizations  and  raw  waste
 loadings   for   mill   effluents  employing  flotation  and
cyanidation in four  mills  are  presented  in  Table  v-31.
These characterizations and loadings are based upon analysis
of raw waste samples collected during site visits.
                            271

-------
TABLE V-30. FLOTATION REAGENTS USED BY THREE MILLS TO BENEFICIATE
          SILVER-CONTAINING MINERAL TETRAHEDRITE (MILLS4401 AND
          4403) AND NATIVE SILVER AND ARGENTITE (MILL 4402)

REAGENT


M.I. B.C. (Methylisobutylcarbinol)
D-52
Z-200 (Isopropl ethylthiocarbamate)
Lime (Calcium oxide)

Sodium cyanide

PURPOSE

MILL 4401
Frother
Frother
Collector
pH Modifier
and Depressant
Depressant
CONSUMPTION
g/metric ton
ore milled

0.00498
0.00746
0.00187
0.109

0.00498
Ib/short ton
ore milled

0.00000995
0.0000149
0.00000373
0.000219

0.00000995
MILL 4402
Cresylic acid
Mineral oil
Dowfroth 250 (Polypropylene glycol
methyl ethers)
Aerofroth 71 (Mixture of 6/9-carbon
alcohols)
Aerofloat 242 (Essentially Aryl
dithiophosphoric acid)
Aero Promoter 404 (Mixture of
Sulfhydryl type compounds)
Z-6 (Potassium amyl xanthate)
Sulfuric acid
Soda ash (Sodium carbonate)
Caustic soda (Sodium hydroxide)
Hydrated lime (Calcium hydroxide)
Frother
Frother
Frother

Frother

Collector

Collector

Collector
pH Modifier
pH Modifier
pH Modifier
pH Modifier
2.83
6.9
0.545

10

90

1.82

70
250
1,260
3.03
320
0.00566
0.0138
0.00109

0.02

0.18

0.00363

0.13
0.49
2.51
0.00605
0.64
MILL 4403
Cresylic acid
Hardwood tar oils
M.I.B.C.
Aerofloat 242
Aerofloat 31 (Essentially Aryl
dithiophosphoric acid)
Xanthate Z-11 (Sodium ethyl xanthate)
Aero S-3477
Zinc sulfate
Sodium sulfite
Frother
Frother
Frother
Collector
Collector

Collector
Collector
Depressant
Depressant
1.25
1.25
3.75
7.51
5.00

2.50
25
150
200
0.0025
0.0025
0.0075
0.015
0.01

0.005
0.05
0.3
0.4
                         272

-------
       TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
                  AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 1 of 2)
MILL
4401
4403
44O2
410S
MILL
4401
4403
4402
4106

MILL
4401
4403
4402
4105

MILL
4401
4403
4402
4106
TSS
CONCEN-
TRATION
(mg/JU
660,000
203,000
•0.000
2
WASTE LOAD
in kg/1 000 metric toni
(lb/1000 rfiart teni)
of conemtnM produced
99.000.000
(108.000,000)
33,901,000
(62,802,000)
•,720.000
119,440.0001
86,000
(1.720.0001
in kg/1000 mitrie torn
(to/1000 ifiorl toiw)
of or* milled
2.476.000
(4,960,000)
1,543.000
(3.086.000)
990,000
(1.980.000)
4
(81
TOC
CONCEN-
TRATION •
(mtlll
22.0
24.0
29.0
11.6
WASTE LOAD
in kg/ 1000 nwtrie tarn
(lb/1000 *ort torn)
of eoncontrn* produced
4.000
(8,000)
4.000
(8,000)
3.130
(8,260)
4,940,000
(9,800,000)
in kg/1000 movie ton*
(to/1000 ihon torn)
of or* milled
100
(2001
180
(360)
320
(640)
46
192)
Cu
CONCEN-
TRATION •
(mg/a)
0.25
0.03
0.22
0.78
WASTE LOAD
in kg/1000 nwtrie torn
(lb/1000 rtiort tons)
of concentrate produced
46
(90)
6
1101
24
1481
335,000
670,000
In kg/1 000 metric toni
(lb/1000 short tarn)
of or* milled
1
(21
0.23
(0.46)
2.4
(4.8)
3.1
(6.2)
Pb
CONCEN-
TRATION •
(mg/U
<0.1
<0.1
0.56
0.06
WASTE LOAD
in kg/1000 nwtrle tont
(lb/1000 short tons)
of eoneentnt* produced
<18
K36I
<17
K34)
60
(120)
25,800
(51.600)
In kg/1000 m*trle Mm
< to/1000 thort torn)
of or* mlll*d
<0.45
K0.90)
<0.8
K1.6I
6.2
(12.4)
0.12
10.24)

CONCEN-
TRATION
(mg/il
470
584
960
1230
TDS
WASTE LOAD
in kg/1000 imtric torn
(lb/1 000 ihort tans)
of concentrate produced
84.600
1169,2001
97.600
(195,000)
104,000
(208.000)
629,000,000
(1.060.000.000)
in kg/1000 m*tric torn
(lb/1000 ihoRtonil
of or* millod
2.110
(4.220)
4,440
(83801
10.600
(21.200I
4300
(9,800)
COO
CONCEN-
TRATION
Img/e)
69.S
15.9
22.7
222
WASTE LOAD
in kg/1000 metric torn
(lb/1000 than ton*)
of eonewitnt* produced
10,700
121,400)
2,700
(6,400)
2.460
14.900)
95,500,000
(191,000,000)
in kg/1000 mMrie to to
(lb/1000 ihort tonil
of or* mill*d
270
1640)
120
(240)
250
(6001
889
(1,800)
Zn
CONCEN-
TRATION
Img/U
0.02
0.17
0.37

WASTE LOAD
in kg/ 1000 nwtrie tons
(lb/1000 ihort ton*)
of eanc*ntr*M produced
3.6
(7.2)
28
(56)
40
(80)
-
In kg/1000 nwtric torn
(lb/1000 ihort tani)
ofor*mill*d
0.09
(0.18)
1.3
(2.6)
4.1
(62)
-
Al
CONCEN-
TRATION
(mg/JU
<0.07
<0.07
0.07
3.60*
WASTE LOAD
In kg/1000 nwtric tont

-------
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
          AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 2 of 2)
MILL
440t
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
1105
(Company Data
onlyl
MILL
4401
4403
4402
4105
{Company Data
only)
MILL
44O1
4403
4402
4105
(Company Data
only)
Hs
CONCEN-
TRATION
tmtfUt
0.0024
0.0008
O.I 490
0004
WASTE LOAD
in kg/1000 metric font
Ub/1 000 thort font)
of concentrate produced
0.4
(0.8)
0,13
10.26)
16
132)
259
1518)
in kg/1000 metric mm
tlb/1OOO>hortton>)
of ore milled
0.01
(0.02)
0.5
11.01
1.6
(3.2)
0.024
10.048)
Ta
CONCEN.
THATION
W HI
<0.3
<0.3
<0.3
"
WASTE LOAD
in kg/IOOO matric tons
(lb/1 000 sh on tons!
of concentrate produced
<54
K108I
 t\
<0.02
<0.02
)
of ore milled
<0.9
(0.81
<4
(<8I
6
(12)
<0.3
K0.6I
Cd
CONCEN-
TRATION
tmg/fcj
<0.02
<0.02
<0.02
<0.01
WASTE LOAD
in kg/IOOO metric tons
(lb/1000 short tons)
of concentrate produced
<36
1C 7. 21
<3.3
K6.6)
<22
K4.4)
< 6,500
K 13,0001
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
<0.09
K0.18)
<0.1S
K0.30I

-------
 Bauxite  Ores

 Water  handling  and  quantity   of  wastewater  flow  within
 surface    bauxite   mines   are   largely   dependent   upon
 precipitation   patterns  and  local topography.  Topographic
 conditions are  often  modified   by  precautionary  measures,
 such   as   diversion    ditching,  disposal  of  undesirable
 materials,  regrading,   and  revegetation.    In   contrast,
 underground   mine   infiltration  occurs  as  a  result  of
 controlled drainage  of the unconsolidated sands in the over-
 burden.  These  sands are under considerable water  pressure,
 and  catastrophic collapses  of sand and water may occur if
 effective  drainage   is  not  undertaken.   Gradual  drainage
 accumulates  in the  mines and is pumped out periodically for
 treatment  and discharge.  As  in  other  mining  categories,
 dewatering is  an   economic,  practical,  and safe-practice
 necessity.

 Beneficiation of  bauxite ores  is  not  currently  practiced
 beyond size reduction, crushing and grinding.  No water use,
 other than dust suppression, results.

 Mining Technique  and Sources of Wastewater

 Open-Pit Mining.   The sequence of operations that occurs in
 a typical  open-pit mining operation is that the mine site is
 cleared of trees, brush, and overburden and then stripped to
 expose the ore.   Timber  values are often obtained from areas
 undergoing site preparation.

 Depending  upon  the  consolidation  of  the overburden, the
 material may be vertically drilled  from  the  surface,  and
 explosive  charges—generally,  ammonium nitrate—are placed
 for blasting.   This  sufficiently  fractures  the  overburden
 material   to  permit  its  removal by earthmoving equipment,
 such as draglines, shovels, and scrapers.  Removal  of  this
 overburden  takes the greatest amount of time and frequently
 requires the largest equipment.

 Following removal of the overburden material, the bauxite is
 drilled,   blasted,   and  loaded  into  haulage  trucks   for
 transport  to   the   vicinity  of  the  refinery.    Extracted
 overburden or spoils are often placed in abandoned  pits  or
 other  convenient  locations,  where some attempts have been
 made at revegetation.

Regardless of the method of mining,  water  use  at  the  two
 existing   operations    is   generally   limited   to   dust
 suppression.   Water removal is required because drainage  is
                            275

-------
a  hindrance  to  mining.   As  such,  mine  dewatering  and
handling are a required part  of  the  mining  plan  at  all
bauxite mines.

The  bauxite  mining  industry  presently  discharges  about
57,000 cubic meters (15 million gallons)   of  mine  drainage
daily  at  two operations.  The open-pit mining technique is
largely  responsible  for  accumulation   of   this   water.
Underground mining accounts for only a fraction of a percent
of  the total.  In association with the open-pit approach to
bauxite mining, water drainage and accumulation occur during
the processes of mine site  preparation  and  during  active
mining.

For  the  open-pit mine represented in Figure V-24, rainfall
and ground water intercepted by the terrain undergoing  site
preparation are diverted to outlying sumps for transfer to a
main  collection  sump.   Diversion  ditching  and  drainage
ditches segregate most surface water, depending upon whether
it has contacted lignite-containing material.   Contaminated
water  is directed to the treatment plant, while fresh water
is diverted  to  other  areas.   At  other  mines,  drainage
occurring during site preparation and mining is not treated,
and segregation of polluted and unpolluted waters may or may
not be practiced.

Water from the main collection sump is pumped to a series of
settling  ponds,  where removal of coarse suspended material
occurs.  These ponds also aid in regulation of flow  to  the
treatment  plant.   A  small  sludge  pond  receives treated
wastewater for final settling before discharge.

Bauxite mining operations characteristically utilize several
pits simultaneously and may practice site  preparation  con-
current  with  mining.   Since  both  bauxite producers have
large land holdings (approximately 4,050 hectares or  10,000
acres), mines and site-preparation activities may be located
in  remote  areas,  where  the  economics of piping raw mine
drainage to a central treatment plant are  unfeasible.   For
larger quantities of mine drainage in remote areas, separate
treatment  plants  appear  necessary.   Portable  and  semi-
portable  treatment  plants  appear  feasible  for  treating
smaller accumulations of wastewater at times when pumping of
mine water for discharge is required.

Underground  Mining.    Underground mining occurs where low-
silica bauxite is located deep enough under the land surface
so that economical removal of overburden  is  not  feasible.
The  underground  operations  have been historically notable
                            276

-------
Figure V-24. PROCESS AND WASTEWATER FLOW DIAGRAM FOR OPEN-PIT BAUXITE
           MINE 5101
   EXPLORATION AND ORE-BODY EVALUATION:
       GEOLOGICAL SURVEY
       TEST DRILLING
SITE PREPARATION:
    CLEARING
    STRIPPING
                                     RUNOFF
                                       AND
                                  GROUND WATER
               MINING:
                   BLASTING
                   LOADING
                   HAULING
                        RUNOFF
                         AND
                    GROUND WATER
        MILLING:
            CRUSHING AND GRINDING
            STORAGE
            BLENDING
^SE

          2.76 m3/metric ton
          (664 gal/short ton)
          BAUXITE
                               SETTLING
         REFINING:
             COMBINATION PROCESS
                                        2.76 m3/metric ton
                                        (664 gal/short ton) BAUXITE
                                           WATER TREATMENT
                                                 PLANT
                                        2.76 m3/metric ton
                                        (664 gal/short ton) BAUXITE
 PRODUCTION - 2,594 metric tons (2,860 short tons) per day
 WATER TREATED DAILY - 7,165 m3 (1,900,000 gal)
                                       2.76 m3/metric ton
                                       (664 gal/short ton) BAUXITE
                                              DISCHARGE
                              277

-------
for relatively high recovery of bauxite under  adverse  con-
ditions   of  unconsolidated  water-bearing  overburden  and
unstable clay floors.   Controlled  caving,  timbered  stope
walls,  and  efficient drainage systems—both on the surface
and  underground—have  minimized  the  problems  and   have
resulted in efficient ore recovery.

Initially,  shafts are sunk to provide access to the bauxite
deposits, and drifts are driven  into  the  sections  to  be
mined.   A room-and-pillar technique is then used to support
the mine roof  and  prevent  surface  subsidence  above  the
workings.   Configurations of rooms and pillars are designed
to consider roof  conditions,  equipment  utilized,  haulage
gradients, and other physical factors.

Ore  is  removed from the deposits by means of a "continuous
miner," a ripping-type machine which cuts  bauxite  directly
from  the ore face and loads it into shuttle cars behind the
machine.   Initial  development  of  the  room  leaves  much
bauxite  in  pillars, and it has been the practice to remove
the pillars and induce caving along a  retreating  caveline.
However,  resultant roof collapse and fracturing can greatly
increase overburden  permeability,  facilitating  mine-water
infiltration   and  subsequently  increasing  mine  drainage
problems.  Recent changes in mining technique have  resulted
in  a cessation of induced caving, but drainage still occurs
in the mines.

Raw mine drainage  accumulates  slowly  in  the  underground
mines  and  is  a  result  of controlled drainage.  The mine
water is pumped to the  surface  at  regular  intervals  for
treatment,   with   subsequent   settling   and   discharge.
Excessive water in the underground mine can lead to  wetting
of  clays  located in drift floors and in resultant upheaval
of the floor.

The most  influential  factor  which  determines  mine-water
drainage  characteristics is mineralization of the substrata
through  which  the  drainage  percolates.    The   existing
underground   mine  receives  drainage  which  has  migrated
through strata of unconsolidated sands  and  clays,  whereas
open-pit  drainage is exposed to sulfide-bearing minerals in
the  soil.   As  shown  in  this   section,   open-pit   and
underground   mine   drainages   differ   qualitatively  and
quantitatively; but, as a factor affecting raw mine-drainage
characteristics,  mineralization  does  not   constitute   a
sufficient basis for subcategorization.
                            278

-------
 Study  of  NPDES permit applications and analysis of samples
 secured during mine visitations revealed  that  the  bauxite
 mining  industry  generates two distinct classes of raw mine
 drainage:  (1)   Acid  or  ferruginous,  and  (2)   alkaline—
 determined  principally  by  the substrata through which the
 drainage flows.  Acid or ferruginous raw  mine  drainage  is
 defined as untreated drainage exhibiting a pH of less than 6
 or  a total iron content of more than 10 mg/liter.   Raw mine
 drainage is defined as alkaline when the untreated  drainage
 has a pH of more than 6 or a total iron content of less than
 10 mg/liter.

 At  present,   the  class  of  raw  mine drainage corresponds
 closely with  mining  technique,   and  open-pit  drainage  is
 characteristically  acid.    Acid  mine  water is produced by
 oxidation of  pyrite contained in lignite present in the soil
 overburden of the area.

 Acid mine drainage with pH generally in the range of 2  to  U
 is produced in  the presence of abundant water.   The sulfuric
 acid  and  ferric  sulfate  formed  dissolve other  minerals,
 including those containing aluminum,  calcium,  manganese,  and
 zinc.

 In areas  undisturbed by mining operations,   these  reactions
 occur  because   the  circulating  ground water contains  some
 dissolved oxygen,  but the   reaction   rate  is   rather  slow.
 Mining activity  which disturbs  the  surface  of the ground
 creates conditions  for a greatly accelerated rate of sulfide
 mineral dissolution.

 Alkaline   mine   water,   characteristic    of   the  existing
 underground  mine,   may  migrate  through the lignitic clays
 located in strata overlying the  mine  before  collecting  in
 the  mine,  but  pH  is  generally  around  7.5.  Data evaluation
 reveals that underground mine  drainage  differs  significantly
 from open-pit mine  drainage  (acid), as  shown in Tables V-32,
 V-33,  and  V-34.

 Though  these mine drainages differ with   respect  to  mining
 technique,  all mine drainages sampled proved to be amenable
 to efficient removal of  selected  pollutant  parameters  by
 liming   and   settling,   as    exhibited  in  Section  VII.
Attainable  treated-effluent  concentrations  are   directly
 related  to treatment efficiency, and these two interrelated
factors do not justify establishment of subcategories.

Due to acid  conditions  and  general  disruption  of  soils
caused  by  stripping  of  overburden  for  open-pit  mines.
                            279

-------
TABLE V-32. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID RAW
           MINE DRAINAGE FROM OPEN-PIT MINE 5101
PARAMETER
pH
Specific Conductance
Acidity
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION 
-------
TABLE V-34. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ALKALINE RAW
          MINE DRAINAGE FROM UNDERGROUND MINE 5101
PARAMETER
PH
Specific Conductance
Alkalinity
IDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sr
Sulfate
Fluoride
CONCENTRATION (mg/e,)
THIS STUDY
7.2*
1,260f
280
780
<2
1.4
0.88
0.8
<0.02
<0.02
1.82
228.8
1.25
INDUSTRY DATA
7.6*

222
862
26
2.3
0.87
<0.05
<0.01
<0.01

246
0.07
NPDES PERMIT
APPLICATION
7.8*
3,281 f
150
550
300
5.0
5.0
2.0
1.6
0.01

50
2.5
   •Value in pH units

    Value in micromhos
                           281

-------
natural revegetation proceeds extremely slowly.   The lack of
vegetative cover aids in accelerating the weathering of  the
unconsolidated  overburden and compounds the acid mine-water
situation.  Extensive furrowed faces  of  exposed  silt  and
sandy  clays  are  evidence of the erosion which infuses the
mine  water  with  particulate  matter.   Fortunately,  this
material  settles  rapidly,  either  in  outlying pits or in
pretreatment settling basins, and presents  no  nuisance  to
properly treated discharges.

Raw Waste Loading

As discussed earlier in this Section, effluents from bauxite
mining operations are unrelated, or only indirectly related,
to  production  quantities  and exhibit broad variation from
mine to mine.  Loadings have been  calculated  for  open-pit
mine 5101 and underground mine 5101, as shown in Tables V-35
and V-36.

Potential  Uses of Mine Water-   Since both domestic bauxite
mines are intimately associated with refineries, the plausi-
bility of utilizing  a  percentage  of  mine  water  in  the
refinery   arises.   Though  the  bauxite  refining  process
intrinsically has  a  substantial  negative  water  balance,
water  is  supplied  from  rainfall on the brown-mud lake or
from fresh-water impoundments.  More importantly, the brown-
mud-lake water posseses a high  pH  (approximately  10)  and
remains amenable to recycling in the caustic leach process.

To  minimize  the effects of dissolved salts in the refining
circuit, evaporators are sometimes used to remove impurities
from  spent  liguor.   However,  mine  water  contains  many
dissolved  constituents  (particularly,  sulfate)  in  large
quantities,  the  effects  of  which  are   detrimental   or
undetermined  at  this  time.   The exacting requirements of
purified alumina, and the specific  process  nature  of  the
refinery,  largely  preclude the  introduction of new intake
constituents  via  alternative  water  sources   (treated  or
untreated mine water) at this time.

Ferroalloy Ores

Waste  characterization  for  the  ferroalloy-ore mining and
milling industry has, of necessity, been based primarily  on
presently   active   operations.   Since  these  comprise  a
somewhat limited set, many types of operations which may  or
will be active in the future were not available for detailed
waste  characterization.   Sites  visited  in the ferroalloy
segment are organized by category and product in Table V-37.
                            282

-------
      TABLE V-35. WASTEWATER AND RAW WASTE LOAD FOR OPEN-PIT MINE 5101
PARAMETER
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION
(mg/& )
IN WASTEWATER
560 to 1290
< 2 to 42
7.0 to 129.1
2.83 to 9.75
2.76 to 52.3
0.82 to 1.19
0.3 to 0.37
490 to 700
0.048 to 1.4
RAW WASTE LOAD
kg/metric ton
1.55 to 3.56
< 0.006 to 0.1 2
0.02 to 0.36
0.008 to 0.027
0.008 to 0.14
0.002 to 0.003
0.0008 to 0.001
1.35 to 1.93
0.0001 to 0.004
Ib/short ton
3.10 to 7.12
< 0.012 to 0.24
0.04 to 0.72
0.016 to 0.054
0.016 to 0.28
0.004 to 0.006
0.001 6 to 0.002
2.70 to 3.86
0.0002 to 0.008
    Daily flow of wastewater = 7,165 m3 (1,900.000 gal)
    Daily mine production » 2,594 metric tons (2,860 short tons)
TABLE V-36. WASTEWATER AND RAW WASTE LOAD FOR UNDERGROUND MINE 5101
PARAMETER
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION
(mg/£)
IN WASTEWATER
550 to 862
< 2 to 300
1.4 to 5.0
0.87 to 5.0
< 0.05 to 2.0
< 0.01 to 1.6
< 0.01 to 0.01
50 to 246
0.07 to 2.5
RAW WASTE LOAD
kg/metric ton
0.12 to 0.18
< 0.0004 to 0.06
0.0003 to 0.001
0.0002 to 0.001
< 0.00001 to 0.0004
< 0.000002 to 0.0003
< 0.000002 to 0.000002
0.01 to 0.05
0.00001 to 0.0005
Ib/short ton
0.24 to 0.36
<0.0008 to 0.12
0.0006 to 0.002
0.0004 to 0.002
<0.00002 to 0.0008
<0.000004 to 0.0006
< 0.000004 to 0.000004
0.02 to 0.10
0.00002 to 0.0010
  Daily flow of wastewater = 83 m3 (22,000 gal)
  Daily mine production - 390 metric tons (430 short tons)
                                 283

-------
      TABLE V-37. TYPES OF OPERATIONS VISITED AND ANTICIPATED-
                  FERROALLOY-ORE MINING AND DRESSING  INDUSTRY
METAL ORE
MINED/MILLED
Chromium
Cobalt
Columbium and
Tantalum
Manganese
Molybdenum
Nickel
Tungsten
Vanadium
MINE
O
X
X
X
V(3)
V(1>»
V(2>
V(1)
MILL
Category 1
(< 5,000 metric tons
[5,512 short tons] per year)






X

Category 2
(Physical
Concentration)
0

X
X

V
V

Category 3
(Flotation)

X
X
X
V(3)
X
V

Category 4
< Leaching)
0

X
X


V
V
(  )  indicates number of operations visited
*    seasonal mine discharge, not flowing during visit
X   likely in the future; currently, not operating
O   most likely process, if ever operated in the U.S.
V   types of operations visited
                                   284

-------
 Since some sites produce multiple  products,   and/or  employ
 multiple  beneficiation  processes,   they are  represented by
 more  than  one  entry  in  the  table.     Where    possible,
 segregated as well as combined waste streams were  sampled at
 such  operations.   Table V-37  also shows  types of  operations
 considered likely in the U.S.  in  the  future   (marked  with
 x's),  as  well  as  those  which  represent likely recovery
 processes for ores not expected to be  worked   soon  (marked
 with  o's).    characteristics   of wastes  from  the  latter two
 groups of operations have been determined,  where   possible,
 from  historical data;  probable ore  constituents and process
 characteristics; and examination of  waste  streams  expected
 to   be  similar (for example,  gravity processors of iron ore
 as  indicators for gravity manganiferous ore operations).

 Treatment of the  individual  process  descriptions  by   ore
 category,  as  adhered  to previously in  this  report,  is not
 used here.   Instead, because of the  wide  diversity  of  ores
 encountered,   the   general  character   of mine   and  mill
 effluents is discussed,  followed by  process descriptions and
 raw  waste    characteristics   of   several   representative
 operations.

 General  Waste Characteristics

 Ferroalloy    mining   and milling  wastewater  streams   are
 generally characterized  by:

     (1)   High suspended-solid  loads

     (2)   High volume

     (3)   Low concentrations  of  most  dissolved pollutants.


 The  large amounts  of material to   be  handled  per  unit  of
 metal  recovered,   the   necessity  to  grind  ore  to  small
 particle  sizes  to  liberate  values,   and   the   general
 application   of  wet separation   and  transport  techniques
 result in the generation  of  large  volumes of effluent  water
 bearing   high concentrations  of  finely divided rock, which
 must be removed  prior to  discharge.  In addition,  the  waste
 stream  is generally contaminated  to some extent by a number
 of dissolved  substances,  derived from the ore  processed  or
 from reagent  additions in the mill.  Total concentrations of
 dissolved    solids   vary  but,  except  where   leaching  is
 practiced, rarely  exceed  2,500 mg/1,  with  Ca-n-,   Na+,  K+,
Mg++,  C03—,  and   S04  accounting for nearly  all dissolved
                            285

-------
materials.  Heavy metals and other notably  toxic  materials
rarely exceed 10 mg/1 in the untreated waste stream.

The  volume  of  effluent  from  both mines and mills may be
strongly influenced by factors of topography and climate and
is frequently subject to seasonal fluctuations.   In  mines,
the  water  flow  depends  on  the  flow in natural aquifers
intercepted and may be highly variable.   Water  other  than
process  water  enters the mill effluent stream primarily by
way of the tailing ponds (and/or settling ponds), which  are
almost  universally  employed.   These  water  contributions
result from direct precipitation on the  pond,  from  runoff
from  surrounding  areas or even from seepage,  and are only
partially amenable to elimination or control.

A number of operations or practices common to  many  milling
operations  in  this  category  involve  the  use of contact
process water and contribute to the  waste-stream  pollutant
load.   These  include  ore washing, grinding, cycloning and
classification, ore and tail transport as a slurry, and  the
use  of  wet  dust-control  methods (such as scrubbers).  In
terms of pollutants contributed to the effluent stream,  all
of  these  processes  are  essentially the same.   Contact of
water with  finely  divided  ore,  gangue,  or  concentrates
results in the suspension of solids in the waste stream, and
in  the  solution of some ore constituents in the water.  In
general, total levels of dissolved material  resulting  from
these  processes  are  quite  low,  but  specific substances
(especially, some heavy metals)  may dissolve to a sufficient
degree to  require  treatment.   These  processes  may  also
result  in  the presence of oil and grease from machinery in
the wastewater stream.  Good  housekeeping  and  maintenance
practice  should  prevent  this  contribution  from becoming
significant.

Ore roasting may be practiced as a part of  some  processing
schemes to alter physical or chemical properties of the ore.
In   current   practice,  it  is  used  to  change  magnetic
properties in iron-ore processing in the  U.S.  and  in  the
past  was  used  to alter magnetic/electrostatic behavior of
columbium and tantalum  ores.   Roasting  is  also  used  in
processing  vanadium ores to render vanadium values soluble.
Although a dry process, roasting generally entails  the  use
of scrubbers for air pollution control.  Dissolved fumes and
ore  components  rendered  soluble  by  roasting  which  are
captured in the scrubber  thus  become  part  of  the  waste
stream.   This  scrubber water may constitute an appreciable
fraction of the total  plant  effluent  and  may  contribute
significantly   to  the  total  pollutant  load.    One  mill
                            286

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surveyed contributes 0.8 ton of contaminated scrubber
water per ton of ore processed.
bleed
Effluents from some ferroalloy mining and milling operations
are  complicated  by  other  operations  performed  on-site.
Thus, smelting  and  refining  at  one  site,  and  chemical
purification  at  another,  contribute  significantly to the
wastewater   generated   at   two   current   ferroalloy-ore
processing  plants.  Since waste streams are not segregated,
and the other processes involve wastes of somewhat different
character then those normally associated with ore mining and
beneficiation, such operations may pose special problems  in
effluent limitation development.

An  additional  component  of  the mill waste stream at some
sites which is not related to the milling process is sewage.
The use of the mill tailing basin as  a  treatment  location
for domestic wastes can result in unusually high levels of a
number  of pollutants in the effluent stream, including NH3,
COD, BOD, and TOC.  At other sites, effluent  from  separate
domestic  waste-treatment  facilities  may  be combined with
mine or mill effluents, raising levels of NH3_, BOD, TOC,  or
residual chlorine.

Sources of_ Wastes  - Mine Effluents

Factors  affecting  pollution  levels  in  mine  water flows
include:

    (1)   Contact with broken rock and dust within the  mine,
         resulting   in  suspended-solid  and  dissolved-ore
         constituents.


    (2)   Oxidation of reduced  (especially,  sulfide)   ores,
         producing acid and increased soluble material.

    (3)   Blasting decomposition products, resulting in  NHJ,
         N03_, and COD loads in the effluent.              "~

    (4)   Machinery operation, resulting in oil and grease.

    (5)   Percolation of water through strata above the mine,
         which may contribute dissolved materials not  found
         in the ore.

As  discussed  previously,  variable  (and,  sometimes, very
high)  flow rates are characteristic of mine  discharges  and
can strongly influence the economics of treatment.   Data for
                            287

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mine  flows  sampled  in the development of these guidelines
are presented in Table V-38.  Observed  mine  flows  in  the
industry  range  from  zero to approximately 36 cubic meters
(9,510 gallons)  per  minute.   Generally,  total  levels  of
dissolved  solids are not great, ranging from 10 to 1400 ppm
in untreated mine waters.   Total  levels  of  some  metals,
however,  can  be  appreciable,  as the data below show, for
some maximum observed levels (in mg/1).

         Al   9.4 ..          Mo   0.5

         Cu   3.8            Pb   0.19

         Fe   17             Zn   0.47

         Mn   5.5

In addition, oil and grease levels as high as 14  mg/1,  and
COD  values  up  to  91  mg/1,  were observed.  Since simple
settling treatment greatly reduces most of the  above  metal
values,  it  is  concluded  that most of metals present were
contributed in the form of suspended solids.   There  is  no
apparent  correlation  between  waste content or flow volume
and production for mine effluents.

Sources of Wastes  - Mill Effluents

Physical  Processing  Mill  Effluents.    In  general,  mills
practicing purely physical ore beneficiation yield a minimal
set  of  pollutants.   Separation  in jigs, tables, spirals,
etc., contributes to pollution in the same  fashion  as  the
general   practices  of  grinding  and  transport--that  is,
through contact of ore and water.  Suspended solids are  the
dominant  waste  constituent,  although,  as in mine wastes,
some  dissolved  metals  (particularly,  those   with   high
toxicity)  may require treatment.  Roasting may be practiced
in some future operations to alter  magnetic  properties  of
ores.   As  discussed  previously,  this  could  change  the
effluent somewhat, by  increasing  solubility  of  some  ore
components, and by introducing water from scrubbers used for
dust   and   fume   control   on   roasting   ovens.   Since
solubilization is generally undesirable in such  operations,
the  very high total dissolved solid values observed at mill
6107 are not anticipated elsewhere.

No sites in  the  ferroalloy  category  actually  practicing
purely  physical  beneficiation  of  ore  using  water  were
visited and sampled in developing  these  guidelines,  since
none  could  be  identified.   A  mine/mill/smelter  complex
                            288

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TABLE V-38. CHEMICAL CHARACTERISTICS OF RAW MINE WATER IN
          FERROALLOY INDUSTRY
MINE
6102
6103
6104
6107
PRODUCT
Mo, W
Mo
W, Mo
V
FLOW
(rn /min (gpm))
2.65 (700)
6.43 (1.700)
34.06 (9.000)
11.3513,000)
pH
4.5
7.0
6.5
7.3
CONCENTRATION (mg/jU
Oil and
GrMM
14
1.0
2.0
•
Nitrate
-
0.15
0.12
•
Fluoride
44.5
4£
0.52
•
As
<0.01
<0,01
<0.07
<0.07
Cd
0.07
<0.01
<0.01
<0.005
Cu
3.8
0.06
<0.02
<0.02
Mn
5.3
5.5
0.21
6.8
Mo
0.5
<0.1
<0.1
<0.1
Pb
0.06
0.19
0.14
-
V
<0.5
<0.5
<0.5
•
Zn
7.0
0.47
0.05
0.09
                       289

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recovering nickel  (mill 6106) which  was  visited,  however,
produces   an   effluent   which  is  felt  to  be  somewhat
representative, since water contacts ore in  belt  washing—
and gangue in slag granulation-operations at that site.  Raw
waste  data  for that operation illustrate the generally low
level  of  dissolved  materials  in  effluents  from   these
operations.   In  general,  these  effluents  pose  no major
treatment problems and are generally suitable for recycle to
the process after  minimal  treatment  to  remove  suspended
solids.

Flotation Mill Effluents.   The practice of flotation adds a
wide variety of process reagents, including acids and bases,
toxicants  (such as cyanide), oils and greases, surfactants,
and complex organics (including amines and  xanthates) .   In
addition   to  finer  grinding  of  ore  than  for  physical
separation, and modified pH,  the presence  of  reagents  may
increase the degree of solution of ore components.

Flotation  reagents  pose  particular  problems  in effluent
limitation and treatment.  Many are complex organics used in
small quantities, whose fates and effects when  released  to
the  environment  are uncertain.  Even their analysis is not
simple (References 28 and 29). Historically,  effluent  data
are  widely  available  only  for  cyanide  among  the  many
flotation reagents employed.   Similarly, in  the  guideline-
development  effort, analyses were not performed for each of
the specific reagents used at the  various  flotation  mills
visited.    The presence of flotation reagents in appreciable
quantities may be detected in elevated values for  COD,  oil
and  grease,  or  surfactants,  as  analytical  data on mill
effluents indicate.  The limitation of reagents individually
appears unfeasible, since the exact suite  of  reagents  and
dosages  is  nearly  unique  to  each  operation  and highly
variable over time.

Current practice in the ferroalloy milling industry includes
flotation of sulfide ores of molybdenum,  and  flotation  of
scheelite  (tungsten  ore).   The ores floated are generally
somewhat complex, containing pyrite  and  minor  amounts  of
lead  and  copper  sulfides.    Reagents  used in the sulfide
flotation  circuits  and  reflected  in  effluents   include
xanthates, light oils, and cyanide (as a depressant).  Since
the  flotation  is  performed  at basic pH, solution of most
metals is at a low level.  Molybdenum  is  an  exception  in
that  it is soluble as the molybdate anion in basic solution
and appears in  significant  quantities  in  effluents  from
several operations.  Tungsten ore flotation involves the use
of  a  quite  different set of reagents—notably, oleic acid
                            290

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and tall oil soaps—and may be performed at acid pH.  At one
major plant, both sulfide flotation for molybdenum  recovery
and  scheelite  flotation  are  practiced,  resulting in the
appearance of both sets of reagents in the effluent.   Visit
sites  included  plants  recovering  both  molybdenum (6101,
6102, and 6103) and tungsten  (6104 and 6105)   by  flotation.
Although  flotation  would  almost certainly be used in such
cases, no currently active processors  of  sulfide  ores  of
nickel or cobalt could be identified in the U.S.

Ore  Leaching.    In  many  ways,  ore  leaching  operations
maximize the pollution  potential  from  ore  beneficiation.
Reagents are used in large quantities and are frequently not
recovered.  Extremes of pH are created in the process stream
and  generally  appear in the mill effluent.   Techniques for
dissolving the material to be recovered  are  generally  not
specific,  and other dissolved materials are rejected to the
waste stream to preserve product purity.   The  solution  of
significant  fractions  of  feed  ore,  and the use of large
quantities of reagents, results  in  extremely  high  total-
dissolved- solids  concentrations.  Because of reagent costs,
and  the  benefits  of  increased   concentration   in   the
precipitation  or  extraction  of  values from solution, the
amount of water used per ton of ore processed by leaching is
generally lower  than  that  for  physical  benefication  or
flotation.    One   ton  of  water  per  ton  of  ore  is  a
representative value.

Effluents for several mills in the ferroalloy industry which
employ leaching were characterized  in  this   study.   Visit
sites  included a vanadium mill (mill 6107)(properly classed
in  SIC  1094,  but  treated  here  because   of   lack   of
radioactives,  end  use  of  product,  and  applicability of
general process to other ferroalloy  ores)   which  practices
leaching as the primary technique for recovering values from
ores, as well as two tungsten mills which employ leaching in
the   process,  though  not  as  the  primary  beneficiation
procedure.  One operation (mill 6105)  leaches a small amount
of concentrate to reduce lime and  phosphorus  content,   and
the   other   (mill   6104)    leaches   scheelite  flotation
concentrates as part of a chemical refining procedure.   Data
for samples from leaching plants in the uranium  and  copper
industries may also be examined for comparison.
                            291

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Process   Description  and  Raw-Waste  Characterization  For
Specific Mines and Mills Visited

Mine/Mill 6101

At mine/mill  6101,  molybdenum  ore  of  approximately  0.2
percent   grade   is   mined  by  open-pit  methods  and  is
concentrated by flotation to yield a 90 percent  molybdenite
concentrate.   The  mine and mill are located in mountainous
terrain, along a river gorge.  The mill is adjacent  to  and
below  the  mine,  the  elevation of which ranges from 2,550
meters (8,400 ft) to 3,000  meters   (10,000  ft)  above  MSL
(mean  sea  level).   The  local climate is dry, with annual
precipitation  amounting  to  28  cm  (11  in.)   and  annual
evaporation of 107 cm (42 in.).

Approximately  22,000  cubic  meters  (6 million gallons) of
water per day are used  in  processing  14,500  metric  tons
(16,000  short  tons)  of ore.  Reclamation of 10 percent of
the water at the mill site, evaporation,  and  retention  in
tails  reduce  the  daily discharge of water to 16,000 cubic
meters (4.3 million gallons).  Process water is  drawn  from
wells  on  the  property and from the nearby river.  No mine
water is produced.

Ore processing consists of crushing, grinding, and  multiple
stages of froth flotation, followed by dewatering and drying
of concentrates.  The complete process is illustrated in the
simplified   flowsheet   of   Figure  V-25.   There  are  no
recoverable  byproducts  in  the  ore.    Reagent   use   is
summarized in Table V-39.

Recovery of molybdenite averages 78 to 80 percent but varies
somewhat,  depending on the ore fed to the mill.  Recoveries
on ore which has been stockpiled  are  somewhat  lower  than
those  achieved  on  fresh ore.  This is, apparently, due to
partial oxidation of the molybdenite to (soluble)  molybdenum
oxide  and  ferrimolybdite,  which  are  not   amenable   to
flotation.   Processing  of  these  oxidized  ores  is  also
accompanied by  an  increase  in  the  dissolved  molybdenum
content  of  the  plant  discharge.   The  final concentrate
produced averages 90 percent KoS2.
                        f
As the flowsheet shows, only one waste stream  is  produced.
Data  for  this  stream, as sampled at the mill prior to any
treatment, are summarized in Table V-40.

High  COD  levels  apparently  result  from  the   flotation
reagents used and provide some indication of their presence.
                            292

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                            Figure V-25.  MILL 6601 FLOWSHEET
                                                         CONCENTRATE-
                                                     CONCENTRATE


                                                      — MIDDLINGS
    CLEANER
     FLOAT

  (6 STAGES Wl TH
  REGRIND AND
INTERNAL RECYCLE)
                                                                                       -TAILS
                                                                                         ' *' LS
                                                                        CONCENTRATE
                                                    UNDERFLOW
     •^•UNDERFLOW
                  OVERFLOW

     22.000 m3/day
     (6,000.000 gpd)
THICKENER

ffinn Ann **A\
OVERFLOW
RECLAIM
WATER
TO TAILING
  POND
                                         293

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TABLE V-39. REAGENT USE IN MOLYBDENUM MILL 6101
REAGENT
Lime
Vapor Oil
Pine Oil
Hypo (Sodium Thiosulfate)
(Na2 8203 • 5H20)
Phosphorus Pentasulfide (?2 85)
MIBC (methyl-isobutyl carbinol)
Sodium Cyanide (Na CN)
DOSAGE
kg/metric ton ore
0.075
0.09
0.015
0.035
0.005
0.02
0.015
Ib/short
ton ore
0.15
0.18
0.03
0.07
0.01
0.04
0.03
 TABLE V-40. RAW WASTE CHARACTERIZATION AND
           RAW WASTE LOAD FOR MILL 6601
PARAMETER
TSS
TOS
OH and Grease
COD
Ai
Cd
Cu
Mn
Mo
Pb
Zn
Fa
Total Cyanide
Fluoride
CONCENTRATION
Img/S, 1
IN WASTEWATER
500.000
2.598
2,0
135
0.01
0.74
51
56.5
5.3
9.8
76.9
1.306
0.02
6.2
TOTAL WASTE
kg/day
14.000,000
42/100
32
2.200
0.16
12
820
900
85
160
1,200
21,000
0.32
99
Ib/day
32,000.000
92,000
70
4,800
0.3S
26
MOO
2,000
190
3SO
2.600
46,000
0.70
220
RAW WASTE LOAD
per unit ore milted
kg/ metric ton
995
3.0
0.0023
0.16
0.000012
0.00086
0.059
0.064
0.0061
0.011
0.086
1.5
0.000023
0.0071
lb/«hort ton
1.990
6.0
0.0046
0.32
0.000023
0.0017
0.11
0.13
0.012
0.023
0.17
3.0
0.000046
0.014
per unit concentrate produced
kg/ metric ton
610.000
1.830
1.4
96
0.0070
0.52
36
39
3.7
7.0
52.4
915
0.014
4.3
Ib/ihon ton
1,200,000
3.670
2.8
190
0.014
1.0
72
79
7.4
14.0
105
1,830
0.028
B.7
                    294

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The  low  cyanide level found reflects significant decreases
in cyanide dosage over earlier operating modes and indicates
almost  complete  consumption  of  applied  cyanide.   Metal
analyses  were performed in acidified samples containing the
solid tailings.  High values may be  largely  attributed  to
metals  which  were  solubilized  from the unacidified waste
stream.

Mine/Mill 6102

Mill 6102 also recovers molybdenite by flotation,  but  mill
processing  is complicated by the additional recovery of by-
product concentrates.  Water use in processing approximately
39,000 metric tons   (43,000  short  tons)  of  ore  per  day
amounts to 90,000 cubic meters (25 trillion gallons) per day.
Nearly  complete  recycle  of  process  water results in the
daily use of only 1,700 cubic meters  (450,000  gallons)  of
makeup  water.  Discharge from the mill tailing basin occurs
only during spring snow-melt runoff,  when  it  averages  as
much as 140,000 cubic meters (38.5 million gallons) per day.

Mining  is  both  underground and open-pit, with underground
operations which began approximately 67 years ago,  and  the
first  open-pit production in 1973.  Recovery of molybdenite
is by flotation in five stages, yielding a final molybdenite
concentrate containing more than 93 percent MoS2.   Tungsten
and  tin  concentrates  are produced by gravity and magnetic
separation, with additional flotation steps used  to  remove
pyrite  and  monazite.  Recovered pyrite is sold as possible
(currently,  about   20  percent  of  production),  with  the
balance delivered to tails.  The monazite float product also
reports  to  the tailing pond, since recovery of monazite is
not profitable for this operation at this time.

The mill operation is located on the continental  divide  at
over  3,353  meters   (11,000  feet)  above  MSL.    The local
terrain is mountainous.  Climate and topography have a major
impact on water-management and  tailing-disposal  practices,
with  a  heavy  snow-melt  runoff  and the presence of major
drainages above tailing-pond areas posing problems.

Mill  Description.    Figure   V-26   presents   a   greatly
simplified  diagram  of  the  flow  cf ore through the mill.
Following crushing and  grinding,  roughing  and  scavenging
flotation  are  used  to  extract  molybdenite from the ore.
Nearly 97 percent of the incoming material—currently, about
39,000 metric tons (43,000 short tons) per  day—is  thereby
rejected  and sent directly to the byproduct recovery plant.
The flotation concentrate, averaging about 10 percent  MoS2[,
                            295

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Figure V-26. SIMPLIFIED MILL FLOW DIAGRAM FOR MILL 6102
     TO
TAILINGS
         •LIGHTS-
                   LIGHTS-
                            CRUSHING
                            (3 STAGES)
                           28% + 3 MESH

                           GRINDING IN
                           BALL MILLS
                                1
                          36% + 100 MESH
                               i
                            FLOTATION
                              i
                            FLOTATION
                         96% OF MILL FEED
                                 GRAVITY SEPARATION
                                 (HUMPHREY'S SPIRALS)
                             PYRITE
                            FLOTATION
                                1
                              TAILS

                             TABLES
                  MONAZITE
                CONCENTRATE
                           MONAZITE
                           FLOTATION
                            MAGNETIC
                           SEPARATION
                  NONMAGNETIC
                      TIN
                  CONCENTRATE
                                    I
                                                         CONCENTRATE
                                                           CLEANER
                                                          FLOTATION
                                                          (4 STAGES)
                                                  DRYING
                                                    I
                                               CONCENTRATE
                                                (93% + MoS2>
                                                                  TO
                                                                  TAILINGS
                                                   /v\ INDICATES
                                                       SAMPLING POINT
                                              MAGNETIC
                                              TUNGSTEN
                                            CONCENTRATE
                             296

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is  fed  to four stages of further flotation.  Reagents used
in the primary flotation step are summarized in Table  V-U1.
Most  are  added  as  the  ore  is fed to the ball mills for
grinding.

Cleaner flotation in four stages and three regrinds yield  a
final   product  averaging  greater  than  93  percent  MoS2
content.  Reagent use in the cleaner grinding and  flotation
circuit is summarized in Table V-42.

Tailings  from  the  rougher flotation are pumped to the by-
products plant, where heavy fractions  are  concentrated  in
Humphreys  spirals.   Pyrite is removed from the concentrate
by flotation at pH 1.5, and the flotation tailings are  then
tabled  to  further concentrate the heavy fractions.  The pH
of the table concentrate is then adjusted  to  1.5  and  its
temperature  raised  to  70  degrees  Celsius  (158  degrees
Fahrenheit), and monazite  is  removed  by  flotation.   The
tailings  from this flotation step are dewatered, dried, and
fed  to  magnetic  separators,  which  yield  separate   tin
(cassiterite)    and   tungsten   (wolframite)  concentrates.
Reagent use in the  flotation  of  pyrite  and  monazite  is
summarized in Table V-43.

Effluent samples were taken at three points in mill 6102 due
to the complexity of the process.  A combined tailing sample
was  taken  representative of the total plant effluent, and,
in addition, effluents were sampled from two points  in  the
process  (marked  19  and 20 on the flowsheet. Figure V-26).
Although flows at these points are very  small  compared  to
the  total  process  flow,  they  were  considered important
because of the acid conditions prevailing in monazite flota-
tion.   Concentrations  and  total  loadings  in  the   mill
effluent,  and  concentrations  in the effluents from pyrite
flotation and monazite flotation, are presented in Tables V-
HH and V-U5.

Considerably heavier  use  of  cyanide  than  at  mill  6101
(almost ten times the dosage per ton of ore)  is reflected in
significantly  higher  levels  in  the untreated mill waste.
Total metal contents are again elevated  by  leaching  solid
particles  in  the tailing stream.   The increase in solution
of most heavy metals as increasingly acid conditions prevail
in processing is evident in the data from the  monazite  and
pyrite flotation effluents.

Mine  water is produced in the underground mine at mill 6102
at an average rate of 4,000 metric tons per day  (700  gpm).
Its  characteristics  are  summarized,  along  with those of
                             297

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   TABLE V-41. REAGENT USE FOR ROUGHER AND SCAVENGER
             FLOTATION AT MILL 6102
REAGENT
Pine oil
Vapor oil
Syntax
Lime (Calcium oxide)
Sodium silicate
Nokes reagent
PURPOSE
Frother
Collector
Surfactant and Frother
Adjustment of pH to 8.0
Slime Dispersant
Lead Depressant
CONSUMPTION
kg/metric top
ore milled
0.18
0.34
0.017
0.15
0.25
0.015
Ib/short ton
ore milled
0.35
0.67
0.034
0.30
0.50
0.03
TABLE V-42. REAGENT USE FOR CLEANER FLOTATION AT MILL 6102
REAGENT
Vapor oil
Sodium cyanide
Nokes reagent
Oowfroth 250
Valco 1801
PURPOSE
Collector
Pyrite and Chal co-
py rite Depressant
Lead Depressant
Frother
Flocculant
CONSUMPTION
kg/metric ton
ore milled
0.45
0.13
0.45
0.015
0.003
Ib/short ton
ore milled
0.90
0.25
0.90
0.03
0.006
                      298

-------
  TABLE V-43. REAGENT USE AT BYPRODUCT PLANT OF MILL 6102
            (Based on total byproduct plant feed)
REAGENT
PURPOSE
CONSUMPTION
kg/metric ton
ore milled
Ib/short ton
ore milled
PYRITE FLOTATION
Sulfuric acid
Z-3 Xanthate
Dowfroth 250
pH Regulation
Collector
Frother
0.018
0.0005
0.0005
0.036
0.001
0.001
MONAZITE FLOTATION
ARMAC C
Starch
Sulfuric acid
Collector
WO2 Depressant
pH Regulation
0.0005
0.0005
0.0005
0.001
0.001
0.001
TABLE V-44. MILL 6102 EFFLUENT CHEMICAL CHARACTERISTICS
          (COMBINED-TAILINGS SAMPLE)
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION

-------
   TABLE V-45. CHEMICAL CHARACTERISTICS OF ACID-FLOTATION STEP
PARAMETER
PH
Cd
Cu
Fe
Mn
Mo
Pb
CONCENTRATION (mg/ 1 } AT INDICATED POINTS OF FIGURE V-26
PYRITE FLOAT (19)
4.5»
0.01
0.2
4.2
4.0
3.0
0.3
MONAZITE FLOAT (20)
1.5
0.042
0.5
490
53.3
4.0
1.34
•Value in pH unit*
                            300

-------
other mine waters, in Table V-38.  At mill  6102,  all  mine
water  is  added  to  the  mill tailing pond and then to the
process circuit.

Mine 6103

Mine 6103 is an underground molybdenum mine which  is  under
development.   Ore from the mine will be processed in a mill
at a site approximately 16 kilometers (10  miles)   from  the
mine  portal.   The mill operation will produce no effluent,
all of the process water being recycled.   Mine  water  flow
presently  averages  9,800 cubic meters per day (1,700 gpm).
Its quality prior to treatment has been summarized in  Table
V-38.

Mine/Mill 6104

This  complex  operation combines mining, beneficiation, and
chemical processing to produce a pure ammonium paratungstate
product as well as molybdenum and  copper  concentrates.   A
total  of 10,000 cubic meters  (2.9 million gallons) of water
are used each day in processing  2,200  metric  tons  (2,425
short  tons) of ore.  The bulk of this water is derived from
the 47,000 cubic meters (13 million gallons) of water pumped
from the mine each day.

The mill process is illustrated in Figures  V-27  and  V-28,
which  also  show  water  flow  rates.   After  crushing and
grinding, sulfides of copper and molybdenum are floated from
the ore, employing xanthate collectors and soda ash  for  pH
modification.   This  flotation  product  is  separated into
copper and molybdenum concentrates in a subsequent flotation
using sodium bisulfide to depress the copper.  Tailings from
the sulfide flotation are refloated using tall oil  soap  to
recover a scheelite concentrate, which is reground and mixed
with purchased concentrates from other sites.  The scheelite
is  digested  and  filtered, and the solution is treated for
molybdenum  removal.   Following  solvent   extraction   and
concentration, ammonium paratungstate is crystallized out of
solution and dried.

Effluent  streams  from  parts of the operation specifically
concerned with  beneficiation  were  sampled  and  analyzed,
along  with the combined discharge to tails for the complete
mill.  Mine water was also sampled, and analyses  have  been
reported  in Table V-38.  Data for a composite effluent from
beneficiation operations,  several  individual  beneficiation
effluents, and the combined plant discharge are presented in
Tables V-46, V-47, V-48, V-49, and V-50.
                             301

-------
Figure V-27. INTERNAL WATER FLOW FOR MILL 6104 THROUGH
         MOLYBDENUM SEPARATION

WET
ORE



WATER
FROM
CREEK





WA
FR
Ml

2,900i
(782.0C
(32.000 gpd) CRUSHING
GRINDING
n3/diy
Mgpdl
3.0B1 m3/d»y
(814.000 gpd)

468m3/dty 11 21 .000 gpdl

r~
16-METER
(60-FT)
THICKENER
l
rcfl
NE 1»m3/d.y

	 ^ SCHEELITE
THICKENER

STEAM
481 m"/d»y
(127.000 gpd)


SLOWDOWN




496 m3/diy
(131.000 gpd)
1 ' I
„ BULKSULFIOE l920'001
FLOTATION
126.000 gpd)


\
COPPEF
MOLYBDE
SEPARAT

NUM
inu ^

MoS2 PRODUCT
I 0.038 m3/d,y
I 110 gpd)
TO STOCKPILE

102 m3/d«y
127,000 gpdl

468 mj/d»v
1121,000 gpd)
TO _^
ATMOSPHERE ^^
DRYING
AND
ROASTING
23 m3/d»y
(8,000 gpd)
Mo03 PRODU
TO
STOCKPIL
646 m3/d.v
1144,000 gpd)
\ <


t



*1
L



2.6 m3/d.y
(700 gpd) I
SOj
SCRUBBER

i
CT
E

6,076 n
(1.606,0
TO
TAILING POND
1,336 m3AJiy
(363.000 gpd)
~/diy 4 .360 m-/diy
"*•" lOllMin (I.IM.000^1
FLOTATION
UNDE

IF LOW
492 m3/diy 1130,000 gpd) __


CONCEI
THICK

PER (26,000 gpd)
ENER
Cu CONCENTRATE PRODUCT
TO STOCKPILE
114 m3/diy
(30.000 gpdl
(
299 m3/diy
(79,000 gpdl
1,106 m'/diy
FILTERING 1292,000 gpd)
WAS


HINO
299 m3/d
-------
       Figure V-28. INTERNAL WATER FLOW FOR MILL 6104
                  FOLLOWING MOLYBDENUM SEPARATION
                                                TO ATMOSPHERE
FROM MOLYBDENUM
  SEPARATION
  (FIGURE V 271
                                                                48 mj/d»v
                                                                112.000 gpdl
                                                                ATMOSPHERE
                                 303

-------
    TABLE V-46. COMPOSITE WASTE CHARACTERISTICS FOR BENEFICIATION
              AT MILL 6104 (SAMPLES 6, 8, 9, AND 11)

PARAMETER

pH
COD
Oil and Grease
As
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
Cyanide
rnMTPMTRATinw
(mg/i > IN
WASTE WATER

10*
238
11.4
<0.07
0.04
4.9
22.5
19.0
0.22
6.3
4.8
0.2

TOTAL WASTE
kg/day

1,100
55
C0.34
0.19
24
110
91
1.1
30
23
0.96
Ib/day

2,400
120
<0.75
0.42
53
240
200
2.4
66
51
2.1
RAW WASTE LOAD
per unit ore processed
kg/ metric ton

0.50
0.025
< 0.0002
0.000086
0.011
0.050
0.041
0.00050
0.014
0.010
0.00044
Ib/short ton

1.0
0.050
< 0.0003
0.00017
0.022
0.10
0.083
0.0010
0.027
0.021
0.00088
per unit total
concentrate produced
kg/metric ton
.
8.1
0.41
<0.003
0.0014
0.18
0.81
0.67
0.0081
0.23
0.16
0.0072
Ib/short ton
.
16
0.81
<0.007
0.0028
0.36
1.6
1.3
0.01G
0.46
0.32
0.014
"Value in pH units
TABLE V-47. WASTE CHARACTERISTICS FROM COPPER-THICKENER OVERFLOW
          FOR MILL 6104 (SAMPLE 5)
PARAMETER
PH
Cd
Cu
Mn
Mo
Pb
Fe
CONCENTRATION
(mq/H) IN
WASTEWATER
11*
0.26
<0.02
1.0
1.2
0.07
26.0
TOTAL WASTE
kg/day
-
0.024
< 0.002
0.091
0.11
0.0064
2.4
Ib/day
-
0.053
<0.004
0.20
0.24
0.014
5.3
RAW WASTE LOAD
per unit ore milled
kg/metric ton
-
0.000011
<0.0000009
0.000041
0.000050
0.0000029
0.001 1
Ib/short ton
-
0.000022
< 0.000002
0.000082
0.00010
0.0000058
0.0022
    "Value in pH units
                              304

-------
TABLE V-48. SCHEELITE-FLOTATION TAILING WASTE CHARACTERISTICS
          AND LOADING FOR MILL 6104 (SAMPLE 7)
PARAMETER
PH
Cd
Cu
Mn
Mo
Pb
Zn
Fe
CONCENTRATION
(mg/£) IN
WASTEWATER
10*
0.32
1.42
41
1.3
0.22
11.2
0.43
TOTAL WASTE
kg/day
—
1.3
5.9
170
5.5
.92
47
1.8
Ib/day
—
2.9
13
370
12
2.0
100
4.0
RAW WASTE LOAD
per unit ore milled
kg/metric ton
-
0.00059
0.0027
0.077
0.0025
0.00042
0.021
0.00082
Ib/short ton
-
0.0012
0.0054
0.15
0.0050
0.00084
0.043
0.0016
 'Value in pH units
TABLE V-49. 50-FOOT-THICKENER OVERFLOW FOR MILL 6104 (SAMPLE 10)
PARAMETER
PH
Cd
Cu
Mn
Mo
Pb
Zn
Fe
CONCENTRATION
(mg/£) IN
WASTEWATER
9»
<0.01
0.31
1.3
21.0
0.04
0.16
7.7
TOTAL WASTE
kg/day
—
< 0.005
0.15
0.61
9.9
0.019
0.075
3.6
Ib/day
—
<0.01
0.33
1.3
22
0.042
0.17
7.9
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
< 0.000002
0.000068
0.00028
0.0045
0.0000086
0.000034
0.0016
Ib/short ton
—
< 0.000005
0.00014
0.00055
0.0090
0.000017
0.000068
0.0033
 •Value in pH units
                           305

-------
TABLE V-50. WASTE CHARACTERISTICS OF COMBINED-TAILING
          DISCHARGE FOR MILL 6104 (SAMPLES 15, 16, AND 17)
PARAMETER
IDS
Oil and Grease
COD
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Total Cyanide
CONCENTRATION
(mg/£) IN
WASTEWATER
2290
14.7
174
< 0.07
0.03
0.03
0.52
50
2.2
<0.02
<0.5
< 0.01
TOTAL WASTE
kg/day
22,900
147
1,740
<0.7
0.30
0.30
5.2
500
22
<0.2
<5.0
<0.1
Ib/day
50.000
320
3,800
<1.5
0.66
0.66
11
1,100
480
< 0.4
< 11
< 0.2
RAW WASTE LOAD
per unit ore processed
kg/metric ton
10.4
0.067
0.79
<0.0003
0.00014
0.00014
0.0024
0.23
0.010
< 0.00009
< 0.002
< 0.00005
Ib/short ton
21
0.13
1.6
< 0.0006
0.00027
0.00027
0.0047
0.45
0.020
< 0.0002
< 0.005
< 0.00009
per unit
concentrate produced
kg/metric ton
170
1.1
13
< 0.005
0.0023
0.0023
0.039
3.7
0.16
< 0.0015
<0.03
< 0.0008
Ib/short ton
340
2.2
26
<0.01
0.0046
0.0046
0.078
7.4
0.32
< 0.003
<0.07
< 0.002
                  306

-------
The  combined-tails  discharge characteristics are not truly
representative of raw waste from the leaching  and  chemical
processing parts of the operation, since advanced treatments
(including  distillation and air stripping) are performed on
parts of the waste  stream  prior  to  discharge  to  tails.
Total  dissolved  solids and ammonia (not determined for the
sample taken), in particular, are greatly reduced  by  these
treatments.

Mine/Mill 6105

Mill  6105,  a considerably smaller operation than mine/mill
6104,  also  recovers  scheelite.   As  shown  in  the  mill
flowsheet   of  Figure  Ill-IB,  a  combination  of  sulfide
flotation, scheelite flotation, wet gravity separation,  and
leaching  is  employed  to  produce  a  65  percent tungsten
concentrate from 0.7 percent  mill  feed.   A  total  of  52
metric  tons   (57  short tons) per day of water drawn from a
well on site are used in processing 46 metric tons  (51 short
tons)  of  ore.   Mill  tailings  are  combined   prior   to
discharge,  providing  neutralization of acid-leach residues
by the high lime content of the ore.  Analytical data for  a
sample  of the combined mill effluent are presented in Table
V-51.

The mine at this site intercepts an aquifer  producing  mine
water, which must be intermittently pumped out (for approxi-
mately  hour every 12 hours).  Total effluent volume is less
than 4 cubic meters (1,000 gallons)  per  day.   Samples  of
this effluent were not obtained because of inactivity during
the  site  visit.  It is expected to be essentially the same
as the mill water-source well, which drains the same aquifer
and which was sampled.

Mine/Mill 6106

Ferronickel is produced at this site by direct smelting of a
silicate ore (garnierite)  from an open-pit mine.  Water  use
is  limited  and is primarily involved in smelting, where it
is used for cooling and for slag granulation.  Beneficiation
of  the  ore  involves  drying,  screening,  roasting,   and
calcining but requires water for belt washing and for use in
wet  scrubbers.   Flow  from  all  uses  combined amounts to
approximately 28 cubic meters (7,700 gallons) per day.  This
combined waste stream was sampled, and its analysis is shown
in Table V-52.

Mine water during wet-weather runoff through a creek bed  to
an  impoundment  used  for  mill  water treatment results in
                             307

-------
  TABLE V-51. WASTE CHARACTERISTICS AND RAW WASTE LOAD AT MILL 6105
           (SAMPLE 19)
PARAMETER
TDS
Oil and Grease
COD
NH3
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
(mg/fc) IN
WASTEWATER
1232
1
39.7
1.4
<0.07
<0.01
0.02
0.52
0.19
0.5
0.02
<0.5
<0.02
0.44
6.9
<0.01
TOTAL WASTE
kg/day
64
0.052
2.1
0.073
< 0.004
< 0.0005
0.0010
0.027
0.0099
0.026
0.0010
<0.03
< 0.001
0.023
0.36
< 0.0005
Ib/day
140
0.11
4.6
0.16
<0.01
< 0.001
0.0022
0.059
0.022
0.057
0.0022
<0.07
< 0.002
0.051
0.79
< 0.001
RAW WASTE LOAD
per unit ore processed
kg/metric ton
1.4
0.0011
0.046
0.0015
<0.0001
<0.00001
0.000022
0.00058
0.00022
0.00057
0.000022
<0.0007
<0.00002
0.00050
0.0078
< 0.00001
Ib/short ton
2.8
0.0022
0.092
0.0030
<0.0002
<0.00002
0.000045
0.0012
0.00043
0.0011
0.000045
<0.001
<0.00004
0.0010
0.016
<0.00002
per unit total
concentrate produced
kg/metric ton
130
0.10
4.2
0.14
< 0.009
< 0.0009
0.002
0.053
0.020
0.052
0.0020
<0.06
< 0.002
0.045
0.71
< 0.0009
Ib/short ton
250
0.20
8.4
0.28
<0.02
< 0.002
0.010
0.11
0.040
0.10
0.010
<0.13
< 0.004
0.091
1.4
< 0.002
  TABLE V-52. CHEMICAL COMPOSITION OF WASTEWATER, TOTAL WASTE, AND
            RAW WASTE LOADING FROM MILLING AND SMELTER EFFLUENT
            FOR MILL 6106
PARAMETER
pH
TSS
TDS
Oil and oraaM
As
Cd
Cu
Mn
Mo
Pb
Zn
Ft
Ni
CONCENTRATION
imt/SL)
IN WASTE WATER
8.6*
226.9
212
3.4
<0.07
< 0.005
< 0.03
0.53
0.5
<0.1
0.06
24
0.4
TOTAL WASTE
kg/day
—
3,600
3,300
54
<1
<0.08
<0.5
8.3
7.9
<2
0.79
380
6.3
Ib/diy
-
7.900
7,300
120
< 2
<0.2
<1
18
17
<4
1.7
840
13.9
RAW WASTE LOAD
par unit or* milled
ko/1000
rnatric tons
—
790
730
12
<0.2
<0.02
<0.1
1.8
1.7
<0.4
0.17
84
1.4
lb/1000 short tons
—
1,600
1,500
24
<0.4
< 0.04
< 0.2
3.7
3.5
<0.9
0.35
170
2.8
par unit conc*ntrata produced
kg/1000
matric tons
—
43,000
39,000
640
<10
< 1
< 6
99
94
<20
9.4
4,500
75
lb/1000 short tons
—
86,000
79,000
1,300
<20
< 2
<10
200
190
<50
19
9,000
150
*Valua in pH units.
                        308

-------
discharges as large as  21,000 cubic meters  (576,000 gallons)
per day  from the  impoundment.  Since the mine was dry during
the site visit, no  samples  of  this   flow  were  obtained.
Company-furnished data  for  the impoundment water quality,
however, reflect  the  impact of mine-site runoff.

Mine/Mill 6107

At this operation, vanadium  pentoxide,  V205_,  is  produced
from  an  open-pit  mine  by  a  complex  hydrometallurgical
process involving roasting,  leaching,  solvent  extraction,
and  precipitation.   The process  is  illustrated in Figure
111-21 and also in Figure V-29   (which  shows  system  water
flows).   In  the mill,  a total of 7,600 cubic meters  (1.9
million gallons)  of   water  are  used   in  processing  1,140
metric  tons   (1,250  short tons) of ore, including scrubber
and cooling wastes and  domestic use.

Ore  from  the  mine  is  ground,  mixed  with   salt,   and
palletized.  Following  roasting at 850  degrees Celsius  (1562
degrees  Fahrenheit)  to  convert  the  vanadium  values  to
soluble  sodium   vanadate,  the  ore  is  leached  and   the
solutions  acidified  to  a pH of 2.5 to 3.5.  The resulting
sodium decavanadate (NajsVlOOJMSJ) is concentrated  by  solvent
extraction,  and  ammonia is  added to precipitate ammonium
vanadate, which is  dried and  calcined  to  yield  a  V2.0I5
product.                                                   ~"

The  most significant effluent streams  are from leaching and
solvent extraction, from  wet scrubbers  on roasters, and from
ore dryers.  Together,  these sources account for  nearly  70
percent  of  the  effluent stream, and essentially all of its
pollutant content.  Analyses for  these waste  streams  are
summarized  in  Tables  V-53, V-54, and  V-55.  Effluents from
the solvent-extraction  and leaching processes are  currently
segregated from the roaster/scrubber effluent, although they
are  both discharged  at the same point, to avoid the genera-
tion of voluminous calcium  sulfate  precipitates  from  the
extremely  high   sulfate  level in the SX stream and the high
calcium level in the  scrubber bleed.  Both  of  these  waste
streams  exhibit  extremely  high dissolved-solid concentra-
tions (over 20,000 mg/1)  and are diluted approximately  10:1
immediately prior to  discharge.

Mercury Ores

Water  flow  and  the  sources,  nature, and quantity of the
wastes dissolved  in  the  water  during  the  processes  of
                            309

-------
                                               Figure V-29. WATER USE AND WASTE  SOURCES FOR VANADIUM MILL 6107
u>
M
O
                                                                                   3300 «/mifl
                                                                                    IKOgpm)
                                                                                                  4.920 /mil.
                                                                                         6910 t fmin II .300 Bwnl
                                                                                          fiaoomi)
                                                                               TO
                                                                           ATMOSPHERE
                                                                      180 £ I
                                                                       ISOgpm)
     WASHING.
 WATER LEACH. AND
SOLVENT EXTRACTION
                                                                                          NEUTRALIZATION
                                                                                                              OVERFLOW IRAIN1
                                                                                                                                           1.135yirun
                                                                                                                                           (300 gpm)
                                                                                                                                   CLEAN-
                                                                                                                                   WATER
                                                                                                                                   DITCH
                                                                                       IT.355.OOO  (3.000.000-9J1)
                                                                                            SCRUBBER
                                                                                            BLEED POND
                                                             3.T8S.OOO-- (1.000.000-5JII
                             tt9.2SO.000- • (SO.OOO.OC3-9I)
                                                                                       113.600.000. ISO.OOO.OOO-gal)
                                                                                              EAST
                                                                                          EFFLUENT POND
                                 NOTE:
                           RUNOFF FROM RAIN
                           IS NOT CONSIDERED
                           EXCEPT WHERE IT
                           ENTERS THE PROCESS
                           (x«J  = SAMPLE NUMBER
                            SAMPLES (TI)AND (72) ARE MINE-MATED SAMPLES

-------
TABLE V-53. WASTE CHARACTERIZATION AND RAW WASTE LOAD FOR MILL 6107
          LEACH AND SOLVENT-EXTRACTION EFFLUENT (SAMPLE 80)
PARAMETER
pH
TDS
Oil and greata
COD
NH3
Ai
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
ft
Ca
Chloride
Fluoride
SuMate
CONCENTRATION
(mg/dl
IN WASTE WATER
3.5*
39.350
94
475
0.16
0.35
0.037
1.15
0.16
54
< 0.1
< 0.05
31
0.52
0.26
206
7.900
4.6
26.000
TOTAL WASTE
kg/day
-
83.000
200
1,000
0.34
0.74
0.078
2.4
0.32
110
<0.2
< 0.1
65
1.1
0.55
430
17.000
9.7
55.000
Ib/diy
-
180.000
440
2.200
0.75
1.6
0.17
5.3
0.7
240
< 0.4
< 0.2
T40
2.4
1.2
950
37.000
21
120,000
RAW WASTE LOAD
per unit ore milled
kg/metric tonH
-
73
0.18
0.88
0.0003
0.00065
0.000068
0.0021
0.00028
0.096
< 0.0002
< 0.0001
0.057
0.00096
0.0005
0.38
15
0.0085
48
Ib/fhort ton
-
146
0.35
1.76
0.0006
0.0013
0.00014
0.0042
0.00056
0.19
< 0.0004
< 0.0002
0.11
0.0019
0.001
0.75
30
0.017
96
          •Value in pH units
                             311

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  TABLE V-54. WASTE CHARACTERISTICS AND WASTE LOAD FOR  DRYER
             SCRUBBER BLEED  AT MILL 6107 (SAMPLE 81)
PARAMETER
PH
TSS
TDS
Oil and Grease
COD
Ammonia
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/£)
IN WASTEWATER
7.8*
—
7,624
15
58.4
2
<0.07
< 0.005
0.25
0.06
4
<0.1
<0.05
29
0.33
27
118
4,220
1.35
255
TOTAL WASTE
kg/day
—
—
4.000
7.8
30.4
1.0
< 0.035
< 0.0025
0.13
0.03
2.1
<0.05
< 0.025
15
0.17
14
61
2,200
0.70
133
Ib/day
—
—
8,800
17
67
2.2
<0.07
< 0.005
0.29
0.07
4.6
<0.1
<0.05
33
0.37
31
130
4,800
1.5
290
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
—
3.5
0.007
0.027
0.0009
< 0.00003
< 0.000002
0.00011
0.00003
0.0018
<0.00004
< 0.00002
0.013
0.00015
0.012
0.054
1.9
0.0006
0.12
Ib/short ton
—
—
7.0
0.014
0.054
0.0018
< 0.00006
< 0.000004
0.00023
0.00006
0.0037
< 0.00009
< 0.00004
0.026
0.00030
0.025
0.11
3.9
0.0012
0.23
•Value in pH units
                           312

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  TABLE V-55. WASTE CHARACTERISTICS AND LOADING FOR SALT-ROAST
             SCRUBBER BLEED AT MILL 6107 (SAMPLE 77)
PARAMETER
PH
TSS
TDS
Oil and Grease
COD
Ammonia
A*
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Chloride
Fluoride
Sulfate
CONCENTRATION
(mglH)
IN WASTEWATER
2.3*
2,000
80,768
5
1,844
0.04
0.08
< 0.005
0.9
<0.03
5.5
—
< 0.05
—
< 0.003
59,500
7.5
780
TOTAL WASTE
kg/day
—
1,900
76,000
4.7
1,700
0.039
0.075
< 0.005
0.86
<0.03
5.2
_
<0.05
—
< 0.003
51,000
7.0
740
Ib/day
—
4,100
160,000
10
3,800
0.086
0.15
<0.01
1.9
<0.07
12
—
<0.1
—
< 0.007
110,000
16
1,600
RAW WASTE LOAD
per unit ore milled
kg/metric ton
—
1.6
67
0.0041
1.5
0.000031
0.000063
< 0.000004
0.00075
< 0.00003
0.0045
—
< 0.00004
—
< 0.000003
45
0.0062
0.64
Ib/short ton
-
3.3
130
0.0085
3.1
0.000063
0.00013
< 0.000008
0.0015
< 0.00006
0.0094
-
< 0.00008
—
< 0.000006
89
0.012
1.3
* Value in pH units
                           313

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mercury-ore  mining  and beneficiation are described in this
section.

Water Uses

Historically, water has had only limited use in the mercury-
ore milling industry.  This is primarily because little,  if
any,  beneficiation  of  mercury  ore was necessary prior to
roasting the ore for recovery  of  mercury.   In  the  past,
mercury ore was typically only crushed and/or ground to pro-
vide  a  properly  sized  kiln  or  furnace  feed.  However,
because high-grade ores  are  nearly  depleted  at  present,
lower-grade  ores  are  being  mined,  and  beneficiation is
becoming more important as a result of the need for  a  more
concentrated furnace or kiln feed.

Currently   in   the  United  states,  one  small  operation
(mine/mill 9201) is using  gravity  methods  to  concentrate
mercury  ore.   In  addition, a large operation (mill 9202),
which opened during 1975, employs  a  flotation  process  to
concentrate  mercury ore.  In both of these processes, water
is a primary  material  and  is  required  for  the  process
operating conditions.  Water is the medium in which the fine
and  heavy  particles  are separated by gravity methods.  In
the flotation  process,  water  is  introduced  at  the  ore
grinding  stage  to  produce  a  slurry which is amenable to
pumping, sluicing, and/or classification for sizing and feed
into the concentration process.

Water is not used in mercury mining operations and  is  dis-
charged,  where it collects, only as an indirect result of a
mining operation.  This water normally results from  ground-
water  infiltration  but may also include some precipitation
and runoff.

Water  flows  of  the  flotation  mill  and  the   operation
employing  gravity  beneficiation  methods  are presented in
Figure V-30.

Sources of Wastes

There are two basic sources of effluents: those  from  mines
and the beneficiation process.  Mines may be either open-pit
or  underground operations.  In the case of an open pit, the
source of the  pit  discharge,  if  any,  is  precipitation,
runoff  and ground-water infiltration into the pit.  Ground-
water infiltration is the primary source of water in  under-
ground  mines.   However,  in some cases, sands removed from
mill tailings are used to backfill stopes.  These sands  may
                            314

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Figure V-30. WATER FLOW IN MERCURY MILLS 9101 AND 9102
  (NO DISCHARGE)
-^ ^N
MILL
RESERVOIR
	 	 ^

^ 16.4 m3/day
(4,320 gpd)
GRAVITY-
MILL
^f TAILING
^V POND
t
                        1,649 m3 day (432.000 gpd)
                          (a) MINE/MILL 9201
  (NO DISCHARGE)

' 1.63m3/min """"
(430 gpm)
FLOTATION
MILL
t
^f TAILING A
5.4 m'/min (1.430 gpm) \^?ONO ^^/
fc /CLARIFICATION
^\ POND

                                     3.8 m3/min (1,000 gpm)

                       •DUE TO BEGIN OPERATION IN 1975.

                          (b) MINE/MILL 9202
  (NO DISCHARGE. WATER NOT USED.
  BENEFICIATION LIMITED TO
  CRUSHING AND/OR GRINDING TO
  PROVIDE FURNACE FEED.)
                  (c) OTHER MERCURY OPERATIONS
                              315

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initially  contain 30 to 60 percent moisture, and this water
may constitute a major portion of the mine effluent.
The particular waste constituents present in a mine or  mill
discharge  are  a  function of the mineralogy and geology of
the ore body and the particular milling process employed, if
any.  The rate and extent to which the minerals  in  an  ore
body  become  solubilized are normally increased by a mining
operation, due to the exposure of sulfide minerals and their
subsequent oxidization to sulfuric acid.  At  acid  pH,  the
potential for solubilization of most heavy metals is greatly
increased.

Wastewater  emanating  from  mercury  mills  consists almost
entirely of process water.   High  suspended-solid  loadings
are  the  most characteristic waste constituent of a mercury
mill waste stream.  This is primarily due to  the  necessity
for fine grinding of the ore to make it amenable to a parti-
cular  beneficiation  process.   In  addition, the increased
surface area of the ground ore enhances the possibility  for
solubilization or suspension of the ore minerals and gangue.
Although  the  total  dissolved-solid  loading  may  not  be
extremely high, the proporition of  dissolved  or  suspended
heavy-metal concentration may be relatively high as a result
of  the highly mineralized ore being processed.  These heavy
metals, the suspended solids, and process  reagents  present
are the principal waste constituents of a mill waste stream.
In  addition, depending on the process conditions, the waste
stream may also have a  high  or  lew  pH.   The  pH  is  of
concern,  not  only  because  of its potential toxicity, but
also because of its effect on the solubility  of  the  waste
constituents.

Quantities of Wastes

The  few  mercury  operations  still active obtain their ore
from open-pit mines.  In the past, however, more than 2/3 of
the domestic production was from ore mined from  underground
mines.   No discharge exists from the open-pit mines visited
or  contacted  during  this  study.    Also,   no   specific
information  concerning  discharges from underground mercury
mines  was  available  during  the  period  of  this  study.
However,  it  is  expected that, where discharges occur from
these underground mines, the particular metals  present  and
the  extent  of  their  dissolution depend on the particular
geology and mineralogy of the ore body and on the  oxidation
potential and pH prevailing within the mine.
                            316

-------
Silica  and  carbonate  minerals  are  the common introduced
gangue  minerals  in  mercury  deposits,  but   pyrite   and
marcasite may be abundant in deposits formed in iron-bearing
rocks.   Another mineral association includes stibnite which
is rare but is more common  than  orpiment.   Other  metals,
such  as gold, silver, or base metals, are generally present
in only trace amounts.

Process Description - Mercury Mining

Mercury  ore  is  mined  by  both  surface  and  underground
methods.   Prior  to  1972, underground mining accounted for
about 60 percent of the ore and 70 percent  of  the  mercury
production  in  the  U.S.   Currently, with market prices of
mercury falling, only a couple of  the  lower-cost  open-pit
operations remain active.

The mode of occurrence of the mercury deposit determines the
method of mining; yet, with either type, the small irregular
deposits  preclude the large-scale operations characteristic
of the majority of the U.S. metals mining industry.

Process Description - Mercury Milling

Processes for the milling of mercury which require water and
result in the waste loading of this water are:

    (1)   Gravity methods of separation

    (2)   Flotation

One mercury operation (mill 9201)  visited  employs  gravity
separation methods of beneficiation; the volume of the waste
stream emanating from this mill is approximately 1,679 cubic
meters  (440,000  gallons)  per  day.   In addition, another
operation (mill 9202)  which began  production  during  early
1975  was  contacted.   This mill employs a flotation process
and discharges 5.5 cubic meters  (1,430 gallons)  of water per
minute.    These  waste  streams  function  to  carry   large
quantities  of  solids  out  of the mill,  while the coarser
material is easily settled out, the very fine  particles  of
ground ore (slimes)  are normally suspended to some extent in
the  wastewater  and  often  present  removal problems.  The
quantity of suspended solids present in a  particular  waste
stream  is  a  function of the ore type and mill process, as
these factors determine how finely ground the ore will be.

In addition to suspended solids, solubilized  and  dispersed
colloidal  or  adsorbed  heavy  metals may be present in the
                            317

-------
waste  stream.   Metals  most  likely  to  be   present   at
relatively high levels are mercury; antimony; and, possibly,
arsenic,  zinc,  cadmium,  and  nickel.  The levels at which
these metals are present depend on the extent to which  they
occur   in   the  particular  ore  body.   Calcium,  sodium,
potassium,   and   magnesium   normally   are    found    at
concentrations of 10 to 200 parts per million.

In the past, little beneficiation of mercury ores was accom-
plished   and  typically  was  limited  to  crushing  and/or
grinding.  In a few cases,  gravity  methods  were  used  to
concentrate  the  ore.   These  practices require no process
reagents.  However, the  operation  (mill  9202)  employs  a
flotation  process,  which  requires  the  use  of flotation
reagents.  These reagents add to the waste  loading  of  the
mill  effluent  as  they  are  consumed in the process.  The
reagents used at this mill are listed in Table V-56.

Mill 9201 currently  beneficiates  mercury  ore  by  gravity
methods.   The ore is first crushed, washed, and screened to
provide a feed suitable for gravity separation.  The ore  is
concentrated  by tabling, which essentially involves washing
the crushed ore slurry across a vibrating  table  which  has
ridges  and  furrows  formed in parallel on its surface.  As
the ore slurry is washed across this surface, the heavy  ore
minerals collect in the furrows, while the fines are carried
across  the  ridges  and  discarded.   The  vibrating action
causes the heavy minerals to travel along the furrows to the
end of the table, where they are collected.

Waste characteristics of mill effluents  of  the  operations
visited are presented in Table V-57.

Uranium, Radium, and Vanadium Ores

Water  use;  flow;  and the sources, nature, and quantity of
wastes during the processes of uranium, radium, and vanadium
ore mining and beneficiation are described in this  section.
For vanadium-ore mining and beneficiation, only those opera-
tions  beneficiating  ores containing source material  (i.e.,
uranium  and  thorium)  subject  to   NRC   licensing,   are
considered here.

Water  Use.   Uranium ores often are found in arid climates,
and water is conserved as an expensive asset in refining  or
milling  uranium,  vanadium,  and  radium  ores,  some mines
yield an adequate water supply for the associated mill,  and
a  wateruse  pattern as shown in part  (a) of Figure V-31 can
be employed.  Here, all or part of the mine water is used in
                             318

-------
TABLE V-56. EXPECTED REAGENT USE AT MERCURY-ORE FLOTATION
          MILL 9202
REAGENT
Aerofloat 25
Z-11 (Sodium isopropyl xanthate)
Sodium Sulfide
PURPOSE
Frother
Collector
Depressing
Agent
CONSUMPTION
kg/metric ton
ore milled
0.05
0.15
0.13
Ib/short ton
ore milled
0.1
0.30
0.25
                     319

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TABLE V-57. WASTE CHARACTERISTICS AND RAW WASTE LOADINGS AT
           MILLS 9201 AND 9202
u
Hg
"tJH CONCEIT
nit, TRATIO
(mg/ft
WASTE LOAD
* in kg/1 000 metric tons
In/1000 short tons)
of coneentrete produced
9201 83
9202 8.2 27.6 44.000

MILL
9201
9202
MILL
9201
•202
MILL
•201
-9202
MILL
•201
•202
MILL
9201
•202
in kg/1000 metric tons
tb/1000 short tons)
of era milled
-
363
(7261
So

TRATION
Img/A)

~
Zn
CONCEN-
TRATION
Img/il
0.14
1.10
WASTE LOAD
in kg/1000 metric tons
(lb/1000 short tons)
of eoneentrete produced
1.930
(3360)
1300
(3.600)
in kg/1000 metric tons
(h/1 000 short tons)
of ore milled
0.014
(0.0281
IB
130)
Cu
CONCEN-
TRATION
Img/^l
-
1.3
WASTE LOAD
In kg/1 000 metric tons
(lb/1000 short tons)
of concentrete produced
-
2.100
14,2001
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
-
17
(34)
Nl
CONCEN-
TRATION
imgll 1

2.4
WASTE LOAD
In kg/1000 metric tons
(lb/1000 short tons)
of concentrate produced
-
3,800
(7,600)
in kg/1000 metric tons
(lb/1000 Short tons)
of ore milled
-
32
(64)
Fe
CONCEN-
TRATION
(me/ 4)
<0.6
2880
WASTE LOAD
in kg/IOOO metric tons
(fc/1000 short tons)
of concentrete produced
< 6,900
K1 3,800)
4310,000
(9.220,000)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<0.06
K0.10)
38,000
(78.000)
As
CONCEN-
TRATION
ImgU)
0.02
0.38
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrete produced
270
(5401
600
11,200)
in kg/1000 metric tons
do/ 1000 short tons)
of or« milled
0002
10304)
5
(10)
Mn
CONCEN-
TRATION
Imo/JU
50.0
7.04
WASTE LOAD
In kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
888300
(1 ,376/XKM
11,300
(23,000) '
in kg/1000 metric tons
(lb/1000 short toni)
of ore milled
6
1101
93
(186)
SULFIDE
CONCEN-
TRATION
Img/f.)

-------
the mill and then rejected to an impoundment, from which  it
is  removed  by  evaporation and seepage.  Mine water — or at
least, that portion not needed in the  mill — is  treated  to
remove  values  and/or  pollutants.   sometimes  the treated
water is reintroduced to the mine for  in- situ  leaching  of
values.   Wastewater from the impoundment is recycled to the
mill when conditions warrant, and additional  recycle  loops
(not  shown  in  the  figure)  may  be  attached to the mill
itself.

When mines are dry or  too  far  from  the  mill  to  permit
economical  utilization of their effluents, the mill derives
water from wells or, rarely, from  a  stream  (part  (b)  of
Figure  V-31).  In these instances, any mine water discharge
may be treated to remove uranium values  and/or  pollutants,
and  these  are then shipped to the mill (part  (c) of Figure
V-31) .

There are completely  dry  underground  mines  and  open-pit
mines  that lose more water by evaporation than they gain by
infiltration  from  aquifers.   All  known  mills  xn   this
industry segment use a hydrometallurgical process.

The  quantity  of  water  used  in  milling  is variable and
depends upon the process used and  the  degree  of  recycle.
From  thSe considerations, the effluent quantities are also
variable.  Acid leach mills generally  produce  between  1.5
and  2.5 tons of liquid per ton of ore; alkaline leach mills
from 0.3 to 0.8 tons of liquid per ton of ore.

Waste constituents

Radioactive Waste constituents.   Radium is one of the  most
potentially ^^ardous  radionuclides.   The  chemistry  of
radiw is similar to that of calcium, barium and  strontium.
The  Environmental  Protection  Agency  has proposed interim
drinkina wate? standards for radium -226 and radium -228  at
5 pCi/;g(pKSur!es per liter, total for both radionuclides.
             = v,ai^-Hfe of 1.620 years, is generated by the
ra           decty of uranium,' which has the very long half-
life^f l.Si billion years.  In uranium  ores  that  are  in
lire 01 4.DJ. uj.**i   j      years,  an  equilibrium  may  be
        heS between ?he rWof decay of uranium into radium
        rate ol decay of radium into  its  daughters.   Once
        uilibrium  is  established,  %•*"££•?*-*}
            i   a.*,-,  va^io  of  the  nan— lives— —i. e. ,   z. /
        ^n  equilibrated ore with a typical grade of 0. 22
        uran?™ £uld contain 0.82 microgram of  radium  per
                            321

-------
           Figure V-31. TYPICAL WATER-USE PATTERNS
 r
IN-SITU LEACH
i
MINE



TREATMENT



MILL
t
i
Rl
	 *•( IMPOUN
ECYCLE
                   (a) WET MINE/MILL COMPLEX
TREATMENT


MILL
                                        I
                                              •RECYCLE
                       (b) SEPARATED MILL
i
      IN-SITU LEACH
MINE


TREATMENT
                 DISCHARGE





                     (c) SEPARATED WET MINE
                          322

-------
kilogram.    Geological   redeposition   reduces  the amount of
radium  in  the  ore.   Because milling processes preferentially
dissolve uranium  and leave radium in solid tailings,  actual
concentrations of   radium  in  tailing-pond  solutions  are
approximately  17  to 81,000  picograms  per  liter.   These
concentrations are  often quoted in curies (Ci)—i.e., 17 to
81,000  picocuries per liter  (pCi/1)—since  the  radioactive
source  strength  of a  quantity  of  radium  in  curies is
essentially  equal to its content  of   radium  by  weight  in
grams.   (Source   strength  unit  for   radionuclides has been
defined as that quantity of radioactive material that decays
at a rate of 37 billion  (3.7 x 10  exp 10)  disintegrations
per  second.)   In   an   acid leach circuit, about 50% of the
thorium and  0.4 to 6.7% of the radium  are dissolved.

Thorium.   There  are other radioactive species  that  result
from the decay of uranium.  Thorium-230, with a half life of
80,000  years,  along with  lead 210  and polonium-210, with
half lives of  222 years  and  139  days,  respectively,  are
considered   along  with  radium-226.   Thorium is observed in
tailing-pond solutions in concentrations from  about  10  to
477 000 pCi/liter.   A maximum concentration for thorium-230
of 2,000 pCi/liter and for radium-226  of  30  pCi/liter  has
been  recommended  by 10 CFR 20 for release to unrestricted
areas   Generally, it has assumed that methods  for  control
of  radium-226 provide adequate control over thorium and the
other radionuclides  of interest.

Chemical  and  Physical   Waste   Constituents.      Chemical
contaminants"  of  milling  wastewaters derive from compounds
introduced in  milling operations or are dissolved  from  ore
in  leaching.   The   common  physical  pollutants—primarily,
suspended solids—figure prominently in discharges from  wet
mines   and  in  the management  of   deep-well disposal and
recycle systems.  One ton of ore containing 4 Ib of U3O8 has
about 515 mCi  of activity from  each   member  of  the  decay
chain   with   a  total  combined  alpha and beta activity of
about'7 200  mCi   About 85% of the total activity ends up in
?he mill waste^and  about 15% is  in   the  uranium  product.
With^ pa^ent'remaining, the thorium-234 and protactium-234
decay   out   of  the   mill  wastes so that,  after a year, the
wastes  contain about  70% of the activity originally  present
in the  ore.

           oollutants (particularly,  metals)  are expected to
           :he waste  streams of specific plants ^tjight be


              individually  but would appear in  waste-stream
                            323

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analysis under  class  headings  (e.g.r  as  TOG,  oils  and
greases,  or  surfactants).  In one specific example, it has
been observed that oils and greases that are known to  enter
alkaline  leach  processes  disappear  and  are  replaced by
approximately   equivalent   quantities   of   surfactants—
presumably,  by saponification (the process involved in soap
manufacture).  Table V-58 shows waste constituents  expected
from mills based upon the process, chemical consumption, and
the  ore mineralogies which are commonly encountered.  These
substances are shown in three groups:  those  expected  from
acid  leach  processes,  those  expected from alkaline leach
processes, and metals expected to be leached  from  the  ore
during  milling  processes.   Table V-59 shows two groups of
constituents  (among  the  sets  of  parameters  which  were
analyzed  both in background waters and waste streams):  (1)
Constituents that were found to exceed background by factors
from three to ten; and (2) Constituents that were  found  to
exceed  background by a factor of more than ten.  Comparison
of Tables V-58 and V-59 illustrates that more,  rather  than
fewer,   pollutants  are  observed  to  be  "added"  by  the
operation than are predicted from process chemistry and  ore
characteristics.     Observed    pollutant    increases   in
conjunction with toxicant lists  were,  therefore,  used  to
select  the parameters on which field sampling programs were
to concentrate.   (See also Section  VI.)    Table  V-59  also
illustrates    some    specific    differences   among   the
subcategories of SIC 1094 that are further explored  in  the
following discussion.

Constituents  Introduced  in  Acid  Leaching.  Acid leaching
(discussed  in   Section   III)   dissolves   numerous   ore
constituents,  approximately  five  percent of the ore, that
appear in the process stream; upon successful extraction  of
uranium  and  vanadium  values,  these  ore constituents are
rejected to tailing solutions.  In plants using a  sulfuric-
acid  leach,  calcium,  magnesium,  and  iron  form sulfates
directly.   Phosphates,  molybdates,  vanadates,   sulfides,
various oxides, and fluorides are converted to sulfates with
the  liberation  of phosphoric acid, molybdic acid, hydrogen
sulfide, and  other  products.   The  presence  of  a  given
reaction  product  depends  on the type cf ore that is being
used; since this is variable, pollutant parameters  must  be
selected  from an inclusive list.  The major pollutant in an
acid leach operation is  likely  to  be  the  sulfuric  acid
itself,  since  a  free  acid  concentration  of  one to one
hundred grams of acid per liter is maintained in the leach.

Excess free acid remaining  in  the  leach  liquors  and  in
solvent  extraction  raffinates (nonsoluble portions) can be
                            324

-------
    TABLE V-58. WASTE CONSTITUENTS EXPECTED
   ACID LEACH PROCESS
 ACID-LEACH CIRCUIT:
   Sulfurie acid
   Sodium chlorate
 LIQUID/SOLID-SEPARATION CIRCUIT
   Polyacrylamides
   Guar gums
   Animal glues
 ION-EXCHANGE CIRCUIT:
   Strong base anionic resins
   Sodium chloride
   Sulfurie acid
   Sodium bicarbonate
   Ammonium nitrate
 SOLVENT-EXTRACTION CIRCUIT:
   Tertiary amines
    (usually, alamine-336)
   Alkyl phosphoric acid
    (usually, EHPA)
   Isodecanol
   Tri butyl phosphate
   Kerosene
   Sodium carbonate
   Ammonium sulfate
   Sodium chloride
   Ammonia gas
   Hydrochloric acid
PRECIPITATION CIRCUIT:
   Ammonia gas
   Magnesium oxide
   Hydrogen peroxide
   ALKALINE LEACH PROCESS
 ALKALINE-LEACH CIRCUIT:
   Sodium carbonate
   Sodium bicarbonate
 ION-EXCHANGE CIRCUIT:
   Strong base anionic resins
   Sodium chloride
   Sulfurie acid
   Sodium bicarbonate
   Ammonium nitrate
 PRECIPITATION CIRCUIT:
   Ammonia gas
   Magnesium oxide
   Hydrogen peroxide
METALS LEACHED FROM ORE
BY MILLING PROCESSES
  Magnesium
  Copper
  Manganese
  Barium
  Chromium
  Molybdenum
  Selenium
  Lead
  Arsenic
  Vanadium
  Iron
  Cobalt
  Nickel
                     SOURCE: Reference 24
                           325

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       TABLE V-59. CHEMICAL AND PHYSICAL WASTE CONSTITUENTS
                  OBSERVED IN REPRESENTATIVE OPERATIONS
MINE/
CATEGORY
9401 /
ALKALINE
9402/
ACID

9403/
ALKALINE
9404
ACID
9405/
ACID
9406/
MINE
CONSTITUENTS THAT EXCEED BACKGROUND*
BY FACTORS BETWEEN THREE AND TEN
Color, Cyanide, Nitrogen as Ammonia, Phosphate,
Total Solids, Sulfate, Surfactants
Pb
Acidity, COD, Color, Dissolved Solids, Phosphate,
Total Solids
Ag. B, Ba, Hg, Zn

Color, Dissolved Solids, Fluoride. Sulfate, Total
Solids, Turbidity
Chloride, Color, Dissolved Solids, Total Solids,
Turbidity
Ag, Hg, K, Mg, Na
Color, Conductivity, Fecal Coliform, Hardness,
Phosphate, Suspended Solids, Total Solids,
Turbidity
Al, As, B, Ba, Be, Ca, Cd. Cr. Cu, Fe, Hg, Mg. Mo,
Ni. Pb, Sb, Se, Zn
Ammonia, Chloride, Hardness, Nitrate, Nitrite, Oil
and Grease, Organic Nitrogen, Sulfate, Total Solids,
Turbidity
As. B, Be, Ca. Mg. Na
CONSTITUENTS THAT EXCEED BACKGROUND*
BY A FACTOR OF MORE THAN TEN
Alkalinity, COD, Fluoride, Nitrate
As, Mo, V
Ammonia, Chloride, Sulfate
Al. As. Be. Cr, Cu. K, Mg, Mn. Mo, Na. Ni. Pb, V

Chloride, COD. Nitrate, Surfactants, Suspended
Solids, TOC
As, Mo, Na. Ti. V
Acidity, Ammonia, Sulfate, Suspended Solids
Al, As, Cr, Fe, Mn, Ni, Pb, Ti, V, Zn
Chloride, COD. Dissolved Solids. Kjeldahl Nitrogen.
Nitrate, Volatile Solids
Co. K, Mn, Na
(Nona among the analyzed items)
'"Background" is defined in text.
                           326

-------
 recycled  to advantage.   In some   operations,   this   acid  is
 used   to   condition  incoming  ores   by   reaction with  acid-
 consuming gangue.   Although this  step aids  in  controlling  pH
 of  raw wastes,  it  does  not reduce the  amount  of   sulfates
 therein.

 Oxidants   are   added  to the  acid   leach  liquor  following
 initial contact with  ore and after reducing gases,   such  as
 hydrogen   and   H2j3,   have been driven from  the slurry.   They
 act in conjunction with an iron content  of  about 0.5 g/1  to
 assure that uranium is in the U(VI)  valence  state.  Sodium
 chlorate  (NaClOJ)  and manganese dioxide   (Mn02j  serve   this
 purpose  in quantities  of  1  to   H g/1.  The  species  of a
 pollutant in the effluent will normally  be  one of   the   more
 oxidized  forms—e.g., ferric rather  than ferrous iron.

 Constituents Introduced in  Alkaline  Leaching.    Alkaline
 leaching  is less likely to solubilize compounds  of  iron  and
 the light metals  and has no effect  on the  common carbonates
 of  the gangue.   Sulfates and  sulfides,  in   the   oxidizing
 conditions  required  for conversion of  U(IV)  to U(VI),
 consume sodium  carbonate and, together with the  sulfate  ion
 generated  in   the common method of sodium removal, pollute
 wastewaters.

 The wastewater  of  an  alkaline leach  mill  is largely derived
 from    two  secondary   processes (Figure  V-32):   tailing
 repulping,  and  purification (or sodium removal) .  The  leach
 itself  is   recycled  via the recarbonaticn  loop.  The wastes
 discarded  to   tailings   often  contain   organic   compounds
 derived  from   the ores.   Oxidizing agents  are   used  in
 leaching,  but air  and oxygen gas  under  pressure have   been
 found   to  serve  as  well as more expensive oxidants and  to
 reduce  pollutant  problems.   The   concentrations  used   in
 alkaline   leach are  only  of  academic  interest because  of
 recycling.   Sodium carbonate concentration varies from 40  to
 50  g/1; sodium  bicarbonate concentration, from 10 to 20  g/1.

An  ammonium  carbonate   process  that   leads  directly  to  a
 sodiumfree   uranium  trioxide product  has been investigated.
 It  is more selective  for  uranium than  the  sodium   carbonate
 process,  but vanadium, while not being recovered, interferes
with    uranium   recovery.   The  process  does   not  require
bicarbonate  and  could   produce  ammonium  sulfate,  as   a
byproduct  (Section  III).  A flow chart of an ammonium car-
bonate  process  is  shown  in Figure V-33.

Constituents Introduced  in Concentration  Processes.    Ion-
exchange   (IX)   resins  are ground into small particles that
                            327

-------
Figure V-32. ALKALINE-LEACH WATER FLOW
                                   TO ATMOSPHERE
          GROUND ORE
           ALKALINE
            LEACH
FRESH WATER
OR
TAILING-SOLUTION
RECYCLE


FILTERING


                               LEACH
                              RECYCLE
                                  I	
     USED
    STACK
     GAS
           PREGNANT
            LEACH
             i
         PRECIPITATION
            CRUDE
           PRODUCT
                               RECARBONATION
  T
               STACK
               GAS
BARREN
 LEACH
                             REPULPED
                             'TAILINGS
FRESH
WATER


PURIFICATION
(SODIUM REMOVAL)
WASTE
WATER
                                            TO TAILING
                                            POND
             END
           PRODUCT
              i
              TO
          STOCKPILE
                328

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Figure V-33. AMMONIUM CARBONATE LEACHING PROCESS


MINING

ORE

AIR

WATER

— ^
—+•

GRINDING
1
PRESSURE
,
r
LEACHING
.
COUNTERCURRENT
DECANTATION
SLU
l
*RY
' •?
j
^ LEACH
"^"SOLUTION
PREGNANT
SOLUTION
TO ATMOSPHERE
1
WASTE GAS
1
AMMONIA AND
CARBON DIOXIDE SI
ABSORPTION TOWERS
i




' t
1
1
1
1
GAS
1
1
STEAM STRIPPI
PRECIPITATIO
SLURRY
FILTRATION

PRODUCT
1

•EAM
..
. .
NG
A
N



TO TAILING 1
POND T
                                               TO
                                            STOCKPILE
                    329

-------
appear  among  suspended  solids  in  raw   waste   streams.
Solvents   are   not  completely  recovered  in  the  phase-
separation step of solvent-exchange (SX) concentration.  The
extent of the contributions of each of these  pollutants  is
difficult to judge by observation of the waste stream, since
there   are   no  specific  analysis  procedures  for  these
contaminants.   Some  prediction  of  the  concentration  is
possible  from  the  observable  loss  of  (IX)  resin and SX
solvents.  Only a small fraction of  IX  resin  is  actually
lost  by the time it is replaced because of breakage; in one
typical operation, the loss amounts to about 100 kg (220 Ib)
per day at a plant that has an inventory of about 500 metric
tons (551 short tons) of resin,  and  handles  3,000  metric
tons  (3,307  short  tons)  per day of ore and about as much
water.  The raw waste concentration of IX resin can thus  be
estimated as about 30 ppm.  Standard tests for water quality
would measure this as a contribution to total organic carbon
(TOC)  which  is  also  due  to  other sources (for example,
organic ore constituents).  Most of this contribution is  in
suspended  solids;  this is illustrated by the fact that TOC
is only about 6 mg/1 in the supernatant  of  the  raw  waste
stream discussed above.

Solvents  are  lost  at  a rate of up to 1/2000 of the water
usage in the SX circuit.  This ratio is set by the  solubil-
ities  of  utilized solvents, which range from 5 to 25 mg/1,
and by the fact that inadequate slime separation can lead to
additional loss to tailing solids.  TOC  of  the  raw  waste
supernatant at mills using SX was found to be 20 to 24 mg/1.
It  is,  again,  impossible  to  determine what part of this
measurement should be ascribed to SX solvents—particularly,
in view of highly carbonaceous ores.

The  most  objectionable  constituents   present   in   mill
effluents  may be the very small amounts (usually, less than
6 ppm)  of the tertiary amines or alkyl  phosphates  employed
in  solvent extraction.  In some cases, these compounds have
been found to be toxic to fish.  An analytic  procedure  for
the entire class of these materials and their decay products
is  not  available,  and they must be identified in specific
instances.

Difficulties in distinguishing among solvents,  ion-exchange
resins, carbonaceous ore constituents, and their degradation
products  made  it  impossible  to  discriminate between the
wastes of mills using SX or IX processes.  Since some of the
solvents have structures with potential for toxic effects in
their degradation products, it would be desirable  to  trace
their fates as well as those of ion-exchange resins.  Future
                            330

-------
research in this field could lead to better characterization
and improved treatment of wastewater.

Process  Descriptions,  Water Use, and Waste Characteristics
for Uranium, Radium, and Vanadium Ore Mining and Milling

Four mine/mill complexes in the licensed segment of the  SIC
1094   category   were   visited  to  collect  data  on  the
utilization of water and  the  characteristics  of  raw  and
treated  wastes.  Water use in the mines and mills is listed
in Table V-60, and treatment systems employed are listed  in
Table V-61.

The  consumption  of  water is seen to vary from 0.75 to 4.3
cubic meters per metric ton (180 to 1,000 gal per short ton)
of ore capacity, with an average of 1.35  cubic  meters  per
metric  ton  (323 gal per short ton).  Two of the operations
(9401 and 9404) derive their water supply  from  wells,  and
one  (9403)  obtains  its water from a stream, in the manner
shown in Figure V-34c.  The fourth operation (9402)  utilizes
mine water,  where mine water is available, at least some of
it is treated by ion exchange  to  recover  uranium  values.
Water  use  in  representative  operations is illustrated in
Figure v-34, and  the  water-flow  configurations  of  these
operations  are illustrated in Figures V-35, V-36, V-37, and
v-38.  While an attempt was made to obtain a  water  balance
in each case, there are some uncertainties.  In Figure V-35,
for  example,  the  loss  from  tailings  by  evaporation is
probably not quite equal to the raw  waste  input  from  the
plant,  and  expansion  of  the  tailing-pond  area  may  be
necessary.  Similarly, it proved difficult  to  account  for
the  rain  water entering the open pit mine of the operation
in Figure V-38.  If and when it rains into this  mine,  some
water  evaporates  immediately  from  the surface, while the
rest  runs  into  a  central  depression   or   seeps   into
underground  aquifiers.   The  first  and  last  effects  in
combination are clearly dominant; less than ten  percent  of
the  calculable  water  input  is seen to evaporate from the
central depression (Figure V-38).

Waste Characteristics  Resulting  From  Mining  and  Milling
Operations.    Two  of  the  operations visited use alkaline
leaching, and two use acid leaching, for extraction of  ura-
nium  values.   Only one operation discharges from the mill,
while two others discharge from mines.  Among the five  NRC-
licensed  subcategories  listed  in  Section  IV, only mills
employing a combination process of acid-and-alkaline  leach-
ing are not represented by the plants visited.   An operation
representing  this  sutcategory  was not visited because its
                            331

-------
     TABLE V-60. WATER USE AND FLOWS AT MINE/MILLS 9401,9402, 9403,
                AND 9404
WATER CATEGORY

Water Supply
Discharge
Supplied to Mill
Recycled to Mill
Lois (Evaporation, etc.)

Makeup Water
Water in Circuit
Discharge
Evaporation and
Seepage
WATER USED
MINE/MILL 9401
m3/day

8,339
3,339
0
8,000
estO
gpd
MINE/MILL 9402
m3/day
9Pd
MINE/MILL 9403
m3/day
MINE PORTION
2,203,000
882,100
0
1,321,000
estO
11,562
4,325
5,307
0
1,920
3,052,000
1,143,000
1,402,000
0
507,200
N/A
N/A
N/A
N/A
N/A
gpd

N/A
N/A
N/A
N/A
N/A
MILL PORTION
2,700
3.200
0
2,700
713,300
845,400
0
713,300
5,307
8,900
0
5,307
1,402,000
2,351.000
0
1,402,000
6.060
6.580
5.400
860
1,601,000
1.738,000
1,427.000
174.400
MINE/MILL 9404
m3/day

est 1,530
0
0
0
est 1,530
gpd

est 404,200
0
0
0
est 404,200

6.300
5.3OO
0
5,300
1,400,000
1,400,000
0
1,400,000
 N/A - Not available
       TABLE V-61. WATER TREATMENT INVOLVED IN U/Ra/V OPERATIONS
FEATURE

Settling Basin
Evaporating Pond
Ion-Exchange Plant
PARAMETER

MINE/MILL
9401
MINE PORTION
ATM in hectares lacrail
Ratantion Tima in hours
Ares in hectares (acres)
U3 Og Concentration in ntg/l
U3 OB Removal in K
0.3 10.74)
est 20
N/A
25
96
9402

0.7 (1.71
est 80
N/A
2 to 12
98
9403

N/A
N/A
N/A
N/A
N/A
9404

N/A
N/A
2 (4.9)
N/A
N/A
MILL PORTION
Tailing Pondlt)
Ion- Exchange Plant
Racarboniiar
DMpWall
Utilization of
Aral in hectares (acres)
Number series-connected
Daily Water Usa in metric torn tthort tons)
Daily Water Usa in metric tons (short tons)
Capacity in metric tons (short tons) water par day
Sand/Slime Separators
Decant Facilities
Filters
Coprecipitation
21 (51.8)
1
490 (540)
1.635(1302)
0
Yes
Yes
TOTAL OPERATION
Ore Handling
Capacity in metric tons (short Ions) per day II 3,200(3.527)
100 (2471
5
N/A
N/A
0
Yes

6,400(7,065)
24 159.3)
3
N/A
520(573)
0
Yes

1,400(1,543)
107(284)
1
N/A
N/A
1,635(1402)
Yes
Yes
Yes

2,700(2,976)
N/A - Not available
                          332

-------
Figure  V-34. WATER  FLOW  IN  MILLS 9401, 9402, 9403, AND  9404


                  	 3,339 ms/d«y   	
   I  8,338 m3/**
     (220.200 gpd)
                    ION EXCHANGE
                                                D«CHAROE
     -IN-tlTU LEACH •
                          5,000 m3/diy
                          (1.320.000 gpd)
c
                            600 m3««v (132.000 gpd)
                                                    ION EXCHANOe
       2.700 M3/div
       (713,300 Bid)
                                MILL
            1,636 m3/d«y
            (431.900 gpdl
     II	1  RECARBONATION  •«-!
                                  3700 m3M«y
                                  (848^00 gpdl
                                                                  TAILING
I 2,700 m3/d«r
 (713,300 gpd)


Lj   LOS,  I
                                      y
(38^00 gpdl







TO ATMC






SPHERE






                                    (d) MILL 9404
                                      333

-------
Figure V-35. FLOWCHART OF MILL 9401
TO ATMOSPHERE
t

Nn Co» . ».


MINING
ORE
\
GRINDING
AND
LEACHING
EVAPORATION MaOH I
Ml 400 m3/day
(105.700 gild)
1
\ r~
PRELIMINARV
PRECIPITATION
1 i


FILTRATION
I
REPULPING
i


STAGE-2
FILTRATION
1
VANADIC ROASTING 	 	
I 	 ACID •« 	 *~D H2S°4 NH3
SOLUTION ™£J
I
\ '
TO
STOCKPILE
2.700 m3/d.y
S- 	 — N. (713,300 gpdl
I WELL 1 ,, .. j, ^ , ton
ER
CH 	
r
*" PURIFICATION


REPULPING
1


STAGE -3
FILTRATION
i
^[ . I J REPULPING
r 	 1 I

VELLOWCAKE | ION ^ 1132,100 gpd) ^
r
TO STOCKPILE
J txtMANUt ^ ^s>>

1
TAIL
por
i
TO ATMOSPHERE
T
1 EVAPORATION 1
«it 400 m-*/d«y ,
1 (10S. 700 gpd)

t

	 «*• REPULPING
«« 1






i i

TO ATMOSPHERE
RAIN
•m m-M.u
(5.280 gpd)
1 4 3
T Mt ft9n mj/«4*u
ING^^
tD s
v (243.000 gpd) _ f"
' L

EVAPORATION
•it 920 m3/diy
(243.000 gpd)
r
ENTRAINMENT
AND SEEPAGE

             334

-------
              Figure V-36. FLOW CHART FOR MILL 9402
                      MINING
     5.300 m3/day
     (1,400,000 gpd)
                       ORE
  CRUSHING
    AND
  GRINDING
"2S°4
NaCIO
 NaCI
 NH,
                     LEACHING
                    THICKENERS
                     SOLVENT
                    EXTRACTION
PRECIPITATION
                   YELLOW CAKE


                   TO STOCKPILE
                                    EVAPORATION
                                    AND SEEPAGE
                                                             5,300 m3/day
                                                             (1,400.000 gpd)
                               335

-------
         Fifura V-37. FLOW CHART OF MILL 9403
                         0.6 m3/div
                          (0.1 gpml
I2SX U3Og + 1.6% U2O5)
                   FLOAT TAILS
                   ALTERNATIVE
  ACID FILTRATION
                                                                 WASH EQUIVALENT TO MOISTURE
                                                                  RETAINED IN CAKE (ELIMINATED
                                                                          IN DRYER)
                         FROM ALKALINE-
                            TAILING SUMP
                              FROM ACID-
                            TAILING SUMP
                                           MAIN TAILING
                                              SUMP
                                                       2,843 m«lnc lom/diy
                                                       I3.134ihontom/diy>
                                                           LIQUID
                                                                                 EVAPORATION
                                                                  TAILING POND WITH 23-HECTARE
                                                                             LIQUID POOL
   TO STOCKPILE
                                 336

-------
                    Figure V-38. FLOW CHART OF MILL 9404
   MILL WATER SUPPLY = 5,300 m3 (1,400,000 gal) per day
                                TO ATMOSPHERE



                                EVAPORATION
                                       .137 m3/day
                                       (36,200 gpd)
  1,530 m3/day
* (404,200 gpd)
            320-hectare (790-acre)
              OPEN-PIT MINE
                       2-hectare (4.9-acre
                         EVAPORATION
                            POND
                                                          CRUSHING
                                                            AND
                                                          GRINDING
                                                         ACID LEACH |
TOTAL LOSS OF
 5,300 m3/day
 (1,400,000 gpd)
                                                      HYDROCYCLONES
               TAILING
                POND
                                               RESIN-IN-PULP
                                               ION EXCHANGE
                                                            PREGNANT
                                                            SOLUTION
                                        BARREN
                                       SOLUTION
                                                           ELLUENT
                                                           (HIGH CD
CAPACITY OF
 1,635 m3/day
 (431,900 gpd)
                                                      PRECIPITATION
                                                          AND
                                                        FILTERING
              DEEP WELL
                                 337

-------
processes were changed recently.  During the visits to these
mills, industry plans that change water use by factors of up
to ten, and which will take place within a year,  were  pre-
sented.   The  data on raw wastes presented in the following
discussion are based mostly on analyses of samples  obtained
during site visits.

The  data  obtained  are  organized into several broad waste
categories:

    1.   Radioactive nuclides.

    2.   Organics, including TOC, oil  and  grease,  surfac-
         tants, and phenol.

    3.   Inorganic  anions,  including   sulfide,   cyanide,
         fluoride,    chloride,    sulfate,   nitrate,   and
         phosphate.

    4.   Light  metals,   relatively   nontoxic,   including
         sodium,  potassium,  calcium,  magnesium, aluminum,
         titanium, beryllum, and the ammonium cation (NH4+).

    5-   Heavy metals, some of which  are  toxic,  including
         silver,  aluminum, arsenic, barium, boron, cadmium,
         chromium, copper, iron, mercury,  manganese,  moly-
         bdenum,  nickel, lead, selenium, strontium, tellur-
         ium, titanium,  thallium,  uranium,  vanadium,  and
         zinc.

This  class  is  further  subdivided into the metals forming
primarily cationic species and those forming anionic species
in the conditions characteristic of raw SIC 1094 wastes  (in
particular, chromium, molybdenum, uranium, and vanadium).

    6.   Other pollutants (general characteristics), includ-
         ding acidity, alkalinity, COD, solids, color, odor,
         turbidity and hardness.

Radioactive Nuclides.   Decay products  of  uranium  include
isotopes  of  uranium,  thorium, proactinium, radium, radon,
actinium, polonium, bismuth, and lead.  These decay products
respond to mining and milling processes in  accordance  with
the  chemistries  of  the  various  elements  and,  with the
exception of the bulk of uranium  isotopes,  appear  in  the
wastes.   Approximately ninety percent or more of the radium
226 remains with solid tailings and sediment  in  mine-water
settling  basins.  Concentrations of raw waste of radium and
uranium observed here should not be released to the environ-
                            338

-------
merit.  The  amounts  that  have  been  observed  under  this
program  are  shown  in  Table  v-62,  where it is seen that
alkaline mills are highest, mines are  second  highest,  and
acid  mills are lowest in the radium content of wastes.  The
high levels encountered at mines are  partially  explainable
by buildup in the recycle accompanying ion-exchange recovery
of  uranium.   Recycle  also  explains the high radium loads
found at alkaline mills.  The low concentrations observed at
acid mills are partially due to the low solubility of radium
sulfate  (formed by reaction with sulfuric acid leach)  and to
the   lack   of   recycle,   but   concentrations-shown   in
parentheses—for  an  evaporation  pond— indicate that such
impoundments may become a pollution hazard  to  ground-water
supplies.

Orqanics.   Organics derived from carbonaceous ores and from
chemicals  added  in  processing  are  measured  as TOC and,
occasionally,  are  distinguishable  as  oils  or   greases,
surfactants,  or phenol.  The small amounts of organics that
are observed are reviewed in Table V-63.

Inorganic Anions.   These  may  be  distinguished  into  two
classes:   (1)  Sulfides,  cyanides, and fluorides, for which
technically  and  economically  feasible  treatments   (e.g.,
oxidation and lime precipitation) are readily available; and
(2) Chlorides, sulfates, nitrates, and phosphates, which are
present  in  fairly  large concentrations in mill wastes and
cannot be removed economically.   Distillation  and  reverse
osmosis,  while  technically  feasible,  raise  the  cost of
recovered water and requires  a  large  energy  expenditure.
Impoundment,  in  effect, results in distillation in regions
like the southwestern  states.   other  anions  are  grouped
together  in  conjunction  with  the  light-metal cations as
total dissolved solids and are found in the levels shown  in
Table V-611.

Light  Metals.   The ions of sodium, potassium, and ammonium
found in wastewaters are subject to inclusion in  the  cate-
gory   of   total   dissolved  solids.   Calcium,  titanium,
magnesium, and aluminum respond to  some  treatments   (e.g.,
lime  neutralization)  and are shown separately.  Table V-65
shows  concentrations  of  aluminum,   beryllium,   calcium,
magnesium,  and  titanium  found  in wastewater effluents of
mines and mills covered in this ore category.

Heavy Metals.  The leach processes in  the  uranium/vanadium
industry  involve  highly  oxidizing conditions that leave a
number of ore metals—specifically, arsenic, chromium, moly-
bdenum,  uranium,  and  vanadium—in  their  most   oxidized
                            339

-------
     TABLE V-62. RADIONUCLIDES IN RAW WASTEWATERS FROM
                URANIUM/RADIUM/VANADIUM MINES AND MILLS
RADIONUCLIDE
and units of measurement
RADIUM 226
in picocuriaj/£
THORIUM
in mg/JZ,
URANIUM
in mg/£
CONCENTRATION
MINES
200 to 3,200
<0.1
4 to 25
ACID MILLS
200 to 700
(4,100)«
(1.1»*
30 to 40
ALKALINE MILLS
100 to 19,000
N/A
4 to 45
'Parentheses denote values measured in wastewater concentrated by evaporation
N/A = Not available
 TABLE V-63. ORGANIC CONSTITUENTS IN U/Ra/V RAW WASTE WATER
PARAMETER
Total Organic Carbon (TOO
Oil and Grease
MBAS Surfactants
Phenol
CONCENTRATION (mg/<>)
MINES
16 to 45
3 to 4
0.001 to 7
<0.2
ACID MILLS
6 to 24
1
0.5
<0.2
ALKALINE MILLS
1 to 450
3
0.02
<0.2
                        340

-------
 TABLE V-64. INORGANIC ANIONS IN U/Ra/V RAW WASTEWATER
PARAMETER
Sulfide
Cyanide
Fluoride
Total Dissolved Solids (TDS)
CONCENTRATION (mg/l)
MINES
<0.5
<0.01
0.45
1,400 to 2,000
ACID MILLS
<0.5
<0.01
< 0.01
15,000 to 36,000
ALKALINE MILLS
< 0.5
< 0.01 to .04
1.4 to 2.1
5,000 to 13,000
TABLE V-65. LIGHT-METAL CONCENTRATIONS OBSERVED IN U/Ra/V
          RAW WASTEWATER

PARAMETER
Aluminum
Beryllium
Calcium
Magnesium
Titanium
CONCENTRATION (mg/l)
MINES
0.4 to 0.5
0.01
90 to 120
35 to 45
0.8 to 1.1
ACID MILLS
700 to 1,600
0.08
220
550
7
ALKALINE MILLS
0.2 to 20
0.006 to 0.3
5 to 3,200
10 to 200
2 to 15
  TABLE V-66. CONCENTRATIONS OF HEAVY METALS FORMING
            ANIONIC SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Arsenic
Chromium
Molybdenum
Uranium
Vanadium
CONCENTRATION (mg/£ )
MINES
0.01 to 0.03
< 0.02
0.5 to 1.2
2 to 25
as to 2.1
ACID MILLS
0.1 to 2.5
2 to 9
0.3 to 16
30 to 180
120
ALKALINE MILLS
0.3 to 1.5
<0.02
<0.3
4 to 50
0.5 to 17
                  341

-------
states,  often as arsenates, chromates, molybdates, uranates
and vanadates.  These anionic species are,  typically,  much
more  soluble  than cations of these metals that precipitate
as hydroxides or sulfides in response to  lime  and  sulfide
precipitation  treatments.   Most  of  these  anions  can be
reduced to lower valences by excess sulfide  and  will  then
precipitate   (actually,  coprecipitate  with each other)  and
stay in solid form if  buried  by  sediment.   The  observed
range  of  concentrations  for  the anionic heavy metals for
mines and mills visited is shown in Table V-66.  One or more
of the heavy metals is observed in  high  concentrations  in
each type of operation.

The  cationic  heavy  metals that had been expected to occur
from data on ores and  processes  include  lead,  manganese,
iron, and copper.  Field sampling results added nickel, sil-
ver, strontium, and zinc to this list.  The observed concen-
trations  of  these metals are shown in Table V-67.  Cadmium
was found in a concentration above the lower detection limit
(20 micrograms per liter) at one alkaline mill discharge.

Other pollutants.  Acid leach mills discharge a  portion  of
the  acid  leach; alkaline leach mills discharge sodium car-
bonate; and mine water is found to  be  well  buffered  with
measurable  acidity  and alkalinity.  Chemical oxygen demand
is occasionally high, and raw wastes, reslurried only to the
extent needed for transport to tailings, carry a  high  load
of  total  solids.   These factors are reflected in the data
shown in Table V-68.  These measures indicate the  need  for
settling,  neutralization, and aeration of the wastes before
discharge.   Those  treatments   alsc   effect   significant
reductions  in other pollutants; for example, neutralization
depresses heavy metals, and aeration reduces organics.

Waste Loads in Terms of  Production.   The  loads  of  those
pollutants that indicated conditions warranting treatment at
the exemplary plants were related to ore production to yield
relative  waste  loads.  The data for three subcategories of
the SIC 1094 segment are presented in Table V-69 (mines)  and
Tables V-70, V-71, V-72, and V-73 (mills).

Occasional large ratios between the parameters  observed  at
differing  operations are believed to be due to ore quality.
The point is illustrated by TOC at mills 9401 and 9403:  The
operators  of  mill  9401  had  contracted  to  run  an  ore
belonging to mine 9404 on a toll basis.  The ore  carried  a
high  carbonaceous material content that caused water at the
9401 mill to turn brown and may have adversely affected  the
concentration process at mill 9404.  Mill 9403, in contrast.
                            342

-------
  TABLE V-67. CONCENTRATIONS OF HEAVY METALS FORMING
            CATIONIC SPECIES IN U/Ra/V RAW WASTE WATER
PARAMETER
Silver
Copper
Iron
Manganese
Nickel
Lead
Zinc
CONCENTRATION (mg/£ )
MINES
<0.01
<0.5
0.2 to 15
< 0.2 to 0.3
< 0.01
0.07 to 0.2
0.02 to 0.03
ACID MILLS
<0.01
0.7 to 3
300
100 to 210
1.4
0.8 to 2
3
ALKALINE MILLS
0.1
<0.5to1
0.9 to 1.6
< 0.2 to 40
0.5
<0.5 to 0.7
0.4
TABLE V-68. OTHER CONSTITUENTS PRESENT IN RAW WASTEWATER
          IN U/Ra/V MINES AND MILLS
PARAMETER
Acidity
Alkalinity
Chemical Oxygen
Demand (COD)
Total Solids
CONCENTRATION 
-------
 TABLE V-69. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE LOAD
             FOR URANIUM MINES 9401 AND 9402
PARAMETER
TSS
COD
TOC
Alkalinity
Ca
Mg
Fa
Mo
V
Ra
Th
U
MINE 9401
CONCENTRATION
(mg/&)
IN WASTEWATER
—
242
15.8
224.4
93
46
0.47
0.5
1.0
3,190*
-
12.1
RAW WASTE LOAD
kg/day
-
2.300
150
2,100
860
420
4
5
9
29.700f
-
113
fc/day
—
5,200
320
4,600
1.900
920
10
11
20
—
-
248
MINE 9402
CONCENTRATION
(mfl/Jl)
IN WASTEWATER
299
600
25
-
117
36
0.23
0.53
<0.5
2,710"
<0.1
11.6
RAW WASTE LOAD
kg/day
640
7,OOO
290
-
1,300
410
3
6
<6
31,100*
0.2
134
fc/day
1.400
15,000
640
-
3,000
910
6
13
<13
—
<2.5
294
 *Valua in picocuria*/£
  Value in picocuriM/day

 TABLE V-70. CHEMICAL COMPOSITION OF RAW WASTEWATER AND RAW WASTE
             LOAD FOR MILL 9401 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COD
TOC
Alkalinity
Cu
F.
Mn
Pb
A«
Mo
V
H«
U
Fluoride
CONCENTRATION
(ing/ )
IN WASTEWATER
294,000
65.6
460
12,200
<0.5
0.92
<0.2

-------
     TABLE V-71. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
                 LOAD FOR MILL 9402 (ACID- OR COMBINED ACID/ALKALINE-
                 MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Acidity
Al
Cu
Mn
Pb
As
Cr
Mo
V
Ra
U
CONCENTRATION
IN WASTEWATER
525,000
63.5
24.0
35,000
1,594
2.7
105
2.1
2.3
9.0
16.0
125
234'
31.1
TOTAL WASTE
kg/day
4,100,000
337
127
185,700
8,460
14
557
11
12
48
85
663
1,240*
165
Ifa/day
9,000,000
743
281
409,500
18.600
32
1,228
25
27
105
187
1,462
» —
364
RAW WASTE LOAD
per unit ore milled
kg/ metric ton
1,000
0.082
0.031
45
2.1
0.003
0.14
0.003
0.003
0.012
0.021
0.16
0.30tf
0.040
Ib/short ton
2,000
0.16
0.062
91
4.1
0.007
0.27
0.005
0.006
0.023
0.041
0.32
0.27*"
0.080
per unit concentrate produced*
kg/metric ton
450,000
37
14
20,400
930
1.6
61
1.2
1.3
5.2
9.3
73
136"
18
to/short ton
900,000
74
28
40,800
1,860
3.1
122
2.4
2.7
10
18.7
146
124***
36
 •On the basis of 1973 production of 98.2% U3O8 and 1.8% MO3
  Value in picocuries/Jl
 ••Value in microcuries/day
  Value in mierocuries/metric ton
•••Value in microcurias/short ton
                                    345

-------
   TABLE V-72. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
               LOAD FOR MILL 9403 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COD
TOC
Alkalinity
Ca
Mo.
Ti
Al
Cu
Fa
Mn
Ni
Pb
Zn
At
Mo
V
Ra
Th
U
Fluorida
CONCENTRATION
IN WASTEWATER
111,000
27.8
<1
1.150.6
3.200
190
0.395
18
1.1
1.6
38
0.52
0.69
< 0.5
1.4
< 0.3
< 0.5
111'
<0.1
3.9
1.4
TOTAL WASTE
kg/day
1,400,000
145
< 5.2
5.980
16.640
990
2.1
94
5.7
8.3
198
2.7
3.6
< 2.6
7.3
< 1.6
< 2.6
580"
<0.5
20
7.3
Ib/day
3,100,000
319
< "
13.190
36,680
2,180
4.5
206
13
18
436
6
7.9
< 6.7
16
< 3.4
< 5.7
<1
45
16
RAW WASTE LOAD
par unit ora mil lad
kg/matrieton
1.000
0.1
< 0.0037
4.3
12
0.71
0.0015
0.07
0.0041
0.0059
0.14
0.0019
0.0026
< 0.0019
0.0052
< 0.0011
< 0.0019
0.41 "
< 0.0004
0.032
0.0052
Ib/ihort ton
2.000
0.2
< 0.0074
8.5
24
1.4
0.0029
0.13
0.0081
0.012
0.28
0.0039
0.0051
< 0.0037
0.01
< 0.0022
< 0.0037
0.37*"
-.0.0008
0.064
0.01
par unit concantrata producad*
kg/ metric ton
1,050.000
109
3.9
4.500
1.3
743
1.5
70
4.3
6.3
149
2.0
2.7
2
6.5
< 1.2
2
431 ft
C0.4
34
5.5
Ib/short ton
2.100.000
217
7.8
9.000
2.5
1,486
3.1
141
8.6
13
297
4.1
5.4
4
11
< 2.3
4
392*"
cO.8
68
11
 •On tha buit ol 1973 production of 67% U,O- and 33% CuS.
 t
  Valua in picocuriat/,'
 • • Valiu in mierocuria>/day
  Valua in microcuriai/matric ton
•"Value in mierocuria*/*hort ton
                                 346

-------
    TABLE V-73. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
                 LOAD FOR MILL 9404 (ACID- OR COMBINED ACID/ALKALINE-
                 MILL SUBCATEGORY)
PARAMETER
TSS
COD
TOG
Acidity
Ca
Mg
Ti
Al
Co
ft
Mn
Ni
Pb
Zn
At
Cr
Mo
V
Ri
U
CONCENTRATION
(mg/S.)
IN WASTEWATER
350.000
629
6.2
4,040
224
550
3
740
0.68
325
210
1.38
0.84
<0.5
0.13
2
< 0.3
120
690'
174.5
TOTAL WASTE
ho/day
2,700,000
3,330
33
21.400
1,190
2,920
16
3,920
3.6
1.720
1,110
7.3
4.5
< 2.7
0.69
11
< 1.6
640
3,660"
925
Ib/day
6,000,000
7,350
72
47.200
2,620
6,430
35
8,650
7.9
3.800
2,450
16
9.8
< 5.8
1.5
23
< 3.5
1,400
-
2.035
RAW WASTE LOAD
p«r unit ora milled
kg/ metric ton
1.000
1.2
0.012
7.9
0.44
1.1
0.0059
1.5
0.0013
0.64
0.41
0.0027
0.0016
< 0.00098
0.00026
0.0039
< 0.00059
0.24
1.35tf
0.38
Ib/ihort ton
2.000
2.5
0.024
15.8
0.88
2.2
0.012
2.9
0.0026
1.28
0.82
0.0054
0.0033
< 0.002
0.00051
0.0079
< 0.0012
0.47
1.23'"
0.75
par unit concantrata produced'
kg/matrie ton
530.000
651
6.4
418
232
569
3.1
766
0.7
336
217
1.4
0.9
0.52
0.13
2.1
0.31
124
718"
180
Ib/ihort ton
1.060.000
1,300
12.8
836
464
1.139
6.2
1,532
1.4
673
435
2.9
1,7
1.0
0.27
4.1
0.62
248
652"*
361
 •On tha basis of 1973 production of 100% U3Og.
  Value in picocurie$/£
 "Value in microcuriet/day
  Value in microcuriei/metric ton
""Value in microcurief/short ton
                                 347

-------
was  concentrating its own, much cleaner, ore.  The ratio of
200:1 in TOC is, therefore, expected.

Metal Ores 3, Not Elsewhere Classified

This section discusses the water uses,  sources  of  wastes,
and  waste loading characteristics of operations engaging in
the mining and  milling  of  ores  of  antimony,  beryllium,
platinum-group metals, rare earth-metals, tin, titanium, and
zirconium.    The  approach  used  in  discussion  of  waste
characteristics of these (SIC 1099) metal processes includes
a general discussion of water uses and sources of wastes  in
the entire group, followed by a description of the character
and  quantity  of wastes generated for each individual metal
listed above.

Water Uses.   The primary use of  water  in  each  of  these
industries  is  in  the  beneficiation  process, where it is
required for the operating conditions of the process.  Water
is  a  primary  material  in  the  flotation  of   antimony,
titanium,  and  rare-earth  minerals;  in  the  leaching  of
beryllium ore; in the concentration of titanium,  zirconium,
and rare-earth minerals  (monazite) from beach-sand deposits;
and  in  the  extraction  of platinum metals from placers by
gravity  methods.   No  primary  tin  ore  deposits  of  any
commercial  significance  are  currently  being mined in the
U.S.  However, a small amount  of  tin  is  recovered  as  a
byproduct  of  a  molybdenum  operation  through  the use of
flotation and magnetic methods.

Water is introduced into  flotation  processes  at  the  ore
grinding  stage  to  produce  a  slurry which is amenable to
pumping, sluicing, or classification  for  sizing  and  feed
into the flotation circuit.  In leaching processes, water is
the  solvent  extraction  medium.   Water also serves as the
medium for gravity separation of heavy minerals.

In underground mining of antimony ore and in open-pit mining
of titanium and beryllium ores, water is not  used  directly
but, rather, is present  (if at all) only as an indirect con-
sequence  of  these  mining  operations.  The mining of sand
placer deposits  for  titanium,  zirconium,  and  rare-earth
minerals  is  done  by dredging, in which a pond is required
for  flotation  of  the  barge.   In  mining  a  placer  for
platinum-group  minerals,  a  barge may be floated either in
the stream or on an on-shore pond, depending on the location
of the ore.
                            348

-------
Water flows of the antimony, beryllium, platinum, rare-earth
titanium,  and  zirconium  mineral  operations  visited  are
presented in Figures V-39, V-40, and V-41.

Sources   of  Wastes.    There  are  two  basic  sources  of
effluents: those from mines or dredging operations  and  the
beneficiation  process.   Mines  may  be  either open-pit or
underground operations.  In the case of  an  open  pit,  the
source  of  the  pit  discharge  (if  any) is precipitation,
runoff, and ground-water infiltration into  the  pit.   Only
one  underground  mine  was  encountered in the SIC 1099 ore
mining   industry—an   antimony   mine—and   no   existing
discharges  have been reported at this time.  Effluents from
beach-sand dredging operations  orginate  as  precipitation,
runoff,   and   groundwater   infiltration.    In  addition,
effluents result from the  fresh  water  used  in  wet  mill
gravity  beneficiation  of  the sands and, subsequently, are
usually discharged into dredge ponds.

The waste constituents present in a mine or  mill  discharge
are functions of the mineralogy of the ores exploited and of
the  milling  or extraction processes and reagents employed.
Acid conditions prevailing at a mine site  also  affect  the
waste  components  by  influencing  the  solubility  of many
metallic components.

Wastewater  from  a  placer  or  sand  mining  operation  is
primarily  water  that  was  used  in a primary or secondary
gravity separation process.  Also, where a placer  does  not
occur  in  a  stream,  water is often used to fill a pond on
which the barge is floated.  The process water is  generally
discharged  into  either  this  pond or an on-shore settling
pond.  Effluents of the settling pond usually  are  combined
with the dredge-pond discharge, and this comprises the final
discharge.  The principal wastewater constituents from these
operations  are  high  suspended solid loadings and coloring
due to high concentrations of humic acids  and  tannic  acid
from  the  decay  of organic matter incorporated into former
beach sands and gravels being mined.

Wastewater  emanating  from  mills  processing   lode   ores
consists  almost entirely of process water.  High suspended-
solid loadings are the most characteristic waste constituent
of a mill waste  stream.   This  is  primarily  due  to  the
necessity  for  fine grinding of the ore to make it amenable
to a particular beneficiation  process.   In  addition,  the
increased  surface  area  of  the  ground  ore  enhances the
possibility for both solubilization and  suspension  of  the
ore  minerals  and gangue.  Although the total dissolved and
                            349

-------
Figure V-39. WATER FLOWS AND USAGE FOR MINE/MILLS 9901 (ANTIMONY) AND
           9902 (BERYLLIUM)
            (NO DISCHARGE!

' 306 TO 382 m3«iy ~~
180.000 TO 100.000 gpd)
FLOTATION
MILL


2M TO 3*3 m3/diy ~~
(7S.OOO TO 90.000 gpd)
TAILING-
IMPOUNDMENT


EVAPORATION
AND
SEEPAGE

                                                                     (NO
                                                                     DISCHARGE)
                         (a) ANTIMONY MINE/MILL 9901
                                                 TO ATMOSPHERE
                        (b) BERYLLIUM MINE/MILL 9902
                                  350

-------
Figure V-40. WATER FLOWS AND USAGE FOR MINE/MILLS 9903 (RARE EARTHS)
           AND 9904 (PLATINUM)
                                                          TO ATMOSPHERE
                                1.6 m3/minutt
                                 (412 gpm)

               NOTE: FOR BYPRODUCT RECOVERV. SEE PART (b) OF FIGURE V-41 (MINE/MILL MOW

                         (a) RARE-EARTH MINE/MILL 9903
RIV
j
OfcJ '
24,730 m3/diy V.^

24,730 m3/d«y
(6.480.000 «pdl
1REDGE
POND
1
^
^^/ 48,500 m3/diV

49.500 m3/diy
I12.S60.000 gpdl
DREDGE WITH
WET GRAVITV
BENEFICIATION



                        (b) PLATINUM MINE/MILL 9904
                               351

-------
Figure V-41. WATER FLOWS AND USAGE FOR TITANIUM MINE/MILLS 9905 AND 9906
OPEN-PIT
MINE

2,668 m3/day


                                                 DISCHARGE
                                                    TO
                                                   RIVER
     FLOTATION AND
  MAGNETIC-SEPARATION
          MILL
             36,069 m3/day
            (9,450.000 gpd)
                     INTERMITTENT
                 DISCHARGE (SEASONAL)
35.19 m3/day
(9,220.000 gpd)
                                            '/day
                                       (230,000 gpd)
                           (a) TITANIUM MINE/MILL 9905
                                                                  TO ATMOSPHERE
»•
ORE
PO
«• -
I
*•
WEL
••«__
DGE \g
NO /^
I
L-OVER
12,099 m3/d
(3.170.000 g
L(S) )

WET
MILL BULK
(GRAVITY "CONCENTRATE"^
SEPARATION)
< /^TAILING
12.099 m-Vday \^ SYSTEM
(3.170.000 gpd) ^*— •— i 	 -*"
•V i i
pd)
12,595 mj/day
(3,300,000 gpd)
RAIN AND
RUNOFF

DRY
MILL
(ELECTROSTATIC
AND
MAGNETIC METHODS)


>L 7,633 m3/day
) (2,000,000 gpd)

» "
17,175 m3/day
(4,500,000 gpd)

(


i
EVAPORATION

DISCHARGE
TO
STREAM
                  (b) TITANIUM/ZIRCONIUM/MONAZITE MINE/MILL 9906
                                  352

-------
suspended solid loading  may  not  be  extremely  high,  the
dissolved  and  suspended  heavy  metal concentration may be
relatively high as a result of the  highly  mineralized  ore
being  processed.  These heavy metals, the suspended solids,
and  process  reagents  present  are  the  principal   waste
constituents of a mill waste stream.  In addition, depending
on  the process conditions, the waste stream may also have a
high or low pH.  The pH is of concern, not only  because  of
its  potential  toxicity,  but also because of its effect on
the solubility of the waste constituents.

Wastewater  emanating  from  a  beach-sand   dredging   pond
consists  of  water in excess of that needed to maintain the
pond at the proper level.  This water also originates as wet
mill effluent and, as a result, contains  suspended  solids.
However,  the  primary waste constituents from these milling
operations  are  the  humic  and  tannic  acids  which   are
indigenous  to  the ore body and which result in coloring of
the water.

Description of Character and Quantity of Wastes

The quantity of wastes resulting  from  mining  and  milling
activities  is  discussed below individually for each of the
SIC 1099 metals.

Antimony

Process Description - Antimony Mining.  Currently, only  one
mine  exists  which ""is  operated solely for the recovery of
antimony ore   (mine  9901).   This  ore  is  mined  from  an
underground mine by drifting (following the vein).

As  indicated  in Figure V-39, no discharge currently exists
from the mine.

Process Description - Antimony Milling.   Only one  mill  is
operating  for  the  recovery of antimony ore as the primary
product.  This  mill  (9901)   employs  the  froth  flotation
process   to   concentrate  the  antimony  sulfide  mineral,
stibnite (Figure III-28).  The particular flotation reagents
used by this mill are listed in Table V-7U.  Water  in  this
operation  is added between the crushing and grinding stages
at the rate of 305 to 382 cubic meters  (80,000  to  100,000
gallons)  per  day.   There  is no discharge, but flow to an
impoundment totals 286 to 343 cubic meters (75,000 to 90,000
gallons) per day.
                            353

-------
TABLE V-74. REAGENT USE AT ANTIMONY-ORE FLOTATION MILL 9901
REAGENT
Dowfroth 250 (Polypropylene glycol
methyl ethers)
Aerofloat 242 (Essentially Aryl
dithiophosphoric acids)
Lead nitrate
PURPOSE
Frother
Collector
Activating
Agent
CONSUMPTION
kg/metric ton
ore milled
0.4
0.1
0.5
Ib/short ton
ore milled
0.8
0.2
1.0
                      354

-------
 Quantities of  Wastes.   Waste  constituents  originate  from two
 sources:  solubilization  and  dispersion  of ore   constituents
 and  consumption  of  the  milling  reagents.

 In   metal   mining  and  milling  effluents,  heavy-metal
 constituents   are   of   primary   concern,   due   to  their
 potentially  toxic  nature.  Metallic minerals known  to occur
 with antimony  in the commercially valuable ore body  of  mine
 9901 are:

         Stibnite             (Sb2S3)
         Pyrite               (FeS2J*
         Arsenopyrite         (FeAsS)
         Sphalerite          (ZnS)
         Argentite            (Ag2_S)
         Cinnabar             (HgS)
         Galena               (PbS)

 The  metals  in  these  minerals are the ones which  would be
 expected to occur at highest  concentrations  in  the  waste
 stream,  and   results   of raw-waste  analysis   support this
 conclusion (Table   V-75).   The  raw-waste characterization
 presented  in  Table  V-75  is  based  upon  the analysis of
 samples collected   during the  mill  visit.   As  would  be
 expected on the  basis of  the  mineralization of the ore body,
 the  metals present at  relatively high concentrations in the
 raw  waste are  antimony  (64.0  mg/1), zinc   (4.35  mg/1),  and
 iron  (18.8  mg/1).   Arsenic is not as high as was  expected
 but  is about an  order of  magnitude greater than  mean  back-
 ground  levels   reported  in  surface  waters of the Pacific
 Northwest Basin.  Waste loadings for important  constituents
 of wastewaters from mill  9901 are listed in Table V-76.

 Beryllium

 Process  Description  -   Beryllium Mining.  Beryllium ore is
 mined on a large scale  at only one domestic  operation.   At
 mine 9902, bertrandite  (H2Be4Si20j9)  is recovered by open-pit
 methods.   A small  amount of  beryl is also mined in the U.S.
by crude open-cut and hand-picking methods.  As indicated in
Figure V-39, no  discharge currently exists at mine 9902.

Process Description - Beryllium Milling.    Currently,  only
one  domestic beryllium operation uses water in a beneficia-
tion process.  This operation is identified as mill 9902 and
employs a proprietary   acid   leach  process  to  concentrate
beryllium oxide  from the ore.
                            355

-------
TABLE V-75. CHEMICAL COMPOSITION OF RAW WASTEWATER DISCHARGED FROM
          ANTIMONY FLOTATION MILL 9901
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TSS
TDS
Hardness
Chloride
COD
TOC
Al
As
Be
Ba
8
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
Total Mn
CONCENTRATION (mfl/£)
8.3"
8.5
11.0
113*
170
149
68
40
1.5
43
7.8
6.2
0.23
< 0.002
<0.3
<0.01
0.103
0.57
0.04
0.12
18.8
0.13
1.93
0.40
PARAMETER
Hfl
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Sb
Mo
Oil and Grease
MBAS Surfactants
Cyanide
Phenol
Fluoride
Total Kjeldahl N
Sulfide
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/JJI
0.0038
0.10
<0.05
<0.2
3.5
0.036
<0.02
2JO
0.11
<0.2
<0.5
4.35
64.0
<02
<1
1.9
<0.01
0.022
<0.1
15
OS
16.6
2.55
0.05
   •Value in pH units

    Value in cobalt units
                          356

-------
TABLE V-76. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD
           AT ANTIMONY MILL 9901
PARAMETER
pH
TSS
COD
TOC
Fe
Pb
Sb
Zn
Cu
Mn
Mo
CONCENTRATION
(mg/£) IN
WASTEWATER
8.3*
997
43
7.8
18.8
0.13
64.0
4.35
0.12
0.40
<0.2
RAW WASTE LOAD
per unit concentrate produced
kg/metric ton
—
74.78
3.22
0.585
1.41
0.0097
4.8
0.366
0.009
0.03
< 0.01 5
Ib/short ton
—
149.56
6.44
1.170
2.82
0.0194
9.6
0.652
0.018
0.06
<0.030
per unit ore milled
kg/metric ton
—
7.48
0.0322
0.059
0.141
0.00097
0.48
0.033
0.0009
0.003
< 0.001 5
Ib/short ton
—
14.96
0.0644
0.118
0.282
0.00194
0.96
0.066
0.0018
0.006
< 0.0030
•Value in pH units
                      357

-------
Quantities   of  Wastes.    As  indicated  in  Figure  V-31,
approximately 3,061 cubic meters  (802,000 gallons)  per  day
of   wastewater   are  discharged  from  mill  9902.   Waste
constituents originate from two sources:  solubilization and
dispersion of ore constituents and  consumption  of  milling
reagents.   However,  because  this  process  involves  acid
leaching, high  solubilization  is  observed  in  the  waste
constituents (Table V-22).

The mineralization of the ore body from which bertrandite is
obtained  is  essentially  that  presented in the tabulation
given below for mine 9902 (beryllium).
    Quartz         SiO.2
    Feldspar       Al silicates with Ca, Kr and Na
    Fluorite       CaF2
    Carbonates
    Iron Oxide Minerals
    Tourmaline     (XY3A16(B03)3(Si6018)(OH)4)
         where     X = Na, Ca; Y = Al, Fe(+3), Li, Mg

Constituents of these minerals are also expected to  be  the
main  constituents  in  the mill waste, and results of waste
analysis support this (Table V-77).  As indicated, the waste
stream from this leaching process is exceptionally  high  in
dissolved   solids   (18,380  mg/1),  consisting  largely  of
sulfate  (10,600 mg/1).  Fluoride  (45 mg/1) is  also  present
at  relatively  high  concentration,  as  are  aluminum  (552
mg/1), beryllium  (36 mg/1), and zinc  (19 mg/1).

Rare Earths

Process Description - Rare-Earth Metals Mining.   The  rare-
earth  mineral  monazite   (Ce,  La,  Th, Y) POUJ is recovered
predominantly as a byproduct  from  sand  placers  mined  by
dredging—primarily,  for  their  titanium  mineral content.
(Refer  to  information  on  mill  9906,  as  described  for
titanium.)   The  rare-earth  mineral  bastnaesite  is  also
currently recovered, as the primary product, by an operation
mining the ore from an open-pit mine  (mine 9903).

As indicated in Figure V-40, no discharge  currently  exists
at mine 9903.

Process Description - Milling.   Monazite is concentrated by
the  wet  gravity  and electrostatic and magnetic separation
methods, discussed in the titanium segment of this section.
                            358

-------
  TABLE V-77. CHEMICAL COMPOSITION OF RAW WASTEWATER FROM
             BERYLLIUM MILL 9902 (NO DISCHARGE FROM TREATMENT)
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
Acidity
Alkalinity
Hardness
COD
TOG
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
17,000*
88f
1.3
18,380
3,035
0
4,000
22
55
<1
0.76
552
0.15
36.0
<5.0
0.65
0.047
43.0
0.20
0.07
<0.5
<0.1
320.0
PARAMETER
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
Cyanide
Phenol
Total Kieldahl N
Sulfide
CONCENTRATION (mg/£)
49.0
0.15
< 0.05
<0.2
77.0
0.062
0.04
270.0
0.22
<0.2
< 0.5
19.0
< 0.2
170
45
10,600
1.25
0.8
< 0.01
< 0.01
0.19
< 0.5
'Value in micromhos/cm

tValue in cobalt units
                          359

-------
A single mill (9903) is currently  beneficiating  rare-earth
minerals   mined   from  a  lode  deposit.   Bastnaesite  is
initially  concentrated  by  the  froth  flotation   process
(Figure  V-42).    Flotation  of rare-earth minerals requires
rigidly  controlled  conditions  and  a  pH  of  8.95,   and
temperature-controlled  reagent  addition is critical to the
successful flotation of these minerals.   Rare-earth  oxides
(REO)  in  the mill heads range from 6 to 11 percent and are
upgraded in the flotation  circuit  to  a  concentrate  that
averages  57  to  65  percent REO, depending upon the heads.
This concentrate is leached with hydrochloric acid to remove
calcium and strontium carbonates, increasing the REO content
in the leached concentrate by as much as 5  to  10  percent.
This  concentrate is processed in a solvent extraction plant
to produce high-purity europium and yttrium oxides; a cerium
hydrate product; a concentrate  of  lanthanum,  praesodymium
and  neodymium;  and a concentrate of samarium and gadolinium
(Figure V-t»3) .

In the solvent extraction plant, the  flotation  concentrate
is initially dried and then roasted to remove carbon dioxide
and  to  convert  the  rare-earths to oxides.  These oxides,
with the exception of cerium oxide, are converted to soluble
chlorides   in   a   hydrochloric-acid   leaching   circuit.
Following  leaching,  the  acid  slurry  is passed through a
countercurrent decantation circuit.  The  primary  thickener
overflow  containing  the chlorides is fed into the europium
circuit, while the leached solids  from  the  countercurrent
decantation circuit make up the feed for the cerium process.

The  leach  liquor  (primary thickener overflow) is clarified
in a carbon filter and  adjusted  to  a  pH  of  1.0  and  a
temperature  of  60 degrees Celsius (140 degrees Fahrenheit)
prior to countercurrent extraction of europium with  organic
solvent  (90  percent  kerosene  and  10 percent ethyl/hexyl
phosphoric acid).  The raffinate from the extraction circuit
makes up the  feed  for  the  lanthanum  circuit,  which  is
discussed later.

After  loading  the  organic  with europium, the europium is
stripped in the solvent extraction  strip  circuit  with  flN
hydrochloric  acid.   The  pregnant  strip solution contains
iron,  which  is  removed  in  precipitation  tanks  by  the
addition  of  soda  ash to lower the pH to 3.0 to 3.5.  This
causes ferric hydroxide to precipitate, and the  precipitate
is  removed  in a pressure filter.  Following removal of the
iron, the europium bearing  solution  goes  through  another
solvent  extraction  and  stripping  circuit, similar to the
previous  one.   The  pregnant  strip   is   pumped   to   a
                            360

-------
Figure V-42. BENEFICIATION OF BERTRANDITE, MINED  FROM A LODE DEPOSIT,
            BY FLOTATION (MILL 9903)
                                CRUSHING
         FRESH
         WATER
                DM m'/mln
                (M|pm)
IBUgpiBl.
       GRINDING
                                         CYCLONE



__ EQUIPMENT 1 1*-** «P"«>







1 . .-
HIGH-INTENSITY


1 	
1 —FROTH FIRS
1

CYCLONE


^ FROTH— ' "

CLASSIFICATION 	 1 AND'MOLVBDENUM

I
i
THIRD STAGE OF _
CONDITIONING "*~
i S«*C
|(206«FI

1

FLOTATION CELLS
UNDERFLOW
* /•
• l*""D ROUGHER ^^ UNDERFLOW — »^te»
FLOTATION CELL "»«•"•»•«" —\^




. I 1 SODIUM
| ORZAN | | CARBONATE

fl
* •'
B°C
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SECOND STAGE OF ^ FIRST STAGE OF
CONDITIONING "™~ CONDITIONING
A SZ«C i
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^— '
FAM
PO
•w
1
	 1 FLOTATIONCEUS
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SECOND, THIRD, AND FOURTH f
CLEANER -p
FLOTATION CELLS
UNDERFLOW
SCAVENGER
FLOTATION CELLS
-FROTH— I

HYDROCHLORIC
ACID
j

BNCENTRATE -|- •» AO^TATORS
1
ALTERNATIVE
i

1






.ING ^
"r
1^6 m yminut*
1618 (pm)
fc LEACHED CONCENTRATE _
" THICKENER
[DRYER I I

ILER
^WAS
ALTERNATIVE
                                     FIN A
                                                               SEPARATION OF RARE-
                                                                EARTH METALS BY
                                                               SOLVENT EXTRACTION
                                                                (SEE FIGURE V-43)
                                      NAL
                                    PRODUCT
                                       I
                                   TO STOCKPILE
                                       361

-------
                       Figure V-43. BENEFICIATION OF RARE-EARTH FLOTATION CONCENTRATE BY
                                 SOLVENT EXTRACTION (MILL 9903}
10
TO WASTE
A.
* 1
1 t
FILTRATE OVERFLOW

FILTER "• — THICKENER
| \
DRYER 1 PRECIPITATION _^__ AMMONIA
' | ' 1 	 '
COARSE OR T
FINE
LANTHANUM CARBON
HYDRATE f'l-TER EUROPIUM
' ) PRODUCT

.• . f"
| I
FILTRATE
1
DRUM
FILTER



FLOTATION
CONCENTRATE
(FROM 1
FIGURE V-421 "


G

DRV
ADOLIMIUM/SAMAHIUM
CARBONATE PRODUCT
SODIUM CARBONATE I



GADOL IMI UM SAMARI UM
PRECIPITATION TANK
i
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BARREN
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|
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	 t REPUL
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.1 	 	 1 "
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                                                                                                ' STOCKPILES

-------
 purification  circuit,  where europium oxide is prepared for
 the market.

 Solutions from the purification circuit are neutralized with
 sodium  carbonate  to  produce   gadolinium   and   samarium
 carbonates, which are collected by a drum filter.

 Returning  to  the  countercurrent  decantation circuit, the
 solids remaining from leaching are  filtered  and  repulped.
 The cerium solids are then thickened, filtered, and dried to
 produce the final concentrate.

 As  mentioned  previously,  the  raffinate  from  the  first
 solvent  extraction  circuit  provides  the  feed  for   the
 lanthanum  circuit.  This raffinate is clarified in a carbon
 filter,  and  ammonia  is  added  to  precipitate  lanthanum
 hydrate.    The  precipitate  is  thickened  and  filtered to
 produce the final concentrate.

 Quantities of Wastes.    As indicated  in  Figure  V-40,   raw
 wastes  are  discharged  at a rate of 1.96 cubic meters  (518
 gallons)  per minute from the flotation circuit  and at a  rate
 of 0.08  cubic  meter  (21  gallons)   per  minute  from   the
 leach/solvent extraction plant.   These waste streams are not
 combined,  and  both are characterized in Table v-78.  These
 data are   based   upon   the  analysis  of   raw-waste  samples
 collected  during  the   mill visit.   Table V-79  presents the
 results of chemical analyses for the rare-earth metals.

 Reagents   used  in   the  flotation,    leach,    and   solvent
 extraction processes of mill 9903  are identified below.

 Flotation Circuit

 Frother            Methylisobutylcarbinol
 Collector               N-80 Oleic Acid
 pH Modifier         Sodium  Carbonate
 Depressants         Orzan,  Sodium Silicofluroide
 Conditioning Agent   Molybdenum Compound

 Leach  Circuit

 Leaching Agent          Hydrochloric Acid

 Solvent-Extraction Circuit

Leaching Agent     Hydrochloric Acid
Precipitants       Sodium Carbonate,  Ammonia, Sodium Hydrosulfide
Solvents           Kerosene, Ethyl/Hexyl Phosphoric Acid
                            363

-------
        TABLE V-78. CHEMICAL COMPOSITION OF RAW WASTEWATER
                   FROM RARE-EARTH MILL 9903
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TDS
TSS
HflrdnoM
coo
TOC
Oil and Grease
MBAS Surfactants
Si02
Al
At
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION  1,500
47
<1
21.2
1.25
<0.1
0.01
0.009
<0.01
< 0.005
2,910
0.04
<0.03
0.03
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfata
Nitrate
Phosphate
Cyanide
Phenol
CONCENTRATION (mg/£)
FLOTATION
.
-
0.5
-
-
< 0.3
-
-
-
-
-
-
•
•
.
-
365

-
-
-
-
LEACH/
SOLVENT
EXTRACTION
<0.05
6.6
3.0
0.85
<0.1
<0.3
94
0.015
0.09
650
4.5
3.36
7.0
< 0.003
<0.1
54,000
<0.1
2.3
1.50
0.09
<0.01
<0.01
•Value in pH units

 Value in cobalt units
                           364

-------
TABLE V-79. RESULTS OF CHEMICAL ANALYSIS FOR RARE-
          EARTH METALS (MILL 9903-NO DISCHARGE)
PARAMETER
Y
La
Ce
Pr
Nd
Sm
Eu
Gd
Th
CONCENTRATION (mg/ I )
LEACH WASTEWATER
—
442
24
6.2
9.6
0.27
< 0.001
< 0.001
< 0.001
FLOTATION RECLAIM WATER
0.014
1.32
2.75
0.27
0.51
0.041
< 0.001
0.006
< 0.001
               365

-------
In   rare-earth   metal   mining   and   milling,   effluent
constituents expected to be present are a  function  of  the
mineralogy  of  the  ore  and  the associated minerals.  The
principal minerals associated with the ore body of mine 9903
are:  bastnaesite (CeFC03, with La, Nd, Pr, SOT, Gd, and Eu);
barite (Bas04) ; calcite  (CaC03); and strontianite  (SrC03).

The dissolved-solid content of the  leach/solvent-extraction
waste  stream  is  extremely  high  (76,162 mg/1) and is due
largely to chlorides (54,000 mg/1).  The metals  present  at
highest  concentrations are those which would be expected on
the basis of known mineralization and use  in  the  process.
These  are  strontium   (4.5  mg/1)  and barium (less than 10
mg/1).  The high concentration of tellurium (3.36  mg/1)  is
unexplained  on  the  basis  of  known  mineralization,  but
mineralization is assumed to be the source of this  element.
Waste  characteristics  and  raw waste loading for the rare-
earth flotation and concentrate leaching/solvent  extraction
processes are given in Table V-80.

Platinum-Group Metals

Process  Description  -  Platinum  Mining.    Production  of
platinum group metals is largely as a byproduct of gold  and
copper  refining,  and  primary  ore  mining is limited to a
single dredging operation  (mine 9904), which  is  recovering
platinum-metal alloys and minerals from a placer deposit.

Process   Description  -  Milling.    Mill  9904  employs  a
physical  separation  process  to  teneficiate  the   placer
gravels  (Figure  III-20).  The dredged gravels are intially
screened, jigged, and tabled to separate the heavy  minerals
from  the  nonmineral lights, which are discarded.  Chromite
and magnetite are separated from  the  platinum-group  metal
alloys  and  minerals  by  magnetic  separation.   The final
platinum-group  metal  concentrate  is  produced  from   the
magnetic-separation product by dry screening and passing the
resultant  material through a blower to remove the remaining
lights.'

Quantities of Wastes .   Wastes resulting  from  the  mining
and   milling   activities   of  this  operation  cannot  be
considered separately, since the wet mill discharges to  the
dredge  pond.   No  reagents  are  required  in  the milling
process, and, as a result, the principal  waste  constituent
from this operation is suspended solids (30 mg/1).  Table V-
81  lists  the  chemical  composition  of the wastewater and
waste loads from mine/mill 9904.
                            366

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TABLE V-80. CHEMICAL COMPOSITION AND RAW WASTE LOAD FROM RARE-EARTH
            MILL 9903
PARAMETER
CONCENTRATION

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TABLE V-81. CHEMICAL COMPOSITION AND LOADING FOR PRINCIPAL WASTE
          CONSTITUENTS RESULTING FROM PLATINUM MINE/MILL 9904
          (INDUSTRY DATA)
PARAMETER
Alkalinity
Conductivity
Hardness
COD
BOD
TS
TDS
TSS
(N) IMH3
Kjeldahl Nitrogen
Al
Cd
Cr
Cu
Total Fe
Pb
Zn
Chloride
Fluoride
Nitrate
Sulfate
Sulfide
CONCENTRATION
(mg/U
WASTEWATER
83
109'
35.6
7.6
3.5
82
52
30
0.18
0.28
0.337
< 0.001
<1.0
<1.0
0.166
0.010
0.028
11.0
0.95
4.5
5.5
1.2
RAW WASTE LOAD
per unit ore milled
kg/1000 metric tons
1.20
—
0.51
0.11
0.05
1.18
0.75
0.43
0.003
0.004
0.005
< 0.00001
<0.01
<0.01
0.002
0.0001
0.0004
0.16
0.01
0.06
0.08
0.02
lb/1000 short tons
2.39
—
1.03
0.22
0.10
2.36
1.50
0.86
0.006
0.008
0.010
< 0.00002
<0.03
<0.03'
0.005
0.0003
0.0008
0.32
0.01
0.13
0.16
0.03
    •Value in micromhos/cm
    TS = Total Solids
                           368

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As indicated  in  Figure  V-40,  24,700  cubic  meters   (6.5
million  gallons)  per  day of water are discharged from the
dredge pond to the river.  The wet milling process  utilizes
49,500 cubic meters  (12.96 million gallons) per day.

The principal associated minerals in this placer  (mine 9904)
are:

    Chromite (FeCrj204)
    Ferroplatinum ~(Fe, Pt, Ir, Os, Ru, Rh, Pd, Cu, Ni) alloy
    Iridium/ruthenium/osmium alloy
    Taurite (Ru, Irr Os) S.2
    Unnamed mineral  (Ir, Rh, Pd)s
    Mertieite  (Pt5(Sb, As)2)
    Sperrylite (PtAs2)
    Gold (Au)

Tin

Tin  is recovered in the U.S. as a byproduct of a molybdenum
operation.  At this mine  (6102), the ore is mined by  glory-
hole  methods,  in which the sides of an open hole are caved
and the broken rock trammed out  through  a  tunnel  at  the
bottom  of  the hole.  No specific waste characteristics and
water uses can,  therefore,  be  assigned  for  this  mining
milling operation.

Titanium

Process   Description  -  Mining.    Titanium  minerals  are
recovered from lode and  sand  deposits.   The  single  lode
deposit  being  exploited  in  the U.S. is mined by open-pit
methods at mine 9905.  Ancient beach-sand placers are  mined
at   several  operations  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
normally are equipped with suction head cutters to mine  the
mineral  sands.   Wastes from dredge ponds and wet mills are
combined; therefore, these operations  are  discussed  under
one heading:  Dredging Operations.

Quantities  of  Wastes;   Mine  9905-    This  is  the  only
existing mine from which titanium lode ore is mined.   Water
is  discharged  from  this open pit at a rate of 2,668 cubic
meters (699,000 gallons) per day.  The chemical  composition
of  this  waste  is  presented in Table V-82.  As these data
show, oils and grease (3.0  mg/1),  fluorides  (3.20  mg/1),
total  Kjeldahl  nitrogen  (2.24  mg/1), and nitrates (15.52
mg/1)  are present at relatively  high  concentrations.   The
                           369

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    TABLE V-82. CHEMICAL COMPOSITION OF RAW WASTEWATER
                FROM TITANIUM MINE 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Total Kjeldahl N
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/W
1,000"
11.3*
0.37
1,240
14
6.4
138.2
546.4
6.4
10.3
3.0
0.32
2.24
0.1
0.1
0.003
< 1
0.01
< 0.002
94.5
<0.01
<0.03
0.33
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Sr
Ag
Na
Se
Te
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/£)
<0.05
26.0
<0.01
<0.01
<0.1
<0.5
13.0
0.129
<0.01
140.0
0.75
< 0.06
<0.2
0.007
<0.1
< 0.1
< 0.01
183.5
3.20
270
15.52
< 0.05
•Value in micromhos/cm

 Value in cobalt units
                            370

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oils and greases undoubtably result from the heavy equipment
used  in  the  mining  operations,  and  the  fluorides  are
indigenous to the ore body.  However,  the  reason  for  the
high   concentrations   of  nitrogen  and  nitrates  may  be
explained in part  by  the  use  of  nitrate-based  blasting
agents.

Process  Description  -  Titanium  Milling;  Mill 9905.  Ore
brought to this mill is beneficiated by a combination of the
magnetic-separation and flotation processes (Figure V-44).

The ore is initially crushed and then  screened.   Both  the
undersize  and  the  oversize screened ores are magnetically
cobbed to remove the nonmagnetic rock, which  is  discarded.
Oversize   magnetic  rock  undergoes  further  crushing  and
screening, while undersize material is fed into the grinding
circuit.  The latter utilizes grinding in rod  mills,  which
are  in circuit with "Ty Hukki" classifiers.  Final grinding
of the undersize material is done in a ball mill.

The magnetite and ilmenite fractions are magnetically  sepa-
rated,  with  the  magnetite  further upgraded by additional
magnetic processing.  The ilmenite sands are  then  upgraded
in  a  flotation  circuit  consisting  of roughers and three
stages of cleaners.  The ilmenite  concentrate  is  filtered
and dried prior to shipping.

Quantities  of  Wastes;   Mill 9905.   Wastes are discharged
from this mill at a rate of 35,191 cubic  meters  (9,220,000
gallons)   per  day.   The  results of a chemical analysis of
this wastewater are presented in Table V-83.  These data are
based on analysis of raw waste samples collected during  the
mill visit.

Reagents  consumed in the flotation circuit of mill 9905 are
identified in Table V-84.  The principal associated minerals
in the ore body of mine  9905  are  listed  in  Table  V-85.
These reagents and constituents of the ore body comprise the
principal constituents of the waste stream.

As  indicated in Table V-84, relatively high levels of iron,
titanium, zinc, nickel,  vanadium,  chromium,   and  selenium
were  observed  in the wastes of mill 9905.  Table V-86 is a
compilation  of  the   concentrations   of   the   principal
constituents of raw wastewater from mill 9905.
                           371

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Figure V-44. BENEFICIATION AND WASTE WATER FLOW OF ILMEIMITE
           MINE/MILL 9905 (ROCK DEPOSIT)
                                            MINING
                                             ORE
                                          CRUSHING
 878 n
(232,000 gpd)
                            36,069 m3/day
                            (9,450,000 gpd)
                                           GRINDING
                                             I
                                        CLASSIFICATION
                                             I
                                          MAGNETIC
                                         SEPARATION
                                  MAGNETICS
                                   NONMAGNETICS-i
                                                          ILMENITE
                                                        AND GANGUE
                                 FILTRATE
                             WATER
                         RETURN TO MILL

                         38,191 m3/day
                                                           I
                                                        FLOTATION
                                                         CIRCUIT
                                                        THICKENER
                                                           I
                                                          FILTER
                                               I
                                              DRIER
                                                       CONCENTRATE



                                                       TO SHIPPING
                          372

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     TABLE V-83. CHEMICAL COMPOSITION OF RAW WASTEWATER
               FROM TITANIUM MILL 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hard nest
COD
TOC
Oil and Grease
MBAS Surfactants
Total Kjaldahl N
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fa
CONCENTRATION (mg/£)
650'
18.0*
2.2
518
26,300
6.0
81.4
344.8
< 1.6
9.0
2.0
0.04
0.65
210
< 0.01
< 0.002
< 0.01
< 0.002
350
0.68
043
500
PARAMETER
Pb
MB
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfata
Nitrate
Phosphate
CONCENTRATION (mg/£)
< 0.05
1875
5.9
1.19
<0.1
2.0
23.7
0.132
0.015
41
0.29
< OHM
2.08
7.6
< 0.1
< 0.1
< 0.01
19.1
32.5
213
0.68
< 0.05
'Value in micromhos/cm

 Value in cobalt units
 TABLE V-84. REAGENT USE IN FLOTATION CIRCUIT OF MILL 9905

REAGENT

Tall oil
Fuel oil
Methyl amyl alcohol
Sodium bifluoride
Sulfuric acid

PURPOSE

Frother
Frother
Frother
Depressant
pH Modifier
CONSUMPTION
kg/metric ton
ore milled
1.33
0.90
0.008
0.76
1.775
Ib/short ton
ore milled
2.66
1.80
0.016
1.52
3.55
                         373

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         TABLE V-85. PRINCIPAL MINERALS ASSOCIATED WITH
                   ORE OF MINE 9905
MINERAL
llmenite
Magnetite
Pyroxene
Feldspar
COMPOSITION
FeTiOa
Fes 04
Complex Ferromagnesium Silicate
Aluminum Silicates with Calcium,
Sodium, and Potassium
TABLE V-86. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD AT MILL 9905
PARAMETER
TSS
TOC
Ni
Ti
Fe
V
Cr
Mn
Se
Cu
Zn
Fluoride
CONCENTRATION
(mg£) IN
WASTE WATER
26,300
9.0
1.19
2.08
500
2.0
0.58
5.9
0.132
0.43
7.6
32.5
RAW WASTE LOAD
par unit concentrate produced
kg/metric ton
462.8
0.158
0.021
0.036
8.8
0.035
0.010
0.103
0.0002
0.008
0.133
0.569
tb/short ton
925.8
0.316
0.042
0.072
17.6
0.070
0.020
0.206
0.0004
0.016
0.266
1.14
per unit ore milled
kg/metric ton
210.4
0.072
0.01
0.017
4.0
0.016
0.006
0.048
0.001
0.0003
.0.061
o.26
Ib/short ton
420.8
0.144
0.02
0.034
8.0
0.032
0.01
0.096
0.002
0.0006
0.122
0.52
                           374

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Figure V-45. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS (RUTILE, ILMENITE,
           ZIRCON, AND MONAZITE) AT MILL 9906
                                                                     TO
                                                                 ATMOSPHERE
    ORE + WATER
                  20,100 mj/day
                  (5,310,000 gpd)
           ORE FED
        FROM DREDGE
                                                                     t
                                                               EVAPORATION
                                    H20
                                    J_
       POND
      WATER
     RECYCLE
          VIBRATING
           SCREENS
                                   7,570 mj/day
                                  (2,000,000 gpd)
—OVERSIZE-
             WASH
            'WATER'
     SPIRALS OR LAMINAR
FLOWS (ROUGHERS AND CLEANERS)
           11,000 m3/day
           (3,170,000 gpd)
             I
                                 TO DRY MILL
                                (FIGURE 111-30)
                                                     -TAILINGS-i
       12,000 m3/day
       (3,170,000 gpd)
                                                           'WASTE (DREDGE)]
                                                                 POND
                                                       15,615 m3/day
                                                      (4,500,000 gpd)
                                                              DISCHARGE
                                    375

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Titanium

Dredging Operations;  Mill 9906 and 9907.   These operations
are  representative of the operations which recover titanium
minerals from beach-sand placers.  Operations 9906 and  9907
utilize a dredge, floating on a pond, to feed the sands to a
wet  mill  (Figure V-45).  The sands are beneficiated in the
wet mill by gravity methods, and  the  bulk  concentrate  is
sent to a dry mill for separation and upgrading of the heavy
minerals.   As  indicated  in Figure V-41, for mill 9906, no
discharge exists from the dry mill.  Water used in  the  wet
mill  is  discharged  to the dredge pond, which subsequently
discharges at a rate of 12,099 cubic  meters  (3.17  million
gallons)   per  day.   Raw  waste  characterization  of  the
combined wet-mill and dredge-pond discharge is presented  in
Table  V-87.    These data are based on analysis of raw waste
samples collected during the visits to these operations.

No reagents are used in the beneficiation of the  sands,  as
gravity  methods  are employed in the wet mill, and magnetic
and  electrostatic  methods  are  used  in  the  dry   mill.
Therefore,   the  principal  waste  constituents,  with  the
exception of waste lubrication oil from the dredge  and  wet
mill,  are  influenced primarily by the ore characteristics.
The ore bodies of operations 9906 and 9907  contain  organic
material which, upon disturbance, forms a colloidal slime of
high  coloring  capacity.   This organic colloid—primarily,
humates  and  tannic  acid—and  the  wasted  oil  are   the
principal  waste  constituents of the pond discharges.  This
is reflected in the high  carbon  oxygen  demand   (COD)  and
total  organic  carbon   (TOC)  values  detected in the waste
streams of operations 9906  and  9907  (Table  V-87).   High
levels  of  phosphate  and  organic  nitrogen are present in
these waste streams also.  The phosphate  and  nitrogen  are
undoubtedly  associated  with the sediments in the ore body.
Raw   waste   concentrations   of    principal    wastewater
constituents discharged from the milling operations at mills
9906 and 9907 are given in Table V-88.

Zirconium

Zirconium  is  recovered  as  a  byproduct of the mining and
milling of sand placer deposits, which have  been  described
under Waste Characteristics of Titanium Ores.  No operations
for  zirconium  alone  are  known in the United States.  The
waste characteristics and water uses accompanying mining and
milling  to  obtain  zircon  concentrate   are,   therefore,
identical to those of the previously described operations.
                            376

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      TABLE V-87. CHEMICAL COMPOSITION OF RAW WASTEWATER
                 AT MILLS 9906 AND 9907
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
COD
TOC
Total KJaldahl N
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
CONCENTRATION (mg/£)
MILL 9906
200*
51.400*
<0.1
1,644
11,000
47.2
47.6
1.338
972
0.65
400
<0.01
69.0
0.05
< 0.002
<0.5
0.10
< 0.002
0.10
0.03
<0.03
MILL 9907
40*
16.240*
0.54
370
209
31.4
3.4
362
321
0.65
40.0
<0.01
15.0
0.03
C0.002
<0.5
0.04
< 0.002
<0.05
<0.01
<0.03
PARAMETER
Total Fe
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Chloride
Fluoride
Phosphate
Phenol
CONCENTRATION (mg/£)
MILL 9906
4.9
<0.05
1.63
0.036
<0.01
<0.1
<0.5
3.5
<0.05
<0.01
27.0
<0.05
<0.06
<0.2
0.014
<0.1
<0.1
30.0
0.03
0.35
<0.01
MILL 9907
033

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TABLE V-88. RAW WASTE LOADS FOR PRINCIPAL WASTEWATER CONSTITUENTS
          FROM SAND PLACER MILLS 9906 AND 9907
PARAMETER
TSS
TOC
COO
Oil and Graau
Ti
Fa
Mn
Cr
Photphata
MILL 9906
CONCENTRATION
(me/ 8.)
IN WASTEWATER
11.000
972
1,337
400
<0.2
4.9
0.36
0.03
0.3S
RAW WASTE LOAD
(par unit total concantrata producad)
kg/matric ton
330
29.2
40.13
12
< 0.006
0.15
0.0011
0.0009
0.011
Ib/thort ton
660
58.4
80.26
24
< 0.012
0.30
0.0022
0.0018
0.022
MILL 9907
CONCENTRATION
(mg/£)
IN WASTEWATER
209
321
361.6
40
0.4
0.93
<0.01
<0.01
0.4
RAW WASTE LOAD
(par unit total concantrata producad)
kg/matric ton
5.01
7.71
8.68
0.96
0.01
0.022
< 0.0024
< 0.0024
0.01
Ib/ihort ton
10.02
1S.42
17.36
1.92
0.02
0.044
v 0.0048
< 0.0048
0.02
                             378

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                         SECTION VI

             SELECTION OF POLLUTANT PARAMETERS
INTRODUCTION

The  water-quality  investigation which preceded development
of recommended effluent guidelines covered a wide  range  of
potential  pollutants.   After considerable study, a list of
tentative control parameters was prepared for each  category
and  sutcategory  represented in this study.  The wastewater
constituents finally selected as being of pollution signifi-
cance for the ore mining and  dressing  industry  are  based
upon  (1)  those  parameters  which  have been identified as
known  constituents  of   the   ore-bearing   deposits   and
overburden,  (2)  chemicals used in processing or extracting
the desired metal (s), and (3)  parameters  which  have  been
identified  as  present  in  significant  quantities  in the
untreated wastewater from each subcategory  of  this  study.
The  wastewater  constituents  are  further divided into (a)
those that have been selected as pollutants of  significance
(with the rationale for their selection), and (b)  those that
are  not  deemed  significant   (with the rationale for their
rejection).  This section is concluded with a  summary  list
of the pollution parameters selected for each category.

GUIDELINE PARAMETER-SELECTION CRITERIA

selection  of  parameters  for  use  in  developing effluent
limitation guidelines was based primarily on  the  following
criteria:

    (1)   Constituents which are frequently present  in  mine
         and  mill  discharges in concentrations deleterious
         to  human,  animal,  fish,  and  aquatic  organisms
          (either directly or indirectly) .

    (2)   The existence of technology for  the  reduction  or
         removal, at an economically achievable cost, of the
         pollutants in question.

    (3)   Research    data    indicating    that    excessive
         concentrations  may  be  capable  of  disrupting an
         aquatic ecosystem.

    (4)   substances which result in sludge deposits, produce
         unsightly  conditions  in  streams,  or  result  in
         undesirable tastes and odors in water supplies.
                            379

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SIGNIFICANCE   AND  RATIONALE  FOR  SELECTION  OF  POLLUTION
PARAMETERS

pH. Acidity, and Alkalinity

Acidity and alkalinity are  reciprocal  terms.   Acidity  is
produced  by substances that yield hydrogen ions upon hydro-
lysis, and alkalinity is produced by substances  that  yield
hydroxyl  ions.   The terms "total acidity" and "total alka-
linity" are often used to express the buffering capacity  of
a   solution.   Acidity in natural waters is caused by carbon
dioxide, mineral acids, weakly dissociated  acids,  and  the
salts  of strong acids and weak bases.  Alkalinity is caused
by  strong bases and the salts of strong  alkalies  and  weak
acids.

The term pH is a logarithmic expression of the concentration
of  hydrogen  ions.  At a pH of 7, the hydrogen and hydroxyl
ion concentrations are essentially equal, and the  water  is
neutral.   Lower  pH  values  indicate acidity, while higher
values indicate alkalinity.  The relationship between pH and
acidity or alkalinity is not necessarily linear or direct.

Waters with a pH below 6.0  are  corrosive  to  water  works
structures,   distribution  lines,  and  household  plumbing
fixtures and can thus  add  such  constituents  to  drinking
water  as  iron,  copper,  zinc,  cadmium,  and  lead.   The
hydrogen ion concentration can affect  the  "taste"  of  the
water.   At a low pH, water tastes "sour."  The bactericidal
effect of chlorine is weakened as the pH increases,  and  it
is  advantageous  to  keep  the pH close to 7.  This is very
significant for providing safe drinking water.

Extremes of pH or rapid pH changes can exert  stress  condi-
tions  or kill aquatic life outright.  Dead fish, associated
algal blooms, and foul strenches are  aesthetic  liabilities
of  any  waterway.   Even moderate changes from "acceptable"
criteria limits of pH are deleterious to some species.   The
relative  toxicity  to  aquatic  life  of  many materials is
increased  by  changes  in  the  water  pH.    Metalocyanide
complexes  can  increase  a thousand-fold in toxicity with a
drop of 1.5 pH units.  The  availability  of  many  nutrient
substances  varies with the alkalinity and acidity.  Ammonia
is more lethal with a higher pH.

The lacrimal fluid of the human eye has  a  pH  of  approxi-
mately 7.0, and a deviation of 0.1 pH unit from the norm may
result  in  eye  irritation  for  the  swimmer.  Appreciable
irritation will cause severe pain.
                            380

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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 load-
ing and adversely affect effluent receiving waters.

Total Suspended solids

Suspended   solids   include   both  organic  and  inorganic
materials.  The inorganic compounds include sand, silt,  and
clay.   The  organic  fraction  includes  such  materials as
grease, oil, tar, animal and vegetable fats, various fibers,
sawdust, hair, and various  materials  from  sewers.   These
solids may settle out rapidly, and bottom deposits are often
a  mixture  of  both  organic  and  inorganic  solids.  They
adversely affect fisheries by covering  the  bottom  of  the
stream  or lake with a blanket of material that destroys the
fish-food bottom fauna  or  the  spawning  ground  of  fish.
Deposits  containing  organic  materials  may deplete bottom
oxygen  supplies  and  produce  hydrogen   sulfide,   carbon
dioxide, methane, and other noxious gases.

In  raw  water  sources for domestic use, state and regional
agencies generally specify that suspended solids in  streams
shall  not  be  present  in  sufficient  concentration to be
objectionable  or  to  interfere   with   normal   treatment
processes.   Suspended  solids  in  water may interfere with
many industrial processes and cause foaming  in  boilers  or
encrustation  on  equipment  exposed to water, especially as
the temperature rises.  Suspended solids are undesirable  in
water  for  textile  industries;  paper and pulp; beverages;
dairy  products;  laundries;  dyeing;  photography;  cooling
systems;  and  power plants.  Suspended particles also serve
as a transport mechanism for pesticides onto clay particles.

Solids may be suspended in water for a time and then  settle
to  the  bed of the stream or lake.  These settleable solids
discharged with manfs wastes may be inert, slowly biodegrad-
able materials, or rapidly decomposable  substances.   While
in  suspension,  they  increase  the turbidity of the water,
reduce light  penetration,  and  impair  the  photosynthetic
activity of aquatic plants.

Solids  in  suspension  are aesthetically displeasing,  when
they settle to form sludge deposits on the  stream  or  lake
bed, they are often much more damaging to the life in water,
and  they  retain  the  capacity  to  displease  the senses.
Solids, when  transformed  to  sludge  deposits,  may  do  a
variety  of damaging things, including blanketing the stream
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or lake bed and thereby destroying  the  living  spaces  for
those  benthic  organisms  that  would  otherwise occupy the
habitat.  When of an organic (and, therefore,  decomposable)
nature,  solids use a portion or all of the dissolved oxygen
available in the area.  Organic materials also  serve  as  a
seemingly  inexhaustible  food  source  for  sludgeworms and
associated organisms.

Turbidity is principally a measure of  the  light-scattering
and  light-absorbing  properties of suspended solids.  It is
frequently used as a substitute method of quickly estimating
the  total  suspended  solids  when  the  concentration   is
relatively low.

High  suspended-solid concentrations are contributed as part
of the mining process, as well as  the  crushing,  grinding,
and  other  processes  commonly used in the milling industry
for most milling operations.  High  suspended-solid  concen-
trations   are  also  characteristic  of  dredge-mining  and
gravity separation operations.

Oil and Grease

Oil and grease exhibit an oxygen demand.  Oil emulsions  may
adhere  to  the  gills  of fish or coat and destroy algae or
other plankton.  Deposition of oil in the  bottom  sediments
can serve to exhibit normal benthic growths, thus interrupt-
ing the aquatic food chain.  Soluble and emulsified material
ingested  by  fish  may  taint the flavor of the fish flesh.
Water-soluble components may exert  toxic  action  on  fish.
Floating oil may reduce the re-aeration of the water surface
and,  in conjunction with emulsified oil, may interfere with
photosynthesis.   Water-insoluble  cc ponents   damage   the
plumage  and  coats  of  water  animals  and fowls.   Oil and
grease in water can result in the formation of objectionable
surface slicks, preventing the full aesthetic  enjoyment  of
the  water.    Oil spills can damage the surface of boats and
can destroy the aesthetic  characteristics  of  beaches  and
shorelines.

Levels  of  oil  and  grease  which  are  toxic  to  aquatic
organisms vary  greatly,  depending  on  the  type  and  the
species  susceptibility.  However, it has been reported that
crude oil in concentrations as low as 0.3 mg/1 is  extremely
toxic  to fresh-water fish.  There is evidence that oils may
persist and have subtle chronic effects.

This parameter is found in discharges of the ore mining  and
dressing  industry  as  a  result  of  the contribution from
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 lubricants and spillage of fuels,  as well as  the  usage   of
 reagents in many milling processes.

 Chemical Oxygen Demand (COD)  and Total  Organic Carbon  (TOG)

 The  chemical  oxygen  demand (COD)  determination provides a
 measure  of the oxygen equivalent  of that  portion of  the
 organic   matter in  a sample that is  susceptible to oxidation
 by  a  strong chemical oxidant.  With  certain  wastes contain-
 ing  toxic  substances,  this  test—or a total  organic  carbon
 determination—may  be the  only method  for  obtaining  the
 organic  load.

 Chemical oxygen demand will result in depletion of dissolved
 oxygen   in  receiving  waters.   Dissolved  oxygen (DO) is a
 water-quality    constituent     that,      in     appropriate
 concentrations,  is  essential,  not only to  keep organisms
 living,  but also to sustain species  reproduction,  vigor, and
 the development of  populations.  Organisms undergo stress  at
 reduced  DO concentrations that  makes them less  competitive
 and  able  to  sustain  their  species   within  the aquatic
 environment.   For example,  reduced  DO   concentrations  have
 been  shown  to  interfere with  fish   populations through
 delayed  hatching of eggs,  reduced  size  and vigor of embryos,
 production of deformities  in  young,  interference  with  food
 digestion,    acceleration  of   blood   clotting,   decreased
 tolerance to  certain toxicants,  reduced food efficiency  and
 growth   rate,   and  reduced maximum sustained swimming speed.
 Fish  food  organisms  are   likewise   affected   adversely   in
 conditions with suppressed  DO.  Since  all aerobic aquatic
 organisms need a certain amount  of oxygen, the  total lack  of
 dissolved oxygen due to  a  high COD can  kill  all  inhabitants
 of  the affected area.

 The   total  organic carbon  (TOC) value  generally falls below
 the true  concentration of organic  contaminants  because other
 constituent elements are excluded.   When an empirical  rela-
 tionship   can   be   established   between  the  total  organic
 carbon,  the biochemical  oxygen  demand,   and   the  chemical
 oxygen   demand,  the TOC provides  a  rapid, convenient method
 of estimating  the other  parameters that express  the  degree
 of  organic contamination.  Forms  of carbon analyzed by this
 test,  among   others,  are:   soluble,  nonvolatile  organic
 carbon;   insoluble,   partially  volatile carbon(e.g.,  oils);
 and insoluble,   particulate  carbonaceous  materials  (e.g.,
 cellulose fibers).

The  final  usefulness   of  the two methods is to assess the
oxygen-demanding load of organic  material  on  a  receiving
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stream.   The widespread use of oil-based compounds, organic
acids, or other organic coumpounds in the flotation process,
as well as the absence of accurate, reproducible tests which
can be routinely performed, points to the use of these tests
as indicators of the levels  of  particular  reagent  groups
which are being discharged.

COD  reflects  the  presence of a variety of materials which
may be present in the effluent from ore dressing operations.
Many flotation reagents exert a chemical oxygen demand,  and
the  presence  of excessive levels of these materials in the
effluent stream will be reflected in  elevated  COD  values.
Higher  COD  values  are  generally  observed  for flotation
effluent streams than  for  those  where  flotation  is  not
practiced.   In  addition,  elevated  COD values reflect the
release  of  significant  quantities  of   chemicals   whose
environmental fates and effects are largely unknown.

Cyanide

Cyanides  in  water  derive  their  toxicity  primarily from
undissociated hydrogen cyanide (HCN), rather than  from  the
cyanide ion  (CN-).  HCN dissociates in water into H+ and CN-
in  a  pH-dependent  reaction.   At a pH of 7 or below, less
than 1 percent of the cyanide is present as CN-; at a pH  of
8, 6.7 percent; at a pH of 9, 42 percent; and at a pH of 10,
87  percent  of the cyanide is dissociated.  The toxicity of
cyanides is also increased by increases in  temperature  and
reductions  in  oxygen  tensions.   A temperature rise of 10
degrees Celsius  (14 degrees Fahrenheit)  produces a  two-  to
three-fold  increase  in  the  rate  of the lethal action of
cyanide.

Cyanide has been  shown  to  be  poisonous  to  humans,  and
amounts over 18 ppm can have adverse effects.  A single dose
of about 50 to 60 mg is reported to be fatal.
Trout and other aquatic organisms are extremely sensitive to
cyanide.   Amounts as small as 0.1 part per million can kill
them.  Certain metals, such  as  nickel,  may  complex  with
cyanide   to  reduce  lethality—especially,  at  higher  pH
values-but  zinc   and   cadmium   cyanide   complexes   are
exceedingly toxic.

When  fish are poisoned by cyanide, the gills become consid-
erably brighter in color than those of normal fish, owing to
the inhibition by cyanide of  the  oxidase  responsible  for
oxygen transfer from the blood to the tissues.
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The  presence  of cyanide in the effluents of the mining and
milling industry is primarily due to the use of cyanide as a
depressant in flotation processes and as a leaching reagent-
particularly, in the gold and silver ore milling categories.

Ammonia

Ammonia is a common product of the decomposition of  organic
matter.   Dead  and  decaying animals and plants, along with
human and animal  body  wastes,  account  for  much  of  the
ammonia  entering  the aquatic ecosystem.  Ammonia exists in
its nonionized form only at higher pH levels and is the most
toxic in this state.  The lower the  pH,  the  more  ionized
ammonia  is formed, and its toxicity decreases.  Ammonia, in
the presence of dissolved oxygen, is  converted  to  nitrate
(NO3)   by  nitrifying  bacteria.  Nitrite (NO2J , which is an
intermediate product between ammonia and nitrate,  sometimes
occurs  in quantity when depressed oxygen conditions permit.
Ammonia can exist in several  other  chemical  combinations,
including ammonium chloride and other salts.

Nitrates   are   considered   to   be  among  the  poisonous
ingredients of mineralized waters,  with  potassium  nitrate
being  more  poisonous than sodium nitrate.   Excess nitrates
cause   irritation   of   the   mucous   linings   of    the
gastrointestinal  tract  and  the  bladder;  the symptoms are
diarrhea and diuresis, and drinking one liter  (1.06  quart)
of  water  containing  500  mg/1  of  nitrate can cause such
symptoms.

Infant methemoglobinemia, a disease characterized by certain
specific blood changes, and cyanosis may be caused  by  high
nitrate concentrations in the water used for preparing feed-
ing formulae,  while it is still impossible to state precise
concentration  limits,  it  has been widely recommended that
water containing more than 10 mg/1 of nitrate nitrogen (N
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all aquatic life within a range of less than 1.0 mg/1 to  25
mg/1,  depending  on  the  pH and the dissolved oxygen level
present.  Ammonia can add to the problem  of  eutrophication
by  supplying nitrogen through its breakdown products.  Some
lakes in warmer climates, and others that are aging quickly,
are  sometimes  limited  by  the  nitrogen  available.   Any
increase  will  speed  up  the  plant  growth  and the decay
process.

No limits on ammonia are set by  the  United  States  Public
Health  Service  (USPHS) Drinking Water Standards of 1962 or
the World Health Organizaton (WHO) International  Standards,
but   the  WHO  European  Drinking  Water  Standards  set  a
recommended limit of 0.5 mg/1 as NH4+.

The odor threshold for ammonia has been  reported  as  0.037
mg/1;  for  the brewing of coffee, the taste threshold is 3U
mg/1.

Ammonia concentrations in the range of 0.3 to 21.4 mg/1 have
been reported to be acutely  toxic  to  various  species  of
fish.   An  indicated  mode  of  toxicity  is  the decreased
ability of hemoglobin to combine with oxygen in the presence
of ammonia and, hence, cause possible suffocation.   Ammonia
concentrations  as  low  as  0.3  mg/1 ahve been observed to
cause a noticable drop in the oxygen content of the blood of
fishes.

Algae, which thrive in high nitrate  concentrations,  appear
to  be  harmed or inhibited when the nitrogen is in the form
of ammonia.  A concentration of 0.5 mg/1 of ammonia nitrogen
has been observed  to  cause  a  complete  disappearance  of
A pha ni zomenon.  The  lethal  concentration  of  ammonia  for
Daphnia has been reported at 8 mg/1.

Solutions containing mixtures of ammonium and  cyanide  ions
are  more toxic to fish than solutions containing either ion
alone.

In leaching operations, ammonia  may  be  used  in  leaching
solutions   (as   in  the  •Dean-Leute'  ammonium  carbamate
process), for  precipitation  of  metal  salts,  or  for  pH
control.   In  the  ore  mining  and dressing industry, high
levels at selected locations may thus be encountered.

Aluminum

Aluminum is one of the most abundant elements on the face of
the earth.  It occurs in many rocks and ores, but never as a
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pure metal.   Although  some  aluminum  salts  are  soluble,
aluminum  is  not likely to occur for long in surface waters
because it precipitates and settles or is absorbed as  alum-
inum  hydroxide,  carbonate, etc.  The mean concentration of
soluble aluminum is approximately 74 micrograms  per  liter,
with values ranging from 1 to 2,760 micrograms per liter.

Aluminum  can  be  found  in  all  soils, plants, and animal
tissues.  The human body contains about  50  to  150  mg  of
aluminum,   and   aluminum   concentrations  in  fruits  and
vegetables range up to 37 mg/kg.  The total aluminum in  the
human  diet has been estimated at 10 to 100 mg/day; however,
very little of the aluminum is absorbed  by  the  alimentary
canal.  Aluminum is not considered a problem in public water
supplies.   Note,  however,  that  excessively high doses of
aluminum may interfere with phosphorus metabolism.  Aluminum
present in surface waters can be harmful to  aquatic  life—
particularly, marine aquatic life.  Marine organisms tend to
concentrate  aluminum  by  a factor of approximately 10,000.
Administration of 0.10 mg/1 of aluminum nitrate for  1  week
proved  lethal  to  sticklebacks.   Approximately  5 mg/1 of
aluminum is lethal to trout when exposed for 5 minutes,  but
the  presence  of  only  1  mg/1  over  the same time period
produces no harmful effects.

Aluminum is generally  a  minor  constituent  of  irrigation
waters.  In addition, most soils are naturally alkaline and,
as  such, are not subject to the toxic effects of relatively
high concentrations of  aluminum.   Where  soils  are  quite
acidic  (pH  below 5.0), aluminum toxicity to plants becomes
very significant.  Aluminum presence is  primarily  observed
in wastewaters from the bauxite-ore mining industry.

Antimony

Antimony  is  rarely  found pure in nature, its common forms
being  the  sulfide,  stibnite  (Sb2S3)    and   the   oxides
cervantite  (Sb20f*)   and  valentinite (Sb^OSJ.  Any antimony
discharged to  natural  waters  has  a  strong  tendency  to
precipitate   and   be   removed   by  sedimentation  and/or
adsorption.

Antimony compounds are toxic to man and  are  classified  as
acutely  moderate  or chronically severe.  A dose of 97.2 mg
of  antimony  has  reportedly  been  lethal  to  an   adult.
Antimony  potassium tartrate, once in use medically to treat
certain parasitic diseases, is no longer recommended because
of the frequency and severity of toxic reactions,  including
cardiac disturbances.
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Various  marine organisms reportedly concentrate antimony to
more than 300 times the amount present  in  the  surrounding
waters.   Few  of  the salts of antimony have been tested in
bioassays; as a result, data on antimony toxicity to aquatic
organisms  are  sketchy.    Antimony   is   commonly   found
associated  with  sulfide  ores  exploited in the silver and
lead  industry,  as  well  as  in  operations  operated  for
antimony primary or byproduct recovery.

Arsenic

Arsenic  is  found  to  a  small  extent  in  nature  in the
elemental form.  It occurs mostly in the form  of  arsenites
of metals or as arsenopyrite (FeS2.
Arsenic  is  normally present in sea water at concentrations
of 2 to 3 micrograms per liter and tends to  be  accumulated
by oysters and other shellfish.  Concentrations of 100 mg/kg
have been reported in certain shellfish.  Arsenic is a cumu-
lative poison with long-term chronic effects on both aquatic
organisms  and  mammalian species, and a succession of small
doses may add up to a final lethal dose.  It  is  moderately
toxic  to plants and highly toxic to animals — especially, as
arsine (AsH3_) .

Arsenic trioxide,  which  also  is  exceedingly  toxic,  was
studied in concentrations of 1.96 to UO mg/1 and found to be
harmful  in that range to fish and other aquatic life.  Work
by the Washington Department of Fisheries on pink salmon has
shown that a level of 5.3  mg/1  of  As20;3  for  8  days  is
extremely harmful to this species; on mussels, a level of 16
mg/1 is lethal in 3 to 16 days.

Severe    human    poisoning    can   result   from   100-mg
concentrations, and 130 mg has proved  fatal.   Arsenic  can
accumulate  in  the  body faster than it is excreted and can
build to toxic levels, from small amounts taken periodically
through lung and intestinal walls from the air,  water,  and
food.   Arsenic  is a normal constituent of most soils, with
concentrations ranging up to 500 mg/kg.  Although  very  low
concentrations  of  arsenates  may  actually stimulate plant
growth,  the  presence  of  excessive  soluble  arsenic   in
irrigation  waters  will reduce the yield of crops, the main
effect appearing to be the destruction of chlorophyll in the
foliage.  Plants grown  in  water  containing  one  mg/1  of
arsenic  trioxides show a blackening of the vascular bundles
in the leaves.  Beans  and  cucumbers  are  very  sensitive,
while   turnips,   cereals,   and   grasses  are  relatively
resistant.  Old orchard soils in Washington that  contain  4
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to 12 mg/kg of arsenic trioxide in the topsoil were found to
have become unproductive.

Arsenic  is known to be present in many complex metal ores—
particularly, the sulfide ores of cobalt, nickel  and  other
ferroalloy ores, antimony, lead, and silver.  It may also be
solubilized  in  mining  and milling by oxidation of the ore
and appear in the effluent stream.

Beryllium

Beryllium is a relatively rare element, found chiefly in the
mineral beryl.  In the weathering process, beryllium is con-
centrated in hydrolyzate and, like  aluminum,  does  not  go
into  solution  to any appreciable degree.  Beryllium is not
likely to be found in natural waters in greater  than  trace
amounts  because of the relatively insolubility of the oxide
and hydroxide at the normal pH range of such waters.

Absorption of beryllium from the alimentary tract is slight,
and  excretion  is  fairly  rapid.   However,  as   an   air
pollutant,  it  is  responsible  for  causing  skin and lung
diseases of variable severity.

concentrations of beryllium sulfate  complexed  with  sodium
tartrate up to 28.5 mg/1 are not toxic to goldfish, minnows,
or  snails.   The  96-hour  minimum toxic level of beryllium
sulfate for fathead minnows has been found to be 0.2 mg/1 in
soft water and 11 mg/1 in  hard  water.   The  corresponding
level  for beryllium chloride is 0.15 mg/1 in soft water and
15 mg/1 in hard water.

In nutrient solution, at acid pH values, beryllium is highly
toxic to plants.  Solutions containing  15  to  20  mg/1  of
beryllium  delay  germination and retard the growth of cress
and mustard seeds in  solution  culture.   The  presence  of
beryllium  in  wastewaters  was  detected  only in raw-waste
effluents from the mining and milling of bertrandite.

Cadmium

Cadmium in drinking water supplies is extremely hazardous to
humans, and conventional  treatment,   as  practiced  in  the
United states, does not remove it.  Cadmium is cumulative in
the liver, kidney, pancreas,  and thyroid of humans and other
animals.   A  severe  bone  and kidney syndrome in Japan has
been associated with the  ingestion  of  as  little  as  600
micrograms per day of cadmium.
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Cadmium  is  an  extremely  dangerous  cumulative  toxicant,
causing insidious progressive chronic poisoning in  mammals,
fish,  and  (probably) other animals because the metal is not
excreted.  Cadmium can form organic compounds which may lead
to mutagenic or teratogenic effects.  Cadmium  is  known  to
have  marked  acute and chronic effects on aquatic organisms
also.

Cadmium acts synergistically with other metals.  Copper  and
zinc   substantially  increase  its  toxicity.   Cadmium  is
concentrated by  marine  organisms--particularly,  mollusks,
which  accumulate  cadmium  in calcareous tissues and in the
viscera.  A concentration factor of 1000 for cadmium in fish
muscle has been reported, as have concentration  factors  of
3,000  in  marine plants, and up to 29,600 in certain marine
animals.  The eggs and larvae of fish are, apparently,  more
sensitive  than  adult  fish  to  poisoning  by cadmium, and
crustaceans appear to be more sensitive than fish  eggs  and
larvae.

Cadmium,  in  general,  is  less toxic in hard water than in
soft water.  Even so, the safe levels of cadmium for fathead
minnows and bluegills in hard water have been  found  to  be
between  0.06 and 0.03 mg/1, and safe levels for coho salmon
fry have been reported to be 0.004 to  0.001  mg/1  in  soft
water.   Concentrations  of  0.0005  mg/1  were  observed to
reduce  reproduction  of  Daphnia  magna  in  one-generation
exposure lasting three weeks.

Cadmium  is  present  in minor amounts in the effluents from
several  ferroalloy-ore  and  copper  mining   and   milling
operations.  It is a common constituent in all zinc ores and
can  be  expected to be present in most lead-zinc operations
especially those where metals are solubilized.

Chromium
Chromium, in its various valence  states,  is  hazardous  to
man.   It  can  produce lung tumors when inhaled and induces
skin  sensitizations.   Large  doses   of   chromates   have
corrosive  effects  on  the  intestinal  tract and can cause
inflammation of the kidneys.  Levels of chromate  ions  that
have  no  effect  on  man appear to be so low as to prohibit
determination to date.

The toxicity of chromium salts toward  aquatic  life  varies
widely  with  the  species,  temperature, pH, valence of the
chromium,  and   synergistic   or   antagonistic   effects—
especially,  that of hardness.  Fish are relatively tolerant
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of chromium salts, but fish-food organisms and  other  lower
forms  of  aquatic  life  are extremely sensitive.  Chromium
also inhibits the growth of algae.

In some  agricultural  crops,  chromium  can  cause  reduced
growth  or  death  of  the  crop.   Adverse  effects  of low
concentrations of chromium on corn, tobacco, and sugar beets
have been documented.

Chromium is present at  appreciable  concentrations  in  the
effluent from mills practicing leaching.  It is also present
as  a minor constituent in many ores, such as those of plat-
inum, ferroalloy metals, lead, and zinc.

Copper

copper salts occur in natural surface waters only  in  trace
amounts,  up to about 0.05 mg/1, so their presence generally
is the result of pollution.  This  is  attributable  to  the
corrosive action of the water on copper and brass tubing, to
industrial  effluents, and—frequently—to the use of copper
compounds for the control of undesirable plankton organisms.

Copper is not considered to be a cumulative systemic  poison
for  humans,  but  it can cause symptoms of gastroenteritis,
with nausea and intestinal irritations,  at  relatively  low
dosages.   The limiting factor in domestic water supplies is
taste.   Threshold  concentrations  for  taste   have   been
generally  reported  in  the  range  of  1.0  to 2.0 mg/1 of
copper, while as much as 5  to  7.5  mg/1  makes  the  water
completely unpalatable.

The   toxicity   of   copper  to  aquatic  organisms  varies
significantly, not only with the species, but also with  the
physical   and   chemical   characteristics  of  the  water,
including  temperature,  hardness,  turbidity,  and   carbon
dioxide  content.   In  hard  water,  the toxicity of copper
salts is reduced by the precipitation of copper carbonate or
other insoluble compounds.  The sulfates of copper and zinc,
and of copper and cadmium, are synergistic  in  their  toxic
effect on fish.

Copper concentrations less than 1 mg/1 have been reported to
be  toxic—particularly,  in  soft  water—to  many kinds of
fish, crustaceans,  mollusks,  insects,  phytopiankton,  and
zooplankton.   Concentrations  of  copper,  for example, are
detrimental to some oysters above 0.1 ppm.  Oysters cultured
in  sea water containing 0.13 to 0.5 ppm of  copper  deposit
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the  metal  in  their  bodies  and  become  unfit  as a food
substance.

Besides,  those  used  by  the  copper  mining  and  milling
industry,  many  other  ore  minerals  in the ore mining and
dressing industry contain  byproduct  or  minor  amounts  of
copper;  therefore,  the waste streams from these operations
contain copper.

Fluorides

As the most reactive non-metal, fluorine is never found free
in nature, but rather occurs as a constituent of fluorite or
fluorspar (calcium fluoride) in sedimentary rocks  and  also
as  cryolite   (sodium  aluminum  fluoride) in igneous rocks.
Owing to their origin only in certain  types  of  rocks  and
only  in a few regions, fluorides in high concentrations are
not a common constituent of natural surface waters, but they
may occur in detrimental concentrations in ground waters.

Fluorides are used as insecticides, for disinfecting brewery
apparatus, as a  flux  in  the  manufacture  of  steel,  for
preserving  wood and mucilages, for the manufacture of glass
and enamels, in chemical industries,  for  water  treatment,
and for other uses.

Fluorides  in  sufficient quantity are toxic to humans, with
doses of 250 to 450 mg giving  severe  symptoms  or  causing
death.

There  are  numerous  articles  describing  the  effects  of
fluoride-bearing waters on dental enamel of children;  these
studies  lead  to  the  generalization that water containing
less than 0.9 to 1.0 mg/1  of  fluoride  will  seldom  cause
mottled  enamel in children; for adults, concentrations less
than 3 or H mg/1 are not likely to cause endemic  "cumulative
fluorosis and skeletal effects.  Abundant literature is also
available  describing  the  advantages of maintaining 0.8 to
1.5 mg/1 of fluoride ion in drinking water  to  aid  in  the
reduction of dental decay—especially, among children.

chronic fluoride poisoning of livestock has been observed in
areas   where   water  contains  10  to  15  mg/1  fluoride.
Concentrations of 30 to 50 mg/1 of  fluoride  in  the  total
ration  of  dairy  cows are considered the upper safe limit.
Fluoride from waters, apparently,  does  not  accumulate  in
soft  tissue  to a significant degree, and it is transferred
to a very small extent into milk and, to a somewhat  greater
degree,  into  eggs.   Data  for  fresh  water indicate that
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fluorides are toxic to fish at  concentrations  higher  than
1.5 mg/1.

Hiah  fluoride levels in the effluents from mines may result
from high levels  in  intercepted  aquifers  or  from  water
contact from rock dust and fragments.  The use of mine water
in  milling,  as  well  as  extended  contact  of water with
crushed and ground ore, may yield high  fluoride  levels  in
mill  effluents.   Levels  may  also be elevated by chemical
action in leaching operations.

Iron

Iron is one of the most abundant constituents of  rocks  and
soils  and,  as  such,  is  often  found  in natural waters.
Although many of the ferric and ferrous salts, such  as  the
chlorides,  are  highly  soluble  in water, ferrous ions are
readily oxidized in  natural  surface  waters  to  insoluble
ferric  hydroxides.  These precipitates tend to agglomerate,
flocculate, and settle or be absorbed  in  surfaces;  hence,
the  concentration  of iron in well-aerated waters is seldom
high.  Mean concentrations of iron in O.S. waters range from
19 to 173 micrograms  per  liter,  depending  on  geographic
location.   When the pH is low, however, appreciable amounts
of iron may remain in solution.

Standards for drinking water are not set for health reasons.
Indeed, some iron is essential  for  nutrition,  and  larger
quantities  of  iron are taken for therapeutic reasons.  The
drinking-water standards are set for esthetic reasons.

In general, very little iron remains in  solution;  but,  if
the  water  is  strongly buffered and a large enough dose is
supplied, the addition of a soluble iron salt may lower  the
pH  of  the  water  to a toxic level.  In addition, a fish's
respiratory channel may  become  irritated  and  blocked  by
depositions of iron hydroxides on the gills.  Finally, heavy
precipitates of ferric hydroxide may smother fish eggs.

The  threshold  concentration for lethality to several types
of  fish  has  been  reported   as   0.2   mg/1   of   iron.
Concentrations of 1 to 2 mg/1 of iron are indicative of acid
pollution  and  other  conditions  unfavorable to fish.  The
upper limit for fish life has been estimated at 50 mg/1.  At
concentrations of iron above  0.2  mg/1,  trouble  has  been
experienced   with   populations   of   the  iron  bacterium
Crenothrix.
                            393

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 Iron is very common in natural waters and  is  derived  from
 common  iron  minerals in the substrata.  The iron may occur
 SJ£2«?0rm?S* susPend*d and dissolved.  The iron mining and
 processing industry inherently increases iron levels present
 ^o^Ce!S-Kr Nine, "aters.  The aluminum-ore mining industry
 also contributes elevated iron levels through mine drainage.,

 Lead

 Lead sulfide and lead oxide are the primary  forms  of  lead
 found  m  rocks.   Certain lead salts, such as the chloride
 and the acetate, are  highly  soluble;  however,   since  the
 oa^°^eH*and  hydr°xide  are  insoluble and the sulfide is
 only slightly soluble, lead  is  not  likely  to   remain  in
 solution   long  in  natural  waters.   m  the  U.S.,  lead
 concentrations  in  surface  and  ground  waters   used   for
 nrovfi^ suPPlies Average 0.01 mg/1.  some natural waters in
 f J n ? H no  mo.untain limestone and galena contain as much
 as u.q to 0.8 mg/1 of lead in solution.
 Lead is highly toxic to human beings  and  is  a  cumulative
 ™io?n;a*-TyPi?al  symPtoms  of  advanced lead poisoning are
 constipation,  loss of appetite, anemia,  abdominal  pain,   and
 ?Sn£i  Paraiysis  «  the muscles.   Lead poisoning  usually
 results from the cumulative  toxic  effects  of lead after
                                         °f  time'  rather
 of   orf      «                 '   The  level at wnich the a"«^nt
 of  bodily lead  intake exceeds the   amount  excreted  by  the
 body is   approximately  0.3 mg/day.   A  total intake of lead
 appreciably in  excess  of  0.6   mg/day  may result  in  the
 accumulation  of   a   dangerous   quantity of lead during a
 11 retime.                                                ^

 The  toxic  concentration of  lead for  aerobic  bacteria  is
 reported   to be 1.0 mg/1; for flagellates and infusoria, 0.5
 mg/1.   Inhibition  of  bacterial decomposition   of  organic
 matter  occurs  at  lead  concentrations  of  0.1 to 0.5 Sg/iT
 Toxic effects of lead on  fish include  the   formation  of  a
              ^Su film °Ver the 9ills~ and, eventually, the
              hl°h  SaUSS!  the  fish  to suffocate.    Lead
              very  dependent on water hardness; in general,
tha     much I*38 toxl5 in hard water.  Some  data  indicate
that  the median period of survival of rainbow trout in soft
water containing dissolved lead is 18 to  2H  hours  at  1 6
J2Xi^ **e K96"hour  ^i™™ toxic level  for fathead minnows
*«*  ?? ha^ been reported as 2.4 mg/1 of lead in soft  water
and  75 mg/1 in hard water.   Toxic levels for fish can range

S2o?^t0 ?5 mg/1 °f I6ad' dePendin9  °n  water  hardnesl,
dissolved  oxygen  concentration,  and  the type of organism
                            394

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studied.   Sticklebacks and minnows  have  not  been  visibly
harmed  when  in   contact  with 0.7 mg/1 of lead in soft tap
water for  3 weeks.  However, the H8-hour minimum toxic level
for  sticklebacks in water containing 1,000 to 3,000 mg/1  of
dissolved  solids  is reported to be 0.3U mg/1 of lead.  The
U.S. Public Health Service Drinking Water Standard specifies
a rejection limit  of 0.05 ppm  (mg/1) for lead.
              i
Elevated concentrations of lead are discharged from lead and
zinc mines and mills, as well as  from  mining  and  milling
operations    exploiting   other   sulfide   ores,   such  as
tetrahedrite  (for  silver and lead); copper ores;  ferroalloy
ore minerals; or mixed copper, lead, and zinc ores.

Manganese

Pure  manganese  metal  is not naturally found in the earth,
but its ores  are  very  common.   Similar  to  iron  in  its
chemical  behavior,  it occurs in the bivalent and trivalent
forms.  The   nitrates,  sulfates,  and  chlorides  are  very
soluble in water, but the oxides, carbonates, and hydroxides
are only sparingly soluble.  The background concentration of
manganese  in most natural waters is less than 20 micrograms
per liter.

Manganese is  essential for the nutrition of both plants  and
animals.   The  toxicological  significance  of manganese to
mammals is considered to be of little consequence,  although
some  cases of manganese poisoning have been reported due to
unusually  high  concentrations.    Manganese   limits   for
drinking  water  have  been  set for esthetic reasons rather
than physiological hazards.

As with most  elements, toxicity to aquatic life is dependent
on a variety  of factors.  The lethal concentration  of  man-
ganese  for   the stickleback has been given at HO mg/1.  The
threshold toxic concentration of manganese for the  flatworm
Polvcelis niqra has been reported to be 700 mg/1 when in the
form  of manganese chloride and 660 mg/1 when in the form of
manganese nitrate.    Trench,   carp,   and  trout  tolerate  a
manganese   concentration  of  15  mg/1  for  7  days;   yet,
concentrations of manganese above 0.005 mg/1  have  a  toxic
effect on some algae.

Manganese  in  nutrient  solutions  has  been reported to be
toxic to many plants,   the  response  being  a  function  of
species  and nutrient-solution composition.   Toxic levels of
manganese in  solution can vary from 0.5 to 500 mg/1.
                            395

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On the basis of the literature surveyed, it appears that the
concentrations of manganese listed below are deleterious  to
the stated beneficial uses.
    a.   Domestic water supply    0.05 mg/1

    b.   Industrial water supply  0.05 mg/1

    c.   Irrigation               0.50 mg/1

    d.   Stock watering          10.0 mg/1

    e.   Fish and aquatic life    1.0 mg/1

Manganese concentrations are found in the effluents of iron-
ore,  lead, and zinc mining and milling operations and would
be expected from any future operations exploiting  manganese
ores.

Mercury

Elemental mercury occurs as a free metal in certain parts of
the  world;  however, since it is rather inert and insoluble
in water, it is not likely to be found  in  natural  waters.
Although  elemental  mercury  is insoluble in water, many of
the mercuric and mercurous salts, as well as certain organic
mercury   compounds,   are   highly   soluble   in    water.
Concentrations  of  mercury  in  surface waters have usually
been found to be much less than 5 micrograms per liter.

The accumulation and retention of mercurial compounds in the
nervous system, their effect on developing tissue,  and  the
ease  of  their  transmittal  across  the placenta make them
particularly dangerous to man.  Continuous intake of  methyl
mercury  at dosages approaching 0.3 mg Hg per 70 kg (154 Ib)
of  body  weight  per  day  will,  in  time,  produce  toxic
symptoms.

Mercury's   cumulative   nature   also  makes  it  extremely
dangerous to aquatic organisms, since they have the  ability
to  absorb  significant  quantities of mercury directly from
the water as well as through the food chain.  Methyl mercury
is the major toxic form; however,  the  ability  of  certain
microbes  to  synthesize  methyl  mercury from the inorganic
forms  renders  all   mercury   in   waterways   potentially
dangerous.  Fresh-water phytoplankton, macrophytes, and fish
are    capable    of    biologically    magnifying   mercury
concentrations from  water  1,000  times.   A  concentration
                            396

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factor  of  5,000  from water to pike has been reported, and
factors of 10,000 or more have been reported from  water  to
brook  trout.   The  chronic  effects  of mercury on aquatic
organisms are not well-known.  The  lowest  reported  levels
which  have resulted in the death of fish are 0.2 micrograms
per liter of mercury, which killed fathead  minnows  exposed
for  six  weeks.  Levels of 0.1 microgram per liter decrease
photosynthesis  and  growth  of  marine   algae   and   some
freshwater phytoplankton.

Mercury  has  been observed in significant quantities in the
wastewater   in   operations   associated    with    sulfide
mineralization,  including mercury ores, lead and zinc ores,
and copper ores, as well  as  precious-metal  operations  of
gold and silver.  It may be liberated in mine waters as well
as in effluents of flotation concentration and acid-leaching
extraction.

Molybdenum

Molybdenum and its salts are not normally considered serious
pollutants,  but the metal is biologically active.  Although
the element occurs  in  some  minerals,  it  is  not  widely
distributed  in nature.  The mean level of molybdenum in the
U.S.  has been reported to be 68 micrograms per liter.   The
most   important   water   quality   aspect  of  MO  is  its
concentration  in  plants  with  irrigation  and  subsequent
possible molybdenosis of ruminants eating the plants.

The  96-hour  minimum  toxic  level  of  fathead minnows for
molybdic anhydride (Mo03) was found to be 70  mg/1  in  soft
water   and   370   mg/1   in  hard  water.   The  threshold
concentration  for  deleterious  effects   upon   the   alga
Scenedesmus occurs at 54 mg/1.  E. coli and Daphnia tolerate
concentrations  of  1000  mg/1  without  perceptible injury.
Molybdenum can be concentrated from  8  to  60  times  by  a
variety   of  marine  organisms,  including  benthic  algae,
zooplankton, mollusks, crustaceans, and teleosts.

Concentrations of a maximum of 0.05 of the  96-hour  minimum
toxic  level are recommended for protection of the most sen-
sitive species in  sea  water,  while  the  24-hour  average
should not exceed 0.02 of the 96-hour minimum toxic level.

Molybdenum  is found in significant quantities in molybdenum
mining and in milling of uranium ores, where  molybdenum  is
sometimes recovered as a byproduct.
                            397

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Nickel

Elemental   nickel  seldom  occurs  in  nature,  but  nickel
compounds are found in many ores and minerals.   As  a  pure
metal,  it is not a problem in water pollution because it is
not affected by, or soluble in, water.  Many  nickel  salts,
however, are highly soluble in water.

Nickel  is extremely toxic to citrus plants.  It is found in
many soils in California, generally in insoluble  form,  but
excessive  acidification of such soil may render it soluble,
causing severe injury to  or  the  death  of  plants.   Many
experiments with plants in solution cultures have shown that
nickel at 0.5 to 1.0 mg/1 is inhibitory to growth.

Nickel salts can kill fish at very low concentrations.  Data
for  the fathead minnow show death occurring in the range of
5 to 43 mg, depending on the alkalinity of the water.

Nickel is present in coastal and open  ocean  concentrations
in  the  range  of 0.1 to 6.0 micrograms per liter, although
the most common values are 2 to  33  micrograms  per  liter.
Marine  animals  contain  up to 400 micrograms per gram, and
marine plants contain up to 3,000 micrograms per gram.   The
lethal limit of nickel to some marine fish has been reported
to  be  as low as 0.8 ppm (mg/1) (800 micrograms per liter).
Concentrations of 13.1 mg/1 have been reported  to  cause  a
50-percent reduction of photosynthetic activity in the giant
kelp   (Macrocvstis   pyrifers)  in  96  hours,  and  a  low
concentration has been found to kill oyster eggs.

Nickel is found in significant quantities as  a  constituent
of  raw wastewater in the titanium, rare-earth, mercury, and
uranium.

Vanadium

Metallic  vanadium  does  not  occur  free  in  nature,  but
minerals  containing  vanadium  are widespread.  Vanadium is
found in many soils and occurs in vegetation grown  in  such
soils.    Vanadium   adversely   affects   some   plants  in
concentrations as low  as  10  mg/1.   Vanadium  as  calcium
vanadate   can   inhibit   the  growth  of  chicks  and,  in
combination with  selenium,  increases  mortality  in  rats.
Vanadium appears to inhibit the synthesis of cholesterol and
to accelerate its catabolism in rabbits.

Vanadium   causes   death   to   occur   in   fish   at  low
concentrations.  The amount needed for lethality depends  on
                            398

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the  alkalinity  of  the  water  and  the  specific vanadium
compound present.  The common  bluegill  can  be  killed  by
about  6  mg/1  in soft water and 55 mg/1 in hard water when
the vanadium is expressed as vanadyl  sulfate.   Other  fish
are similarly affected.

Limitation  and  control  of  vanadium  levels  appear to be
necessary  in  the  effluents  from   operations   employing
leaching methods to extract vanadium as a primary product or
byproduct.   As  treated  here,  it  can  be  expected to be
contributed by the ferroalloy industry, where high  vanadium
levels were observed both in barren solutions from a solvent
extraction  circuit and in scrubber waters from ore roasting
units.  High vanadium values are also found associated  with
uranium  operations,  where  vanadium  is also obtained as a
byproduct.

Zinc

Occurring abundantly in rocks  and  ores,  zinc  is  readily
refined into a stable pure metal and is used extensively for
galvanizing, in alloys, for electrical purposes, in printing
plates,  for  dye  manufacture and for dyeing processes, and
for many other industrial purposes.  Zinc salts are used  in
paint    pigments,    cosmetics,   pharamaceuticals,   dyes,
insecticides,  and  other  products  too  numerous  to  list
herein.   Many  of these salts (e.g., zinc chloride and zinc
sulfate)  are highly soluble in water; hence,  it  is  to  be
expected  that  zinc  might occur in many industrial wastes.
On the other hand, some zinc  salts  (zinc  carbonate,  zinc
oxide,   and   zinc   sulfide)    are   insoluble  in  water;
consequently, it is to  be  expected  that  some  zinc  will
precipitate  in  and  be  removed  readily from most natural
waters.

In zinc-mining areas, zinc  has  been  found  in  waters  in
concentrations  as high as 50 mg/1; in effluents from metal-
plating works and small-arms ammunition plants, it may occur
in significant concentrations.  In most surface  and  ground
waters,  it is present only in trace amounts.  There is some
evidence  that  zinc  ions   are   adsorbed   strongly   and
permanently on silt, resulting in inactivation of the zinc.

Concentrations of zinc in excess of 5 mg/1 in raw water used
for drinking water supplies cause an undesirable taste which
persists  through  conventional treatment.  Zinc can have an
adverse effect on man and animals  at  high  concentrations.
In  soft  water, concentrations of zinc ranging from 0.01 to
0.1 mg/1 have been reported to be lethal to fish.   zinc  is
                            399

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thought  to exert its toxic action by forming insoluble com-
pounds with the mucous that covers the gills, by  damage  to
the  gill  epithelium,  or possibly by acting as an internal
poison.   The  sensitivity  of  fish  to  zinc  varies  with
species,  age,  and conditions, as well as with the physical
and   chemical   characteristics   of   the   water.    Some
acclimatization to the presence of zinc is possible.  It has
also  been  observed  that the effects of zinc poisoning may
not become apparent  immediately,  so  fish  relocated  from
zinc-contaminated  water  to  zinc-free  water, after 4 to 6
hours of exposure to zinc, may  die  48  hours  later.   The
presence  of  copper  in  water may increase the toxicity of
zinc to aquatic  organisms,  but  the  presence  of  calcium
(hardness)  may decrease the relative toxicity.

Observed values for the distribution of zinc in ocean waters
very  widely.   The  major  concern  with  zinc compounds in
marine water is not one of acute toxicity, but rather of the
longterm sublethal effects of  the  metallic  compounds  and
complexes.     From   an   acute-toxicity   point   of  view,
invertebrate marine animals seem to be  the  most  sensitive
organisms  tested.   The  growth  of  the  sea  urchin,  for
example, has been retarded by as little as 30 micrograms per
liter of zinc.

Zinc sulfate has also  been  found  to  be  lethal  to  many
plants, and it could impair agricultural uses.

Elevated zinc levels were found at operations for the mining
and  milling of lead and zinc ores; at copper mines and flo-
tation mills;  at  gold,  silver,  titanium,  and  beryllium
operations;  and  at  most ferroalloy-ore mining and milling
sites.

Radiation and Radioactivity

Exposure to ionizing radiation at levels substantially above
that of general background levels  has  been  identified  as
harmful  to  living  organisms.   Such  exposure  may  cause
adverse somatic effects such as cancer and  life  shortening
as well as genetic damage.  At environmental levels that may
result  from  releases  by  industries  processing naturally
radioactive materials, the existence of such adverse effects
has not been  definitely  confirmed.   Nevertheless,  it  is
generally agreed that the prudent public health policy is to
assume  a  non-threshold health effect response to radiation
exposure.  Furthermore, a linear response curve is generally
assumed which enables statistical  estimates  of  risk  made
                            400

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from  observed  effects  occurring at higher exposures to be
applied at low levels of exposure.

The half-life of the particular  radionuclides  released  to
the  environment  by  an  industry is extremely important in
determining  the  significance  of  such   releases.     Once
released  to  the  biosphere,  radionuclides with long half-
lives can persist for hundreds and thousands of years.  This
fact coupled with their possible buildup in the  environment
can  lead  to  their  being a source of potential population
exposure for many years.  Therefore, in  order  to  minimize
the  potential  impact  of these radionuclides, they must be
excluded from the biosphere as much as possible.

Plants and animals that  incorporate  radioactivity  through
the biological cycle can pose a health hazard to man through
the  food  chain.  Plants and animals, to be of significance
in the cycling of radionuclides in the  aquatic  environment
must  assimilate  the  radionuclide,  and  retain  it.  Such
processes may lead to bioconcentration of the  radioactivity
so  that  the  activity per gram of food is greater than the
activity per gram of  water.   Bioconcentration  factors  as
great  as  several  thousand have been observed.  Even if an
organism is not eaten before it dies, the radionuclides  may
remain  in the biosphere continuing as a potential source of
exposure.

Aquatic life may  assimilate  radionuclides  from  materials
present  in  the  water,  sediment,  and  biota.  Humans can
assimilate radioactivity through  many  different  pathways.
Among  them are drinking contaminated water, and eating fish
and shellfish that have radionuclides incorporated in  them.
Where  fish  or  other  marine  products that may accumulate
radioactive materials  are  used  as  food  by  humans,  the
concentrations  of  the  radionuclides  in the water must be
restricted to provide assurance that  the  total  intake  of
radionuclides  from  all sources will not exceed recommended
levels.

Naturally  occurring  radionuclides,  particularly  of   the
uranium-238   and   thorium-232  series,  can  be  found  in
appreciable concentrations  in  several  types  of  minerals
throughout the country.  Radium-226, a member of the uranium
series,    is    the    radionuclide   against   which   the
radiotoxicities of most other bone-seeking radionuclides are
compared.  This is due to the relatively high dose delivered
to bones from incorporated radium and the wealth of data  on
the  effects  of  radium-226  on  humans  as  the  result of
numerous medical and industrial exposures.   However,  other
                            401

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radionuclides  in  the  uranium  and  thorium  series may be
important,  particularly  if  released  into  water.   These
include  radium-228,  uranium,  and  lead-210 and its alpha*
emitting daughter, polonium-210.  Radium-228,  a  member  of
the  thorium  series,  has been designated as a radionuclide
for  which  ingestion  should  be  controlled  in   proposed
drinking  water  regulations.   The  isotope  lead-210 is of
particular interest.  Although it is a bone-seeker, a  small
fraction  of  its  daughter,  polonium-210,  is released and
distributed  to  soft   tissue,   where   it   concentrates,
particularly  in  the  liver  and  gonads.   The  levels  of
radionuclides other than radium-226 and uranium  present  in
process streams and treated effluents are generally not well
detailed.    Consequently,  no  other  effluent  limits  are
considered  at  this  time.   However,  because   of   their
potential public health significance, an effluent limitation
on radium-228, lead-210 and polonium-210 may be warranted in
the future.

Radium-226

Radium-226  is a member of the uranium decay series.  It has
a half-life of 1620 years.  This radionuclide  is  naturally
present   in   soils   throughout   the   United  States  in
concentrations ranging from 0.15 to 2.8 picocuries per gram.
It is also naturally present in ground  waters  and  surface
streams in varying concentrations.  Radium-226 is present in
minerals  in  the  earth1s  crust.  Minerals contain varying
concentrations  of  radium-226  and   its   decay   products
depending upon geological methods of deposition and leaching
action   over  the  years.   If  ingested,  the  human  body
incorporates radium into bone  tissue  along  with  calcium.
Some  plants  and animals also concentrate radium so that it
can significantly impact the food chain.

As a result of  its  long  half-life,  radium-226  which  is
present  in minerals extracted from the earth may persist in
the biosphere for many years after its introduction  through
effluents or wastes.  Therefore, because of its radiological
consequences, concentrations of this radionuclide need to be
restricted to minimize potential exposure to humans.

Flotation Reagents

The   toxicity  of  organic  floation  agents—particularly,
collectors and their decomposition products—is an  area  of
considerable   uncertainty,   particularly  in  the  complex
chemical environment present  in  a  typical  flotation-mill
discharge.  Standard analytical tests for individual organic
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 reagents  have  not   evolved to  date.   The tests  for COD and
 TOC are  the  most reliable  tests   currently  available  which
 give  indications of  the presence  of  some of the flotation
 reagents.

 Data available on the fates and  potential toxicities of many
 of   the  reagents indicate that only a  broad range  of
 tolerance  values is  known.  Table VI-1 is a list of some of
 the  more  common flotation   reagents   and   their   known
 toxicities as  judged  from  organism tolerance information.

 Asbestos

 "Asbestos"   is a generic term  for a  number of fire-resistant
 hydrated silicates that, when  crushed or processed, separate
 into flexible  fibers  made  up   of  fibrils  noted  for  their
 great tensile  strength.   The   asbestos minerals differ in
 their metallic elemental content, range of fiber  diameters,
 flexibility, hardness,  tensile strength, surface  properties,
 and  other   attributes  which may  affect their respirability,
 deposition,    retention,    translocation,    and    biologic
 reactivity.

 Asbestos is toxic by inhalation  of  dust particles, with the
 tolerance being 5 million  particles  per cubic foot  of  air.
 Prolonged  inhalation can  cause cancer  of the lungs, pleura,
 and peritoneum.   Little is  known   about  the  movement  of
 asbestos fibers   within   the  human body,  including their
 potential entry through the gastrointestinal  tract.   There
 is   evidence   that bundles  of   fibrils  may be  broken down
 within the body to individual  fibrils.   Asbestos  has  the
 possibility  of  being  a   hazard  when waterborne in large
 concentrations; however, it is insoluble in water.

 To  date, there is  little   data  on  the  concentrations  of
 asbestos  in   ore mining  and  milling  water   discharges.
 Knowledge of the  concentrations in water  that  pose  health
 problems is   poorly  defined.  Currently, this area is being
 investigated by many  researchers concerning themselves  with
 health,  movement,  and analytical techniques.

 Because   of   public  reports  concerning  the  presence  of
 asbestos  in   wastewater  from  an   iron-ore   beneficiation
 operation,    a   reconaissance  analysis  for  asbestos  was
 performed on samples collected as part  of  site  visits  to
 four discharging iron-ore beneficiation operations.   The raw
wastewater  and  effluent  of tailing ponds at each facility
were examined  for the presence or  absence  of  asbestos  or
asbestos-like   fibers.   The  method  of  analysis  used for
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 TABLE Vl-l. KNOWN TOXICITY OF SOME COMMON FLOTATION REAGENTS
                USED IN ORE MINING AND MILLING INDUSTRY
TRADE NAME
Aerofloat 25
Aerof loat 31
Aerofloat 238
Aerofloat 242
Aerof roth 65
Aerofroth 71
Aero Promoter
404
Aero Promoter
3477
AROSURF
MG-98A
Oowfroth 250
DowZ-6
DowZ-11
Dow 2-200
Jaguar
M.I.B.C.
-
Superfloc 16
CHEMICAL COMPOSITION
EtMirtMly aryl dithtophosphoric acid
EaMntlally aryl dithiophosphorie acid
Sodium di-secondary butyl
dithiophoiphate
EaMntially aryl dithiophoiphorlc acid
Polyglycol type compound
Mlxtura of 6-9 carbon alcohol*
Mixture of nilfhydryl type compound!
Unknown
Unknown
Chromium salts (ammonium, potassium,
and sodium chromate and ammonium,
potassium, and sodium dichromatel
Copper sulfate
Cresylic acid
Polypropylene glycol methyl ethers
Potassium amyl xanthate
Sodium isopropyl xanthata
Isopropyl ethylthionocarbsmate
Based on guar gum
Lime (calcium oxidet
Methylisobutylearbinol
Pine oil
Potassium ferricyanida
Sodium farrocyanide
Sodium hydroxide
Sodium oleata
Sodium silicate
Sodium sulfide
Sulfuric acid
Polyacrylamide
FUNCTION
Collector/Promoter
Collector/Promoter
Collector /Promoter
Collector/Promoter
Frother
Frother
Collector/Promoter
Collector/Promoter
Collector/Promoter
Depressing agent
Activating agent
Frother
Frother
Collector /Promoter
Collector/Promoter
Collector/Promoter
Floeculant
pH modifier and
flocculant
Frothar
Frother
Depressing agent
Depressing agent
pH modifier
Frother
Depressing agent
Activating agent
pH modifier and
flocculant
Flocculant
KNOWN TOXIC
RANGE (mg/Jl)
1000 to 10,000
10 to 1000
1000 to 10.000
>1000
1 to 100
100 to 1000
10 to 1000
0.01 to 1.0
0.1 to 1.0
>1000
0.1 to 200
0.2 to 2.0
10 to 100
10 to 1000
>1000
1 to 100
0.25 to 2.5
to 1000
to 1000
to 1000
00 to 1000
to 100
to 100
>1000
TOXICITV*
Low
Moderate
Low
Low
Moderate
Moderate
Moderate
High
High
Low
Moderate to High
High
Moderate
Moderate
Low
Moderate
Moderate to High
Moderate
Moderate
Moderate
Moderate
Moderate
Moderate
Low
Toxicity   Tolerance Level
 High
Moderate
 Low
 <1.0mg/£
1.0to1QOOmg/£
 >1000mg/£
NOTE:  Toxic range is a function of organism tested and water quality, including hardness
       and pH. Therefore, toxicity deta presented in this table are only generally indica-
       tive of reagant toxicity. Although the toxicity ranges presented here are based on
       many different organisms, much of the data are presented in relation to salmon,
       fathead minnows, sticklebacks, and Daphnia.
                                    404

-------
 detection  was  one  based  upon  published  literature  and
 employed scanning electron microscopy.


 Fibers  were  not  detected  in  any of the samples with the
 exception of the influent to  the  tailing  pond  from  Mill
 1107.   Energy-dispersive x-ray analysis indicated, however,
 that the fiber was not of an asbestos type.   Both  raw  and
 treated  wastewaters  from  mills 1107, 1108, 1109, and 1110
 were examined, and no  asbestos  or  asbestos-like  minerals
 were found.

 While  the  results  of  the  survey indicate the absence of
 asbestos fibers at  each  of  the  sites  investigated,  the
 presence  or  absence  of asbestos at other locations in the
 iron-ore  mining  and  beneficiation  industry   cannot   be
 confirmed.     It  does  not  appear  possible  to  recommend
 effluent levels or treatment technology at this time.   It is
 recommended, however, that a  reconaissance  evaluation  for
 asbestos   be   performed   at   each  iron-ore  mining  and
 beneficiation  operation  to  determine   whether   possible
 asbestos levels of concern are present.

 SIGNIFICANCE   AND  RATIONALE  FOR  REJECTION  OF  POLLUTION
 PARAMETERS

 A  number of pollution parameters  besides those  selected  and
 just  discussed  were  considered  in each category but were
 rejected for one or  more of these reasons:

     (1)   Simultaneous reduction   is   achieved  with another
          parameter which is limited.

     (2)   Treatment   does  not "practically"  or economically
          reduce the  parameter.

     (3)   The   parameter was  not    usually   observed   in
          quantities    sufficient   to   cause  water-quality
          degradation.

     (4)   There  are insufficient data  on water-quality degra-
          dation or   treatment  methods   which   might   be
          employed.

Because  of  the  great   diversity of the ores mined and the
processes employed in the ore mining and dressing  industry,
selections   for  subcategories  of  the  parameters  to  be
monitored and controlled—as well  as  those  rejected—vary
considerably.    Parameters   listed  in  this  section  are
                            405

-------
parameters which have been rejected for the ore  mining  and
dressing industry as a whole.

Barium and Boron

Barium and boron are not present in quantities sufficient to
justify consideration as harmful pollutants.

Calcium, Magnesium, Potassium, Strontium, and Sodium

Although these metals commonly occur in effluents associated
with  ore  mining  and  dressing  activities,  they  are not
present in  quantities  sufficient  to  cause  water-quality
degradation,  or  there  are  no practical treatment methods
which can be employed on a  large  scale  to  control  these
elements.

Carbonate

There  are  insufficient  data  for  dissolved  carbonate to
justify consideration of this ion as a harmful pollutant.

Nitrate and Nitrite

There are  insufficient  data  for  dissolved  nitrates  and
nitrites   to   justify   their   consideration  as  harmful
pollutants, although nitrogen and nitrate contributions  are
known  to  stimulate  plant  and  algal growth.  There is no
treatment available to practically reduce these ions.

Selenium

The levels of selenium  observed  in  the  wastewaters  from
mines and mills are not sufficiently high for selenium to be
considered as a harmful pollutant.

Silicates

Silicates  may  be  present  in the wastewaters from the ore
mining and dressing industry, but the levels encountered are
not sufficiently high to warrant classification as a harmful
pollutant.

Tin

Tin does not exist in sufficient quantities  from  mines  or
mills to be considered a harmful pollutant.
                            406

-------
Zirconium

There  is  no  information  available  which  indicates that
significant levels of zirconium are present in the  industry
to be classed as harmful.

Total Dissolved solids

High dissolved-solid concentrations are often caused by acid
conditions  or  by the presence of easily dissolved minerals
in the  ore.   since  economic  methods  of  dissolved-solid
reduction  do  not exist, effluent limitations have not been
proposed for this parameter.

SUMMARY OF POLLUTION PARAMETERS SELECTED BY CATEGORY

Because of the wide variations observed with respect to both
waste components  discharged  and  loading  factors  in  the
different  segments of the ore mining and dressing industry,
a single, unified list of all parameters  selected  for  the
industry  as  a whole would not be useful.  Therefore,  Table
VI-2  summarizes  the   parameters   chosen   for   effluent
limitation guidelines for each industry metal category.
                            407

-------
TABLE VI-2. SUMMARY OF PARAMETERS SELECTED FOR EFFLUENT
         LIMITATION BY METAL CATEGORY
PARAMETERS
pH (Acidity/Alkalinity)
Total Suspended Solids (TSS) , .
Settteable Solids
Chemical Oxygen Demand (COD)
Cyanide
Ammonia
Aluminum
Antimony
Arsenic

Cadmium
Chromium
Copper
Iron
Lead
Mercury
Molybdenum
Nickel
Vanadium
Zinc
Radium
Uranium
PARAMETERS SELECTED FOR EFFLUENT LIMITATIONS
6
*
0











O








£
6
i
a
Q
f
Q


0





0

0






Q


Lead and Zinc Ores
L
f
£


9





0

f






0


o>
O
•o
3
U
f



£





£

^

0
9



6


Silver Ores
*
•
0

0





^

^

^
•



0


$
6
0)
X
I






t






0





t


Ferroalloy Ores
9
9

9




0

0
•
0

t

0


f


Mercury Ores
^
^













^

0




1 Uranium. Radium,
and Vanadium Ores
4
0

f

0


0

0





afc

•
f
I

Metal Ores. Not Elsewhere Classified
1 Antimony
Ores
4
0





0
0




A





9


E
Is
mo

%
n
•5
o
N
limitation
o separate
z




I
"» M
5 2
510
|

















-
M
(=6
jdenum
>>
o
x
i
*-
U
•o ,
O
a
M
re

O
2
o


fa"
o
IV
•- s
II
ss
zr
Titanium
Ores
0
0

f









0



0
^


Rare-Earth
Ores
£
s
'•S
o
N
8
0
n
1
ra
1
1
Zirconium
Ores
' as byproduct with titanium. 1
c
o
1
o
E
3
C
O
u
M
C
o
1
£
C
                  408

-------
                                 TABLE

                                METRIC TABLE

                              CONVERSION TABLE
 MULTIPLY (ENGLISH UNITS)

     ENGLISH UNIT  ABBREVIATION
 acre
 acre - feet
 British Thermal
   Unit
 British Thermal
   Unit/ pound
 cubic feet/minute
 cubic feet/second
 cubic feet
 cubic feet
 cubic Inches
 degree Fahrenheit
 feet
 gallon
 gallon/minute
 horsepower
 Inches
 Inches  of mercury
 Pounds
 "ill 11 on gallons/day
Pound/sguare
  Inch (gauge)
square feet
square Inches
ton (short)
yard
                     ac
                     ac ft

                     BTU

                     BTU/lb
                     cfm
                     cfs
                     cu ft
                     cu ft
                     cu in
                     °F
                     ft
                     gal
                     gpm
                     hp
                     1n
                     In Hg
                     lb
                    ml

                    pslg
                    sq ft
                    sq 1n
                    ton
                    yd
                                        by                TO OBTAIN (METRIC UNITS)

                                   CONVERSION   ABBREVIATION   METRIC UNIT
 hectares
 cubic meters

 kilogram -  calories

 kilogram calories/kilogram
 cubic meters/minute
 cubic meters/minute
 cubic meters
 liters
 cubic centimeters
 degree Centigrade
 meters
 liters
 liters/second
 killowatts
 centimeters
 atmospheres
 kilograms
 cubic  meters/day
 kilometer

 atmospheres  (absolute)
 square meters
 square centimeters
metric ton (1000 kilograms)
meter
0.405
1233.5
0.252
0.555
0.028
1.7
0.028
28.32
16.39
0.555(°F-32)*
0.3048
3.785
0.0631
0.7457
2.54
0.03342
0.454
3,785
1.609
(0.06805 psig +1)*
0.0929
6.452
0.907
0.9144
ha
cu m
kg cal
kg cal/kg
cu m/min
cu m/min
cu m
1
cu cm
°C
m
1
I/sec
kw
cm
atm
kg
cu m/day
km
atm
sq m
sq cm
kkg
m
* Actual conversion, not a multiplier
                                    409

-------