EPA 440/1-75/061
  GROUP II
       Development Document for
   Interim Final and Proposed Effluent
  Limitations Guidelines and New Source
      Performance Standards for the
     Ore Mining and Dressing Industry
         Point Source Category
                  Vol. I
                 vF

                
-------
          DEVELOPMENT DOCUMENT

                  for

       INTERIM FINAL AND PROPOSED

    EFFLUENT LIMITATIONS GUIDELINES

                  and

    NEW SOURCE PERFORMANCE STANDARDS

                for the

        ORE MINING AND DRESSING

         POINT SOURCE CATEGORY

       VOLUME I - SECTIONS I  - VI
            Russell E.  Train
             Administrator

      Andrew W.  Breidenbach, Ph.D
   Acting Assistant Administrator for
     Water and Hazardous Materials
                        \
                        ^
              Allen Cywin
 Director, Effluent Guidelines Division

             Donald C.  Gipe
            Project Officer

            Ronald G.  Kirby
       Assistant Project Officer

      Effluent Guidelines Division
Office of Water and Hazardous Materials
  U.S. Environmental Protection Agency
        Washington, D.C.  20460
              October 1975

-------

-------
                          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   (33  U.S.C.  1551, 1314, and 1316, 86 Stat. 816 et.
seq.)  (the "Act") .

Effluent limitations guidelines contained herein  set  forth
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,   1983,   respectively.   The  standards  of
performance  and  pretreatment  standards  for  new  sources
contained  herein set forth the degree of effluent reduction
which is achievable through  the  application  of  the  best
available   demonstrated   control   technology,  processes,
operating methods, or other alternatives.

Based upon the application of the best  practicable  control
technology  currently  available, 14 of the 41 subcategories
for which separate limitations are suggested can be operated
with no discharge of process waste  water.   With  the  best
available  technology  economically achievable, 21 of the 41
subcategories for which separate  limitations  are  proposed
can  be operated with no discharge of process waste water to
navigable waters.   No  discharge  of  process  waste  water
pollutants  is  also  achievable as a new source performance
standard for 21 of the 41 subcategories.

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) .
                           111

-------
                    CONTENTS (VOLUME I)

Section                                                       Pa§e

I        CONCLUSIONS                                           !

II       RECOMMENDATIONS                                       3

III      INTRODUCTION                                          H

              PURPOSE AND AUTHORITY                            H

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

              SUMMARY OF ORE-BENEFICIATION PROCESSES           17

              GENERAL DESCRIPTION OF INDUSTRY BY ORE           29
              CATEGORY

IV       INDUSTRY CATEGORIZATION                               141

              INTRODUCTION                                     141

              FACTORS INFLUENCING SELECTION OF                 143
              SUBCATEGORIES IN ALL METAL ORE CATEGORIES

              DISCUSSION OF PRIMARY FACTORS INFLUENCING        148
              SUBCATEGORIZATION BY ORE CATEGORY

              SUMMARY OF RECOMMENDED SUBCATEGORIZATION         169

              FINAL SUBCATEGORIZATION                          169

V        WASTE CHARACTERIZATION                                173

              INTRODUCTION

              SPECIFIC WATER USES IN ALL CATEGORIES            175

              PROCESS WASTE CHARACTERISTICS BY ORE             176
              CATEGORY

VI       SELECTION OF POLLUTANT PARAMETERS                     373

              INTRODUCTION                                      373

              GUIDELINE PARAMETER-SELECTION CRITERIA            373
                               v

-------
SIGNIFICANCE AND RATIONALE FOR SELECTION          374
OF POLLUTION PARAMETERS

SIGNIFICANCE AND RATIONALE FOR REJECTION          398
OF POLLUTION PARAMETERS

SUMMARY OF POLLUTION PARAMETERS SELECTED          400
BY CATEGORY
                VI

-------
                    CONTENTS (VOLUME II)

Section                                                         Page

VII      CONTROL AND TREATMENT TECHNOLOGY                        403

              INTRODUCTION                                       403

              CONTROL PRACTICES AND TECHNOLOGY                   404

              TREATMENT TECHNOLOGY                               419

              EXEMPLARY TREATMENT OPERATIONS BY ORE              460
              CATEGORY

VIII     COST, ENERGY, AND NONWATER-QUALITY ASPECTS              567

              INTRODUCTION                                       567

              SUMMARY OF METHODS USED                            567

              WASTEWATER-TREATMENT COSTS FOR IRON-ORE            573
              CATEGORY

              WASTEWATER TREATMENT COSTS FOR COPPER-ORE          581
              CATEGORY

              WASTEWATER-TREATMENT COSTS FOR LEAD- AND           588
              ZINC-ORE CATEGORY

              WASTEWATER-TREATMENT COSTS FOR GOLD-ORE            600
              CATEGORY

              WASTEWATER-TREATMENT COSTS FOR SILVER-ORE          621
              CATEGORY

              WASTEWATER-TREATMENT COSTS FOR BAUXITE             631
              CATEGORY

              WASTEWATER-TREATMENT COSTS FOR FERROALLOY-         634
              ORE CATEGORY

              WASTEWATER TREATMENT COSTS FOR MERCURY-            658
              ORE CATEGORY

              WASTEWATER TREATMENT COSTS FOR URANIUM-            670
              ORE CATEGORY

              WASTEWATER TREATMENT COSTS FOR METAL               685
              ORES,  NOT  ELSEHWERE  CLASSIFIED

              NON-WATER  QUALITY  ASPECTS                         6"
                                vii

-------
IX       BEST PRACTICABLE CONTROL TECHNOLOGY CURRENTLY          703
         AVAILABLE, GUIDELINES AND LIMITATIONS

              INTRODUCTION                                      703

              GENERAL WATER GUIDELINES                          705

              BEST PRACTICABLE CONTROL TECHNOLOGY               707
              CURRENTLY AVAILABLE BY ORE CATEGORY
              AND SUBCATEGORY

X        BEST AVAILABLE TECHNOLOGY ECONOMICALLY                 763
         ACHIEVABLE, GUIDELINES AND LIMITATIONS

              INTRODUCTION                                      753

              GENERAL WATER GUIDELINES                          764

              BEST AVAILABLE TECHNOLOGY ECONOMICALLY             766
              ACHIEVABLE BY ORE CATEGORY AND SUBCATEGORY


XI       NEW SOURCE PERFORMANCE STANDARDS AND                   795
         PRETREATMENT STANDARDS

              INTRODUCTION                                      795

              GENERAL WATER GUIDELINES                          796

              NEW SOURCE STANDARDS BY ORE CATEGORY              796

              PRETREATMENT STANDARDS                            801

XII      ACKNOWLEDGMENTS                                        809

XIII     REFERENCES                                             813

XIV      GLOSSARY                                               821

              LIST OF CHEMICAL SYMBOLS                          846

              CONVERSION TABLE                                  847
                             Vlll

-------
                      TABLES (VOLUME I)

No.,                             Title                           Pa2<


II-l     Summary of Recommended BPCTCA Effluent Limitations      4
              By Category and Subcategory—Ores for Which
              Separate Limitations Are Proposed
II-2     Summary of Recommended BATEA Effluent Limitations       6
              By Category and Subcategory—Ores for Which
              Separate Limitations Are Proposed
II-3     Summary of Recommended NSPS Effluent Limitations        8
              By Category and Subcategory—Ores for Which
              Separate Limitations Are Proposed
HI-1    Iron-Ore Shipments for United States                    30
III-2    Crude Iron-Ore Production for U.S.                      31
III-3    Reagents Used for  Flotation of Iron Ores                36
III-4    Various Flotation  Methods Available for Pro-            37
              duction of High-Grade  Iron-Ore Concentrates
III-5    Total Copper-Mine  Production of Ore by Year             42
III-6    copper-Ore Production from  Mines  by State  (1972)        42
III-7    Average Copper Content of Domestic Ore                  44
III-8    Average Concentration of Copper in Domestic ores        44
              by Process  (1972)
III-9    Copper Ore Concentrated in  the United States            45
              by Froth Flotation, Including LPF Process
               (1972)
111-10   Heap or Vat Ore Leached in  the United States  (1972)     48
III-11   Average Price Received from Copper in the               50
              United States
III-12   production of Copper  from Domestic Ore by               51
              Smelters
111-13   Mine Production of Recoverable Lead in the              53
              United States
111-14   Mine Production of Recoverable Zinc in the              54
              United States (Preliminary)
III-15   Domestic  Silver Production  from Different               66
              Types of Ores
111-16   Silver  Produced at Amalgamation and Cyanidation        67
              Mills in the  U.S. and  Percentage  of  Silver
              Recoverable  from All  Sources
III-17   Production of Bauxite in  the  United  states              71
111-18   Production of Ferroalloys  by  U.  S. Mining  and          73
               Milling  Industry
111-19   Observed  Usage  of  Some Flotation  Reagents               95
III-20   Probable  Reagents  Used in Flotation  of  Nickel           gg
               and  Cobalt  ores
                                  IX

-------
                            TABLES  (cont.)

 —                            Title

 111-21   Domestic Mercury Production Statistics
 III-22   Isotopic Abundance of Uranium
 III-23   Uranium Milling Activity by State, 1972
 111-24   Uranium Concentration in IX/SX Eluates
 III-25   Decay Series of Thorium and Uranium
 111-26   Uranium Milling Processes
 III-27   Uranium Production
 III-28   Vanadium Production
 111-29   Vanadium Use                                            125
 III-30   Production of Antimony from Domestic Sources
 111-31   Domestic Platinum-Group Mine Production and Value
 111-32   Production and Mine Shipments of Titanium               136
               Concentrates from Domestic Ores in the US
 I v-1      summary of Industry Subcategorization Recomended
 IV-2      Final Subcategorization
 V-l       Historical Constituents of Iron-Mine Discharges
 V-2       Historical Constituents of Waste water from Iron-
               Ore Processing
 V-3       Chemical Compositions  of Sampled Mine Waters
 V-4       Chemical Compositions  of Sampled Mill Waters
 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
 V-6       Chemical Characteristics of Discharge Water
               from Mine 1108
 V-7       Characteristics of Mill  1108  Discharge Water
 V-8       Principal Copper Minerals Used in the United States
 V-9       Mine-Water Production  from Selected  Major Copper-
               Producing Mines and Fate(s)  of  Effluent
 V-10      Summary  of Solid Wastes  Produced  by  Plants
               Surveyed
 V-ll      Raw Waste Load in  Water  Pumped from  Selected
               Copper Mines
 V-12      1973  Water  Usage in  Dump,  Heap,  and  In-Situ
               Leaching  Operations
 V-l3      Chemical Characteristics  of Barren Heap,  Dump, or
               In-Situ Acid  Leach  Solutions (Recycled:  No
               Waste Load)
 V-14     Water Usage in Vat Leaching Process  as  a  Function      71Q
               of Amount of  Product  (Precipitate  or Cathode
              Copper) Produced
V-15     Chemical Characteristics of Vat-Leach Barren
              Acid Solution  (Recycled:  No Waste Load)

-------
                            TABLES  (cont.)

No.                             Title

V-16          Miscellaneous Wastes from Special Handling of      221
              Ore Wash Slimes in Mine 2124  (No Effluent)
V-17          Examples of Chemical Agents Which May be Employed  225
              In copper Flotation
V-18          Water Usage in Froth Flotation of Copper           227
V-19          Raw Mill Waste Loads Prior to Settling in Tailing  228
              Ponds
V-20          Waste water Constituents and Waste Loads Resulting 233
              from Discharge of Mill Process Waters
V-21          Range of Chemical Characteristics of Sampled Raw   241
              Mine Water from Lead/Zinc Mines 3102, 3103,
              and 3104
V-22          Range of Chemical Characteristics of Raw Mine      242
              Waters from Four Operations in Solubiliza-
              tion-Potential Subcategory
V-23          Ranges of constituents of Waste waters and Raw  Waste248
              Loads for Mills 3102, 3103, 3104, 3105, and
              3106
V-24          Chemical Composition  of Raw Mine Water from Mines  253
              4105 and 4102
V-25          Process Reagent Use at Various Mills Beneficiating256
              Gold Ore
V-26          Minerals Commonly Associated  with Gold Ore         256
V-27          Waste characteristics and Raw Waste Loads at Four  258
              Gold Milling Operations
V-28          Raw Waste Characteristics of  Silver Mining         263
              Operations
V-29          Major Minerals Found  Associated with Silver Ores   266
V-30          Flotation Reagents Used by  Three Mills to Bene-   267
              ficiate Silver-Containing Mineral Tetrahedrite
               (Mills 4401 and 4403) and Native Silver and
              Argentite  (Mill 4402)
V-31          Waste characteristics and Raw Waste Loads at Mills 268
               4401, 4402, 4403, and 4105
V-32          Concentrations of selected  Constituents  in Acid   275
              Raw Mine Drainage from Open-Pit Mine  5101
V-33          Concentrations of Selected  constituents  in Acid   275
              Raw Mine Drainage from Open-Pit Mine  5102
V-34          concentrations of Selected  constituents  in Alkaline276
              Raw Mine Drainage from Underground Mine  5101
V-35          Waste water and Raw waste Load  for Open-Pit Mine  5101  278
V-36          Waste water and Raw Waste Load  for Underground     278
              Mine  5101
V-37          Types of Operations Visited and Anticipated—      279
               Ferroalloy-Ore Mining and Dressing Industry
                                   XI

-------
                            TABLES  (cont.)

 No.
                                                                 Page
 V~38          Chemical Characteristics of Raw Mine Water in
               Ferroalloy Industry
 V-39          Reagent Use in Molybdenum Mill 6101                289
 V"40          *aw Waste Characterization and Raw Waste Load      289

 V~41          Reagent Use for Rougher and Scavenger Flotation    292
               at Mill 6102
                       Use for Cleaner Flotation at Mill 6102
   ^          Reagent Use at Byproduct Plant of Mill 6102  (Based 293
               on Total Byproduct Plant Feed)
 V~44          Mflj- 6102 Effluent Chemical Characteristics  (com-  293
               bined-Tailings Sample)
 Y~f5          Chemical Characteristics of Acid-Flotation Step    295
 V~46          Composite Waste Characteristics for Beneficiatior 299
               at Mill 6104 (Samples 6, 8, 9,  and 11)
 V~47          Waste Characteristics from Copper-Thickener over-  299
               flow for Mill 6104 (Sample 5)
 V-48          Scheelite-Flotation Tailing Waste Characteristics  300
               and Loading for Mill 6104  (Sample 7)
 JJ"?«          50-Foot-Thickener Overflow for  Mill 6104 (Sample 10)300
 V  50          waste Characteristics of Combined-Tailing Discharge 301
               for Mill 6104 (Samples  15,  16,  and 17)
 V~5i          Waste Characteristics and  Raw Waste Load at Mill
               6105 (Sample 19)
 v~52          Chemical Composition of Waste water,  Total Waste,  302
               and Raw Waste Loading from Milling and  Smelter
               Effluent for Mill 6106
 V~53          Waste Characterization  and  Raw  Waste  Load for      306
               Mill 6107 Leach and Solvent-Extraction  Effluent
               (Sample 80)
 v~54           Waste Characteristics and Waste  Load  for Dryer     307
               Scrubber Bleed  at Mill  6107  (Sample 81)
 v~55           Waste Characteristics and Loading  for Salt-Roast   3ns
               Scrubber Bleed  at Mill  6107  (Sample 77)
 v~56           Expected Reagent  Use  at  Mercury-Ore Flotation
               Mill  9202
v~57           Waste  Characteristics and Raw Waste Loadings  at
               Mills  9201 and  9202
 V-58           Waste  Constituents  Expected
v~59           Chemical and Physical Waste Constituents  observed
               in Representative Operations
                                 Xll

-------
                           TABLES (cont. )
No.

V-60          Water Use and Flows at Mine/Mills 9401, 9402, 9403, 326
              and 9404                                  .
V-61          Water Treatment Involved in U/Ra/V Operations       J^b
V-62          Radionuclides in Raw Waste waters from Uranium/     334
              Radium/Vanadium Mines and Mills
V-63          Organic Constituents in U/Ra/V Raw Waste  water      334
V-64          Inorganic Anions in U/Ra/V Raw Waste water          336
V-65          Light-Metal Concentrations Observed in U/Ra/V       336
              Raw Waste water
V-66          Concentrations of Heavy Metals Forming Anionic      336
              Species in U/Ra/V Raw Waste water
V-67          concentrations of Heavy Metals Forming Cationic     337
              Species in U/Ra/V Raw Waste water
V-68          Other constituents Present in Raw Waste water in    337
              U/Ra/V Mines and Mills
V-69          Chemical Composition of Waste water and Raw  Waste   339
              Load for Uranium Mines 9401 and  9402
V-70          Chemical Composition of Raw Waste water and  Raw     339
              Waste Load for Mill 9401  (Alkaline-Mill
              Subcategory)
V-71          Chemical Composition of Waste water and Raw  Waste   340
              Load for Mill 9402  (Acid- or Combined  Acid/
              Alkaline-Mill Subcategory)
V-72          Chemical Composition of Waste water and Raw  Waste   341
              Load for Mine 9403  (Alkaline-Mill Subcategory)
V-73          Chemical Composition of Waste water and Raw  Waste   342
              Load for Mill 9404  (Acid- or Combined  Acid/
              Alkaline-Mill Subcategory)
V-74          Reagent Use at Antimony-Ore Flotation  Mill 9901     349
V-75          Chemical Composition of Raw Waste water Discharged  350
              From Antimony Flotation Mill 9901
V-76          Major Waste Constituents  and Raw Waste Load  at      351
              Antimony Mill 9901
V-77          Chemical Composition of Raw Waste water from       353
              Beryllium Mill 9902  (No Discharge from
              Treatment)
V-78          Chemical Composition of Raw Waste water from       353
              Rare-Earth Mill  9903
V-79          Results of Chemical Analysis for Rare-Earth         359
              Metals  (Mill 9903 — No  Discharge)
                                  xin

-------
                            TABLES (cont.)

 —                            ^^                            Page
 V~80           Chemical  Composition and  Raw Waste Load from       361
               Rare-Earth  Mill  9903
 V~81           Chemical  Composition and  Loading for Principal     363
               Waste Constituents  Resulting from Platinum
               Mine/Mill 9904  (Industry  Data)
 V~82           Chemical  Composition of Raw  Waste water from       364
               Titanium  Mine 9905
 v~83           Chemical  Composition of Raw  Waste water from       366
               Titanium  Mill 9905
 V~8£*           Reagent Use in Flotation  Circuit of  Mill  9905       ^*
 V~°l           Principal Minerals  Associated with Ore  of  Mine 9905 367
 v~bb           Major Waste Constituents  and Raw Waste  Load  at     367
               Mill 9905
 v~87           Chemical  Composition of Raw  Waste water at Mills    371
               9906 and  9907
 V~88          Raw Waste Loads for  Principal Waste  water  Consti-   372
               tuents from Sand Placer Mills 9906 and  9907

VI~1          Known Toxicity of Some Common Flotation Reagents    396
              Used in Ore Mining and Milling Industry
VI~2          Summary of Parameters Selected for Effluent  Limi-   401
              tation by Metal Category
                                xiv

-------
                     FIGURES (VOLUME I)


Mo.                            Title

III-l    Beneficiation of Iron Ores                              ^
III-2    Iron-Ore Flotation-Circuit Flowsheet                    *>
III-3    Magnetic Taconite Beneficiation Flowsheet               J°
III-U    Agglomeration Flowsheet                                 ^9
III-5    Major copper Mining and Milling Zones of the U.S.       4J-
III-6    General Outline of Methods for Typical Recovery         46
              of Copper from Ore
III-7    Major Copper Areas Employing Acid Leaching in
              Heaps, in Dumps, or In Situ
III-8    Lead/Zinc-Ore Mining and Processing Operations          56
III-9    cyanidation of Gold Ore:  Vat Leaching of Sands         61
              and 'Carbon-in-Pulp1 Processing of Slimes
111-10   Cyanidation of Gold Ore;  Agitation/Leach               63
              Process
III-11   Flotation of Gold-Containing Minerals with              64
              Recovery of Residual Gold Values by
              Cyanidation
111-12   Recovery of Silver .Sulfide Ore by Froth                 68
              Flotation
III-13   Gravity-Plant Flowsheet for Nigerian Columbite          81
111-14   Euxenite/Columbite Beneficiation-Plant Flowsheet       82
III-15   Representative Flow Sheet for Simple Gravity            83
              Mill
111-16   Simplified Molybdenum Mill Flowsheet                    86
111-17   Simplified Molybdenum Mill Flow Diagram                 88
111-18   Simplified Flow Diagram for Small Tungsten              91
              Concentrator
III-19   Mill Flowsheet for a Canadian columbium                 92
              Operation
III-20   Flowsheet of Tristage crystallization  Process           95
              for Recovery of Vanadium, Phosphorus,  and
              Chromium from Western Ferrophosphorus
III-21   Arkansas Vanadium Process Flowsheet                    96
III-22   Flowsheet of Dean-Leute Ammonium  Carbamate              97
              Process
111-23   Pachuca Tank for Alkaline Leaching                      108
111-24   Concentration Processes and Terminology                 112
III-25   Simplified Schematic Diagram of Sulfuric  Acid           118
              Digestion of Monazite  Sand for Recovery
              of Thorium, Uranium, and Rare  Earths

-------
                           FIGURES  (cont.)

 —                             Title

 111-26   Simplified Schematic Diagram of Caustic Soda            119
               Digestion of Monazite Sand for Recovery
 TTT __    _.  of Thorium, Uranium, and Rare Earths
 III-27   Effect of Acidity on Precipitation of Thorium,          120
               Rare Earths and Uranium from a Monazite/
               Sulfuric Acid Solution of Idaho and
               Indian Monazite Sands
 111-28   Generalized Flow Diagram for Production of              124
               Uranium, Vanadium, and Radium
 111-29   Beneficiation of Antimony Sulfide Ore by                TOO
               Flotation
 TTT"^?   Gravity Concentration of Platinum-Group Metals          133
 TTT oi   Beneficiation of Heavy-Mineral Beach Sands              130
 111-32   Beneficiation of Ilmenite Mined from a Rock Deposit     140
 v-i      Flow Scheme for Treatment of Mine Water                 183
                               aca                           183
 v-j       Water Balance for Mine/Mill 1105 (September 1974)
 v-4       Concentrator Flowsheet for Mill 1105
 V-5       Flowsheet for Mill 1104 (Heavy-Media Plant)             19Q
 V-6       Simplified Concentration Flowsheet for Mine/Mill 1108  194
 V-7       Waste water Flowsheet for Plant 2120-B Pit             TQQ
V-8       Flowsheet  of  Hydrometallurgical Process Used in        207
              Acid  Leaching  at  Mine 2122
V-9       Reactions  by  Which  Copper Minerals Are Dissolved in    209
              Dump,  Heap,  or In-Situ Leaching
V~10          Typical  Design of Gravity  Launder/Precipitation   ?in
              Plant
v~11          Cutaway  Diagram of Cone Precipitator              211
V~12          Diagram  of Solvent Extraction Process  for Recovery 213
„ __          of Copper by Leaching of Ore  and  Waste
v~13          Vat Leach Flow Diagram (Mill  2124)                 217
V~1{|          Flow  Diagram for  Flotation of Copper (Mill 2120)   222
v 15          Addition of  Flotation Agents  to Modify Mineral    994
              Surface
V~16          Flowsheet for  Miscellaneous Handling of Flotation
              Tails  (Mill  2124)
v~17          Dual Processing of Ore  (Mill 2124)                 23fi
v~18          Leach/Precipitation/Flotation Process              007
v~19          Water Flow Diagram for Mine 3105
                                  .          .
v~20          Water Flow Diagram for Mine 3104
v~21          Flow Diagram for Mill 3103                         247
                                 xvi

-------
                           FIGURES  (cont. )
No.

V-22          Water Flow in Four selected Gold Mining and        251
              Milling Operations
V-23          Water Flow in Silver Mines and Mills
V-24          Process and Waste water Flow Diagram for Open-Pit
              Bauxite Mine 5101
V-25          Mill 6601 Flowsheet
V-26          Simplified Mill Flow Diagram for Mill 6102
V-27          Internal Water Flow For Mill 6101 Through
              Molybdenum Separation
V-28          Internal Water Flow for Mill 6104 Following        297
              Molybdenum Separation
V-29          Water Use and Waste Sources for Vanadium Mill  6107 304
V-30          Water Flow in Mercury Mills 9101 and 9102          310
V-31          Typical Water-Use Patterns                         316
V-32          Alkaline-Leach Water Flow                          322
V-33          Ammonium Carbonate Leaching Process                324
V-34          Water Flow in Mills 9401, 9402, 9403, and  9404    327
V-35          Flowchart of Mill 9401                             328
V-36          Flow Chart for Mill 9402                           329
V-37          Flow Chart of Mill 9403                            330
V-38          Flow Chart of Mill 9404                            331
V-39          Water Flows and Usage for Mine/Mills 9901  (Antimonyp44
              and 9902  (Beryllium)
V-40          Water Flows and Usage for Mine/Mills 9903          345
               (Rare Earths) and 9904  (Platinum)
V-41          Water Flows and Usage for Titanium  Mine/Mills      346
              9905 and 9906
V-42          Beneficiation of Bertrandite,  Mined  from  a Lode   355
              Deposit by Flotation  (Mill 9903)
V-43          Beneficiation of Rare-Earth Flotation  Concentrate 356
              by Solvent Extraction  (Mill 9903)
V-44          Beneficiation and Waste  Water  Flow  of  Ilmenite    365
              Mine/Mill 9905  (Rock Deposit)
V-45          Beneficiation of Heavy-Mineral Beach Sands (Rutile,370
              Ilmenite, Zircon, and Monazite)  at  Mill 9906
                                  xvn

-------
                         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  ores,  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 14 of 41 subcategories for which separate limitations
are  proposed  can  be operated with no discharge of process
waste   water.    With   the   best   available   technology
economically   achievable,   facilities  in  21  of  the  41
subcategories can be operated with no discharge  of  process
waste  water  to  navigable waters.  No discharge of process
waste water  is also achievable as a new  source  performance
standard for facilities in 21 of the 41 subcategories.

Examination  of the waste water 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 waste water 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.

-------
                         SECTION II

                      RECOMMENDATIONS


The  recommended effluent limitation guidelines based on the
best  practicable  control  technology  currently  available
(BPCTCA)  are summarized in Table II-l.  Based on information
contained  in  Sections  III through VIII, it is recommended
that facilities in 14 of the  41  subcategories  achieve  no
discharge of process waste water.

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 waste water by
1983.

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 four
subcategories,  new   source   performance   standards   are
identical   to   BPCTCA   and   BATEA  recommended  effluent
limitations.

-------
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

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
1X6
GOLD ORES
Mines
Mills
{CyankJation Process
Amalgamation Process
Flotation Process
• Gravity Separation

X
IX-7
IX-8
1X9
IX 10
SILVER ORES
Mine*
Mills
{Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation

X
IX 11
IX 12
IX 13
IX-14
BAUXITE ORE
I
IX-15

-------
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 DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mines/Mills
Mills (
> 6,000 metric tontt/yeer
< 5,000 metric tonst/year
> 5,000 metric tonsVyear by Physical Processes
> 5,000 metric tons* /year by Flotation
Leaching


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

X
X
IX-22

ANTIMONY ORES
Mines
Mills -
Flotation Process

X
IX-23

BERYLLIUM ORES
Mines
Mills
X
X


PLATINUM ORES
Mines or Mine/Mills
IX-24
RARE-EARTH ORES
Mines
Mills -
Flotation or Leaching
X
X


TITANIUM ORES
Mines
Mills 
-------
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 < Pnvsical/Cnemical Separation
| Magnetic and Physical Separation

X
X-1
X-2
COPPER ORES
Mines / Open-Pit, Underground, Stripping
} Hydrometallurgical (Leaching)
{Vat Leaching
Flotation
X
X
X
X-3

LEAD AND ZINC ORES
Mines
Mills

X
X-4

GOLD ORES
Mines
{Cyarudation Process
Amalgamation Process
Flotation Process
Gravity Separation

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

X
X
X
X-6
(Same as BPCTCA)
BAUXITE ORE
Mines I!
X-7

-------
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
EFFLUENT
ZERO LIMITATIONS
DISCHARGE RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mine/Mills
Mills /
> 5,000 metric tonst/year
< 5,000 metric tons^/year
> 5,000 metric tons* /year by Physical Processes
> 5,000 metric tonsVyear by Flotation
Leaching
X-8
(SameasBPCTCA)
X-9
X-10
X 11
MERCURY ORES
Mines
Mills <
II X'12
Gravity Separation II X
Flotation Process II X
URANIUM, RADIUM, VANADIUM ORES
Mines
Mills \
Acid or Acid/Alkaline Leaching
Alkaline Leaching
X-13
X
X
ANTIMONY ORES
Mines
Mills
Flotation Process
(SameasBPCTCA)
X
BERYLLIUM ORES
Mines
Mills
X
X
PLATINUM ORES
Mines or Mine/Mills
(SameasBPCTCA)
RARE EARTH ORES
Mines
Mills -
Flotation or Leaching
X
X
TITANIUM ORES
Mines
Mills 1
Electrostatic/Magnetic and Gravity/Flotation Processes
Physical Processes with Dredge Mining
(SameasBPCTCA)
X
(SameasBPCTCA)

-------
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
j Physical/Chemical Separation
| Magnetic and Physical Separation

X
(SameasBATEA)
(Same as BATEA)
COPPER ORES
Mines
Mills
{Open-Pit, Underground .Stripping
Hydrometallurgical (Leaching)
f Vat Leaching
< Flotation
X
X
X
(SameasBATEA)

LEAD AND ZINC ORES
Mines
Mills

X
(Same as BATEA)

GOLD ORES
Mines
Mills
{Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation

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

X
X
X
(SameasBATEA)
(Same as BPCTCA)
BAUXITE ORE

(Same as BPCTCA)

-------
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
Mino/MilK
Mills |
> 5,000 metric tons'/veer
< 5,000 metric tonst/year
> 5,000 metric tons^/year by
> 5,000 metric tonsVyear by
Leaching

Physical Processes
Flotation


XI-1
XI-2
XI-3
(Same as BATEA)
MERCURY ORES
Mines
Mills <
Gravity Separation
Flotation Process


X
X
(Same as BPCTCAl

URANIUM, RADIUM, VANADIUM ORES
Mines
Mills j
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 as BPCTCA)
RARE-EARTH ORES
Mines
Mills -
Flotation or Leaching

X
X


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

X
(Same as BPCTCA)
(Same as BPCTCA)

-------
                        SECTION III

                        INTRODUCTION
PURPOSE AND AUTHORITY

The United States Environmental Protection Agency  (EPA)   is
charged under the Federal Water Pollution Control Act Amend-
ments  of  1972 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
lr 1983, 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

-------
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
categories.  Publication of an amended list will  constitute
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 will be amended when proposed regulations  for  the
Ore  Mining  and  Dressing  Industry  are  published  in the
Federal Register.

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

The guidelines in this document identify, in  terms  of  the
chemical,   physical,   and  biological  characteristics  of
                          12

-------
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
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,  waste  water  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.

Data Base

Each  SIC category was first examined to determine the range
of activities incorporated by the  industry  classification.
                          13

-------
 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
 waste   water   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.

 Categorization 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 waste water 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 waste  waters in the plant, and  (2)  the
 constituents    (including   thermal)  of  all  waste  waters,
 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
                          14

-------
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,  waste  water  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  waste  water  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
waste  waters  which  might  be  expected  in effluents from
mining and milling operations.  In addition  to  examination
of    candidate   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.

Cost 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
                           15

-------
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
technologies.


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
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
                           16

-------
presently employed in the industry, but included also  those
processes in pilotplant 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
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-
                           17

-------
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
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, CaWCW).
                           18

-------
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 (FeTiO_3) and monazite
(a rare-earth phosphate).  They consist of a helical conduit
(usually, of five turns) about a vertical axis.  A slurry of
ore is fed to the conduit at the  top  and  flows  down  the
spiral  under gravity.  The heavy minerals concentrate along
the inner edge  of  the  spiral,  from  which  they  may  be
withdrawn through a series of ports.  Wash water may also be
added  through  ports  along  the  inner edge to improve the
separation efficiency.  A single spiral may,  typically,  be
                          19

-------
used  to  process 0.5 to  2.4 metric tons  (0.55 to  2.64 short
tons) of ore per hour; in large plants, as many  as  several
hundred spirals may be run in parallel.

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
"HuntingtonHeberlein"  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,
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
                           20

-------
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  si2e  be  small,
that reagents compatible with the mineral to be recovered be
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
                           21

-------
phenomenon,  it  can often be made highly specific, allowing
production of high-grade  concentrates  from  very-low-grade
ore  (e.g., over 95-percent MoS_2 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  (NaS^cOC2H5).  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
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
                           22

-------
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.

     (4)  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).
         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.
                           23

-------
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
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.
                          24

-------
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 of treating an  enriched
gold-   or   silver-   bearing   concentrate.   The  gravity
concentrate is ground for several hours in an amalgam barrel
(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:
                             25

-------
         + 8NaCN + 2H20  + 02 — > 4NaAu (CN) 2 +

 and  subsequently sorbed onto activiated carbon ("Carbon-in-
 J*ulp» process)  or precipitated with metallic zinc  according
 to the reaction:

     2NaAu(CN)2  +  Zn --- >  Na2Zn(CN)   +  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
 SK>lubilized 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
 Jmall 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.

 j?retreatment 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
 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
 ^regnant leach solutions by the methods discussed above.

 Different types of gold/silver ore reguire  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
                          26

-------
"kerosene  or  fuel  oil  to  inhibit adsorption of gold from
•solution.

Other refractory ores  are  those  which  contain  sulfides.
Roasting to liberate the sulf ide-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.

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

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, Cl- , NO3-,  HSOU-,
or  CO3_ —  (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:
    4RX  +  (U02 (804)3)  ---- >    R^UO^ (S(Ht) 3  +  4X-
                          27

-------
 This reaction proceeds from left to  right  in  the
 process.   Typical  resins adsorb about ten percent      e
 mass in  uranium  and  increase  by  about  ?en  percent  in
 density.   m a concentrated solution of the mobile ion (for
 rSveTsId  -  N-^drochl0--  -eld).  the  reaction 'can (* be
 reversed,  and  the  uranxum  values  are  eluted  (in  this
                                     acid) '   m general.  the
                                        for a metaiiic cation
                    Cr++ + >  Mg-n-  >   Na+

 andr because of decreasing ionic radius,  with atomic number:

                    92U  >   42Mo   >   23V

 and the separation of hexavalent 92U cations by ion exchange
 or solvent extraction should prove to be  easier than that of
 any other naturally occurring element.

 Uranium,  vanadium, and molybdenum (the latter being a common
 ore constituent)  almost always appear in   aqueous   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  (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
 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  inter ferent.   Unfortunately,  the  solvents  are easily
polluted by slimes, and complete liquid/solid separation  is
necessary.   lonexchange and solvent-extraction circuits can
                           28

-------
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
shipments  is  shown in Table III-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.
                           29

-------
                  TABLE II1-1. IRON-ORE SHIPMENTS FOR UNITED STATES
                METRIC TONS    LONG TONS
                                                                         METRIC TONS    LONG TONS
  REGION
Great Lakes
Northeastern
Southern
Western
TOTAL U.S.
 REGION
                                  a. QUANTITIES SHIPPED BY REGION
                                                   AMOUNT SHIPPED
                           1968
65,093.239
 3,602,706
 3.474,203
10.566,860
                 82,736,905
64,065,185
 3,545,805
 3,419,333
10,399,972
                                                        1969
72,534,630
 3,453,486
 4,733,087
10,454,364
71,389,050
 3,398,943
 4,658,335
10,289,252
70,180,666
 3,043,857
 5,022,369
10,544,782
69,072,263
 2,995,784
 4,943,048
10,378,242
                                                  AMOUNT SHIPPED
                           1971
                 62,766,873
                  2,859,973
                  4,240,720
                  8,253,243
              61,775,561
               2,814,804
               4,173,744
               8.122,895
                                                       1972
               METRIC TONS    LONG TONS
                                            METRIC TONS   LONG TONS
                                                                         METRIC TONS    LONG TONS
              65,759,357
               2,362,067
               4,032,651
               7,397,815
             64,720,783
              2,324,762
              3,968,961
              7,266,471
              77,504,865
               2,405,456
               3,923,518
               8,462,579
             76,280,787
              2,367,465
              3,861,552
              8,328,925
b. SHIPMENTS FROM GREAT LAKES REGION AS PERCENTAGES OF TOTAL U S 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
487
506
52 3
56 3
61 8
61.7

CATEGORY
=======
Direct Shipping
Coarse Ores
Fine Ores
Screened Ores
Concentrates
Agglomerates

c. PERCENTAGES OF TOTAL U.S. SHIPMENTS

1968
8.2


3.2
28.3
60.3
100.0

1969
='" ' — -
7.0


3.1
27.5
62.4
100.0
YEAR
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
•
y A.

12.9
12.9

71.8
100.0
    SOURCE: Reference 1
                                               30

-------
         TABLE II1-2. CRUDE IRON-ORE PRODUCTION FOR U.S.
                          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
9,575.011
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

REGION
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
                                  31

-------
 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  on  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 oresr 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  mav
 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.                                           ^     y

 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   makes   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
                         32

-------
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  III-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.

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,  briquetting,  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
                           33

-------
     Figure 111-1. BENEFICIATIOIM OF IRON ORES



t t
WASHING JIGG



SINTERING
1
ORE
CRUSHING AND
SCREENING
t
BLENDING
V
CONCENTRATING PROCESSES:
PHYSICAL
1 t
|NG MAGNETIC IfAXr
SEPARAT.ON SEp^^ON
+
AGGLOMERATION
PROCESSES

PELLETIZING NODULIZING


CHEMICAL
|
FLOTATION
1


BRIQUETTING
f
          TO STOCK PILE AND/OR SHIPPING
                   34

-------
Figure 111-2.   IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
               DENSIFYING THICKENER
                   UNDERFLOW
                  CONDITIONERS
                      1


ROUGHER FLOTATION
ROUGHER
CONCENTRATE
(10 CELLS)
^.,1
I
1
FROTH OF
FIRST 2 CELLS
I

CLEANER FLOTATION
CLEANER
TAIL CLE/
1 CONCEIT
ROUGHER
TAIL


1
FROTH OF
XNER FIRST 2 CELLS
JTRATE | _
(8 CELLS)

RECLEANER
1
RECLEANER
TAIL
FLOTATION
1
RECLEANER
CONCENTRATE
(7 CELLS)
I
	 to»
                                                       TO
                                                       TAILING
                                                       BASIN
                                                TOTAL
                                              FLOTATION
                                             CONCENTRATE
                                                  J
                                          TO AGGLOMERATION
                                             (FIGURE 111-4)
                    35

-------
          TABLE 111-3. REAGENTS USED FOR FLOTATION OF IRON ORES
                    represent approximate maximum usages.  Exact chemical composition of reagent



  '•    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 lb/,hort 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)


 *•   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)
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
                                         36

-------
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
                                    37

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

                            I
                COBBER MAGNETIC SEPARATION
             CONCENTRATE
              BALL MILL
                 ±
     CLEANER MAGNETIC SEPARATION
             CONCENTRATE

            	I
            HYDROCYCLONE
         OVERSIZE    UNDERSIZE
                HYDROSEPARATOR
            CONCENTRATE
                 i
     FINISHER MAGNETIC SEPARATION
 CONCENTRATE

     i
     KEI

     I
    FILTER
     T
TO PELLETIZING
 (FIGURE 111-1}
                                              TAILING
TO TAILING BASIN
                           38

-------
  Figure 111-4. AGGLOMERATION FLOWSHEET
CONCENTRATE FILTER CAKE
                 BALLING DRUM
                    SCREEN
              UNDERSIZE  OVERSIZE
                                   FUEL
                                    I
                  AGGLOMERATION FURNACE
                 PELLETS         EXHAUST GASES

                    t                I
              TO STOCK PILE     TO ATMOSPHERE
              AND/OR SHIPPING
                39

-------
 formation  of  pellets or  balls  of  iron  ore  fines,  followed  by
 heating.   (Figure  III-4  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
 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  1840's,
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
                          40

-------
Figure 111-5.   MAJOR COPPER MINING AND MILLING ZONES OF THE U.S.
                            MINING AND MILLING COPPER AS A PRIMARY METAL




                       I:*:*:] 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

-------
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.   Th-  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 229,471 metric
tons  (253,000 short tons) of  copper.   Recovery  of  copper
from  leach  solutions  by  iron precipitation accounted for
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.
                          43

-------
     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 III-8. 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
—
0.07*
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

-------
Figure 111-6. GENERAL OUTLINE OF METHODS FOR TYPICAL RECOVERY OF
          COPPER FROM ORE
                   ORE K01%Cu)
                                          »- OVERBURDEN AND
                                            WASTE DUMPS
                                                       TO MARKET
                                                      (OR REFINERY)
                          46

-------
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  in-situ   (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 extract!on-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.

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-11).   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
                           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
                            LEACHING ZONES

-------
       TABLE 111-11. AVERAGE PRICE RECEIVED FROM COPPER
                    IN THE UNITED STATES
YEAR
1865 - 1874
1907
1910
1915
1917
1920
1925
1930
1932
1935
1940
1945
1950
1955
1960
1965
1970
1972
1973
PRICE IN CENTS PER KILOGRAM (CENTS PER POUND)
LAKE COPPER*
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(18.62-24.62)
66.00- 94.60(30.00-43.00)
66.00- 72.60(30.00-33.00)
74.80 - 83.60 (34.00 - 38.00)
116.6 -132.0(53.00-60.00)
109.7 -114.7(49.88-52.13)
110.3 -159.2(50.13-72.38)
ELECTROLYTIC COPPERt
.
-
-
.
.
.
-
-
.
-
-
.
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
                           50

-------
 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
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
                    51

-------
 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
 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
                           52

-------
       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.
                                    53

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


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


1973

RANK


4
5


7
1

6

2

3
9
8



%


11.94
9.55


4.13
17.27

6.94

17.4

13.32
3.48
3.51


Total
Daily average*
1973
JAN.-DEC. TOTALS
METRIC TONS
7,638
16
51^33
41,216
4,823
245
17343
74,576
379
29355
11,147
73361
17,104
57,474
15,023
15,131
5,768
7365
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
18358
63,367
16364
16,682
6,359
8,672
475,853
1,304
1
1974
JAN. TOTALS
METRIC TONS
600
_.
3,961
3,279
224
_.
1,238
6^89
82
2,361
863
6,961
1,575
7,239
1,130
1,281
528
733
38,644
1,246
SHORT TONS
662

4,367
3,615
247

1,365
7,265
90
2,603
951
7,675
1,737
7381
1,246
1,412
582
808
42,606
1,374
'Based on number of days in month without adjustment for Sundays or holidays.
                                   54

-------
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  111-14,  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.

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) or
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 employs the 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.
                           55

-------
     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
                               ZINC ROUGHER
  FINAL LEAD
 CONCENTRATE
                                        ZINC ROUGHER
                                        CONCENTRATE
  THICKENING
AND FILTRATION
         USUALLY
         RECYCLED
        TO PROCESS
       WATER SYSTEM
                                       AND FILTRATION
                           WATER
                           DISSOLVED SOLIDS
                           SUSPENDED SOLIDS
                           EXCESS REAGENTS
                    TO SUBSURFACE
                      DRAINAGE
          CONCENTRATE
            TO ZINC
            SMELTER
USUALLY RECYCLED
   TO PROCESS
  WATER SYSTEM
                                     56

-------
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  40
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
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   pH
adjustments  with  hydrated  lime (CaOH2) are common.  Other
reagents commonly used and their purposes are:
                            57

-------
          Reagent                       Purpose

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

 The finely ground ore slurry is introduced into a series  of
 flotation  cells,   where  the  slurry is agitated and air is
 introduced.   The  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 tailings  pond.    In
some  cases,  the  coarse  tailings  are  separated  using a
cyclone separator and pumped to the mine for backfilling.
                           58

-------
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
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.   Domestic  production  of  gold  for  1972 was  45.1
million  grams   (1.45  million  troy  ounces).    Of   this,
approximately  76%  come   from  four producers, while the 25
leading producers accounted  for  98%  of  production.   The
domestic production of gold has been on a downward  trend for
the  last   20 years, largely as a result of reduction in the
average grade of ore being mined,  ore  depletions  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  (approximately $70  in  1972  to  nearly
$200  in  1974)  has   stimulated  a  widespread  increase in
prospecting and exploration activity.  As a result  of   this,
the  recovery  of  gold   from  low-grade  ore may now become
economically  feasible, and an increase in  production   might
be expected in the near  future.

Mining  Practices.  Gold  is mined  from  two types of  deposits:
placers   and   lode or  vein deposits.   Placer mining consists
of excavating  gold-bearing   gravel   and   sands.   This is
currently   done  primarily by  dredging but, in  the  past, has
included hydraulic mining and  drift  mining  of buried placers
too deep  to  strip.    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
                           59

-------
 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,  sometimes in  connunction  with
 amalgamation.

 Prior  to 1970,  amalgamation  was the process  used to recover
 nearly 1/4 of the gold produced   domestically.    Since  that
 txme,  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 0.395 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 54.695  in 1972.

 Current practice for  the amalgamation process  (as 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  g9ld/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.

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
                          60

-------
Figure 111-9. CYANIDATION OF GOLD ORE: VAT LEACHING OF SANDS
         AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
                      ORE
                                           TO REFINERY
     TO SMELTER
                         61

-------
 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 111-10)   The

 SiSkeners?    10n   ±S  C°llected  *»  Banting  frU,  the
 Cyanidation  of  slimes  generated  in  the  course  of  wet
           ±SK cufrently  bei"9  done by a recently developed
          carbon-in-pulp  (FigUre  II:[-9).   The  slimes  are
              a^CyanXdS  solution  in  large  tanks, and the
             gold cyanide is  collected  by  adsorption  onto
 activated  charcoal.    Gold  is  stripped7 from the charcSa?
 using a small  volume  of  hot  caustic;   an  electrowinning
 process  is used for  final  recovery of the gold in the SI?
 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 go Id- containing minerals (Figure III-
 11).   The  single  operation  currently  using  this  method
 further processes the tailings from the flotation circuit  bv
 the   agitation/cyanidation   method  to recover the  residual
 gold  values.

 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  trov
 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
                         62

-------
Figure 111-10. CYANIDATION OF GOLD ORE: AGITATION/LEACH PROCESS
      ORE
     J_
    CRUSHING
    GRINDING
  CONDITIONING
COUNTERCURRENT
   LEACHING IN
   THICKENERS
  PRECIPITATION
  OF GOLD FROM
  LEACHATE BY
  ADDITION OF
  ZINC DUST
       I
  COLLECTION OF
  PRECIPITATE IN
   FILTER PRESS
       1
   PRECIPITATE
   FILTERED AND
    THICKENED
   TO SMELTER
                        REAGENTS (CN)
BARREN
 PULP
TAILING
 POND
TAILING-POND
   DECANT
  RECYCLED
      BARREN SOLUTION
          RECYCLED
                           J
                            63

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

       ORE
    CRUSHING
    GRINDING
   CONDITIONING
       i
    SELECTIVE
     FROTH
    FLOTATION
  CONCENTRATE
    FILTERED
  AND THICKENED
   TO SMELTER
                         FLOTATION CIRCUIT
                             TAILINGS
-fr
                  REAGENTS (CN)
                           LEACHING IN
                            THICKNERS
               .BARREN
                 PULP
TO TAILING
POND
                       PRECIPITATION OF GOLD
                         FROM LEACHATE BY
                       ADDITION OF ZINC DUST
                               I
                          COLLECTION OF
                          PRECIPITATE IN
                           FILTER PRESS
                       PRECIPITATE FILTERED
                          AND THICKENED
                               T
                           TO SMELTER
                            64

-------
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
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 III-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)12Sb5S13.    A   tetrahedrite    concentrate    contains
approximately  25  to  32%  copper in addition to  the   25.72  to
                             65

-------
   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
                                66

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

YEAR

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
*Less than 1/2 unit


SOURCE: REFERENCE 2
                           67

-------
 Figure 111-12. RECOVERY OF SILVER SULFIDE ORE BY FROTH FLOTATION
                                 ORE
                                NO. 1
                          FLOTATION CIRCUIT
                                NO. 2
                         FLOTATION CIRCUIT
     RETREATMENT
        CIRCUIT
                       PYRITE
                     CONCENTRATE
                               NO. 3
                         FLOTATION CIRCUIT
                                I
       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
                     FINAL PYRITE
                    CONCENTRATET
             CONTAINS 3.43 KILOGRAMS PErt
              METRIC TON (100 TROY OUNCES
              PER SHORT TON)
                              68

-------
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.

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.

Less than 1 percent of the current  domestic  production  of
silver   is   recovered   by   amalgamation  or  cyanidation
processes.  These  processes  have  been  described  in  the
discussion of gold ores of this report.

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  (2,045,344
tons) of ore in 1973.  Both operations are  associated  with
bauxite   refineries   (SIC  2819) ,  where  purified  alumina
(A12O_3) 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.
                           69

-------
 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  crusSng?  ore
 blending,  and  grinding  in  preparation  for refining.  in
 1972, less than 10 percent of the bauxite used  for  primarv
 aluminum  production  was  of  domestic  origin    with  the
 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

 rS?i"rC:Lent  P°  meet  ?reS6nt  needS  °f the nine domestic
 ™i? rieS'.^  Re°ent  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 ton)
 per  day    pilot    plant    which    converts    alunite
 K2A16(OH)12(SO«)4,   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

 nJS"^   5r°m   % PSak  year   in  1970'  and  Preliminary
 production figures for  1974 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
                            70

-------
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
                             71

-------
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 or 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
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  5  ).   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.
                           72

-------
           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 6
 ^Reference 3
••Reference 7
tt
  Reference 5
                           73

-------
A~    ^   "T.I-18  shows,  molybdenum  mining  and   milling
constxtute  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  world's
:uc  'bde-ium supply, with two mines doir.nating  the  industry.
, L c    ,vio  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
least  one major new operation in Colorado is expected to be
in operation soon.

The  only  commercially  important  ore  of  molybdenum   is
molybdenite,  MoS2.  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  7).   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 7).  Projected  demand  for
tungsten  will exceed supply before the year 2000 at present
prices, and production from currently inactive deposits  may
be anticipated (Reference 7).

Commercially  important  ores  for  tungsten  are  scheelite
(CaWOU_)  and the wolframite series,  wolframite ((Fe, Mn)W04J,
ferberite  (FeWOjf) ,  and  huebnerite  (MnWO^.     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
                          74

-------
about 0.6  percent  tungsten,  and  concentrates  containing
about  70  percent W0_3 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.

Manganese  and  nickel  ores  are each recovered at only one
active operation in the U.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 ore milling are mixed
with effluents from an on-site  smelter  and  with  seasonal
mine-site  runoff.  Water-quality impact from the mining and
milling of these  two  metals  is  thus  presently  minimal.
Future  production  of manganese and nickel, however, may be
expected 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
dominate   the   world  manganese  industry  (Reference  8).
Additional uses include material for battery electrodes  and
agents  for  impurity removal in glassmaking.  Domestic pro-
duction 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
7).   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.

The U.S. Bureau of Mines divides manganese-bearing ores into
three classes  (Reference 7):

    (1)   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
of the total production of 43,536 metric tons (48,000  short
tons)   was in the form of true manganese ores (Reference 7).
Future domestic production is likely on a significant  scale
from manganiferous ores — particularly, on the Cuyuna Range
in  Minnesota,  where  preparations  for  the  resumption of
                            75

-------
 production are  currently  underway.   This   area,   although
 currently  quiescent,  accounted  for  85  percent of  domestic
 production in 1971  (Reference  7).

 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  9, 10, 11, 12), and ammonium carbamate   leaching
 (Reference   13).    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  lateritic deposit is currently being
 mined.   Some sulfide nickel  ore  deposits  with  commercial
 possibilities  have  been found in Alaska (Reference 2).   If
 they are developed, 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 14).

 There  are   no  mines  or mills currently active in the U.S.
 producing ores or concentrates of chromium,   cobalt,  colum-
 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  under
 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),
 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  near
                          76

-------
future.   At these sites, as at essentially all sites around
the world, cobalt is  a  coproduct  or  byproduct  of  other
metals,  and  the  production rates and world price of these
other metals, particularly copper and nickel, exert  primary
influence   on  the  cobalt  market   (Reference  5).   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 5).

Columbium  and  tantalum  concentrates have in the past been
produced at as many as six sites in the U.S. (Reference 15),
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 5).   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
                           77

-------
the same type of process used for vanadium.  There  is  also
some similarity to tungsten, molybdenum, and chromium.

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 5).  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.
                           78

-------
while  the lighter gangue is kept suspended by the agitation
and is removed with the overflow.  Often, a bed  of  coarser
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 5).
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 16).

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  O.OOU  to  0.08 inch)  (Reference 5  ).  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 17).

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 5 )  or the separation  of
cassiterite  and  wolframite  (Reference 18).   Although they
                           79

-------
are both most frequently implemented as dry processes,  wet-
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 111-14 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  18   ).     its
specificity also  allows separation of different ore minerals
(e.g.,  CuS  and   MoS_2)   where  desired,   and operation with
minimum reagent consumption  since  reagent  interaction  is
typically  only  with the particular materials to be floated
or depressed.
                              80

-------
Figure 111-13. GRAVITY-PLANT FLOWSHEET FOR NIGERIAN COLUMBITE
                                                        -OVERFLOW-
                                                        -OVERFLOW-
                                                                           f T°
                                                                           WASTE
— »J


15-cm (6-in.)
GRAVEL
PUMP

CYCLONES
{OPTIONALI
*
4
SHAKING TABLES
                                                            - OVERFLOW -
                                                        •MIDDLINGS
                                                           -MIDDUNGS-
                                                           I
                                                      -HEADS-


                                                        - OVERSIZE-
                           (BLED PERIODICALLY)-
                           SOURCE: REFERENCE 19
                                  81

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


J DREDGE
1
HEAVY MINERAL
CONCENTRATE
t
[_ STORAGE

1

o
L o^cc™ j 	 -q nv,^™,i.L| -T """""=""= [""^STORAGE







1 | * ^ WASTE
TO WASTE-* — SLIME

TO -*- QUARTZ «
STORAGE ^^

STORAGE"^" GARNET «
WASTE -*-TAILINGS-*
-4 	 CLASSIFIER L>JATTRITIC


t
INDUCED-ROLL
MAGNETIC SEPARATOR
t_

)NER ^ SEP A^ATO^I -MCLASSIFIER ~»- DRYER

*
MAGNETIC SEPARATOR I
(LOW-INTENSITY)
1
| SCREEN • - . n

INDUCED-ROLL
MAGNETIC SEPARATOR
t | 	 1


r 	 NONCONDUCT
ELECTROSTATIC
SEPARATOR
1
NONCONDUCTORS
. INDUCED-ROLL
MAGNETIC SEPARATOR
T
. CROSSBELT MAGNETIC
SEPARATOR
1 1 MID
NONMAGNETICS
SCREEN J

[ FURNACE |
ELECTROSTATIC
SEPARATOR
ORS 1 1 — CONDUCTORS 	 1
[ SCREEN |
< 28 MESH '

CROSSBELT MAGNETIC
SEPARATOR
DLINGS 	 *-
r
CROSSBELT MAGNETIC ' I 	 1
SEPARATOR 	 ' 	 ILMENITE f—
> 35 MESH < 35 MESH 1 NONMAGNETICS
4 X NONMAGNETICS I
T T ALTERNATE | _ f
FEE

| WET TABLE j 	 DRIED CONC

r
AIR TABLE
	 *
CONCENTRATE
.ENTRATE 	 M
CROSSBELT MAGNETIC 1 	 NONMAGNETICS
SEPARATOR J AND MIDDLINGS *"^
I" — ' L- 	 1
fc TO
STORAGE
	 ^_ TO
STORAGE
| MONAZITE | | EUXENITE 1 COLUMBITE
                               TO STORAGE
                             82

-------
   Figure 111-15. REPRESENTATIVE FLOW SHEET FOR SIMPLE GRAVITY MILL
                            MINING
                              ORE
   TO
WASTE
TAILS -
                         GRINDING AND
                           CRUSHING
                              I
                           SCREENING
                              FINE
                       MIDDLINGS
                                            COARSE
                                    - HEADS
                            •TAILS-
                                                   MIDDLINGS
                                          SHAKING
                                           TABLES
                                             I
                                               • CONCENTRATE
                                   83

-------
 Details  of  the  flotation  process — exact  suite   and   dosage
 of   reagents,   fineness   of  grinding,  number   of regrinds,
 cleaner-flotation  steps etc., — 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 18).  Collect-
 ors  are most often alkali xanthates with two to  five   carbon
 atoms — for example, sodium ethyl xanthate (C2H50 . NaCS2).
 Frothers  are   generally  organics  with   a soluble hydroxyl
 group and a  "non-wettable"  hydrocarbon   (Reference   17  ).
 Pine   oil   (C6H12OH),   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-
                             84

-------
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
Nokes 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
Oleic 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*
	 	

0.04 to 0.8
 •IN USE AT ONLY ONE KNOWN OPERATION, NOT NOW ACTIVE
                         85

-------
             Figure 111-16. SIMPLIFIED  MOLYBDENUM MILL FLOWSHEET
                                          MINING
                                           ORE

                                            I
                                        CRUSHINGS,
                                       WEIGHING, AND
                                        SCREENING
                                           BALL
                                          MILLS
                                         CYCLONES
                                        OVERFLOW
                                      ROUGHER FLOAT
                                        MIDDLINGS
                       	MIDDLINGS —
                                                     	UNDERFLOW
                                                         -CONCENTRATE-
              SCAVENGER
               FLOAT
            (4 STAGES WITH
             REGRIND AND
           INTERNAL RECYCLEI
OVERFLOW
—]
       -UNDERFLOW-
     OVERFLOW
    -RECLAIM —
      WATER
TO TAILINGS
   POND
                                                     — CONCENTRATE
                                                          MIDDLINGS -
                                       CLEANER
                                        FLOAT
                                     (6 STAGES WITH
                                      REGRIND AND
                                   INTERNAL RECYCLE)
                                                                                     — TAILS-
               TAILS

               _*_
                                                                        CONCENTRATE
              CYCLONES
                                                    [ -
                             UNDERFLOW -
                                                                        MOLYBDENUM
                                                                          PRODUCT
                                            86

-------
tion  is  shown  in  Figure  III-17.   Pyrite  flotation and
monazite flotation are accomplished at acid pH (U.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 Ib/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 O.U 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,
depending  primarily   on the accessory  mineralization in the
                             87

-------
Figure 111-17. SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM

' 	 CONCENTRATE 	
	 LIGHT TO TAILS 	
MONAZITE
	 CONCENTRATE 	
TO TAILS

CRUSHING
(3 STAGES)

28% + 3 MESH
1
GRINDING
BALL MILLS
36% + 100 MESH
^l
+
FLOTATION
t
FLOTATION
I
96% OF MILL FEED
*
GRAVITY
HUMPHREY'S SPIRALS
*
PYRITE
FLOTATION
1
TAILS
*
TABLES
*
MONAZITE
FLOTATION
1
MAGNETIC
SEPARATION
1
NONMAGNETIC
TIN CONCENTRATE i r


1
CONCENTRATE
1
CLEANER
FLOTATION
(4 STAGES)
1
DRYING
*
MOLYBDENUM
CONCENTRATE
(93% + MoS2)
                                                 -TAILINGS-
                      MAGNETIC TUNGSTEN
                         CONCENTRATE
                     88

-------
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
                             89

-------
 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  IH-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  lb/short ton)

 Soap  skimmings                no  kg/metric ton
                                   (80 lb/short ton)

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

 sol                            5  kg/me trie 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 cleanina
 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
the flowsheet for a Canadian pyrochlore  (NaCaCb206F) mill in
Figure III-19.                                 	
                          90

-------
Figure 111-18. SIMPLIFIED FLOW DIAGRAM FOR SMALL TUNGSTEN CONCENTRATOR
                  ORE
                   i
                                 SULFIDE
                                FLOTATION
                                 CYCLONE
                                75% SANDS
                                   i
                                 GRAVITY
                                 TABLES
                              TAILINGS
                                              25%
                                            SLIMES
                                                            OVERFLOW
THICKENER
SCHEELITE
FLOTATION
                                                       HCI LEACH
                                                     (15 TO 20% OF
                                                       FRACTION)
                                                    TUNGSTEN
                                                   CONCENTRATE
                                 91

-------
Figure  111-19. MILL FLOWSHEET FOR A CANADIAN COLUMBIUM OPERATION
                           CONCENTRATE
                                     ~L
                                    MAGNETIC
                                    SEPARATION
                                    DESLIMING
                                    UNDERFLOW

                                       I
                                  BULK FLOTATION
                                   CONCENTRATE
                                    DESLIMING
                                    UNDERFLOW
                                    MAGNETIC
                                   SEPARATION
                                 PRIMARY CLEANING
                                   (FIRST STAGE)
                                   CONCENTRATE
                                 PRIMARY CLEANING
                                  (SECOND STAGE)
                                  CONCENTRATE
                                SECONDARY CLEANING
                                  (THREE STAGES)
                                  CONCENTRATE

                                      I
                                                         ~l
                                                 - TAILS -
                                    REVERSE
                                   FLOTATION
— CONCENTRATE -
                    MAGNETITE
                                                                                       TO
                                                                                     STOCKPILE
                             —.      OVERFLOW
                              I      (CALCITE)
                     SPIRALS
                                                                     TAILS
                                                                    TABLING
                                                                  CONCENTRATE
                                                                                 CONCENTRATE
                                  SOURCE:REFERENCE 5
                                       92

-------
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   legs   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 6  ), 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 vanadium's 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% P205,  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^CO_3) and salt, and roasted at 750 to
800 degrees  Celsius   (1382  to  1472  degrees  Fahrenheit).
Phosphorus,  vanadium,  and chromium are converted to water-
soluble trisodium phosphate, sodium metavanadate,  and sodium
                           93

-------
 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  V205  (redcake) by  acidification,  and  chromium
 is  precipitated  as   lead  chromate.   By  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  13).   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  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 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
                           94

-------
Figure 111-20. FLOWSHEET OF TRISTAGE CRYSTALLIZATION PROCESS FOR
            RECOVERY OF VANADIUM, PHOSPHORUS, AND CHROMIUM
            FROM WESTERN FERROPHOSPHORUS
        FERROPHOSPMORUSj

                        SETTLING AND OECANTATION
                                         SOLIDS
                                                  WATER WASH
             PRIMARY CRYSTALLIZATION
              PRIMARY CENTRIFUGING
                                                    WASH
                                                    LIQUOR
        PREGNANT
        SOLUTION
                             CRYSTALS
                                    -DISSOLUTION-

                                        J_
                                | SECONDARY CRYSTALLIZATION [
          RED CAKE  MOTHER LIQUOR
[ FUSION j |
| FILTRATION |


VANADIUM PRODUCT
Ci«
T
WA
TO
	 • ^STOCKPlLC
3HI2 SOLUTION mnnf
j i
[ FILTRATION J
1 \ 	 ^_ SOLUTION
	 I 	 TO WASTF
0 CHROMIUM ^ TO
>TE PRODUCT ' STOCKPILE

-------
    Figure 111-21. ARKANSAS VANADIUM PROCESS FLOWSHEET
                        1.5 -2.0% V2O5
                             t
       6-10%
       NaCL
TERTIARY
 AMINES
                          GRINDING
                         PELLETIZING
                          ROASTING
                 850°C (1562°F)
                           NaVO,
                             £
H2S04


•
LEACHING AND
ACIDIFICATION
                                      pH 2.5 - 3.5
                         (Na6V10028
                   SODIUM DECAVANADATE)
                            I
SOLVENT EXTRACTION
                                                 NH4OH
                       PRECIPITATION
                            I
                          NH4V03

                        PRECIPITATE
                            I
                        CALCINING
                       V2O5 PRODUCT


                       96

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

                       RAW SIZED ORE -- 1.9 cm (0.75 in.)
                            REDUCING FURNACE
                                                     -CO,
                                   i
                            GRINDING (30 MESH)
WEAK Mn
SOLUTION
                                   i
                 LEACHING
           (TWO 11.356- i (3000-GAL))
              REACTION TANKS
                 IN SERIES
                                   I
                           9.14-m (30-ft) CLARIFYING
                                THICKENER
7.6-m (25-ft) COUNTERCURRENT
    WASHING THICKENERS
               LIVE
              STEAM
  SLURRY
   STILL
NH4NH2CO2
 TAILINGS
                                        NEW LEACH LIQUOR
                                           LEACH LIQUOR
                                           REGENERATION
                TW011.356-X
                  (3000-GAL)
               PRECIPITATION
              TANKS (IN SERIES)
-NH,
                                                           T   r
                                   MnCO,
                                CLARIFYING
                                THICKENER
                                   MOTHER
                                  " LIQUOR "
                 70% SOLIDS
                               ROTARY DRYER
                                    NH4NH2CO2
                                  PRODUCT
                                  97

-------
  TABLE 111-21. DOMESTIC MERCURY PRODUCTION STATISTICS

CATEGORY


Production in metric tons
(flasks)
Dollar value (thousands)
YEAR

1969

1.029
(29,640)
$14,969

1970

948
(27,296)
$11,130

1971
56
621
(17,883)
$ 5,229


21
253
(7,286)
$1,590


6



SOURCE: REFERENCE 2
                      98

-------
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  low  of  3,000  flasks  (104 metric tons, or 115 short
tons) by 1985.

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) .

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 is 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.   Flotation  of
mercury  ore  has  not  been used to date in the  United  States.
However,  an operation scheduled to  begin  in  Nevada  later  in
1975  will  concentrate  mercury  ore  by  flotation.   This
concentrate will  be  furnaced,   and annual  production  of
mercury   from the  operation  is   expected   to  reach  20,000
flasks  (695 metric tons, or  765 short tons).

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
                           99

-------
 always  been  obtained  from  uranium  ores.  In the oast 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
 orTmarltv  for a P°llutant in the »*stes.   Uranium is  mLel
 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 Is a
 ferroalloying metal and,  in the form  of  its  oxide,  as  a
 S?SiS A •  V^adium  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),  o-
 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

 asnh^J10no With-.Karb°naCeOUS   fossils-   i.e.,  lignite and
 asphalts,  ores  with  a grade in excess of a  fraction  of  a
 oS^be   uranium are  rare (80% of  the  industry operates with


 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
 uranium  for  the  nuclear industry.  The processes  of  crushing
 and  grinding,   conventionally  associated  with  a mill, are
 intimately connected with the  hydrometallurgical  processes
 tnat   yield   the  concentrate,  and  both processes normally
 share  a waste water  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  auite
 profitable.                                             ^

 The uranium oxide concentrate, whose grade is usually quoted
 in  percent  of  U308   (although  that  oxide figures in the
 assay,  rather than in the product), is generated by  one  of
 several hydrometallurgical processes.   For purposes of waste
water 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
                           100

-------
         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  0.14  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 UOJ2, 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,
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  (U(VI)) compounds, notably carbonates.  Typically,   a
primary  deposit  is  weathered  by  oxidized water, forming
hydrated oxides of uranium  with  compositions  intermediate
between UO^ and UO^.  The composition U.30§ — i.e., U02.2U03
    is  particularly stable.  The process occasionally  stops
at gummite  (U0_2. H20) , an orange or red,  waxy  mineral,  but
usually   involves   further  oxidation   and  reactions  with
alkaline   and   alkaline-earth   oxides,   silicates,   and
phosphates.   The transport leads to  the  surface uranium ores
of  arid  lands,  including carnotite  (K2(U02)2(VOU)2.3H20),
uranophane        (CaU2Si2011.7H£0) ,       and       autunite
 (Ca(U02J2.(PO^)_2.10-12H20)  and,  if  reducing conditions are
encountered,  to  the  redeposition  of   U(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.

-------
TABLE 111-22. ISOTOPIC ABUNDANCE OF URANIUM
ISOTOPE
U238
U 235
U234
U236
HALF-LIFE (YEARS)
451 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)
                102

-------
 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  mine
 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  U.S.
 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  respect  to
 modern  grade that   makes  it   profitable to   mine  a  given
 deposit  by open-pit methods at one point and by underground
 mining  at others.

 Exploration is conducted initially with airborne and surface
 radiation sensors that delineate promising  regions  and  is
 followed  by   exploratory  drilling, on a 60-m (200-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   specified
 in   terms   of   ore  that  can yield uranium at  $18 per kg  (2.2
 Ib) , a  price paid by the government for  stockpiling.   Recent
 increases in price  and the possibility of increased  uranium
 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  Torts)
New Mexico         137,108             66             69
Wyoming             28,300             14             14
Utah and
Colorado           11,400              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
                           103

-------
 Texas gulf coast,  dominate  the  industry.   in  1974
 Mexico provided 43 percent and Wyoming 32 percent of
 production.   In  1974, the U. S. production was 7.1
 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
 r££f£  X t0° XKW f°5 economic recovery of uranium, which is
 recovered as a byproduct only in Florida.  Vanadium, in ores
 r2aJ    ^ co?tain Cranium 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
 ?«r^?S  Sf  *  ?anuary 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
 occasionally  recycled through the  mine to recover values  bv
 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-leachinq)
 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
"^cn  J°nS   (495  Short  t0ns> 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.
                           104

-------
TABLE 111-23. URANIUM MILLING ACTIVITY BY STATE, 1972

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
Kin nc MII i c


3
7
3
2
3
1
1
20
                       105

-------
           CJu_§hirK[,  ^nd Roasting.  Ore from the  mine tends
c& be quite viriaJle 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 "alu   j the proper source
?nd  to  rchieve  uniform  ~r-\er and .*•   blending to a:iire
aniform  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
(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 (CaCO_3)  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   	>  2UOJ3

         2W)3 + 2H2SOU. + 5H_2O	> 2 (UO2SOJ4)  . 7H2O
                            106

-------
Uranyl  sulfate   (UO2SO4)  forms a complex, hydrouranyl tri-
sulfuric acid  (H^UO2(SOJ4) .3) r 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 111-23).  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:

                     2UO2 + 02 	> 2U03
                        (oxidation)

      3Na2(CO.3) + UO3 + H2O	> 2NaOH + Na4_ (U02J (C03)3
                         (leaching)

               2NaOH + CO2	> Na2C0.3 + H20
                    (recarbonization)

   2Na4JU Na.2U2O7   + 6Na2_CO3_ + 3H20
                     (precipitation)

The efficient utilization of water  in  the  alkaline leach
circuit  has  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  waste  water  but are often ignored.
Purification to remove sodium ion is necessary both to  meet
the  specifications  of  American uranium processors  and for
                            107

-------
         Figure 111-23. PACHUCA TANK FOR ALKALINE LEACHING
                                                        LEACH
                                                       AIRLIFT
=////=:////=
                          108

-------
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.,  Na2lJ2p2 = Na_2O +
2UO3I.

The process of generating UO2:  fuel pellets  from  yellow-cake
feed  involves reduction by gaseous ammonia at  a  temperature
of a few hundred degrees C.  At   this temperature,  ammonia
thermally  decomposes into hydrogen, which reduces the UO3_
component to UO_2 and  nitrogen  (which acts as an   inert  gas
and reduces the  risk  of  explosion in and around the reducing
furnace).   With sodium diuranate  as   a  feed,  the process
results in a mix of UOJ  and Na_2O  that is difficult to  purify
(by water leaching of NaOH) without  impairing  the ceramic
qualities  of  uranium dioxide.   When, in contrast, ammonium
diuranate is used as  feed,  all byproducts are  gaseous,   and
pure  U Na2SO^  + 3H20  + 2 (UQ2) S04

and  the  uranium  values   are precipitated  with  ammonia and
filtered, to yield a  yellow cake  (ammonium diuranate or UOj!)
that is low in or free of sodium.

         UO2SCW  + H.2O +  2NH3_	> (NHU) .23(34  + UOJ3

The reactions leading to this product are   interesting for
their  byproduct—namely, sodium  sulfate.  The  latter,  being
classed approximately in the  same   pollutant  category  as
sodium   chloride,  requires  expensive   treatment for its
removal.  Ammonium-ion discharges  which  might result from  an
ammonium carbonate leaching circuit   that would   yield the
desired  product immediately  are viewed with  more concern,
even though there is  a demand for ammonium  sulfate ;o   fer-
tilize  alkaline southwestern soils.  Ammonium  sulfate  could
be generated by  neutralizing the  wastes  of the  ammonium loop
with  sulfuric   acid  wastes  from  acid leaching  wastes.
Opponents  of a  tested ammonium process  argue that nitrites,
an intermediate  oxidation product  of  accidentally  discharged
                           109

-------
ammonium ion, present a present health  hazard  more  severe
than that from sulfate ion.

Vanadium  Recovery.   Vanadium, found in carnotite  CK2(\3O2)2
(VO4J 2 . 3H2!O) as well as in  heavy  metal  vanadates—e.g.,
vanadinite   (9PbO  .  3V2O_5 . PbCl_)—is converted  to sodium
orthovanadate (Na^SVO^) , which is water-soluble  by  roasting
with  sodium  chloride  or  soda  ash (Na2CO3).  After water
leaching, ammonium chloride is  added,  and  poorly soluble
ammonium vanadates are precipitated:

      NaSVOU + 3NH4C1 + H2O	> 3NaOH +NIMVO3 + 2NH4OH
                 (ammonium metavanadate)

           Na3VO4 * 3NH4C1	> 3NaCl + (NHU)3VOi»
                 (ammonium orthovanadate)

The  ammonium  vanadates  are  thermally decomposed to yield
vanadium pentoxide:

         2(NH!)!VQ1	>  6NH3   + 3H2O   + V2O5

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 + Fe_2O_3 + 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.
                         110

-------
Concentration and Precipitation.   To 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 0.2%.  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 III-24.

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-
24a) .  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-2Ub,
c).  Sands are often repulped, by the addition of some waste
water  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 waste water 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  III-24b-e)  are
based  on  the same principle:  Polar organic molecules tend
to exchange a mobile ion in their  structure  —  typically,
C1-,   N0_3-r  HSO_4-,  C0_3— (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
                         111

-------
  Figure 111-24. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 1 of 2)
        FROM •
        LEACH
PREGNANT
LEACH LIQUOR

                                            SLIMES
               CLEAR LEACH LIQUOR
               TO COLUMN IX OR SX

                    a) LIQUID/SOLID SEPARATION
                                                         SLIMY PULP TO
                                                         RESIN-IN-PULP IX
                                                       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
                                 112

-------
Figure 111-24. 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
PREGNANT
LEACH
LIQUOR
1
        °0
        Oo
    LOADING
                       SOLVENT
                               LOADED
                               ORGANIC
                            CD-
BARREN  STRIPPED
ELUANT  SOLVENT
                                                            SOLVENT
                    BARREN
                    LIQUOR
                   »
            PHASE
            SEPARATION         STRIPPING

               e) SOLVENT EXTRACTION
                                                    Vy ^PREGNANT
                                                     X-X ELUATE
                 If
                                                         PHASE
                                                         SEPARATION
          PRECIPITATION

 f)  ELUEX PROCESS
                                            RECYCLE
                         BARREN
                         ELUANT

IX

ELUANT
~\ f ^
1 1
/ \

IX

                                                   PREGNANT
                                                     ELUATE
                                  PARTIALLY  STORAGE)
                                  STRIPPED   \     /   LOADED
                                  RESIN       ^ -~     RESIN

                                        g) SPLIT ELUTION
                               113

-------
TABLE 111-24. URANIUM CONCENTRATION IN IX/SX ELUATES
PROCESS
U3O8 CONCENTRATION (%)
Ion exchange
Resin-in-pulp
Fixed-bed IX:
Chloride elution
Nitrate elution
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
                      114

-------
brittle, radioactive, and magnetic, permitting concentration
by magnetic means.  There are some deposits of  consolidated
monazite sands in Wyoming.

Hydrometallurgical  processes are used to separate a thorium
and rare-earth concentrate  from " magnetically  and  gravity
concentrated sands  (Figures 111-25 and 111-26).  Either acid
or alkaline leach processes may be used, but cationic rather
than  anionic  species predominate in the leach, in contrast
with otherwise analogous uranium processes.  Thorium  preci-
pitates  from  sulfuric  acid  solution  at  a  pH below one
(Figure III-27), in contrast to  rare  earths  and  uranium;
this fact, as well as its reduced solubility in dilute mona-
zite    sulfate    solution,   is   utilized   for   thorium
concentration.   The  latter  process,  when   used   alone,
requires  as  much  as  300  liters   (318  qt)  of water per
kilogram  (2.2 lb)  of  monazite  sulfate  and  is  not  very
economical.   When  used  in  conjunction  with neutralizing
agents as a fine control on pH, it is very effective.

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.
                           115

-------
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
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 halflives 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  (Tho;2) .   American
ores  from the North and South Carolinas,  Florida, and Idaho
contain 1.2 to 7 percent ThO^, with a typical value  of  3.4
percent.  Monazite, a phosphate of cerium and lanthanum with
some  thorium  and  some  uranium  and other rare earths,  is
found  in  granites  and  other  igneous  rocks,   where  its
concentration  is  not economically extractable.   Erosion of
such rocks concentrates the monazite sands,  which constitute
about 0.1 percent of the host  rock,  in  beach  and  stream
deposits.    Mining  often  is  combined  with the recovery of
ilmenite,   rutile,  gold,   zircon,  cassiterite,   or   other
materials  that  concentrate  in a similar way.   Monazite is
                           116

-------
X be the mobile ion.  Then, the exchange  reaction  for  the
uranyltrisulfate complex is

         URX +  (U02 (S0.4) 3_)   --->    R4U02(SOJi)l + 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+++   Mg+ +   Na+) and,
because of decreasing ionic radius, with atomic number   (92U
42  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
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, of the pregnant solutions
influence the adsorption  of  heavy  metals.   For  example,
seven  times  more  vanadium than uranium is adsorbed on one
resin at pH 9; at pH 11, the  ratio  is  reversed,  with  33
times  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-24d,  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-24e)
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
                          117

-------
Figure 111-25. 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,
SIUICA, ZIRCON, AND RUTILE)
                    FILTRATE
                 (R.E., U, AND P205)
                    SELECTIVE
                   PRECIPITATION
                    AT pH 2.3
1
FILTRATE

*
PRECIPITATE OF
U AND P2OS
(BYPRODUCT)
       TO WASTE
                    TO SHIPPING
                                                  MONAZITE
                                                    SAND
                                                  GRINDING
                                                  OPERATION
                                                  DIGESTION
                                                 DISSOLUTION
                                               Th, R.E., U, AND P2O5
                                                  SELECTIVE
                                                 PRECIPITATION
                                                  AT pH 1.05
                                                  PRECIPITATE
                                                (Th, R.E.. AND P2O5)
                                         PURIFICATION BY SOLVENT EXTRACTION,
                                          SELECTIVE PRECIPITATION, OR FRAC-
                                             TIONAL CRYSTALLIZATION
                                                 CONCENTRATES
                                                  TO SHIPPING
                                           SOURCE: REFERENCE 20
                                    118

-------
 Figure 111-26. SIMPLIFIED SCHEMATIC DIAGRAM OF CAUSTIC SODA DIGESTION
            OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
            AND RARE EARTHS
          MAIN STREAM
                                                  MONAZITE
                                                    SAND
                                     NaOH
                                H2O
                  I	
              FILTRATE
           (NaOH AND Na3P04)
                 ±
           CRYSTALLIZATION

           	1
                         i
                       RESIDUE
                       (Na3P04)
                     (BY-PRODUCT)
                          1
                       FILTRATE
                     (RARE EARTHS)
                       SELECTIVE
                     PRECIPITATION
                FILTRATE
                                  i
                        1
                     GRINDING
                     OPERATION
                                                  DIGESTION
                                 138°C
                                 (280°F)

              HYDROUS METAL-OXIDE CAKE
                    (Th, U, AND R.E.)
PRECIPITATE OF
 RARE EARTHS
 (BY-PRODUCT)
                TO WASTE

SOURCE: REFERENCE 20
      t
 TO STOCKPILE
                                                   SELECTIVE
                                                 PRECIPITATION
                                                  PRECIPITATE
                                                   (Th AND U)
  PURIFICATION BY
SOLVENT EXTRACTION
                                                       T
  	"—	1    TO
  CONCENTRATES |-»" STOCKPILE
                                   119

-------
 Figure 111-27. EFFECT OF ACIDITY ON PRECIPITATION OF THORIUM, RARE
            EARTHS AND URANIUM FROM A MONAZITE/SULFURIC ACID
            SOLUTION OF IDAHO AND INDIAN MONAZITE SANDS
     100
Q
ui
<
oc
a.
UJ
D

O
                                        IDAHO MONAZITE SAND
                                       D INDIAN MONAZITE SAND
                                       A
20 -
                               ACIDITY (pH)

                 AGITATION TIME:      5 MINUTES
                 DILUTION RATIO:      H2O: SAND = 45:1 TO 50:1
                 DIGESTION RATIO:     93% H2SO4: SAND = 1.77
                 NEUTRALIZING AGENT: 3.1% NH4OH

                 SOURCE: REFERENCE 20
                               120

-------
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 being
Ra228 at 6.7 years.

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 111-23, also
Supplement B).  A general flow chart, to  be  used  in  con-
junction  with  Table 111-26, is presented in Figure 111-28.
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  U_3O8
per year) since 1968.

Table  111-27  shows  uranium production for the period 1968
through 1972, expressed in terms of both  ore  movement  and
UJOJ3 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.
                          121

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

ELEMENT OR
NAME


SYMBOL(S)


HALF-LIFE
ENERGY OF RADIATION
(MeV)
(X.
fl I r
Thorium Series
Thorium
Mesothorium 1
Mesothorium 2
Radiothorium
Thorium X
Thoron
Thorium A
Thorium B
Thorium C
Thorium C'
Thorium C"
Thorium D
90Th
88Ra228(MsTh1)
89Ac228 (MsTh2)
^Th228 (RdTh)
ggRa224 (ThX)
ggRn220 (Tn)
^Po216 (ThA)
82Pb212 (ThB)
83Bi212(ThC)
^Po^lThC')
81TI208 (ThC")
82Pb2°8(ThD)
1.34x 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
-
-
-
0.053
1.55
-
-
-
ft
0.36
2.20
-
1.82
-
-
-
-
r
-
-
-
-
r
—
2.62
-
Uranium Series
Uranium
Thorium
Protactinium

Uranium
Thorium
Radium
Radon
Polonium
Lead

Bismuth
Polonium
Thallium
Lead
Bismuth
Polonium
Lead
92u238 (uu
9QTh234(UX1)
91Pa234(UX2)
7*14.
92u234 (uio
90Th230(.o)
88Ra226
86Rn222
^Po218 (RaA)
82Pb214 (RaB)
O1A
83Bi2U (RaC)
MPo214 (RaC')
81TI210(RaC")
82Pb210 (RaD)
83Bi210(RaE)
84Po210 (RaF)
82Pb206 (RaC)
4.55 x 109 years
24.1 days
1.14 minutes
C
2.69 x 10° years
8.22 x 104 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
5.49
5.99
-

5.50
7.68
-
-
-
5.30
-
-
0.13
2.32

—
-
-
-
ft
0.65

3.15
-
1.80
0.025
1.17
-
-
-
0.09
0.80

—
r
0.19
-
-
r

1.8
-
-
0.047
-
r
-
                 122

-------
           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
5








SOURCE: REFERENCE 21
                             123

-------
Figure 111-28. GENERALIZED FLOW DIAGRAM FOR PRODUCTION OF URANIUM
           VANADIUM, AND RADIUM                         unMmum,
                   MINING
                    1
               ORE TREATMENT
                  LEACHING
                    I
                LIQUID/SOLID
                 SEPARATION
             I"
             I
  ION EXCHANGE
           SOLVENT EXTRACTION
     PATH I
       I
   PATH IH
                 PATHn
               PRECIPITATION
      TO
STOCKPILE
                    1
                            •-H
  URANIUM
CONCENTRATE
VANADIUM
BYPRODUCT
RECOVERY
TO
STOCKPILE
                            124

-------
                    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
U308 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
U3O8 RESERVES*
1000
METRIC TONS
146
185
224
248
248
251
SHORT TONS
161
204
247
273
273
277
•At $18.000 per metric ton ($16,340 per short ton).
                    TABLE III-28. VANADIUM PRODUCTION


YEAR
1968
1969
1970
1971
1972
u.s. V2o5
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
V205 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
_
Ferrovanadium
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

METRIC
TONS
•'.• 	
4,084
172
43
453
1972
SHORT
TONS
4,502
190
47
499

%
86
4
1
9
                                   125

-------
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 28).

Thorium production in the U.S. during 1968  was  100  metric
tons (110 short tons) as was demand, mostly for the chemical
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
metric  tons  (35  short tons)—i.e., was in surplus by 1433
metric tons  (1577 short tons).

M§£ai 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
                          126

-------
       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 CONTENTI
METRIC TONS
    1,179
    1,065
     542
     751
     468
SHORT TONS
    1,300
    1,174
     598
     828
     516
 includes product-on from antimony ores and concentrates and byproduct recovery from silver concentrates.
 tByproduct produced at lead refineries in the United States.
                                         127

-------
antimony content.  Nearly all  of  this  production  can  be
attributed  to  a  single  operation  which is using a froth
flotation process to concentrate  stibnite  (Sb2S3)  (Figure
111-29) .                                         ~

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,
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
(Be3_Al_2(Si6O.18))  was produced in Colorado and South  Dakota.
The   largest   domestic   source  of  beryllium  ore  is  a
bertrandite (Be4Si2O7  (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.
                          128

-------
Figure 111-29. BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
                      MINING
                       ORE
                       i
                     CRUSHING
                    GRINDING
                       1
                  CLASSIFICATION
     ROUGHER
    FLOTATION
      FROTH
        I
     •TAILS
 •TAILS-
SCAVENGER
 FLOTATION
 CLEANER
FLOTATION
  FROTH
  	I
                             I
                           FROTH
TO
WASTE
                      FILTER
                                        FILTRATE
                       1
                    THICKENER
                                         WASTE
                       I
                      FINAL
                  CONCENTRATE
                   TO SHIPPING
                             129

-------
 A  sulfuric  acid  leach  process  is  employed  to  recover
 beryllium  from  the  Spor Mountain bertrandrite.  This is a
 P^^1?^ary  Process'  however,  and  further  details   are
 wxthheld.  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
 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
 SKSt*,??6^1??'  with Production also from platinum placers.
 Table 111-31  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
 trom  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) .
                         130

-------
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
                          131

-------
 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 III-30.

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

 Byproduct platinum-group metals from gold or copper ores are
 sometimes refined by electrolysis and  chemical  means.   In
 the  Sudbury District of Canada, sulfide ore is processed by
 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
Jersey,  Texas,  Utah,  and 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
 39,  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 (CeFCO_3)  and
monazite (Ce,  La,  Th,  Y) P04_.    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
                          132

-------
Figure 111-30. GRAVITY CONCENTRATION OF PLATINUM-GROUP METALS
                 DREDGE
                (SCREENING,
               JIGGING, AND
                 TABLING)
                 TABLING
                 MAGNETIC
                SEPARATION
CMflOMITE/
MASMITITE
                  DRYING
                SCREENING
                  SIZING

                 11
                  BLOWER
             90% CONCENTRATE
         (PLATINUM GROUP AND GOLD)
                    1
               TO SHIPPING
                 TO
                 WASTE
 TO
' SHIPPING
                 TO
                 WASTE
                         133

-------
 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, bastnaestite  is  mined
 by  open-pit   methods.   The ore,  containing 7 to 10  percent
 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 PEO.

 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 Meixco 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-
                           134

-------
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
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  TiO^;  however,  some  sand  deposits yield  altered
ilmenite  (leucoxene) containing 60 percent or more TiO.2,  as
well as rutile containing 90 percent or more Ti02-
                           135

-------
 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
887,506
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
                         136

-------
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 Ti02 and are  mined  with
floating suction or bucket-line dredges handling up to 1,088
metric  tons   (1,200  short tons)  of material per hour.  The
sand is treated by wet 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
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 40 percent
iron, and rutile has 4 to 10 percent iron.

Tailings from  the HT separators  (nonconductors)  may  contain
zircon  and  monazite   (a  rare-earth mineral).  These heavy
minerals  are  separated  from    the   other   nonconductors
 (silicates) by various 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 111-31.
                          137

-------
      Figure 111-31. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS
WET MILL

DRY MILL
   r
      MAGNETICS
                   NONMAGNETICS
MONAZITE
  TO
SHIPPING
                           FLOWS (ROUGHERS AND CLEANERS)
                      NONCONDUCTORS
                                138

-------
Ilmenite is also currently mined from a rock deposit in  New
York  by  conventional  open-pit  methods.   This  ilmenite/
magnetite ore, averaging 18 percent  T±02,  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 111-32.

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
(ZrSiO_4) , 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, rutile, 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) .
                         139

-------
Figure 111-32. BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
                              MINING
                               ORE
                              JL
                            CRUSHING
                            GRINDING
                               i
                         CLASSIFICATION
                              I
                            MAGNETIC
                           SEPARATION
         I
                 MAGNETICS-
                              i
NONMAGNETICS
      MAGNETITE
         i
              ILMENITE
            AND GANGUE
      DEWATERER
                i
                                                 FLOTATION
                                                  CIRCUIT
                                                 THICKENER
                                                   i
                                   TO
                                 WASTE
               FILTER
                                                   i
                                                  DRIER
                                               CONCENTRATE
                                                    *
                                                TO SHIPPING
                           140

-------
                         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 contains 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;

          (14) Facility age;

          (15) Land availability.
                          141

-------
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
extraction  under NRC licensing surveillance  (SIC 1094); 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
                          142

-------
 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.

     (U)   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 SUBCATEGORIES   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 waste water  because  of  differing  water
 quality,  flow  rate or treatability.  Levels of pollutants in
mine  waters  are generally lower or less complex than those
in mill process  waste  waters.   Mine  water  contact  with
finely  divided  ores, (especially oxidized ores) is minimal
                           143

-------
and mine water is not exposed to the suite of process  water
reagents   often  added  in  milling.   Waste  water  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 waste
water and  mill  waste  water  separately,  there  are  some
situations  in which combining the mine waste water and mill
process waste water 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 waste water 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.

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  some  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.
                          144

-------
 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 waste water 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 waste water 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
 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
                          145

-------
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
runoff  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
guality 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
                           146

-------
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 waste water.  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
basis   for   general  subcategorization,  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
                           147

-------
 The  general  geologic  setting determines  the  type of mine—
 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
 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.
                        148

-------
 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
         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.
                          149

-------
  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

In preparation for selection of  sites  for  visitation  and
sampling,  the  operations  were  further  classified on the
basis of  size, relative age, and  whether  they  had  closed
water systems or produced an effluent from either the mining
or processing operation:
Operation A
    High tonnage
    Open-pit
    Iron formation

    Physical processing

Operation B
    Medium tonnage
    Open-pit
    Iron formation

    Physical processing
Operation C
    Medium tonnage
    Open-pit
    Natural ore
    Physical processing

Operation D
    Low tonnage
    Open-pit
    Natural ore
Older plant  (1957)
Mine produces effluent
Processing plant has closed water
     system
Medium age plant (1965)
Mine produces effluent
Processing plant has closed water
     system
     Older plant
     No effluent
(1948)
     Older plant (1953)
     Mine produces effluent
     Processing plant produces effluent
                          150

-------
    Physical processing

Operation E
    High tonnage
    Open-pit
    Iron formation
    Physical processing
                             Medium age plant (1967)

                             Mine produces effluent
                             Processing plant has closed
                                  water system
                             Medium age plant  (1967)
                             No effluent
Operation F
    High tonnage
    Open-pit
    Iron formation
    Physical processing

Operation G
    Low tonnage
    Open-pit
    Iron formation

    Physical and chemical
         processing

Operation H
    Medium tonnage
    Open-pit
    Iron formation
    Physical and chemical
         processing

Operation I
    Medium tonnage
    Open-pit
    Iron formation
    Physical and chemical
         processing

Operation J
    Low tonnage
    Underground
    Iron formation
    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
                             Older plant  (1959)
                             Mine produces effluent
                             Processing plant  produces
                                  effluent
                             Older plant  (1956)
                             Mine produces effluent
                             Processing plant  produces  effluent
                             Medium age plant  (1964)
                             Mine produces  effluent
                             Processing plant  produces  effluent
                             Older plant  (1958)
                             Mine produces effluent
                             Processing plant  produces  effluent
                          151

-------
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.

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  Subcategorization.    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
                            152

-------
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  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.

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 hydrometallurgical processes

    III. Copper mills employing the vat-leaching process

    IV.  Copper mills employing froth flotation

Problems in Subcategorizing 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 tailings
lagoon or use the water in the leach circuits.   A  decrease
                            153

-------
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.

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 utilized 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
                           154

-------
   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

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.
                            155

-------
The  treatability  of  mine  waste  water  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^) to ferrous sulfate
and sulfuric acid.  This oxidation  requires  both  moisture
and  air   (oxygen source) to occur.  The acid generated then
leaches hea\*y 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
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
                          156

-------
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   employed.    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-
         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
         40  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
                           157

-------
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 Subcateqorization

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
              recovery of gold.

    III. Mill(s)  or mine/mill complex(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 SiIver-Ore Subcateqorization

Based 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
                           158

-------
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 of silver.

Bauxite  Ores

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 waste water.

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  minedrainage  characteristics.   In  addition,   an
evaluation   of   removal   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  Subcategorization

 Based  on   the  results    of    intensive   study,    facility
 inspections,   NPDES  permit  applications, and communication
                            159

-------
 with the industry,  it was concluded that  the bauxite  mining
 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 Subcategorization

 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.
                         160

-------
     V.  Mills practicing ore  leaching and  associated
              chemical beneficiation  techniques.

 The  subcategory   including mills processing  less  than  5,000
 metric tons  of ore per year is representative of   operations
 which  are typically both intermittent in operation and eco-
 nomically  marginal.   This  subcategory  is   believed  to
 contain,   at  present,  almost  exclusively  processors of
 tungsten ores.

 Purely physical processing provides   the  minimum  pollution
 potential  consistent  with  recovery of values from an ore
 using water.  All  mills falling into  this  subcategory are
 expected  to share the same major pollution problem—namely,
 suspended solids generated by  the  need  for  crushing and
 grinding.    The  exposure of finely divided ore  (and gangue)
 to water may also  lead to solution of some  material, but, in
 general, pretreatment levels   will  be  low and   treatment,
 relatively simple.  The dissolved material  will clearly vary
 with  the  ore being processed, but treatment is expected to
 be essentially the same  in  all  cases  and  to   result in
 similar  maximum   effluent  levels.   There are currently no
 active major water using physical processors in  the  ferro-
 alloy industry except in the case of  nickel, where water use
 is  not  really  in the process.  Information has  been  drawn
 heavily, therefore,  from  past  data  and  related  milling
 operations—particularly,  in  the  iron  ore industry.   The
 close relationship between iron ores  and manganiferous  ores,
 where such production is likely in the near future, as   well
 as  the  nature  of  the  data  itself,  makes this transfer
 reasonable.  These  milling processes  are   fully  compatible
 with recycle of all mill water.

 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 solubility of
 ore components.   Reagents added to effect the flotation  may
 include major pollutants.  Cyanide, for example,  is commonly
 used  and,   though  usage is low,  may necessitate treatment.
 The  added  reagents  may  have  secondary  effects  on  the
 effluent  as  well;  the formation of cyanide complexes, for
 example,  may 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 process changes or serious  recovery
 losses.
                         161

-------
 Ore  leaching  operations differ substantially from physical
 processors and flotation plants in  effluent  character  and
 treatment  requirements.   The  use  of larqe quantities (in
 relation to ore handled) of  reagents,  and  the  deliberate
 solubilization of ore components,  characterizes these opera-
 tions.    The  solubilization  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+
 SOU.,  etc.)  may be present in sufficient quantities to  cause
 major  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
 One  facility was 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 and other  uses.   very
 little  beneficiating  of  mercury  ores  is  known  in  the
 industry.  Common practice for most producers   (since  rela-
tively  low  production  characterizes most operators) is to
 feed the cinnabar-rich ore directly to  a  kiln  or  furnace
without  beneficiation.  Water use in most of the operations
                          162

-------
is at a minimum, although a rather large  (20,000-flask-per-
year,  or 695-metric-ton-per-year or 765-short-ton-per-year)
flotation operation with high water use is  expected  to  be
operating  in  the  near  future.   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   (U.S.
Bureau of Mines projection).

Final Mercury Ore Subcategorization

Since   most   mercury   operations   are  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 is  expected in
the near future 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  waste
water, 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.

The  milling  processes   of  this  industry   involve   complex
hydrometallurgy.    Such  point  discharges  as might  occur in
milling  processes (i.e.,  the production of  concentrate)   are
expected   to  contain  a  variety of pollutants that need to be
 limited.   Mining, for the ores,  is expected   to  lead  to  a
                            163

-------
 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.    Th4
 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 hydrometallurgical 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
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  waste waters from mills using acid leaching remain acid
at the process discharge (not to be confused  with  a  point
                         164

-------
discharge),  retain  various heavy metals, and are generally
not  suitable   for   recycling   without   additional   and
specialized   treatment.   Those  from  the  alkaline  leach
process are normally  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 rawwaste 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
must be made between mills using acid leaching (or both acid
and alkaline leaching)  of  ore  and  mills  using  alkaline
leaching of ore only.

While other differences between ores and  processes, in addi-
tion  to   those mentioned above, can have an effect on waste
water 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)  retained   reflect  not  only  differences in waste
water characteristics but also  (a) differences in the volume
of waste water that  must be  stored and managed  in  a   zero-
effluent   condition   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 waste water from most current  mining and  ore
 dressing   operations  in the uranium industry.   In addition,
 other milling operations that now discharge waste water plan
 to terminate their discharges within a year  or two.

 Ore  characteristics   were   considered   and,    within   a
 subcategory,    cause   short-term   effect  on  waste  water
 characteristics    that    does    not    justify    fur^e*
 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
                            165

-------
 treatability  of  waste  water,   since  all  are  interrelated.
 Vanadium operations  which are  not extracting  radioactive ore
 or covered under government licensing   regulations   (NPC  or
 agreement  states),   are   subcategorized  in  the  ferroalloys
 section.

 Final  Subcategorization 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
              or combined  acid-and-alkaline leaching.

   III.    Mills which process uranium,   radium,  or   vanadium
              ores    to    yield    concentrates  by   alkaline
              leaching only.

 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
extraction,  producers  are  often  penalized  for  antimony
content at a smelter.
                          166

-------
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 other 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.
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.
                             167

-------
     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(es) 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
              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
                            168

-------
              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 waste water
characteristics  and  treatability.    The   other   factors
recognized  as  causing differences in the wastes discharged
do not significantly effect the treatability of  the  wastes
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.
                           169

-------
TABLE IV-1. SUMMARY OF INDUSTRY SUBCATEGORIZATION  RECOMMENDED
              CATEGORY
                                                          SUBCATEGORIES
                                            MINES
         IRON ORES
                                            MILLS
                Physical and Chemical Separation,
                Physical Separation Only
                                                            Magnetic and Physical Separation
                                            MINES
                                                            Open-Pit, Underground, Stripping
                                                            Hydrometallurgical (Leaching)
         COPPER ORES
                                                            Vat Leaching
                                            MILLS
                                                            Flotation Process
          LEAD AND ZINC ORES
          GOLD ORES
                                                            Cyanidation Process
                                                            Amalgamation Process
                                                            Flotation Process
                                                            Gravity Separation
                                                            Byproduct of Base-Metal Operation
                                            MINES
          SILVER ORES
          BAUXITE ORE
                                            MILLS
                                            MINES
                                            MINES
                                                            Flotation Process
                                                            Cyanidation Process
                                                            Amalgamation Process
                                                            Gravity Separation
                                                            Byproduct of Base-Metal Operation
          FERROALLOY ORES
                                                            < 5,000 metric tons (5,512 short tonsl/year
                                                            > 5,000 metric tons/year by Physical Processes
                                                            > 5,000 metric tons/year by Flotation
                                                            Leaching
          MERCURY ORES
                                            MILLS
                                                            Gravity Separation
                                                            Flotation Process
                                                            Byproduct of Base/Precious-Metal Operation
          URANIUM. RADIUM,
          & VANADIUM ORES
             ANTIMONY ORES
MINES

MILLS

MINES

MILLS
                                                            Acid or Acid/Alkdline Leaching
                                                            Alkaline Leaching
                                                            Flotation Process
                                                         	      	
                                                            Byproduct of Basc/Precious-Metal Operation
             BERYLLIUM ORES
              PLATINUM ORES
MINES
MILLS
                                    MINES OR MINE/MILLS
                                            MINES
             RARE EARTH ORES
                                                            Flotation 01 Leaching
                                                            Dredging or Hydraulic Methods
             TIN ORES
                                            MILLS
             TITANIUM ORES
MINES

MILLS
                                                            Electrostatic/Magnetic and Gravity/Flotation Processes
                                                            Physical Processes with Dredge Mining
             ZIRCONIUM ORES
                                    MILLS OR MINE/MILLS
                                                  170

-------
TABLE IV-2. FINAL RECOMMENDED INDUSTRY SUBCATEGORIES
SUBCATEGORY SUBDIVISION
1 ron Ores
1
1
^
1
Bauxite Ore
Ferroalloy Ores
Mercury Ores
Uranium, Radium,
& Vanadium Ores
Antimony Ores
Beryllium Ores
Rare-Earth Ores
Titanium Ores
Mines
Mills
Mines
(Open Pit, Under-
ground, Stripping)
Mines
Mills
Mills
Mills
Mills
Mine or
Mine/Mills
Mines
Mines
Mills &
Mines
Mills
Mills
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mines
Mills
Mills or
Mine/Mills
Physical and Chemical Separation
Physical Separation Only
Magnetic and Physical Separation
Copper
Lead and zinc
Gold
Silver
Hydrometallurgical (Leaching) (Copper)
Vat Leaching (Copper)
Flotation Process (Copper)
Flotation Process (Silver)
Flotation Process (Lead and zinc)
Flotation Process (Gold)
Cyamdation Process (Gold)
Cyamdation Process (Silver)
Amalgamation Process (Gold)
Amalgamation Process (Silver)
Gravity Separation (Gold)
Gravity Separation (Silver)
Gravity Separation (Platinum)
Gravity Separation (Tin)

>5,000 metric tons
<5,000 metric tons (5,512 short
tons/year
>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 Process





Electrostatic/Magnetic and Gravity/
Flotation Processes
Physical Processes with Dredge Mining
Zirconium Ores
Dredging or Hydraulic Methods
(Monazite)
                       171

-------
                         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 or 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 subcategory, with the waste characteristics and
loads identified for each,  concludes this section.

Because of widely varying waste  water  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
industry and various regulatory  and  research  bodies,  and
from   current  data  for  effluent  samples  collected  and
analyzed  during  this  study.   The  waste  water  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.
                           173

-------
 Mine Water

 The waste water 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  and is present only in placer  operations operating
 by  gravity  methods,  in  hydraulic  mining,  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.    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
 arevent, as much as possible,  excess water from entering the
 forking  area.    The quantity of water from an ore mine thus
 LS unrelated,  or  only  indirectly  related,   to  production
 quantities.    Therefore,  raw waste loadings are expressed in
 zerms of concentration rather than units  of  production  in
 -he ore categories discussed in  Section  IV.

 Cn  addition  to handling and treating often massive volumes
 af mine drainage during active mining operations,  metal  ore
 ;nine  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
 ill mining operations have ceased.   The   long-term  drainage
 :rom  tailing  disposal  also presents   long-term potential
 >roblems.   Surface mines,  on the other  hand,   are  somewhat
 tore  predictable  and less permanent in their production of
 line drainage period.   Water handling within a surface  mine
 .s  fairly  uniform   throughout  the life of the mine.   It is
 tighly   dependent   upon   precipitation     patterns     and
 »recautionary methods employed,  such  as  the use of diversion
 .itches,  burial of toxic  materials, and  concurrent regrading
 nd revegetation.

 Because  mine   drainage does not necessarily cease with mine
 losure,  a decision  must  be made as to the point  at which a
 .ine    operator  has   fulfilled    his    obligations   and
 -esponsibilities for a particular   mine  site.    This   point
                           174

-------
      be  further  discussed  in  Section  VII, "Control and
Treatment Technology."

SPECIFIC WATER USES IN ALL 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

Woncontact  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
 ased in or  resulting from  the process.

 |Moncontact  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.

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

 Transport Water
                           175

-------
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.

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.
                            176

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

    (1)  Waste water 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.

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.
                            177

-------
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.

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

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

Processing Wastes.   The processing of ore  from the mine may
result in the presence of a number of waste materials in the
waste water.  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
contri bution s.

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-
tuents of waste water.  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.
                          178

-------
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, TOG, 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-l; 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
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-4.    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
                            179

-------
    TABLE V-1. HISTORICAL CONSTITUENTS OF IRON-MINE DISCHARGES
PARAMETER
TSS
TDS
COD
pH
Oil and Grease
Al
Ca
Cr
Cu
Fe
Pb
MO
Hfl
Ni
Na
Mn
Zn
Chloride
Cyanide
CONCENTRATION 
-------
       TABLE V-3. CHEMICAL COMPOSITIONS OF SAMPLED MINE WATERS




PARAMETER



pH
Alkalinity
COD
TSS
TDS
Conductivity
Total Fe
Dissolved Fe
Mn
Sulfate





o
£
5
73"
204
274
2
455
440*
0.04
<002
0.21
85





|
UJ
z
5
7.2"
-
482
2
505
400*
<002
<002
<002
175





M
O
UJ
z
5
7.5'
-
9.2
5
609
700*
<0.02
<002
040
215
CONCENTRATION (mj/i ) in WASTEWATER FROM MINE

*~
8,

1
•&
'
7.2"
176
4.5
30
246
310*
<002
<0.02
<0.02
45
£
3

a
&
£
£
'
7.4"
-
1.0
<1
281
320'
<0.02
<0.02
<0.02
28
CM
01
n
|
(5
£
S

7.4'
_
18.3
21
169
21 5t
<0.02
<0.02
0.059
21

CM
&
IT
1
|

7.6"
211
22.8
20
271
340*
0.18
•C0.02
<0.02
26




r*.
o
Ul
z
1
8.4"
218
18
6
1.302
1,960*
4.50
<0.02
3.20
152
1
i
&

E
I
§

71"
37.4
4.5
10
118
110*
2.80
<002
0.025
11.2
£
I
s
o
i
1

7.2"
118
9.0
2
440
550*
1.30
0.04
0.054
33.2




i
UJ
z
i
8.3"
181
27 .5
12
308
342*
0.30
0.02
0.65
36.7




o
UJ
2
I
7.9"
66.0
<10
48
1790
1.125*
1.10
0.08
<0.02
780
1

1
5
§
|
*
7.54"
151
16.7
13.25
499.5
S66.8*
0.86
0.027
0.39
134
V»luem pH units
Value in micromhos/cm
     TABLE V-4. CHEMICAL COMPOSITIONS OF SAMPLED MILL WATERS


PARAMETER



PH
Alkalinity
COD
TSS»*
TDS
Conductance
Total Fe
Dissolved Fe
Mn
Sulfate

1102


Tailing-
Pond
Influent
_
-
9.3
30
533
-
210.0
<0.02
330.0
175
CONCENTRATION (mg/t, ) IN WASTEWATER FROM MILL
1103


Tailing-
Pond
Influent
7.5*
-
<1.0
12
198
350f
90.0
0.06
37.50
40
1105


Tailing-
Pond
Influent
8.2*
-
9.2
50
287
3751
1180.0
0.10
320.0
55
1107


Tailing-
Pond
Influent
*
7.6
-
22.5
15
712
-
0.70
< 0.02
67.0
236
1108


Tailing-
Pond
Influent
7.3*
-
13.5
12
230
1301
8.20
0.16
7.50
19.5
1109


Tailing-
Pond
Influent
9.5*
238
13.5
55
360
262 T
0.04
0.04
16.0
20.7
1110
Tailing-
Pond
Influent
and
Minewater
9.00*
13.4
11.9
22
2,360
1 ,900f
< 0.62
< 0.02
0.032
475



Average
(All
Mills)
8.2*
125.7
11.5
28
669
603f
212.8
0.06
111.1
146
 Value in pH units
 Value in micromhos
* Expressed in %
                              181

-------
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
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
                            182

-------
     Figure V-1. FLOW SCHEME FOR TREATMENT OF MINE WATER
                           MINE WATER
                       SEDIMENTATION BASIN
                     SETTLED
                     SOLIDS
                       TO
                      WASTE
                      CLARIFIED
                      EFFLUENT
                     TO RECEIVING
                       WATERS
                                                         t
                           TO PROCESS
                             WATER
  Figure V-2. WATER FLOW SCHEME IN A TYPICAL MILLING OPERATION
                              WATER
      TO
STOCKPILE'
PROCESS.
PRODUCT
PROCESS PLANT
                COAGULANT
                              f
                              TO
                            WASTE
                                         PROCESS
                                         TAILING
                                           i
                                   RECYCLE (80-97%)
                       SEDIMENTATION
                           BASIN
                       P
                     SETTLED
                      SOLIDS
                                               CLARIFIED
                                               EFFLUENT
                              f
                         TO RECEIVING
                            WATERS
                              183

-------
  impoundment structures,  loss of water to the  tailinas   anri
  evaporation or water loss during processing.   taxllngs'   and

  Process  Descriptions
                                      flow within
 Mine  and  Mill  1105.    Mine  and  mill  1105 is a  typical
 taconite operation.   Open-pit  mines  associated  with  ?he
 Crude magnetic taconxte 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
 i^Tar^' iJ?6 ^^ magnetite fro<« the mill is agglomerate!
 me^rif tons  oT  "I v  ^^ *W™*^eIy 2?6H million
 metric  tons  (2.6  million  long  tons)   of  oxide  oellPt«?
 annually for blast-furnace feed.                      pellets

 The mine  mill  and pelletizing plant are located on a large
 (20 oS^acr^f  £y the operating company, with 8094 hectares
 ^  an?  £   }  utxllzed at Present.   An initial tailing pond
 of  405  hectares  (1000  acres)   has beeji filled.   A second
 l,619hectare (4,000-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-l   Settling

                 *° C°ntaln ^ "
 to  wo   akes                                    ±S

 The mill water system is a closed  loop.   Plant  processes  use
oer    nn                       mgd) ' with  189  cubic   mers
per  minute  (72 mgd) returned from the 91.4-meter  (300-foot)
diameter tailing thickener overflow  and  15.1   cubic   me?2rs
per  minute   (5.7  mgd)  returned  from  the tailing  ponTo?
basin.  The tailing thickener receives waste or tailings   in
a  slurry  from  the  concentrate pellet plant.  A non?oxic
anionic polyacrylamide flocculant is added  to the  ?h7cken"er
to  assist in settling out solids.   Tailing thickener  unde?-
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
used   ?o  1oT b^St±n2  agentS-    Electric:  shoved are
S f  i, ?  . ?     the   broken  ore  into  100-ton-capacitv
diesel/electric trucks for haulage to the primary crusher.
                           184

-------
       Figure V-3.  WATER BALANCE FOR MINE/MILL 1105  (SEPTEMBER 1974)
              3.4 m3/mm (900 gpm)
               (INTERMITTENT)
  SETTLING BASINS
CREEK
              J)
                                    MINE DEWATERING
                                                  17 to 32 m3/min (4,500 to 8,500 gpm)
                                       MAX. 8.23 m3/mm (2,200 gpm)
                                       (INTERMITTENT)
Q   LAKE 1    J
                   2 PUMPS
                 @ 11.4 m3/mm
                 (3,000 gpm) EA.
                (INTERMITTENT)
                                   PLANT STORAGE TANK
                                            15.1 m3/min (4,000 gpm)
                               PLANT PROCESSES
                                            204 m3/min (54,000 gpm)
                                   TAILINGS THICKENER
                                                          189 m3/min
                                                          (50,000 gpm)
                                       15.1 m3/min (4,000 gpm)
                               s*—       "x  15-
                               ( TAILING POND %—
15.1 m3/min (4,000 gpm)
                                                                            SETTLING BASINS
                                                                              17to32m3/mm
                                                                              (4,500 to 8.500 gpm)
C   LAKE 2   ^
                                           185

-------
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  crusher,  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
m-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-
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  'green  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.
                           186

-------
Figure V-4. CONCENTRATOR FLOWSHEET FOR MILL 1105
                       FEED
                       1
        DRY SEMIAUTOGENOUS GRINDING MILLS
                       I
             VERTICAL DRY CLASSIFIER
         OVERSIZE
UNDERSIZE

   I
                            CYCLONE CLASSIFIER
                         OVERSIZE
        UNDERSIZE

            t
                                          WET MAGNETIC COBBING
    OVERSIZE  UNDERSIZE

        I         t
       CONCENTRATE
                        TAIL
             DRY MAGNETIC ROUGHER
             TAIL
                         CONCENTRATE
                              I	
     DRY MAGNETIC SCREENING
      I
  MIDDLING
       WET SECONDARY
       GRINDING MILLS
                                      HYDROCYCLONES
                                     UNDERSIZE   OVERSIZE
             TAIL
        i
                                    HYDROSEPARATION
                                CONCENTRATE
                                                 TAIL
                        WET FINISHER MAGNETIC SEPARATION
                  CONCENTRATE
                                                   TAIL
                       TO
                     PELLET
                     PLANT
             TAILING THICKENER
             I
        UNDERFLOW
                                                        OVERFLOW
             TO
            WASTE
            TO
       TAILING POND
     TO
REUSE WATER
                             187

-------
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/limonite/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
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 U5 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.
                          188

-------
   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
                              189

-------
                Figure V-5.  FLOWSHEET FOR MILL 1104 (HEAVY-MEDIA PLANT)
                                           CRUDE ORE
                                            (37% IRON)
            DOUBLE-DECK
              SCREEN
-< 0.64 cm (0.26 in.)-
      > 15.2 cm
                  10.2 to 15.2 cm
                   (4 to 6 in.)

                  .	I
DOUBLE-DECK
  SCREEN
                J<24%F«


          TO ROCK-REJECTS
             STOCKPILE
                      j3C%Fi
                                                               42% p.
                                                     CONE
                                                    CRUSHER
                                                    SCREEN
                                             > 0.64 em
                                             O0.2S in.)
                          < 0.64 cm
                          K0.2B in.)
!

HEAVY-MEDIA
SURGE PILE
                      0.64 to 3.2 cm (0.25,10 1.25 m.l-
                                     < 0.64 cm
                                     K 0.26 in.)
                                          <0.64cm(<0.2Sin.l-
                                                                  RAKE CLASSIFIER
                                                            48m.it. to
                                                         (0.64 cm (0.25 in.)
                                            OVERFLOW


                                                I 31% Fe
                                                                                TO
                                                                            TAILING POND
HEAVY-
MEDIA
SEPARATOR
(SP.GR. - 3.10)
14% Fi

HEAVY-
MEDIA
SEPARATOR
(SP.GR. - 2.90)
58% Ft


16% F. 60% Ft
i '
1
FLOAT REJECTS   FLOAT REJECTS


           » i<
  TO FLOAT-REJECTS
    STOCKPILE
                                                 190

-------
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
rem
depending on the quality of the  oversize  material.   P^nt
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
heavymedia surge pile.  The fraction which is ^ss 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.

Concentrates  produced  in the  plant are shipped  by rail and
boat to the  lower  Great Lakes.   The  5 8-percent- iron  heavy-
media   concentrate   serves  as   blast-furnace  feed.   The  58-
percentiron 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.
                             191

-------
 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
 114 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 waste water 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


 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
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
                           192

-------
TABLE V-6. CHEMICAL CHARACTERISTICS OF DISCHARGE WATER
          FROM MINE 1108
PARAMETER
pH
Alkalinity
COD
TSS
TDS
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION
(rna/H )
7.2»
118
9.0
2
440
1.3
0.04
0.054
33.2
        •Value in pH units
                         193

-------
Figure V-6. SIMPLIFIED CONCENTRATION FLOWSHEET FOR MINE/MILL 1108
                                   MINING
                                     I
                                 CRUDE ORE
                                                             9.3 m3/min (5.5 cfs
                                      24,500 metric tons
                                      (20,700 long tons)
                                      per day
                                REGRINDING,
                           FLOTATION, THICKENING,
                              AND FILTRATION
                                     I
                                SECONDARY
                               CONCENTRATE
                                   i
                                PELLETIZING
                                OPERATION
                                 PELLETS
                                    t
                               TO STOCKPILE
                                                          17 m3/min (10 cfs)
                                                        41.6 m-Vmin (24.5 cfs)
                                                     I '    14.4 m3/min (8.5 cfs)
17 m3/min (10 cfs)
                                                        0.85 m3/min (0.5 cfs)
                                                                      8.1% SOLIDS
                                                            100 m3/min (59 cfs)
                 I
                                                                      TO TAILINGS
                                      194

-------
    TABLE V-7. CHARACTERISTICS OF MILL 1108 DISCHARGE WATER
PARAMETER
pH
Alkalinity
COD
TSS
TDS
Tout F.
Dtaolv«lF>
Mn
Sulfttl
PROGRAM SAMPLE
CONCENTRATION
W8.HN
WASTE WATER
7.1*
62.0
22.5
10
160
2.06
0.93
0.06
5
WASTE LOAD
in g/nwtric ton
(Ib/fhort ton)
PRODUCT
-
213 10,42)
77.4 (0.1S)
3.4 (0.0071
560 11.08)
7.06 (0.013)
3.2 (0.0061
0.17 (0.0003)
17.2 (0.034)
104MMNTH AVERAGES
AVERAGE
CONCENTRATION
(mg/JU
7.0-
-
8.6
-
0.76
0.66
~
WASTE LOAD
in g/nwtrk: ton
(Ib/lhort ton)
PRODUCT
-
-
20.7 (0.040)
-
1.83 (0.0036)
1.68 10.0031)
~
HIGH
CONCENTRATION
(mg/fc)
7.8*
-
S3
-
3.60
5.80

LOW
CONCENTRATION
(mj/m
6.5'
~
1
~
0.01
0.01

•Vlhn in pH unitl
                               195

-------
                    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
 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 too low in copper  content
                           196

-------
TABLE V-8. PRINCIPAL COPPER MINERALS USED
           IN THE UNITED STATES
MINERAL

Chalcocite
Chalcopyrite
Bornite
Covellite
Enargite
COMPOSITION
SULFIDES
Cu2S
CuFeS2
Cu5FeS4
CuS
Cu3AsS4
OCCURRENCE*

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
                   197

-------
 or  as oxide ore, which does not concentrate economically bv
 flotation.                                                 J

 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 waste water  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 ton)  of 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.

 Air  quality  control   within  open-pit  mines  consists  of
 spraying water on  roads for dust  control.  Underground mines
 may   employ   scrubbers,   which    produce   a    sludge  of
 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  (FeS2.) .  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
 4.0  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.
                           198

-------
Figure V-7. WASTEWATER FLOWSHEET FOR PLANT 2120-B PIT
 NATURAL DRAINAGE,
    SEEPAGE, AND
       RUNOFF
       MINING
      (DRILLING,
      BLASTING,
        AND
      LOADING)
         ORE-
 16,560,000 metric tons/year
(18,250,000 short tons/year)
TO
MILL
       EXCESS
     MINEWATER
                «J
           0.06 m /metric ton
           (14.4 gal/short ton)
                          LIME PRECIPITATION
           0.06 m3/metric ton
           (14.4 gal/short ton)
    DISCHARGED
                          199

-------
TABLE V-S. 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)f
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
gjIS.OIavg)*
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.
         0 to 81.1 m3/metr,c 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-m3/19.432-gal) months.
     N/A = not available
     N.E. = no effluent
                                         200

-------
TABLE V-10. SUMMARY OF SOLID WASTES PRODUCED BY PLANTS SURVEYED
MILL
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)
33,623353*
82,737 (UG)
33,112,800*
O(UG)
88,452,000*
10386,400 *
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
12,000,000t
20,000 (UG)
35,557,000*
0(UG)
37,063,000*
91,200(UG)
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
3362^74
1,567,460
3,712,262
1,480^50
635,040
9,383,475
130,386
471,375
11,465,193
1,211,680
16,656,192
19,935,266
23,342,256
8,059,688
34,745,760
1,970,438
7,912398
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
83.5T
0.04
2.81
-
2.02
0.004
1.42
-
2.55
5.53
1.94
* All or a portion leached
  Stripping operation
N/A = Not available
UG = Underground
                                     201

-------
                               TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM

                                         SELECTED COPPER MINES (Sheet 1 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb

CONCENTRATION
(mg/K,l
42,01 3.5m3/day
9.64*
544
8
1
5
<10
0.2
0.5
< 0.05
< 0.003
< 0.50
< 0.07
< 0.05
< 0.2
3.80
< 0.05
< 0.05
< 0.10
< 0.2
0.13
0.0008
< 0.05
MINE 2119
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 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
< 0.0385
< 0.0385
< 0.077
< 0.154
0.10
0.00062
< 0.038
lb/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
5.846
< 0.0770
< 0.0770
< 0.154
< 0.308
0.20
0.00124
< 0.076
MINE 2120-K
CONCENTRATION
(mg/U
27,524.5m3/day
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
0.24
< 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
O5.17
7.13
303.4
1.52
1 ,395.6
4.85
N/A
< 3.03
< 1.06
2,609.2
< 7.59
30,340
1,517
5.01
3.64
< 7.59
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
606.8
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
o
to
         *Value in pH units

-------
                       TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                  SELECTED COPPER MINES (Sheet 2 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pta
MINE 2120 B
CONCENTRATION
ImgU)
2,725.2m3/day
6.1*
2,152
40
< 1.0
3.2
<10
0.04
5.30
0 1
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*
258.6
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
< 0.01 2
MINE 2120 CE
CONCENTRATION

-------
                              TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                        SELECTED COPPER MINES (Sheet 3 of 4)
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
HS
Pb

CONCENTRATION

-------
                              TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                        SELECTED COPPER MINES (Sheet 4 of 4)
O
Ul
PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2123
CONCENTRATION
(mg/i'l
409m3/day
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 tons
75 m3/10OO metric tons
6.96"
101
0.2
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
< O.OOOOO8
< 0.04
lb/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
< O.O016
0.016
< 0.08
< 0.016
0.14
< 0.004
< 0.008
< O.O4
0.12
< O.O00016
< 0.08
                               *Valu^in pH units

-------
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  (H)  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
plants surveyed discharge all of their mine water to surface
waters.  Half  of  these  treat  the  water  first  by  lime
precipitation and settling.

Process  Description-Hydrometallurgical Extraction Processes
(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
                            206

-------
   Figure V-8.  FLOWSHEET OF HYDROMETALLURGICAL PROCESSES USED IN
               ACID LEACHING AT MINE 2122
TO ATMOSPHERE
                                                TO ATMOSPHERE
      43 nr/nxtric ton
      (10,282 gri/ihort ton)
241 m3/mttric ton
(57,834 galMiort ton)
EVAPORATION
                                                 EVAPORATION


SEWAGE,
SEEPAGE,
RUNOFF,
MINE WATER,
AND WELL WATER
MAI
WA
{
296m3
(70,879 pi
EUP
TER
\
— ^""RESERVOIR

27 m /mttric ton _
(6,426 pl/ihort ton)

2,294 m3/m«tric to
'm.tric ton (549,873 gd/ihort tor
/then ton)
BA
360 m3/RMtric ton SOL
(86,303 j«l/«hort ton) MAKEUP
WATER '
107 m3/nwtric ton r—
U. (25,704 gri/thort ton)



DUMP
LEACH
i ,
PREG
", SOLI
RREN
UTION
L
r
NANT
TION
2,386 m3/m«tric ton
(571.914 ^l/ihort ton)
IK m3/mMric t™
PRECIPITATION ^ (3,727 jrt/ihort ton! poTABLE
PLANT
^ WATER
1
CEMENT
COPPER
                                                      03 m3/metric ton
                                                      (64 |il/ihort ton)
                                                     TO
                                                  STOCKPILE
                                     207

-------
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
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.
                            208

-------
Fiflure V-9. REACTIONS BY WHICH COPPER MINERALS ARE DISSOLVED IN
          DUMP, HEAP, OR IN-SITU LEACHING
                                AZURITE

            Cu3(OH)2-(C03)2 + 3H2S04 ^   * 3CuSO4 + 2CO2 + 4H2O


                              MALACHITE

              Cu2(OH)2-C03 + 2H2S04 ~ — ^ 2CuSO4 + CO2 + 3H2O


                             CHRYSOCOLLA

               CuSi03'2H20 + H2S04 ^Z± CuSO4 + SiO2 + 3H2O


                                CUPRITE

                     Cu20 + H2S04 ~  ^ CuS04 + Cu + H20

           Cu20 + H2S04 + Fe2(S04)3 ^Z± 2CuSO4 + H2O + 2FeSO4


                             NATIVE COPPER

                    Cu + Fe2(SO4)3  ^  ^ CuSO4 + 2FeSO4


                              TENORITE


                      CuO + H2SO4  v  ^ CuSO4 + H2O


            3CuO + F«2(S04)3 + 3H20  ^Z± 3 CuSO4 + 2Fe(OH»3

         4CuO + 4FeS04 + 6H2O + O2  ^— ^ 4CuSO4 + 4Fe(OH)3


                             CHALCOCITE


                  Cu2S + Fe2(SO4)3 ~  ^ CuS + CuSO4 + 2FeSO4


                Cu2S + 2Fe2(S04)3 ^   ^ 2CuSO4 + 4FeSO4 + S


                              COVELLITE

                   CuS + Fe2(SO4)3 ~  ^ CuSO4  + 2FeSO4 + S



        Chalcopyrite will slowly dissolve in acid ferric sulfate solutions and also
         oxidize according to:
                     CuFeS2 + 2O2 7~^ CuS + FeSO4;

                        CuS + 2O2 ~  V CuS04.

      Pyrite oxidizes according to:


               2FeS2 + 2H20 + 7O2 ^Z± 2FeSO4 + 2H2SO4.

-------
Figure V-10. TYPICAL DESIGN OF GRAVITY LAUNDER/PRECIPITATION PLANT
            DRAINS
        SIDE VIEW
                                         CELL
         SOLUTION

  CELL    FLOW *ft
                                            •z-
                                             o
TOP VIEW


CANS
                                                    SOLUTION

                                                      FLOW
                                                        PERFORATED

                                                        SCREEN
                           END VIEW


                     SOURCE:  REFERENCE 23

-------
       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 23
                               211

-------
                                                 •=*.-*

               CuS04 + Fe -r-> cu  +  FeS04

 Scrap iron of other forms and sponge iron may be employed.

                         :         -     •—•
 cone,  into the  shredded iron scrap.  This
 percent    The  resulting  cement copper is  85  to 99 percent
 pure  and 1S  sent to the smelter for further  purification?
 ^lbarre£ ^lution from * Precipitation  plant  is  recycled
 from  a  holding  pond  to  the  top  of  the ore body  after
 sulfunc acid and makeup water are added,  if necessary.


 Leach solutions containing greater than 25 to 30  grams  per

 ties   T^e         " U8ua11* Sent to  electrowinning
Solvent  extraction  of  copper  from acid leach solutions by

        "6       ±S raPidly Becoming an important methoS  of
                          liquors contain less than 30 grams

                                1S m°St
            .extra?tiQn,   a  reagent  with high affinity for

ainitv  fn°\Sn  W6ak 3Cid  solutionsr  and  with7 low
?t  is  DifSd  t   •  ions; is carried in an or^anic medi™-
It  is  placed  in  intimate  contact  with   copper   leach
solutions,  where  H+  ions  are  exchanged for Cu(++)  ion?
This regenerates the acid, which is recycled  lo  the  dump"
The  organic  medium,  together  with  copper,  is sent to  -
stripping  cell,   where  acidic  copper  sulfate
                        212

-------
Figure V-12. DIAGRAM OF SOLVENT EXTRACTION PROCESS FOR RECOVERY OF
          COPPER BY LEACHING OF ORE AND WASTE
                                      MINE
                                      DUMP
                                    WEAK CuSO4
                                  LEACH SOLUTION
                       RAFFINATE_
                       RECYCLED
 SOLVENT
EXTRACTION
  PLANT
                                        I
                                     Cu++ ON
                                 ORGANIC CARRIER
             RECYCLED
              ORGANIC
            CARRIER (H+)
                                       i


STRIPPING
AREA


                  RECYCLED
              ACIDIC ELECTROLYTE
                    (H2S04)
     I
   CuS04
ELECTROLYTE
    i
                                   ELECTROLYTIC
                                    RECOVERY
                                      PLANT
                                     CATHODE
                                      COPPER
                                        I
                                        TO
                                     STOCKPILE
                              213

-------
exchange H+ ion  for Cu(++).  This regenerates the organic/m-
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  0.454  kg   (1
Ib) 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-
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 technigues 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.
                          214

-------
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: Copper - A
         Position Survey, 1973, Reference 24)
       t Production taken from NPDES permit application
       N/A - Production not available; only flow available
                             215

-------
      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 Grease
S
As
B
Cd
Cu
Fe
Pb
Mn
Hg
Ni
Tl
Se
Ag
Te
Zn
Sb
Au
Co
Mo
Sn
Cyanide
CONCENTRATION (mg/H) 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
<0.05
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
<0.5
<0.05
3.80
0.75
-
<0.01
2123
3.56"
44,368
162
80
27.5
2.0
<0.5
0.07
<0.01
5.55
97.0
650.0
0.1
123.5
0.0010
5.68
<0.1
0.030
<0.1
1.8
33.0
<0.5
<0.05
7.3
1.33
-
<0.01
2122
2.49*
83,226
34
385.1
46.0
5.0
<0.5
<0.01
0.08
4.50
72.0
3,500.0
1.14
190.0
0.0003
31.1
<0.1
< 0.003
0.038
2.5
74.5
<2.0
<0.05
72.0
0.35
2.40
•C0.01
2125
4.24*
29,494
218
440.0
11.0
<1.0
<0.5
<0.07
0.03
0.20
7.00
3,688.0
<0.1
1494
0.0007
6.90
<0.1
< 0.020
<0.1
1.1
21.0
<0.5
<0.05
13.70
0.5
-
<0.01
2104
3.39*
-
-
-
.
-
-
0.04 to 0.60
-
0.56
52.25
-
0.68
-
0.0003
-
-
0.13
-
-
-
-
-
-
-
-
-
•Value in pH units
                            216

-------
                    Figure V-13.  VAT  LEACH  FLOW  DIAGRAM (MILL 2124)
                                            TO ATMOSPHERE
ORE WASH WATER
         16i m3/nwtric ton
         (38,463 (ri/ihort ton)
                                                                                       TO ATMOSPHERE
                                                     I m /metric ton
                                                    (8,166a«l/thort tonl
                                             EVAPORATION
    CRUSHER
H2S04


1.1 rrr/rmtric ton
(2«4 wl/ihon ton
ORE MAKEUP _
280 m /nwtric ton
i

\


LEACH
TANKS
1
l<
2
                    8.506 gal/short ton)
     SLIMfS
                                           TAILS
                                                                BARREN SOLUTION-
                                                                176 m3/m»tric ton
                                                                142.181 oil/ikoit ton)
        COPPER-RICH 	
        ELECTROLYTE
       287 m3/metric ton
      (48,578 gil/ihort ton)
                                    31 m3/mrtric ton
                                    (7,397 ad/short ton)
                                                                                      ELECTROWINNING
        W§ m /iMtric ton
        (38,463 (ri/ihort ton)
                          28 m3/mttric ton
                          (5,620 o>l/thort ton)
                               10 m3/nwtric ton
                             (2,411 o>l/ihort ton)
    TO MILL
                                                       WASH
                                                      WATER
                      TO MILL
                                         TO WASTE
188 m3/rr»tric ton
(45,176 9.1/ihort ton)
                                                                                            TO
                                                                                         STOCKPILE
                                                  TO LEACH DUMPS
                                                 AND PRECIPITATION
                                                      PLANT
                                          217

-------
 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-14  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.

 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-11.)

Typically,  coarse  ore is delivered to the mill for two- or
three-stage reduction by truck, rail or conveyor and is then
                           218

-------
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)
m^/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 24
       were used to calculate a value which may be a low estimate of water use.
     ••Product is cathode copper
                               219

-------
TABLE V-16. CHEMICAL CHARACTERISTICS OF VAT-LEACH BARREN ACID
          SOLUTION (RECYCLED: NO WASTE LOAD, MILL 2124)
                             CONCENTRATION (mg/e )
        *Value in pH units
                      220

-------
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
                          221

-------
       Figure V-14. FLOW DIAGRAM FOR FLOTATION  OF COPPER (MILL 2120)
                                    MINING  I
                                     ORE
                                                                                TO ATMOSPHERE
                                                          195 m3/metric ton
                                                          (46.735 9»l/»hort ton)
                 27 m /metric ton
                (6,491 Hl/ihort ton)
                                                                                       15 m3/nwtric ton
                                                                                       (3,709 pl/ihort ton)
                                                                        EVAPORATION
                                                                              (3,709 gal/thort ton)
RECYCLES
©*©'
                                           OVERFLOW
                                            (IF ANY)
                                                                          DISCHARGE
118 m'/mctric ton
(28,189 gil/ihort ton)
                                        222

-------
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:
Reagent
  type

pH control
collector
collector

frother
Example of
 Reagent

lime
Xanthate
'Minerec
compounds
MIBC
Ib/short ton
 mill feed

    10.0
     0.01
     0.03

     0.02
kg/metric ton
  mill feed	

    5.0
    0.005
    0.015

    0.04
 The  specific  types  of  reagents  employed  and   amounts   needed
 vary  considerably   from  plant  to   plant,  although  one  may
 classify  them,  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
                            223

-------
Figure V-15. ADDITION OF FLOTATION AGENTS TO MODIFY MINERAL SURFACE
       REAGENTS TO
        ADJUST pH
        COLLECTOR
                      PULP FROM GRINDING CIRCUIT
                            (25-45% SOLIDS)
CONDITIONER
                                i
                                                 WETTING AGENT
                                                   DISPERSANT
                            CONDITIONER
                •FROTH-
          TO
      ADDITIONAL
      PROCESSING
                       ACTIVATOR
                     (OR DEPRESSANT)
CONDITIONER
 FLOTATION
   CELLS
                                            -TAILINGS-
                          TO
                        WASTE
                            224

-------
TABLE V-17. EXAMPLES OF CHEMICAL AGENTS WHICH MAY BE
          EMPLOYED IN COPPER FLOTATION
MINERAL
Bornite
Chalcocite
Ghalaopvrite
Native Copper
Azurite
Cuprite
Malachite
PRECIPITATION AGCNT
—
—
—
—
Sodium monotulfide
Sodium monoiulfide
Bedium menatulfide
PH
REGULATION
Lime
Lime
Lime
Lime
Sodium carbonate
Sodium carbonate
Sodium carbonate
DtSPERCAMT
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
DEPRISSANT
Sodium cyanide
Sodium cyanide
Sodium cyanide
Sodium cyanide
Quebracho
Quebracho
Tannie acid
ACTIVATOR
—
—
—
—
PotytuMide
PotytuMide
Polyiulfio.
COLLECTOR
Xanthate
Aerofloats
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
acid
Pine oil.
Vapor oil,
Cresylic
acid
Pine oil.
Vapor oil,
Cresylic
acid
                   Source: Reference 25
                      225

-------
 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 Se
 floated from copper flotation feed; gold and silver  may  be
 leached   by  cyanide  from  the  copper  concentrate,  with
 precipitation by zinc dust.

 W^ter 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
 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
 tailing-pond   (settled) decant water are summarized in Table
V-19.   Residual flotation agents or their  degradation  pro-
 ducts   may  be  harmful  to  aguatic  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.
                          226

-------
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 24
                          227

-------
        TA*tE V-U. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
                  TAILING PONDS (Sh«t 1 of 4)

II
pH
•M'

TM


QMandGnMt

CQNC€MT RATION
213.080 m3/d.y
0.1J-

1.4N
4

<1.0
Al < 1.0
At
Cd



<0.07
<0.06
0.28
0,06
< 0.6
I
Ag

Sr
Zn


< 0.803
< 0.1

0.83

O»WMinf

< 0.2
< 0.01
MILL 2110 || 	 |ortMM
48,200 »H/*ort ton
0.12*
—
478.883
II CONCENTRATION
JGacMnaiGaB
146,080 m3/d.y
(28,008.000 1.1/d.y)
11.00'
36%
2,082
1 4«K fl

< 336.2
< 336.2
< 23.63
< 16.81
04.13
310.38
V00.1

-------
       TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
                  TAILING PONDS  (Sheet 2 of 4)

PARAMETER
Flow
pH
SES*
TDS
TSS
Oil and Grease
SiO2
Al
A<
Cd
Cu
ft
Pta
Mn
Hg
Ni
S>
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphate
Cyanide
Operating
dayi/yaar
Annual
Production

MILL 2121 (Slime Tails)
CONCENTRATION
(mg/il
64.964 m3/day
17,161 ,000 gal/day)
9.3-
30%
438
202
1.0
4.75
5.9
<0.07
<0.02
3.50
10.05
0.22
0.25
0.0098
<0.05
0.022
<0.1
0.07
0.9
<0.5
<0.04
<0.05
<0.5
0.24
<0.01

RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
104.7 m3/metric ton
9.3'
-
45.863
21,151
104.7
49,737
617.8
<7,33
<2.09
366.48
1,052.23
23.04
26.18
1.0261
<5.24
2.304
<10.5
7.33
94.24
<52.4
<4.19
< 524
<52.4
25.13
<1 OS
lb/1000 short tons
25.097 gal/short ton
9.3"
-
91.725
42.302
2O9.4
99.474
1.236
< 14.66
< 4.19
732.96
2.104.65
46.07
52.35
2.0523
< 10.47
4.607
<20.9
14.66
188.48
< 104.7
< 8.38
< 10.47
< 104.7
50.26
< 2.O9
•utn


223,318 metric torn (246,162 short tons)

MILL 2121 (lands)
CONCENTRATION
Img/JU
32,825 m3/day
18.672,400 gal/day)
9.28*
10%
310
6
<1
-
900
<0.07
0.03
46
1.216
0.40
48
O.O001
1.72
< O.O03
<0.1
0.06
8.50
<0.5
1.1
<005
<0.5
-
<0.01
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
52.9 m^/metnc ton
9.28*
—
16,404
317
< 52.9
—
47.624
<3.70
1.59
2,434.1
64.345.3
21.17
2.539.9
0.0053
91.01
< 0.159
< 5.3
317
449.78
<265
58.21
< 2.65
<26.5
-

-------
         TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
                     TAILING PONDS (Sheet 3 of 4)
PARAMETER
Flow
pH
SES*
TDS
TSS
Oil end Grin*
&02
Al
As
Cd
Cu
ft,
Pta
Mn
Hg
Ni
Se
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphete
Cyanide
Operating
days/year
Annual
Production
of Concentrate

CONCENTRATION
Img/il
278.084 m3/d.v
(73,470,000 gal/day)
8.54'
IS*
4,276
24
3
12.26
<1
<0.07
<0.05
0.08
<0.1
2.79
0.047
0.0002
<0.1
0.022
<0.1
1.81
<0.05
<1.0
0.08
<0.05
<0.2
0.15
<0.01
MILL 2122
RAW WASTE LOAD PER UNIT PRODUCT
kg/1000 metric tons
134 m3/metnc ton
8.54-
-
573,188
3,217
402.1
1,675.60
<134
<9.38
<6.7
10.72
<13.4
373.99
6.3
0.0268
<134
2.949
< 13.4
242.63
< 6.7
<134
10.72
< 67
< 26.8
20.11
< 1.34
lb/1000 ihort tons
32,079 gel/short ton
8.54*
-
1,146,376
6,434
804.3
3,351 19
< 268.1
< 18.77
< 13.4
21.45
< 26.8
747.99
12.6
0.0536
<26.8
5.898
< 26.8
485.25
<13.4
< 268.1
21.45
<13.4
< 53.6
4021
< 268
357
740.602 metric tons (81 7,636 short tonl)
MILL 2123
CONCENTRATION
(mg/£)
11,446 m3/day
(3,024,000 gal/dayl
13.00*
30%
2,494
20
10
27
1
<0.07
<0.03
0.77
0.15
<0.1
< 0.06
0.0019
< 0.05
0.07
<0.1
2.26

-------
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
           TAI LING PONDS (Sheet 4 of 4)

PARAMETER

Flow
pH
SES'
TDS
TSS
Oil and Grease
Si02
Al
As
Cd
Cu
ft
Pb
Mn
Hg
Ni
Sa
Ag
Sr
Zn
Sb
Co
Au
Mo
Phosphate
Cyanide
Operating
days/year
Annual
Production
of Concentrat
MILL 21 24
CONCENTRATION

-------
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  Mi11  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 process water.  Also, of
the  six,  three  have  sophisticated  lime   and   settling
treatment,  or  are installing it, to protect the guality of
the discharge.

Three of the copper mills surveyed , all of which  discharge
water  from  the tailing pond, are compared in Table V-20 as
to the quality  of,  and  the  amount  of  loading  in,  the
discharged  decant  water.    In  the  calculations  made  to
present these data,  no  allowance  was  made  for  incoming
process water.

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.
                         232

-------
                        TABLE V-20. WASTEWATER CONSTITUENTS AND WASTE LOADS RESULTING
                                  FROM DISCHARGE OF HILL PROCESS WATERS
PARAMETER

TUB

MmriGrMM







M»




CyMM*
CONTROL
TREATMENT

CONCENTRATION
Inn/ « ) IN WASTEWATER
9.6**'
3.3B*

15.0
<0.07
< 0.01
< 0.0*6
0.06
< 0.10
< 0.1
0.03
0.6*11
< 8.06
0.043
2.40

                    Md tonh soluMn
"•In

ttf
             **tt Mn>

        \W«. in pH urn*

-------
Figure V-16. FLOWSHEET FOR MISCELLANEOUS HANDLING OF FLOTATION
          TAILS (MILL 2124)
                   SULFIDE
               FLOTATION CELLS
                  TAILINGS
                     i
               HYDROSEPARATOR
                   SLIMES
                     i
                    ACID
                    LEACH
                 (THICKENER)
                      PREGNANT
                      SOLUTION
              BARREN
             SOLUTION
                PRECIPITATION
                   PLANT
                     I
                   CEMENT
                   COPPER
     CONCENTRATE
  •SANDS-
UNDERFLOW
  (SLIMES)  "
TO
STOCKPILE
TO
TAILING
POND
                      TO
                      STOCKPILE
                              234

-------
 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
 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/Precipitation/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
 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 +  FeSO^J
                   (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:
                          235

-------
        Figur. V-17. DUAL PROCESSING OF ORE {MILL 2124)
                   MINING
                    ORE
                  CRUSHERS
                     i
                   LEACH
RECYCLED
 WATER      ORE
            LEACH
            TAILS
                            _
                       .   ACID
                        SOLUTION"
                              RECYCLED AGIO-
 DIRECT
SULFIDE
 MILL
 FEED
             i
        CONCENTRATOR
                    RECYCLED
                   'OVEftFLOW'
                  .RECYCLED
                    DECANT
                                              ELECTROWINNING
                              TAILING
                             THICKENERS
                                                         TO
                                                      STOCKPILE
                           236

-------
Figure V-18. LEACH/PRECIPITATION/FLOTATION PROCESS
              COPPER SULFIDE CONCENTRATE
                       AND
                   SPONGE COPPER
                   TO STOCKPILE
                       237

-------
    Reagent             kg/metric ton       Ib/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) .

The chemical characteristic of  raw  waste  water  from  the
milling  operation  appear  to be considerably less variable
from facility to facility than mine waste water.  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  per
day   (79,070  gal/short  ton/day)  to 1,100 cubic meters per
metric  ton  per  day   (263,566   gal/short   ton/day)   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.

     (4)  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
                            238

-------
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  equipment   utilized   (and,
         often, maintained) underground.

     (3)  Hardness and alkalinity  associated   with  the  host
         rock and ore.

     (4)  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
waste  water  flow   for  a  mining  operation  exhibiting  no
solubilization  potential   is    shown   in   Figure   V-19.
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-21.  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  waste
water   pumped   from  the  mine  contains  the  same  waste
parameters as that from the preceding subcategory  but  also
contains  substantial  soluble metals.   Table V-22 shows the
                           239

-------
       Figure V-19. WATER FLOW DIAGRAM FOR MINE 3105
SEEPAGE
  DRILL
COOLING'
  WATER   270 m^/day
          (72,000 gpd)
                            MINE
                           PUMPING
                               7,600 m3/day
                               (2,000,000 gpd)
                    (Ml
MILL FEED-WATER
   RESERVOIR
                      FUEL AND LUBRICANT
                      SPILLAGE AND
                      LEAKAGE
                      EXPLOSIVE
                      WASTE PRODUCTS
                             240

-------
TABLE V-21. 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
IDS
COD
TOC
Oil and Grease
P
NH3
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
<0.02to0.90
37 to 63
3 to 57
0.3 to 1.2
          Value in pH units
          Data may reflect influence of acid stabilization on sediment
                        241

-------
TABLE V-22.  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
IMH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
i 	 	 	
CONCENTRATION (mg/£ )
IN RAW MINE WATER
3.0 to 8.0*
14.6 to 167
178 to 967
< 2 to 58
260 to 1,722
15.9 to 95.3
1 to 11
Oto3
0.020 to 0.075
< 0.05 to 4.0
0.0001 to 0.001 3
< 0.0001 to 0.0001
0.1 to 0.3
1.38 to 38.0
< 0.02 to 0.04
0.01 6 to 0.055
0.17 to 0.42
< 0.02 to 57 .2
0.12 to 2.5
48 to 775
< 0.01 to 220
0.06 to 0.80
             *Value in pH units
                         242

-------
 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 water (or water vapor)  as the limiting condition:

 (A)  Assuming  that  the  process  takes  place  in   a   dry
     environment,   an  equal amount of sulfur dioxide will be
     generated with the formation of  (watersoluble)   ferrous
     sulfate:

          FeS_2 + 302     	>   FeS04 + S02

 (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:

          2F6S2. +  2H20 + 70J2   	>  2FeS04  + 2H2SO<4

 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:

 4FeS04 +  2H2S04 + 02    —>    2Fe2(S04)3 +  2H20

 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
 ferrobacillusf  Thiobacillus   sulfooxidans)^   are  involved~IH
the above  reaction   and,  at   least,  are  responsible   for
 accelerating  the  oxidation  of  ferrous iron to the ferric
 state.

Reaction 3_—Precipitation of Iron
                          243

-------
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_(SOU)J + 6H20   --- >   2Fe(OH)3 + 3H2S04

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:

    Fe2 (S04)jl + FeS2 + H2O  --- >   3FeSOU + 2S

              S + 30 + HO   --- >   R2SQ
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 O.U5 micrometer  (0.000018  inch)
in  size  and,  thus, appear as soluble metals under current
analytical practice.

Conditions compatible with solubilization of certain metal s-
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-
                           244

-------
20.    The   characteristics   of   mine  waters  from  this
subcategory are illustrated by Table V-22,  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 waste water 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.

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.

The  raw  and  treated  waste  characteristics of four mills
visited during this program are  presented  in  Table  V-23.
                           245

-------
        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)
                                              FUEL AND LUBRICANT
                                              SPILLS AND LEAKAGE
                                              EXPLOSIVE WASTE
                                              PRODUCTS
                             246

-------
                        Figure V-21. FLOW DIAGRAM FOR MILL 3103
                                                                    WATER
                                                                  FROM MILL
                                                                 FEED RESERVOIR
                                                                       9,500 m3/diy
                                                                       (2,500.000 gpdl
        Zn SCAVENGERS  Zn ROUGHERS  <   Zn CONDITIONER
TO TAILING-
POND SYSTEM
                                  TO STOCKPILES

                                       (b) MILL PROCESS
                                            247

-------
 TABLE V-23. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE
            LOADS FOR MILLS 3102, 3103, 3104, 3105, AND 3106
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
MBAS Surfactants
P
Ammonia
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Cyanide
Sulfate
Chloride
Fluoride
RANGE OF
CONCENTRATION
lmg/2 I
IN WASTEWATER
lower limit
7.9-
26
310
<2
670
71.4
11
0
0 18
0042
<005
< 00001
<0 1
0 12
<002
0005
<002
<002
005
< 001
295
21
0 13
upper limit
88-
609
1.760
108
2.834
1.535
35
8
3.7
0 150
14
0 1
1 9
046
036
0.011
067
008
0 53
003
1 825
395
0 26
RANGE OF RAW WASTE LOAD
per unit ore milled
kg/1000 metric tons
lower limit
_
410
460
7
940
6
635
5
0236
0 108
0064
<0 00013
<0127
0089
<0026
0008
<0026
<0026
0064
< 0013
130
20
0370
upper limit
_
1.600
4.700
285
8,500
4.800
130
21
13
0876
26.4
00026
69
17 2
0158
0018
1 77
0290
1 16
0109
4.800
870
0944
lb/1000 short tons
lower limit
_
820
920
14
1.840
12
13
10
047
021
0 125
< 0.00026
<0.25
019
<0052
0016
< 0.052
<0052
0 000129
< 0026
260
40
074
upper limit
_
3.200
9.400
570
17.000
9.600
260
42
26
1 76
528
00052
138
344
0316
0036
344
0580
232
0218
9.600
1,740
1 88
per unit concentrate produced
kg/1000 metric tons
lower limit
_
1.450
2.290
30
4,800
30
30
30
205
054
032
< 000168
<0900
062
< 0 18
<0 18
< 0 18
< 045
0012
0.091
1,260
210
203
upper limit
_
10.200
32,500
2,000
50,900
50,000
580
130
607
254
185
0130
322
860
1 96
885
1 36
100
0198
0509
33,700
4.070
545
lb/1000 short tons
lower limit
_
2,900
4,580
60
9,600
60
60
60
570
1.08
064
< 000336
< 1 8
1 24
< 036
< 036
<036
< 0.90
< 0.024
0 182
2,520
420
4.06
upper limit
_
20,400
65,000
4.000
101,800
100,000
1,160
260
121 4
508
370
0.260
644
172
392
17 7
272
20
0396
1 18
67,400
8,140
10.9
•Value in pH untts
                              248

-------
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
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.  Wash water is nor-
mally  combined  with  the  process   waste   effluent   out
constitutes  only  a  small  fraction of 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 dredgind 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 materials or the
mine and is normally discharged  separately  from  the  mine
effluent.
                            249

-------
 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
 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.

 Waste   water  from a placer operation is  primarily water  that
 was used in  a gravity  separation process.   Where a  placer
 does  not  occur  in a  stream, water  is used  to fill a pond on
 which a  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 constitutes the
 final discharge.   The  principal  waste  water   constituents
 from placer operations are high suspended  solids.

 Waste water 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
                           250

-------
Figure V-22.   WATER  FLOW IN  FOUR SELECTED GOLD
                MINING AND MILLING OPERATIONS
                           (a) MINE/MILL 4101
UNDERGROUND
MINE

3317 m3/d«Y
DISCHARGE TO STREAM


' 2.290 m3/d.y ~"
(600,000 gpd)
AMALGAMATION
MILL*

                                                   DISCHARGE TO STREAM
       •AMALGAMATION OF GRAVITY-SEPARATED SANDS; FINES AND GOLD-EXTRACTED SANDS
        ARC FLOATED FOR RECOVERY OF BASE METALS.
                          (b) MINE/MILL 4102
                                     4,947 m3/d.¥
                                     11,296.000 gpd)
     'GOLD VALUES PRESENT IN BASE-METAL CONCENTRATES. RECOVERED AT SMELTER OR REFINERY.
                          (c) MINE/MILL 4103
UNDERGROUND
MINE


DISCHARGE
IMPOUNDED
                                      964 m3/div
                                      (250.000 gpd)
      '824 m3/d«v
       (210,000 gpd)
1,908 m3/d.y
1600.000 tpd)
                                                     INTERMITTENT DISCHARGE
                                           DURING APRIL AND MAY
                          (d) MINE/MILL 4104
                               251

-------
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  but,  in  the past, has included hydraulic and
drift mining of buried placers  too  deep  to  strip.   Lode
deposits are mined either by either underground  (mines 4102,
4104,  and 4105) or open-pit (mine 4101)  methods, the parti-
cular 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.

The  chemical  composition  of raw mine effluent measured at
two of the 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).

Process Descriptions - 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.
                           252

-------
    TABLE V-24. CHEMICAL COMPOSITION OF RAW MINE WATER FROM
                MINES 4105 AND 4102
PARAMETER
pH
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 Fa
Pb
CONCENTRATION (mg/£)
MINE 4105

275
34f
2.40
1,190
1,176
14
733
35.01
12.0
1
0.095
<0.2
0.03
< 0.002

-------
 In general,  the  cyanidation process  involves   solubilization
 of  gold with cyanide  solution,  followed by precipitation of
 gold from solution with  zinc dust.   (See Figure  m-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
constructed   pad  to  support  the  rock  and  collect   the
 leachate.
                          254

-------
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  gold-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 1104 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 410U are further
processed  by  the  cyanidation/agitation-leach  process  to
recover residual gold values.

In  addition  to  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 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  a  potential  solubilization  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 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.   Calcium, sodium, potassium, and
magnesium are  found at concentrations 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
                           255

-------
TABLE V-25. PROCESS REAGENT USE AT VARIOUS MILLS BENEFICIATING
           GOLD ORE
MILL
4105
4105
4101
4102
MILL PROCESS
Cyanidation/Leach
Cyanidation/Char-in-pulp
Cyanidation/Agitation Leach
Amalgamation
REAGENT CONSUMPTION
CYANIDATION
kg/metric ton
ore milled
0.13
0.58
0.18
-
Ib/short ton
ore milled
0.26
1.16
0.35
-
AMALGAMATION
kg/metric ton
ore milled
_
	
	
0.001
Ib/short ton
_
_
_
0.002
         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
Cu FeS
PbS
ZnS
CdS
HgS
(Fe, Ni)gS8
Au Te2
(Au, Ag) Te2
Au
Se*
                  'Accompanies sulfur in sulfide minerals
                          256

-------
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/"carbon-
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.

Water  flows  of  some  silver mining and milling operations
visited during this program are presented in Figure V-23.
                           257

-------
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
          MILLING OPERATIONS (Sheet 1 of 2)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
4105
(Vat Leach)
4105 (Carbon
in-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Leach)
(Vat Leach)
4105 (Carbon
m-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
4105
(Vat Laaeh)
tn-Pulp)

MINE/MILL
4102
(Amalgamation)
(Agitation Laach)
(Vat Laach)
in-Pulp)

CONCEN-
TRATION
(mg/i)
495,000
545,000

485,000

CONCEN-
TRATION
(mg/2)
34.3
500

97.0

CONCEN-
TRATION
(ma/£l
0.03
0.17

2.0

CONCEN-
TRATION
(ma/Hi
1.5
<0.5

77.0
TSS
WASTE LOAD
in kg/IOOO metric tons
(lb/10OO short tons)
of concentrate produced
61.695,315,000
(123,390,630)
11,541,465.000
(23.082,930.000)
-
4.7 x ID!]
9.4 x 1011
in kg/1000 metric tons
(to/1000 short tons)
of ore milled
2.871,000
(5,742,000)
436,000
(872,000)
_
4. 17 1.OOO
(8.342,0001
TOC ' 	
WASTE LOAD
in kg/1000 metric tons
lib/1000 abort tonil
of concentrate produced
4,275,000
18,550,000)
1,059,000
(2,1 18.000)
-
94,100,000
1188.200,000)
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
199
1398)
40
(80)
-
830
11,660)
Cu
WASTE LOAD
m kg/1000 metric tons
(lb/1000 abort tonil
of concentrate produced
3,740
(7,480)
3.600
(7,200)
~
1,941,000
(3,882,000)
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
0.2
(0.4)
01
(02)
-
17
134)
Fe
WASTE LOAD
in kg/IOOO metric tons
(lb/1000 short tons)
of concentrate produced
187,000
(374,000)
< 10,600
K2 1.2OOI
-
74,700,000
(149.400.000)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
8.7
(17.4)
•>0.4
K0.8)
-
660
(1,320)
TDS
CONCEN-
TRATION
(mg/£l
462
4.536
-
886
WASTE LOAD
in kg/1000 metric tons
(lb/1000 short tons)
of concentrate produced
19,942.000
(39,884,000)
96,060,000
1192,120,000)
-
859300.0OO
(1,719,800,000)
in kg/1000 metric tons
(lta/1 000 short tons)
of ore milled
930
(1.860)
3,600
	 (7,200) 	

7.600
(15.200)
COD
CONCEN-
TRATION
Img/dl
1142
43
-
17894
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
1,423,000
(2,847 000)
911,000
11.822,000)
_
173,700,000
(347,400.000)
in kg/1000 metric tons
(Ib/IOOO short tons)
of ore milled
66
11321
34
1681
-
1,540
1 3.080)
As
CONCEN
TRATION
(mg/)l)
<0.07
005
3.5
-
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
< 8.700
107,400)
106
1212)
1,510.000
(3,020,000)
-
in kg/1000 metric tons
(lb/1000 short tons)
of ore milled
<0.4
K0.8I
0.04
10.081
14
(281
-
Zn
CONCEN-
TRATION
(mg/dl
1.3
3.1
-
0.22
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
162,000
1324,000)
65.600
(131,200)
-
213,000
(426,000)
in kg/1000 metric tons
(Ib/IOOO short tons)
of ora milled
7.5
(15.1)
2.5
(51

2
(41
                          258

-------
• ABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
          MILLING OPERATIONS (Sheet 2 of 2)

MINE/MILL
4102
(Amalgamation)
4101
(Agitation LMCh]
4105
IV.t LMCh}
4105 (Carbon
m-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
4105
(Vat Laach)
4105 (Carbon-
in-Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laachl
4105
(Vat Laach)
4105 (Carbon
in-Pulp)
Pb
CONCEN-
TRATION
(mgU)
<0.1
<0.1
—
<0.1
WASTE LOAD
in kg/1000 matric torn
lib/1000 Ihort tonil
of concantrata producad
<12,500
K 25.000!
< 2.1 00
K 4,2001
—
< 97.000
«1 94,000)
in kg/1000 matric toni
(t>/1000 short tons)
of ora millad
<0.6
kl.21
<0.08
K 0.1 61
-
<0.9
K1.8I
Hg
CONCEN-
TRATION
Img/i)
0.0011
—
0.004
0.0042
WASTE LOAD
in kg/1000 matrk tons
(Ib/IOOOahorttons)
of concantrata producad
137
(274)
_
1,700
(3.400)
4,070
(8.140)
in kg/IOOO matric toni
lfc/1000 ihort torn)
of on millad
0.0064
(0.0128)
-
0.016
10.032)
0.036
10.072)
SULFIDE
CONCEN-
TRATION
(mg/il
<0.5
<0.5
0.2
1.7
WASTE LOAD
in kg/1000 matric tom
(Ib/IOOOihorttoiw)
of concantrata producad
< 62.000
K 124.000)
< 10.600
K21.200)
86.000
(172,000)
1.650.000
(3.300.000)
in kg/1000 matric ton.
(fc/1 000 short tons)
of ora millad
<2.9
K5.8I
•c-0.4
K0.8)
0.8
(1.6)
15
(30)

CONCEN-
TRATION

<0.02
0.10
<0.01
<0.02
Cd
WASTE LOAD
in kg/1000 matric torn
(lb/1000 ihort torn)
of concantrata producad
< 2,500
K 5.000)
2.100
(4700)
< 4.300
K8.600I
< 19.400
« 38,800)
in kg/1000 matric tons
lib/1000 short toni)
of ora millad
<0.1
K0.2)
0.08
(0.16)
•C0.04
X0.08)
<0.17
K0.34)
CYANIDE
CONCEN-
TRATION
(mg/fcl
<0.01
5.06
-
0.06


WASTE LOAD
in kg/1000 matric tons
I to/1000 ihort torn!
of concantrata producad
< 1.250
K2.500)
107,000
(214.0001
_
58,000
(116,000)
in kg/1000 matric tons
(lb/1000 short tons)
of ora millad
<0.06
K0.12)
4
18)
-
0.52
(1.04)

                            259

-------
Figure V-23. WATER FLOW IN SILVER MINES AND MILLS
                     1.132 mj (299.060 pl)/d«y
                     (a) MINE/MILL 4401
UNDERGROUND
MINE
DISCHARGE
2,933 m3/day
(775,000 gpd)
( r-.acfK \
V ./ 545 m3/
^ 	 (144.00C

^ FinTATinM
day "~ MILL
L_


RAIN
12m3/day
1 ' (3,340 gpd)
<-^ ^--^ _ __.

1.500 m3/d.y *^^ POND ) , " , »• EVAPOflATION
(390.000 ypd) 	 	 (715 to 914 gpd)
954 n.3/day
(252.000 gpd)
                   (b) MINE/MILL 4402
                  260

-------
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.  Waste water  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 into either a barge pond  or an  onshore
settling  pond.  The effluent of the settling pond usually is
combined  with  the barge pond discharge,  and  this comprises
the  final discharge.  The principal waste  water   constituent
from  any  placer operations, whether silver,  gold, or other
materials, is high loadings of  suspended  solids.

Waste water 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   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
mineralization  of  the  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
                          261

-------
 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)   ama Igamat ion.

 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
 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
                         262

-------
           TABLE V-28. RAW WASTE CHARACTERISTICS OF SILVER
                      MINING OPERATIONS
PARAMETER
pH
Acidity
Alkalinity
Color
Turbidity (JTUI
TOS
TDS
TSS
Hardness
COO
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
47t
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
<5t
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
Te
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
6.3
0.0004
0.06
<0.1
<02
14.5
0.068
<0.02
12.0
0.21
<0.3
<0.5
0.03
<0.2
<0.2
1.15
338
0.10
0.25
<0.01
<0.01
0.21
-
<0.5
13.0
•Value in pH units

 Value in cobalt units

TOS = Total Solids
                            263

-------
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
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
                           264

-------
 some extent in the waste water  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.

 Bauxite Ores

Water  handling  and  quantity  of  waste  water  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,
                          265

-------
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)i2Sb4S13
(Cu, Fe,Ag)12As4S13
PbS
ZnS
Cu FeS2
FeS
Ag2S
CdS
(Mg, Ni) O- Si O2 • x H2O
(FefNi)9S8
Bi
Ag2S
Ag5 Sb 84
Sb2S3
                266

-------
TABLE V-30. FLOTATION REAGENTS USED BY THREE MILLS TO BENEFICIATE
          SILVER-CONTAINING MINERAL TETRAHEDRITE (MILLS 4401 AND
          4403) AND NATIVE SILVER AND ARGENTITE (MILL 4402)

REAGENT


M.I. B.C. (Methylisobutylcarbinol)
D-52
2-200 (Isopropl ethylthiocarbamate)
Lime (Calcium oxide)

Sodium cyanide

PURPOSE

CONSUMPTION
g/metric ton
ore milled
MILL 4401
Frother
Frother
Collector
pH Modifier
and Depressant
Depressant
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
Cresylicacid
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
Ml LL 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
                            267

-------
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
           AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 1 of 2)

MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
41 OS
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
TSS
CONCEN-
TRATION
(ing/ HI
550,000
203.000
90.000

WASTE LOAD
in kg/1000 matric tons
(lfa/1000ihorttoiul
of concantrata producad
99.000.000
(198.000.000)
33.901.000
(62.802.000)
9.720.000
(19,440.000)
-
HI kg/1000 matric tona
Id/1000 Ihort torn)
of or. millad
2.475.000
(4.950.OOOI
1.543.000
(3.086,000)
990.000
(1.980.000)
-
TOC
CONCEN-
TRATION
(mo/KI
22.0
24.0
29.0

WASTE LOAD
in kg/1 000 matric toni
lib/1000 short tons)
of concantrata producad
4.000
(8.000)
4.000
(8,000)
3,130
16.260)
-
m kg/1000 matric tona
(*>/1000 short tons)
of ora millad
100
(2001
180
(360)
320
(6401
-
Cu
CONCEN
TRATION

-------
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE  LOADS
           AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 2 of 2)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
4401
4403
4402
4105
(Company Data
only)
Hg
CONCEN
TRATION

<2
K4I
-
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<009
K018I
<015
K030I
<0.2
(CO 4)
-
Mo
WASTE LOAD
in kg/1000 metric tons
lib/1000 short tons)
of concentrate produced
<36
«72I
<30
K60I
58
11161
< 32,400
K64.800)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled
<09
«1 8)
<4
«8I
6
(121
<0.3
(<06)
Cd
CONCEN-
TRATION
lm«/£)
<002
<002
<002
<0.01
WASTE LOAD
in kg/1000 metric tons
(lb/1000 short tons)
of concentrate produced
<3.6
K7.2)
<3.3
(<6.6)
<2.2
K44)
< 6.500
K 13,000)
in kg/1000 metric tons
lib/1000 short tons)
of ore milled

<2
K4I

                          269

-------
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 Waste water

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
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  locations.  The open-pit mining technique is
largely  responsible  for  accumulation   of   this   water.
Underground mining accounts for only a fraction of a percent
                             270

-------
 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
 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
                           271

-------
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
2.76 m3/metric ton
(664 gal/short ton)
BAUXITE
         Q SETTLING PONDS  )
          REFINING:
              COMBINATION PROCESS
                    2.76 m3/metric ton
                  i ' (664 gal/short ton) BAUXITE
                                             WATER TREATMENT
                                                   PLANT
                                                       2.76 m3/metric ton
                                                       (664 gal/short ton) BAUXITE
                                           Q  SLUDGE POND  J
  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 tort) BAUXITE
                                                DISCHARGE

-------
 mined.   A room-and-piliar technique is then used to support
 the mxne 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 charges 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.    Underground   mines
 receive   drainage  which   has migrated   through  strata  of
 unconsolldated  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
 subcategoriz ation.

 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.
                           273

-------
The  class  of  raw  mine  drainage corresponds closely with
raining    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  4
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  underground  mines,
may  migrate  through  the  lignitic clays located in strata
overlying the mines before collecting in the mines,  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,
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.
                            274

-------
TABLE V-32. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID RAW
           MINE DRAINAGE FROM OPEN-PIT MINE 5101
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
IDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
2.8*
1,000*
397
0
560
<2
7.2
3.5
23.8
0.82
0.3
500
0.29
INDUSTRY DATA
* *#
3.4 •
250
0
617
2
15.4**
59.6**
5.9**
25.3**
0.31
490
0.048
NPDES PERMIT
APPLICATION
3.5*
1,903f
—
40
1,290
10
7.0
4.2
38
1.0
0.37
700
1.4
 "Value in pH units
Value in micromhos
''Averages of five samples
  TABLE V-33. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID
             RAW MINE DRAINAGE FROM OPEN-PIT MINE 5102
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sr
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
3.2*
1,580*
782.0
0
1,154
< 2
64.0
7.7
88.0
0.36
0.063
0.1
887.5
0.59
INDUSTRY DATA**
2.8*
2,652t
533
-
-
416
62.2
-
44.6
-
-
-
726
-
NPDES PERMIT
APPLICATION
3.0*
2,000t
—
0
96
1,280
20.6
9.0
51.0
0.8
0.01
—
226
0.26
*Value in pH units    ^Value in micromhos   **Averages of eight or more grab samples taken in 1974
                               275

-------
TABLE V-34. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ALKALINE RAW
          MINE DRAINAGE FROM UNDERGROUND MINE 5101
PARAMETER
PH
Specific Conductance
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sr
Sulfate
Fluoride
CONCENTRATION (mg/£)
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
                              276

-------
 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   liquor.   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.
 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
                          277

-------
     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/ 2, )
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
< 0.01 2 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 waste water = 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)
                                     278

-------
      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
15,51 2 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)
O

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
                                     279

-------
 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  waste  water  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++,  Na+,  K+,
 Mg++,  C0_3—,  and  SO^.   accounting  for nearly  all dissolved
 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.
                           280

-------
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 waste water 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
 surveyed contributes  0.8 ton of contaminated  scrubber  bleed
water per ton  of ore processed.

 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
 waste   water   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.
                           281

-------
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  NH_3,
         N0.3, 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
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

         Mri   5.5
                            282

-------
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
(m /min (gpml)
2.66 (700)
6.43(1,700)
34.06 (9,000)
11.36 13,000)
pH
4.5.
7.0
6.5
7.3
CONCENTRATION (ms/£)
Oil and
Gr*aw
14
1.0
2.0

Nitrate

0.15
0.12

Fluoride
44.5
4.5
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
021
6.8
Mo
0.5
<0.1
<0.1
<0.1
Pb
O.CX
0.19
0.14
-
V
<0.5
<0.5

-------
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
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.
                          284

-------
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 26 and 27). 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
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-
                           285

-------
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.

Process   Description  and  Raw-Waste  Characterization  For
Specific Mines and Mills Visited

Mine/Mi11 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.
                          286

-------
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 MoS2:.

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.
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 million 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 MoS^.    Tungsten
                            287

-------
                            Figure V-25. MILL 6601 FLOWSHEET
                        22.000 m3/d§y
                        (6.000.000 gpd)
                                                 -UNDERFLOW -1
                                       OVERFLOW
           -TAILS



, - ,.
—MIDDLINGS 	
J
ROUGHER FLOAT

MIDDLINGS
SCAVENGER
FLOAT
                                                          -CONCENTRATE-
                                     (4 STAGES WITH
                                      REGRIND AND
                                   INTERNAL RECYCLE)
                                         TAILS
                                                      -CONCENTRATE—I

                                                          MIDDLINGS-
                                              CLEANER
                                               FLOAT

                                           (6 STAGES WITH
                                            REGRIND AND
                                         INTERNAL RECYCLE)
                                                                                           -TAILS-
                                                                           CONCENTRATE
                                                      UNDERFLOW
    M+UNDERFLOW
     •^•UNDERFLOW

     22.000 m3/*y
     (•,000,000 gpd)
2,200 m3/
-------
TABLE V-39. REAGENT USE IN MOLYBDENUM MILL 6101
REAGENT
Lime
Vapor Oil
Pine Oil
Hypo (Sodium Thiosulfate)
(Na2 $203 • 5H2O)
Phosphorus Pentasulfide (?2 85)
MIBC (methyl-isobutyl carbinol)
Sodium Cyanide (Na CN)
DOSAGE
g/ 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
TDS
Oil and Green
COD
At
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Total Cyanide
Fluoride
CONCENTRATION
(mo/Jt )
IN WASTEWATER
600JOOO
2,598
2.0
136
0.01
0.74
51
56.5
5.3
9.8
76.9
1,305
0.02
6.2
TOTAL WASTE
ke/day
14,000,000
42,000
32
2.200
0.16
12
820
900
85
160
1^00
21,000
0.32
99
lb/d.y
32,000.000
92,000
70
4300
0.35
26
1300
2,000
190
350
2,600
46,000
0.70
220
RAW WASTE LOAD
per unit ore milled
kg/metric ton
995
3.0
O.O023
0.16
0.000012
0.00086
0.059
0.064
0.0061
0.011
0.086
1.5
0.000023
0.0071
Ib/short ton
1390
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
1330
1.4
96
0.0070
0.52
36
39
3.7
7.0
52.4
915
0.014
4.3
Ib/ihort ton
1,200,000
3,670
2.8
190
0.014
1.0
72
79
7.4
14.0
105
1330
0.028
8.7
                   289

-------
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  of 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  MoS^,
is  fed  to four stages of further flotation.  Reagents used
in the primary flotation step are summarized in Table  V-m.
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 4.5, 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,
                           290

-------
Figure V-26. SIMPLIFIED MILL FLOW DIAGRAM FOR MILL 6102

CRUSHING
(3 STAGES)
28% ••• 3 MESH
\

GRINDING IN
BALL MILLS
1
36% + 100 MESH
*|
t


FLOTATION



* *


FLOTATION
1
96% OF MILL FEED


TO «
TAILINGS | /~^
V9)

MONA7ITF _--
^ CONCENTRATE
V

RAVITY SEPARATK
lUMPHREY'SSPIRAI
\
PYRITE
FLOTATION
CLEANER Tn
FLOTATION 	 ^- '" ...-.«.
(4 STAGES) TAILINGS
1

DRYING
,
Y
CONCENTRATE
(93% + MO^J

TAILS
TABLES


i^ /O\ INDICATES 1
MONAZITE
FLOTATION


"*~V°)
MAGNETIC
SEPARATION
\< V


NONMAGNETIC MAGNETIC
TIN TUNGSTEN
CONCENTRATE CONCENTRATE
                           291

-------
   TABLE V-41. REAGENT USE FOR ROUGHER AND SCAVENGER
             FLOTATION AT IV ILL 8102
REAGENT
Pin* oil
Vapor oH
Syntax
Lime (Calcium oxide)
tedium silicate
Nokes reagent
PURPOSE
Frother
Collector
Surfactant and Frother
Adjustment of pH to 8.0
Slime Dispenent
Lead Depfeasant
CONSUMPTION
kg/metric ton
ore milled
0.18
0.34
0.017
0.15
0.25
0.01S
Ib/ihortton
OM miMed
0.35
0.«7
0.034
0.30
0.50
0.03
TABLE V-42. REAGENT USE FOR CLEANER FLOTATION AT MILL 6102
RfAQfNT
Vapor oil
Sodium cyanide
Nokas reafurt
Dowfroth250
Valco 1861
PURPOSE
Collector
Pyrite and Chatco-
pyrite Depreuawt
Lead Deprewant
Frother
Fteocuiant
CONSUMPTION
kg/metric ton
oremiMed
0.45
0.13
0.45
0.015
0.003
Ib/ihortton
nm ijiiHari
vw IVIICT^V
0.90
OJ5
O.M
0.03
0.006
                      292

-------
  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-TAI LINGS SAMPLE)
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
(mg/i) IN
WASTE WATER
150.000
2,254
4
23.8
<0.1
0.19
21.0
50
17.5
2.1
25.0
1,500
11.7
0.45
TOTAL WASTE
kg/day

200,000
360
2,100
<9
17
1,890
4,500
1,600
190
2,250
135,000
1,100
41
Ib/day

440,000
790
4.600
<20
37
4,200
9.900
3.500
418
4,950
300,000
2/.00
90
RAW WASTE LOAD
per unit ore processed
kg/metric ton
998
4.7
0.0080
0.049
< 0.0002
0.00040
0.047
0.10
0.037
0.0044
0.052
3.1
0.026
0.00095
Ib/short ton
1996
9.3
0.016
0.098
<0.0004
0.00080
0.088
0.21
0.074
0.0088
0.10
6.3
0.052
0.0019
per unit total
concentrate produced
kg/metric ton

2,700
4.6
28
<0.1
0.23
25
58
21
2.5
30
1,800
15
0.55
Ib/short ton

5,400
9.2
56
<0.2
0.46
50
120
43
5.0
60
3,600
30
1.1
                     293

-------
 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-
 44 and V-45.

 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
 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
                           294

-------
    TABLE V-45. CHEMICAL CHARACTERISTICS OF ACID-FLOTATION STEP
PARAMETER
PH
Cd
Cu
Fe
Mn
Mo
Pb
CONCENTRATION (mg/ A ) AT INDICATED POINTS OF FIGURE V-26
PYRITE FLOAT (19)
4.5«
0.01
02
4.2
4.0
3.0
0.3
MONAZITE FLOAT (20)
1.5
0.042
05
490
53.3
4.0
1.34
•Value in pH units
                          295

-------
Figure V-27. INTERNAL WATER FLOW FOR MILL 6104 THROUGH
         MOLYBDENUM SEPARATION

WET
ORE



WATER
FROM
CREEK





WA
FR
Ml




121 m3/di
132400*

ren
NE
461 m3/*y
(127.000 ffd

i
«" CRIM
)l— *• Al
ORIN
(782400 gpd)
MING (614000 91
EN NO
4H m3/*v (121,000 fpd)
'


16-MtTER
(60-FT)
THICKENER
i i
136 m3/OWN
23 m3Afcy
MoO3
STO

E J2.6 m3/dw
|(700gpd) '
YING SO,
ISTING ^ SCRUBBER
|
) '

i 6,076 m
(1.606,01
T
TO TO
CKPILE TAILING POND
1.336 m3Ahy
(363.000 gpd)
•/diy
**' to SCHEELITE
FLOTATION
UNDE

IFLOW
492m3/.


CONCEI
THICK

PER
WRATE
ENER
4 MO m /ifav
0,162,000 |pd)
toy (130400 gpd)
96 m3/diy
(26400 gpd)

Cu CONCENTRATE PRODUCT
TO STOCKPILE
114 m3/d«y
130400 gpd)
1
FILTI
WAS


299m3/
(79,000
RING
4D
KING
toy
H>d)
1,106 m3/d«y
(292.000 gpd)

2*9m3/dw
(79400 gpd)
MOLYBDENUM
P SEPARATION



3/diy
Mgpd)

409m3/d>y
(106,000 gpd)


TO
» SOLVENT
EXTRACTION
(FIGURE V 28)
                     296

-------
       Figure V-28. INTERNAL WATER FLOW FOR MILL 6104
                 FOLLOWING MOLYBDENUM SEPARATION
                                              TO ATMOSPHERE
FROM MOLYBDENUM
  SEPARATION
  (FIGURE V-27)
                          297,

-------
 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-U6, V-47,  V-48, V-49, and  V-50.

 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  111-18,   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 aguifer  producing  mine
water,  which must be intermittently pumped out (for approxi-
mately  hour every 12 hours).   Total effluent volume is less
                           298

-------
   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
CONCENTRATION

-------
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
                         300

-------
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
WASTE WATER
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
to/short ton
340
2.2
26

-------
 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/i) 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
O.OS2
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 8106
PARAMETER
pH
TSS
TDS
Oil and pease
A*
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Ni
CONCENTRATION
Una/ JO
IN WASTEWATER
8.6*
226.9
212
3.4
< 0.07
< 0.005
<0.03
0.53
0.5
<0.1
0.05
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/day
-
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
metric torn
—
790
730
12
<0.2
<0.02
<0.1
1.8
1.7
<0.4
0.17
84
1.4
lb/1000 short toni
—
1,600
1,500
24
< 0.4
<0.04
< 0.2
3.7
3.5
<0.9
0.35
170
2.8
par unit concentrate preduced
kg/1000
metric 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
•Value in pH units.
                               302

-------
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
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  hydr©metallurgical
process involving roasting,  leaching,  solvent  extraction,
and  precipitation.    The  process  is illustrated in Figure
III-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
pelletized.   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 (Na^VKK)^) is concentrated  by  solvent
extraction,   and  ammonia  is  added to precipitate ammonium
vanadate, which is  dried  and  calcined  to  yield  a  V205
product.
                           303

-------
                                      Figure V-29. WATER  USE AND WASTE SOURCES  FOR VANADIUM  MILL 6107
U)
o
                          NOTE
                     RUNOFF FROM RAIN
                     IS NOT CONSIDERED
                     EXCEPT WHERE IT
                     ENTERS THE PROCESS
                     ©
   -- SAMPLE NUMBER

SAMPLES (7l;AND (72JARE MINE-WATER SAMPLES

-------
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-5U, 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
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 is accomplished 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),
due to open during 1975, will employ 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.
                          305

-------
TABLE V-53. WASTE CHARACTERIZATION AND RAW WASTE LOAD FOR MILL 6107
          LEACH AND SOLVENT-EXTRACTION EFFLUENT (SAMPLE 80)
PARAMETER
pH
TDS
Otl and grease
COD
NH3
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fe
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/SL)
IN WASTEWATER
3.5*
39,350
94
475
0.16
0.35
0.037
1.15
0.15
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/day
-
180.000
440
2.200
0.75
1.6
0.17
5.3
0.7
240
< 0.4
< 0.2
140
2.4
1.2
950
37.000
21
120,000
RAW WASTE LOAD
per unit ore milled
kg/metric ton
-
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/short 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
                              306

-------
   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
C0.00004
< 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
                                 3<"»-i
                                 u/

-------
  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
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Chloride
Fluoride
Sulfate
CONCENTRATION
(mg/£)
INWASTEWATER
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
                              308

-------
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
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.

Waste  water  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 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.   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
                          309

-------
Figure V-30. WATER FLOW IN  MERCURY MILLS 9101 AND 9102
  (NO DISCHARGE)
^S
MILL
RESERVOIR
	 — -^

' 16.4 m3/day
(4,320 gpd]
GRAVITY-
SEPARATION
MILL
t
fc ( TAILINfi
" \ POND
^T
  (NO DISCHARGE)
                        1,649 m3 day (432,000 gpd)


                          (a) MINE/MILL 9201
MILL
WELLISI
< 	 -^

/ 1.63 m3/min
(430 gpm)
FLOTATION
MILL
I
k
^f TAILINR A
5.4 m3/min (1,430 gpm) \^^ POND J^
x-^ ^
__to/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
                             310

-------
toxicity,  but  also because of its effect on the solubility
of the waste constituents.

Quantities of Wastes

The few mercury operations still active in late  1974  were,
for  the most part, obtaining 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.


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.  Stibnite 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 U.S. mining.

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
                           311

-------
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 new
plant (mill 9202)  due to begin production during early  1975
was  contacted.   This mill will use a flotation process and
expects to discharge 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  waste water 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
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) due to open
during 1975 will use a flotation process, which will require
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 which are expected to be 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.
                            312

-------
TABLE V-56. EXPECTED REAGENT USE AT MERCURY-ORE FLOTATION
          MILL 9202
REAGENT
Dowfroth 250 (Polypropylene glycol methyl ethers)
Z-11 (Sodium isopropyl xanthate)
Lime (Calcium oxide)
Sodium silicate
PURPOSE
Frother
Collector
Depressing
Agent
Depressing
Agent
CONSUMPTION
kg/metric ton
ore milled
0.15
0.13
0.05
0.10
Ib/short ton
ore milled
0.30
0.25
0.10
0.20
                       313

-------
 Sometime  during   the  spring  or early summer of  1975, mill
 9202  is to begin operation for the concentration of  mercury
 sulfide ore by a froth flotation process.

 Waste characteristics  of  mill  effluents of the operation
 visited and of a pilot-plant operation using  the  flotation
 process 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
 the mill and then rejected to an impoundment, from which  it
 is  removed  by  evaporation  and,  possibly, 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.    Waste  water 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  in  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 these considerations,  the effluent quantities are  also
                            314

-------
TABLE V-57. WASTE CHARACTERISTICS AND RAW WASTE LOADINGS
          AT MILLS 9201 AND 9202
pH
MILL inpH
unto
•201 6.5
9202
(Pilot Operation! ~

MILL
9201
9202
(Pilot Operation)
MILL
9201
9202
(Pilot Operation)
MINE
9201
9202
(Pilot Operation)
H9
WASTE LOAD
CONCEN-
TBAT,I,PN i" kg/1000 metrie font
In* HI (lb/1000 short tons)

v virisuv ii aae. f»>
-
0.0072 11
122)
in kg/tOOO metric tont
(b/1000 short tons)
of on milled
-
0.094
(0.188)
So
CONCEN-
TRATION
(mg/i)
<0.5
0.03
WASTE LOAD
in kg/1000 nwtrie torn
(lb/1000 short tons)
of concentrate produead
< 6.900
l<1 3,800)
SO
(100)
in kg/1000 nwtrie toni
(to/1000 Ihort tont)
of on milM
<0.06
K0.10)
0.4
10.8)
Tt
CONCEN-
TRATION
(mg/i)
<0.08
"
WASTE LOAD
in kg/1 000 metric tons
(lb/1000 short tons)
of concentrate produoid
< 1.100
K2.200)
-
in kg/IOOO nwtrie tons
(b/1000 ihort tom)
of Of. milwd
< 0.008
(< 0.01 61
-
Zn
CONCEN-
TRATION
(mg/£)
0.14
"
WASTE LOAD
in kg/1000 nwtrie tons
(lb/1000 short tons)
of concentrate produced
1.930
(3.860)
-
m kg/ 1000 nwtrie toni
ll>/1 000 abort tom)
of ora millad
0.014
(0.028)
-
Fa
CONCEN-
TRATION
Imo/d)

-------
          Figure V-31. TYPICAL WATER-USE PATTERNS
I
     IN-SITU LEACH
MINE


TREATMENT
\


MILL
i
i
	 ^/ IMPOUN
                  (a) WET MINE/MILL COMPLEX
TREATMENT


MILL
                                              RECYCLE
                      (b) SEPARATED MILL
I
     IN-SITU LEACH
MINE


TREATMENT
                DISCHARGE




                     (c) SEPARATED WET MINE
                            316

-------
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  hazardous  radionuelites.   The  chemistry   of
radium  is similar to that of calcium, barium and strontium.
The Environmental Protection  Agency  has  proposed  interim
drinking  water standards for radium -226 and radium -228 at
5pci/l  (picocuries per liter) total for both radionuclides.

Radium, with a half-life of 1,620 years, is generated by the
radioactive decay of uranium, which has the very long  half-
life  of  4.51  billion  years.  In uranium ores that are in
place  for  billions  of  years,  an  equilibrium   may   be
established between the rate of decay of uranium into radium
and  the  rate  of decay of radium into its daughters.  Once
this equilibrium is established, the  ratio  of  uranium  to
radium   equals  the  ratio  of  the  half-lives—i.e.,  2.7
million.  An equilibrated ore with a typical grade  of  0.22
percent  uranium  would contain 0.82 microgram of radium per
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  ciruit,  about  50%  of  the
thorium and .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-210r  with
half  lives  of  222  years  and 139 days, respectively, are
considered along with radium-226.  Thorium  is  observed  in
tailings  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.
                           317

-------
Chemical   and   Physical   Waste  Constituents.    Chemical
contaminants of milling waste waters 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 U3OJ5 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
the  mill  waste,  and  about 1536 is in the uranium product.
With no parent 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.

Additional pollutants (particularly, metals) are expected to
appear in the waste streams of specific plants that might be
using  unusual  ores.    Certain   compounds,   particularly
organics,  are  expected  to  undergo  changes  and  are not
identifiable individually but would appear  in  waste-stream
analysis  under  class  headings  (e.g.,  as  TOC,  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.
                           318

-------
  TABLE V-58. WASTE CONSTITUENTS EXPECTED
  ACID LEACH PROCESS
ACID-LEACH CIRCUIT:
  Sulfuric acid
  Sodium chlorate
LIQUID/SOLID-SEPARATION CIRCUIT:
  Polyacrylamides
  Guar gums
  Animal glues
ION-EXCHANGE CIRCUIT:
  Strong base anionic resins
  Sodium chloride
  Sulfuric 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
  Sulfuric 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 28
                            319

-------
       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, COO, 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
At, 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
(None among the analyzed items)
•"Background" is defined in text.
                               320

-------
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 of 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
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  liguor  following
initial  contact  with ore and after reducing gases, such as
hydrogen and H2S, 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   (NaClO.3)  and  manganese dioxide (MnO^) serve this
purpose in quantities of 1 to  4  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
waste waters.

The waste water 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 recarbonation loop.  The wastes
discarded  to  tailings  often  contain  organic   compounds
                           321

-------
Figure V-32. ALKALINE-LEACH WATER FLOW
         GROUND ORE
             i
                                    TO ATMOSPHERE
                                         T
       I
          ALKALINE
            LEACH
FRESH WATER
OR
TAILING-SOLUTION
RECYCLE


1
FILTERING


                                  I
                               LEACH
                              RECYCLE
                                  I	
          PREGNANT
           LEACH
             i
        PRECIPITATION
           CRUDE
          PRODUCT
     USED
     STACK
     GAS


       I
       I
                               RECARBONATION
  t
               STACK
               GAS
BARREN
 LEACH
                            REPULPED
                            "TAILINGS"
FRESH
WATER


PURIFICATION
(SODIUM REMOVAL)
WASTE _
WATER ^
                                           k TO TAILING
                                            POND
            END
          PRODUCT
            TO
         STOCKPILE
              322

-------
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
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
                            323

-------
Figure V-33. AMMONIUM CARBONATE LEACHING PROCESS
       MINING
        ORE
                                  TO ATMOSPHERE
      GRINDING
 LEACH
SOLUTION"
                                    WASTE GAS
                                        1
               AMMONIA AND
               CARBON DIOXIDE
             ABSORPTION TOWERS
  PRESSURE LEACHING
  COUNTERCURRENT
    DECANTATION
	PREGNANT
  SOLUTION
      SLURRY
                          •FILTRATE-
        TO TAILING
          POND
                                              t
STEAM
                         STEAM STRIPPING
                          AND URANIUM
                          PRECIPITATION
                                                  SLURRY
                                                FILTRATION
                                                  PRODUCT
                                                    TO
                                                 STOCKPILE
                        324

-------
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
research in this field could lead to better characterization
and improved treatment of wastewater.

Process  Descriptionsr  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
                           325

-------
      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
Loss (Evaporation, etc.)

Makeup Water
Water in Circuit
Discharge
Evaporation and
Seepage

MINE/MILL 9401
m3/day
gpd
WATER USED
MINE/MILL 9402
m^/day
gpd
MINE/MILL 9403
m3/day | gpd
MINE/MILL 9404
m3/day
gpd
MINE PORTION
8,339
3,339
0
5,000
estO

2,700
3,200
0

2,700
2,203,000
882,100
0
1,321,000
est 0
11,552
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
N/A
N/A
N/A
N/A
N/A
MILL PORTION
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

660
1,601,000
1,738,000
1,427,000

174,400
est 1330
0
0
0
est 1,530
est 404,200
0
0
0
est 404,200

5,300
5,300
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/RaA/ OPERATIONS
FEATURE
PARAMETER
MINE/MILL
9401
9402
MINE PORTION
Settling Basin
Evaporating Pond
Ion-Exchange Plant
Area in hectares (acres) II 0.3 (0.74)
Retention Time in hours || est 20
Area in hectares (acres) || N/A
U3 OB Concentration in mg/l II 25
U3 OB Removal in % I 96
0.7(1.7)
est 80
N/A
2 to 12
98
MILL PORTION
Tailing Pond(j)
Ion-Exchange Plant
Rtcarbonizer
Deep Well
Utilization of
Area in hectares (acres)
Number series-connected
Daily Water Use in metric tons (short tons)
Daily Water Use in metric tons (short tons)
Capacity in metric tons (short tons) water per day
Sand/Slime Separators
Decant Facilities
Filters
Coprecipitation
21 (51.8)
1
490 (540)
1,635(1302)
0
Yes
Yes
100(247)
5
N/A
N/A
0
Yes


N/A
N/A
N/A
N/A
N/A

24 (59.3)
3
N/A
520(573)
0
Yes


N/A
N/A
2 (4.9)
N/A
N/A

107 (264)
1
N/A
N/A
1,635(1,802)
Yes
Yes
Yes
TOTAL OPERATION
Ore Handling
Capacity in metric tons (short tons) per day
3,200(3,527)
6.400(7,055)
1,400(1,543)
2,700(2,976)
N/A - Not available
                                326

-------
Figure V-34. WATER FLOW IN MILLS 9401, 9402, 9403, AND 9404


                          3,339 m3/dtv
I Ml
i
^*WE
Nc I i |kj ION EXCHANGE L?8?jll!0^K" y» DISCHARGE

1220,200 gpd)
1 IN-SITU LEACH 	




LUST}— — — - 	 1».
(713,300 gpd)
1,836 m3/d«y
1431.900 gpd)
6,000 m3/diy
(1.320.000 gpd)
	
Fflfl miMiy (1*7 0"" grit

MILL 	 |»/
3,200 m3/div
1 W4S.400 gpd)





S TAILING X
>» 	 POND ^^/
2,700 m3/d«y
(713.300 gpdl
I—*- LOSS
                             (a) MILL 9401
   2.136 m3/day
MINE1! I '"^ flp"'
9/»17 m3W
u,..ce | (2.500.000 gpd)
|

^^^ -/ 71V_Ju-, I

(584.000 Jpdi


7^25 m3/diy
(1,900,000 gpdl
1,920 m3/diy
(507^00 gpd)
5,307 m3/d.y
1 11.400.000 gpdl

3,800 m3/d»y
(951.000 gpd)
(578,500 gpdl



ESS USES AND LOSS

MILL
t


(1 ,400.000 gpdl 	 "
QE
4,326 m3/diy
"(1,140.000 gpd)
GE

•S i 5.307 m3/d«y 1
l< 11. 400.000 gpdl
                            (b) MILL 9402
                                                               TO i
                                                          ATMOSPHERE 1
                                                              EVAPORATION
FAM^. 	

r—^ 6,080 m3/diy
(1,600,000 gpdl
6,400 m3/d«y
(1, 430.000 gpdl


i
i

860 m3/d« ^^ 	
I (174,400 gpd)
' 520 mj/diy
(137,400 gpd) 1


3 POHO\
••***"^
DISCHARGE f* -- 	 	 	 COPRECIPITATION
                            (c) MILL 9403
t
OPEN-PIT
MINE
_JL_
EVAPORATIOI
POND
	
	 HA
1,630 m3/dly I 	
1404,200 gpd)
1,393 m3/dty "~
(388.000 gpdl
rx
^ 137 m3/diy
(36^00 gpdl
	 1 TO ATMOSPHERE
UNACCOUNTED LOSS








                           (d) MILL 9404
                          327

-------
                Figure V-35.  FLOWCHART OF MILL 9401
                                                             TO ATMOSPHERE
                                                                  L
TO ATMOSPHERE
       TO STOCKPILE
                            328

-------
               Figure V-36. FLOW CHART FOR MILL 9402
                       MINING
      5,300 m3/day
     (1,400,000 gpd)
                       ORE
CRUSHING
  AND
GRINDING
H2S04
NadO
 NaCI
                     LEACHING
                                   •SANDS
                    THICKENERS
                     SOLVENT
                   EXTRACTION
                  PRECIPITATION
                  YELLOW CAKE


                  TO STOCKPILE
                                 EVAPORATION
                                 AND SEEPAGE
                                                             5,300 m3/day
                                                             (1,400,000 gpd)
                                329

-------
      Figure V-37. FLOW CHART OF MILL 9403
                                                         WASH EQUIVALENT TO MOISTURE
                                                         RETAINED IN CAKE (ELIMINATED
                                                                 IN DRYER)
TO STOCKPILE
^-"44-HECTARE (108-ACREI"
f TAILING POND WITH 23-HECTARE
N>	IBS-ACRE) LIQUID POOL_
                       330

-------
Figure V-38. FLOW CHART OF MILL 9404
MILL WATER SUPPLY = 5,300 n
RAIN
1,530 m3/day
1 (404,200 gpd)
n3 (1,400,000 gal) per day
TO ATMOSPHERE
EVAPORATION
137 m3/day
(36,200 gpd)
C^^-^^x
320-hectare (790-acre) /^-hectare (4.9-acreWN
OPEN-PIT MINE I EVAPORATION ) )
^^KMV*S/
^^^


TOTAL LOSS OF
5,300 m3/day
It Ann nnn ~**A\ .
U,HUU,UUUgpa/ |\3
LH


	 [ TAILING A
V POND )
IT
FILTER
CAPACITY OF
1,635 m3/day
(431,900 gpd)
\ •
( DEEP WELL J



MINING
ORE
^^

GRINDING «•—
( WELL 2 J
^~. _-^
1
ACID LEACH
RIM
— ^ * "
EPULP U« 	 SANDS 	 1 HYDROCYCLONES
T
BARREN
SOLUTION
1 \AfPI 1 d \


1
SLIMES

RESIN-IN-PULP '
ION EXCHANGE -^
1
PREGNANT
SOLUTION
1ELLUE

SE
NT
(HIGH CD
PRECIPITATION
AND
FILTERING

3_j: 	
WAQHiwr;

-LOW c r-J
            331

-------
 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  subcategory  was not visited because its
 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
                        332

-------
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-
ment.  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.

Organics.   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  tht  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
                          333

-------
     TABLE V-62. RADIONUCLIDES IN RAW WASTEWATERS FROM
                URANIUM/RADIUM/VANADIUM MINES AND MILLS
RADIONUCLIDE
and units of measurement
RADIUM 226
in picocuries/l
THORIUM
in mg/£
URANIUM
in mg/a
CONCENTRATION
MINES
200 to 3,200
<0.1
4 to 25
ACID MILLS
200 to 700
(4,100)*
(1.D*
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 (TOC)
Oil and Grease
MBAS Surfactants
Phenol
CONCENTRATION (mg/°)
MINES
16 to 45
3to4
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
                         334

-------
total dissolved solids and are found in the levels shown  in
Table V-64.

Light  Metals.   The ions of sodium, potassium, and ammonium
found in waste waters 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 waste water 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
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,
a.- jn, 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   also   effect   significant
reductions  in other pollutants; for example, neutralization
depresses heavy metals, and aeration reduces organics.
                           335

-------
 TABLE V-64. INORGANIC ANIONS IN U/Ra/V RAW WASTEWATER
PARAMETER
Sulfide
Cyanide
Fluoride
Total Dissolved Solids (TOS)
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/RaA/
          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/X )
MINES
0.01 to 0.03
<0.02
0.5 to 1.2
2 to 25
0.5 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
                        336

-------
  TABLE V-67. CONCENTRATIONS OF HEAVY METALS FORMING
            CATIONIC SPECIES IN U/Ra/V RAW WASTEWATER
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
<02to40
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 (mg/£)
MINES
2
200 to 230
<10to750
200 to 10,000
ACID MILLS
4,000
0
30
300,000 to 500,000
ALKALINE MILLS
0
1,000 to 5,000
10
100,000 to 300,000
                         337

-------
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-71f 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 TOG 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,
was  concentrating its own, much cleaner, ore.  The ratio of
200:1 in IOC is, therefore, expected.

Metal Ores - 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
                          338

-------
TABLE V-69. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE LOAD
            FOR URANIUM MINES 9401 AND 9402

PARAMETER

TSS
COD
TOC
Alkalinity
Ca
Mg
Fe
Mo
V
Ra
Th
U
MINE 9401
CONCENTRATION
(mg/£)
IN WASTEWATER
—
242
15.8
224.4
93
45
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
29,700+
-
113
Ib/day
-
5,200
320
4,600
1,900
920
10
11
20
-
-
248
MINE 9402
CONCENTRATION
IN WASTEWATER
299
600
25
-
117
36
0.23
0.53
<0.5
2,7 10»
<0.1
11.6
RAW WASTE LOAD
kg/day
640
7,000
290
-
1,300
410
3
6
<6
31,100*
<1.2
134
Ib/day
1,400
15,000
640
-
3,000
910
6
13
03
—
<2.5
294
 •Value in picocuries/£
  Value in picocuries/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
Fe
Mn
Pb
At
Mo
V
Ra
U
Fluoride
CONCENTRATION
(ma/ )
IN WASTEWATER
294,000
55.6
450
12,200
<0.5
0.92
<0.2
<0.05
0.33
<0.3
17
19,000t
43.9
2.1
TOTAL WASTE
kg/day
3,200,000
150
1,215
32,940
<1.4
2.5
<0.54
<0.14
0.89
<0.81
46
51,300"
118
5.7
Ib/day
7,000,000
331
2.680
72,620
<3
5.5
< 1.2
< 0.3
2
< 1.8
101
261
13
RAW WASTE LOAD
per unit ora milled
kg/metric ton
1,000
0.047
0.38
10
< 0.00042
0.00078
< 0.00017
< 0.000042
0.00028
< 0.00025
0.014
16n
0.041
0.0018
Ib/ihort ton
2,000
0.094
0.76
2)
<0.00084
0.0016
<0.00034
< 0.000084
0.00056
< 0.00051
0.029
33*"
0.081
0.0035
par unit concentrate produced
kg/ metric ton
550,000
26
211
5.720
<0.23
0.43
< 0.094
< 0.023
0.15
<0.14
8,870tf
22
0.98
Ib/snort ton
1,100,000
52
422
11,440
<0.47
0.86
<0.19
< 0.047
0.31
< 0.28
16
20,100"*
45
2.0
tt
•On the basis of 1973 production of 94.5% Uj

^Value in picocur)es/£
 Value in microcurim/day
 Value in microcuries/metric ton
 Value in microcuries/ihort ton
                            and 5.5% V
                                     339

-------
      TABLE V-71. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
                  LOAD FOR MILL 9402 (ACID- OR COMBINED ACID/ALKALINE-
                  MILL SUBCATEGORY)

PARAMETER
=
TSS
COD
TOC
Acidity
Al
Cu
Mn
Pb
As
Cr
Mo
V
Ra
U
CONCENTRATION
(mg/£ )
IN WASTEWATER
=====5
525,000
63.5
24.0
35.000
1,594
2.7
105
2.1
2.3
9.0
16.0
125
234t
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

Ib/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
0.30n
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
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
0.27*** 136tf
0.080
18
Ib/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% U_O0 and 1 8% MO
  t                           o B         3
  Value in picocunes/?
 "Value in microcuries/day
  Value in microcuries/metric ton
***Value in microcuries/short ton
                                    340

-------
 TABLE V-72. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
              LOAD FOR MILL 9403 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COP
TOC
Alkalinity
C.
Mg
Ti
Al
Cu
F.
Mn
Ni
Pb
Zn
As
Mo
V
Ha
Th
U
Fluoride
CONCENTRATION
(mg/ei
IN WASTEWATER
111,000
27.8
<1
1,150.6
3.200
190
0.395
1 18
1.1
1.6
38
0.52
0.69
< 0.5
1.4
< 0.3
< 0.5
111f
<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
< H
13,190
36.680
2,180
4.5
206
13
18
436
6
7.9
< 5.7
16
< 3.4
< 5.7
-
<1
45
16
RAW WASTE LOAD
per unit on milled
kg/ metric ton
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
0.411t
< 0.0004
0.032
0.0052
Ib/short 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 concentrate produced*
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
5.5
< 1.2
2
431 "
<0.4
34
5.5
Ib/ihort 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
i -
< 2.3
4
392'"
<0.8
68
11
 •On the basil o< 197 3 production of 67% Uj
 'value in picocuries/..
 "Value in microcunes/day
 **Value in microcunes/metnc ton
•"Value in microcuries/short ton
                           . and 33% CuS.
                                     341

-------
   TABLE V-73. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE
                LOAD FOR MILL 9404 (ACID- OR COMBINED ACID/ALKALINE-
                MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Acidity
Ca
MO
Ti
Al
Cu
Ft
Mn
Ni
Pb
Zn
As
Cr
Mo
V
Ra
U
CONCENTRATION
lmg/2,1
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
kg/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
< 27
0.69
11
< 1.6
640
3,660"
925
Ib/day
6,000,000
7,360
72
47.200
2.620
5.430
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
per unit ore 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
1.35rt
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
O.OOO51
0.0079
< 0.0012
0.47
1.23***
0.75
per unit concentrate produced*
kg/ metric 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/lhort 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 the basis of 1973 production of 100% Ll-j
  Value in picocunes/£
 "Value in microcuries/day
  Value in microcuries/metric ton
•••Value in microcuries/short ton
                                    342

-------
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.

Water flows of the antimony, beryllium, platinum, rare-earth
titanium,  and  zirconium  mineral  operations  visited  are
presented in Figures V-39, V-UO, 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.

Waste water 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  waste  water  constituents from
these operations  are  high  suspended  solid  loadings  and
                           343

-------
Figure V-39. WATER FLOWS AND USAGE FOR MINE/MILLS 9901 (ANTIMONY) AND
           9902 (BERYLLIUM)
            (NO DISCHARGE)

' 306 TO 3(2 m3««y
(80.000 TO 100,000 gpdl
FLOTATION
MILL

286 TO 343 m3/diy
TAILING-
IMPOUNDMENT



EVAPORATION
AND
SEEPAGE

                                                                      (NO
                                                                      DISCHARGE)
                         (a) ANTIMONY MINE/MILL 9901
                                                  TO ATMOSPHERE
                         (b) BERYLLIUM MINE/MILL 9902
                               344

-------
Figure V-40. WATER FLOWS AND USAGE FOR MINE/MILLS 9903 {RARE EARTHS)
           AND 9904 (PLATINUM)
                                                           TO ATMOSPHERE
ARGE)
0.08 m'/mmuu
121 gpm)
0.36 m3/minute
(95 gpm)

LEACH CON
'
FLOTAT
i



'



1 .6 m3/mmute

1 0.08 m3/minut« 1
T (20 gpm) 7
EVAPORATION
^1 EVAPO
0.08 m3/minut. ~~ \. K
(20 gpm) ^*— —
, - ' TAI
136m3/minute
(518 gpm)
RECLAIM
TANK
RATION*
)ND )
-ING >
^V POND J





EVAPORATION


0.2 m /minute
                                 (412 gpml

               NOTE. FOR BYPRODUCT RECOVERY, SEE PART (b) OF FIGURE V-41 (MINE/MILL 9906)

                         (a) RARE-EARTH MINE/MILL 9903
RIV
I
'Eft J— |fc{ '
J 24,730 m3/d.v V.

24,730 m3/d«v
(6,480,000 gpdl
>REDGE
POND


./ 49,500 m3/d.y
DREDGE WITH
WET GRAVITY
BENEFICIATION

49300 m3/d«y
(12.960,000 gpd)

                        (b)  PLATINUM MINE/MILL 9904
                                    345

-------
Figure V-41. WATER FLOWS AND USAGE FOR TITANIUM MINE/MILLS 9905 AND 9906
             2,668 mj/day
             (699,000 gpd)
      FLOTATION AND
  MAGNETIC-SEPARATION
  	.MILL
              36,069 m3/day
             (9,450,000 gpd)
                      INTERMITTENT
                  DISCHARGE (SEASONAL)
(9,220,000 gpd)
                                         878 m°/day
                                         (230,000 gpd)
                             (a) TITANIUM MINE/MILL 9905
                                                                     TO ATMOSPHERE
                                      BULK
                                  CONCENTRATE"
          12,099    ay
          (3,170,000 gpd)
                           12,099 m3/day
                           (3,170,000 gpd)
  TAILING
   POND
  SYSTEM
                                   12,595 mj/day
                                  (3,300,000 gpd)
                  DRY
                  MILL
             (ELECTROSTATIC
                  AND
           MAGNETIC METHODS)
                                                        7,633 irr/day
                                                       (2,000,000 gpd)
                                                                      EVAPORATION
             17,175 m-Vday
             (4,500,000 gpd)
                                                                       DISCHARGE

                                                                           TO
                                                                          STREAM
                    (b)  TITANIUM/ZIRCONIUM/MONAZITE MINE/MILL 9906
                                      34C

-------
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.

Waste  water  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  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.  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.

Waste  water  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
                            347

-------
product.  This  mill   (9901)  employs  the  froth  flotation
process   to   concentrate  the  antimony  sulfide  mineral,
stibnite  (Figure 111-28).  The particular flotation reagents
used by this mill are  listed in Table V-74.  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.

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            (Sb2S_3)
         Pyrite              (FeS^)
         Arsenopyrite        (FeAsS)
         Sphalerite          (ZnS)
         Argentite           (Ag^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 waste waters 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  (H2BeU^Si209)  is recovered by open-pit
methods.   A small amount of beryl is also mined in the U.S.
                          348

-------
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


0.1
0.5

Ib/short ton
ore milled


0.2
1.0

                         349

-------
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
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
Total Mn
I
CONCENTRATION (mg/£)
8.3"
8.5
11.0
113*
170
149
68
40
1.5
43
73
6.2
0.23
< 0.002
<0.3
<0.01
0.103
057
0.04
0.12
183
0.13
1.93
0.40
PARAMETER
Hg
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Sb
Mo
Oil and Great*
MBAS Surfactants
Cyanide
Phenol
Fluoride
Total Kjeldahl N
Sulfide
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/£)
0.0038
0.10
<0.05
<0.2
3.5
0.036
<0.02
2.0
0.11
<0.2

-------
TABLE V-76. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD
           AT ANTIMONY MILL 9901
PARAMETER
sass^^sss^^^^ssss^^^ss
PH
TSS
COD
TOC
Fe
Pb
Sb
Zn
Cu
Mn
Mo
CONCENTRATION
(mg/£) IN
WASTEWATER
5==
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
                         351

-------
 by crude open-cut and hand-picking methods.  As indicated in
 Figure V-39f 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.

 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, K,  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) PO^t) 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).
                        352

-------
 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
COO
TOC
Oil and Grease
MBAS Surfactant*
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
CONCENTRATION (mg/£)
17,000"
88*
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

-------
 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.

 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 REOr 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-43) .

 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  fe "
discussed later.
,_	c—^^-^ wi~J_^./ ,   J.HV- j.aj-j-j.uauc J..LUIII tue extraction circ
makes up the  feed  for  the  lanthanum  circuit,  which  is
After  loading  the  organic  with europium, the europium is
stripped in the solvent extraction  strip  circuit  with  4N
hydrochloric  acid.   The  pregnant  strip solution contains
                          354

-------
Figure V-42. BENEFICIATION OF  BERTRANDITE, MINED FROM A  LODE DEPOSIT.
              BY FLOTATION (MILL 9903)
                  0.36 in3/m*i  1M nT/mimm
                   (W gpml     <61« «pm>
        -UNDERFLOW
    CYCLONE
                         FBOTH	  FIRST AND SECOND ROUGHER
                       - FROTH _|    FLOTAT,ON CELLS
                                      UNDERFLOW
                          - FROTH
                   FIRST CLEANER
                  FLOTATION CELLS
                     FROTH
              SECOND. THIRD, AND FOURTH
                    CLEANER
                 FLOTATION CELLS
                   UNDERFLOW
   -UNDERFLOW
SCAVENGER
FLOTATION CELLS
ROTH— I
I

1.96 m'/mmutt
  (518 gpm)
   LEACHED CONCENTRATE
       THICKENER
.TO
 WASTE
                                                                                 ALTERNATIVE
                                                                        SEPARATION OF RARE-
                                                                          EARTH METALS BY
                                                                        SOLVENT EXTRACTION
                                                                          (SEE FIGURE V-43)
                                         TO STOCKPILE
                                     355

-------
                   Figure V-43. BENEFICIATION OF RARE-EARTH FLOTATION CONCENTRATE BY
                             SOLVENT EXTRACTION (MILL 9903)
Ul
TO WASTE
1 <
\ t
FILTRATE OVERFLOW
1

FJLTER •*— THICKENER
L— t
DRYER mtCf"J£>0>> -*— AMMONIA
COARSE OR T
FINE
HYDRATE FILTER EUROPIUM
' J PRODUCT


1 "
1 Clllf
THICKENER DRUM _____
FILTER
FILTRATE
1
DRUM
FILTER

FLOTATION
CONCENTRATE
(FROM
FIGURE V-421







Y
DRYER
GADOLIMIUM/SAMARIUM

CARBONATE
SODIUM CARBONATE
1 >
GADOLIMIUIV
PRECIPITA
EURO
I/SAMARIUM
MON TANK
>IUM





PURIFICATION
1
REL
»0
IDE




SODIUM
CARBONATE

' ' ' 1 ' SOLVENT EXTRACTION
i STRIPPING UNIT
3 STOCKPILES , i
SODIUM HYOROSULFIDE
AND AMMONIA
f
PRECIPITATION
TANKS "• 	 RAFFINATE

LANTHANUM CIRCUIT 1

t 1
LOADED BARREN
EUROPIUM ORGANIC
ORGANIC
r
SOLVENT EXTRACTION
UNITS




1ESIOL
PRODUCT







ORGANIC
STORAGE
TANK

£
<•



THICKENER
1
1
OVERFLOW
T
CARBON
FILTERS

1
PREPARATION
TANKS





ORGANIC

HYDROCHLORIC
ACID

EUROPIUM CIRCUIT 1


FEED _.
TANKS
IRON FREE
PREGNANT
EUROPIUM STRIP
SOLUTION

H HEARTH
ROASTER


1 '
| COOLER

uuo m /mm 121 gpml HYDROCHLORIC
' WAI EH ACID SOLUTION
t
LEACH
TANKS

fc THICKENER
"" 2
OVERFLOW





fc CARBON
FILTER
AND — J " " t'~
HEPULPEH Jf
'

CERIUM 1 ' 	 1 '
CIRCUIT 1
1 ' 1 '
j-T* 	 ,[1
*~ SOLVENT
EXTRACTION

I
LOADED
ORGANIC
SOLVENT EXTRACTION
STRIPPING UNITS
PREGNANT EUROPIUM
STRIP SOLUTION
IRON PRECIPITATION
TANKS
Y
PRESSURE LEAF
FILTER
AGHOUSE 4-, 1 Y '
1 I FILTER ) /~ CERIUM\
1 1 . r .J ( STnoARF 1
CYCLONE CERIUM \-"OIMOS/
CIRCUIT 	 •. 	
" T i

COARSE OR )
1 	 ^l 	 FINE CERIUM fc 1
i HYDRATE 1 TO
1 	 * 	 1 BLENDED r STOCKPILES
BLENDER | CERIUM fc

SODIUM
"*~ CARBONATE
FERRIC HYnROXinF ^ TO
CAKE WASTE

-------
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
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-UO,  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            Methylisobutylcarblnol
Collector               N-80 Oleic Acid
pH Modifier        Sodium Carbonate
Depressants        Orzan, Sodium  Silicofluroide
Conditioning Agent   Molybdenum Compound

Leach Circuit

Leaching Agent          Hydrochloric Acid
                          357

-------
        TABLE V-78. CHEMICAL COMPOSITION OF RAW WASTEWATER
                   FROM RARE-EARTH MILL 9903
PARAMETER
pH
Acidity
Alkalinity
Color
Turbidity (JTU)
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
SiO2
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/£)
FLOTATION
9.02*
.
;
.
14,476
360,000
-
-
3,100
-
-
-
-
-
-
-
-
-
0.35
-
-
LEACH/
SOLVENT
EXTRACTION
8.23*
345
2,125
80f
5.2
76,162
786
7,220
>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
A8
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfate
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
                                358

-------
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
                   359

-------
Solvent- Extraction Circuit

Leaching Agent     Hydrochloric Acid
Precipitants       Sodium Carbonate, Ammonia, Sodium Hydrosulfide
Solvents           Kerosene, Ethyl/Hexyl Phosphoric Acid

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 (CeFCCH, with La, Nd, Pr, Sm, Gd, and Eu) ;
barite  (BaSO^) ; calcite (CaC03J ; and strontianite
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/i) .  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  beneficiate  the   placer
gravels   (Figure  111-20) .  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
                           360

-------
TABLE V-80. CHEMICAL COMPOSITION AND RAW WASTE LOAD FROM RARE-EARTH
            MILL 9903
PARAMETER

PH
TSS
TOC
Cr
Mn
V
Fluoride

PH
TSS
TOC
SiO2
Cr
Mn
V
Te
Ni
CONCENTRATION
(mg/£) IN
WASTEWATER
RAW WASTE LOAD f
per unit of concentrate
kg/metric ton ] Ib/short ton
per unit ore milled
kg/metric ton | Ib/short ton
(a) Flotation Mill
9.02*
360,000
3,100
0.35
0.5
<0.3
365
_
9,335
80.4
0.009
0.013
< 0.0078
9.46
_
18.670
160.8
0.018
0.026
< 0.01 6
18.93
_
933.5
8.04
0.0009
0.0013
< 0.0008
0.95
^
1,867.0
16.08
0.0018
0.0026
<0.0016
1.89
(b) Leach/Solvent-Exchange Mill
8.23*
786
47
1.25
0.04
3.0
<0.3
3.36
0.85
_
0.833
0.047
0.00125
0.00004
0.003
<0.0003
0.003
0.001
_
1.67
0.094
0.00250
0.00008
0.006
< 0.0006
0.006
0.002
	
_
_
—
_
_
_
	
-
__
— _
_ ,
—
_
^
t 	
,__„.
-
      Value in pH units

      Based upon maximum production achievable (part a) or estimated amount of flotation concentrate
      produced (part b)
                                  361

-------
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 waste water and
waste loads from mine/mill 9904.

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 (FeCr204)
    Ferroplatinum (Fe, Pt, Ir, Os, Ru, Rhr Pd, Cu, Ni)  alloy
    Iridium/ruthenium/osmium alloy
    Taurite (Ru, Ir, Os)S2
    Unnamed mineral  (Ir, Rh, Pd)S
    Mertieite (Pt5(Sb, As) 2)
    Sperrylite (PtAsj?)
    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
                            362

-------
TABLE V-81. CHEMICAL COMPOSITION AND LOADING FOR PRINCIPAL WASTE
          CONSTITUENTS RESULTING FROM PLATINUM MINE/MILL 9904
          {INDUSTRY DATA)
PARAMETER
Alkalinity
Conductivity
Hardness
COD
BOO
TS
TDS
TSS
(N) NH3
Kjeldahl Nitrogen
Al
Cd
Cr
Cu
Total Fe
Pb
Zn
Chloride
Fluoride
Nitrate
Sulfate
Sulfide
CONCENTRATION
(mg/ £ )
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
                          363

-------
     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/£)
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
Sa
Ta
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
1 	 — 	 1
CONCENTRATION (mg/K)
<0.05
26.0
<0.01
<0.01
<0.1
<05
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
                           364

-------
Figure V-44. BENEFICIATION AND WASTE WATER FLOW OF ILMENITE
           MINE/MILL 9905 (ROCK DEPOSIT)
                                            MINING
                                               I
                                              ORE
                                           CRUSHING
              878 mj/day
             (232,000 gpd)
36,069 nT/day
(9,450,000 gpd)
                                            GRINDING
                                         CLASSIFICATION
                                              I
                                           MAGNETIC
                                          SEPARATION
                                   MAGNETICS	«-NONMAGNETICS-i
                                                           ILMENITE
                                                         AND GANGUE
                                 FILTRATE
                              WATER
                         "RETURN TO MILL

                          35,191 m3/day
                          (9,220,000 gpd)
                                                         FLOTATION
                                                           CIRCUIT
                                                          THICKENER
                                                            I
                                                           FILTER
                                1
                               DRIER
                                                            I
                                                        CONCENTRATE

                                                             f
                                                         TO SHIPPING
                                 365

-------
   TABLE V-83. CHEMICAL COMPOSITION OF RAW WASTEWATER
              FROM TITANIUM MILL 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardness
COO
TOC
Oil and Grease
MBAS Surfactants
Total Kjeldahl N
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/H)
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
0.58
0.43
500
PARAMETER
Pb
Mg
Total Mn
Ni
T(
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Phenol
Chloride
Fluoride
Sulfate
Nitrate
Phosphate
CONCENTRATION (mg/£)
< 0.06
187.5
5.9
1.19
<0.1
2.0
23.7
0.132
0.015
41
0.29
< 0.06
2.08
7.6
< 0.1
< 0.1
< 0.01
19.1
32.5
213
0.68
< 0.05
•Value in micromhos/cm

tValue 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
SSSSSSSSSSSSSS^S^SSS^S^Sl
2.66
1.80
0.016
1.52
3.55
                            366

-------
         TABLE V-85. PRINCIPAL MINERALS ASSOCIATED WITH
                   ORE OF MINE 9905
MINERAL
llmenite
Magnetite
Pyroxene
Feldspar
COMPOSITION
Fe Ti 03
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/2,) 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
per 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
Ib/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
0.26
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
                         367

-------
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
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-UU) .

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 waste water 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  waste  water from mill  9905.
                            368

-------
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   waste  water
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.
                         369

-------
Figure V-45. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS (RUTILE, ILMENITE,
           ZIRCON, AND MONAZITE) AT MILL 9906
                                                                      TO
                                                                 ATMOSPHERE
    ORE + WATER
        I
20,100 m3/day
(5,310,000 gpd)
                 ORE FED
               FROM DREDGE
                                                                     i
                   EVAPORATION
       POND
      WATER
     RECYCLE
                   H2O
                  J_
                VIBRATING
                 SCREENS
             WASH
            WATER*

           11,000 m3/day
           (3,170,000 gpd)
                                     I
              7,570 m3/day
             (2,000,000 gpd)
—OVERSIZE-
            SPIRALS OR LAMINAR
      FLOWS (ROUGHERS AND CLEANERS)
                   I
                                 TO DRY MILL
                                (FIGURE III-30)
       WELL
        t—TAILINGS-i
       12,000 m3/day
       (3,170,000 gpd)
                                          WASTE (DREDGE)
                                               POND
                                                       15,615 m3/day
                                                      (4,500,000 gpd)  ,
                                                              DISCHARGE
                                    370

-------
      TABLE V-87. CHEMICAL COMPOSITION OF RAW WASTEWATER
                 AT MILLS 9906 AND 9907
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
COO
TOC
Total Kieldahl 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
0.6S
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
< 0.002
<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 ((07
033
<0.05
0.66
0.01
<0.01
<0.1
<05
1.3
<0.05
' < 0.01
5.0
<0.05
0.15
0.40
0.002
<0.1
<0.1
15.0
<0.01
040
<0.01
•Value in micromhos/cm

 Value in cobalt units
                            371

-------
TABLE V-88. RAW WASTE LOADS FOR PRINCIPAL WASTEWATER CONSTITUENTS
          FROM SAND PLACER MILLS 9906 AND 9907

PARAMETER
TSS
TOC
COO
Oil and Grease
Ti
Fe
Mn
Cr
Phosphate
MILL 9906
CONCENTRATION

IN WASTE WATER
11.000
972
1,337
400
<0.2
4.9
0.36
0.03
0.35
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
330
29.2
40.13
12
< 0.006
0.15
0.0011
0.0009
0.011
Ib/short ton
660
58.4
80.26
24
< 0.012
0.30
0.0022
0.0018
0.022

CONCENTRATION
(mfl/2)
IN WASTEWATER
209
321
361.6
40
0.4
0.93
<0.01
<0.01
0.4
MILL 9907
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
5.01
7.71
8.68
0.96
0.01
0.022
< 0.0024
< 0.0024
0.01
Ib/short ton
10.02
15.42
17.36
1.92
0.02
0.044
< 0.0048
< 0.0048
0.02
                           372

-------
                          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 subcategory represented in this study.  The waste  water
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  waste  water from each subcategory of this study.
The waste water 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-SEL ECTION 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.
                           373

-------
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 ma-
result  in  eye  irritation  for  the  swimmer.  Appreciab"
irritation will cause severe pain.
                           374

-------
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 man's 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
or lake bed and thereby destroying  the  living  spaces  for
                         375

-------
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
gravityseparation 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  components   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
lubricants and spillage of fuels,  as well  as  the  usage  of
reagents in many milling processes.
                          376

-------
 Chemical Oxygen Demand (COD)  and Total Organic Carbon

 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.
 Pish 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
 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
                          377

-------
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.

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.
                           378

-------
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  (NC>2) , 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 (NO_3-
N) not be used for infants.  Nitrates are  also  harmful  in
fermentation  processes and can cause disagreeable tastes in
beer.  In most natural water, the  pH  range  is  such  that
ammonium  ions  (NH4+)   predominate.   In  alkaline  waters,
however,  high  concentrations  of  un-ionized  ammonia   in
undissociated  ammonium  hydroxide  increase the toxicity of
ammonia solutions.   In streams polluted with sewage,   up  to
one half of the nitrogen in the sewage may be in the form of
free  ammonia,  and  sewage may carry up to 35 mg/1 of total
nitrogen.  It has been shown that, at a level of 1.0  mg/1 of
un-ionized ammonia,  the ability  of  hemoglobin  to  combine
with  oxygen  is impaired, and fish may suffocate.   Evidence
indicates that ammonia exerts a considerable toxic effect on
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
                           379

-------
increase  will  speed  up  the  plant  growth  and the decay
process.  In leaching operations, ammonia  may  be  used  in
leaching   solutions   (as   in  the  "Dean-Leute1  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
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 waste waters from the bauxite-ore mining industry.

Antimony

Antimony  is  rarely  found pure in nature, its common forms
being  the  sulfide,  stibnite  (Sb2S_3)    and   the   oxides
cervantite   (Sb204J   and  valentinite (Sb20^) .  Any antimony
discharged to  natural  waters  has  a  strong  tendency  to
                          380

-------
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.

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  (FeS/N FeAs2J .

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 (AsH_3) .

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  As^0_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
                          381

-------
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  U
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  waste  waters  was detected only in raw-waste
effluents from the mining and milling of bertrandite.
                           382

-------
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.

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
                           383

-------
 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
 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.
                          384

-------
 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,  phytoplankton,  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
 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
 fluoridebearing 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
                          385

-------
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
fluorides are toxic to fish at  concentrations  higher  than
1.5 mg/1.

High  fluoride levels in the effluents from mines may result
from high levels  in  intercepted  aguifers  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 U.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
guantities  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
                          386

-------
experienced   with   populations   of   the  iron  bacterium
Crenothrix.

Iron is very common in natural waters and  is  derived  from
common  iron  minerals in the substrata.  The iron may occur
in two forms:  suspended and dissolved.  The iron mining and
processing industry inherently increases iron levels present
in process or mine waters.  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  in  rocks.   Certain lead salts, such as the chloride
and the acetate, are  highly  soluble;  however,  since  the
carbonate  and  hydroxide  are  insoluble and the sulfide is
only slightly soluble, lead  is  not  likely  to  remain  in
solution   long  in  natural  waters.   In  the  U.S.,  lead
concentrations  in  surface  and  ground  waters  used   for
domestic supplies average 0.01 mg/1.  Some natural waters in
proximity  to  mountain limestone and galena contain as much
as 0.4 to 0.8 mg/1 of lead in solution.

Lead is highly toxic to human beings  and  is  a  cumulative
poison.   Typical  symptoms  of  advanced lead poisoning are
constipation, loss of appetite, anemia, abdominal pain,  and
gradual  paralysis  in  the muscles.  Lead poisoning usually
results from the cumulative  toxic  effects  of  lead  after
continuous ingestion over a long period of time, rather than
from  occasional small doses.  The level at which the amount
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
lifetime.

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 mg/1.
Toxic effects of lead on fish include  the  formation  of  a
coagulated  mucus  film over the gills—and,  eventually, the
entire body—which  causes  the  fish  to  suffocate.   Lead
toxicity  if  very  dependent on water hardness; in general,
lead is much less toxic in hard water.  Some  data  indicate
that  the median period of survival of rainbow trout in soft
water containing dissolved lead is 18 to  24  hours  at  1.6
mg/1.   The  96-hour minimum toxic level for fathead minnows
to lead has 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
                          387

-------
 from 0.1 to 75 mg/1 of lead, depending  on  water  hardness,
 dissolved  oxygen  concentration,   and  the type of organism
 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 a 8-hour minimum toxic level
 for sticklebacks in water containing 1,000 to 3,000 mg/1  of
 dissolved  solids  is reported to  be 0.34 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.

 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 40 mg/1.   The
 threshold toxic  concentration of manganese  for the  flatworm
 Poly_celis  nigra  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.
                          388

-------
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
factor  of  5,000  from water to pike has been reported, and
                           389

-------
 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
 waste  water   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.  EA 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.
                          390

-------
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 U3 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 liter, and
marine plants contain up to 3,000 micrograms per liter.  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   (Macrocystis  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 waste water 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
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
                           391

-------
 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
 contrxbuted 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
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
                           392

-------
species,  agef  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
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
                           393

-------
 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
 -I." v c J_ S •

 Naturally  occurring  radionuclides,  particularly  of   the
 uranium-2-38    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 radiotoxicity
 ot 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 radionuclides  in  the
 uranium and thorium series may be important, particularly if
released into water.   These include radium-228,  uranium, and
 iead-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
                          394

-------
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  earth's  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
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
                            395

-------
  TABLE Vl-l. KNOWN TOXICITY OF SOME COMMON FLOTATION REAGENTS
                USED IN ORE MINING AND MILLING INDUSTRY
TRADE NAME
Airofloat 25
Aerof loat 31
Aerof loat 238
Aerofloat 242
Aarofroth 65
Aarof roth 71
Aero Promoter
404
Aero Promoter
3477
AROSURF
MG-98A
Oowf roth 250
DowZ-6
DowZ-11
Dow Z 200
Jaguar
M.I.B.C.
-
Superfloc 16
CHEMICAL COMPOSITION
Esuntially aryl dithtophosphoric acid
Essentially aryl dithiophosphoric acid
Sodium di-iecondary butyl
dithiophosphate
Esuntially aryl dithiophosphork acid
Polyglycol typa compound
Mixture of 6-9 carbon alcohols
Mixture of sulfhydryl type compounds
Unknown
Unknown
Chromium salts (ammonium, potassium,
and sodium chromate and ammonium,
potassium, and sodium dichromate)
Copper sulfate
Crasylic acid
Polypropylene glycol methyl ethers
Potassium amyl xanthate
Sodium isopropyl xanthate
Isopropyl ethylthionocarbamate
Based on guar gum
Lime (calcium oxide)
Mathylisobutylcarbinol
Pine oil
Potassium ferr'Ryanide
Sodium ferrocyanide
Sodium hydroxide
Sodium oleate
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
Flocculant
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 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
1 to 1900
1 to 1000
1 to 1000
100 to 1000
1 to 100
1 to 100
>1000
TOXICITY*
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
 O.Omo/Jl
1.0to 1000mg/£
 > 1000 mgli
NOTE:  Toxic range is a function of organism tested and water quality, including hardness
       and pH. Therefore, toxicity data presented in this table are only generally indica-
       tive of reagent 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.
                                       396

-------
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  waste  water  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
waste  water  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
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  waste  waters from mills 1107, 1108, 1109, and 1110
                          397

-------
 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
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.
                          398

-------
 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 waste  waters  from
 mines and  mills  are  not sufficiently high  for  selenium to be
 considered as a  harmful pollutant.

 Silicates

 Silicates   may  be   present  in  the waste waters  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.

 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
                          399

-------
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.
                           400

-------
TABLE Vl-2. SUMMARY OF PARAMETERS SELECTED FOR EFFLUENT
          LIMITATION BY METAL CATEGORY

PARAMETERS
pH (Acidity/Alkalinity)
Total Suspended Solids (TSS)

Chemical Oxygen Demand (COD)
Cyanide
Ammonia
Aluminum
Antimony
Arsenic

Cadmium
Chromium
Copper
Iron
Lead
Mercury
Molybdenum
Nickel
Vanadium
Ztnr
Radium
Uranium
PARAMETERS SELECTED FOR EFFLUENT LIMITATIONS
0»
6
c
0
•
•











+







Copper Ores
•
•


•





•

•








8
6
u
c
N
•o
c
(Q
•o
s




•





•

•

•
•





I/I
0>
6
5
o
O
•



•





+

*

•
t





Silver Ores




•





4

*

0
•





Bauxite (Al) Ores






•






^






£
0
>
_o
1
5
u.
•
•

•

i


•

•

•

9

•




Mercury Ores
•
•













•

•




Uranium, Radium,
and Vanadium Ores



f




•

•





+

^
4^
•
Metal Ores, Not Elsewhere Classified
1 Antimony
Ores







•
•




•




A
v

1
Is
£6

No separate limitations; zero discharge

1 Platinum
1 Ores





















i/i
c £
KO
bdenum
o
E
c.
\
u
3
?
a
»
.0
W)
re
>
c
o
TS
0}
O
tt
c
(fl
o
CD
•- £
ss
CO —
sy
22;
1 Titanium
Ores
:

t









i



•
^
V

1 Rare-Earth
Ores
No separate limitations; zero discharge

» Zirconium
Ores
E
3
E
re
£
1
' 0
1
a
>.
a
i/i
re
>
c
O
•o
£
«
>
o
"i
E
3
C
0
u
1 N
s
0
13
i
V
E
&
$
0
                         401
                                 ft U. S. GOVERNMENT PRINTING OFFICE : 1975 O - 596-127

-------
Pate I*»«e

-------