DEVELOPMENT DOCUMENT FOR
   EFFLUENT LIMITATIONS GUIDELINES
   AND STANDARDS OF PERFORMANCE
              FOR THE
ORE MINING AND DRESSING INDUSTRY
       POINT SOURCE CATEGORY
             £  £* \
             | ^9&7 3
             V
      UNITED STATES ENVIRONMENTAL PROTECTION AGENCY

               APRH 1975

-------
                                   DRAFT
                                  NOTICE
The  attached document is a DRAFT CONTRACTOR'S REPORT.  It includes tech-
nical information and recommendations submitted by the Contractor to the
United States Environmental Protection Agency ("EPA") regarding the sub-
ject industry.  It is being distributed for review and comment only.  The
report is not an official EPA publication, and it has not been reviewed
by the Agency.

The report, including the recommendations, will be undergoing extensive
review by EPA, Federal and State agencies, public interest organizations,
and other interested groups and persons during the coming weeks.  The
report—and, in particular, the contractor's recommended effluent limita-
tion guidelines and standards of performance—is subject to change in any
and all respects.

The regulations to be published by EPA under Section 304 (b) and 306 of
the Federal Water Pollution Control Act, as amended, will be based to a
large extent on the report and the comments received on it.  However,
pursuant to Sections 304 (b) and 306 of the Act, EPA will also consider
additional pertinent technical and economic information which is developed
in the course of review of this report by the public and within EPA.  EPA
is currently performing an economic impact analysis regarding the subject
industry, which will be taken into account as part of the review of the
report.  Upon completion of the review process, and prior to final pro-
mulgation of regulations, an EPA report will be issued setting forth EPA's
conclusions concerning the subject industry, effluent limitation guide-
lines, and standards of performance applicable to such industry.  Judg-
ments necessary to promulgation of regulations under Sections 304 (b) and
306 of the Act, of course, remain the responsibility of EPA.  Subject to
these limitations, EPA is making this draft contractor's report available
in order to encourage the widest possible participation of interested
persons in the decision making process at the earliest possible time.

The report shall have standing in any EPA proceeding or court proceeding
only to the extent that it represents the views of the Contractor who
studied the subject industry and prepared the information and recommenda-
tions.  It cannot be cited, referenced, or represented in any respect in
any such proceedings as a statement of EPA's views regarding the subject
industry.

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

-------
                              ERRATA
Draft Development Document for Effluent Limitations Guidelines ana
Standards of Performance for the Ore Mining and Dressing Industry
Point Source Category dated April 1975.

Pg IX-12       Line 18, 1st word:  change evaporation to precipitation.

Pg IX-13       Change 30-day average Hg waste load per unit ore milled
               from 0.00035 kg/1000 metric tons to 0.0035 kg/1000
               metric tons and from 0.0007 lb/1000 short tons to 0.007 lb/1000
               short tons.

Pg IX-13       Change 24-hour maximum Ilg waste load  per unit ore milled
               from 0.0007 kg/1000 metric tons to 0.007 kg/1000 kg/1000 metric
               tons and from 0.0015 lb/1000 short tons to 0.015 lb/1000
               short tons.

Pg IX-21       Change 30-day average Zn concentration from 0.15 to 0.1.

Pg IX-21       Change 24-hour maximum  Zn concentration from 0.25 to 0.2.

Pg IX-23       Change 30-day average Zn concentration from 0.15 to 0.1.

Pg IX-23       Change 24-hour maximum  Zn concentration from 0.25 to 0.2.

Pg IX-23       Change 30-day average Zn waste load per unit ore milled
               from 0.87  kg/1000 metric tons to 0.58 kg/1000 metric tons
               and from 1.45 lb/1000 short tons to 1.16 lb/1000 short  tons.

Pg IX-23       Change 24-hour maximum  Zn waste load per unit ore milled
               from 1.75  kg/1000 metric tons to 1.16 kg/1000 metric
               tons and from 2.90  lb/1000 short tons to 2.32 lb/1000
               short tons.

Pg IX-25       Change 30-day average Zn concentration from 0.15 to 0.1.

Pg IX-25       Change 24-hour maximum  Zn concentration  from 0.25  to 0.2.

Pg IX-25       Change  30-day average Zn waste load per  unit ore milled
               from 0.045 kg/1000  metric tons to  0.03 kg/1000 metric  tons
               and from 0.09 lb/1000 short tons to 0.06 lb/1000 short tons.

Pg IX-25       Change  24-hour maximum  Zn waste  load per unit  ore  milled
               from 0.074 kg/1000  metric tons  to  0.06 kg/1000 metric  tons
               and from 0.15 lb/1000 short tons to 0,12 lb/1000 short tons.

Pg IX-29       Change  30-day average Zn  concentration from 0.15  to 0.1.

-------
Pg IX-29       Change 24-hour maximum Zn concentration from 0.25 to 0.2.

Pg IX-32       Change 30-day average Zn concentration from 0.15 to 0.1.

Pg IX-32       Change 24-hour maximum Zn concentration from 0.25 to 0.2.

Pg IX-32       Change 30-day average Zn waste load per unit ore milled
               from 0.87 kg/1000 metric tons to 0.58 kg/1000 metric  tons
               and from 1.45 lb/1000 short tons to 1.16 lb/1000 short tons.

Pg IX-32       Change 24-hour maximum Zn waste load per unit ore milled
               from 1.45 kg/1000 metric tons to 1.16 kg/1000 metric tons
               and from 2.90 lb/1000 short tons to 2.32 lb/1000 short tons.

Pg  I X-38      Change 24-hour maximum As concentration from 0.7 to 0.8.

Pg IX-38       Change 24-hour maximum Cd concentration from 0.07 to 0.1.

Pg IX-38       Change 30-day average Zn concentration from 0.15 to 0.1.

Pg IX-47       Change 30-day average As concentration from 0.5 to 0.4.

Pg IX-47       Change 24-hour maximum As concentration from 0.7 to 0.8.

Pg IX-47       Change 30-day average As waste load per unit ore milled
               from 2.3 kg/1000 metric tons  to 1.8 kg/1000 metric tons
               and from 4.6 lb/1000 short tons to 3.7 lb/1000 short tons.

Pg IX-47       Change 24-hour maximum As waste load per unit ore milled
               from 3.22 kg/1000 metric tons to 3.7 kg/1000 metric tons
               and from 6.44 lb/1000 short tons to 7.4 lb/1000 short tons.

Pg  X-13       Identification of BATEA for Gold Mines  (alone) should
               read:  The best available technology economically
               achievable is chemical  (lime  or sulfide) precipitation
               with settling ponds.   (Same as BPCTCA.)

Pg  X-27       Change 30-day average As concentration from 0.5 to 0.4.

Pg  X-27       Change 24-hour maximum As concentration from 0.7 to 0.8.

Pg X-27        Change 30-day average As waste load per unit ore milled
               from 2.3 kg/1000 metric tons  to 1.8 kg/1000 metric tons
               and from 4.5  lb/1000 short tons to 3.7 lb/1000 short  tons.

Pg X-27        Change 24-hour maximum As waste load per unit ore milled
               from 3.2 kg/1000 metric tons  to 3.7 kg/1000 metric tons
               and from 6.4  lb/1000 short tons to 7.4  lb/1000 short  tons.

-------
Pg X-32        Change 30-day average TSS concentration from 25 to 20.

Pg X-32       Change 30-day average TSS waste load per unit ore milled
               from 0.36 kg/1000 metric tons to 0.3 kg/1000 metric tons
               and from 0.72 lb/1000 short tons to 0.6 lb/1000 short tons.

Pg XI-5        Change 30-day average Cd concentration from 0.5 to 0.05.

-------
           DRA^T
      DEVELOPMENT DOCUMENT

              for

EFFLUENT LIMITATIONS GUIDELINES

              and

    STANDARDS OF PERFORMANCE
ORE MINING AND DRESSING INDUSTRY
          Prepared By:

      Calspan Corporation
         P. 0. Box 235
    Buffalo, New York  14221

    Contract No. 68-01-2682

           April
           DRAFT

-------
                              DRAFT
                            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 appJl-
cation of the best practicable control technology currently
available (BPCTCA) and the degree of effluent reduction attain-
able 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 tech-
nology, processes, operating methods, or other alternatives.

Based upon the application of the best practicable control
technology currently available, 16 of the 43 subcategories for
which separate limitations are proposed can be operated with  no
discharge of process wastewater.  With the best available
technology economically achievable, 22 of the 43 subcategories
for which separate limitations are proposed can be operated
with no discharge of process wastewater to navigable waters.
No discharge of process wastewater pollutants is also achiev-
able as a new source performance standard for 22 of the 43
subcategories.

Supporting data and rationale for development oC the proposed
effluent limitation guidelines and standards of performance
are contained in this report.
                               I I I
                              DRAFT

-------
                                DRAFT
                               CONTENTS

Section                                                          Page

I         CONCLUSIONS                                            L-l

II        RECOMMENDATIONS                                        I I-I

III       INTRODUCTION                                           Ill-l

          PURPOSE AND AUTHORITY                                  JIJ-I

          SUMMARY OF METHODS USED FOR DEVELOPMENT OF
          EFFLUENT LIMITATION GUIDELINES AND STANDARDS
          OF TECHNOLOGY                                          III-3

          SUMMARY OF ORE-BENEFICIATION PROCESSES                 III-7

          GENERAL DESCRIPTION OF INDUSTRY BY ORE CATEGORY        II1-20
               Iron Ore                                          111-20
             ^ Copper Ore                                        111-31
               Lead and Zinc Ores                                111-38
              ^-Gold Ore                                          111-50
              ^Silver Ores                                       JII-54
               Bauxite                                           111-60
              'Ferroalloy Ores                                   II[-61
               Mercury Ores                                      111-85
               Uranium, Radium, and Vanadium Ores                II1-91
            v.  Metal Ores, Not Elsewhere Classified              1EI-1L8

IV        INDUSTRY CATEGORIZATION                                1V-J

          INTRODUCTION                                           IV-1
          MINE                                                   IV-2
          MILL                                                   IV-2

          FACTORS INFLUENCING SELECTION OF SUBCATEGORIES IN
          ALL METAL ORK CATEGORIES                               IV-)

          DISCUSSION OK 1'KIMAKY FACTORS INFLUENCING
          SUBCATEGORI/.ATLON BY ORK CATEGORY                      IV-9
               Iron Ore                                          IV-10
               Copper On-                                        IV-1A
               Lead and X.inc Ori?s                                IV-13
               Gold Ores                                         IV-11)
               Silver Ores                                       JV-21
               Bauxite Ores                                      1V-2J
                               DRAFT

-------
                              DRAFT
                             CONTENTS  (cont.)

Section                                                          Page

IV (cont.)     Ferroalloy Ores                                   IV-23
               Mercury  Ores                                      IV-26
               Uranium,  Radium, and Vanadium Ores                IV-27
               Metal  Ores, Not  Elsewhere Classified              IV-32

          SUMMARY OF  RECOMMENDED SUBCATEGORIZAT10N               IV-35

V         WASTE CHARACTERIZATION                                 V-l

          INTRODUCTION                                           V-l

          SPECIFIC WATER USES IN ALL  CATEGORIES                  V-3
               Noncontact Cooling Water                          V-3
               Wash Water                                        V-4
               Transport Water                                   V-4
               Scrubber  Water                                    V-4
               Process and Product Consumed Water                V-4
               Miscellaneous Water                               V-4

          PROCESS WASTE  CHARACTERISTICS BY ORE CATEGORY          V-5
               Iron Ore                                          V-5
               Copper Ore                                        V-21
               Lead and  Zinc. Ores                                V-69
               Gold Ores                                        V-78
               Silver Ores                                       V-87
               Bauxite Ores                                      V-102
               Ferroalloy Ores                                   V-lll
               Mercury Ores                                      V-137
               Uranium,  Radium, and Vanadium Ores                V-149
               Metal  Ores - Not Elsewhere Classified             V-173
                    Antimony                                     V-l83
                    Beryllium                                    V-185
                    Rare Earths                                  V-188
                    Plat fnum-Groitp MeLals                        V-196
                    Tin                                          V-J98
                    Titanium                                     V-200
                    Zirconium                                    V-207

VI        SELECTION OK POLLUTANT I'AKAMKTKRS                      VI-1

          INTRODUCTION                                           VI-1
                             DRAFT

-------
                             DRAFT
                               vil
                              DRAFT
                           CONTENTS (cont.)

Section

VI        GUIDELINE PARAMETER-SELECTION CRITERIA
(cont.)
          SIGNIFICANCE AND RATIONALE FOR SELECTION OF
          POLLUTION PARAMETERS                                   VI-2

          SIGNIFICANCE AND RATIONALE FOR REJECTJON OF
          POLLUTION PARAMETERS                                   VI-2'J

          SUMMARY OF POLLUTION PARAMETERS SELECTED BY
          CATEGORY                                               VI-JI

VII       CONTROL AND TREATMENT TECHNOLOGY                       Vll-l

          INTRODUCTION                                           VI1-2

          CONTROL PRACTICES AND TECHNOLOGY                       VII-2
               Mining Techniques                                 VII-2
               Surface-Water Control                             VII-6
               Segregation or Combination of Mine and Mill
                    Wastewaters                                  VII-7
               Regrading                                         VII-8
               Erosion Control                                   VII-10
               Revegetatlon                                      VII-12
               Exploration, Development, and Pilot-Scale
                    Operations                                   VI1-15
               Mine and Mill Closure                             VI1-17

          TREATMENT TECHNOLOGY                                   VI L -1l)
               Impoundment Systems                               VI1-20
               Clarifiers and Thickeners                         V1I-23
               Flocculation                                      VII-25
               Centrifugation                                    VII-26
               Hydrocyclones                                     VII-26
               Filtration                                        V1I-27
               Neutralization                                    VI]-28
               Chemical Free I pi tjt Ion Processes                  VT1-.10
               ReducLlon                                         VII-4J
               Oxidation, Aeration, and Air Stripping            Vll-AJ
               Adsorption                                        VIT-44
               Ion Exchange                                      V1I-40
               UlLral I I tratlon and Reverse Osmosis               VTl-S/i
               High-Dens Ity-SLudge Acid Neutral Iz.iL Ion           Vll-SS

-------
                             DRAFT
Section
VII (cont.)
VIII
                           CONTENTS  (cont.)
     Solvent Extraction                                VII-57
     Evaporation and Distillation                      VII-57
     Techniques for Reduction of Wastewater Volume     VII-58
     Electrodialysis                                   VII-6]
     Freezing                                          VII-62
     Biological Treatment                              VII-62

EXEMPLARY TREATMENT OPERATIONS BY ORE CATEGORY         VII-64
     Iron Ore                                          VII-64
     Copper Ores                                       VII-69
     Lead and Zinc Ores                                VII-84
     Gold Ores                                         VII-98
     Silver Ores                                       VII-109
     Bauxite Ore                                       VII-117
     Ferroalloy Ores                                   VII-121
     Mercury Ores                                      VII-148
     Uranium, Radium, and Vanadium Ores                VII-150
     Metal Ores, Not Elsewhere Classified              VII-159
          Antimony Ores                                VII-159
          Beryllium Ores                               VII-161
          Platinum-Group Metals                        VII-161
          Rare-Earth Ores                              VII-162
          Tin Ores                                     VII-164
          Titanium Ores                                VII-164
          Zirconium Ores                               VII-171

NONWATER-QUALITY ENVIRONMENTAL ASPECTS                 VII-171

COST, ENERGY, AND NONWATER-QUALITY ASPECTS             VIII-1

INTRODUCTION                                           VIII-1

SUMMARY OF METHODS USED                                VIII-2
     Capital Costs                                     VITI-2
     Annual Costs                                      VI LI-b

WASTEWATER-TREATMENT COSTS FOR IKON-OKK CATKCORY       V1II-8
     Iron-Ore Mines                                    Vlll-S
     Iron-Ore Mills Employing Chemlc.il/IMiys U:«l
          Separation                                   VIII-12
                              viJI
                             DRAFT

-------
                             DRAFT
                           CONTENTS (cont.)
Section
VIII      WASTEWATER TREATMENT COSTS FOR COPPER-ORE CATEGORY     VIII-17
(cont.)        Copper Mines                                      VIII-17
               Copper Mills Using Froth Flocculation
                    (Precipitation Minus Evaporation = -76 cm
                    to Positive) (-30 in. to Positive)           VlII-20

          WASTEWATER-TREATMENT COSTS FOR LEAD- AND ZINC-ORE
          CATEGORY                                               VII1-25
               Lead/Zinc Mines with No Solubility Potential      VITI-25
               Lead/Zinc Mines with Solubility Potential         VIII-28
               Lead/Zinc Mills                                   VIIT-32

          WASTEWATER-TREATMENT COSTS FOR GOLD-ORE CATEGORY       VIII-38
               Gold Mines (Alone)                                VIII-38
               Gold Mills or Mine/Mills (Cyanidation Process)    VIII-43
               Gold Mills (Amalgamation Process)                 VIII-47
               Gold Mills (Flotation)                            VIII-52
               Gold Mine/Mills Employing Gravity Separation      VIII-57

          WASTEWATER-TREATMENT COSTS FOR SILVER-ORE CATEGORY     VIII-62
               Silver-Ore Mines                                  VIII-62
               Silver Mills Employing Cyanidation, Amalgama-
                    tion, Gravity Separation, and Byproduct
                    Recovery                                     VIII-67
               Silver Mines Employing Flotation Process          VIII-68

          WASTEWATER-TREATMENT COSTS FOR BAUXITE CATEGORY        VII1-7L
               Bauxite Mines                                     VIII-7]
          WASTEWATER-TREATMENT COSTS FOR FERROALLOY-ORE
          CATEGORY
               Ferroalloy-Ore Mines
               Ferroalloy Mine/Mills Annually Processing
                    Less Than 5,000 Metric Tons (5,500
                    Short Tons) Ore by Methods Other Than
                    Ore Leaching
               Ferroalloy Mills Annually Processing More
                    Than 5,000 Metric Tons (5,500 Short
                    Terns) Ore by 1'hyslcal Methods
               Ferroalloy Mills Annually Processing More
                    Than 5.000 Metric Tons (5.500 Short
                    Tons) Ore by Flotation
               Ferroalloy Mills Practicing Ore Leaching
VIII-7-S
vr.r I-TJ
VII1-7S
VLIT-83
VII1-87
VI1I-95
                               ix
                             DRAFT

-------
                                 DRAFT
Section

VIII
(cont.)
IX
                            CONTENTS  (cont.)
WASTEWATER TREATMENT  COSTS  FOR  MERCURY-ORE CATEGORY
     Mercury-Ore  Mines
     Mercury Mills  Employing  Flotation Process
     Mercury Mills  Employing  Gravity Separation

WASTEWATER TREATMENT  COSTS  FOR  URANIUM-ORE CATEGORY
     Uranium Mines
     Uranium Mills  Using Acid or Combined Add/
          Alkaline  Leaching
     Uranium Mills  Using Alkaline Leaching

WASTEWATER TREATMENT  COSTS  FOR  METAL ORES, NOT
ELSErfWERE CLASSIFIED
     Antimony Mines
     Titanium Mines
     Titanium Mills Employing Electrostatic
          and/or  Magnetic Separation with Gravity
          and/or  Flotation  Process
     Platinum Mine/Mills Employing Dredging

BEST PRACTICABLE  CONTROL TECHNOLOGY CURRENTLY
AVAILABLE, GUIDELINES AND LIMITATIONS

INTRODUCTION

GENERAL WATER GUIDELINES
     Process Water
     Cooling Water
     Storm-Water  Runoff

BEST PRACTICABLE  CONTROL TECHNOLOGY CURRENTLY
AVAILABLE BY ORE  CATEGORY AND SUBCATEGORY
     Iron Ores
     Copper Ores
     Lead and ZJnc Ores
     Gold Orc.s
     Silver Orus
     Bauxite Ores
     Ferroalloy Ores
     Mercury Ores
     Uranium, Kudlum, ami Vaiim! lum Ore.::
VIII-103
VIII-103
VIII-107
V1II-112

V1II-115
VTII-115

VI11-125
VI IL-128
                                                                 VIII-132
                                                                 VIII-132
                                                                 VIII-135
                                                                 VIII-138
                                                                 VIII-1A1
                                                                 IX-1

                                                                 IX-1

                                                                 IX-2
                                                                 IX-2
                                                                 IX-2
                                                                 IX-3
                                                                 IX-4
                                                                 IX-4
                                                                 1X-8
                                                                 IX-14
                                                                 IX-18
                                                                 IX-26
                                                                 TX-J5
                                                                 IX-J5
                                                                 IX-4 6
                                                                 1X-51
                                DRAFT

-------
                                DRAFT
                           CONTENTS (cont.)

Section                                                          Page

IX (cont.)     Metal Ores, Not Elsewhere Classified              IX-55
                    Antimony Ores                                IX-55
                    Beryllium Ores                               IX-58
                    Platinum Ores                                IX-59
                    Rare-Earth Ores                              JX-59
                    Tin Ores                                     IX-62
                    Titanium Ores                                !X-d2
                    Zirconium Ores                               1X-66

X         BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE,
          GUIDELINES AND LIMITATIONS                             X-]

          INTRODUCTION                                           X-J

          GENERAL WATER GUIDELINES                               X-2
               Process Water                                     X-2
               Cooling Water                                     X-3
               Storm-Water Runoff                                X-3

          BEST AVAILABLE TECHNOLOGY ECONOMICALLY ACHIEVABLE,
          BY ORE CATEGORY AND SUBCATEGORY                        X-5
               Iron Ores                                         X-5
               Copper Ores                                       x~7
               Lead and Zinc Ores                                X-10
               Gold Ores                                         X-13
               Silver Ores                                       X-15
               Bauxite Ores                                      X-J 7
               Ferroalloy Ores                                   X-19
               Mercury Ores                                      X-26
               Uranium, Radium, and Vanadium Ores                X-28
               Metal Ores, Not Elsewhere Classified              X-29
                    Antimony Oros                                X-29
                    Beryllium Ores                               X-.ll
                    PLitinum Ores                                *--*'
                    Raro-Kartli Oros                              X-J3
                    Tin Orus                                     X-JJ
                    Titanium Ores                                X-34
                    Zirconium Ores                               X-J5
                                  xl


                                DRAFT

-------
                                DRAFT
Section
XI
XII

XIII

XIV
                 CONTENTS (cont.)



NEW SOURCE PERFORMANCE STANDARDS AND PRETREATMENT
STANDARDS

INTRODUCTION

GENERAL WATER GUIDELINES

NEW SOURCE STANDARDS BY ORE CATEGORY
     Ferroalloy Ores
     Uranium Ores

PRETREATMENT STANDARDS

ACKNOWLEDGMENTS

REFERENCES

GLOSSARY

CHEMICAL ELEMENTS

CONVERSION TABLE
                                                                Page
XI-1

XL-1

XI-2

X[-2
XI-4
XI-6

XI-9

XII-1

XIII-1

XIV-1

XIV-26

XIV-27
                                  xlJ


                                DRAFT

-------
                               DRAFT
                               FIGURES
No.                             Title                            Page

III-l     Beneficiation of Iron Ores                             111-25
III-2     Iron-Ore Flotation-Circuit Flowsheet                   TTI-26
III-3     Magnetic Taconite Beneficiation Flowsheet              IJ1-29
II1-4     Agglomeration Flowsheet                                FI1-30
III-5     Major Copper Mining and Milling Zones of the U.S.      111-32
III-6     General Outline of Methods for Typical Recovery
               of Copper from Ore                                111-37
III-7     Major Copper Areas Employing Acid Leaching in
               Heaps, in Dumps, or In Situ                       111-40
IH-fl     Lead/Zinc-Ore Mining and Processing Operations         111-47
III-9     Cyanldation of Gold Ore:  Vat Leaching of Sands
               and "Carbon-in-Pulp1 Processing of Slimes         111-52
111-10    Cyanidation of Gold Ore;  Agitation/Leach
               Process                                           111-53
III-ll    Flotation of Gold-Containing Minerals with
               Recovery of Residual Gold Values by
               Cyanidation                                       I11-55
111-12    Recovery of Silver Sulfide Ore by Froth
               Flotation                                         111-59
111-13    Gravity-Plant Flowsheet for Nigerian Columbite         111-72
111-14    Euxenite/Columbite Beneflciation-Plant Flowsheet       111-73
111-15    Representative Flow Sheet for Simple Gravity
               Mill                                              111-74
111-16    Simplified Molybdenum Mill Flowsheet                   111-77
111-17    Simplified Molybdenum Mill Flow Diagram                111-79
111-18    Simplified Flow Diagram for Small Tungsten
               Concentrator                                      111-82
111-19    Mill Flowsheet for a Canadian Columbium
               Operation                                         111-83
111-20    Flowsheet of Tristage Crystallization Process
               for Recovery of Vanadium, Phosphorus, and
               Chromium from Western Ferrophosphorus             111-86
HI-21    Arkansas Vanadium Process Flowsheet                    111-87
111-22    Flowsheet of Dean-Leute Ammonium CarbamaLe
               Process                                           111-88
IIL-23    Pachuca Tank for Alkaline Leaching                     111-99
If1-24    Concentration Processes and Terminology                111-104
111-25    Simplified Schematic Diagram of Sulfurlc Acid
               Digestion of Monazlte Sand for Recovery
               of Thorium, Uranium, and Rare Earths              III'-109
                                xlll


                               DRAFT

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

 No.                              Title                           Page

 111-26    Simplified Schematic Diagram of Caustic Soda
                Digestion of Monazite Sand for Recovery
                of Thorium, Uranium, and Rare Earths              I1I-1LO
 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                             Ill-Ill
 111-28    Generalized Flow Diagram for Production of
                Uranium, Vanadium, and Radium                     ILI-JIb
 111-29    Beneficiation of Antimony Sulfide Ore by
                Flotation                                         JII-120
 111-30    Gravity Concentration of Platinum-Group Metals          III-124
 HI-31    Beneficiation of Heavy-Mineral Beach Sands             III-130
 111-32    Beneficiation of Ilmenite Mined from a Rock Deposit     III-131
 V-l       Flow Scheme for Treatment of Mine Water                V-12
 V-2       Water Flow Scheme in a Typical Milling Operation        V-12
 V-3       Water Balance for Mine/Mill 1105 (September 1974)       V-14
 V-4       Concentrator Flowsheet for  Mill 1105                   v-16
 V-5       Flowsheet for Mill 1104 (Heavy-Media Plant)             V-20
 V-6       Simplified Concentration Flowsheet for Mine/Mill 1108   V-23
 V-7       Wastewater Flowsheet for Plant 2120-B Pit               V-28
 V-8       Flowsheet of Hydrometallurgical Process Used in
                Acid Leaching at Mine  2122                        V-37
 V-9       Reactions by Which Copper Minerals Are Dissolved In
                Dump,  Heap,  or In-Situ Leaching                   v-38
 V-10       Typical Design of Gravity Launder/Precipitation
                Plant                                              v-40
 V-ll       Cutaway Diagram of Cone Precipitator                   v-41
 V-12       Diagram of  Solvent  Extraction  Process for  Recovery
                of Copper by Leaching  of  Ore  and Waste             V-43
 V-13       Vat  Leach Flow Diagram (Mill 2124)                      V-4 7
 V-14       Flow Diagram for  Flotation  of  Copper (Mill  2120)        V-S3
 V-15       Addition of  Flotation Agents to Modify Mineral
                Surface                                            V-54
 V-16       Flowsheet for Miscellaneous Handling of Flotation
                Tails  (Mill  2124)                                  V-6S
 V-17       Dual  Processing of  Ore  (Mill 2124)                      V-h(>
 V-18       Leach/Pr€-clpItation/KLoLiit Ion  Process                  V-dH
V-19       Water  FJow Diagram  for  Mine 3105                        V-71
V-20       Water  Klow D Lag rum  for  Mint-  3104                        V-7(.
V-21       Flow  Diagram  for  Mill  'IJ01                              V-7'l
                                xlv


                               DRAFT

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

—                               Title

V-22       Water  Flow In Four Selected  Gold Mining and
               Milling  Operations                                 V-82
V-23       Water  Flow in Silver  Mines and  Mills                    V-92
V-24       Process  and Wastewater Flow  Diagram for Open-Pit
               Bauxite  Mine  5101                                 V-104
V-25       Mill 6601  Flowsheet                                     V-121
V-26       Simplified Mill  Flow  Diagram for Mill  6102              V-124
V-27       Internal Water Flow For  Mill 6104  Through
               Molybdenum  Separation                              V-130
V-28       Internal Water Flow for  Mill 6104  Following
               Molybdenum  Separation                              V-131
V-29       Water  Use  and Waste Sources  for Vanadium Mill  6107      V-138
V-30       Water  Flow in Mercury Mills  9101 and 9102               V-143
V-31       Typical  Water-Use  Patterns                              V-150
V-32       Alkaline-Leach Water  Flow                              V-157
V-33       Ammonium Carbonate Leaching  Process                     V-J58
V-34       Water  Flow in Mills 9401, 9402, 9403,  and 9404          V-161
V-35       Flowchart  of  Mill  9401                                 V-163
V-36       Flow Chart for Mill 9402                                V-164
V-37       Flow Chart of Mill 9403                                 V-165
V-38       Flow Chart of Mill 9404                                 V-166
V-39       Water  Flows and Usage for Mine/Mills 9901 (Antimony)
               and 9902  (Beryllium)                               V-179
V-40       Water  Flows and Usage for Mine/Mills 9903
               (Rare  Earths) and 9904  (Platinum)                  V-180
V-41       Water  Flows and Usage for Titanium Mine/Mills
               9905  and  9906                                     V-181
V-42       Beneficiation  of Bertrandite, Mined from a Lode
               Deposit  by Flotation (Mill 9903)                   V-191
V-43       Beneficiation  of Rare-Earth  Flotation  Concentrate
               by  Solvent Extraction (Mill 9903)                  V-192
V-44       Beneficiation  and  Waste Water Flow of  IJmenite
               Mine/Mill 9905 (Rock Deposit)                      V-202
V-45       Beneficiation  of Heavy-Mineral  Bcat-li Sands (RuLilc,
               1Imenite, Zircon, and Monazltc) aL Mill 9906       V-J06
VII-1      Lime Neutralization and Prcrlplt.it Ion  I'TOCL-HH  for
               Treatment o[  Mine Wat«.-r I'rlur to  Dlscluir^i-         VI I-12
VII-2     Theoretical Solubilities of MoLuJ.Lons .is .1
               Function  of pll                                     VI I - H
                               xv
                             DRAFT

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

No.                             Title                            Page

VII-3     Minimum pH Value for Complete Precipitation
               of Metal Ions as Hydroxides                       VII-34
VII-4     Heavy-Metal Precipitation vs pH for TaJ]ing-Pond
               Effluent pH Adjustments by Lime Addition          VI1-16
VII-5     Diagram of Modified Desal Process                      VI1-51
VII-6     Mill 1105 Water-Use System (Zero Discharge)            VTI-68
VII-7     Control of Effluent by Reuse of Mine Water in
               Leaching (Mine 2122)                              V1I-70
VII-8     Control of Mine-Water Effluent by Reuse in the
               Concentrator (Mine/Mill 2119)                     VII-71
VII-9     Control of Effluent Through Reuse of Mill Flotation-
               Process Water in Other Facilities
               (Mine/Mill 2124)                                  VII-78
VII-10    Reduction in Waste Pollutant Load in Discharge
               by Separation of Minewater From Tailing Pond
               for Separate Treatment (Mill 2121)                VII-80
VII-11    Schematic Diagram of Treatment Facilities at
               Mine 3107                                         VII-89
VII-12    Schematic Diagram of Water Flows and Treatment
               Facilities at Mill 3103                           VII-92
VII-13    Schematic Diagram of Water Flow and Treatment
               Facilities at Mill 3102 (Tailing Pond/Stilling
               Pond/Biological Treatment/Polishing Pond)         VII-95
VII-14    Schematic Diagram of Water Flow and Treatment
               Facilities at Mill 3105                           VII-97
VII-15    Schematic Diagram of Treatment Facilities at
               Mill 3101                                         VII-99
VII-16    Lime-Neutralization Plant for Open-Pit Mine 5102       VII-119
VII-17    Water-Flow Schematic Diagram for Mill 6102             VII-133
VII-18    Ion Exchange for Mercury and Uranium at Low
               Loadings and Concentrations                       VII-154
VII-19    Chemical Changes in a Sequence of Tailing
               Impoundments at Mill 9402                         VII-158
                                xv i
                               DRAFT

-------
                             DRAFT
No.
                                                                 Page
II-l      Summary of Recommended BPCTCA Effluent Limitations
               By Category and Subcategory—Ores for Which
               Separate Limitations Are Proposed                 J1-2
II-2      Summary of Recommended BATEA Effluent Limitations
               By Category and Subcategory—Ores for Which
               Separate Limitations Are Proposed                 11-4
II-3      Summary of Recommended NSPS Effluent Limitations
               By Category and Subcategory—Ores for Which
               Separate Limitations Are Proposed                 [1-6
III-l     Iron-Ore Shipments for United States                   111-21
III-2     Crude Iron-Ore Production for U.S.                     111-22
III-3     Reagents Used for Flotation of Iron Ores               111-27
III-4     Various Flotation Methods Available for Pro-
               duction of High-Grade Iron-Ore Concentrates       111-28
III-5     Total Copper-Mine Production of Ore by Year            111-33
III-6     Copper-Ore Production from Mines by State (1972)       111-33
III-7     Average Copper Content of Domestic Ore                 111-35
III-8     Average Concentration of Copper in Domestic Ores
               by Process (1972)                                 111-35
III-9     Copper Ore Concentrated in the United States
               by Froth Flotation, Including LPF Process
               (1972)                                            IH-36
111-10    Heap or Vat Ore Leached in the United States (1972)    111-39
III-ll    Average Price Received from Copper in the
               United States                                     III-A1
111-12    Production of Copper from Domestic Ore by
               Smelters                                          111-42
111-13    Mine Production of Recoverable Lead in the
               United States                                     111-44
111-14    Mine Production of Recoverable Zinc in the
               United States (Preliminary)                       111-45
111-15    Domestic Silver Production from Different
               Types of Ores                                     111-57
111-16    Silver Produced at Amalgamation and Cyanldation
               Mills in the U.S. and Percentage of Silver
               Recoverable from All Sources                      111-58
111-17    Production of Bauxite in the United States             IH-62
111-18    Production of Ferroalloys by U.  S. Mining and
               Milling Industry                                  111-64
111-19    Observed Usage of Some Flotation Reagents              111-76
111-20    Probable Reagents Used in Flotation of Nickel
               and Cobalt Ores                                   111-80
                              xvii
                             DRAFT

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

No.                             Title                            Page

111-21    Domestic Mercury  Production  Statistics                 111-89
111-22    Isotopic Abundance of Uranium                          ITI-93
111-23    Uranium Milling Activity  by  State, 1972                ITI-97
111-24    Uranium Concentration in  IX/SX Eluates                 ITI-103
111-25    Decay  Series  of Thorium and  Uranium                    111-113
111-26    Uranium Milling Processes                             III-114
111-27    Uranium Production                                    III-117
111-28    Vanadium Production                                   111-117
111-29    Vanadium Use                                           III-117
111-30    Production of Antimony from  Domestic Sources           III-119
111-31    Domestic Platinum-Group Mine Production aiul Value      I FT-123
111-32    Production and  Mine Shipments of Titanium
               Concentrates from Domestic Ores in the U.S.       111-128
IV-1      Summary of Industry Subcategorlzatlon Recommended      fV-36
V-l       Historical Constituents of Iron-Mine Discharges        V-9
V-2       Historical Constituents of Wastewater from Iron-
               Ore Processing                                   V-9
V-3       Chemical Compositions of  Sampled Mine Waters           V-10
V-4       Chemical Compositions of  Sampled Mill Waters           V-10
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-18
V-6       Chemical Characteristics  of  Discharge Water
               from Mine  1108                                   V-22
V-7       Characteristics of Mill 1108 Discharge Water           V-24
V-8       Principal Copper  Minerals Used in the United States    V-26
V-9       Mine-Water Production from Selected Major Copper-
               Producing Mines and  Fate(s) of Effluent           V-29
V-10      Summary of Solid  Wastes Produced by Plants
               Surveyed                                          V-30
V-ll      Raw Waste Load  In Water Pumped from Selected
               Copper Mines                                     V-.12
V-12      1973 Water Usage  Ln Dump, Heap, and Tn-SLtu
               Leaching Operations                               V-4S
V-13      Chemical Character!stJos  of  Barren Heap, Dump, or
               In-Situ Acid Leach Solutions (Recycled:  No
               Waste Load)                                       V-4h
V-14      Water Usage In Vat Leaching  Process as a Function
               of Amount of Product (Precipitate or Cdtliode
               Copper) Produced                                  V-49
V-15      Chemical Character 1stIcs  of  Vat-Leach Barren
               Acid Solution (Recycled:  No Waste Loud)          V-50
                                xvlll
                               DRAFT

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

No.                              Title                           Page

V-16      Miscellaneous Wastes from Special Handling uf
               Ore Wash Slimes in Mine 2124 (No Effluent)        V-51
V-17      Examples of Chemical Agents Which May be Employed
               In Copper Flotation                               V-55
V-18      Water Usage in Froth Flotation of Copper               V-58
V-19      Raw Mill Waste Loads Prior to Settling in Tailing
               Ponds                                             V-59
V-20      Wastewater Constituents and Waste Loads Resulting
               from Discharge of Mill Process Waters             V-63
V-21      Range of Chemical Characteristics of Sampled Raw
               Mine Water from Lead/Zinc Mines 3102, 3103,
               and 3104                                          V-72
V-22      Range of Chemical Characteristics of Raw Mine
               Waters from Four Operations in SolubiJiza-
               tion-Potential Subcategory                        V-77
V-23      Ranges of Constituents of Wastewaters and Raw Waste
               Loads for Mills 3102, 3103, 3104, 3105, and
               3106                                              V-80
V-24      Chemical Composition of Raw Mine Water from Mines
               4105 and 4102                                     V-84
V-25      Process Reagent Use at Various Mills Beneficiating
               Gold Ore                                          V-88
V-26      Minerals Commonly Associated with Gold Ore             V-88
V-27      Waste Characteristics and Raw Waste Loads at Four
               Gold Milling Operations                           V-89
V-28      Raw Waste Characteristics of Silver Mining
               Operations                                        V-95
V-29      Major Minerals Found Associated with Silver Ores       V-97
V-30      Flotation Reagents Used by Three Mills to Bene-
               ficiate Silver-Containing Mineral Tetrahedrite
               (Mills 4401 and 4403) and Native Silver and
               Argentite (Mill 4402)                             V-99
V-31      Waste Characteristics and Raw Waste Loads at Mills
               4401, 4402, 4403, and 4105                        V-IOO
V-32      Concentrations of Selected Constituents in Acid
               Raw Mine Drainage from Open-Pit Mine 5101         V-J07
V-33      Concentrations of Selected Constituents in Acid
               Raw Mine Drainage from Open-Pit Mine 5J02         V-J07
V-34      Concentrations of Selected Constituents In Alkaline
               Raw Mine Drainage from Underground Mine 5101      V-108
V-35      Wastewater and Raw Waste Load for Open-Pit Mine 5101   V-110
V-36      Wastewater and Raw Waste l.ot-ul for Underground
               Mine 5101                                         V-110
V-37      Types of Operations Visited and Anl ic Ipaf eel —
               Ferroal Joy-Ore Mini UK "»(1 Dressing Industry       V-112
                                xlx


                              DRAFT

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

No.                             Title                            page

V-38      Chemical Characteristics  of  Raw Mine Water in
               Ferroalloy  Industry                               V-116
V-39      Reagent Use  in Molybdenum Mill 6101                    V-122
V-AO      Raw Waste  Characterization and Raw Waste Load
               for Mill 6601                                    V-122
V-41      Reagent Use  for  Rougher and  Scavenger Flotation
               at Mill 6102                                      V-J25
V-42      Reagent Use  for  Cleaner Flotation at Mill 6102         V-125
V-43      Reagent Use at Byproduct  Plant of Mill 6102 (Based
               on Total Byproduct Plant Feed)                    V-127
V-44      Mill 6102  Effluent Chemical  Characteristics (Com-
               bined-Tailings Sample)                            V-127
V-45      Chemical Characteristics  of  Acid-Flotation Step        V-128
V-46      Composite  Waste  Characteristicis for Beneficiation
               at Mill 6104  (Samples 6, 8, 9, and 11)            V-132
V-47      Waste Characteristics from Copper-Thickener Over-
               flow  for Mill 6104 (Sample 5)                     V-132
V-48      Scheelite-Flotation Tailing  Waste Characteristics
               and Loading for Mill 6104 (Sample 7)              V-133
V-49      50-Foot-Thickener  Overflow for Mill 6104 (Sample 10)   V-133
V-50      Waste Characteristics of  Combined-Tailing Discharge
               for Mill 6104 (Samples  15, 16, and 17)            V-134
V-51      Waste Characteristics and Raw Waste Load at Mill
               6105  (Sample  19)                                  V-136
V-52      Chemical Composition of Wastewater, Total Waste,
               and Raw Waste Loading from Milling and Smelter
               Effluent for  Mill 6106                            V-136
V-53      Waste Characterization and Raw Waste Load for
               Mill  6107 Leach and  Solvent-Extraction Effluent
               (Sample 80)                                       V-139
V-54      Waste Characteristics and Waste Load for Dryer
               Scrubber Bleed at Mill  6107 (Sample 81)           V-140
V-55      Waste Characteristics and Loading for Salt-Roast
               Scrubber Bleed at Mil]  6107 (Sample 77)           V-I4I
V-56      Expected Reagent Use .it Mercury-Ore Flotation
               Mill  9202                                         V-147
V-57      Waste Characteristics and Knw Waste Loadings at
               Mills 9201  and 9202                               V-148
V-58      Waste Constituents Expected                            V-153
V-59      Chemical and PliysicaJ Waste  Constituents Observed
               in Representative Operations                      V-L54
                               DRAFT

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

No.                             Title                            Page

V-60      Water Use and Flows at Mine/Mills 9401, 9402, 9403,
               and 9404                                          V-162
V-61      Water Treatment Involved  In U/Ra/V Operations          V-162
V-62      Radionuclides in Raw Waste-waters from Uranium/
               Radium/Vanadium Mines and Mills                   V-169
V-63      Organic Constituents in U/Ra/V Raw Wastewater          V-169
V-64      Inorganic Anions in U/Ra/V Raw Wastewater              V-171
V-65      Light-Metal Concentrations Observed in U/Ra/V
               Raw Wastewater                                    V-171
V-66      Concentrations of Heavy Metals Forming Anionic
               Species in U/Ra/V Raw Wastewater                  V-171
V-67      Concentrations of Heavy Metals Forming Cationic
               Species in U/Ra/V Raw Wastewater                  V-172
V-68      Other Constituents Present in Raw Wastewater in
               U/Ra/V Mines and Mills                            V-172
V-69      Chemical Composition of Wastewater and Raw Waste
               Load for Uranium Mines 9401 and 9402              V-174
V-70      Chemical Composition of Raw Wastewater and Raw
               Waste Load for Mill  9401 (Alkaline-Mill
               Subcategory)                                      V-174
V-71      Chemical Composition of Wastewater and Raw Waste
               Load for Mill 9402 (Acid- or Combined Acid/
               Alkaline-Mill Subcategory)                        V-175
V-72      Chemical Composition of Wastewater and Raw Waste
               Load for Mine 9403 (Alkaline-Mill Subcategory)    V-176
V-73      Chemical Composition of Wastewater and Raw Waste
               Load for Mill 9404 (Acid- or Combined Acid/
               Alkaline-Mill Subcategory)                        V-177
V-74      Reagent Use at Antimony-Ore Flotation Mill 9901        V-184
V-75      Chemical Composition of Raw Wastewater Discharged
               From Antimony Flotation Mill 9901                 V-186
V-76      Major Waste Constituents and Raw Waste Load at
               Antimony Mill 9901                                V-187
V-77      Chemical Composition of Raw Wastewater from
               Beryllium Mill 9902  (No Discharge from
               Treatment)                                        V-189
V-78      Chemical Composition of Raw Wastewater from
               Rare-Earth Mill 9903                              V-194
V-79      Results of Chemical Analysis for Rare-Earth
               Metals (Mill 9903—No Discharge)                  V-195
                                 xxl


                               DRAFT

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

 No.                             Title                            page

 V-80      Chemical Composition and  Raw Waste Load from
                Rare-Earth Mill 9903                              V-197
 V-81      Chemical Composition and  Loading for Principal
                Waste Constituents Resulting from Platinum
                Mine/Mill 9904 (Industry Data)                    V-199
 V-82      Chemical Composition of Raw Wastewater from
                Titanium Mine 9905                                V-201
 V-83      Chemical Composition of Raw Wastewater From
                Titanium Mill 9905                                V-204
 V-84      Reagent Use in Flotation Circuit of Mill 9905          V-204
 V-85      Principal Minerals Associated with Ore of Mine 9905    V-205
 V-86      Major Waste Constituents and Raw Waste Load at
                Mill 9905                                         V-205
 V-87      Chemical Composition of Raw Wastewater at Mills
                9906 and 9907                                     V-208
 V-88      Raw Waste Loads for Principal Wastewater Consti-
                tuents from Sand Placer Mills 9906 and 9907       V-209

 VI-1      Known Toxicity of Some Common Flotation Reagents
                Used in Ore Mining and Milling Industry           VI-27
 VI-2      Summary of Parameters Selected for Effluent Limi-
                tation by Metal Category                          VI-32

 VII-1      Results of Coprecipitation Removal of Radium
                from Wastewater                                   VII-40
 VII-2      Properties of Ion Exchangers for Metallurgical
                Applications                                      V1I-48
 VII-3      Analytical Data for Modified Desal Process             VIT-53
 VII-4      Rejection of Metal Salts by Reverse-Osmosis
                Membranes                                         VTJ-56
 VII-5      Chemical Characteristics of Settling-Pond  Dis-
                charge at Mine 1105                               VTI-65
 VII-6      Chemical Compositions  of Raw and Treated  Waste-
                loading at Mine/Mill  1109                          VTI-67
 VII-7      Concentration of Parameters I'rosi>nf  Ln  Raw
                Wastewater iind  Effluent Following  Lime
                Precipitation nL  Mine 212011                       VI1-72
 VII-8      Concentration of I'araineters I'ro.soiit  In  Knw W.isLe-
                water  and Effluent  Following Lime  I'reciplta-
                tion at  Mine 2120C                                 VI1-73
VII-9      Dump, Heap,  and In-SJ tu  l.cdrh-Solut Inn  Control
                and  Treatment Pr«.ii_tii-i- (I97J)                      VI 1-75
                                 xx I i
                                DRAFT

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

No-                              Title                           Page

VII-10    Solution-Control Practice in Vat Leaching of
               Copper Ore                                        VII-77
VII-11    Reduction of Pollutants in Concentrator Tails
               by Settling at Various pH Levels                  VII-82
VII-12    Efficiency of Coagulation Treatment to Reduce
               Pollutant Loads in Combined Waste (Includ-
               ing Mill Waste) Prior to Discharge (Pilot
               Plant)                                            VII-83
VII-13    Chemical Compositions of Raw and Treated Mine-
               waters from Mine 3105 (Historical Data Pre-
               sented for Comparison)                            VII-85
VII-14    Chemical Compositions of Raw and Treated Waste-
               waters from Mine 3107 (Historical Data Pre-
               sented for Comparison)                            VII-87
VII-15    Chemical Compositions of Raw and Treated Mine
               Waters from Mine 3101                             VII-90
VII-16    Chemical Compositions and Waste Loads for Raw and
               Treated Mill Wastewaters at Mill 3103             VII-93
VII-17    Chemical Composition and Waste Loading for Raw
               and Treated Mill Wastewater Mill 3102             VII-96
VII-18    Waste Compositions and Raw and Treated Waste Loads
               Achieved at Mill 4102 by Tailing-Pond Treat-
               ment                                              VII-103
VII-19    Chemical Compositions of Mill Wastewater and
               Tailing-Pond Decant Water at Mill 4101 (No
               Resultant Discharge)                              VII-105
VII-20    Waste Compositions and Raw and Treated Waste
               Loads at Mill 4401 (Using Tailing-Pond
               Treatment and Partial Recycle)                    VII-113
VII-21    Chemical Compositions of Mill Raw Wastewater
               and Tailing-Pond Decant Water at Mill 4402        V1I-116
VII-22    Chemical Compositions of Raw and Treated Mine
               Waters at Mine 5101                               VII-120
VII-23    Chemical Compositions of Raw and Treated Mine
               Waters at Mine 5102                               VII-122
VII-24    Chemical Compositions of Raw Mine Wastewater
               and Treated Effluent at Mine 6103                 VII-124
VII-25    Chemical Compositions of R.iw and Treated Mine
               Waters at Mine 6104 (Cl.irl riorcuLatnr
               Treatment)                                        VII-L26
VII-26    Chemical Compositions of KMW and Treated Waste-
               waters at Mine 6L07                               VII-127
                             xxlil

                            DRAFT

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

No.                              Title                           Page

VII-27    Analyses of  Intake and Discharge Waters From Mill
               6101  (Company Data)                               VIT-130
V1I-28    Chemical Composition of WasLt-wa tor ami W;iste
               Loading for  Kill 6101                             Vll-lll
VII-29    Chemical Composition and Calculated Waste Load  for
               Mill  6102 Tailing-Pond Surface Water, with
               Analytical Data for Mi11-Reservoir Water          VII-135
VII-30    Chemical Composition and Waste Loading for Discharge
               at Mill 6102 (Company Data)                       VII-135
VII-31    Chemical Composition and Treated Waste Loads for
               Overflow from First Settling Pond at Mill  6106    VII-138
VII-32    Characteristics of Surface Water from Second Settling
               Pond  at Mill 6106                                 VII-138
VII-33    Chemical Composition and Treated Waste Loads from
               Final Effluent for Mine/Mill 6106 During
               Rainy Season (Company Data)                       VII-139
VII-34    Chemical Composition ami Waste Loading from Area
               Runoff  and Reclamation-Pond Seepage at Mill
               6107  (Company Data)                               VII-139
VII-35    Chemical Composition and Waste Loading for Cooling
               Water Effluent at Mill 6107 (Company Data)        VIi-140
VII-36    Chemical Composition and Waste Loading for Process
               Effluent After AmmonLa Treatment at Mill 6107     VII-142
VII-37    Chemical Composition and Waste Loading for Drier
               Scrubber Bleed Water After Settling Treatment
               at Mill 6107                                     VII-143
VII-38    Chemical Composition and Waste Loading for Holding-
               Pond  Effluent  (Process Water and Drier Scrubber
               Bleed)  at Mill 6107 (Company Data)                VII-144
VII-39    Chemical Composition and Waste Loading for Roaster
               Scrubber Bleed Water After Settling at Mill
               6107                                              VII-145
VII-40    Chemical Composition and Waste Loading for Roaster
               Scrubber Bleed Water After Settling .it Mill
               6107  (Company Data)                               VI1-146
V1I-41    Chemical Composition an
-------
                                DRAFT
                            TABLES (cont.)

No.                             Title                            Page

VII-43    Chemical Compositions of Raw and Treated
               Wastewaters at Mine 9402 (001)                    VII-154
VII-44    Chemical Compositions of Raw and Treated
               Wastewaters at Mine 9402 (002)                    VII-155
VII-45    Chemical Compositions of Raw and Treated Waste-
               waters and Effluent Waste Loading at Mill  9403
               (Settling and BaC12^ Coprecipitation)              Vfl-160
VII-46    Chemical Composition of Treated Effluent and
               Waste Load from Mine/Mill 9904 (PJatinum)          VTL-163
VII-47    Chemical Compositions of Raw Wastewatcr and Treated
               Recycle Water at Mill 9903 (No Discharge)          VII-163
VII-48    Chemical Compositions of Raw Wastewater and
               Treated Recycle Water at Mill 9905                VII-165
VII-49    Chemical Compositions of Raw and Treated
               Wastewaters at Mill 9906                          VII-167
VII-50    Chemical Compositions of Raw and Treated
               Wastewaters at Mill 9907                          VII-168
VII-51    Wastewater Composition and Treated Waste Load
               With Acid Flocculation and Settling at
               Mill 9906                                         VII-169
VI1-52    Wastewater Composition and Treated Waste Load
               With Acid Flocculation and Settling at
               Mill 9907                                         VII-170
VIII-1    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mine 1105          VIII-9
VIII-2    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mill 1107          VIII-13
VIII-3    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mine 2120          VII1-18
VIII-4    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mill 2121          VIII-21
VIII-5    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Typical
               Mine (Hypothet Jciil)—l-cad/Zinc, No Solubility     V11I-26
VIII-6    Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Typical
               Mine (Hypotheticdl)--Lead/Zlnc, Solubility        VIII-29
VIII-7    Water Effluent Treatment Costs and Resulting
               Waste-Loud Characteristics For Typical
               Mill (Hypothetical)—J.c.id/Zinc                    VllI-3/i
VIII-8    Water Effluent Treatment COHLS and Result Ing
               Waste-Load Character 1 sites fur Typical
               Mine (Hypothetical )—flo Id                         V1II-J9
                                 XXV
                               DRAFT

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

No.                              Title                            Page

VIII-9     Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Mil] 4105          VI FT-44
VIII-10    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Mill 4102          V11F-48
VIII-11    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristirs for Mill 4104          V1JI-5J
VIII-12    Water Effluent Treatment Costs and Kt-sultiug
                Waste-Load Characteristics fur Typical
                Mine/Mill (Hypothc-t Leal) — Gold/Gravity            VI11-58
VIII-13    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Typical
                Mine (Hypothetical)—Silver                       VIII-63
VIII-1A    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Mill 4401          VIII-69
VIII-15    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Mine 5102          VIII-72
VIII-16    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Typical
                Mine (Hypothetical)—Ferroalloy                   VIII-76
VIII-17    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Typical
                Mine/Mill (Hypothetical)— Ferroalloy/Limited      VUI-79
VIII-18    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristics for Typical
                Mill (Hypothetical)—Ferroalloy/Physical          VII1-84
VIII-19    Water Effluent Treatment Costs and Resulting
                Waste-Load Characteristies for Typical
                Mill (Hypothetical)— Fcrroalloy/F] oLation         VI1I-88
VIII-20    Water Effluent Treatment Costs and ResuJting
                Waste-Load Characteristics for Typical
                Mill (Hypothetical)--Ferroalloy/Laaching          VIII-96
VIII-21    Water Effluent Treatment Costs and Resulting
                Waste-Load CharacteristIvs for Typlval
                Mine (Hypothetical ) — Mc.n-ury                      VL 11-104
VIIT-22    Water Ef fluent Trt-.H iiii-nt (Justs .ind Rot.ii I L Inn
                W.isU-l.oml Ui.ira. u-rlsLl. s for Mill 4J02          VI1I-HW
Vm-23    Water Kl'l'luunl  TriMlmeiH Cn--,!-. ,md Kvsullinn
                W;i.sti—Lo.itl Cli.ir.un>-risL U's lurMIII ').U)I          VT11-.N3
VI 11-24    Wntor lifllutMit  Tivnt-iniMH d'KLs .nul RosnUlnj;
                WasLc-Lo.id ClULMcLoi i ht fi:^ Cur Typical
                Mine (llypotlu-Llc.il)—ULMII iinn                      VUI-LIG
                                 xxv I
                                DRAFT

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

No.                             Title                            Page

VIII-25   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mill 9405          VIII-126
VIII-26   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mill 9403          VI1I-130
VIII-27   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Typical
               Mine (Hypothetical)—Antimony                     VIII-133
VIII-28   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mine 9905          VIII-136
VIII-29   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mill 9905          VIII-139
VIII-30   Water Effluent Treatment Costs and Resulting
               Waste-Load Characteristics for Mine/Mill 9904     VIII-142

IX-1      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Iron-Ore Mines            TX-5
LX-2      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Iron-Ore Mills
               Employing Physical Methods and/or
               Chemical Reagents                                 IX-7
IX-3      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Copper Mines              IX-10
IX-4      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Copper Mills Using
               Froth Flotation (Net Evaporation Less Than
               76.2 cm (30 in.) per year)                        IX-13
IX-5      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Lead and Zinc
               Mines Having No Solubilization Potential          IX-15
IX-6      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Lead and Zinc Mines
               Having Solubilization Potential                   IX-17
IX-7      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Lead and/or Zinc Mills    IX-19
IX-8      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Gold Mines                IX-21
IX-9      Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Gold Mines Using
               Amalgamation Process                              IX-23
IX-10     Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Gold Mills Using
               Flotation Process                                 IX-25
                                xxv I i
                               DRAFT

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

No.                              Title                           page

IX-11      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Gold Mines  or Mills
                Using Gravity-Separation Methods                  1X-27
IX-12      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Silver Mines  (Alone)      IX-29
IX-13      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Silver Mills Using
                Amalgamation Process                             IX-32
IX-14      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Silver Mills Using
                Gravity Separation                               IX-34
IX-15      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Bauxite Mines (Acid
                or Alkaline Mine Drainage)                        IX-36
IX-16      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Ferroalloy-Ore Mines      IX-38
IX-17      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Ferroalloy-Ore Mills
                Treating Less Than 5,000 Metric Tons (5,512
                Short Tons)  Per Year                             IX-40
IX-18      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Ferroalloy-Ore Mills
                Treating More Than 5,000 Metric Tons (5,512
                Short Tons)  Per Year by  Physical Processing       IX-43
IX-19      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Ferroalloy-Ore Mills
                Using Flotation Process                           IX-45
IX-20      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Ferroalloy-Ore Mills
                Using Leaching Process                           IX-47
IX-21      Parameters Selected and Effluent  Limitations
                Recommended for BPCTCA—Mercury Mines             IX-49
IX-22      Parameters Selected and Effluent  Limitations
                Recommended  for BPCTCA—Uranium Mines             IX-53
IX-23      Parameters Selected and Effluent  Limitations
                Recommended  for BPCTCA—Antimony Mines            IX-57
IX-24      Parameters Selected and Effluent  Limitations
                Recommended  for BPCTCA—Platinum Mills            IX-60
IX-25      Parameters Selected and Effluent  Limitations
                Recommended  for BPCTCA—Titanium Mines            IX-63
                               xxviii


                               DRAFT

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

No.                             Title                            Page

IX-26     Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Titanium Mills            IX-65
IX-27     Parameters Selected and Effluent Limitations
               Recommended for BPCTCA—Titanium Dredge Mine
               With Wet Separation Mill                          IX-67
X-l       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Iron-Ore Mines             X-6
X-2       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Iron-Ore Mills Employing
               Physical Methods and/or Chemical Reagents         X-8
X-3       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Lead and Zinc Mines
               Having Solubilization Potential                   X-12
X-4       Parameters Selected and Effluent Limitations
               Recommended for Alkaline Mine Drainage
               BATEA—Bauxite Mines (Acid or Alkaline Mine
               Drainage)                                         X-18
X-5       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Ferroalloy-Ore Mines       X-20
X-6       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Ferroalloy-Ore Mills
               Treating Less Than 5,000 Metric Tons (5,512
               Short Tons) Per Year                              X-22
X-7       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Ferroalloy-Ore Mills
               Treating More Than 5,000 Metric Tons (5,512
               Short Tons) Per Year by Physical Processing       X-22
X-8       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Ferroalloy-Ore Mills
               Using Flotation Process                           X-25
X-9       Parameters Selected and Effluent Limitations
               Recommended for BATEA—Ferroalloy-Ore Mills
               Using Leaching Process                            X-27
X-10      Parameters Selected and Kffluent Limitations
               Recommended for BATKA—Ur.mium Mines              X-JO
X-ll      Parameters Selected and Kf fluent Li mi tat Urns
               Recommended Cor HATKA—Platinum Mi Us             X- 12
XI-1      Parameters Selected and lifflueut Limitations
               Recommended for NSl'S — Ferroalloy-Ore Mines        X I-5
XI-2      Parameters Selected and Effluent Llmlt.it ions
               Recommended for NSPS—Ferroalloy-Ore Mills
               Using Flotation Process                           XT-7
XI-3      Parameters Selected and Effluent Limitations
               Recommended for NSPS—Uranium Mines               XI-8
                                xxix
                               DRAFT

-------
                                   DRAFT
                                 SECTION 1

                                CONCLUSIONS


      To  establish  effluent  limitation guidelines and standards i>f
      performance,  the ore mining and dressing industry was divided
      into 43 separate categories and subcategories for which separ-
      ate limitations were recommended.   This report deals with the
      entire metal-ore mining  and dressing industry and examines
      the industry  by ten major categories:   iron ure;  copper ore;
      lead and zinc ores; gold ore;  silver ore;  bauxite ore;  ferro-
      alloy-metal ores; mercury ores; uranium,  rudltun and vanadium
      ores; and metal ores,  not elsewhere classified (orus of anti-
      mony, beryllium, platinum,  rare earths,  tin,  tJtjn lum,  and
      zirconium).   The subcatcgorlzation of  the  ore c.ilcgnrlc.s Is
      based primarily upon ore mineralogy and  processing or extrdc-
      tion 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 16 of  43 subcategories  for  which  separate  limitations
      are proposed  can be operated with  no discharge of process
      wastewater.   With the  best  available technology economically
      achievable, facilities in 22 of the 43  subcategories can be
      operated with no discharge  of  process wastewater  to navigable
      waters.  No discharge  of process wastewater is also achiev-
      able as a new source performance standard  for facilities in
      22 of the 43  subcategories.

      Examination of the wastewater  treatment methods employed in
      the ore mining and dressing industry indicates tli.it mil ing
      ponds or other types of  sedimentation  impoundments an-  the
      most commonly used methods  of  suspended-sol id remov;i I,  .and
      that these Impoundments  provide the additional  benefit  of
      reduction of dissolved parameters  as well.  Tailing impound-
      ments also servo to equalize- flow  rates and concent r.it inns of
      wastewater parameters.

      It is concluded that,   for areas of  excess water l>.i 1 .nu-e,  the
      practices of runoff diversion,  segregdtiou  of wdsie *troams,
      and reduction in the use of process  water will  assist in  the
      attainment  of no discharge  for  the  speciI led  snlu  .n cgoi Ics.
      Effective chemlcdl -treatment, methods which will result  in
      significant Improvement  in  d ischaryo-wat IT <|n
-------
                                DRAFT
                              SECTION II

                            RECOMMENDATIONS
    The recommended effluent limitation guidelines based on the
    beat 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 16 of  the 43 subcategorles achieve no discharge of process
    wastewater.

    The recommended effluent limitation guidelines based upon the
    best available technology economically achievable (BATEA) are
    summarized in Table IJ-2.  Of the 43 subcategorics listed for
    which separate limitations are recommended, it Is recommended
    that facilities in 22 subcategories achieve no discharge of
    process wastewater by 1983.

    The new source performance standards (NSPS) recommended for
    operations begun after the promulgation of recommended guide-
    lines for the ore mining and dressing industry are summarized
    in Table  II-3.  With the exception of three subcategories, new
    source performance standards are identical to BPCTCA and BATEA
    recommended effluent limitations.
NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
                                   11-1

                                 DRAFT

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

X
IX-1
IX-2
COPPER ORES
Mines
Mills
( Open-Pit, Underground, Stripping
1 Hydrometallurgical (Leaching)
{Vat Leaching
Flotation (Net Evaporation >76.2 em'}
Flotation (Net Evaporation < 76.2 cm'}
X
X
X
IX-3
1X4
LEAD AND ZINC ORES
Mines
( No Solubilization Potential
I SolubHIzation Potential
Mills


IX-5
1X6
IX-7
GOLD ORES
Mines
Mills
! Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation

X
IX-8
IX-9
1X10
IX-11
SILVER ORES
Mines
Mills
{Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation

X
X
IX-12
IX 13
IX 14
BAUXITE ORE
Mines ||
IX 15
   •76.2 cm - 30 In.

NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
                                   II-2

                                  DRAFT

-------
                                  DRAFT


   TABLE 11-1. SUMMARY OF RECOMMENDED BPCTCA EFFLUENT LIMITATIONS BY
             CATEGORY AND SUBCATEGORY -ORES FOR WHICH SEPARATE
             LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills <
< 6.000 metric tonjt/year
> 6.000 metric tons* /year 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 Leeching
Alkaline Leaching

X
X
IX-22

ANTIMONY ORES
Mines
Mills -
Flotation Process

X
1X23

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
RfllfMI
Mills i
Electrostatic/Magnetic end Gravity/Flotation Processes
Physical Processes with Dredge Mining


1X25
IX-26
IX 27
    6.000 metric tons • 6,612 short tons

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

                                   II-3
                                  DRAFT

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

X
X-1
X-2
COPPER OHfcS
Mines
Mills
{Open-Pit, Underground. Stripping
Hydrometallurgical (Laachmgl
{Vat Leaching
Flotation (Net Evaporation > 76.2 cmVyear)
Flotation (Net Evaporation < 76.2 cmVyear)
X
X
X
X
(Same ai BPCTCA)

LEAD AND ZINC ORES
Mines
J No Solubilization Potential
| Solubilization Potential
Mills

X
(SameasBPCTCA)
X-3

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

X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
SILVEH ours
Mines
Mill*
I Flnt.ilitiii Procuss
* Cydiiiclulion I'rniuss
\ Amalgamation PIOIIISS
' Gravity Sepdrnlitui

X
X
X
(SameasBPCTCA)
(Same as BPCTCA)
UAUXITE OHt
Mines
X-4
  •76.2 cm • 30 in

NOTICE: THESE ARE TENTATIVE RECOMMENDATIONS BASED UPON INFORMATION IN THIS REPORT AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVED AND FURTHER INTERNAL REVIEW BY EPA
                                 IT-4

                                DRAFT

-------
                                DRAFT
 TABLE 11-2. SUMMARY OF RECOMMENDED BATEA EFFLUENT LIMITATIONS BY
           CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
           LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
!< 5,000 metric tons'/year
> 5.000 metric tons'/year by Physical Processes
> 5.000 metric tons* /year by Flotation
Leaching


X-5
X-6
X-7
X-8
X-9
MERCURY ORES
Mines
Mills
{Gravity Separation
Flotation Process


X
X
(Same es BPCTCA)

URANIUM. RADIUM. VANADIUM ORES
Mines
Mills
1 Acid or Acid/Alkaline Leaching
| Alkaline Leaching

X
X
X-10

ANTIMONY ORES
Mines
Mills
— Flotation Process


X
(Same as BPCTCA)

BERYLLIUM ORES
Mines
Mills
X
X


PLATINUM ORI S
Mines or
Mum/Milk
1
XII
HAHt L AMI HOMES
Minns
Mills
1 loltifioit 01 LiMitimg

X
X


TITANIUM OKCS
Mines
Mills
j Eluclrusiiilic/M.iflnnli< ,IIH| Gr.ivilv/Floldliuii 1'iticusvs
| Physical I'liicunes willi (JitKlgu MIIIMII)

X
(Same as BPCTCAI
(Same as BPCTCA)
  5.00O miitrir Ions  b.SIV slum Inns
NOTICE THESE ARE TENTATIVE HECOMMtNDATIONS BASED UPON INFUKMATIOM IN THIS REPOR T AND ARE
SUBJECT TO CHANGE BASED UPON COMMENTS RECEIVf.D AND FURTHER INTERNAL REVIEW HY EPA

                                  I l-'i
                                DRAFT

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

X
(Same as BPCTCA)
(Same as BPCTCA)
COPPER ORES
Mines
Mills
J Open-Pit. Underground. Stripping
| Hydrometallurgical (Leaching)
/ Vat Leaching
< Flotation (Net Evaporation > 76.2 cmVyear)
' Flotation (Net Evaporation < 76.2 cm'/year)
X
X
X
X
(Same as BPCTCA)

LEAD AND ZINC ORES
Mines
{No Solubilization Potential
Solubilization Potunlial
Mills

X
(Same as BPCTCA)
(SameasBATEA)

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

X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
SILVER ORES
Mines
Mills
(Flotation Process
Cyanidalion Process
1 Amalgamation PIOCIISI
' Gravity Separation

X
X
X
(Same as BPCTCA)
(Same as BPCTCA)
BAUXITE ORE
Mines
I
(Same as BATEA)
  •76 2 cm - 30 in.


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

                                DRAFT

-------
                                DRAFT
   TABLE 11-3. SUMMARY OF RECOMMENDED NSPS EFFLUENT LIMITATIONS BY
             CATEGORY AND SUBCATEGORY - ORES FOR WHICH SEPARATE
             LIMITATIONS ARE PROPOSED (Sheet 2 of 2)
CATEGORY/SUBCATEGORY
ZERO
DISCHARGE
EFFLUENT
LIMITATIONS
RECOMMENDED
IN TABLE
FERROALLOY ORES
Mines
Mills
!< 5.000 metric tons' /year
> 5.000 metric tons' /year by
> 5,000 metric tons'/year by
Leaching
Physical Processes
Flotation


Xl-l
(Same us BATEA)
(Same as BATEA)
XI-2
(Same as BATEA)
MERCURY ORES
Mines
Mills
{Gravity Separation
Flotation Process


X
X
(Same as BPCTCA)

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

X
X
XI-3

ANTIMONY ORES
Mines
Mills
— Flotation Process


X
(Same as BPCTCA)

BERYLLIUM ORES
Mines
Mills
X
X


PLATINUM ORES
Mines or Mine/Mills
(Same as BATEA)
RARE 1 AHTHORES
Mines
Mills
Flolnlion <» Limrhing

X
X


TITANIUM ORES
Mines
Mills
{Elactrostatic/Miignetir .nicl Clrnviiy/riolnlioii Pinrussiit
Physical Processes with Oimlgo Mining

X
(Some ns BPCTCA)
(Sumo us BPCTCAI
   5.000 mulnc tons - b.512 ihon tout

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

                                  I 1-7
                                DRAFT

-------
                             DRAFT
                          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
 navigable water of the United  States.

 Section 301(b) of the Act requires the achievement, by not
 later than July 1,  1977,  of  effluent  limitations for point
 sources,  other than publicly owned treatment works, which
 are based on the application of the best practicable control
 technology currently available as  defined by the Adminis-
 trator pursuant to Section 304(b)  of  the Act.   Section 301(b)
 also requires the achievement,  by  not later than July 1, 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 deter-
 mined 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 stan-
 dard 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 for the Ore
Mining and Dressing Industry point source category.
                            III-l


                            DRAFT

-------
                                 DRAFT
Section 306 of the Act requires the Administrator, within one year
after a category of sources  is included in a list published pur-
suant 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 Admin-
istrator to promulgate 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 24 source
categories.  Publication of an amended list will constitute announce-
ment 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 Manage-
ment 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 pollutants, the lovol
of pollutant reduction attainable through application of tin? best
practicable control technology currently available, and best avail-
able technology economically achievable.  Standards of perfornuinre
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).


                                 III-2

                                 DRAFT

-------
                            DRAFT
 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 cate-
 gories.   Development  of reasonable Industry categories nnd
 subcategories and establishment of effluent guidelines and
 treatment standards require  a sound  understand Ing and know-
 ledge  of the Ore  Mining and  Dressing Industry,  i he mining
 techniques and milling processes Involved, tin-  mineralogy
 of  the ore deposits,  water use, wastewater generation and
 characteristics,  and  the  capabilities  of existing control
 and treatment technologies.

 Approach

 This report describes the results  obtained from application
 of  the above approach to  the mining  of metals and ore min-
 erals  for the ore mining  and dressing  industry.  The survey
 and extensive sampling and analysis  covered a wide range of
 processes,  products,  and  types of  wastes.  In each SIC cate-
 gory,  slightly different  evaluation  criteria were applied
 initially,  depending  upon the nature of  the extraction pro-
 cesses 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.
 Information  used  as a data base for detailed examination
 of  each category  was obtained from a wide variety of sources
 including  published data  from journals and trade literature,
 mining industry directories, general business publications,
 texts on mining/milling technology, texts on industrial
 wastewater control, summaries of production of  the partIculur
metals of  interest, U.S. Bureau of Mines annual snmniarU-s,
                            III-3

                           DRAFT

-------
                            DRAFT
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 tech-
nology.  Various mining company personnel, independent
researchers, and state and  federal environmental officials
also supplied requested information.

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 wastewater usage and characteristics
for each subcategory as developed in Section IV and discussed
in Section V included:  (1) the source and volume of water
used in the particular process employed and the source of
waste and wastewaters in the plant, and (2) the 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 t lie-
industry and present In measurable quantities won- se I cited
as pollutants subject to effluent limitation guide.! Ines .nul
standards.

Site Visits and Sampling Program

Based upon information gathered as part of the dssembly ol
a data base, examination of NPDES permits and permit appli-
cations, surveys by trade associations, and examination ol
texts, Journals, and the literature available on treatment
practices in the Industry,  selection of mining and milling
                            III—'«


                           DRAFT

-------
                             DRAFT
operations which  were  thought  to  embody exemplary  treatment
practice was  made for  the  purpose of  sampling and  verifica-
tion, and to  supplement  compiled  data.   All factors  poten-
tially  influencing industry  subcategorization were represented
by the  sites  chosen.   Detailed information on production,
water use, wastewater  control,  and water treatment practices
was obtained.  As a result of  the visits,  many subcategorles
which had been tentatively determined were found  to  be  unnec-
essary.  Flow diagrams were  obtained  indicating the  course of
wastewater streams.  Control and  treatment plant design and
detailed cost data were  compiled.

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

Raw waste characteristics  were  then identified for each sub-
category.  This included an  analysis  of all constituents of
wastewaters which might  be expected in  effluents from mining
and milling operations.  In  addition  to examination  of  can-
didate control parameters, a reconnaissance investigation
of some 55 chemical parameters  was performed upon  raw and
treated effluent  for each  site  visited.   Additionally,  limited
sampling of mine  waters  for  several 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 liase

Cost information  contained in  this report  was obtained
directly from industry during  plant visits, from enj; I neei -
ing firms, equipment suppliers, .uul from Liu- 1 11 era t ure .
The lnforin.il ion obtained I ruin  them- sourei-s h.is lu-en n^i-il
tci develop goner.11  capital,  opera l lug, ami  overall  iosl^
for etich 11:0. it men L  mid control  method.   Where d.ila was  l.u-k-
ing, costs were developed  p.irameir n-.i I I v I rom knowledge nl
equipment required, processes employed,  constriu-iIon, and
maintenance requirements.  This generalized cost ilain plus
the specific  information obtained  from  plant visits  was then
used for cost effectiveness  estimates in Section Vlll .uul
wherever else costs nre mentioned  iti  this  report.
                             I I L-"J


                             DRAFT

-------
                            DRAFT
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 environ-
mental 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 treat-
ment 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 con-
trol 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 presently employed In
the industry, but Included also those processes in pilot-
plant or laboratory-research stages.
                            LII-6


                            DRAFT

-------
                             DRAFT
SUMMARY OF ORE-BENEF1CIATION 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 mostly the  mineral to be
recovered or mostly gangue.  The separation of the particles
on the basis of some difference between the ore mineral and
the gangue can then yield a concentrate high in metal value, as
well as waste rock (tailings) containing very little metal.
The separation is never perfect, and the degree of success
which is attained is generally described by two numbers:
(1)  percent recovery and (2) grade of the 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 concen-
trate produced.  Frequently, depending on end use, a particular
grade of concentrate is desired, and specific gangue components
are limited as undesirable impurities.

Many properties are used as the basis for separating valuable
minerals from gangue, including:  specific gravity, conductivity,
magnetic permeability, affinity for certain chemicals, 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-inclusive;
                              III-7


                              DRAFT

-------
                             DRAFT
rather, its  purpose is to discuss Che 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."

Gray it y-Conc en t ra t ion Processes

General.   Gravity-concentration proe-esses exploit di f rereiuA-s
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 partii-le in .1
gravity separator.  Smaller particles of ore mlncrdJ may he-
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*.  lroi  sufficiently
small partle-le.s, eve«n the smallest t urhu le-nce* suspends tin*
p.i rl It-It-  lor long p<>r Joels of liim-, ren-inl le>ss ol lU-nsily.
Sin li slimes, ont-i- lormeil, i-.mnnt. lie-  I n i>vi-i i-il l*v j.'.i.ivllv
Le-chn I c|ii«". .mil  m.iy cause very  low recuve i i e-.  in f.i.-ivilv
professing ol  IllK-'ly frl.iMe* ores, sin li .1:. -,i heel Me  li.ili I inn
I unysl 
-------
                             DRAFT
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 Jlghter 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 concen-
trate 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), whiJe
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.  Repro-
cessing 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,  a relatively recent'
development, provide an efficient means of gravity separation
for large, volumes of material between 0.1 mm and 1 mm (up to
approximately O.OJ  In.) In diameter und have been wldoJy
appiled—particularly, in the processing of heavy sands for
llmenlte (FeTlOJ)  and mon.i7.lte. (a rare-earth phosphate) .
They consist of  a  helical conduit (usually, of five turns)
about a vertical axis.  A slurry of ore is feil to the conduit
                              T1I-9


                             DRAFT

-------
                                 DRAFT
 at the top and flows down the spiral under gravity.   The
 heavy minerals concentrate along the inner edge of the spiral,
 from which they may be withdrawn through a series of ports.
 Wash water may also be added through ports along the inner
 edge to improve the separation efficiency.  A single spiral
 may, typically, be used to process 0.5 to 2.4 metric tons
 (0.55 to 2.64 short tons) of ore per hour; In large  plants,
 as many as several hundred spirals may be run in parallel.

 Sink/Float Separation.   Sink/float 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 are
 milling industry  are suspensions of very fine ferrosilicon or
 galena  (PbS)  particles.  Ferrosilicon particles may  be  used to
 achieve medium specific gravities as high  as  3.5 and are used
 in  "Heavy-Medium  Separation."  Galena, used in the "Huntington-
 Heberlein" process,  allows the achievement  of somewhat  higher
 densities.   The particles are maintained in suspension  by a
modest  amount  of  agitation in the separator and  are  recovered
 for  reuse  by  washing both values and gangue after  separation.

Magnetic  Separation

Magnetic  separation  is  widely  applied  In the  ore milling Industry,
 both for  the  extraction of values from orc>  mid  fur llu-  separation nl"
different  valuable minerals  recovered  from  complex orr^.  Kxl t;nsl ve
use of magnetic, separation Is  inudc  In  I IK-  processing ol  ores <>l  lion,
columbium and  tantalum, and  tungsten,  in mime  a  lew.   The Repartition Is
                                 III-IO


                                DRAFT

-------
                               DRAFT
 based on differences in magnetic permeability (which,  although
 small, is measurable for almost all  materials)  and is  effective
 in handling materials not normally considered magnetic.   The
 basic process involves the transport of ore  through a  region  of
 high magnetic-field gradient.   The most magnetically permeable
 particles are attracted to a moving  surface,  behind which is
 the pole of a large electromagnet, and  are carried by  it  out  of
 the main stream of  ore.  As the surface leaves  the high-field
 region,  the particles drop off—generally, into a  hopper  or onto
 a conveyor leading  to further processing.

 For large-scale applications—particularly,  in  the iron-ore
 industry—large,  rotating drums surrounding  the magnet are
 used.  Although dry separators  are used for  rough  separations,
 these drum separators are most  often run wet  on the slurry
 produced in grinding mills.  Where smaller amounts of material
 are handled,  wet  and crossed-belt  separators  are frequently
 employed.

 Electrostatic Separation

 Electrostatic separation is  used to  separate  minerals on  the
 basis  of their  conductivity.  It is  an  inherently  dry process
 using  very high voltages (typically,  20,000  to  40,000 volts).
 In  a typical  implementation, ore is  charged  to  20,000 to  40,000
 volts, and the  charged  particles are  dropped  onto  a conductive
 rotating drum.  The conductive  particles discharge very rapidly
 and  are  thrown  off  and  collected,  while  the non-conductive
 particles  keep  their  charge  and adhere  by electrostatic attraction.
 They may then be  removed from the  drum  separately.

 Flotation  Processes

 Basically,  flotation  is a  process  whereby particles of one
 mineral  or  group  of minerals are made, by addition of chemicals,
 to adhere  preferentially to  air bubbles.  When  air is forced
 through  a  slurry  of mixed minerals,  then, the rising bubbles
 carry with  them the particles of the mineral(s)  to  be separated
 from the matrix.  If  a  foaming agent  Ls added which prevents
 the  bubbles from  bursting when they  reach the surface,  a  layer
 of mineral-laden  foam  Is buJlt up  at  flie surface of  the flotation
 cell which may be removed to recover  the mineral.   Requirements
 for  the  success of  the  operation are  Lluit particle  size be
 small, that reagents compatible with  the mineral to he recovered
 he u«fd,  and  that water  conditions in the cell noL  Interfere
with attachment of  reagents  to mineral or Lo nlr bubbles.
                              TII-11


                               DRAFT

-------
                              DRAFT
Flotation concentration has become a mainstay of the ore
milling industry.  Because it  is adaptable to very fine
particle sizes  (less than 0.001 cm), it allows high rates of
recovery from slimes, which are inevitably generated in crushing '
and grinding and which are not generally amenable to physical
processing.  As a physico-chemical surface phenomenon, it can
often be made highly specific, allowing production of high-grade
concentrates from very-low-grade ore (e.g., over 95-percent
MoS2. concentrate from 0.3-percent ore).  Its specificity also
allows separation of different ore minerals (e.g., CuS, PbS,
and ZnS), where desired, and operation with minimum reagent
consumption, since reagent interaction is typically only with
the particular materials to be floated or depressed.

Details of the flotation process—exact suite and dosage of
reagents, fineness of grinds, number of regrinds, cleaner-
flotation steps, etc.—differ at each operation where it is
practiced and may often vary with time at a given mill.
A complex system of reagents is generally used, including
four basic types of compounds:  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
specific kinds of particles to the air bubbles, and depressants
prevent it.  Frequently, activators are used to allow flotation
of ore depressed at an earlier stage of the milling process.
In almost all cases, use of each reagent in the mill is low
(generally, less than 0.5 kg—approximately 1 Ib—per ton of
ore processed) , and the bulk of the reagent adheres to tailings
or concentrates.

Sulfide minerals are all readily recovered by flotation using
similar reagents in small doses, although reagent requirements
and ease of flotation do vary  throughout the class.  Sulfide
flotation is most often carried out at alkaline pH.  Collectors
are most often alkaline xanthates having two to five carbon
atoms—for example, sodium ethyl xanthate (NaS2COC2H5).
Frothers are generally organics with a soluble hydroxyl group
and a "non-wettable" hydrocarbon.  Pine oil (C6H120H), for
example, is widely used to allow separate recovery of metal
values from mixed-sulfide ores.  Sodium cyanide is widely
used as a pyrite depressant.  Activators useful in sulfide-ore
                               111-12

                               DRAFT

-------
                             DRAFT
flotation may include cuprous sulfide and sodium sulfide.
Sulfide minerals of copper, lead, zinc, molybdenum, silver,
nickel, and cobalt are commonly recovered by flotation.

Many minerals in addition to sulfides may be, and often are,
recovered by flotation.  Oxidized ores of iron, copper,
manganese, the rare earths, tungsten, titanium, and columbium
and tantalum, for example, may be processed in this way.
Flotation of these ores Involves a very different suite of
reagents from sulfide flotation and has, in some cases, required
substantially larger dosages.  Experience has shown these
flotation processes to be, in general, somewhat more sensitive
to feed-water conditions than sulfide floats; consequently,
they 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 beneficlated 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 anlonic 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 Jcaching
           In water.

      (3)  Uranium ores are only mildly solnliio In water, hut
           they dissolve quickly Jn ai-id or .ilk.-illne solutions.

           Native gold which Is found Ln a Iliu-ly divided
           Is soluble in mercury and can hi- t-xlrdited by
                             TII-13


                             DRAFT

-------
                                 DRAFT
            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.

 Leaching solutions can be categorized as  strong,  general
 solvents (e.g.,  acids) and weaker, specific solvents (e.g., cyanide)
 The acids dissolve any metals present, which often include
 gangue constituents (e.g., calcium from  limestone).  They are
 convenient to use, since the ore does not have to be 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 Le.icli  In n  v.irieLy of w;iys.  In  v.iL
 leaching,  the process  Is  carried otiL  In .1 eontalner (v.it),
 often  equipped with  facilities for .1^1 Kit Ion,  he;Uln>',,  .if r;iL inn.
 and presaurizat Ion  (e.g.,  P.ieliuc .1  (.inks).  .i"----jj_iti le.uhlnn
 takes  place in the ore body,  with  Uu- lejeli .ipplloil either  by
 plumbing  or by percolation  through overburden.   The pregnant
 leach  solution 1s  pumped  to  the recovery facility ;incl c.in oil en
 be  recycled.  In-situ  leaching Is  most  economic-iij  when  Lhe  ore
body is surrounded by  an  impervious nuiLrlx.  When water  sufflees
                                11 I - I /•


                                DRAFT

-------
                              DRAFT
 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, and  uranium, with sulfuric
 acid leach,  are recovered  in this fashion.

 Amalgamation.    Amalgamation is the process by which mercury
 is alloyed with some other metal to produce an amalgam.  This
 process is applicable to free milling  of 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 highly 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.  Js 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
                              111-15


                              DRAFT

-------
                              DRAFT
 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 hy the
 solution according to the reaction:

       4Au + SNaCN + 2H20   	>   4NaAu(CN)2_ + 4NaOH

 and  subsequently sorbed onto activiated carbon ("Carbon-.ln-1'u I p"
 process) or precipitated with metallic zinc according to the-
 react ion:

 NaAu(CN)^ + NaCN + Zn + 2H20   	>   NaZn(CN)^ + Au + ftf  + NaOH

 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  solu-
 tion is  the Carbon-in-Pulp process.   This process was developed
 to provide economic  recovery of gold from low-grade  ores or
 slimes.   In this process,  gold which has  been  solubilized  with
 cyanide  is brought  into contact with 6  x  16 mesh  activated
 coconut  charcoal in  a series of tanks.  The pulp  and  enriched
 carbon are air lifted and  discharged on small  vibrating  screens
 between  tanks, where the carbon is separated and  moved to  the
 next adsorption  tank,  counter-current to  the pulp flow.  Gold-
 enriched carbon  from the last  adsorption  tank  Is  leached wjth
 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  nri-  deposited
onto stainless sU:i-l  wool  cathodes.

 l»rotreatment of  ores  containing only llnt-ly divided gold ,nul
sllvcir usually  Includes  multistage crushing, I Iiu-  grind Lug,  .uul
classification of  the  ore  pulp into  SHIR!  
-------
                             DRAFT
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 counter-
current decantation processing.  Gold or silver is then
recovered from the pregnant leach solutions by the methods
discussed above.

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

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

The cyanidation process is comparatively simple, and is applicable
to many types of gold/silver ore, but efficient low-cost
dissolution and recovery of the gold and silver are possible onJy
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 chlorine treatment.

      (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 aL a level that wiJl minimize
           consumption of ryanJde by the dissolution of oi hor
           meta] •+•". Ur Inn mlner.ils.
                              111-17


                              DRAFT

-------
                             DRAFT
       (A)  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 die  same
principle:  Polar organic molecules tend to exchange  a
mobile ion in their structure — typically, C1-,  NCtt-,  HSCM-, or
C03. —  (anions)  or H+ or Na+ (cations) — for  an ion with a
greater charge  or a smaller ionic radius.   For  example,  let
R be the remainder  of the polar molecule (in the case, of a
solvent) or polymer (for  a resin),  and  let  X be the mobile
ion.  Then, the exchange  reaction for the example of  the
uranyltrisulfate complex  is:

      4RX + (UO£(S04J3_) ----  ~> R4U02^S04)3_ +  4X-
Thls 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-l-H- > Mg-H- >  Na+

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

                     92U > 42Mo > 23V

and the separation of hexavalent  92U cations  by  Lon 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) a J most always appear  In aqueous  solutions us
oxidized Lons (uranyl,  vanadyl , or molylulale  radicals), wll.li
uranium and vanadium additionally cuinplcxc.il with  anlonlc radicals
to form trisul fates  or  trlcarbonates In Lhc leach.  The com-
plexes react an Ionic-ally, and  the affinity n| exchange, resins
                             111-18


                             DRAFT

-------
                             DRAFT
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 pK 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, solvent-
extraction 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.  Ion-
exchange and solvent-extraction circuits can be combined to take
advantage of the slime resistance of resin-in-pulp ion exchange
and of the separatory efficiency of solvent extraction (Eluex
process).
                            IIL-19


                             DRAFT

-------
                              DRAFT
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 arid 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 pro-
cessing 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% (Refer-
ence 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 break-
down of crude iron-ore production in the U.S. is shown in
Table III-2.  A breakdown of U.S. iron-ore shipments by pro-
ducing company is given in Supplement B to this document.
Except for a very small tonnage, iron ores are beneficiated
before shipping.

Beneficlation of iron ore includes such operations as crush-
ing, screening, blending, grinding, concentrating, classify-
ing, and agglomerating and is most 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 and to recover more of the secondary minerals in an ore
deposit.  As with many other natural resources, future avail-
ability will largely be a matter of cost rather than of abso-
lute depletion as these lower-grade ores are utilized.  Bene-
fication methods have been developed to upgrade 20-30% iron
'taconite' ores into high-grade materials.

In most cases, open-pit mining is more economical than con-
ventional underground methods.  It provides the lowest cost
operation and is employed whenever the ratio of overburden
(either consolidated or unconsolidated) to ore does not
                             111-20


                             DRAFT

-------
                                 DRAFT
         TABLE 111-1. IRON-ORE SHIPMENTS FOR UNITED STATES
                       a. QUANTITIES SHIPPED BY REGION
REGION
Great Lakes
NortfiBBStB rn
Southern
•Hf--»-— —
YYMiarn
TOTAL U.S.

REGION

Groat Lakes
Northeastern
Southern
UUAM*IUM
VW5I6fn
TOTAL U.S.
AMOUNT SHIPPED
1968
METRIC TONS
66.093,239
3.602.706
3.474.203
10.666.860
82.736.806
LONG TONS
64,065.185
3.645.805
3.419.333
10.399.872
81.430.195
1969
METRIC TONS
72.S34.630
3.453.486
4.733.087
10.464.364
91.176.567
LONG TONS
71,389.050
3.398.943
4.658.335
10.289.252
89.735.580
1970
METRIC TONS
70.180.666
3.043357
5.022.36B
10.644.782
88.791.674
LONG TONS
69,072.263
2,995,784
4.943.048
10.378.242
87.389.337

AMOUNT SHIPPED
1971
METRIC TONS
62.766.873
2.B59.973
4.240,720
8,253,243
78.120,810
LONG TONS
61.775.561
2.814304
4.173.744
8.122,895
76.887,004
1972
METRIC TONS
65.769.357
2.362.067
4.032,651
7,397,815
79.537,162
LONG TONS
64.720.783
2.324,762
3.966.961
7,266,471
78,280.977
1873
METRIC TONS
77.604365
2,405,456
3.923,518
8.462,679
92.296,418
LONG TONS
76.280,787
2.367,465
3,861 ,562
8,328,925
90,838.729
    b. SHIPMENTS FROM GREAT LAKES REGION AS PERCENTAGES OF TOTAL US. SHIPMENTS
YEAH
1968
1869
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
73.6
GREAT LAKES AGGLOMERATES
AS PERCENTAGE OF TOTAL
U.S. SHIPMENTS
48.7
60.6
62.3
66.3
61.8
61.7
                    c. PERCENTAGES OF TOTAL U.S. SHIPMENTS

CATEGORY
Direct Shipping
CoaneOres
Fine Ores
Screened Ores
Concentrates
Agglomeratei

YEAR
1968
8.2


3.2
283
60.3
100.0
1968
7.0


3.1
27.6
62.4
100.0
1670
6.0


27
28.2
64.1
100.0
1971
4.3


3.1
23.7
689
100.0
1972
2.0
12.8
11.9


73.3
100.0
1973
2.4
129
129


71.8
100.0
SOURCE: Reference 1
                                 111-21
                                 DRAFT

-------
                                 DRAFT
           TABLE II1-2. CRUDE IRON-ORE PRODUCTION FOR U.S.
                          a. QUANTITIES PRODUCED
YEAR
1968
1969
1970
1971
1972
1973
PRODUCTION BY REGION
GREAT LAKES
METRIC TONS
169.349.027
169.328.625
172.799.698
161.947.509
1 68,183.907
186,627.840
LONG TONS
166.832.339
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
6.916.338
LONG TONS
10.075.038
9.575.01 1
9.028.913
7.651.428
6.615.513
6.806.120
SOUTHERN
METRIC TONS
7.743.542
9.135.961
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
207463.916
212.498.366
197558,696
187.686.069
220,263,451
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
6.0
9.4
1000
1971
82.0
3.9
4.8
93
100.0
1972
84.3
3.6
6.0
7.1
1000
1973
84.7
32
39
82
1000
SOURCE. Reference 1
                                111-22
                                DRAFT

-------
                             DRAFT
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 ores, it may be as high as 6
or 7 to 1; in the case of taconite, a stripping ratio of less
than 1/2 to 1 may become necessary.  Stripping the overburden
necessitates continually cutting back the pit walls to permit
deepening of the mine to recover ore in the bottom.  Power
shovels, draglines, power scrapers, hydraulicking, and hydraulic
dredging are used to recover ore deposits.  Drilling and blast-
ing may be 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 con-
veyors, skip hoists, or a combination of these.  It is then
transferred to a loading dock for hauling to a crushing plant
for size reduction, to a screening plant for sizing, or to a
concentrating plant for treatment by washing (wet size classi-
fication and tailings rejection) or by gravity separation.

Special problems are associated with the mining of taconite.
The extreme hardness of the ore necessitates additional drill-
ing/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 tech-
niques 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 work-
Ings,  and elaborate pumping facilities usually requires expen-
sive capital Investments.  Production in terms of iron ore/day
is much lower than in the case of open pit production, neces-
sitating the presence of very high grade ores for economic
recovery.   The health, and safety of operators and the ecological
Impacts of underground mining operations are of increasing concern
to the public, while pressures from the public sector have recently
been forcing Important changes in mine operations and planning.
                            111-23

                            DRAFT

-------
                             DRAFT
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.

Crushing and screening reduce the size of crude ore not
requiring processing  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 prod-
uct 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 attracted
to ores  conditioned with flotation reagents separate out iron
ore in the froth during the  flotation process, while magnetic
separation techniques are used  where ores containing magnetite
are encountered.

At the present time, there are  only three chemical process-
ing 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 II1-4.  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.  Initially higher grade ores
containing hematite, on the  other hand, while containing more
iron cannot be upgraded much above 55% iron.

Agglomerating processes follow  concentration operations and
serve to Increase permeability  of furnace feed and reduce
"fines"  which normally would be lost in the flue gases.
Sintering, pelletizing, briquetting, and nodullzlng 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, with a reduction Ln impuri-
ties.  Pelletizing  involves  the formation of pellets or ballh
of iron  ore fines,  followed  by  heating.  (Figure IIT-4 illus-
trates a typical palletizing operation.)  Nodules or lumps are
                              111-24

                             DRAFT

-------
                   DRAFT
    Figure 111-1. BENEFICIATION OF IRON ORES


ORE
CRUSHING AND
SCREENING
*
BLENDING
t
CONCENTRATING PROCESSES:
PHYSICAL
* i
WASHING' JIGGIN


1
| SINTERING 1
* *
r MAGNETIC VJfpn
G SEPARATION SEpM*°
*
AGGLOMERATION
PROCESSES

PELLETIZING NODULIZING


CHEMICAL
*
'Y-
A FLOTATION
TION
*

1
BRIQUETTING
t
T
          TO STOCK PILE AND/OR SHIPPING
                   111-25






                   DRAFT

-------
                      DRAFT
Figure 111-2.   IRON-ORE FLOTATION-CIRCUIT FLOWSHEET
               DENSIFYING THICKENER
                   UNDERFLOW
                            CONDITIONERS
                                 I
               ROUGHER FLOTATION
                       ROUGHER
                         TAIL
                       TO
                       TAILING
                       BASIN
          ROUGHER
        CONCENTRATE
          (10 CELLS)
              1
           FROTH OF
         'FIRST 2 CELLS
              I
               CLEANER FLOTATION
      CLEANER
        TAIL
      	I
  CLEANER
CONCENTRATE
  (8 CELLS)
  FROTH OF
FIRST 2 CELLS
                      1
              RECLEANER FLOTATION
           RECLEANER
              TAIL
         RECLEANER
        CONCENTRATE
          (7 CELLS)
             I
                                               TOTAL
                                              FLOTATION
                                            CONCENTRATE
                                                 J
                                         TO AGGLOMERATION
                                            (FIGURE 111-4)
                      111-26


                      DRAFT

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


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

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


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

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


3.   Cetionlc Flotation of Hematite (from crude ore)

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


4.   Cationic Flotation of Silica (from crude ore)

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


5.   Cationic Flotation of Silica (from magnetite concentrate)

     Amine: 5 g/metric ton (0.01 Ib/short ton)
     Methylisobutyl carbinol: as required
                                           111-27


                                          DRAFT

-------
                                     DRAFT
TABLE 111-4. VARIOUS FLOTATION METHODS AVAILABLE FOR PRODUCTION
             OF HIGH-GRADE IRON-ORE CONCENTRATES
             1.   Anionic flotation of specular hematites


             2.   Upgrading of natural magnetite concentrated by cationic flotation


             3.   Upgrading of artificial magnetite concentrates by cationic flotation


             4.   Cationic flotation of crude magnetites


             5.   Anionic flotation of silica from natural hematites


             6.   Cationic flotation of silica from non-magnetic iron formations
                                     III-28


                                     DRAFT

-------
                            DRAFT
  Figure 111-3. MAGNETIC TACONITE BENEFICIATION FLOWSHEET
                    CRUSHED CRUDE ORE
                         T  _
                        ROD MILL |
               COBBER MAGNETIC SEPARATION
            CONCENTRATE
                 i
            | BALL MILL |
                 V
     CLEANER MAGNETIC SEPARATION
            CONCENTRATE

                 I
           HYDROCYCLONE
        OVERSIZE   UNDERSIZE
                HYDROSEPARATOR
            CONCENTRATE
     FINISHER MAGNETIC SEPARATION
CONCENTRATE
 THICKENING
     I
     LTI
     T
TO PELLETIZING
 (FIGURE MM)
                                             TAILING
                                             TAILING
                                             TAILING
                                             TAILING
                                                 TO TAILING BASIN
                            111-29


                            DRAFT

-------
                   DRAFT
  Figure 111-4. AGGLOMERATION FLOWSHEET
CONCENTRATE FILTER CAKE
                                 BENTONITE
                 BALLING DRUM
                      1
                    SCREEN
                            I
              UNDERSIZE   OVERSIZE

                            I
                           FUEL
                           I
                  AGGLOMERATION FURNACE
                 PELLETS        EXHAUST GASES

                    t                *
              TO STOCK PILE      TO ATMOSPHERE
              AND/OR SHIPPING
                   III-JO



                   DRAFT

-------
                                 DRAFT
 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 grind-
 ing 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 beneficiating  the  ores  and
 wastes by hydrometallurgical and/or physical-chemical pro-
 cesses.   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
               Nevada
               MLchlgan
               Tennessee
A series of tables follow which give statistics for the U.S.
copper industry.  Table III-5 lists total copper mine produc-
tion of ore by year, and Table III-6 gives copper ore produc-
tion by state for 1972.  The average copper content of domestic

                                 111-31


                                 DRAFT

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

                                          |y:::j::| MINING AND MILLING COPPER AS A COPRODUCT
                                                                                                                   O
                                                                                                                   33

-------
                        DRAFT
  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 111-6. 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
                        III-'JJ





                        DRAFT

-------
                                 DRAFT
ores  is  given by Table III-7.   The average concentration of
copper recovered from domestic  ores,  classified by  extraction
process,  is  listed in Table  III-8.  Copper concentrate produc-
tion  by  froth flotation is given in Table  III-9, while pro-
duction  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 remain-
der 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.   Electrowon copper amounted to 13.5%.

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 beneficiation.  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 moly-
bdenum byproduct ore,  18,824 metric tons (20,750 short tons)
of  byproduct  molybdenum were produced.

Processes Employed to  Extract Copper  from  Ore.   The mining
processes 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.

Beneficiation  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.   Hydromotallurgica.1 pro-
cesses currently employ  sulfurlc  acid (5-10%)  or iron su]fate
to dissolve  copper from  the  oxldlc or mixed oxldlc-sulfidic
                                 IIL-34


                                 DRAFT

-------
                          DRAFT
     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 111-8. AVERAGE CONCENTRATION OF COPPER IN DOMESTIC
          BY PROCESS (1972)
ORES
STATE
ARIZONA
UTAH
NEW MEXICO
MONTANA
NEVADA
MICHIGAN
IDAHO
TENNESSEE"
COLORADO
ALL OTHER
TOTAL U.S.
CONCENTRATION (%)
FLOTATION*
0.51
0.58
0.70
0.55
0.54
0.82
-
0.64
-
1.35
0.55
DUMP/HEAP
LEACH
0.47
1.10
-
-
0.38
N/A
-
N/A
-
-
0.47
DIRECT SMELTER
FEED
1.94
—
0.07f
4.06
0.68
-
2.65
-
10.24
2.30
1.68
    • INCLUDES FROTH FLOTATION AND LEACH-REDUCTION/FLOTATION
    •• FROM COPPER/ZINC ORE
    t JUST AS A FLUXING MATERIAL
      SOURCE: REFERENCE 2
                          111-35
                          DRAFT

-------
                         DRAFT
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
                         TII - 10


                         DRAFT

-------
                           DRAFT
Figure 111-6. GENERAL OUTLINE OF METHODS FOR TYPICAL RECOVERY OF
         COPPER FROM ORE
ORE KOIHCul

WASTE
DUMP "«
LEACH ««
IACIDI IAC
ACID ACID
SOLUTION SOL'N
ACID
RECVCLED
1 1
PRECIPI
PLA

ORE 10 144% Cu) ORE ^ OVERBURDEN AND
ORE -'»""« WASTE DUMPS

1 i
AP MMTJJ CRUSHER *j SCREENING
101 IACIDI MINED I

i ORE CRUSHER
ACID 1 1
Ml'M «
RECVCLED RECVCLED J SCREENING
I 	 |
[y,""1 HEAP _^. TERTIARY
HT* [^ CRUSHER MIXED
(, ACID i &ULFID
ACID RECYCLED 'OLUTION [_j ^^
SPONGE IRON
CEMENT 1 "•"•".'"..""" COPPER AND
rnPVFR 1 KUAHI 100N
n
TOSM
1
REFIf

ELTER TAII.IMO 1 , TmrnriiirnT ^. FLOTATION ^ TA
POND P^ THICKENERS ^ CEtLS ••
* 1 t
,ERV RECYCLED WATER «IIHI*»IH»IB
THICKENERS 'j'^
REFINERY
TO SMtLTER
CEMENT
COPPER
t
PRECIPITATION
PLANT

OXIDE/
E ORE


WASH
WATER
VAT LEACH
IACIDI
LS '
BARREN
PF
SI
l
ELECTRO
WINNING
FACIIITV
ARD '"
OW
T SULFIDE
| FILTER | 4
1 TO DUMP
BYPRODUCT COPPEHISI
MOLYBDENUM CONCENTRATE
t TO SMELTER
TO SMELTER |
Y REFINERY
MARKET
t

CATHOOI
COPPER

EGNANT
3LUTION



1UMAKKLI
ion HI IINERVI
                           111-37
                           DRAFT

-------
                            DRAFT
 ores in dumps, heaps, vats or in-situ (Table 111-10).   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 extrac-
 tion-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,  tel-
 lurium, and iron.   These are removed in the refinery.

 A 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.  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 electro-
 winning to  recover copper dissolved from oxide  ores by  acid.

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

 Lead  and Zinc Ores

 Lead  and  zinc mines and mills  in the U.S.  range in age  from
 over  one  hundred years to essentially new.   The size of
 these operations ranges from several hundred metric tons
 of  ore  per  day to  complexes capable of moving about ten
 thousand  metric tons  of ore per  day.   Lead  and  zinc ores
 are produced almost exclusively  from underground  mines.
 There are some deposits wlilcli  are amenable  to open pit  oper-
 ations; a number of minus  during Llielr early opening stages
 of  operation are started «ns open-pit  mines  and  then developed
 into  underground mines.   At  present.,  only one small open-pit
mine  is  in  operation,  nnd  Its  useful  life  Is estimated  in
months.   Therefore,  for all  practical  purposes, jll miniiiR
                            II1-J8
                            DRAFT

-------
                         DRAFT
TABLE 111-10. HEAP OR VAT ORE 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
                          111-39





                         DRAFT

-------
0
3J
     --
     o
                         Figure 111-7.  MAJOR COPPER AREAS EMPLOYING ACID LEACHING IN HEAPS.
                                    IN DUMPS, OR IN SITU
O
3J
                                                 ma LEACHING ZONES

-------
                                   DRAFT
          TABLE 111-11. AVERAGE PRICE RECEIVED FROM COPPER
                        IN THE UNITED STATES
   YEAR
                           PRICE IN CENTS PER KILOGRAM (CENTS PER POUND)
   LAKE COPPER*
ELECTROLYTIC COPPERt
1865-1874
   1907
   1910
   1916
   1917
   1920
   1926
   1930
   1932
   1936
   1940
   1946
   1960
   1966
   1960
   1966
   1970
   1972
   1973
     60.94 (27.70)
     46.86 (21.30)
     28.86 (13.12)
     38.81 (17.64)
     64.20 (29.18)
     39.62 (18.01)
     31.77 (1444)
     29.48 (1340)
     13.00 ( 6.91)
     19.62 ( 8.92)
     26.66(11.66)
     26.40 (12.00)
 40.96- 64.16(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(63.00-60.00)
109.7 -114.7(49.88-62.13)
110.3 -169.2(60.13-72.38)
  42.90-  63.90(19.60-24.60)
  69.30-  94.60(31.50-43.00)
  66.00 •      (30.00)
  77.00-  81.40(35.00-37.00)
 116.9  -132.3(63.12-60.12)
 111.4  -116.8(60.63-62.63)
 116.9  -151.1 (63.13-68.70)
  • 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
                                 111-41

                                   DRAFT

-------
                    DRAFT
 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
                     111-42
                    DRAFT

-------
                              DRAFT
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 conver-
sion 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
sulflde).  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 zlncite (zinc oxide), zinc willemite
(silicate), and franklinite (an iron, zinc, manganese oxide
complex).  Sphalerite is often found in association with sul-
fides of iron and lead.  Other elements often found in associa-
tion 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 pre-
liminary 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 jreu
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,

                              111-43


                              DRAFT

-------
                                   DRAFT
        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.-OEC.
METRIC TONS
S
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
251371
51
1.078
1.331
5,674
1.359
443
596
486
300.163
1.658
SHORT IONS

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.
                                   lit-44


                                  DRAFT

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


STATE
Arizona
California
Colorado
Idaho
Illinois
Kantueky
Main*
Missouri
Montana
New Jartay
New Mexico
New York
Pennsylvania
Tennessee
Utah
Virginia
Washington
Wisconsin


1973

RANK


4
5


7
1

B

2

3
9
8


%


11.94
9.65


4.13
17.27

6.94

17.4

13.32
3.48
331


Total
Daily average*
1973
JAN.-OEC. TOTALS
METRIC TONS
7.638
16
51.533
41.216
4323
245
17343
74376
379
29,955
11.147
73361
17.104
57.474
15323
15.131
5,768
7365
431399
1.183
SHORT TONS
8,421
18
66317
45,442
5.318
270
19.672
82,223
418
33327
12,290
81,435
18358
63,367
16364
16.682
6.359
8.672
475.853
1.304
1974
JAN. TOTALS
METRIC TONS
600
-
3.961
3,279
224
-
1,238
6389
82
2.361
863
6361
1,575
7,239
1,130
1.281
528
733
38.644
1,246
SHORT TONS
662
-
4,367
3,615
247
...
1.365
7.266
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 ad|ustment for Sundays or holidays.
                                 111-45
                                  DRAFT

-------
                              DRAFT
 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 mineral-
 containing  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 hoppers.  Primary
 or rough crushing equipment  is often  operated underground.
 The drilling and  transportation equipment is, of course,
 highly mechanized and  employs diesel  power.. At some locations,
 the equipment  is  maintained in underground shops, constructed
 in tnined-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 loca-
 tions, water enters with sand or tailings used in hydraulic
 backfill operations.

 The water is pumped from the mine at  a rate necessary Lo
maintain operations in  the mine.  The amount of water puinpcJ
 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 miJJIon liters (JO to 40
million gallons)  per day.   In many rases, llirre Is 
-------
                                      DRAFT
     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
 CONCENTRATE
  FINAL LEAD
 CONCENTRATE
  THICKENING
AND FILTRATION
         USUALLY
         RECYCLED
        TO PROCESS
       WATER SYSTEM

1
TAILING ZINC "OUGHER
Wl> CONCENTRATE
i *
TAILING
THICKENER

ZINC CLEANER
FLOTATION
1
TO TAILING FINAL ZINC
DAM CONCENTRATE

r
o ,
VCLE
t
~1
THICKENING
AND FILTRATION
' EFFLUENT 4
                                                         WATER
                                                         DISSOLVED SOLIDS
                                                         SUSPENDED SOLIDS
                                                         EXCESS REAGENTS
                                                                              I
                    TO SUBSURFACE
                      DRAINAGE
CONCENTRATE
  TO ZINC
  SMELTER
USUALLY RECYCLED
   TO PROCESS
  WATER SYSTEM
                                      111-47
                                      DRAFT

-------
                              DRAFT
 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 precon-
 centrated using mechanical devices based on specific gravity
 principles.   The ore or preconcentrate is initially crushed
 to a size suitable for introduction into fine grinding equip-
 ment, such as rod mills and ball mills.   These mills run wet
 and are usually run in circuit with rake or cyclone classi-
 fers 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  (CaOH^)  are  common.   Other  reagents
 commonly used and  their  purposes are:

          Reagent                        Purpose

Methyl  Isobutyl-carbinol                 Frother
 Propylene Glycol Methyl  Ether            Frother
Long-Chain Aliphatic Alcohols            Frother
Pine Oil                                 Frother
Potassium Amyl Xanthate                  Collector
Sodium Isopropol Xanthate                Collector
Sodium Ethyl  Xantliate                    Collector
Oixanthogen                              Collector

                             TII-48


                              DRAFT

-------
                              DRAFT
          Reagent                       Purpose

Isopropyl Ethyl Thionocarbonate         Frother
Sodium Dlethyl-dlthiophosphate          Frother
Zinc Sulfate                            Zinc Depressant
Sodium Cyanide                          Zinc Depressant
Copper Sulfate                          Zinc Actlvant
Sodium Dichrornate                       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 hydropho-
blc (Incapable of dissolution In water) 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.

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 waste stream from a lead/zinc flotation ml]1
contains the residual solids from the original ore which
have been finely ground to allow mineral recovery.  The
                             111-49


                             DRAFT

-------
                                  DRAFT
stream also  contains dissolved solids and excess  mill  rea-
gents.   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 indus-
try is concentrated  in eight states:  Alaska, Montana, New
Mexico, Arizona,  Utah,  Colorado,  Nevada, and South Dzikota.
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 cither underground or
open-pit methods,  the  particular method  chosen depending on
such factors as size and shape of the deposit, ore grade,
physical and mineralogical character of  the ore and surround-
ing rock, and depth  of the deposit.

Milling Practices.   Milling pr.icliccs for Lhc recovery anil
benef id at ion of  gold  and goLd-i/nuld i" ing ores .iro ry.inul.i-
tion, amalgamation,  flotntlun,  ;mn
of ore mined from lode dcpusl Ls.  I'l.urr u|>ur;iL Inns, lu'wi'vi-i .
employ only  gravity  methods, SOIIK-I hw. In con juncL Inn  with
amalgamation.
                                 I 1 I-SO
                                  DRAFT

-------
                                 DRAFT
Prior to 1970, amalgamation was the process used to recover
nearly 1/4 of the gold produced domestically.  Since that
time, environmental concerns have caused restricted use of
mercury.  As a result, the percent of gold produced which
was recovered by the amalgamation process dropped from 20.3%
in 1970 to 0.3% 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 domes-
tically was recovered by cyanidation, and this increased to
54.6% 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 gold/silver-bearing gravels are beneficiated
by gravity methods, and, in the past, the precious metal-
bearing sands generally were batch amalgamated in barrel
amalgamators.  However, amalgamation in specially designed
sluice boxes was also practiced.

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

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 agita-
tion 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 cyanide solution
is collected by decanting from the thickeners.

Cyanidation of slimes generated  In the course of wee  grinding
is currently being done by a recently developed process,
carbon-in-pulp  (Figure  111-9).  The slimes are mixed  with a
cyanide solution in large tanks, and  Lhe solubilized  gold
cyanide is collected  by adsorption onto activated charcoal.
Gold is stripped from the charcoal using a small volume  of
                                  111-51


                                  DRAFT

-------
                            DRAFT
Figure 111-9. CYANIOATION OF GOLD ORE: VAT LEACHING OF SANDS
          AND 'CARBON-IN-PULP' PROCESSING OF SLIMES
                                            O HI IINIIIV
     lOSMIimt
                           II 1-51'
                           DRAFT

-------
                            DRAFT
Figure 111-10. CYANIDATION OF GOLD ORE: AGITATION/LEACH PROCESS
      ORE
CRUSHING





   GRINDING
      I
 CONDITIONING
      I
COUNTERCURRENT
  LEACHING IN
  THICKENERS
 PRECIPITATION
 OF GOLD FROM
 LEACHATE WITH
   ZINC DUST
       *
 COLLECTION OF
 PRECIPITATE IN
  FILTER PRESS
      I
  PRECIPITATE
  FILTERED AND
   THICKENED
       1
  TO SMELTER
   REAGENTS (CN)
BARREN
 PULP
TAILING
 POND
TAILING-POND
  DECANT
 RECYCLED
      BARREN SOLUTION
         RECYCLED
                              111-53


                             DRAFT

-------
                                  DRAFT
 hot  caustic;  an  electrowinning process  is  used  for  final
 recovery  of  the  gold in Lhe  mill.   Bullion is subsequently
 produced  at  a refinery.

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

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

 Silver Ores

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

 Current domestic  production  of new silver  is derived almost
 entirely  from exploitation of  low-grade  and complex primary
 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
                                 111-54


                                 DRAFT

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





GRINDING
L
I


CONDITIONING
*

SELECTIVE
FROTH
FLOTATION
L
t


CONCENTRATE
FILTERED
AND THICKENED
I
TOSME
LTER




BARREN
SOLUTION
RECYCLED





WATER

I1EAACMTC






FLOTATION CIRCUIT
TAILINGS
J
\



LEACHING IN
THICKNERS
j
PRECIPITATION OF
GOLD FROM LEACHATE
WITH ZINC DUST
1

COLLECTION OF
PRECIPITATE IN
FILTER PRESS

\


PRECIPITATE FILTERED
AND THICKENED

BARREN ^ TO TAILING
PULP POND
                             \
                          TO SMELTER
                             111-55
                            DRAFT

-------
                               DRAFT
silver is produced are listed  in Table 111-15.

Present extractive metallurgy  of silver was  developed  over
a period of more  than  100  years.   Initially,  silver, as  the
major product, was recovered from rich, yet  simply  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  liber-
ated—gold/silver ores, processed by amalgamation and  cyani-
dation, now contribute only 1  percent of the domestic  silver
produced.  Primary sulfide ores, processed by flotation  and
smelting, account for  99 percent (Table 111-16).

Selective froth  flotation  processing can effectively and effi
ciently beneficiate almost any type and grade of  sulfide ore.
This process employs various well-developed  reagent combina-
tions 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, promotors,  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)J.2Sb4S_3.   A tetrahedrite concentrate contains approxi-
mately 25 to 32%  copper in addition to the 25.72 to 44.58
kilograms pi-r mi-trie  Inn  (750  Lo 1300 troy ounce per  ton)
of silver.  A  low-j;r.idr (i.4'5  k« per metric  ton;  J 00  trny
o* per Lou) .'>M v«T/pvr He  i nncenLrntii Is prndufiid .it one mill.
Antimony ni.iy cuiiipr i so  up  to I H'X. nl  the Li-Lralied r i t c conci-ii-
Lrate and ni.iy or  m.iv  iu»t hi- exl rncLud prlur  lo  shipment  to
j smelte.r.

Various oilier s U viir-eunui Ininj; minerals are re-covered MS
byproducts ol primary  copper,  load, and/or zinc uporiiL ions.
Wherrj Lhis occur*!., Lhf usu-il pracllce is to  ultimately rm-ovi-
tin? silver Irnm  Lht- bfisc-muLJl flotation concentrates  at Lhe
smelter or rn finery.
                              r I l-.r»f*


                              DRAFT

-------
                                 DRAFT
   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
                                 111-57

                                 DRAFT

-------
                         DRAFT
    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


YEAH
1968
1969
1970
1971
1972
SILVER BULLION AND PRECIPITATES RECOVERABLE BY
AM ALG AM AT ION
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

SILVER RECOVERABLE FROM ALL SOURCES (%)
AMALGAMATION
028
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
tLess than 1/2 unit


SOURCE:  REFERENCE 2
                           I I »-'>»

                          DRAFT

-------
                              DRAFT
Figure 111-12. RECOVERY OF SILVER SULFIDE ORE BY FROTH FLOTATION
                                ORE
                               NO. 1
                         FLOTATION CIRCUIT
                               NO. 2
                         FLOTATION CIRCUIT
     RETREATMENT
        CIRCUIT
                               NO. 3
                         FLOTATION CIRCUIT
                                T
       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
                       PYRITE
                    CONCENTRATE
                     FINAL PYHITE
                    CONCENTRATE1
             CONTAINS 3.43 KILOGRAMS Phil
              METRIC TON (100 TROY OUNCLS
              PER SHORT TON)
                               I LI-59

                              DRAFT

-------
                             DRAFT
Less than 1 percent of the current domestic production of
silver is recovered by amalgamation or cyanldation 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 (A1203)
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 chemi-
cal 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 rained 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 uncon-
solidated sands and clays, but some slumping does occur.

Underground mining occurs at one Arkansas facility, and this
operation provides the low-silica ore essential to the com-
bination process of refining.  Although this type of mining
is relatively costly, it is a viable alternative to the pur-
chase of foreign ores at elevated prices.  However, one of
the operations utilizes imported bauxite for blending of
ore grades.  Milling of the bauxite ore involves crushing,
ore blending, and grinding in preparation for refining.  In
1972, less than 10 percent of the bauxite used for primary
aluminum production was of domestic origin.  With the Increas-
ing 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
                             111-60

                             DRAFT

-------
                             DRAFT
insufficient to meet present needs of the nine domestic
refineries.  Recent price increases in foreign bauxite
supplies aid in assuring the future of domestic bauxite
operations, regardless of the limited national reserves.

The search for potential economic sources of aluminum per-
sists, and many pilot projects have been designed to produce
aluminum.  Currently, the most notable attempt to utilize
an alternative source of aluminum Is a 9 metric ton (10 ton)
per day pilot plant which converts alunite,  K2A16^0H)12(S04)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.  Addi-
tionally, the processing of alunite creates  no significant
"red mud" (leach residue) environmental problems.  Currently
alunite mining is in the exploratory stages, with a commercial-
scale refinery is slated for construction in 1975.  Full-scale
mining will entail drilling, blasting, and hauling using
bench mining techniques.  From all indications, alunite may
provide an economical new source of aluminum.

Bauxite production in the United States has declined recently
from a peak year in 1970, and preliminary production figures
for 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 these operations Is based on the use of a portion of their
end product in the production of ferroaJloys (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
past and are included in present reserves.  Beneficiation
                            111-61


                             DRAFT

-------
                            DRAFT
TABLE 111-17. PRODUCTION OF BAUXITE IN THE UNITED STATES
YEAH
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
                           111-62


                           DRAFT

-------
                              DRAFT
techniques include numerous gravity processes, jigging,
tabling, sink-float, Humphreys spirals; flotation, both
basic-sulfide and fatty-acid; and a variety of ore leach-
ing 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.  Pro-
duction figures for the ferroalloy mining and milling industry
since 1945 are summarized in Table II1-18.

As Table 111-18 shows, molybdenum mining cind milJJng constitute
the largest and most stable segment of Llie ferroalloy ore
mining and milling industry in the United States.  The U.S.
produces over 85% of the world's molybdenum supply, with two
mines dominating the industry.  These two mines cire among tho
25 largest mining operations In the U.S.  Production is
expected to Increase in the near future wirli i-.xp.mded output.
from existing facilities,  and at least mio major in-.w oper.i-
tion in Colorado is expected to be in operation soon.

The only commercially important ore of molybdenum Is
molybdenite, MoS^.  It is mined by both open-pit and under-
ground methods and is universally concentrated by flotation.
Commercially exploited ore currently ranges from O.I- to 0. )
                             Ill-bJ


                             DRAFT

-------
                              DRAFT
            TABLE HI-IS. PRODUCTION OF FERROALLOYS BY
                        U.S. MINING AND MILLING INDUSTRY
METAL
Chromium
CoJumbJum 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
(15,1501
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
" Reference 5
                               111-64

                              DRAFT

-------
                             DRAFT
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 con-
centrate); with the sale of some stockpiled material, only
about 50 mines operated with a price of $43 per unit (Refer-
ence 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 (
and the wolframite series, wolframite ((Fe, Mn)WOM , ferberite
(FeWOji) , and huebnerite (MnWO^.  Underground mining predomin-
ates, and concentration is by a wide variety of techniques.
Gravity concentration, by jigging, tabling, or sink float
methods, is frequently employed.  Because sliming due to the
high friability of scheelite ore (most U.S. ore is scheelite)
reduces recovery by gravity techniques, fatty-acid flotation
may be used to increase recovery.  Leaching may also be
employed as a major beneficiation step and is frequently
practiced to lower the phosphorus content of concentrates.
Ore generally contains about 0.6 percent tungsten, and
concentrates containing about 70 percent W03^ are produced.
A tungsten concentrate Is also produced as a byproduct of
molybdenum milling at one operation in a process involving
gravity separation, flotation, and magnetic separation.

Managanese 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.
                             111-65


                             DRAFT

-------
                              DRAFT
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 deoxldizer, 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
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 carbanwto leach-
ing (Reference  13).   It is most  likely  that heavy-media
separation will  be practiced  in the immediate future.
                            111-66


                             DRAFT

-------
                              DRAFT
 Nickel  ores  are not widely  available  in  the U.S.  The
 lateritic  deposit which  is  currently  being mined  is the
 only  known domestic deposit of  its kind.  Some  sulfide
 nickel  ore deposits with commercial possibilities have
 been  found in Alaska  (Reference 2).   If  they are devel-
 oped, processes entirely different from  those in use at
 the present  operation will  be employed.  Most likely,
 processing will involve  selective  flotation with reagent
 and water  usage and pollution problems quite similar to
 those of Canadian nickel operations (Reference  14) and
 domestic copper sulfide  deposits (Reference 5).

 There are  no mines or mills currently active in the U.S.
 producing  ores or concentrates  of chromium, cobalt, colum-
 blum, 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,
 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 (Reference 5).

 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
 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 influ-
ence on the cobalt market (Reference 5).  Known domestic
ore from which cobalt might be  recovered is a complex copper
                             111-67


                             DRAFT

-------
                            DRAFT
 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.   Concentra-
 tion techniques varying widely from fairly  simple gravity
 and  hand  picking techniques through magnetic  and electro-
 static  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 parti-
 cular, hydrometallurgical  processes like  those used in vana-
dium 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 met.ils  may be bend'tciated In
the same  type of  process used  for vanadium.  There  Is also
some similarity to tungsten, moJybdenum,  and  chromium.

Ferroalloy Ore Beneficlatlon Processes

Ore processing in the  ferroalloys industry varies widely.
and even ores bearing  the  same ore  mineral may be coneentrdied
                             111-68


                            DRAFT

-------
                             DRAFT
by widely differing techniques.  There is thus no scheelite
recovery process or pyrolusite concentration technique per £>e_.
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 func-
tions (such as crushing and grinding ore) that are common to
nearly all ore concentration procedures.  Fundamental ore
beneficlation 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
beneficiatlon 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, sink floatation,  Humphreys spirals, and
tabling.  Jigging Is applicable to fairly coarse ore, ranging
in size from 1 mm to 13 mm (approximately 0.04 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, 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
                             111-69

                             DRAFT

-------
                              DRAFT
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 applica-
tion 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 peffect, and the concentrate grades
into gangue  at the edge of  the  table through a mixed product
called middlings,  which is  generally collected separately
from concentrate and gangue and then retabled.  Frequently,
several sequential stages of tabling are required  to produce
a concentrate  of the desired grade.   Particle  size, as well
as density,  affects the behavior of particles  on a shaking
table, and the table feed generally must be well classified
to ensure  both high ore recovery and a good concentration
ratio.  Humphreys spiral separators are  useful  for ore
ground to  between 0.1 mm and 2  mm  (approximately 0.004 to
0.08 inch) (Reference 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 cassiterlr.'
columbite, and monazite (Reference 5  ) or the  separation
of casslterite and wolframite (Reference 18).  Although
they are both  most frequently Implemented as dry processes,
wet-belt magnetic  separators are used.   Since  ore  particle:-.
are charged  to 20,000 to 40,000 volts  for electrostatic
                              111-70


                             DRAFT

-------
                              DRAFT
 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 grind-
 ing 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
MoS^. concentrate from  0.3 percent) (Reference 18 ).
 Its specificity also allows separation of different  ore
minerals (e.g.,  CuS and MoSjO  where desired,  and operation
with minimum reagent consumption since reagent interaction
is typically only with the particular materials to be
floated or depressed.
                            111-71


                             DRAFT

-------
                             DRAFT
Figure 111-13. GRAVITY-PLANT FLOWSHEET FOR NIGERIAN COLUMBITE
                                                                    TO
                                                                   WASTE
                                                           	*- 111 Mil I
                        SOURCE Rtl!tl4CNCir 19
                             111-72
                            DRAFT

-------
                                DRAFT
   Figure 111-14. EUXENITE/COLUMBITE BENEFICIATION-PLANT FLOWSHEET
TO
| DREDGE""]
HE AW MINERAL
CONCENTRATE
t
[ STORAGE |


i
1

__l
[ SCREEN | >| ROD







M.LL r*j MAGNETITE }-+. '»,„,.,
r


ro
VASTE
TO WASTE -«— SLIME -4 	 JCLASSIFIER [^ATTRITIONER^ SEPARATOR f*jCLASSIFIER|-^{ ORVER


RAGE-"- °OARTZ

RAGE"*- OARN"



t
INDUCED-ROLL
MAGNETIC SEPARATOI
L
1


'
	 1 MAGNETIC SEPARATOR
^~J (LOW INTENSITY)
'
,- .— J SCR
!_


1
MJ



>


*

*
INDUCED-ROLL
\GNETIC SEPARATOR



\ f URNACE |
i NONCONDI
H ELECTROSTATIC
SEPARATOR
NONCONDUCTORS
— . INDUCED-ROLL
v MAGNETIC SEPARATOR
T
4 CROSSBELT MAGNETIC
SEPARATOR
ELECTROSTATIC 1
SEPARATOR |
ICTORS— 1
1 1— MIDDLINGS — — »•
NONMAGNETICS
SCREEN "1
1— CONDUCTORS 	 •,
f SCREEN |
< 28 MESH 1

}
1 CROSSBELT MAGNETIC
SEPARATOR

CROSSBELT MAGNETIC
SEPARATOR 	
1
	 ' 1 ILMENITE |—
> 36 MESH < 36 MESH 1 NONMAGNETICS
tX NONMAGNETICS J.
f ALTERNATE 1 f

'

| AIR TABLE |



TE -^-TAILINGS -4-| WET TABLE | 	 DRIED CC


CONCENTRATE
INCENTRATE » 1
ICHOSSBELI MAGNETIC 1
SEPARATOR [-


, 	 1
| MONAZITE ] |

FUXFNITE ] C
NONMAUNFTICS 	
AND MIDDLINGS •"
~l
9LUMBIII |
1
»
-^.'
                                                                TO
                                                                STORAGE
                                                               STORAGE
                                  TO STORAGE


                                111-73
                               DRAFT

-------
                                  DRAFT
    Figure 111-15. REPRESENTATIVE FLOW SHEET FOR SIMPLE GRAVITY MILL
   TO
WASTE
TAILS-
                             MINING
                              ORE
                              i
                         GRINDING AND
                           CRUSHING
                              I
                           SCREENING
                              FINE
                           SHAKING
                            TABLE
                      MIDDLINGS-
                                            COARSE
                         - HEADS
                           •TAILS-
                                                  MIDDLINGS
                                          SHAKING
                                          TABLES
                                            J
                                                        •CONCENTRATE
                                 111-74


                                 DRAFT

-------
                             DRAFT
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, activa-
tors, 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 (C2H5G . NaCSjZ).
Frothers are generally organics with a soluble hydroxyl group
and a "non-wettable" hydrocarbon (Reference 17 ).  Pine oil
(C6H120H), 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 I11-16.
Water use in this operation currently amounts to approxi-
mately 1.8 tons of water per ton of ore processed, essentially
all of which is process water.  Reclaimed water from thick-
eners at the mill site (shown on the flowsheet) amounts to
only 10 percent of total use.

Where byproducts are recovered with molybdenite, a somcwh.it
more complex mill flowsheet results, although the
                             111-75

                             DRAFT

-------
                        DRAFT
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
NokM 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 aminei
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 SO*
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
                        111-76




                        DRAFT

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


BALL
MILLS
\


1
CYCLONES 1 	 UNDERFLOW 	 1

                                    OVERFLOW
                                  ROUGHER FLOAT

                                    MIDDLINGS

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


 	MIDDLINGS •
    »«-UNDERFLOW
      -UNDERFLOW-I THICKENER
1

CLEANER
FLOAT
16 STAGES WITH
REGRIND AND
INTERNAL RECYCLE)
                                                                    T
                                                                 CONCENTRATE
                                                                            — TAILS -
                                                  UNDERFLOW-
                                                                   ORYER
                                                                 MOI YBDtNUM
                                                                  PHODUri
TO TAILINGS
  FOND
                                        111-77
                                       DRAFT

-------
                             DRAFT
molybdenite  recovery circuits  themselves  remain quite
similar.  A  very simplified  flow diagram  for  such an opera-
tion is shown  in Figure  111-17.   Pyrite flotation and
monazite  floation are accomplished  at  acid  pH (4.5 and 1.5,
respectively), somewhat  increasing  the likelihood of solu-
bllizing  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 flotat-
tion from complex  ores in processes involving multiple flota-
tion steps.  Some  of the most likely reagents to be used in
these operations are presented in Table 111-20, 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 pro-
cesses to be,  in general, somewhat more sensitive to feed-
water conditions than sulfide floats;  consequently,  they are
less frequently  run  with recycled water.

In current U.S.  operations,  scheelite  is  rccovurt-d hy
flotation using  fatty acids as collectors.  A Lyplral 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/sliorc
ton), and sodium oleate  (0.1 to 0.2 kg/metric ton, or 0.2 to
0.4 Ib/short ton).   In addition,   materials such as copper
sulfate or acid  dichromate may be used in small to moderate
                            111-78


                             DRAFT

-------
                         DRAFT
 Figure 111-17. SIMPLIFIED MOLYBDENUM MILL FLOW DIAGRAM
                      CRUSHING
                      (3 STAGES)
                     28% + 3 MESH
                      GRINDING
                     BALL MILLS
                    36% + 100 MESH
                     FLOTATION
 •CONCENTRATE
                         1
                      FLOTATION
                  96% OF MILL FEED
•LIGHT TO TAILS-
                      GRAVITY
                  HUMPHREY'S SPIRALS
                         i
                       PYRITE
                     FLOTATION
                       TAILS
                         t
LIGHT TO TAILS
  MONAZITE
CONCENTRATE
  TO TAILS
-i
                       TABLES
                      MONAZITE
                      FLOTATION
                     MAGNETIC
                     SEPARATION

              NONMAGNETIC
            TIN CONCENTRATE
                                             I
                                        CONCENTRATE

                                         CLEANER
                                         FLOTATION
                                         (4 STAGES)
                                            I
                                                   -TAILINGS
                                          DRYING
                                            t
                                        MOLYBDENUM
                                        CONCENTRATE
                                         (93% + MoS2)
                        MAGNETIC TUNGSTEN
                           CONCENTRATE
                         111-79


                         DRAFT

-------
                              DRAFT
TABLE 111-20. PROBABLE REAGENTS USED IN FLOTATION OF
             NICKEL AND COBALT ORES
                   Lime
                   Amyl Xanthate
                   Isopropyl Xanthate
                   Pine Oil
                   Methyl liobutyl Carbinol
                   Triathoxybutane
                   Dextrin
                   Sodium Cyanide
                   Copper Sulfate
                   Sodium Silicate
                                111-80

                                DRAFT

-------
                              DRAFT
amounts as conditioners and gangue depressants.  Scheelite
flotation circuits may run alkaline or acid, depending
primarily on the accessory mineralization in the ore.
Flotation of sulfides which occurs with the scheelite is also
common practice.  Sulfide float products may be recovered
for sale or simply removed as undesirable contaminants for
delivery to tails.  Frequently, only a portion of the ore
(generally, the slimes) is processed by flotation, the
coarser material being concentrated by gravity techniques
such as tabling.  A simplified flow diagram for a small
tungsten concentrator Illustrating these features is shown
in Figure 111-18.  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.  Approxi-
mately four tons of water are used per ton of ore processed
in this operation.

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

Reagents used include:

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

Soap skimmings                40 kg/metric ton (80 Ib/short ton)

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

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

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

While the flotation processes are similar, columbium and
tantalum flotation plants are likely to possess an unusual
degree of complexity due to the complex nature of their ores,
which necessitates multiple processes to effectively sepa-
rate the desired concentrates.  This is illustrated in the
flowsheet for a Canadian pyrochlore (NaCaCb£06_F) mill in
Figure 111-19.
                             111-81

                              DRAFT

-------
                                  DRAFT
Figure 111-18. SIMPLIFIED FLOW DIAGRAM FOR SMALL TUNGSTEN CONCENTRATOR
                  ORE
                                 SULFIDE
                                FLOTATION
                                 CYCLONE
                                 75% SANDS
                                   i
                                 GRAVITY
                                 TABLES
                              TAILINGS
                                              25%
                                             SLIMES
                                                            OVERFLOW
    THICKENER
    SCHEELITE
    FLOTATION
                                                       HCI LEACH
                                                      (15 TO 20% OF
                                                        FRACTION)
  TUNGSTEN
CONCENTRA1L
                                 111-82


                                 DRAFT

-------
                                              DRAFT
    Figure  111-19. MILL  FLOWSHEET  FOR A CANADIAN COLUMBIUM OPERATION
             I   MINING  I
                 ORE
               PRIMARY
               CRUSHING
      > lA-Cffl
      » i*-em  r
    r~<0.7B.|nJ~l

H SECONDARY I
 CRUSHING |
               3
               SCREENING


J SECONDARY |   < ' •—«» 1*-M
  CRUSHING
                MILL
               ORE BIN
              ROD MILLING
            I SCREENING 1—
                40 MESH

           -I BALL MILLING I
               PVRITE
              FLOTATION
 CYCLONE
OVERFLOW

                             •	CONCENTRATE
                                           ~L
                                         MAGNETIC
                                        SEPARATION
                        TAILS
                                         DESLIMING
                                                     J
     UNDERFLOW
        r
I   BULK FLOTATION  L
    CONOENTRATE
                                         DESLIMING
                                                     J
                                        UNDERFLOW

                                            *
                                         MAGNETIC
                                        SEPARATION
                                            I
                                      PRIMARY CLEANING
                                        (FIRST STAGE)
                                TAILS    CONCENTRATE
                                       PRIMARY CLEANING  I
                                        (SECOND STAGE!   I


                                        CONCENTRATE


                                             ^	
                                     SECONDARY CLEANING
                                       ITHREE STAGES)
                                                      — TAILS-
                                        CONCENTRATE

                                            t
                                                                      _»J MAGNETITE
                                                                                              1
                                         TABLING
                                                     1 - TAILS -»J
                                         REVERSE
                                        FLOTATION
                                                    >
                                                                                             TO
                                                                                           STOCKPILE
                   CONCENTRATE-*-

                                       TAILS
                                      _i_
                                      TABLING
                                                                        CONCENTRATE
                 - TAILS-
                               TO
                            ' TAILING
                              POND
                                                                                       CONCENTRATE
                                       SOURCE: REFERENCES
                                              111-83


                                             DRAFT

-------
                            DRAFT
Ore Leaching  Processes.   While not a predominant practice
in the ferroalloys  industry, ore leaching has played a. part
in a number of operations and is likely to increase as seg-
ments of the industry process ores of lower grade or which
are less easily beneficiated.  A number of leaching processes
have been developed for manganese ores in the search for
methods of exploiting plentiful, low-grade, difficult-to-
concentrate 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 chromice 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 well-
documented 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% V2_05^ and
some Cr, Ni, and Mo — yields vanadium as a byproduct 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 (Na2_C03_) 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
                            111-84

                           DRAFT

-------
                              DRAFT
metavanadate, and sodium chromate, while the iron remains  in
insoluble form and  is not extracted  in a water leach following
the roast.

Phosphate values are removed  from the leach in three stages
of crystallization  (Figure 111-20).  Vanadium can be recovered
as V2_0|5_ (redcake) by acidification,  and chromium is precipi-
tated as lead chromate.  By this process, 85 percent of vanadium,
65 percent of chromium, and 91 percent phosphorus can be ex-
tracted.

Another, basically non-radioactive,  vanadium ore, with a grade
of 1 percent VjZO^, is found in a vanldlferous, mixed-layer
montmorillonlte/illite and goethite/montroseite matrix.
This ore is opened up by salt roasting, following extrusion
of pellets, to yield sodium metavanadate, which is concen-
trated 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 carhamate 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 111-21).
During this time, the environmental hazards and extremely
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
Insectlde, Fungicide, and Rodentlclde Act.   In addition,
registration has been suspended for mercury alkyl compounds
                             111-85

                            DRAFT

-------
                             DRAFT
Figure 111-20. FLOWSHEET OF TR1STAGE CRYSTALLIZATION PROCESS FOR
           RECOVERY OF VANADIUM, PHOSPHORUS, AND CHROMIUM
           FROM WESTERN FERROPHOSPHORUS
       I FERROPHOSPHORUS |
                    | NlCt |
                          ROASTING

I SETTLING AN


|
| PRIMARY CRYSTALLIZATION]
I
[ PRIMARY CENTRIFUGINC
PREGNANT CHVS
SOLUTION 1 M PO i
1 *~^ ^


| TERTIARY CRVSTALLIZATION ]


| TERTIARY CENTRIFUGING
L
•« 	 %£??. . 1 C.O. 1 iOtUTIO
>
DOECANTATION |
1
SOLIDS WATEB WAJM
1 A
T T
| FILTRATION |
j |
RESIDUE WASH ,.
1 LIQUOR
, . ^
TALS [ WATER ]
•— ^^— DISSOLUTION 	 *
•
[ SECONDARY CRYSTALLIZATION]


| SECONDARY CENTRIFUCING 1
] 1 1
T t
CRYSTALS SECONDARY
EGNANT I SOLUTION — *"
V 1
                                                         • TO WASTE
        RED CAKE  MOTHER LIQUOR


                  * 	
       ' FUSION •
          i
                I
FILTRATION
     BLACK-CAKE
     VANADIUM PRODUCT
                          CHflOMAlE
                          SOLUTION
                  TO
                 WASTE
| FILTRATION
J
cuanuinu 1
PRODUCT j
ZJ
1 » SOLUTION
TO WASTE

» 1U
"" STOl'KPIlE
                             LII-86
                            DRAFT

-------
                          DRAFT
   Figure 111-21. ARKANSAS VANADIUM PROCESS FLOWSHEET

                        1.5 -2.0% V2O5

                             t
      6-10%
      NaCL
TERTIARY
 AMINES
                         GRINDING
                        PELLETIZING
                            I
                         ROASTING
                            I
                 850°C (1562°F)
                           NaVO,
H2S04


LEACHING AND
ACIDIFICATION
                            I
                                     pH 2.5 • 3.5
                         
-------
                               DRAFT
  Figure 111-22. FLOWSHEET OF DEAN-LEUTE AMMONIUM CARBAMATE PROCESS

                      RAW SIZED ORE - 1.9 cm (0.75 in.)



                                                 -CO,
            REDUCING FURNACE
                          GRINDING (30 MESH)
 WEAK Mn
 SOLUTION
                                 I
                LEACHING
           (TWO 11,356-j£(3000-GAL))
              REACTION TANKS
                IN SERIES
                                 I
                         9.14-m (30-ft) CLARIFYING
                              THICKENER
7.6-m (25-ft) COUNTERCURRENT
   WASHING THICKENERS
             LIVE
             STEAM
                                    • NEW LEACH LIQUOR
                                     -ill
                                        LEACH LIQUOR
                                        REGENERATION
               TWO 11.356- X.
                (3000-GAL)
              PRECIPITATION
             TANKS ON SERIES)
-NH,
                                MnCO,
 SLURRY
  STILL
   I
NH4NH2CO2
               CLARIFYING
               THICKENER
 MOTHER
' LIQUOR "
TAILINGS
                              70% SOLIDS
                                 i	
                            ROTARY DRYER
                                •NH4NH2CO2
                                 I
                              PRODUCT
                               111-88


                               DRAFT

-------
                        DRAFT
  TABLE 111-21. DOMESTIC MERCURY PRODUCTION STATISTICS
CATEGORY
No. of producing mines
Production in metric tons
(flasks)
Dollar value (thousands)
YEAR
1969
109
1,029
(29,640)
$14.969
1970
79
948
(27.296)
$11.130
1971
56
621
(17,883)
$ 5.229
1972
21
253
(7.286)
$1.590
1973
6


SOURCE: REFERENCE 2
                        111-89
                        DRAFT

-------
                             DRAFT
and nonalkyl uses on rice seed, in laundry products, and in
marine antifouling paint.  An immediate effect of this has
been a substantial reduction in the demand for mercury for
paints and agricultural applications.  However, future growth
in the consumption of mercury is anticipated for electrical
apparatus, Instruments, and dental supplies.  From consider-
ation 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 pro-
duction 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 operat I-in
scheduled to begin in Nevada later in 1975 will concentrate
mercury ore by flotation.  This concentrate will be furnaci-u,
and annual production of mercury from the operation is e.\p.
to reach 20,000 flasks (695 metric tons, or 765 short tons).
                            111-90


                             DRAFT

-------
                             UHAFT
Uranium, Radium, and Vanadium Ores

The mining and milling of uranium, vanadium, and radium
constitute one industry segment, because uranium and
vanadium are often found in the same ore and because radium,
resulting from the radioactive decay of uranium, has always
been obtained from uranium ores.  In the past 20 years, the
demand for radium has vanished as radioactive isotopes
(e.g., Co 60, Pu 240) with tailored characteristics as sources
of radiation have become available.  Radium is now treated
as a pollutant in the wastes.  Uranium is mined primarily
for its use in generating energy and isotopes in nuclear
reactors.  In the U.S., vanadium is primarily generated as a
byproduct of uranium mining for use as a ferroalloying metal
and, in the form of its oxide, as a catalyst.  Vanadium used
as a ferroalloy metal has been discussed in the Ferroalloys
Section.

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

Because It would be uneconomical to transport low-grade
uranium ores very far, mines are closely associated with mills
that yield a concentrate containing about 90 percent uranium
oxide.  This concentrate is shipped to plants that produce
compounds of natural and isotopically enriched uranium for
the nuclear industry.  The processes of crushing and grinding,
conventionally associated with a mill, are intimately connected
with the hydrometallurgical processes that yield the concen-
trate, and both processes normally share a wastewater disposal
system.  Mine water, when present, is often treated separately
and is sometimes used as a source of mill process water.
Mine water frequently contains a significant amount of
uranium values, and the process of cleaning up mine water
not only yields as much as one percent of the product of some
mines but is also quite profitable.
                             111-91


                             DRAFT

-------
                             DRAFT
The uranium oxide concentrate, whose grade is usually quoted
in percent of \J3Q^  (although that oxide figures in the assay,
rather than in the product), is generated by one of several
hydrometallurgical processes.  For purposes of wastewater
categorization, they may be distinguished as follows:

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

     (2)  Values in the leachate are sometimes concentrated
          by ion exchange, either in a solid resin (IX) or
          by solvent extraction (SX).  They are then precipi-
          tated as the concentrate, yellowcake.

Some vanadium finds are not associated with significant uranium
concentrations.  There, as in vanadium byproduct operations,
vanadium oxide (V_205) is made by water leaching following a
salt roasting operation.  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 111-22.

Primary deposits of uranium ore contain uraninite, the U(IV)
compound U02^ and are widely distributed in granites and
pegmatites.  Pure speciments of this compound, with density
ranging to 11, are rare, but its fibrous form, pitchblende,
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
V02_ and UOJK  The composition U30£ — i.e., U0£.21)03^ — is
particularly stable.  The process occasionally stops at gumniiU'
(U02^H2_0) , an orange or red, waxy mineral, but usually
involves further oxidation and reactions with alkaJine and
alkaline-earth oxides, silicates, and phosphates.  The trnns-
                              111-92


                             DRAFT

-------
                 UHAM
TABLE 111-22. ISOTOPIC ABUNDANCE OF URANIUM
ISOTOPE
U238
U235
U234
U236
HALF-LIFE (YEARS)
451 x 109
7.13 X108
2.48 x 105
2.39 x 107
ABUNDANCE
9927%
0.72%
0.0057%
Traces Identified
(Moon-1972; Earth-1974)
                 111-93





                 DRAFT

-------
                             DRAFT
 port  leads  to the surface uranium ores of  arid  lands,  including
 carnotite  (K2/U02)2/V04;2.3H2p),  uranophane (CaU2Si20n^7H20) ,
 and autunite (Ca(U02)2_(P04)£.10-12H2_0) 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.

 A reducing  environment  is often  provided by  decaying biologi-
 cal 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 encoun-
 tered 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  (197-ft) grid,
and development drilling,  on  a 15-m (49-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  costs  and the  possibility of increased
uranium demand due to the  current  energy situation have
resulted in  the mining,  for storage, of  ore  below this thres-
hold 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

               New Mexico           44%
               Wyoming              39%
               Utah                 4%
               Colorado             4%
               Texas                4%
               Others               5%
                             111-94


                             DRAFT

-------
                             L/ru+r i
 The number  of  separate  deposits  in  the  western  United States
 exceeds  1000,  but  half  of  the  reserves  lie  in 15 deposits.
 Four  of  these,  in  central  Wyoming,  on the border between
 Colorado and Utah,  in northwestern  New  Mexico,  and on the
 Texas gulf  coast,  dominate the industry.  In 1970, New Mexico
 provided 46 percent, Wyoming 26  percent, Colorado 12 percent,
 and Utah 7  percent  of uranium  production, for a total of 91
 percent  of  U.S. production.

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

 Mining practice is  conventional.  There are 122 underground
 mines as of 1 January 1974, with a typical depth of 200 m
 (656  ft).   Special  precautions for the  ventilation of under-
 ground mines reduce the exposure of  miners to radon, a short-
 lived, 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 by
 leaching and ion exchange.

 Because  of  the small size  of pockets of high-grade ore, open-
 pit 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 stock-
 piles.   On January  1974, 33 open pit mines were being worked,
 and 20 other (e.g., heap-leaching) sources were in operation.

Most mines ship ore to  the mill by truck.   In at least one
 Instance, a short  (100-km, or  62-mi.) railroad run is involved.
Most mining areas share at least two mill processes, one
 using acid leaching and the other, for high limestone content,
using alkaline leaching.
                             111-95


                             DRAFT

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

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

 Ore high in vanadium is sometimes roasted  with sodium chloride
 at this stage Co convert insoluble heavy-metal vanadates (vanadium
 complex) and carnotlte to more soluble sodium vanadate, which is
 then  extracted  with water.  More  often,  this process  is deferred
 until after concentration of  uranium/vanadium values.  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 70  micrometers  (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 (CaC03_) 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 'J(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
                              111-96


                             DRAFT

-------
                     DRAFT
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

NO. OF MILLS
3
7
3
2
3
1
1
20
                     111-97
                     DRAFT

-------
                           DRAFT
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:
                          ^   0 £ — > 2U03_

          2U03. + 2H2SO4 + 5H20->2 (U02S04) . 7H20

Uranyl sulfate (U02S04) forms a complex, hydrouranyl tri-
sulfuric acid  (H4U02/S04)3) , in the leach, and the anions
of this acid are extracted  for value.

Alkaline Leach .   A solution of sodium carbonate (40 to
50 g per liter) in an oxidizing environment selectively
leaches uranium and vandium values from their ores.  The
values may be precipitated  directly from the leach by rais-
ing the pH with the addition of sodium hydroxide.  The super-
natant 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 cen-
tral 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:
                        (oxidation)
       3Na2_(C03J + UO^ + H20 — > 2NaOH
                         (leaching)
                2NaOH + C0£ -> Na2C02 + H20
                     (recarbonization)
    2Na(U02)(C03)2 + 6NaOH  ->Na2y207^   + 6Na2C03_ + 3H2
                      (precipitation)
                            111-98


                           DRAFT

-------
                    DRAFT
Figure 111-23. PACHUCA TANK FOR ALKALINE LEACHING
                                              LEACH
                                              AIRLIFT
                   111-99





                   DRAFT

-------
                             DRAFT
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 wastewater but are often  ignored.
Purification to  remove sodium ion is  necessary both to  meet
.the  specifications of  American uranium processors and for
the  preparation  of natural uranium dioxide fuel.   The latter
process  will be  used to Illustrate the problem caused by
excess sodium.  Sodium diuranate may  be considered as a mix-
ture of  sodium and uranyl oxides — i.e., Na2.U2.07, --- Na2_0  +
The process  of  generating U02_ 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 U03^
component to U02^  and  nitrogen (which  acts as an inert gas
and reduces  the risk  of explosion in  and around the  reducing
furnace).  With sodium diuranate as a feed, the process
results in a mix  of U0£ and  Na20 that is difficult to purify
(by water leaching of NaOH)  without impairing  the  ceramic
qualities of uranium  dioxide.  When,  in contrast,  ammonium
diuranate is used as  feed,  all byproducts are  gaseous, and
pure  U02_  remains.  The structural integrity of this  ceramic
is immediately  adequate for  extended  use in the popular
CANDU (Canadian deuterium-uranium)  reactors.   Sodium ion,
as well as vanadium values,  can  be  removed  from raw  yellow
cake  (sodium diuranate) produced by alkaline leaching in
two steps.   In  the first step, the  yellow cake is  roasted,
and some  of  the sodium ion forms water-soluble sodium vanadate,
while organics  are carbonized and burned off.   The roasted
product is water  leached,  yielding  a  V2_0_5. concentrate as des-
cribed below.   The remaining sodium diuranate  is redissolved
in sulfuric  acid,

      Na2jU2_0;7_ +  3H2_SCM ->Na2S(M + 3H20 + 2(U02_)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.
                             III-100

                             DRAFT

-------
                            DRAFT
           U02SCM + H20 + 2mi_ - > (NH4) 2804^ + U03_

 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 a hypothe-
 tical 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 to fer-
 tilize alkaline southwestern soils.   Ammonium sulfate could
 be generated by neutralizing the wastes of the hypothetical
 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 ammonium ion, present a present health hazard
 more severe than that from sulfate ion.
 Vanadium Recovery.    Vanadium,  found in carnotite (K2(U02J
 (V04)2^ .  3H20)  as well  as  in heavy metal vanadates — e.g.,
 vanadinite  (9PbO .  3V205^ .  PbC12)— is converted  to  sodium
 orthovanadate (Na.3V04) ,  which is  water-soluble by roasting
 with  sodium chloride  or  soda ash  (Na2C03) .   After water
 leaching, ammonium  chloride is  added,  and  poorly soluble
 ammonium vanadates  are  precipitated:

       Na3V(M + NH4C1 +  H20 — ^ 2NaOH + NaCl + NH4V03_
                  (ammonium metavanadate)

             Na3VO^  +  3NH4C1 — > 3NaCl  + (NH4)J3V(M
                  (ammonium orthovanadate)

 The ammonium vanadates are  thermally  decomposed  to  yield
 vanadium pentoxide:

             2(NH4)3V04  — > 6NH3    +  3H20
A significant fraction  (86 to 87%) of V205^ is used in
the ferroalloys industry.  There, ferrovanadium has been
prepared in electric furnaces by the reaction:

              V205^ + FeZp^ + 8C — > SCO   + 2FeV

or by alumino thermic reduction (See Glossary) in the
presence of scrap iron.

The remainder of y2_05^ production is used in the inorganic
chemical industry, and its processing is not within the
scope of these guidelines.
                           111-101
                           DRAFT

-------
                             DRAFT
Since  the mining  and  beneficiation  of vanadium ores not con-
taining uranium values  present an excellent example of
hydrometallurgical  processes  in  the mining and ore dressing
of ferroalloy metals  (under SIC  1061),  it will be explored
further under that  heading.   Because of the chemical similarity
of vanadium  to columbium, tantalum, and other ferroalloy
metals, recovery  processes for vanadium are likely to be
quite  similar to  hydrometallurgical processes that will be
used in the  ferroalloys mining industry when it becomes
more active  again.

Concentration and Precipitation.    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 pre-
cipitated 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 concentra-
tions  in the eluate of  some of these processes are shown in
Table  111-24.

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-24,i) .
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 HI-24b, c).  Sands
are often repulped, by  the addition of some wastewater streau
or another,  to facilitate flow to the tailing pond as much
                            III-102


                             DRAFT

-------
                      DRAFT
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
12 to 1.6
IX/SX combination
Eluex process
3.0 to 7.5
                     III-103






                     DRAFT

-------
                                DRAFT
  Figure 111-24. CONCENTRATION PROCESSES AND TERMINOLOGY (Sheet 1 of 2)
        FROM
        LEACH
WATER
            ¥»
                  PREGNANT
                  LEACH LIQUOR
                 »*«»o
                         REPULPING
               CLEAR LEACH LIQUOR
               TO COLUMN IX OR SX

                    a)  LIQUID/SOLID SEPARATION
                                                        SLIMY PULP TO
                                                        RESIN-IN PULP IX
                                                      SAND
                                          <>»    Q~
                                                    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
                                III-104


                                DflAFT

-------
                             DRAFT
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
PREG
LEAC
LIQU
0

H
u
o Ao«
Op°'/
W Oo
' o°
BARREN STI
* ELUANT SO
I
LOADED
i—1 l— i ORGANIC |4J— ' !-i
SOLVENT °o o
- * • •

••M^^^^M^
DIPPED
LVENT
SOLVENT
~ 	 ~ —

HV J^ ([BARREN ^y JPREGNANT! i
^~* '| LIQUOR ^-^ ELUATE 11'
PHASE PHASE
LOADING SEPARATION STRIPPING SEPARATION
e) SOLVENT EXTRACTION
LEACH
f) E

IX
f
SX
RECYCLE
BARREN t» " ELUANT
ELUANT ' ^i f
IX IX
. I* \
PREGNANT
.ELUATE

PARTIALLY STORAGE)
STRIPPED \ . LOADED
RESIN ^ _ ' RESIN
g) SPLIT ELUTION


f PRECIPITATION
LUEX PROCESS
                             III-105
                             DRAFT

-------
                             DRAFT
as a few kilometers  away.   Consequently,  there  is some latitude
for the selection of the wastewater  sent  to the tailing pond,
and mill operators can  take advantage of  this fact in selecting
environmentally  sound waste-disposal procedures.

Ion exchange and solvent extraction  (Figure III-24b-e) are
based on the same principle:  Polar  organic molecules tend to
exchange a mobile ion in their  structure  — typically,
C1-, N03_-, HS(M-, C03_— (anions), or H+ or Na+  (cations) —
for an ion with  a greater charge or  a smaller ionic radius.
For example, let R be the remainder  of the polar molecule  (in
the case of a solvent)  or polymer (for a  resin), and let X
be the mobile Ion.   Then, the exchange reaction for the
uranyltrisulfate complex Is

          4RX +  (U02SCM)3) ---- * - - (141102504   + 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-f-H- > Mg-H- > 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 anlonic
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, wirli
33 times as much uranium as vanadium being captured.  Thesi-
variations In affinity, multLple columns,  and control of
leaching time with respect uo  breakthrough (the time when
the interface between loaded and regenerated resin, l''ij;urc
III-24d, arrives at the end of the column) are used ti> makv
                             III-106


                              DRAFT

-------
                           DRAFT
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 counter-
current SX concentrators that are useful even when each
stage effects only partial separation of a value from an
interferent.  Unfortunately, the solvents are easily polluted
by slimes, and complete liquid/solid separation is necessary.
IX and SX circuits can be combined to take advantage of both
the slime resistance of resln-ln-pulp ion exchange and the
separatory efficiency of solvent exchange (Eluex process).
The uranium values are precipitated with a base or 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 actlnide  (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  con-
tain  from  1  to  10  percent thoria (Th02).  American ores  from
                            II1-107

                            DRAFT

-------
                            DRAFT
the North and South Carolinas, Florida, and Idaho contain
1.2 to 7 percent Th02^ 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 concentra-
tion 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, casslterite, or other materials that concen-
trate in a similar way.  Monazite is brittle, radioactive,
and magnetic, permitting concentration by magnetic means.
There are some deposits of consolidated monazite sands in
Wyoming.

Hydrometallurglcal 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 111-27), 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 Ib) of mona-
zite sulfate and is not very economical.  When used in con-
junction 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 environ-
mental 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
                            III-108
                            DRAFT

-------
                           DRAFT
Figure 111-25. SIMPLIFIED SCHEMATIC DIAGRAM OF SULFURIC ACID DIGESTION
          OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
          AND RARE EARTHS
TO
STOCKPILE
Fll
IU. A

SE
PREC
Al

*
FILTRATE
1
TO WASTE

— MAIN STREAM


I
RESIDUE
._ (UNDIGESTED MONAZITE SAND.
SILICA. ZIRCON. AND RUTILEI



FILTRATE
(R.E , U. AND P2O6)
|
SELE
PRECIP
AT(
CTIVE
IAIIUN •*
H23
I
t
-TRATE
ND P20g)
\
.ECTIVE
pH 60
t
*
PRECIPITATE OF
R E AND P2O5




}
PRECIPITATE OF
U AND PjOB
(BYPRODUCT)
TO SHIPPING
MONAZITE
SAND
*
GRINDING
OPERATION
1
DIGE!
1
DISSOl
1
1





1



Th. R E . U. AND PjOj 1 HjO
\
SELE
PRECIPI
ATp


(

TATION


NH


PRECIPITATE
ITh, R E . AND PjOjl
*
PURIFICATION BY SOLVENT EXTRACTION.
SELECTIVE PRECIPITATION. OR FRAC
TIONAL CRYSTALLIZATION

l

1
IO°C
i»FI

,OH

CONCENTRATES
TO SHIPPING
SOURCE- REFERENCE 20
                           III-109


                           DRAFT

-------
                             DRAFT
Figure 111-26. SIMPLIFIED SCHEMATIC DIAGRAM OF CAUSTIC SODA DIGESTION
          OF MONAZITE SAND FOR RECOVERY OF THORIUM, URANIUM,
          AND RARE EARTHS
MONAZITE
SAND
^^•^•B MAIN STREAM l •

IMSOH OPERATION
1 f
i
^___^ DIGESTION



f HYDROUS METAL-OXI
CRYSTALL.ZATION 1 Hh. U. AMU HA
* \
±Y DISSOLUTION

1 	 NaOH (Na3PO4) 1

.1
' ' fc SELECTIVE
^ PRECIPITATION

+
FILTRATE
(RARE EARTHS)
1 PRECIPITATE
(Th AND U)
^ SELECTIVP
"~ PRECIPITATION
1
1 *
FILTRATE! PRECIPITATE OF PURIFICATION
	 1 RARE EARTHS SOLVENT EXTRAC

JL (BY-PKODUCI) I





138°C
(280°F)
DE CAKE
.)





BY
;TIOIM

TO STOCKPILE CONCENTRATES h^- STOCKPILE
tllRrp- REFERENCE 2D 	 . 	 1
                             III-110
                             DRAFT

-------
                               DRAFT
 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
o
LU
DC
SI
O
                                        IDAHO MONAZITE SAND
                                      a INDIAN MONAZITE SAND
                                      A
                                             I	I	1	1
40 -
     20 -
                               ACIDITY (pH)
                 AGITATION TIME:
                 DILUTION RATIO:
                 DIGESTION RATIO:
                 NEUTRALIZING AGENT:

                 SOURCE: REFERENCE 20
                              5 MINUTES
                              H2O: SAND - 45:1 TO 50:1
                              93% H2S04: SAND - 1.77
                              3.1% NH4OH
                             Ill-Ill
                              DRAFT

-------
                             DRAFT
 uranium diuranate in Che 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.   Although 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 its analogous subcategories should adequately
 cover future operations for thorium.

 Radiation parameters of thorium and uranium daughters are
 somewhat different.   The two decay series are compared  in
 Table 111-25.   The uranium series is  dominated by radium,
 which—with a  halflife of over  1,600  years and chemical
 characteristics that are distinctly different from those  of
 the actinides  and lanthanldes—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, Ra228,  at 6.7  years,  finds
 use in luminous watches but lacks the long-term potency of
 Ra226.   In addition,  by equilibrium rules,  it can  only attain
 a concentration of less than one part in 2 billion.  Radium
 226 occurs at  one part in 3 million in an equilibrium of
 uranium and its daughters.   Thus,  radiation  problems in
 thorium mining are present  but  less critical than  in the
 case of uranium.

 Industry Flow  Charts.    Of  the  sixteen mills operating in
 1967  (Table 111-26) ,  no two used identical leaching concen-
 tration,  and precipitation  steps.   The same  was  probably
 true  of the 20 mills  operating  in  1972 (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,  vanadium,  and radium pro-
duction data (U.S. Bureau of Mines  1972  data  published in
1974) show  that  uranium production  increased  slightly in
anticipation of  uranium use in  commercial reactors.  This
trend is  expected  to  intensify  until  about 1975—fspecia]Jy,
in  response to  the energy situation of  1974.  Thereafter,
delays  in reactor  licensing,  the shortage of  waste-fuel
storage space,   and the  difficulties involved  in nuclear-waste
                             III-112


                             DRAFT

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

ELEMENT OR
NAME


SYMBOL (S)

Thorium
Mesothorlum 1
Mesotttorlum 2
Radiothorlum
Thorium X
Thoron
Thorium A
Thorium B
Thorium C
Thorium C'
Thorium C"
Thorium D
Th232
90Th
ggRa228 (MsTh,)
ggAe^^ (MsThg)
gjjTh"^ (RrfTh)
B8Ra224 (ThX>
oaRn (Tn)
MPo218 (ThA)
82Pb212 (ThB)
83Bi212 (ThC)
84Po212 (ThC'J
g,™208 (ThC"l
gjPb208 (ThD)


HALF-LIFE
ENERGY OF RADIATION
(MeV)
* \ f
r
Thorium Series
1.34 x 1010 years
6.7 years
6.13 hours
1.90 years
3.64 days
64.5 seconds
0.158 seconds
10.6 hours
60.5 min
3 x 10* second
3.1 minutes
Stable
4.20
-
4.5
BM
6.68
6.28
6.77
-
6.05
8.77
-
-
—
0.053
1.55
-
-
—
^
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
gjU238 (Ul)
gOTh^lUX,)
91Pe234 (UX2I
gjU234 
82Pb206
-------
                             DRAFT
           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
Karr-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 a 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 (HjOj)
VANADIUM RECOVERY



7
3
4

1
3
3
8
1
6


SOURCE. REFERENCE 21
                             III-114
                             DRAFT

-------
                               DRAFT
Figure 111-28. GENERALIZED FLOW DIAGRAM FOR PRODUCTION OF URANIUM,
          VANADIUM. AND RADIUM
                  MINING
                    I
               ORE TREATMENT
                    I
                 LEACHING
                    1
                LIQUID/SOLID
                SEPARATION
  ION EXCHANGE
           SOLVENT EXTRACTION
                     	I
     PATH!
     I
   PATHUI
PATHH







\


1
1
' T '
DDCfMDITATiriK











      TO
STOCKPILE
                    I
  URANIUM
CONCENTRATE
                                              I
        VANADIUM
       BYPRODUCT
        RECOVERY
TO
STOCKPILE
                              III-115


                              DRAFT

-------
                             DRAFT
 disposal technology may cause some reduction in demand.

 Table 111-27 shows uranium production for the period 1968
 through 1972, expressed in terms of both ore movement and
 UJto8  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-
 tonlum)  will become available if and when environmental
 problems of fuel recycling are resolved—partlcuarly,  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 differ-
 ently,  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  illus-
 trated in Table 111-29.

 Radium is traded from foreign sources,  but not mined,  in
 quantities of about 40 grams (or curies)  (0.14 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.

 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).   Ihe  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).
                             III-116

                             DRAFT

-------
                                  DRAFT
                    TABLE 111-27. URANIUM PRODUCTION
YEAR
1968
1969
1970
1971
1972
1973
ORE MOVEMENT
1000
METRIC TONS
5.861
6.367
6.749
5.708
6.834
6.162
1000
SHORT TONS
6,461
6.916
6,337
6,292
6,431
6.781
U3Og PRODUCTION
1000
METRIC TONS
11.244
10.664
11.732
11.167
11.727
12.032
1000
SHORT TONS
12.394
11.634
12.932
12.298
12.927
13.263
U3Og RESERVES*
1000
METRIC TONS
146
186
224
248
248
261
1000
SHORT TONS
161
204
247
273
273
277
•At $18.000 per metric ton ($16,340 par diort ton).
                   TABLE III-28. VANADIUM PRODUCTION


YEAR


1968
1969
1970
1971
1972
u.s.v2os
PRODUCTION
1000
METRIC
TONS
6,590
5.369
5.085
4,812
4,771
1000
SHORT
TONS
6,192
5.918
6,605
5,304
5,259
XOF
WORLD

46
31
27
28
26
WORLD V20g
PRODUCTION
1000
METRIC
TONS
12.119
16.892
18.337
16,883
18,136
1000
SHORT
TONS
13,359
18,620
20.213
18.610
19,990
V2OB VALUE

PER
METRIC
TON
$3,910
$5.190
$7.216
$7.887
$6.941
PER
SHORT
TON
$3.547
$4.708
$6.546
$7.156
$6.297
                        TABLE 111-29. 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
1972
METRIC
TONS
4,084
172
43
453
SHORT
TONS
4,502
190
47
499
%
86
4
1
9
                                  III-117
                                  DRAFT

-------
                              DRAFT
Metal Ores,  Not  Elsewhere Classified

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

Antimony Ores

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

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

Only slightly more than 13 percent of the antimony produced
in  1972 was  recovered from ore  being mined primarily for its
antimony content.  Nearly all of this production can be
attributed to  a  single operation which is using a froth
flotation process to  concentrate stlbnite (Sb2£3_) (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).
                             III-118


                              DRAFT

-------
                                      DRAFT
       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
S.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 LEAOt
(ANTIMONY CONTENT)
METRIC TONS
1.179
1.065
542
751
468
SHORT TONS
1.300
1.174
598
828
516
'Include! production from antimony ores and concentrate! and byproduct recovery from silver concentrate!
tByproduct produced at lead relmeriei in the United State*
                                     III-119
                                      DRAFT

-------
                              DRAFT
Figure 111-29. BENEFICIATION OF ANTIMONY SULFIDE ORE BY FLOTATION
    ROUGHER
    FLOTATION
      FROTH
        I	
     -TAILS'
                     MINING
                      ORE
                      J_
                    CRUSHING
                    GRINDING
                      I
                  CLASSIFICATION
 •TAILS
^—

SCAVENGER
FLOTATION


FINAL
TAILINGS
                           TO
                           WASTE
 CLEANER
FLOTATION
 FROTH
	I
                          FROTH
                     FILTER
                      I
                   THICKENER
                     FINAL
                  CONCENTRATE
                       F
                  TO SHIPPING
                              III-120


                              DRAFT

-------
                             DRAFT
Antimony Is also contained  in lead concentrates and is ulti-
mately recovered as a byproduct at lead smelters.  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 (Be3A12.(S16pl8))
was produced in Colorado and South Dakota.  The largest
domestic source of beryllium ore is a bertrandite (Be4S12p_7_
(OH)2} mine in the Spor Mountain district of Utah.  Domestic
beryl prices were negotiated between producers and buyers and
were not quoted in the trade press.

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

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

Platinum-Group Metal Ores

The platinum-group metal ore mining and milling industry is
defined for this document as those operations which are
Involved in the mining and/or milling of ore for the primary
or byproduct/coproduct recovery of platinum,  palladium,
Iridlum, 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
                             111-121


                             DRAFT

-------
                             DRAFT
platinum-group metals  is principally as a byproduct of copper
smelting, 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 glaciofluvlal outwash.  The other
domestic source of platinum is as a byproduct of refining
copper from porphyry and other copper deposits and from lode
and placer gold deposits, although the grade is extremely
low.

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

Beneficiation of Ores.   The mining and processing techniques
for recovering crude platinum from placers in the U.S. are
similar to those used  for recovering gold.  The bulk of the
crude placer platinum  is recovered by large-scale bucket-line
dredging, but small-scale hand methods are also used in
Columbia, Ethiopia, and (probably) the U.S.S.R.  A flow
diagram for a typical  dredging operation is presented as
Figure 111-30.

In the Republic of South Africa, milling and beneficiatlon
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.
                            III-122

                             DRAFT

-------
                           DRAFT
TABLE 111-31. DOMESTIC PLATINUM-GROUP MINE PRODUCTION AND VALUE
YEAR
1968
1969
1970
1971
1972
MINE PRODUCTION
KILOGRAMS
460.1
671.4
638.6
660.8
532.2
TROY OUNCES
14.793
21,588
17,316
18,029
17,112
VALUE
$1,600,603
$2.094,607
$1.429.621
$1,369,676
$1,267,298
       SOURCE: REFERENCE 2
                          III-123
                          DRAFT

-------
                             DRAFT
Figure 111-30. GRAVITY CONCENTRATION OF PLATINUM-GROUP METALS
                 DREDGE
                (SCREENING,
               JIGGING, AND
                 TABLING)
                 TABLING
                MAGNETIC
               SEPARATION
CHROMITE/
(MAGNETITE
                 DRYING
               SCREENING
                  IT
                 SIZING
                BLOWER
            90% CONCENTRATE
        (PLATINUM GROUP AND GOLD)

              TO SHIPPING
                 TO
                 WASTE
 TO
' SHIPPING
                ,TO
                WASTE
                            III-124


                           DRAFT

-------
                             DRAFT
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 (CeFC03_) and monazlte (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
rate-earth elements.  In 1972, approximately 10,703 metric
tons (11,800 short tons) of rare-earth oxides were obtained
in flotation concentrate from 207,239 metric tons (approxi-
mately 228.488 short tons) of bastnaesite ore mined and milled
(Reference  2  ).  Monazlte 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 1.8 metric tons (2.0 short tons) of byproduct
monazlte 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
                             III-125

                              DRAFT

-------
                            DRAFT
and countercurrent decantation.  The bastnaesite is not
dissolved by this treatment, and the concentrate is further
upgraded to 72 percent REO.  Finally, the leached product
is usually roasted to remove the carbon dioxide from the
carbonate, resulting in a product with over 90 percent REO.

Monzazite is recovered from heavy-mineral sands mined primarily
for their titanium content.  Beneficiatlon of monazite is by
the wet-gravity, electrostatic, and magnetic techniques dis-
cussed 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 no known 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.
Feasability studies continue for mining and milling facilities
for a 4,065-metric-ton-per-day (4,472-short-ton-per-day)
open-pit fluorite tin/tungsten/beryllium mine in Alaska's
Seward Peninsula which is to open by 1976.  Reserves at the
prospect area represent at least a 20-year supply.  As tech-
nological improvements in beneficiation are made and demands
for tin increase, large deposits considered only submarginal
resources, in which tin in only one of several valuable
commodities, are expected to be brought into production.

In general, crude casslterlte 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.  Cassiterlte concentrate
recovery takes place after flotation of molybdenum ore by
                            III-126

                             DRAFT

-------
                             DRAFT
 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 (FeT102) and rutile (T102) .   The United States
 is a major source of ilmenite but not of rutile.   Since 1972,
 however,  a new operation in Florida  has  been producing a
 small amount (less than approximately 2  metric tons,  or 2.2
 short tons,  per year) of 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 opera-
 tion 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_2_;  however,  some sand  deposits  yield altered
 ilmenite  (leucoxene)  containing 60 percent  or more  T102^  as
 well  as rutile  containing 90  percent  or more
The method of mining and beneficiatlng titanium minerals
depends upon whether the ore to be mined is a sand or rock
deposit.  Sand deposits occurring in Florida, Georgia, and
New Jersey, contain 1 to 5 percent TiO£ and are mined with
floating suction or bucket-line dredges handling up to 1,088
metric tons (1,200 short tons) of material per hour.  The
sand is treated by wet gravity methods using spirals, cones,
sluices, or jigs to produce a bulk, mixed, heavy-mineral
concentrate.  As many as five individual marketable minerals
are then separated from the bulk concentrate by a combination
of dry separation techniques using magnetic and electrostatic
(high-tension) separators, sometimes in conjunction with dry
and wet gravity concentrating equipment.
                            Ill-127

                            DRAFT

-------
                          DRAFT
  TABLE 111-32. PRODUCTION AND MINE SHIPMENTS OF TITANIUM
              CONCENTRATES FROM DOMESTIC ORES IN THE U.S.
YEAR
1968
1969
1970
1971
1972
PRODUCTION*
METRIC TONS
887,508
884,641
787,235
619.549
618,251
SHORT TONS
978,509
931,247
867,955
683,075
681,644
SHIPMENTS*
METRIC TONS
870,827
809,981
835,314
647,244
661.591
SHORT TONS
960,118
893,034
920.964
713.610
729.428
•Includes a mixed product containing rutile, leucoxene. and altered ilmenite.
SOURCE: REFERENCE 2
                         III-128
                         DRAFT

-------
                             DRAFT
 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, there-
 fore, 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.

 Ilmenite is also currently mined from a rock deposit in New
 York by conventional open-pit methods.  This  ilmenite/
 magnetite ore, averaging 18 percent Ti02_, is  crushed and ground
 to  a small particle size.  The ilmenite and magnetite fractions
 are separated in a magnetic separator, the magnetite being
 more magnetic due  to its greater iron content.  The Ilmenite
 sands are further  upgraded in a  flotation circuit.   Beneficia-
 tion 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
                            III-129


                             DRAFT

-------
                        DRAFT
Figure 111-31. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS
ORE FED
FROM DREDGE
\ |
VIBRATING
SCREENS
+
SPIRALS OR LAMINA
FLOWS (ROUGHERS AND CL
WET MILL
DRV MILL 1
SCRUBBER PLANT

*
DRIER
*
ELECTROSTATIC
SEPARATORS
|

SPIRALS AND/OR
TABLES
1
MAGNETIC

— -T— 1 ""
*
• 	 MAGNETICS 	 ' 	 NONMAGNETICS— 1
1 MONAZITE | ( ZIRCON
___[ 	 ' 	 1 	
TO TO 1
SHIPPING SHIPPING SHIf

	
TO
POND





"• 	 SODIUM
^ HYDROXIDE







MAGNETIC
SEPARATOR
wnuunriurTirT. — .. — ™ 	 MAGNETICS 	 1

TII c I 1 ILMENITE 1
J 1 '


, , ,
) TO
>P|NG SHIPPING
                       III-130






                       DRAFT

-------
                              DRAFT
Figure 111-32. BENEFICIATION OF ILMENITE MINED FROM A ROCK DEPOSIT
                              MINING
                               ORE
                                t
                            CRUSHING
                             GRINDING
                               ±
                          CLASSIFICATION
                               I
                             MAGNETIC
                            SEPARATION
                 •MAGNETICS-
                               I
NONMAGNETIC*
       MAGNETITE
          i
              ILMENITE
            AND GANGUE
       DEWATERER
                i
             FLOTATION
               CIRCUIT
                                  TAILINGS
                                    T
                                                     i
             THICKENER
                                    TO
                                   WASTE
               FILTER
                                                     ±
                                                    DRIER
                                                     1
                                                 CONCENTRATE
                                                  TO SHIPPING
                               III-131


                                DRAFT

-------
                            DRAFT
mining and/or milling of zirconium or for its primary or
byproduct/coproduct recovery.

The principal mineral source of zirconium (ZrS104_), is zircon
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 monazlte) which are separated
from one another by a cominatlon of electrostatic and mag-
netic 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 esti-
mated 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).  Zircon occurs with
titanium minerals In beach-sand deposits.
                            III-132

                            DRAFT

-------
                            DRAFT
                          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 subcategorlzation 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.



                             IV-1


                             DRAFT

-------
                            DRAFT
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 limitation
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, 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, loading and shipping facilities."
                            IV-2


                           DRAFT

-------
                             DRAFT
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 repre-
sented.   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 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 sub-
categories which have been made as part of these recommenda-
tions :

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

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

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

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

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

FACTORS INFLUENCING SELECTION OF SUBCATEGORIES IN ALL METAL
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.   Rationale for selection or
rejection of each as a pertinent subcategorization criterion
is included.
                             IV-3

                            DRAFT

-------
                             DRAFT
Designation  as  a_ Mine or Mill

There are  many  reasons for  recommending mine water discharge
limits different from mill  effluents  limits.  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 some instances where mine water is
used in a  process, or is discharged to a treatment system
associated with a mill,  the effluent  limitations of the mill
will prevail.   In many instances, it  may be desirable to seg-
regate mine  water from mill process water because of differing
water quality or ease of treatment with respect to water flow
volume, because of the parameters contained, or because of
parameter  concentration.  In general,  levels of pollutants
in mine waters  are lower or less complex than those in mill
process waters.   Contact with finely  divided ores (especially,
oxidized ores)  is minimal,  and mine water is not exposed to
the suite  of process  water  reagents often added in milling.
The much smaller suspended-solid loads present in mine water
may be effectively removed  in thickeners and later transported
to tailing ponds or other disposal sites, or may be settled
or treated in relatively permanent Impoundments without the
extensive  dam raising necessary  in.tailing disposal practice.
In addition, the volume  of  mine  water  is almost totally beyond
the operator's  control and  may greatly exceed mean water
flow.  Effluent volume reduction is often not a viable option
in mine water treatment.

Type £f_ 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.  Although  interception of aquifers
occurs with  both open-pit and underground mines and can create
difficulties associated  with mine dewatering, it is largely
uncontrollable  with respect to mine-operator choice of loca-
tion.  Categorization as  an open-pit or  underground mine,
although the two  types differ in the methods used for extraction
of ore, did  not  result in a usable scheme — that is, one which
reflects differences  in  water quality, ore mineralogy, or
other pertinent  factors.  Placer mines exploit unconsolidated
deposits for metal ores  that can be concentrated largely by
gravity methods.  Therefore, the mineralogy and extraction
methods become  the dominant factors in determining subcategoi.-
ization.   In addition, high suspended-solid loads can be
controlled by existing technology to levels equivalent to
existing underground  and  open-pit mine treated effluents.
                             IV-4

                            DRAFT

-------
                            DRAFT
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 treat-
ment technologies for mine waters from the above sources,
designation of the type of mine was not selected as a suitable
basis for subcategorization in the Industry.

Type of_ Processing (Beneficlation, 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 pollu-
tion 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 gener-
ated 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 impor-
tant parameters being nearly identical for many subcategories.

The practice of flotation significantly changes the character
of mill effluent in several ways.  Generally, mill water pH
is altered or controlled to increase flotation efficiency.
This, together with the fact that ore grind is generally finer
than for physical processing, may have the secondary effect
of substantially increasing the solubility of ore components.
Reagents added to effect the flotation may include major
pollutants.  Cyanide, for example, is used in several sub-
categories.  Although usage is usually low, its presence in
effluent streams has potentially harmful effects.  Oils are
also common flotation reagents which are undesirable in
effluent streams.  The added reagents may have secondary
effects on the wastewater as well, such as in the formation of
cyanide complexes.  The result may be to increase solubility
of some metals and decrease treatment effectiveness.  Some
flotation operations may also differ from physical processors
in the extent to which water may be recycled without major
process changes or serious recovery losses.
                             IV-5

                            DRAFT

-------
                             DRAFT
Ore leaching operations differ substantially  from  physical
processors  and flotation plants In wastewater character and
treatment requirements.  The use of large quantities  (in
relation to ore handled) of  reagents,  and the deliberate
solubillzation of ore components characterizes these  opera-
tions.  Wide diversity of leaching and chemical extraction
processes,  therefore, affects the character and quantities
of water quality constituents,  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  subcate-
gorizing each  major ore category.

Mineralogy  of  the Ore

The mineralogy and  host rock present greatly  determine the
beneflciation  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 beneficlatlon process adopted is based
upon the mineralogical characteristics of the ore; therefore,
the waste characteristics  of the  mine  or  mill reflect both
the ores mined and  the extraction process used.  For these
reasons, ore mineralogy was  determined to be  a primary
factor affecting subcategorization  In  all categories.

End Product

The end product  shipped  is closely allied to  the mineralogy
of the ores  exploited;  therefore, mineralogy and processing
were found  to  be more advantageous methods of subcategorization.
Two ores, vanadium  ores  and  monazite ores, are the exceptions
treated here which  were  based upon considerations of end
product or  end use.   Therefore, end product was not found
to be a suitable basis  for categorization of  the Industry
as a whole.
                             IV-6


                            DRAFT

-------
                            DRAFT
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, such as many other industry
sedments.  Therefore, climate and rainfall variations must
De 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.  Two ore categories (i.e., the uranium
and copper ore industries) make primary use of the process-
ing method, followed by the secondary factors of climatic
conditions or rainfall as the basis for subcategorization.
Treatment of large volumes of water by evaporation in many
areas of the United States cannot be utilized where topo-
graphic 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 diffi-
cult than in a semi-arid area.  Therefore, limited use was
made of climate and rainfall as secondary subcategorization
factors.

Production and Size

The variation of size and production of operations in the
industry ranges from small hand cobbing operations to those
mining and processing millions of tons of ore per year.
The size or production of a facility has little to do with
the quality of the water or treatment technology employed,
but have considerable influence on the water volume and costs
incurred in attainment of a treatment level in specific
cases.  Mills processing less than 5,000 short tons (4,535
metric tons) of ore per year in the ferroalloys industry
(most notably, tungsten) are typically intermittent in opera-
tion, have little or no discharge, and are economically mar-
ginal.  Pollution potential for such operations is relatively
low due to the small volume of material handled if deliberate
solution of ores Is not attempted.  Few of the operations are
covered by NPDES permits.  Accordingly, size or production
was used in a limited sense for subcategorization in the
ferroalloys categories but was not found to be suitable for
the industry as a whole.
                            IV-7

                          DRAFT

-------
                            DRAFT
Reagent Use

The use of reagents in many segments of the industry, such
as different types of froth flotation separation processes,
can potentially affect the quality of wastewater.  However,
the types and quantities of reagents used are a function of
the mineralogy of the ore and extraction processes employed.
Reagent use, therefore, was not a suitable basis for subcate-
gcrizatlon 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. In mines, however, mineralogy influences the
chemical nature of the wastewater.  This characteristic was
used in subcategorization of lead and zinc mines.

Water Use and/or Water Balance

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

Treatment Technologies Employed

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

General Geologic Setting

The general geologic setting determines the type of 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.
                            IV-8


                           DRAFT

-------
                            DRAFT
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
or 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.

Facility Age

Many mines and mills are currently operating which have oper-
ated 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 con-
centrating 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.   The categorization presented here is believed to
be the least number of groups having significantly different
                            IV-9

                           DRAFT

-------
                             DRAFT
water-treatment problems and water-pollution potentials.
The best technology, from a water-quality standpoint, would
be a completed closed system with all mining and processing
water being recycled.  This technology should be considered
for any new sources and for existing operations which have
the favorable processes and mineralogies, water flows and
quality requirements, location, and land availability that
will allow economic practice.

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 character-
istics 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 informa-
tion 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
                             IV-10

                            DRAFT

-------
                              DRAFT
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.

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

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

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

Examination of  the preliminary subcategorlzation 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        Older plant (1957)
     Open-pit            Mine produces effluent
     Iron formation      Processing plant has closed water
                              system
     Physical processing

Operation B
     Medium tonnage      Medium age plant (1965)
     Open-pit            Mine produces effluent
     Iron formation      Processing plant has closed water
                              system
     Physical processing
                              IV-11

                             DRAFT

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

Operation D
     Low tonnage
     Open-pit
     Natural ore
     Physical processing

Operation E
     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
Older plant (1948)
No effluent
Older plant (1953)
Mine produces effluent
Processing plant produces effluent
Medium age plant (1967)

Mine produces effluent
Processing plant has closed
     water system
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
                             IV-12

                            DRAFT

-------
                              DRAFT
 The mines visited and sampled had a 1973 production of approxi-
 mately 43,853,450 metric tons (48,350,000 short tons), or 47.5
 percent of the total United States production of iron ore.

 One of the Initial goals of this study was determination
 of the validity of the Initial categorization.   The primary
 source of the data utilized for this evaluation was information
 obtained during this study, plant visits,  and sampling program.
 This Information was supplemented with data obtained through
 personal interviews and literature review and with historical
 effluent quality data from NPDES permit applications and monitor-
 ing 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 pre-
 sently used concentrating technology.

 The milling categorized into four distinct  classes  based
 on  the type of  ore,  geographical  location,  and the  type of
 processing.

       Category  Ha.  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 and not
                     beneficiating ores of  the Minnesota
                     Biwablk formation.

      Category  lid.  Mills using magnetic separation for the
                     benefication of iron formations in the
                     Minnesota Biwabik formation.

Final Iron-Ore Subcategorization.   Based on the types of
discharges found from all mills, the first three subcategories
                            IV-13

                            DRAFT

-------
                            DRAFT
 can be  grouped  into a single segment.   Mills  in  the Minnesota
 Mesabi  Range  have demonstrated that  a  distinct subcategory
 can be  made because of hydrological  characteristics of the
 area, type of ore,  and the mode of beneficlation.

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

 II.   Iron ore  mills employing physical and/or chemical
      separation

 III.  Iron ore  mills employing magnetic and physical separa-
      tion (Mesabi  Range)

 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  subcategorles based  upon process resulted
 in grouping operations with similar water quality character-
 istics.   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-pre 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.
                            IV-14


                            DRAFT

-------
                            DRAFT
      II.  Copper mines employing hydrometallurgical processes

      III. Copper mills employing the vat-leaching process

      IV.  Copper mills employing froth flotation (areas
          where net evaporation equals or exceeds 76.2
          centimeters (30 inches) per year

      V.   Copper mills employing froth flotation (areas
          where net evaporation is less than 76.2 centi-
          meters (30 inches) per year.

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, but the froth flota-
tion  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, mine drainage is sent to tailing lagoons, although
a decrease in excess water problems can be realized in many
cases if mine water,is treated separately from mill process
water.  For this reason, the mining subcategory remains inde-
pendent of mill and hydrometallurgical beneficiating processes.

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.  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*,
                            IV-15

                            DRAFT

-------
                            DRAFT
land availability, and  topography, a preliminary scheme for
subcategorlzation 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 flota-
tion 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 cf 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.  These factors which
effect the economics  of treatment at specific facilities, if
considered as a basis for subcategorizatlon, would, if carried
to their logical completion, result in individual considera-
tion for each facility.

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

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

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

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

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

      V.   Water Balance:
           (a)  Total recycle possible
           (b)  Total recycle not possible
                            IV-16

                            DRAFT

-------
                             DRAFT
 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
 cont>iderably.   No basic differences In raw waste characteristics
 or 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 dewaterlng 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 sulflde
 to the tailing  retention area.   The only ore concentration
 process currently practiced in the lead/zinc industry is froth
 flotation.   Subcategorization based on milling  process is,
 therefore,  not  necessary.

 The  treatability  of mine wastewater is significantly affected
 by the occurrence of local  geological conditions which cause
 solubilizatlon  of undesirable metals or salts.  A common,
 and well-understood, example is acid mine drainage caused
 by the oxidation  of pyrite  (FeS2) to ferrous sulfate and
 sulfuric acid.  This oxidation requires both moisture and
 air  (oxygen  source) to occur.   The acid generated then leaches
 heavy metals from the exposed rock as 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
                             IV-17


                            DRAFT

-------
                               DRAFT
by later contact with  limestone  or  dolomitic limestone, but
the pH level attained  may  not be high  enough to cause pre-
cipitation 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 solubiliza-
tion conditions do not occur.  The  treated  effluents from
mines in this  subcategory  often  exhibit higher levels of
heavy metals in solution than untreated mine waters from
mines where solubilization conditions  do not occur.

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 subcategorlzation on this basis.

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

1.    Lead and/or  zinc mines having no solubilization potential

II.   Lead and/or  zinc mines having solubilization potential

III.  Lead and/or  zinc mills

For purposes of the subcategorization  recommended here, solu-
bilization potential in the lead/zinc mining industry is defined
as total heavy metal concentration  in  untreated mine wastewater
equal to or exceeding  2 mg/1 for the sum of the concentrations
of Cd, Cr, Cu, Pb  and  Zn.

Problems Affecting Subcategorization.  The  separation of a
specific discharge into the mine or mill subcategory is simple
and straightforward.   The  further subcategorization of mine
water based on the local geologic conditions which lead to
solubilization is, clearly, not  so  straightforward.  It is
necessary, however, because a treatment practice appropriate
for mine waters not affected by  this condition is relatively
ineffective in controlling wastes from mines in this subcate-
gory.  Indeed, treatment consisting of chemical precipitation,
followed by sedimentation,  is required to attain effluent
quality approaching the raw waste chemical  composition dis-
charged from the typical non-affected mine.  Assignment
of a mine to this  subcategory will  be  straightforward
if the pH of the mine  water is significantly below neutral
(i.e., if it is 6.0 or below).   Those  situations where the
                               IV-18

                              DRAFT

-------
                            DRAFT
mine water is alkaline are not so easily classified.   The
manifestation of a mine falling in this category is high
soluble metals (particularly, zinc) in the effluent.   The
cause, as previously discussed, is related to the geology
of the formation being mined.

In general, the type of treatment practices currently utilized
and those envisioned are not energy-intensive technologies;
therefore, energy consideration on an overall basis has
minor impact.  In specific cases, however, topographical con-
sideration may create instances where significant energy
would be expended in pumping water for treatment.

In general, it is prudent to minimize the volume of the waste
stream which must be treated.  It is, therefore, good practice
to segregate runoff from non-contaminated areas from that of
areas which will require treatment.

Gold Ores

The most important factors considered in determining whether
subcategorization was necessary for the gold ore category
were ore mineralogy, general geologic setting, type of
processing, wastes and waste treatability, water balance, and
final product.  Upon intensive background data compilation
(as discussed in Section III), mill inspections, and communi-
cations with the industry, most of the factors were found to
reduce to mineralogy of the ore (and, thus, product) and mill-
ing 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 charac-
           teristic difference between mine effluents and
           mill/mine-mill effluents is their quantitative
           and qualitative pollutant loadings.  This differ-
           ence between mines and mills makes necessary the
           application of differing waste-treatment tech-
           nologies and/or the segregation of sources for
                             IV-19


                            DRAFT

-------
                            DRAFT
           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 tech-
           nology.

      (2)  The specific beneficiation process adapted is
           based on the geology and mineralogy of the ore.
           The waste characteristics and treatabllity of
           the mill effluent are a function of the particu-
           lar 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
           treatabllity; for example, cyanide removal requires
           different technology than that used for metal
           removal.

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

Final GoId-Ore Subcategorization

On the basis of the rationale developed above and previously
discussed in the introductory portion of this section, six
                             IV-20

                            DRAFT

-------
                           DRAFT
subcategories were identified for the gold mining and  milling
industry:

      I.   Hlne(s) alone.

      II.  Mill(s) or mine/mill complex(ea) 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).

      VI.  Mill(s) where gold is a byproduct of a base-metal
                operation.  (Gold values are present in  the
                base metal concentrate and are recovered at
                the smelter.)

Silver Ores

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

Final Silver-Ore  Subcategorization

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
                            IV-21


                            DRAFT

-------
                              DRAFT
 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.

       VI.  Mill(s) or mine/mill complex(es) where silver is
                 recovered as a byproduct of a base-metal
                 operation.   (Silver is present in the  base-
                 metal concentrate(s)  and is recovered  at the
                 smelter.)

Bauxite  Ore

In the bauxite mining industry,  most  criteria for subcate-
gorizatlon 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 min-
ing 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 emit-
ting no  wastewater.

After  the  site visits to  both operating mines, it  was evident
that the mining  technique is  closely associated with the
characteristics  of  the mine drainage,  and  that mineralization
is directly  responsible for mining-technique  and  raw mine-
drainage  characteristics.   In addition,  an  evaluation of
                              IV-22


                             DRAFT

-------
                             DRAFT
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 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 discharging water.
                               IV-23


                              DRAFT

-------
                          DRAFT
      II.  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.

       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
                           IV-24


                          DRAFT

-------
                           DRAFT
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 reason-
able.  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 compon-
ents.  Reagents added to effect the flotation may include
major pollutants.  Cyanide, for example, is commonly used
and, though usage is low, may necessitate treatment.  Oils
are also common flotation reagents which are undesirable in
effluent streams.  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.

Ore leaching operations differ substantially from physical
processors and flotation plants in effluent character and
treatment requirements.  The use of large quantities (in
relation to ore handled) of reagents, and the deliberate
solubllization of ore components, characterizes these opera-
tions.  The solubllization 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+,
SOA^, 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 leach-
ing, as in the uranium industry, for example), but the limited
current activity and data available at this time do not sup-
port such a division.
                           IV-25

                          DRAFT

-------
                            DRAFT
 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  dis-
 charge 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 tails.
 On  the other hand, it  may be unattainable despite a negative
 annual balance due to  severe seasonal fluctuations,  coupled
 with soil porosity,  which renders diversion ineffective,  or
 other topographic factors.   Particular situations may  also
 exist where other environmental effects,  such as massive
 energy consumption,  make zero discharge unfeasible.  It is
 anticipated that, under the Impetus of effluent limitations
 established under PL 92-500, and the resultant pollution  con-
 trol costs,  many mills in the defined subcategories will
 choose the often less  expensive option of discharge  elimina-
 tion.

 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  beneficiatiug 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
 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 pro-
jection) .
                            IV-26

                           DRAFT

-------
                            DRAFT
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 excep-
tions.  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.

     IV.   Mill(s) where mercury is recovered as a byproduct
                of base- or precious-metal concentrates.
                (The recovery takes place at a refinery or
                smelter.)

Uranium. Radium, and Vanadium Ores

The primary factors evaluated in consideration of Subcate-
gorization of the uranium, radium, and vanadium mining and
ore dressing Industry are:  end product,  type of processing,
ore mineralogy, waste characteristics, treatability of
wastewater, and climate, rainfall, and location.  Based upon
an intensive literature search,  plant inspections, NPDES permits,
and communications with the industry, this category is
categorized by milling process,  mineralogy (and, thus,
product), and climatic factors.   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 hydromecal-
lurgy.   Such point discharges as might occur in milling processes
                            IV-2 7

                           DRAFT

-------
                              DRAFT
 (i.e., the production of concentrate) are expected to con-
 tain a variety of pollutants that need to be limited.  Mining,
 for the ores, is expected to lead to a smaller set of con-
 taminants.  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-radio-
 active vanadium ores; and (4) The processing of the ores of
 the third subcategory to yield vanadium concentrate.

 A careful distinction will be drawn  between the radioactive
 processes and the vanadium industry  by including in the
 former all operations within SIC 1094 that are licensed by
 the U.S.  Nuclear Regulatory Commission (NRC, formerly AEC,
 Atomic Energy Commission)  or by agreement states.   The agree-
 ment states,  Including the uranium producing states of Colorado,
 Texas, and Washington, have assumed  all licensing,  record-
 keeping,  and  inspection responsibilities for radioactive
 materials from the U.S.  government upon establishing  regula-
 tions regarding radioactive materials that are at  least as
 stringent as  those of the  NRC(AEC).   The licensing requirements,
 as  set forth  in the code of Federal  Regulations, Title 10,
 Part 20 <10CFR20),  already constitute restrictions  on the
 discharge of  radioactive nuclides  that form a minimum (larg-
 est-discharge)  standard  superseding  any less-stringent regu-
 lations.

 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  ferroalloy-
metal  ore  mining and dressing,  in  another  portion of  these
 guidelines.  The vanadium  subcategories  are  summarized there
as members  of  the mining and  hydrometallurgical process sub-
categories.

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
                              IV-28

                             DRAFT

-------
                            DRAFT
mills into fewer subcategories.  This simplification is based
on the observations discussed below.

Raw wastewaters from mills using acid leaching remain acid
at the process discharge (not to be confused with a point
discharge), retain various heavy metals, and are generally
not suitable for recycling without additional and specialized
treatment.  Those from the alkaline leach process are normally
recycled in part, since the leach process is somewhat selec-
tive for uranium and vanadium, and other metals remain in the
solid tailings.  At one time, It was expected that mills
using solvent exchange would have a radically different raw-
waste character due to the discharge of organic compounds.
The fact that mills not using solvent exchange often process
ore that is rich In organics make this distinction less
important.  As a result, a distinction 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 wastewater
characteristics, they are not believed to Justify further
subcategorization.  For example, there are some uranium/radium
ores that contain molybdenum and others that do not.  Efflu-
ent 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 wastewater characteristics but also
(a) differences in the volume of wastewater that must be
stored and managed in a zero-effluent condition and (b) dif-
ferences 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 con-
sidered relative to present and future exploitation of uran-
ium reserves In the United States.  Under the present defi-
nition of point discharge by the U.S. Environmental Protection
Agency (40 CFR125, para. K of the Federal Register), seepage
into the ground (and evaporation) is not counted as contri-
buting to a point discharge.  With this factor taken into
consideration, and in view of the observation that economi-
cally exploitable uranium reserves are found in arid climates
                             IV-29

                            DRAFT

-------
                            DRAFT
for geological/chemical reasons, no point discharges are
needed to manage the raw wastewater from most current mining
and ore dressing operations in the uranium industry.  In
addition, other operations that now discharge wastewater
plan to.terminate their discharges within a year or two.  To
make possible a zero discharge from a sealed mine/mill pond,
the annual evaporation of water must exceed the annual pre-
cipitation of water.  Since control of discharges to ground
water may become the goal of future state or federal legis-
lation, seepage should not in the future be counted on to
remove effluents.

There will be fluctuations in rainfall and evaporation from
year to year that may result in temporary accumulation of
excess water (or, conversely, in low levels) in wastewater
evaporating ponds.  It is common engineering practice to
accommodate such fluctuations up to but not including the
abnormal "storm" that is observed only once every 25 years
or once every 100 years.  In view of the useful life of mines
and mills, the integrated fluctuations short of those caused
by a 25-year storm should be stored in the holding ponds of
a zero-discharge operation without overflow.  Beyond this
level, no fluctuations should result in progressive pond
failure, including at least those due to any documented
storm and also those due to the 100-year storm.

Ore characteristics were considered and, within a subcategory,
cause short-term effect on wastewater characteristics that
does not justify further subcategorization.  Waste character-
istics were, as described above, considered extensively, and
it was found difficult to distinguish whether the acid/alkaline
leach distinction is based on process, mineralogy, waste
characteristics, or treatability of wastewater, since all
are interrelated.  Vanadium operations which are not extracting
radioactive ore or covered under government licensing regula-
tions (NRC or agreement states)  are subcategorized in the
ferroalloys section.
                             IV-30

                            DRAFT

-------
                           DRAFT
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 not concentrate) ores
                of uranium, radium, or vanadium under NRC
                (formerly AEC) or Agreement State license.

     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 yeild concentrates by alkaline
                leaching only.

Problems Involved in Subcategorization.   Those milling
operations surveyed which currently have a point discharge
either are changing the extraction process employed or
making plans to attain zero discharge.  All of these opera-
tions are In arid areas.  As uranium ore in the eastern U.S.
is exploited however, present methods of obtaining zero
discharge may have to be reexamined because of the humid
climate expected to be encountered.
                            IV-31

                          DRAFT

-------
                              DRAFT
 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 o* 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

 Antimony mining and milling are practiced  at two locations
 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.

 Final Antimony-Ore  Subcategorization

 The antimony ore mining and dressing  Industry  has  been separated
 into three 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.

     III.   Mill(s)  or mine/mill  complex(es)  obtaining antimony
                 as  a byproduct of a base-  or precious-metal
                 milling operation;  antimony  present in the
                 base- or precious-metal concentrate is recovered
                 at  a smelter or  refinery (antimony extraction
                 plant).

Beryllium Ores

Beryllium mining and  milling in  the United States are repre-
sented by one operating facility.  Therefore, subcategorization
                              IV-32

                             DRAFT

-------
                           DRAFT
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 beneflclatlon.

Rare-Earth Ores

Rare-earth ores currently are obtained from two types of
mineralogies:  bastnaesite and monazite.  Monazlte is an
ore both of thorium and of rare-earth elements, such as
cerium.  The subcategorization which follows is based pri-
marily upon division into mines and mills, as well as on
the type of processing employed for extraction of the rare-
earth elements.

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

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

    III.   Mlll(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).
                           IV-33

                           DRAFT

-------
                            DRAFT
Tin Ores

A single operating location currently produces tin as a
byproduct of molybdenum mining and beneficlatlon.  Placer
deposits of tin are found in the world and could be exploited
if discovered in the U.S.  Therefore, a single subcategory
for miuing and one subcategory for milling are listed:

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

     II.   Mill(a) or mine/mill complex(es) using gravity
                methods, flotation, magnetic separation
                and/or electrostatic 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, llmenite, 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
                methods used In conjunction with electrostatic
                and/or magnetic methods for the primary or
                byproduct recovery of titanium minerals.
                              IV-34


                             DRAFT

-------
                           DRAFT
Zirconium Ores

Zirconium ±s obtained from the mineral zircon in conjunction
with dredging operations.  Mo additional subcategorizatlon
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 electro-
               static 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
subcategorles, a summary of categories and subcategorles
recommended for the ore mining and dressing industry is
presented here In Table IV-1.
                            IV-35

                            DRAFT

-------
                            DRAFT
TABLE IV-1. SUMMARY OF INDUSTRY SUBCATEGORIZATION RECOMMENDED
CATEGORY
IRON ORES
COPPER ORES
LEAD AND ZINC ORES
GOLD ORES
SILVER ORES
BAUXITE ORE
FERROALLOV ORES
MERCURY ORES
URANIUM. RADIUM
& VANADIUM ORES
Q ANTIMONY ORES
UJ
iL
5 BERYLLIUM ORES
U
u PLATINUM ORES
jjj RARE-EARTH ORES
IU
H
o 	
* TIN ORES
Ul
O
_j
< TITANIUM ORES
3
ZIRCONIUM ORES
SUBCATEGORIES
MINES





MILLS

MINES
Physical/Chemical Separation
Magnetic and Physical Separ anon (Mesabi Range)
Open-fit. Underground. Stripping
Hydrometellurgical (Laaehing)
Vat Leaching
Flotation
Process
Net Evaporation 2,76 2 cm I3O in I/year
Net Evaporation < 76 2 cm (30 in I/year
No Solubilmtion Potinlial
Solubiliiation Potantial
MILLS
MINES
MILLS
Cyanidation Process
Amalgamation Process
Flotation Process
Gravity Separation
Byproduct of Base Maial Operation
MINES
MILLS
Flotation Process
Cyanidation Process
Amalgamation Process
Gravity Separation
Byproduct ol Base Metal Operation
MINES
MINES
MILLS
< 5.000 mi
itnc tons (5.512 short tonsl/year
> 5.000 metric tons/year by Physical Processes
> 5.000 m
Btric tons/year by Flotation
Leaching
MINES
MILLS
Gravity Separation
Flotation Process
Byproduct of Base/Precious-Metal Operation
MINES
MILLS
Acid or Acid/Alkaline Leaching
Alkaline Leaching
MINES
MILLS
Flotation Process
Byproduct ol Bese/Procious Metal Operation
MINES
MILLS
MINES OR MINE/MILLS
MINES
MILLS
Flotation 01 Leaching
Dredging or Hydraulic Methods
MINES
MILLS
MINES
MILLS
Electrostatic/Magnetic and Gravity/Flotation Procos&o
Physical Processes with Dredge Mining
MILLS OR MINE/MILLS
                              IV-36





                             DRAFT

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

                          DRAFT

-------
                            DRAFT
Mine Water

The effluent 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 industry, process water is present
only in placer operations operating by gravity methods, in
hydraulic mining, in washing operations, and in dust control.
Therefore, water is not normally utilized in the actual mining
of ores.  It is an unwanted natural feature or problem that
presents problems to much of the mining industry.  It enters
mines by ground-water infiltration and surface runoff and
becomes polluted by contact with materials in the host rock,
ore, and overburden.  The polluted mine water then requires
treatment before it can be safely discharged into the surface
drainage network.  These effluent quantities of ore mines
thus are unrelated, or only indirectly related, to production
quantities, except as noted above.  Therefore, raw waste
loadings are expressed in terms of concentration rather than
units of production in many of the ore categories discussed
in Section IV.

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

Because mine drainage with pollutants does not
necessarily cease with mine closure, a decision must be
                             V-2

                           DRAFT

-------
                           DRAFT
made as to the point at which a mine operator has fulfilled
his obligations and responsibilities for a particular mine
site.  This point will be further discussed in Section VII,
"Control and Treatment Technology."
SPECIFIC WATER USES IN 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
                                  vater

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

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

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

Noncontact Cooling Water

The largest use of noncontact cooling water in the ore mining
and dressing industry is for the cooling of equipment, such
as crusher bearings, pumps, and air compressors.
                            V-3

                          DRAFT

-------
                             DRAFT
Wash Water

Wash water comes  into direct contact with either the raw
material, reactants, or products.  An example of this type
of water usage is ore washing to remove fines.  Waste efflu-
ents can arise from these washing sources because the resul-
tant solution or  suspension may contain dissolved salts,
metals, or suspended solids.

Transport Water

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

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 contami-
nated 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.
                              V-4

                            DRAFT

-------
                            DRAFT
PROCESS WASTE CHARACTERISTICS BY ORE CATEGORY

Iron Ore

The nature and quantity of pollutants discharged in wastewater
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 depen-
dent on the deposit mined and the substrata through which
the water flows prior to discharge.

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

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

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

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

In most casea, 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 wastewater 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
                             V-5

                           DRAFT

-------
                             DRAFT
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.

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.  The principal contaminants
of 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)  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
wastewater.  Some of these are derived from the ore itself,
and others are added during processing.  Still others are
not intentionally added but are inadvertent and inherent
contributions.
                              V-6

                           DRAFT

-------
                            DRAFT
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-
tue-'ts of wastewater.  Oil and grease from machinery and
equipment may also contaminate the water.

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

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

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

       (1)  Solid  loadings  of 10  to 50  percent (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 flo-
ation  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  indi-
cation of  the presence of  organic  reagents,  but  no definitive
information  is related by  these parameters.
                              V-7


                            DRAFT

-------
                             DRAFT
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 concen-
trations of particular parameters tend to associate with a
particular mining area or ore body.  For example, the dis-
solved 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 benef iciation of iron-containing minerals, as much
as 27.2 cubic meters of water per metric ton  (7,300 gallons
per long ton) and as little as 3.4 cubic meters of water per
metric ton (900 gallons per long ton) of concentrate may be
used.  The average amount of water per metric ton of ore
produced is approximately 11.8 cubic meters  (3,200 gallons per
long ton) .  Most processing water in beneficiation operations
is recycled to some extent.  The amount of recycle is depen-
dent 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 divalent 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 impound-
ment systems.
                               V-8

                            DRAFT

-------
                              DRAFT
    TABLE V-1. HISTORICAL CONSTITUENTS OF IRON-MINE DISCHARGES
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cr
Cu
Fe
Pb
Mg
Ho
Ni
Na
1UW
win
Zn
Chlonda
Cyanide
CONCENTRATION Img/W
BEFORE TREATMENT
MIN
1.000
140.0
0.200
5.00*
1.800
0.003
0.003
0.001
0.001
0.060
0.001
0.020
0.002
0.003
0.023
0.001
0.001
1.000
0.010
MAX
5.000.0
1380.0
36.0
8.40*
9.000
0.350
256.0
0.010
1.000
178.0
0.100
118.0
2.000
0.100
15.0
18.0
8.0
120.0
0.02
AVG
371.51
436.18
6.470
7.45"
4.511
0.066
85.39
0.007
0.167
13.3
0.018
39.35
1.001
0.024
7.511
2.462
1.869
27.143
0.013
NO.
19
17
10
IB
9
7
3
9
12
14
9
3
2
6
2
14
9
14
4
AFTER TREATMENT
MIN
1.000
100.0
0.026
6.800*
0.400
0.007
0.002
0.010
0.005
0.008
0.008
0.008
-
0.010
.
0.001
0.010
0.900
0.005
MAX
30.0
1,090.0
42.0
8.500*
20400
0.350
0.158
0.010
0.370
2.100
0.100
0.029
•
0.075
-
6.900
0.340
180.00
0.020
AVG
10.693
390.10
12.116
7.652*
4.313
0.131
0.045
0.010
0.120
0446
0.023
0.017
-
0.023
•
1.720
0.185
33.225
0.011
NO.
27
20
14
21
16
9
4
6
10
11
8
3
-
5
-
11
5
20
4
•Value in pH units
     TABLE V-2. HISTORICAL CONSTITUENTS OF WASTEWATER FROM
               IRON-ORE PROCESSING
PARAMETER
TSS
TDS
COO
PH
Oil and Grease
Al
Ca
Cu
Fe
Pb
Ni
Mn
Zn
Chloride
Cyanide
CONCENTRATION (mg/£)
BEFORE TREATMENT
MIN
1.20
0.500
0.200
5.000*
0.030
0.030
5S.O
-
0.200
0.100
0.010
0.007
0.006
1.000

MAX
9.999.0
356.0
36.0
8.300*
40400
5.000
250.0
-
10.0
5.0
0.050
20.0
10.0
110.0

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

NO.
11
10
7
12
8
6
£
•
9
3
3
9
5
11

AFTER TREATMENT
MIN
0400
0.300
0.200
6.000*
0.100
0.009
82.0
0.010
0.050
0.045
0.010
0.016
0.010
0.350
0.008
MAX
200.0
1,090.0
90.0
8.300*
90.0
0.270
181.0
0.450
1.610
0.250
0.200
2.100
0.115
180.0
0.020
AVG
25.133
393.27
19.518
7.259*
12.0
0.107
131.5
0.230
0.453
0.111
0X187
0.529
0.056
42.875
0.013
NO.
15
16
12
16
13
8
2
2
10
4
3
10
4
16
4
'Value in pH units
                               V-9

                             DRAFT

-------
                                  DRAFT
        TABLE V 3. CHEMICAL COMPOSITIONS OF SAMPLED MINE WATERS



PAflAMETEH


pH
Alkalinity
COD
TSS
TOS
Conductinly
Total ft
Diuolrad F«
M»
Sulfate
CONCENTRATION (nq/ll IN EFFLUENT FROM MINE



o
£
73'
204
274
2
455
440'
004
<002
021
85


a
jj
i
72'
482
2
505
400*
<002
<002

-------
                              DRAFT
 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 ±s
 gixen in Figure V-l.

 Process  operations  generally recycle high percentages of their
 water.   Water in the  plant process is  used to wash and trans-
 port the ore through  grinding processes.   After separation
 of  the concentrate, the  tailings are discharged to a tailing
 pond, where  the coarse and fine waste  rock particles settle
 (Figure  V-2) .   Clarified water  is returned to be used in further
 processing,  and a portion  is discharged to receiving waters.

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

 Process  Descriptions

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

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

Crude magnetic  taconite is mined, mainly  from the lower
cherty member  of the Minnesota Biwabik  formation, by con-
ventional open-pit methods and then milled to produce a
fine magnetite.  The fine magnetite from  the mill is
agglomerated in a grate-kiln system to  produce approximately
                               V-ll

                              DRAFT

-------
                             DRAFT
       Figure V-1. FLOW SCHEME FOR TREATMENT OF MINE WATER
                             MINE WATER
                         SEDIMENTATION BASIN
                      SETTLED
                       SOLIDS
                        TO
                       WASTE
           CLARIFIED
           EFFLUENT
         TO RECEIVING
            WATERS
                                       TO PROCESS
                                         WATER
  Figure V-2. WATER FLOW SCHEME IN A TYPICAL MILLING OPERATION
                              WATER
       TO
STOCKPILE
PROCESS
PRODUCT
PROCESS PLANT
                COAGULANT
                             TO
                            WASTE
             n
                                         PROCESS
                                         TAILING
                                                  RECYCLE (80-97%)
                                    SETTLED    CLARIFIED
                                    SOLIDS    EFFLUENT
             TO RECEIVING
               WATERS
                              V-12

                            DRAFT

-------
                             DRAFT
2.64 million metric Cons (2.6 million long tons) of oxide pellets
annually for blast-furnace feed.

The mine, mill, and pelletizing plant are located on a large site
controlled by the operating company, with 8094 hectares
(20,000 acres) utilized at present.  An initial tailing pond
of 405 hectares (1000 acres) has been filled.  A second 1,619-
hectare (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-3.  Settling
basins are used to contain the water before it is discharged
to two lakes.

Chemical analysis of the mine discharge water after settling
shows the following chemical constituents:

          Parameter                Concentration (mg/1)

          pH*                                7.4
          TSS                               17
          IDS                              281
          Iron (total)                       0.10
          Iron (dissolved)         less than 0.02
          Manganese                less than 0.02
          Sulfate                           36
          *  value in pH units

The mill water system is a closed loop.  Plant processes use
204 cubic meters per minute (54,000 gpm), with 189 cubic
meters per minute (50,000 gpm) returned from the 91.4-meter
(300-foot) diameter tailing thickener overflow and 15.1
cubic .meters per minute (4,000 gpm) returned from the tailing
pond or basin.  The tailing thickener receives waste or tailings
in a slurry from the concentrate pellet plant.  A nontoxic,
anionic polyacrylamide flocculant is added to the thickener
to assist in settling out solids.  Tailing thickener under-
flow is pumped to the tailing basin.

Rotary machines are used in the mine to prepare blast holes
for the ammonium nitrate-fuel oil (ANFO) and metallized slurry
blasting agents.  Electric shovels are used to load the
broken ore into 100-ton-capacity diesel/electric trucks for
haulage to the primary crusher.
                               V-13

                              DRAFT

-------
                      DRAFT
V 3. WAI tR BALANCE FOR MINE/MILL 1105  (SEPTEMBER 1974)
3 4 mj/min (900 gpntl
\
(INTERMITTENT)
>
3 SETTLING BASINS ^X
(IN SERIES) J
\

^ CREEK ^
i
Q LAK
2 PUMPS
C 11 4 m /mm
(3.000 gpm) EA
~~ 	 s^llNTtRMI TTENT)
LL-^
p-
MINE OEWATERING
\
17 to 32 m3/mm (4.500 to 8.500 gpm)

MAX. 8.23 m3/mm (2.200 gpm)
(INTERMITTENT)
PLANT STORAGE TANK
i




15.1 m3/mm (4.000 gpm)
i
PLANT PROCESSES
1


204 m3/mm (54.000 gpm)
' 189 m3/min
TAILINGS THICKENER
\
^TAILIN
15 1 m3/mm
t
BPON^V^

(50.000 gpm)


4.000 gpm)
1 m3/mm (4.000 gpm)


\
L
{ 3 SETTLING BASINS^N
<^^ (IN SERIES) J
\
17 to 32 m3/mm
(4.600 to 8.500 gpm)
^" LAKE 2 ^
                       V-14





                      DRAFT

-------
                             DRAFT
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
14-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
                              V-15

                             DRAFT

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

           I
                            CYCLONE CLASSIFIER
                         OVERSIZE
               UNDERSIZE

                   t
         SCREEN
                                          WET MAGNETIC COBBING
    OVERSIZE  UNDERSIZE

       I         t
                   I
              CONCENTRATE
    I
   TAIL
            DRY MAGNETIC ROUGHER
             TAIL
             i
CONCENTRATE
     I	
     DRY MAGNETIC SCREENING
  MIDDLING
 	I
                                      WET SECONDARY
                                      GRINDING MILLS
                                      HYDROCYCLONES
                                     UNDERSIZE  OVERSIZE
            TAIL
               i
                                    HYDROSEPARATION
                               CONCENTRATE
                                   i
                         i
                        TAIL
                         I	
                       WET FINISHER MAGNETIC SEPARATION
                       I
                  CONCENTRATE
                         TAIL
          J
                      TO
                     PELLET
                     PLANT
                    TAILING THICKENER
                                       UNDERFLOW
                                  I
                              OVERFLOW
             TO
           WASTE
                  TO
              TAILING POND
     TO
REUSE WATER
                            V-16


                          DRAFT

-------
                           DRAFT
in closed circuit with screens to return undersize material
to the drum and to control the green ball size.

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

Green balls are dried in an updraft and downdraft section
of the grate.  Dried balls then pass through a preheat section
on the grate.  The magnetite begins to oxidize, and the
balls to 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  (an iron ore that contains moisture) 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
an average rate of 7,086 cubic meters per day  (1,872,000 gpd)
and at a maximum 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 45 cubic meters
(]2,000 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.
                              V-17

                            DRAFT

-------
                             DRAFT
    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/l ) 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
                              V-18
                             DRAFT

-------
                             DRAFT
Mining is carried out by conventional open-pLr methods.
Ammonium nitrate explosives are used in bLat.t-i.ng.  Shovels
load the ore into trucks for transport to the plant.

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

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

Material from the heavy-media surge pile is split into
fractions which are 3.2 cm by 1.6 cm  (1.25 inches x 0.63 inch)
and 1.6 cm x 0.64 cm (0.63 inch by 0.25 inch).  Both fractions
go to identical sink/float treatment  in a ferrosilicon
suspension.  Float rejects or tailings from the heavy  sus-
pension treatment are trucked to a stockpile.  Concentrates
go directly to a railroad loading picket.  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 lakes.  The 58-percent-iron heavy-media
concentrate serves as blast-furnace feed.  The 58-percent-
iron jig concentrate is later sintered at the steel plant
before entering the blast furnace.
                               V-19

                              DRAFT

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

                                          I
                                       CRUDE ORE
                                       (J7» IRON)


                                          *
           DOUBLE DECK
             SCREEN
             --
           OOUBLE DECK
              SCREEN
      > 16 2 cm
      06 m)

        I	
10 2 la 16 2
 (4lo6in)
                                                  *•* f*  42% Ft
                               CONE
                              CRUSHER
         TO HOCK REJECTS
            STOCKPILE
                                         >064«n
                                         OOJSin)
                                    < 064 em
                                    k026m)
1

HEAVY MEDIA
SURGE PILE
                    0 6« to 3 2 cm 10 25 u IX in )-
                                                              <064cm
                                                              KOJSml
FLOAT REJECTS
       \

 TO FLOAT REJECTS
    STOCKPILE
              TO
             TAILING
             POND
       RECYCLE TO
       HEAVY MEDIA
       SEPARATORS
CONCENTRATES TO TRANSPORTATION
     I- 32% OF CRUDE ORE)
                                                V-20

                                              DRAFT

-------
                            DRAFT
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 consti-
tuents appear in the tailing deposit.  The concentration plant
processes approximately 21,000 metric tons (20,700 long tons)
per day of low-grade hematite at 35.5 percent iron to produce
approximately 9,850 metric tons (9,700 long tons) per day of
concentrated ore at 65.5 percent iron.  The remaining 11,200
metric tons (12,346 short tons), at approximately 10 percent
total iron, are discharged to the tailing basin.

Mine water is currently pumped from the actively mined pit
and discharged directly.  The chemical constituents of the
discharged water are given in Table V-6.

Water in the concentration process is utilized at a rate of
114 cubic meters (30,000 gallons) per minute.  Ore is first
ground to a fine state  (80 percent less than 325 mesh) and
the argillaceous slime materials removed by wet cycloning.
A simplified flow scheme is included in Figure V-6.  Subse-
quently, the concentrated ore is floated using tall oil -
fatty acid.  The flotation underflows are discharged to a
tailing stream, which is discharged directly to a 385-hectare
(950-acre) tailing basin.  Approximately 80 percent of the
water from the tailing pond is returned to the concentrating
plant as reuse water (untreated) .  The remaining 20 percent
is discharged, after treatment, to a local creek.  This dis-
charged wastewater is first treated with alum, then with a
long-chain polymer to promote flocculation.  It then passes
to a 8.5-hectare (21-acre) pond, where the flocculated particles
settle.  The concentration of chemical parameters and the
waste loading in this discharge are given in Table V-7.

Copper Ore

Frequently, discharged wastes encountered in the copper ore
mining and dressing industry include waste streams from
mining, leaching, and milling processes.  These waste
streams are often combined for use as process water or
treated together for discharge.  Other wastes encountered
in 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.
                              V-21

                            DRAFT

-------
                       DRAFT
TABLE V-6. CHEMICAL CHARACTERISTICS OF DISCHARGE WATER
          FROM MINE 1108
PARAMETER
PH
Alkalinity
COD
TSS
TDS
Total Fe
Dissolved Fe
Mn
Sulfate
CONCENTRATION
(mg/£ )
7.2*
118
9.0
2
440
1.3
0.04
0.054
33.2
       •Value in pH units
                        V-22

                      DRAFT

-------
                                    DRAFT
Figure V-6. SIMPLIFIED CONCENTRATION FLOWSHEET FOR MINE/MILL 1108
                                  MINING
                                     T
                                 CRUDE ORE
                                      24.500 metric tont
                                      (20.700 long tons)
                                      per day
                                REGRIIMOING.
                           FLOTATION. THICKENING.
                              AND FILTRATION
                                    I
                                SECONDARY
                               CONCENTRATE
                                   i
                                PEL..ETIZING
                                OPERATION
                                 PELLETS
                                    t
                                                            9 3 m3/min (5 5 cfs)
                                                         17 m3/min (10 cfi)
                                                        41 6 m3/min (24 5 cfi)
                                                        14 4 m3/rnin (3 5 cfs)
17 m-Vrnin (10 cfs)
                                                        0 85 m3/min (0 5 cfs)
                               TO STOCKPILE
             8 1% SOLIDS
   100 m3/mm (59 cfs)

                 Y
            TO TAILINGS
                                       V-23


                                     DRAFT

-------
                            DRAFT
    TABLE V-7. CHARACTERISTICS OF MILL 1108 DISCHARGE WATER
PARAMETER
pH
AllMlmlly
COD
TSS
TDS
TotilF*
DaaohwIF*
Mi
SulUM
PROG RAM SAMPLE
CONCENTRATION
Ima/D IN
WASTE WATER
7.1*
82.0
22.6
10
16O
2.06
O93
005
5
WASTE LOAD
in g/nwlric ton
(Ib/lhort ton)
PRODUCT
-
213 (O42I
77 « (01SI
3.4 100071
660 (108)
7 OS (0 013)
32 (0008)
0 17 (0 0003)
172 (0.034)
10MONTH AVERAGES
AVERAGE
CONCENTRATION
Iml/tl
70-
-
-
86
-
-
078
066
—
WASTE LOAD
in o/iMtric ton
lib/Ikon ton)
PRODUCT
_
-
-
20.7 10040)
-
-
183(00036)
1 66(00031)
-
HIGH
CONCENTRATION
lino/I)
79-
-
-
S3
-
-
360
680
-
LOW
CONCENTRATION
Img/ll
66*
-
-
1
-
-
001
001
-
•WlminpHunili
                             V-24





                            DRAFT

-------
                             DRAFT
            I. Mining:
               a. Cooling
               b. Dust control
               c. Truck washing
               d. Sanitary facilities
               e. Drilling

           II. Hydrometallurgical processes associated with
               mining:  Dump, heap, and in situ leaching
               solutions.

          III. Milling Processes:
               a. Vat leach
                    1. crusher dust control
                    2. Vat leach solution
                    3. Wash solutions
               b. Flotation
                    1. Crusher dust control
                    2. Carrier water for flotation

Copper Ore Mining.   Most of the domestic copper is mined in
low-grade ore bodies in the western United States.  All mining
and milling activities adjust to the type of copper minerali-
zation 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 block-
caving or open-pit methods.  The choice of method is deter-
mined 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 over-
burden 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.
                               V-25

                             DRAFT

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

Chalcocite
Chalcopyrite
Bornito
Covell ite
Enargite
COMPOSITION
OCCURRENCE*
SULFIDES
Cu2S
CuFeS2
Cu5FeS4
CuS
Cu3A$S4
SW.NW. ••
SW. NW. ••
NW,SW
NW,SW
NW
OXIDES
Chryiocolla
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 11 Cu
NC.SW"
 *SW = Southwest U.S.
 NW = Northwest U.S.
 NC = Northcentral U.S.

"Major minerals
                     V-26

                  DRAFT

-------
                                DRAFT
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
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
or as oxide ore, which does not concentrate economically
by flotation.

Water Sources and Usage.   In the mining of copper ores,
water collected from the mines may originate from subsurface
drainage or seepage 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 and seepage must be pumped from the
mines.  Table V-9 lists the amount of mine water pumped from
selected mines and the ultimate fate of this wastewater at
surveyed mines.  Open-pit mines pumped 0 to 0.27 cubic meter
per metric ton (0 to 64.7 gallons per short ton) of ore produced,
while underground mines pumped 0.008 to 3.636 cubic meters
per metric ton (1.91 to 871 gallons per short ton) of ore
produced.

Solid wastes produced are summarized in Table V-10 as metric
tons (or short tons) of waste (actually, overburden and
wastes) per metric ton (short 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.
                                V-27

                                DRAFT

-------
                          DRAFT
Figure V-7. WASTEWATER FLOWSHEET FOR PLANT 2120-B PIT
 NATURAL DRAINAGE.
     SEEPAGE. AND
       RUNOFF
         i
       MINING
      (DRILLING,
      BLASTING,
         AND
      LOADING)
                        ORE
                16,560,000 metric tons/year
               (18.250,000 short tons/year)
                              TO
                              MILL
       EXCESS
     MINEWATER
                o
           0.06 m /metric ton
           (14.4 gal/short ton)
  c
POND
LIME PRECIPITATION
           0.06 m3/metric ton
           (14.4 gal/short ton)
    DISCHARGED
                            V-28
                          DRAFT

-------
                                       DRAFT
TABLE V-9. MINE-WATER PRODUCTION FROM SELECTED MAJOR COPPER-PRODUCING
            MINES AND FATE(S) OF EFFLUENT
MINE
2101
2102
2103
2104
2107
2108
2109
2110
211.1
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
0270
0.008
NE.
0.086
N/A
N.E.
N.E
N.E
N.E
0.015
40 5 (avg|T
1769
0.030
0886
0014
0.654
0.486
0170
0034
0075
N.E.
gal/short ton
ore produced
64.7
1 85
N.E
20.6
N/A
N.E.
NE
N.E.
NE
35
9.715 0
-------
                                    DRAFT
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
38.32 1.250*
21.532.400>
57.213,000
22,129,279'
0 (UG)
12,566.400*
26.700.000
115.000
9.420.000
50.000 (UG)
19.773.594
12,000,000*
20.000 (UG)
35,557.000*
0(UG)
37.063.000*
91.200IUG)
36.500.000*
0(UG)
97.500.000*
12.000.000*
16.909.000
short tons
42,241,897
23.735.379
63.066.462
24.393,325
0
13352,068
29.431.677
126,766
10.383.760
56.116
21.796.630
13.227.720
22^)46
39,194,836
0
40354.915
100,531
40.234,315
0
107,475,220
13.227.720
18,638,959
MILL ORE
metric tons
7.934.320
8.782.600
15,407,000
8,10.1,784
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
8384,136
38,300,000
2.172,000
8.722,000
(1973)
short tons
8,746.080
9.681.148
16,983,290
8330.678
N/A
4,328,771
1304371
4310.663
1.798370
771,617
11,401363
158,427
572.753
13330393
1.472.274
20,238.411
24.222.111
28.362.436
9.793.072
42.218.473
2.394.217
9.614.348
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
                                      V-30


                                     DRAFT

-------
                             DRAFT
Air quality control within open-pit nines 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.

Wastewater 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 under-
ground shafts, either in use or abandoned on the property.
This source of water is valuable and is used for other
copper-producing processes.  In contrast, mine water in Utah,
Montana, Colorado, Idaho, Oklahoma, Michigan, Maine, and
Tennessee—especially, in underground mines—is often unwanted
excess, which must be disposed of if reuse in other processes
(such as leaching and flotation) is not possible.

The primary chemical characteristics of mine waters are:
(1) occasional presence of pH of 2.0 to 9.5; (2) high dissolved
solids; (3) oils and greases; and (4) dissolved
metals.  Often, mine water is characterized by high sulfate
content, which may be the result of sulfide-ore oxidation
or of gypsum deposits.  Mine water—particularly, acid mine
water—may cause the dissolution of metals such as aluminum,
cadmium, copper, iron, nickel, zinc, and cobalt.  Selenium,
lead, strontium, titanium, and manganese appear to be
indicators of local mineralogy and are not solubilized
additionally by acid mine water.

Handling of Mine Water.   As shown in Table V-9, mine
waters are pumped to leach and mill operations as a water
source for those processes whenever possible.  However, four
                             V-31

                             DRAFT

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

                               SELECTED COPPER MINES (Sheet 1 of 4)
O
30    <
>    I
•n    W
PARAMETER
Flow
pH
IDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
S«
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2119
CONCENTRATION
(mg/ ;i
42.0135m3/day
964*
544
8
1
5
<10
02
OS
< 005
< 0003
< 050
< 007
< 005
< 02
380
< 005
< 005
< 010
< 02
013
00008
< 005
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
752 m3/1000 metric tons
964'
4185
62
077
385
<769
0 154
0385
<0038
<0002
<0385
<0054
<0038
< 0 154
2923
< 00385
< 00385
<0077
<0 154
0 10
000062
<0038
lb/1000 short tons
180.332 gal/1000 short tons
964'
8370
124
1 54
770
<1538
0308
0770
< 0076
< 0004
< 0770
< 0108
< 0076
< 0308
5846
< 00770
< 00770
< 0154
< 0308
020
0 00124
< 0076
MINE 2120-K
CONCENTRATION
(mg/5.1
27,524 Sm3/day
349*
4.590
4
<1.0
31
20
010
920
032
N/A
< 002
<007
1720
< 05
2.0000
100
033
024
< 05
1 35
00784
< 01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
15.173 m3/1000 metric tons
349'
69.630.3
607
<1517
713
3034
1 52
1.3956
485
N/A
< 303
< 1 06
2.6092
< 759
30.340
1.517
501
364
< 759
2048
1 19
< 152
lb/1000 short tons
3.635.997 gal/1000 short tons
3.49*
139.2606
1214
< 3034
1426
606.8
304
2.791 2
97
N/A
< 606
< 212
5.2184
< 1517
60.680
3.034
100?
728
< 15 17
4096
2.38
< 304
                                                                                                   O
                                                                                                   33
                                                                                                   >
                                                                                                   •n
•Value in pH units

-------
                                   TABLE V-11. RAW WASTE LOAD IN WATER PUMPED FROM
                                              SELECTED COPPER MINES (Sheet 2 of 4)
O
3)    <-
>    f
•Tl    CO
PARAMETER
Flow
PH
TOS
TSS
Oil and Grease
TOC
COO
B
Cu
Co
Se
Te
Ai
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
MINE 2120-B
CONCENTRATION

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

                                            SELECTED COPPER MINES (Sheet 3 of 4)
D
3D
     f
     to
PARAMETER
Flow
pH
TOS
TSS
Oil and Greaie
TOC
COO
B
Cu
Co
Se
Te
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Si
HE
Pb
MINE 2121
CONCENTRATION
Img/Zl
3.8153m3/day
737'
29.250
69
<10
<45
819
219
087
<004
<0077
060
<007
28
<05
<01
222
<002
<005
<05
119
< 0 0001
<01
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric Ions
17 28 m /1 000 metric tons
737-
5.053 9
11 9
<0173
<0778
141 5
0378
0150
<0007
<0013
0104
<0012
0484
<0086
<0017
0384
<0003
<0009
<0086
206
< 0 00002
<0017
Ib/ 1000 short tons
4.141 gal/1000 short tons
737-
10,107 8
238
<0346
< 1 556
283
0756
03
<0014
<0026
0208
<0024
0968
<0 172
<0034
0768
<0006
<0018
<0172
41.2
< 000004
<0034
MINE 2122
CONCENTRATION
(mg/2)
3.274m3/day
761'
2.288
2
3
21
389
Oil
1.90
190
<0003
0.2
<007
133
<02
95
083
<005
013
<02
0.83
< 00001
<05
RAW WASTE LOAD PER UNIT ORE MINED
kg/1000 metric tons
34 m3/1000 metric tons
7.6V
7869"
0069
0103
0722
134
0004
0065
0065
< 00001
0007
< 0002
0046
< 0007
0327
0029
< 0002
0004
< 0007
0.029
< 0 000003
< 0017
lb/1000 short tons
8.053 gal/1000 short tons
76V
15738
0 138
0206
1 444
268
0.008
0130
0130
< 00002
0014
<0004
0092
<0014
0654
0058
<0004
0.008
<0014
0.058
< 0000006
<0034
o
a
            •Value m pH unit*

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

                                              SELECTED COPPER MINES (Sheet 4 of 4)
1

PARAMETER
Flow
pH
TDS
TSS
Oil and Grease
TOC
COD
a
Cu
Co
Se
Te
fa
Zn
So
Fe
Mn
Cd
Mi
Mo
Sr
Hg
Pb
MINE 2123
CONCENTRATION

                                    • Value in pH units

-------
                            DRAFT
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 sur-
faces and to utilize the natural slope of ridges and valleys
for the recovery and collection of pregnant liquors.  In
some cases, dumps have been placed on specially prepared
surfaces.  The leach material is generally less than 0.61
meter (2 feet) in diameter, with many finer particles.
However, it may include large boulders.  Billions of tons
of material are placed in dumps that are shaped as truncated
cones.

The leach solution is recycled from the precipitation or other
recovery operation, along with makeup water and sulfuric
acid additions (to pH 1.5 to 3.0).  It is pumped to the top
of dumps and delivered by sprays, flooding, or vertical pipes.
Factors such as climate, surface area, dump height, mineralogy,
scale of operation, and size of leach material affect the
choice of delivery method.  Figure V-9 summarizes the reactions
by which copper minerals are dissolved in leaching.
                              V-36

                             DRAFT

-------
                                            DRAFT
    Figure V-8. FLOWSHEET OF HYDROMETALLURGICAL PROCESSES USED IN
                 ACID LEACHING AT MINE 2122
TO ATMOSPHERE
                                                        TO ATMOSPHERE
      43 m3/RMlnc ton
      110.282 oriMiort ton)
                    241 m3/nwtnc ton
                    (57.834 pl/tfiort ton)
EVAPORATION
                                                         EVAPORATION
                               27 m3/mitnc ton
                               (6.426 gri/lhort ton)
                                              -SEEPAGE —
                 296 m3/nutnc ton
               I70.B79 gal/dion ton)
 2.294 m3/nwtnc ion
(549.873 pl/ihort ton)
                           360 m3/nwtnc Ion
                           (86,303 gri/inort ton)


                               107 n>3/m«nc Ion
                              <25.704 gd/ihort ton)
                                                                   2.386 m3/Rwtnc Ion
                                                                   (571.914 gri/riiort ton)
                                16 m3/m«tnc ton
                               (3.727 ojlhhon ton)
                                                               OJ m3/metnc ton
                                                               (64 gil/ihort ton)
                                               V-37


                                            DRAFT

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

            Cu3(OH)2-(CO3)2 + 3H2SO4 ~^ 3CuSO4 + 2CO2


                               MALACHITE

               Cu2(OH)2-CO3 + 2H2SO4 ~—^ 2CuSO4 + C02 + 3H2O


                              CHRYSOCOLLA

                CuSiO3-2H20 + H2SO4 ~—^ CuSO4 + Si02 + 3H2O


                                 CUPRITE

                      Cu2O + H2SO4     ^ CuSO4 + Cu + H2O

            Cu2O + H2SO4 + Fe2(SO4)3     ^ 2CuSO4 + H2O + 2FeSO4


                              NATIVE COPPER

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


                                TENORITE

                       CuO + H2SO4      "* CuSO4 + H2O

             3CuO + Fe2(SO4)3 + 3H2O "^"^ 3 CuSO4 + 2Fe(OH)3

           4CuO + 4FeSO4 + 6H2O + O2 ^~~^ 4CuSO4 + 4Fe(OH)3


                               CHALCOCITE

                   Cu2S + Fe2(S04)3  ~   ** CuS + CuSO4 + 2FeSO4

                  Cu2S + 2Fe2(SO4»3  "   *" 2CuSO4 + 4FeSO4 + S


                               COVE L LITE

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


          Chalcopyrite will slowly dissolve in acid ferric sulfate solutions and also
        will oxidize according to:

                       CuFeS2 + 2O2  ^     CuS + FeSO4;

                          CuS + 2O2  ^     CuSO4.

        Pyrite oxidizes according to:

                 2FeS2 + 2H2O + 7O2  ~  ^ 2FeSO4


                                  V-38


                                DRAFT

-------
                            DRAFT
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 surface leaching operation is similar
to a heap or dump leach.

Recovery o_f_ Copper From Leach Solutions.   Copper dissolved
in leach solutions may be recovered by iron precipitation,
elettrowinning, or solvent extraction (liquid ion exchange).
Hydrogen reduction has been employed experimentally.

Copper is often recovered by iron precipitation as cement
copper.  Burned and shredded scrap cans are most often used
as the source of iron, although other iron scrap and sponge
iron may also be used.  In 1968, 12 percent of the domestic
mine copper production was in the form of cement copper re-
covered by iron precipitation.  Examples of iron launders
and cone precipitators are shown in Figures V-10 and V-ll.

The pregnant copper solution (0.5 to 2.2 g/1) is passed over
shredded or burned iron scrap and precipitates copper by
replacement according to the reaction:

                CuS04 + Fe«=^Cu + FeS04
                              V-39

                             DRAFT

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

, - C
' - \
' ' \
...
'\
. -j!
i i-
u.
f CELL
SOLUTION
j- CELL FLOW -
5*
i
4*- _
i
r-
J- > *
C
T 0
1- 3
r
j*
j
T
-1- *-
J

J
\
x
N
J
TOP VIEW
                                       CANS
                                FLOW
     DRYING
      PAD
                                                        PERFORATED
                                                        SCREEN
DECANT
BASIN

'•v.':;.r-n\W^v ^\-^^-w* vwl-
1 / -:-N W •*' ' •
r
                           END VIEW

                      SOURCE: REFERENCE 23


                              V-40
                            DRAFT

-------
                            DRAFT
       Figure V-11. CUTAWAY DIAGRAM OF CONE PRECIPITATOR
  BARREN
  SOLUTION
     COOPER
SETTLING AND
  COLLECTION
       ZONE
   COPPER DISCHARGE
                                                         SCRAP IRON
DYNAMIC
ACTION
ZONE
                                              COPPER-BEARING
                                              SOLUTION
                       SOURCE: REFERENCE 23
                               V-41


                             DRAFT

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

Gravity iron launders employ gravity to allow solutions to
trickle over and through iron scrap.  Spray water washes
remove copper frequently from the can surfaces.  Occasionally,
solution is introduced  from below and flows upward through
the iron to produce a coarser, but highly pure, cement copper.
(See Figure V-10.)

Cone precipitators may  be employed for copper recovery.  Solu-
tion is injected, through nozzles at the bottom of the cone,
into the shredded iron  scrap.  This injection, under pressure,
both precipitates copper rapidly and removes it from the iron
surface by the turbulent action.  (See Figure V-ll.)

Precipitated copper is  recovered by draining and scooping out
the solids.  Recovery from pregnant solution may be 60 per-
cent.  The resulting cement copper is 85 to 99 percent pure
and is sent to the smelter for further purification.

The barren solution from a precipitation plant is recycled from
a holding pond to the top of the ore body, after sulfuric
acid and makeup water are added, if necessary.

Leach solutions containing greater than 25 to 30 grams per
liter of copper are usually sent to electrowinning facili-
ties.  The cathode copper produced is highly pure and
does not require smelting.

Solvent extraction of copper from acid leach solutions by
organic reagents is rapidly becoming an important method of
recovery.  When pregnant liquors contain less than 30 grams
of copper per liter, the process is most applicable.  (See
Figure V-12.)

In solvent extraction,  a reagent with high affinity for
copper and iron in weak acid solutions, and with low affinity
for other ions, is carried in an organic medium.  It is
placed in intimate contact with copper leach solutions,
where H+ ions are exchanged for Cu(-H-) ions.  This regener-
ates the acid, which is recycled to the dump.  The organic
medium, together with copper, is sent to a stripping cell,
where acidic copper sulfate solutions exchange H+ ion for
Cu(-H-) .  This regenerates the organic/H+ media and passes
copper to the electrolytic cells, where impurity-free copper
                              V-42

                             DRAFT

-------
                              DRAFT
Figure V-12. DIAGRAM OF SOLVENT EXTRACTION PROCESS FOR RECOVERY OF
          COPPER BY LEACHING OF ORE AND WASTE
                                      MINE
                                      DUMP
                                        I
                                   WEAK CuSO4
                                 LEACH SOLUTION
                       RAFFINATE
                      'RECYCLED "
                                      i
  SOLVENT
EXTRACTION
   PLANT
                                     Cu"1"1" ON
                                ORGANIC CARRIER
              RECYCLED
              ORGANIC
            CARRIER (H+)
                                      i


STRIPPING
AREA


                 RECYCLED
             ACIDIC ELECTROLYTE
                  (H2S04)
     T
   CuS04
ELECTROLYTE
    i
                                  ELECTROLYTIC
                                   RECOVERY
                                     PLANT
                                    CATHODE
                                     COPPER
                                       I
                                       TO
                                    STOCKPILE
                               V-43


                              DRAFT

-------
                            DRAFT
 (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 recycl-
ing.

Table V-13 lists the chemical characteristics of barren leach
solutions at selected plants.  This solution is always recycled
and is almost always totally contained.

Other metals, such as iron, cadmium, nickel, manganese, zinc,
and cobalt, are often found in high concentrations in leach
solutions.  Total and dissolved solids often build up so that
a bleed is necessary.  A small amount of solution may be sent
to a holding or evaporation pond to accomplish the control
of dissolved solids.

Handling and Treatment of  Water.   No discharge of pollutants
usually occurs from leaching operations, except for a bleed,
which may be evaporated in a small, nearby lagoon.

Process Description - Mill Processing

Vat Leaching.   Vat leaching techniques require crushing and
grinding of high-grade oxide ore (greater than 0.4 percent Cu).
(See Figure V-13.)  The crushed ore, either dry or as a
slurry, is placed in lead-lined tanks, where it is leached
                             V-44

                            DRAFT

-------
                           DRAFT
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
                              V-45

                            DRAFT

-------
                            DRAFT
      TABLE V-13. CHEMICAL CHARACTERISTICS OF BARREN HEAP,
                DUMP, OR 1N-SITU ACID LEACH SOLUTIONS
                (RECYCLED: NO WASTE LOAD)
PARAMETER
pH
TS
TSS
COD
TOC
Oil and Qreasa
S
At
B
Cd
Cu
Fa
Pb
Mn
Hg
Ni
Tl
Se
Ag
Ta
Zn
Sb
Au
Co
Mo
Sn
Cyanide
CONCENTRATION (mg/£l IN LEACH SOLUTION FROM MINE
2120
3.66*
28.148
14
61 5 X
1.3
<1.0
<0.6
<0.07
0.11
7.74
38.0
2.880.0
0.1
260.0
0.0009
2.40
<1.0
< 0.003
<0.1
1.0
940.0
<0.6
<0.06
3.30
.
.
<0.01
2124
2.82 •
47.764
188
1.172
28.0
8.0
<0£
023
031
0.092
146.0
6.300.0
<0.1
94.0
O.O012
720
< 0.1
< 0.040
<0.1
1.0
2BS
<0£
<0.05
3.80
0.76
-
<0.01
2123
3.56*
44,368
162
80
27.6
2.0
<0.6
0.07
<0.01
6.66
97.0
660.0
0.1
123.6
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
249*
83.226
34
386.1
46.0
6.0
<0.6
<0.01
0.08
4.60
72.0
3.600.0
1.14
190.0
0.0003
31.1
<0.1
<0.003
0.038
2.6
74.6
<2JO
<0.06
72.0
0.36
2.40
<0.01
2126
4.24*
29/494
218
440.0
11.0
0.0
<0*
<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
<0.6
<0.06
13.70
0.6
•
<0.01
2104
3J9-
•
•
•
•
•
"
0.04 to 0.60

0.66
6226
•
0.68
•
0.0003
•
"
0.13
~
•
•
"
'
•
•
•

•Value In pH units
                              V-46
                             DRAFT

-------
                                            DRAFT
               Figure  V-13. VAT LEACH FLOW DIAGRAM  (MILL  2124)
                                        TO ATMOSPHERE
                                                34 m /metric ton
                                                <8.166gal/ihort ton)
                         1 1 m"/metnc ton
                         (264 gaJ/ihort ton)
                290 m'/ralric ton
              (69.506 oil/ihort ton)
                                                                                  TO ATMOSPHERE
                                                                                   EVAPORATION
 SLIMES
    166 m /metric ton
    (39.453 gil/ihart ton)
                      28 m'/metric tan
                      (5.620 pl/ihort ton)
                           10 m /metric ton
                         (2.411 gal/ihort ton)
TO MILL
                  TO MILL
                                     TO WASTE
      BARREN SOLUTION •
      176 m3/metnc ton
      (42.181 B»l/ihort ton)
     _ COPPER-RICH 	
       ELECTROLYTE
       207 m'/metric ton
     (49.578 gal/ihort ton)
189 mj/metnc ton
(45.176 gal/short ton)
                                                                                          31 m /metric ton
                                                                                          (7.397 gal/ihon ton)
                                                                                  ELECTROWINNING
                                                                                        TO
                                                                                     STOCKPILE
                                              TO LEACH DUMPS
                                             AND PRECIPITATION
                                                  PLANT
                                               V-47


                                            DRAFT

-------
                           DRAFT
with sulfuiic acid for approximately four days.   This method
is applicable to nonporous oxide ores and is employed for
better recovery of copper in shorter time periods.

The pregnant copper solution, as drawn off the tanks, contains
very high concentrations of copper, as well as some other
metals.  The copper may be recovered by iron precipitation
or by electrowinning.

Water is utilized in the crusher for dust control, as leach
solution, and as wash water.  The wash water is low in copper
content and must go to iron precipitation for copper recovery.
Table V-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 solu-
tions 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.
                             V-48

                            DRAFT

-------
                                 DRAFT
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
12.568 1
49,578
METHOD OF RECOVERY
Solvent Extraction/Iron
Precipitation*
Electrowinni ng* *
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
                                   V-49

                                  DRAFT

-------
                         DRAFT
TABLE V-15. CHEMICAL CHARACTERISTICS OF VAT-LEACH BARREN ACID
          SOLUTION (RECYCLED: 1MO WASTE LOAD)
PARAMETER
pH
IDS
TSS
COD
TOC
Oil and Grease
Al
Cd
Pb
Cr
Cu
Fe
Mn
Ni
V
Tl
Se
Ag
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/t )
1.1*
169,000
515
331
96
1.0
1,540.0
0.42
2.0
17.0
27,800
4,800.0
47.3
1.70
2.50
<0.03
< 0.003
0.17
11.5
51.0
2.0
< 0.01
          "Value in pH units
                            V-50

                          DRAFT

-------
                         DRAFT
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
H8
Ni
Se
Ag
Ti
Zn
Co
Mo
Cyanide
CONCENTRATION (mg/J.)
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
                           V-51

                         DRAFT

-------
                            DRAFT
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 floation at the concentrator.  The process includes
crushing, grinding, classification, flotation, thickening,
and filtration.   (See Figure V-14.)

Typically, coarse ore is delivered to the mill for two- or
three-stage reduction by truck, rail or conveyor and is
then fed to a vibrating grizzly feeder, which passes its over-
size material to a jaw crusher.  The ore then travels by con-
veyor 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        Example of          lb/short ton   kg/metric ton
 type           Reagent             mill feed       mill feed

pH control     lime                    10.0           5.0
collector      Xanthate                 0.01          0.005
collector      Minerec                  0.03          0.015
               compounds
frother        MIBC                     0.02          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.
                              V-52

                            DRAFT

-------
                                             DRAFT
      Figure V-14.  FLOW  DIAGRAM FOR FLOTATION OF COPPER  (MILL 2120)
                                     MINING
                                      ORE
                                    CRUSHERS
                          IBS m3/metrie ton
                        (46.736 g>l/diart ton)
                   RECYCLE
          REAGENTS
                          196 m3/matrw Ion
                         (46.736 ffl/ihort ton)
                                                 PROCESS
                                    BALL MILL
    CONCENTRATION


 CuS
FROTH
                              i
IATEH


196m3/
(46.735

WATER
SOFTENING
                                                                                   TO ATMOSPHERE
                                                                 TAILINGS-
                          THICKENERS
   FRESH WATER
                  27 m3/nwtrw ton
                 (6.491 gal/ihort ton)
                                0.06 m3/metne ton
                                (11.1 gdMrart ton)
RECYCLES (*)+ (1»V  118 m3/n»trh ton
         \*-S  V-X    128.189 mri/ihort «
                      (28.189 ga)/ihort ton)
                                                                RECYCLE-
                                                                            THICKENER
                                                THICKENED
                                                  TAILS

                                       108 m3/nwtric ton
                                     (25.964 QriMiort ton)

                               77 m3/mitiic ton         .
                               (18^46 orf/dnit ton) 	J
      IS m3/mtrle ton
      (3.709 gri/diort ton)
EVAPORATION
                                                                              TAILING
                                                                               POND
                                                                                 Il6m3/matne ton
                                                                                 (3.709 gal/ihort ton)
                                             OVERFLOW
                                              (IF ANY)
                                                                             DISCHARGE
                                                  V-53


                                               DRAFT

-------
                               DRAFT
Figure V-15. ADDITION OF FLOTATION AGENTS TO MODIFY MINERAL SURFACE
                      PULP FROM GRINDING CIRCUIT
                            (25-45% SOLIDS)

REAGENTS TO
ADJUST pH



'
CONDITIONER
i



WETTING AGENT

DISPERSANT
                             CONDITIONER
                                 I
COLLECTOR


CONDITIONER
                •FROTH-
                                i
           TO
       ADDITIONAL
       PROCESSING
FLOTATION
  CELLS
                      ACTIVATOR
                    (OR DEPRESSANT)
-TAILINGS-
                        TO
                       WASTE
                                 V-54

                               DRAFT

-------
                       DRAFT
TABLE V-17. EXAMPLES OF CHEMICAL AGENTS WHICH MAY BE
          EMPLOYED IN COPPER FLOTATION
MINERAL
Bornite
Chalcocite
Chalcopyrite
Native Copper
Azunte
Cuprite
Malachite
PRECIPITATION AGENT
—
—
—
—
Sodium momnulfide
Sodium monosulfidc
Sodium monosulfide
PH
REGULATION
Lime
Lime
Lime
Lime
Sodium urbonate
Sodium carbonate
Sodium carbonate
DISPERSANT
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
Sodium silicate
DEPRESSANT
Sodium cyanide
Sodium cyanide
Sodium cyanide
Sodium cyanide
Quebracho
Quebracho
Tannic acid
ACTIVATOR
—
—
—
—
Polysuliide
Polysulfidc
Polysulfide
COLLECTOR
Xanthdte
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats
Xanthate
Aerofloats.
Fatty acids
and salts
Fatty acids
and sdlts.
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
                        V-55

                      DRAFT

-------
                             DRAFT
 Rougher-cell concentrate is cleaned in cleaner flotation
 cells.   The overflow is thickened,  filtered,  and  sent  to
 the  smelter.  Tailings (sands)  from the cleaner cells  are
 returned to the mill for regrinding.   Tailings from the
 rougher cells are sent to the tailing pond for settling of
 solids.  Scavenger cells, in the last cells of the  rougher
 unit,  return their concentrate  (overflow)  to  one  of the
 first  rougher cells.

 In flotation, copper sulfide minerals are  recovered in the
 froth  overflow.  The underflow  retains the sands  and slimes
 (tailings).  The final,  thickened and filtered, concentrate
 contains 15 to 35 percent copper (typically,  25 to  30  percent)
 as copper sulfide.  Copper recoveries average 83  percent,
 so a significant portion of the copper is  discarded to tailing
 ponds.   Tailings contains 15 to 50  percent solids (typically,
 30 percent) and 0.05 to  0.3 percent copper.

 Selective or differential flotation is practiced  in copper
 concentrators,  which (for example)  are used for separation
 of molybdenum from copper concentrate,  for separation  of
 copper  sulfide from pyrite,  and for separation of copper
 sulfide from copper/lead/zinc ore.  Silver may be floated
 from copper flotation feed;  gold and  silver may be  leached
 by cyanide from the copper concentrate, with  precipitation
 by zinc dust.

 Water Usage iri 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  at  this time.
                              V-56

                            DRAFT

-------
                            DRAFT
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 aquatic biota, although their consti-
tuents and toxicity have not been fully determined.  Their
presence (if any), however, does not appear to hamper the
recycling of tailing decant water to the mill process.  Water
is characterized by 1 to 4 grams per liter of dissolved
solids and by the presence of alkalinity, sulfate, surfactant,
and fluoride.  Dissolved metals in decant water are usually
low, except for calcium (from lime employed in flotation
process), magnesium, potassium, selenium, sodium, and
strontium—which do not respond to precipitation with lime.
On occasion, cyanide, phenol, iron, lead, mercury, titanium,
and cobalt are detectable in the decant.  However, in these
cases, the water is either recycled fully or partially dis-
charged.

Handling or Treatment of Decanted Water From Mill Tailing
Ponds.   The majority of the industry recycles all mill pro-
cess water from the thickeners and the tailing pond due to
the need for water in the areas of major copper-ore production.
Of the balance of the industry, which includes approximately
six major copper producing facilities and an undetermined
number of operations producing copper as a byproduct, at
least half (50 percent) are currently working toward attaining
recycle of mill process water.  Also, of the six, three
have sophisticated lime and settling treatment, or are
installing it, to protect the quality of the discharge.

Three of the copper mills surveyed , aJl ->t wh^ch discharge
water from the railing pond, are compared  Ln Table V-20 as
to the quality of, and the amount of loading in, the discharged
                              V-57

                            DRAFT

-------
                           DRAFT
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
                             V-58

                            DRAFT

-------
                              DRAFT
      TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
                TAILING PONDS (Sheet 1 of 4)
PARAMETER
Flow
pH
SES'
TOS
TSS
Oil >nd Giuw
s.o2
Al
A>
Cd
Cu
ft
Pb
Mn
Hg
Mi
S.
Ag
Sf
Zn
Sb
Co
Au
Mo
PfMMpfWM
Cyan*.
OpiMtina
din/rMr
Annual
Production
oi Conctnttata
MILL 2119
CONCENTRATION
\rntU)
21 3,966 m3/d.¥
156.630.000 B»l/dav)
912-
3OX
1.422
4
<10
-
< 10
<007

105.980 m3/d.v
I28.OOO.OOO gal/d>*l
11 08-
35%
2.652
< 2
30
-
-
-
<002
077
520
<01
007
00008

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

PARAMETER

Flo*
pN
SES<
TDS
TSS
Oil and Gru»
s.o2
Al
Ai
Cd
Cu
Ft
Pb
Mn
Hg
Mi
S.
Ag
Si
Zn
Sb
Co
Au
Mo
Phosphate
Cvinid.
Operating
dar»/y«
Annual
Product ion
of Conc«ntrM«

CONCENTRAIION
l/ihort ion
928-
_
32.808
63S
<10S8
_
95248
< 741
317
4.8682
128.6907
4233
50799
00106
18203
<0317
< 106
635
89956
< 529
11641
<529
<529
_

-------
                               DRAFT
        TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
                   TAILING PONDS (Sheet 3 of 4)
PARAMETER
Flo-
PH
SESr
IDS
TSS
Oil ind GIMM
S*>?
Al
A.
Cd
Cu
Ft
Pb
Mn
Hg
Ni
$•
A«
S>
Zn
Sb
Co
A»
Mo
PhoiptMt.
Cyanida
OpH«.no
dayi/yur
Anniul
Product 100
ol Concnural*
MILL 2122
CONCENTRATION
(mg/ei
278 O84 n>3/d«y
<73.470.000oil/d«vl
854'
16*
4,778
24
3
122S
<1
<007
<005
008
<01
278
OM7
OOO02
<01
0022
<01
181
lua in pH unili
                                 V-61


                                DRAFT

-------
                       DRAFT
TABLE V-19. RAW MILL WASTE LOADS PRIOR TO SETTLING IN
           TAI LING PONDS (Sheet 4 of 4)
PARAMETER
FlOT.
pH
SES'
IDS
TSS
Oil and Graan
ao2
Al
A.
Cd
Cu
F.
Pb
Mn
HO
Ni
Sa
Al
Sr
Zn
Sb
Co
Au
Mo
Phosphata
Cvanida
Oparalmg
dayWyaar
Annual
Production
ol Concarwau
MILL 212*
CONCENTRATION
Img/EI
19.322 n.3/day
(5.1O4800galfdavl
1005*
50%
2.848
6
1
4675
<05
<007
005
912 6
1.982
035
31
00006
28
hort lam
24 ITOgil/irionton
10 OS1
-
573.989
1J2IO
2017
9.428 67
<1O08
< 14 12
1008
184.035 G
399 735 5
7059
6.263 2
0 1210
56471
< 06OS
< 20 17
24202
1 12942
<1008
33883
<1008
5.90729
4.195 0
< 202
362
69.362 metric lent {76.457 then iwul
       • Valuo in pH uniti
                          V-62


                        DRAFT

-------
                                 TABLE V-20. WASTEWATER CONSTITUENTS AND WASTE LOADS RESULTING

                                             FROM DISCHARGE OF MILL PROCESS WATERS
O
30
     o>
     10
PARAMETER
FLOW
pH
TDS
TSS
Oil end Graase
At
B
Cd
Cu
Fa
Pb
Mn
Hg
Ni
Se
Si
Zn
Co
Cyanide
CONTROL
TREATMENT
MILL 2120-
CONCENTRATION
15.140"
96'"
3,328
8
150
(007
<001
<0005
006
< 0 10
< 01
003
00011
< 005
0043
240
/10OO them loni
7.243.078' "
839'"
176.002
968
2418
<0604
968
<030
726
56.21
12088
362
<0006
<604
1.82
5924
<302
7.26
<0604
35* RECYCLE
NONE
O
a
             •Influenced by acid mma water and laach solution



             'influancad by mine water and smeller wastes


             "Inm3/day


             "in n>3/1000 metric tons


            •"In gal/1000 ihort tons



            111 Value in pH unit!

-------
                            DRAFT
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.

Variations in Mill Processes

Dual Process.   Ores which contain mixed sulfide and oxide
mineralization  in  equal ratios (greater than 0.4 percent copper
sulfides or oxides) may be treated with vat leaching, as well
as with froth flotation, in a dual process (Figure V-17).

Ore is crushed  and placed in vats for leaching with sulfuric
acid, as described under "vat leaching."  The leachate is
sent to iron precipitation or electrowinning plants for
recovery of copper.  The residue, or tails, remaining in the
vats contains nonleachable copper sulfides and is treated
by froth flotation to recover the copper, as described under
"Froth Flotation."

Water usage and tailing-water quality are similar to the
processes of vat leaching and froth flotation.  No discrete
discharge differences result from this variation compared to
vat leaching and froth flotation.
                              V-64

                            DRAFT

-------
                            DRAFT
Figure V-16. FLOWSHEET FOR MISCELLANEOUS  HANDLING OF FLOTATION
          TAILS (MILL 2124)
                     ORE
                   SULFIDE
               FLOTATION CELLS
                      I
                   TAILINGS
CONCENTRATE
                      TO
                      STOCKPILE
                     i
               HYDROSEPARATOR
  •SANDS-
                    SLIMES
                     i
                    ACID
                    LEACH
                 (THICKENER)
UNDERFLOW
  (SLIMES)
                  TO
                  TAILING
                  POND
                       PREGNANT
                       SOLUTION
               BARREN
              SOLUTION
                 PRECIPITATION
                    PLANT
                   CEMENT
                   COPPER
                      TO
                      STOCKPILE
                               V-65
                             DRAFT

-------
                     DRAFT
 Figure V-17. DUAL PROCESSING OF ORE (MILL 2124)
             MINING  •
             ORE
RECYCLED
 WATER      ORE
            LEACH
            TAILS
                          ACID
                        SOLUTION'
RECYCLED ACID-
                ELECTROWINNING
                               TAILING
                             THICKENERS
             RECYCLED
             DECANT
                                                CATHODE
                                                 COPPER
                                                  TO
                                               STOCKPILE
                       V-66


                    DRAFT

-------
                           DRAFT
Leach/Precipitation/Flotation (LPF) Process.    Mixed sulfide
and oxide mineralization may also be handled  by the leach/pre-
cipitation/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 + FeS04^
                      (excess)

Copper precipitates as a sponge, and the entire copper sponge,
together with pulp-sponge iron feed, is carried to flotation
cells.  Flotation recovers both sponge copper and copper
sulfide in the froth by means of the proper conditioning re-
agents, such as Minerec A as a collector and pine oil as a
frother.  Flotation is accomplished at a pH of 4.0 to 6.0
(+0.5).  The concentrate is thickened and filtered before
it is shipped to the smelter.  Copper recovery may be as high
as 91 percent.  An example of reagent consumption for this
process is:

      Reagent             kg/metric ton       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
                              V-67

                            DRAFT

-------
                       DRAFT
Figure V-18. LEACH/PRECIPITATION/FLOTATION PROCESS
             COPPER SULFIDE CONCENTRATE
                      AND
                  SPONGE COPPER
                   TO STOCKPILE
                         V-68


                       DRAFT

-------
                              DRAFT
 Lead  and  Zinc  Oret.

 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 (52.84 million gallons per day).

 The chemical characteristic of raw wastewater from the milling
 operation appear to be considerable less variable from
 facility  to facility than mine wastewater.  The volume of
 mill  discharge varies from as little as 1000 cubic meters
 per day (264,200 gallons per day) to as much as 16,000 cubic
 meters per day (4.23 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 seepage.

      (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 j.n situ before any metals
are solubilized.   Therefore, the extent of heavy metals in
 solution is minimal.   The principal contaminants of such mine
waters are:
                               V-69

                              DRAFT

-------
                               DRAFT
       (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
wastewater  flow for a mining operation exhibiting no solubil-
ization 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 seven 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 in this subcategory are  the same as those
for mines with  no  solubilization  potential.   The key differ-
ence in this  category is the local geologic  conditions that
prevail at  the  mine.   These conditions  lead  to  either gross
or localized  solubilization caused by  acid generation or
solubilization  of  oxidized minerals.   The resultant wastewater
pumped from the mine contains the  same  waste  parameters as
that from the preceding subcategory  but  also  contains sub-
stantial soluble metals.
                               V-70

                              DRAFT

-------
                    DRAFT
Figure V-19. WATER FLOW DIAGRAM FOR MINE 3105
SEEPAGE . . •*-
DRILL
WATER 270mJ/day
(72,000 gpd)

MINE
1



                   PUMPING
                       7.600 m3/day
                       (2.000,000 gpd}
               MILL FEED-WATER
                  RESERVOIR
                                     FUEL AND LUBRICANT
                                     SPILLAGE AND
                                     LEAKAGE
                                     EXPLOSIVE
                                     WASTE PRODUCTS
                      V-71


                    DRAFT

-------
                        DRAFT
TABLE V-21. RANGE OF CHEMICAL CHARACTERISTICS OF SAMPLED
          RAW MINE WATER FROM LEAD/ZINC MINES 3102, 3103,
          AND 3104
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/S, )
7.4 to 8.1 •
180 to 196
200 to 330
2 to 138
326 to 510
< 10 to 631
<1to4
3 to 29
0.03 to 0.15
< 0.05 to 1.0
< 0.0001 to 0.0001
< 0.2 to 4.9
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
                          V-72
                         DRAFT

-------
                            DRAFT
The following reactions are the basic chemical reactions that
describe an acid mine-drainage situation:

Reaction 1^ — Oxidation of Sulf ide 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 (water-
     soluble) ferrous sulfate:

          FeS£ + 302^   - >   FeSCM + 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:

          2FeS2^ + 2H20 + 70,2  - =»  2FeS04. + 2H2S04^

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:
                                 2Fe.2( 504)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
f errobacillus , Thiobacillus sulfooxidans) are involved in
the above reaction and, at least, are responsible for
accelerating the oxidation of ferrous iron to the ferric
state.
                             V-73

                            DRAFT

-------
                             DRAFT
 Reaction  3_ — Precipitation of Iron

 The  ferric iron associated with the sulfate  ion  commonly
 combines  with the hydroxyl ion of water to form  ferric hydrox-
 ide.   In  an acid environment,  ferric hydroxide is  largely
 insoluble and precipitates:

     Fe2_(S04).3_ + 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)_3 + FeS£ + H20  - >   3FeS04 + 2S

               S + 30 + H20  - >   H2S04_

The fact that very little  "free"  sulfuric acid is found in
mine waste drainage is probably due  to  the  reactions between
other soluble mineral species and  sulfuric  acid.

In some ore bodies,  such reactions — and subsequent solubili-
zation of metals — may occur  in local regions in which little
or no limestone  or dolomite  is available for neutralization
before the harmful solubilization  occurs.   Once a metal such
as copper, lead,  or zinc is  in solution, the subsequent
mixing and neutralization  of that  water may not precipitate
the appropriate  hydroxide  unless a rather high pH is obtained.
Even if some of  the metal  is precipitated,  the particles may
be less than 0.45  micrometer (0.000018  inch) in size and,
thus, appear as  soluble metals under current analytical
practice.

Conditions compatible with solubilization of certain metals —
particularly, zinc — are associated with heavily fissured ore
bodies.  Although  the minerals being  recovered are sulfides,
fissuring of the ore body  allows the  slight oxidation of the
ore to oxides, which are more soluble then  the parent minerals.
                               V-74


                             DRAFT

-------
                            DRAFT
When conditions exist which provide a potential for solubili-
zation, the mine water resulting is of a quality which requires
treatment beyond conventional sedimentation.  The best current
practice suggests that the treated mine water is likely to
be of a quality inferior to raw discharge from mines where
the potential for such solubilization does not exist.

A flow diagram illustrating flows encountered in a mine of
the type described in this subcategory is shown as Figure
V-20.  The characteristics of mine waters from this subcategory
are illustrated by Table V-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 wastewater from a lead/zinc flotation mill consists
principally of the water utilized in the flotation circuit
itself, along with any housecleaning water used.  The waste
streams consist of the tailing streams (usually, the under-
flow of the zinc rougher flotation cell), the overflow from
the concentrate thickeners, and the filtrate from concentrate
dewatering.  The water separated from the concentrates is
often recycled in the mill but may be pumped with the tails
to the tailing pond, where primary separation of solids
occurs.  Usually, surface drainage from the area of the
mill is also collected and sent to the tailing-pond system
for treatment.

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.
                               V-75

                             DRAFT

-------
                            DRAFT
        Figure V-20. WATER FLOW DIAGRAM FOR MINE 3104
SEEPAGE-
       MINE
(ALL WATER REQUIRED
 FOR DRILLING FROM
      SEEPAGE)
                         T
                       PUMPING
                           3.460 m3/day
                           (915,000 gpd)
                 XSEDIMENTATIONN
                 \^^  BASIN ^^/
                      DISCHARGE
                                              FUEL AND LUBRICANT
                                              SPILLS AND LEAKAGE
                                              EXPLOSIVE WASTE
                                              PRODUCTS
                               V-76


                             DRAFT

-------
                            DRAFT
TABLE V-22. RANGE OF CHEMICAL CHARACTERISTICS OF RAW MINE WATERS
          FROM FOUR OPERATIONS IN SOLUBILIZATION-POTENTIAL
          SUBCATEGORY
PARAMETER
PH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
NH3
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£ )
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
1to11
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
                            V-77

                           DRAFT

-------
                          DRAFT
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.
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.
                            V-78

                           DRAFT

-------
                                          DRAFT
                       Figure V-21. FLOW DIAGRAM FOR MILL 3103
                                                               WATER
                                                              FROM MILL
                                                            FEED RESERVOIR
                                                                  9 500 m3/d»y
                                                                  (2.500.000 gpd)
        Zn SCAVENGERS  Zn ROUCHERS
TO TAILING-
POND SYSTEM
                               TO STOCKPILES

                                    (b) MILL PROCESS
                                           V-79


                                         DRAFT

-------
                             DRAFT
 TABLE V-23. RANGES OF CONSTITUENTS OF WASTEWATERS AND RAW WASTE
           LOADS FOR MILLS 3102, 3103, 3104, 3105, AND 3106
PARAMETER
pH
Alkalinity
Hwdnn
TSS
TDS
COD
TOC
OilwdGf«n»
MBAS SurfKUnti
P
Ammonia
Hg
Pb
Zn
Cu
Cd
C>
Mn
ft
Cv«n«J.
Sulhu
Chloral*
Fluorri*
RANGE OF
CONCENTRATION
Img/CI
IN WASTEWATER
lam limit
79-
26
310
<2
670
714
11
a
018
OO42
r unit concMitraH produad
kg/1000 mi
tomf limit
_
1450
2.290
30
4300
30
30
30
205
054
032
< 000168
<0900
062
<018
<018
<018
< 045
0012
0091
1.260
210
203
UK lam
upon limit
-
10700
32500
2.000
50300
50000
580
130
607
254
IBS
0130
32.2
860
1.96
885
1 36
100
0198
0509
33.700
4070
545
Ib/IOOO dion tani
tomrhmn
_
2.900
4380
60
8.600
60
60
60
570
108
064
< 000336
< 18
124
< 036
< 036
<036
<090
<0024
0182
2.620
420
406
upper limit
-
20.400
66.OOO
4OOO
101JBOO
lOOjOOO
1 160
260
1214
508
370
0260
644
172
392
177
272
20
0396
1 18
67.4OO
8140
109
•VlIlM in pH umli
                               V-80


                              DRAFT

-------
                            DRAFT
 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 but 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 (dredging),  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.

 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  seepage into the  pit.  Ground-water
 seepage  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 parti-
cular 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.
                              V-81

                            DRAFT

-------
                              DRAFT
Figure V-22.   WATER FLOW IN  FOUR SELECTED  GOLD
                MINING AND MILLING OPERATIONS
                           (a) MINE/MILL 4101
             3.817 n,3/
-------
                              DRAFT
 Wastewater  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  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 constitutes
 the  final discharge.  The principal wastewater constituents
 from placer operations are  high suspended solids.

 Wastewater  emanating from mills consists  almost entirely
 of process  water.  High suspended-solid loadings are the most
 characteristic waste constituent of a mill waste stream.
 This is primarily  due to the necessity for fine grinding of
 the  ore to  make  it amenable to a particular beneficiation
 process.  In addition, the  increased surface area of the ground
 ore  enhances the possibility for solubilization of the ore
 minerals  and gangue.  Although the  total  dissolved-solid
 loading may not be extremely high,  the dissolved heavy-metal
 concentration may be relatively high as a result of the
 highly  mineralized ore being processed.   These heavy metals,
 the  suspended solids, and process reagents present are the
 principal waste constituents of a mill waste stream.
 Depending on the process conditions, the  waste
 stream  may  also have a high or low  pH.  The pH is of concern,
 not  only  because of its potential toxicity, but also because
 of the  resulting effect on the solubility of the waste
 constituents.

 Process Description - Mining

 Gold  is mined from two types of deposits:  placers and lode
 (vein)  deposits.  Placer mining consists  of excavating gold-
 bearing gravel and sands.   This is  currently done primarily
 by dredging  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,  S04—, Fe,  Mn,  and Zn) .
                              V-83

                              DRAFT

-------
                                 DRAFT
      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
Be
B
Cd
Ca
Cr
Cu
Total Fa
Pta
CONCENTRATION (mg/£)
MINE 4105
.
275
34*
2.40
1,190
1,176
14
733
35.01
12.0
1
0.095
<0.2
0.03
< 0.002

-------
                           DRAFT
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.

In general, the cyanidation process involves solubilization
of gold with cyanide solution, followed by precipitation of
gold from solution with zinc dust.  (See Figure III-9.)

The agitation-leach process employed by mill 4401 requires
water to slurry the ground ore.  Cyanide solution is added
to this pulp in tanks, and this mixture is agitated to main-
tain maximum contact of the cyanide with the ore.  Pregnant
solution is separated from the leached pulp in thickeners,
and gold is precipitated from this solution with zinc dust.
(See Figure 111-10.)

The vat leaching process is employed by mill 4105.  In this
process, vats are filled with ground ore slurry, and the
water is allowed to drain off.  Cyanide solution is then
sprayed into the vats, and gold is solubilized by cyanide
percolating through the sands.  Pregnant solution is collected
at the bottom of the vats, and gold is precipitated with zinc
dust.
                             V-85

                            DRAFT

-------
                            DRAFT
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.

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 4101 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
                              V-86

                            DRAFT

-------
                           DRAFT
upgraded by filtering and thickening.   Tailings from the
flotation process of mill 4101 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 pollu-
tant 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 con-
centrations of less than 100 mg/1 to over 1000 mg/1.

High levels of soluble metals usually result from the leaching
processes, and this is well-illustrated by the cyanide leach
process in the gold industry.  Table V-27 summarizes the
chemical composition and raw waste loads resulting from four
gold milling operations.  The processes represented include
amalgamation, cyanidation/agitation-leach, cyanidation/vat
leach, and the cyanidation/"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.
                             V-87

                            DRAFT

-------
                             DRAFT
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
ore milled
-
-
-
0.002
         TABLE V-26. MINERALS COMMONLY ASSOCIATED
                    WITH GOLD ORE
MINERAL
Arsenopynte
Pyrite
Chalcopynte
Galena
Sphalerite
Greenockite
Cinnabar
Pentlandite
Calverite
Sylvanite
Native Gold
Selenium
COMPOSITION
FeAsS
FeS
Cu FeS
PbS
ZnS
CdS
HgS
(Fe. Ni)g SB
Au Te2
(Au. Ag) Te2
Au
Se«
                   •Accompanies sulfur in sulfide minerals
                               V-88
                             DRAFT

-------
                               DRAFT
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
           MILLING OPERATIONS  (Sheet 1 of 2}
MINE/MILL
41O2
(Amaloinutnnl
4101
(Agnation Lose hi
4105
(Vat Luch)
4105 (Carbon
MI Pulp)
MINE/MILI
4102
(AmalgamBtioiO
41O1
(Agitation Luch)
4105
(W.I Laach)
4105 (Carbon
in Pulp)
MINE/MILL
4102
(Amalgamation)
4101
(Agitation Laach)
41OS
(Vit Loch)
4105 (Carbon
in Pulp)
MINEflMLL
4102
(Amalgamation)
4101
(Agnation LMdil
4105
(Vil Luch)
4105 (Carbon
m-Pulpl
TSS
CONCEN
TRATION
(mat I)
495.000
545 OOO
48!>6ob ~
WASTE LOAD
in ha/1000 matnc tons
llb/IOOO shorl tons)
ol ooneentuta produced
61 695 31 5 OOO
1123.3906301
11 541.485000
123 082 93O.OOO)
-
47, I0]|
94 x 1011
rn ki/1000 matnc toni
(Ib/tOOO short torn)
of on millad
2.871 000
(5 742 000)
436.0OO
1872.000)
_
4.I71.0OO
(8.342.000)
IOC
CONCEN
TRATION
fm»/ I t
343
50 O
~~
970

CONCEN
TRATION
Ima/fcl
003
017
~
20

CONCEN
TRATION
lrng/41
1 5
l
of ors millad
87
11741
-04
< 081
_
66O
(9.320)
TDS
CONCEN
TRATION
(mg/£)
462
4.536
-
886
WASTE LOAD
in kft/IOOO matnc ions
(Ib/IOOO short ions)
of concantrata producad
I9^42j000
(39884.000)
96 060 OOO
1192120000)
_
8b9,900 OOO
(1719800000)
in kg/1000 malnc tons
(lb/1000 short tons)
of ora millad
930
(1.860)
3600
(7.2001
-
7.600
(15,2001
COD
CONCEN
TRATION
Img/ei
1142
43
-
17894
WASTE LOAD
in ko/IOOO matric tons
llb/IOOO titan ions)
of concanlr ata producod
1423000
12,847 000)
9 11. OOO
(1822000)
-
173.700.000
1347400000)
in kg/ 1000 matric ions
llb/tOOO short tons)
of or« millod
66
(1321
34
1681
_
1 540
130801
Ai
CONCEN
TRATION
(mg/ei

-------
                              DRAFT
TABLE V-27. WASTE CHARACTERISTICS AND RAW WASTE LOADS AT FOUR GOLD
           MILLING OPERATIONS (Sheet 2 of 2)
MINE/MILL
4102
(Amalgamation)
4IO1
(Agitation Loch)
41OS
(V.I Laach)
4105 (Carbon-
in-Pulp)

MINE/MILL
41O2
(Amalgamation)
4101
(Agitation Loch)
4105
(V.t Laachl
41 OS (Carbon.
in Pulp)

MINE/MILL
4102
(Amalgamation)
41O1
(Agnation L«achl
4105
(V.I Laachl
4105 (Carbon-
m-Pulp>
Pb
CONCEN
TRATION
(ma/ HI
<01
/10OOdionlora)
ol concantr.ta produoad
< 2.500
K5.000I
2.100
(4^00)
< 4.300
K 8.6001
< 19.400
K 38.800)
in kg/ 1000 matrn tdni
(Ib/IOOO short tons)
ol ora millad
<0 1
CO 2)
008
(016)
<004
«0 08)
<017
K0.34)
CVANIOE
CONCEN
TRATION
(mg/ei
<001
506
-
006
WASTE LOAD
in kg/IOOO matric tons
(Ib/IOOO shon tons)
ol coneantrala producad
<1J50
K2300)
107000
(214.000)
__
58000
1116.000)
in kg/1000 matric tons
(Ib/IOOO short ion si
ol ora millad
<006
K012I
4
18)
-
052
(104)

                               V-90

                               DRAFT

-------
                            DRAFT
 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.

 Sources of Wastes

 There are two basic sources of effluents containing pollutants:
 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 precipita-
 tion,  runoff and ground-water seepage into the pit.  Ground-
 water seepage is the primary source of water in underground
mines.   However, in some cases,  sands removed  from mill
 tailings are vised to backfill stopes.  These sands may initially
                              V-91

                            DRAFT

-------
                        DRAFT
Figure V-23. WATER FLOW IN SILVER MINES AND MILLS
             S49m3/day
             (145.000 gpd)
         FLOTATION
                                                      M 264 m3/day
                                                      (nt 67.000 gpd)
./ 1.109 m3/d«¥
i
\
MILL
t

3.161m3/d»Y ^V rmm J ' V POI
(835^00 gpd) ^ 	 ^ ^^-__
1.635 m3 (432.000 gri)Atay

                       1.132 m3 (299.000 gall/dcy

                      (a) MINE/MILL4401
UNDERGROUND
MINE
DISCHARGE
2.933 m3/d«v
I (776,000 gpd)
V J 646 m3/
^ 	 ^ <144,0a

... ^ FLOTATION
^ ^ MILL
igpd» ' 	 1 '


RAIN
I 12m3/d>v
4(3,340 gpd)

1.600 m3/dav V *OH° / 7 7 n, J B m3/rtlv
(396.000 gpd) ^^. — __— " (715 to 914 gpd)
' 964m3/div
(252.000 gpd)
                     (b) MINE/MILL 4402
                          V-92


                        DRAFT

-------
                            DRAFT
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 is obtained for silver from placer
deposits as a byproduct of gold recovery.  Wastewater placer
operations utilize primarily the water that was used in the
gravity separation process.  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 dredge pond discharge, and this comprises the final
discharge.  The principal wastewater constituent from any
placer operations, whether silver, gold, or other materials, is
high suspended solids.

Wastewater emanating from silver mills consists almost entirely
of process water.  High suspended-solid loadings are the most
characteristic waste constituent of silver-mill waste streams.
This is caused by fine grinding of the ore, making it amenable
to a particular beneficiation process.  In addition, the in-
creased 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 pro-
cessed.  These heavy metals, the suspended solids, and process
reagents present are the principal waste constituents of a
mill waste stream.  In addition, depending on the process
conditions, the waste stream may also have a high or low pH.
The primary method of ore beneficiation in the silver-ore
milling industry is flotation.  As a result, mill waste
streams can be expected to contain process reagents.
                             V-93

                           DRAFT

-------
                              DRAFT
 Process Description - Mining

 As discussed previously, very little water use is encountered
 ins 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, total
 manganese, and antimony.

 Process Description - Milling

 Milling processes of silver ore which require water  and result
 in the waste loads present in mill water are:

       (1)   flotation,

       (2)   cyanidation,  and

       (3)   amalgamation.

 The selective froth flotation process can effectively  and
 efficiently beneficiate  almost any type and  grade  of sulfide
 ore.   This process is  employed by  mills 4401 and 4403  to
 concentrate the silver-containing  sulfide mineral  tetrahedrite
 and by mill 4402  to concentrate free  silver  and the silver
 sulfide mineral argentite.   In this  flotation  process, water
 is  added  in the ore grinding circuit  to produce a  slurry for
 transporting the  ore through the flotation circuit.  This
 slurry first flows through tanks (conditioners), where various
 reagents  are added to  essentially  cause the  desired mineral
 to  be  more amenable to flotation and  the undesired minerals
 and  gangue to be  less  amenable.  These  reagents are generally
 classified as collectors,  depressants,  and activators, according
 to  their  effect on the ore minerals and  gangue.  Also, pH
modifers are  added as  needed  to  control  the  conditions of the
reaction.   Following conditioning,  frothing  agents are added,
and  the slurry  is  transported  into  the  flotation cells,
where  it is  mixed  and  agitated  by  aerators at  the bottom
of  the  cells.   The collector and activating  agents cause the
                               V-94

                             DRAFT

-------
                              DRAFT
          TABLE V-28. RAW WASTE CHARACTERISTICS OF SILVER
                     MINING OPERATIONS
PARAMETER
PH
Acidity
Alkalinity
Color
Turbidity (JTU)
TOS
TDS
TSS
Hardness
COD
TOC
Oil and Grease
MBAS Surfactants
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
MINE 4401
8.0«
10.2
85.0
47t
2.0
504
504
<2
2403
11.9
17
4
0.085
<0.2
<0.07
< 0.002
<0.6
0.11
<0.02
46.0
C0.1
<0.02
0.33
<0.1
27.5
MINE 4403
.
4.2
76.2
<5*
23
622
622
<2
424.8
19.8
16
2
0.030
<0.2

-------
                             DRAFT
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
                               V-96

                              DRAFT

-------
                  DRAFT
TABLE V-29. MAJOR MINERALS FOUND ASSOCIATED
          WITH SILVER ORES
MINERAL
Tetrahedrhe
Tennantite
Galena
Sphalerite
Chalcopyrite
Pyrite
Naumannite
Greenockite/
Xanthochroite
Garnierite
Pentlandite
Native Bismuth
Argenite
Stephanite
Stibnite
COMPOSITION
(Cu, Fe. Ag)i2 AS4$13
(Cu, Fe. Ag)i2Sb4Si3
PbS
ZnS
CuFeS2
FeS
Ag2S
CdS
(Mg, Ni) 0- Si 02 • x H2O
(Fe, Ni)g S8
Bi
Ag2S
Ag5 Sb $4
Sb2S3
                  V-97
                 DRAFT

-------
                               DRAFT
 settled, the very fine particles of ground ore (slimes)  are
 normally suspended to some extent in the wastewater and  often
 present removal problems.  The quantity of suspended solids
 present in a particular waste stream is a function  of the  ore
 type and mill process because these factors determine how
 finely ground the ore is.

 Heavy metals present in the minerals listed in Table V-29
 may also be present in dissolved or dispersed colloidal  form
 in the mill waste stream.  Since the settlable solids, and
 most suspended solids, are collected and retained in tailing
 ponds, the dissolved and dispersed heavy metals present  in
 the final discharge are of concern.   Depending on the extent
 to which they occur in the ore body,  particular heavy metals
 may be present in a mill waste stream in the range  of from
 below detectable limits to 2 to 3 ppm.   Calcium, sodium,
 potassium,  and magnesium normally are found at concentrations
 of 10 to 250 ppm each.  In addition to  the suspended  solids
 and dissolved metals,  reagents used in  the mill beneficiation
 process also add to the pollutant loading of the waste stream.
 The particular reagents used are a function of the  process
 employed to concentrate the ore.   In the silver milling
 industry,  the various  flotation reagents (frothers,  collectors,
 pH modifiers, activating agents,  and depressants) are the most
 prominent  reagents  of  the flotation process.   Table V-30
 indicates  the quantity of these reagents consumed per ton of
 ore milled.   A portion of these reagents which are consumed
 in the process is present in the waste  stream.  Note  that a
 large number of compounds fall under the more  general categories
 of frothers,  collectors,  etc.   At any one mill, the particular
 combination  of reagents  used is normally chosen on the basis
 of research conducted  to  determine the  conditions under which
 recovery is  optimized.  While  flotation processes are
 generally similar,  they  differ specifically  with
 regard to the particular  reagent  combinations.  This is
 attributable,  in part,  to the  highly  variable  mineralization
 of  the ore  bodies exploited.   Waste  characterizations and raw
waste  loadings  for mill effluents employing  flotation and
 cyanidation  in  four mills are  presented  in Table V-31.  These
characterizations and  loadings are based  upon  analysis of raw
waste  samples  collected during site visits.
                               V-98

                              DRAFT

-------
                             DRAFT
TABLE V-30. FLOTATION REAGLNTS 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)
0-52
Z-200 (Isopropl ethylthiocarbamate)
Lime (Calcium oxide)

Sodium cyanide

PURPOSE

MILL 4401
Frother
Frother
Collector
pH Modifier
and Depressant
Depressant
CONSUMPTION
g/metric ton
ore milled

0.00498
0.00746
0.00187
0.109

0.00498
Ib/short ton
ore milled

0.00000995
0.0000149
0.00000373
0.000219

0.00000995
MILL 4402
Cresylic acid
Mineral oil
Dowfroth 250 (Polypropylene glycol
methyl ethers)
Aerofroth 71 (Mixture of 6/9-carbon
alcohols)
Aerofloat 242 (Essentially Aryl
dithiophosphoric acid)
Aero Promoter 404 (Mixture of
Sulfhydryl type compounds)
Z-6 (Potassium amyl xanthate)
Sulfuric acid
Soda ash (Sodium carbonate)
Caustic soda (Sodium hydroxide)
Hydrated lime (Calcium hydroxide)
Frother
Frother
Frother

Frother

Collector

Collector

Collector
pH Modifier
pH Modifier
pH Modifier
pH Modifier
2.83
6.9
0.545

10

90

1.82

70
250
1.260
3.03
320
0.00566
0.0138
0.00109

0.02

0.18

0.00363

0.13
0.49
2.51
0.00605
0.64
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

250
25
150
200
0.0025
0.0025
0.0075
0.015
0.01

0.005
0.05
0.3
0.4
                             V-99
                            DRAFT

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


4106
(Company Data
only)
MILL
4401
4403
4402
410S
(Company Data
only!

MILL
440.1

4402
it OS
(Company Gala
only!
MILL
44O1
4403
44O2
41OS
(Company OiU
only)

CONCEN
TRATION
Img/lt
650.000
7O3.000
9000O


CONCEN-
TRATION
Img/ei
220
240
290


CONCEN
TRATION
img/l)
026
OO3
022


CONCEN
TRATION

-------
                           DRAFT
TABLE V-31. WASTE CHARACTERISTICS AND RAW WASTE LOADS
           AT MILLS 4401, 4402, 4403, AND 4105 (Sheet 2 of 2)
MILL
44O1
4403
4402
4105
(Company Dill
only)
MILL
4401
4403
4402
4105
(Company Data
only)
MILL
44O1
4403
4402
4105
(Company Data
only)

MILL
4401
4403
4402
41O5
(Company Date
only)

CONCEN
TRATION
Img/tl
00024
00008
01490
OOO4

CONCEN
TRATION
Img/l)
<03
<03
•CO 3


CONCEN
TRATION
(mg/ei

(lb/1000 short tont)
of concantrata producad
<36
«72)
<30
K60)
58
(116)
< 32.400
K 64.800)
in kg/1000 matric toni
(lb/1000 diorl torn)
of ora millad
<09
«1 8)
<4
«8)
6
1121
<03
«06)
Cd
CONCEN
TRATION
Img/t)
<002
<002
<002
<001
WASTE LOAD
m kg/1000 matric tons
(lb/1000 short tons)
of concanf rat* producad
<36
K72I
<33
K66)
<22
K44)
< 6,500
K 13.000)
in kg/1000 matric tons
(lb/1000 short tons)
of ora millad
<009
K018)
<015
K030)
<02
K04I
<006
K012)
Sa
CONCEN-
TRATION
lmg/£)
0154
0144
-
—
WASTE LOAD
m kg/1000 matric tons
(lb/1000 short tons)
of concantrata producad
28
(56)
24
(48)

-
in kg/IOOO matric tons
(lb/1000 short tons)
of ora milled
07
(1 4| 	
1 1
12.21

-
Ni
CONCEN
TRATION
I me/ 11
0 14
005
010
010
WASTE LOAD
in kg/1000 matric tons
lib. 1000 short tons)
of concantrata producad
250
(5001
8
1161
11
(22)
64.700
1129400)
m kg/1000 malnc Ions
(lb/1000 short tons)
of ora milled
063
1126)
04
108)
1
(21
06
112)
So
CONCEN
TRATION
Img/ll
185
230
••02
-
WASTE LOAD
m kg/1000 matric Ions
lib/1000 short tons)
of concantrata producad
333
(6661
384
1768)
<22
K44)
-
in kg/IOOO metric tons
(lb/1000 short tons)
of ore milled
83
(166)
17
134)
<2
K4)
1
                          V-101

                           DRAFT

-------
                              DRAFT
Bauxite Ores

Water handling and quantity of wastewater flow within surface
bauxite mines are largely dependent upon precipitation patterns
and local topography.  Topographic conditions are often modi-
fied by precautionary measures, such as diversion ditching,
disposal of undesirable materials, regrading, and revegeta-
tion.  In contrast, underground mine seepage 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 accu-
mulates 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 ojE Wastewater

Open-Pit Mining.   The sequence of operations that occurs
in a typical open-pit mining operation is that the mine site
is cleared to 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 explo-
sive 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.
                              V-102


                               DRAFT

-------
                              DRAFT
Regardless of the method of mining, water use at the two
existing operations is generally limited to dust suppression,
and water removal is required because it results in 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 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 segre-
gation 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 waste-
water 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 waste-
water at times when pumping of mine water for discharge is
required.
                              V-103


                              DRAFT

-------
                                    DRAFT
Figure V-24. PROCESS AND WASTEWATER FLOW DIAGRAM FOR OPEN-PIT BAUXITE
           MINE 5101
   EXPLORATION AND ORE-BODY EVALUATION:
        GEOLOGICAL SURVEY
        TEST DRILLING
                     i
              SITE PREPARATION:
                  CLEARING
                  STRIPPING
    RUNOFF
     AND
GROUND WATER
                MINING:
                    BLASTING
                    LOADING
                    HAULING
    RUNOFF
     AND
GROUND WATER
         MILLING:
             CRUSHING AND GRINDING
             STORAGE
             BLENDING
                   2.76 m'/metric ton
                   (664 gal/short ton)
                   BAUXITE
          REFINING:
              COMBINATION PROCESS
                    2.76 ra3/metric ton
                 ' '  (664 gal/short ton) BAUXITE
                                             WATER TREATMENT
                                                   PLANT
                                                            o
                                                       2.76 m /metric ton
                                                     '  (664 gal/short ton) BAUXITE
  PRODUCTION = 2,594 metric tons (2,860 short tons) per day
  WATER TREATED DAILY = 7.165 m3 (1,900,000gal)
                   2.76 m3/metric ton
                   (664 gal/short ton) bAUXiTfc'
                                                DISCHARGE
                                     V-104

                                    DRAFT

-------
                            DRAFT
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 under-
ground—have minimized the problems and have resulted in
efficient ore recovery.

Initially, shafts are sunk to provide access to the bauxite
deposits, and drifts are driven into the sections to be mined.
A room-and-pillar technique is then used to support the mine
roof and prevent surface subsidence above the workings.
Configurations of rooms and pillars are designed to consider
roof conditions, equipment utilized, haulage gradients, and
other physical factors.

Ore is removed from the deposits by means of a "continuous
miner," a ripping-type machine which cuts bauxite directly
from the ore face and loads it into shuttle cars behind  the
machine.  Initial development of the room leaves much bauxite
in pillars, and it has been the practice to remove the pillars
and induce caving along a retreating caveline.  However,
resultant roof collapse and fracturing can greatly increase
overburden permeability, facilitating mine-water infiltration
and subsequently increasing mine drainage problems.  Recent
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 seepage 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  unconso-
lidated sands and clays, whereas open-pit drainage  is  exposed
                              V-105

                             DRAFT

-------
                            DRAFT
to sulfide-bearing minerals  in the soil.  As shown in this
section, open-pit and underground mine drainages differ
qualitatively and quantitatively; but, as a factor affecting
raw mine-drainage characteristics, mineralization does not
constitute a sufficient basis for subcategorization.

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 rag/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.

The class of raw mine drainage corresponds closely with
mining technique, and open-pit drainage is characteristically
acid.  Acid mine water is produced by oxidation of pyrite
contained in lignite present in the soil overburden of the
area.

Acid mine drainage with pH generally in the range of 2 to 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.
                            V-106

                           DRAFT

-------
                              DRAFT
TABLE V-32. CONCENTRATIONS OF SELECTED CONSTITUENTS IN ACID RAW
           MINE DRAINAGE FROM OPEN-PIT MINE 5101
PARAMETER
PH
Specific Conductance
Acidity
Alkalinity
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION (mg/£)
THIS STUDY
2.8*
1.000f
397
0
560
<2
7.2
3.5
23.8
0.82
0.3
500
0.29
INDUSTRY DATA
3.0*
250
0
617
2
21.8
3.23
18.6
1.19
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
  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*
1.580f
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*
2.000f
—
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
                              V-107
                              DRAFT

-------
                             DRAFT
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 img/i)
THIS STUDY
7.2»
1.260t
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*
150
550
300
5.0
5.0
2.0
1.6
0.01

50
2.5
    •Value in pH units

     Value in micromhos
                              V-108

                              DRAFT

-------
                            DRAFT
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.

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-like 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
                             V-109


                            DRAFT

-------
                                     DRAFT
     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
(rng/H)
IN WASTEWATER
560 to 1290
< 2 to 10
7.0 to 21 .8
3.23 to 4.2
18.6 to 38
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 0.028
0.02 to 0.06
0.01 to 0.01
0.05 to 0.10
0.002 to 0.003
0.0008 to 0.001
1.35 to 1.93
0.0001 to 0.004
Ib/short ton
3.10 to 7.12
< 0.012 to 0.056
0.04 to 0.1 2
0.02 to 0.02
0.10 to 0.20
0.004 to 0.006
0.001 6 to 0.002
2.70 to 3.86
0.0002 to 0.008
   Daily flow of wastewater = 7,165 m3 (1.900,000 gal)
   Daily mine production = 2,594 metric tons (2,860 short tons)
TABLE V-36. WASTEWATER AND RAW WASTE LOAD FOR UNDERGROUND MINE 5101
PARAMETER
TDS
TSS
Total Fe
Total Mn
Al
Zn
Ni
Sulfate
Fluoride
CONCENTRATION
(mg/&)
IN WASTEWATER
550 to 862
< 2 to 300
1.4 to 5.0
0.87 to 5.0
< 0.05 to 2.0
< 0.01 to 1.6
< 0.01 to 0.01
50 to 246
0.07 to 2.5
RAW WASTE LOAD
kg/metric ton
0.12 to 0.18
< 0.0004 to 0.06
0.0003 to 0.001
0.0002 to 0.001
< 0.00001 to 0.0004
< 0.000002 to 0.0003
< 0.000002 to 0.000002
0,01 to 0.05
0.00001 to 0.0005
Ib/short ton
0.24 to 0.36
<0.0008 to 0.12
0.0006 to 0.002
0.0004 to 0.002
<0.00002 to 0.0008
< 0.000004 to 0.0006
< 0.000004 to 0.000004
0.02 to 0.10
0.00002 to 0.0010
   Daily flow of wastewater = 83 m  (22,000 gal)
   Daily mine production = 390 metric tons (430 short tons)
                                      V-110
                                      DRAFT

-------
                           DRAFT
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  intro-
duction 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 to be similar (for example, gravity
processors of iron ore as indicators for gravity manganiferous-
ore operations).

Treatment of the individual process descriptions by ore
category, as adhered to previously in this report, is not
used here.  Instead, because of the wide diversity of ores
encountered, the general character of mine and mill effluents
is discussed, followed by process descriptions and raw waste
characteristics of several representative operations.

General Waste Characteristics

Ferroalloy mining and milling wastewater streams are generally
characterized by:

     (1)  High suspended-solid loads

     (2)  High volume

     (3)  Low concentrations of most dissolved pollutants.
                            V-lll

                            DRAFT

-------
                                     DRAFT
      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(D*
V<2)
V{1)
MILL
Category 1
(< 5.000 metric tons
[5.512 short tons] per year)






X

Category 2
(Physical
Concentration)
O

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
                                     V-112

                                     DRAFT

-------
                            DRAFT
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 dis-
charge.  In addition, the waste stream is generally contam-
inated 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-H-, Na+, K+, Mg++, C03—, 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 inter-
cepted 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 direct
precipitation on the pond, from runoff from surrounding areas,
or even from seepage and are only partially amenable to
elimination or control.

A number of operations or practices common to many
milling operations in this category involve the use of contact
process water and contribute to the waste-stream pollutant
load.  These include ore washing, grinding, cycloning and
classification, ore and tail transport as a slurry, and the
use of wet dust-control methods (such as scrubbers).  In terms
of pollutants contributed to the effluent stream, all of these
processes are essentially the same.  Contact of water with
finely divided ore, gangue, or concentrates results in the
suspension of solids in the waste stream, and in the solution
of some ore constituents in the water.  In general, total
levels of dissolved material resulting from these processes
are quite low, but specific substances (especially, some
heavy metals) may dissolve to a sufficient degree to require
treatment.  These processes may also result in the presence
                            V-113

                            DRAFT

-------
                              DRAFT
of oil and grease from machinery in the wastewater stream.
Good housekeeping and maintenance practice should prevent
this contribution from becoming significant.

Ore roasting may be practiced as a part of some processing
schemes to alter physical or chemical properties of the ore.
In current practice, it is used to change magnetic properties
in iron-ore processing in the U.S. and in the past was used
to alter magnetic/electrostatic behavior of columbium and
tantalum ores.  Roasting is also used in processing vanadium
ores to render vanadium values soluble.  Although a dry process,
roasting generally entails the use of scrubbers for air
pollution control.  Dissolved fumes and ore components rendered
soluble by roasting which are captured in the scrubber thus
become part of the waste stream.  This scrubber water may
constitute an appreciable fraction of the total plant effluent
and may contribute significantly to the total pollutant load.
One mill 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 wastewater gen-
erated at two current ferroalloy-ore processing plants.  Since
waste streams are not segregated, and the other processes
involve wastes of somewhat different character then those
normally associated with ore mining and beneficiation, such
operations may pose special problems in effluent limitation
development.

An additional component of the mill waste stream at some
sites which is not related to the milling process is sewage.
The use of the mill tailing basin as a treatment location for
domestic wastes can result in unusually high levels of a
number of pollutants in the effluent stream, including NH3_,
COD, BOD, and TOG.  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.
                              V-114

                              DRAFT

-------
                              DRAFT
       (2)  Oxidation of reduced (especially, sulfide) ores,
           producing acid and increased soluble material.

       (3)  Blasting decomposition products, resulting in NH3_,
                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

           Mn   5.5

In addition, oil and grease levels as high as 14 mg/1,  and
COD values up to 91 mg/1, were observed.  Since simple settling
treatment greatly reduces most of the above metal values,
it is concluded that most of metals present were contributed
in the form of suspended solids.  There is no apparent
correlation between waste content or flow volume and production
for mine effluents.

Sources £f Wastes  - Mill Effluents

Physical Processing Mill Effluents.  In general, mills practicing
purely physical ore beneficiation yield a minimal set of
                             V-115

                             DRAFT

-------
                         DRAFT
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 /mm (apm)l
2.66 (700)
6.43 (1.700)
34.06 (9.000)
11.3613.000)
pH
4.5
7.0
6.5
7.3
CONCENTRATION (mg/£)
Oil and
Grsau
14
1.0
2.0

Nitrate
-
0.15
0.12
-
Fluor Ida
443
4.5
0.52
-
As
<0.01
<0.01
<0.07
<0.07
Cd
007
<0.01
<001
<0005
Cu
3.8
0.06
<0.02
<0.02
Mn
5.3
5.5
0.21
63
Mo
05
<01
<0.1
<01
Pb
0.06
0.19
0.14
•
V
<0.5
<0.5
<0.5

Zn
70
047
0X15
0.09
                         V-116
                        DRAFT

-------
                              DRAFT
 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 opera-
 tions 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.

Flotation reagents pose particular problems in  effluent limit-
ation 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
                             V-117

                             DRAFT

-------
                             DRAFT
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-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
                              V-118


                             DRAFT

-------
                            DRAFT
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 compar-
ison.

Process Descriptions and Raw-Waste Characterizations For
Specific Mines and Mills Visited

Mine/Mill 6101

At mine/mill 6101, molybdenum ore of approximately 0.2 percent
grade is mined by open-pit methods and is concentrated by
flotation to yield a 90 percent molybdenite concentrate.
The mine and mill are located in mountainous terrain, along
a river gorge.  The mill is adjacent to and below the mine,
the elevation of which ranges from 2,550 meters  (8,400 ft)
to 3,000 meters (10,000 ft) above MSL (mean sea level).
The local climate is dry, with annual precipitation amounting
to 28 cm (11 in.) and annual evaporation of 107 cm (42 in.).

Approximately 22,000 cubic meters (6 million gallons) of
water per day are used in processing 14,500 metric tons
(16,000 short tons) of ore.  Reclamation of 10 percent of
the water at the mill site, evaporation, and retention in
tails reduce the daily discharge of water to 16,000 cubic
meters (4.3 million gallons).  Process water is drawn from
wells on the property and from the nearby river.  No mine
water is produced.
                             V-119

                             DRAFT

-------
                           DRAFT
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
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.
                            V-120

                            DRAFT

-------
                                           DRAFT
                          Figure V-25. MILL 6601  FLOWSHEET
                                                       CONCENTRATE-
                 •— MIDDLINGS •
   SCAVENGER
     FLOAT

  14 STAGES WITH
  REGRIND AND
INTERNAL RECYCLE)
                                                  -CONCENTRATE -

                                                    —MIDDLINGS-
    CLEANER
     FLOAT

  (8 STAGES WITH
  REGRIND AND
INTERNAL RECYCLE)
                                                                                     -TAILS-
                                      TAILS
                                       i
                                                                      CONCENTRATE
                          OVERFLOW—
                                                  -UNDERFLOW
    ^UNDERFLOW
                  OVERFLOW
                    i
    ^UNDERFLOW

    22.000 m3/
-------
                   DRAFT
TABLE V-39. REAGENT USE IN MOLYBDENUM MILL 6101
REAGENT
Lime
Vapor Oil
Pine Oil
Hypo (Sodium Thiosulfate)
(Na2 $203 • 5H20)
Phosphorus Pentasulfide (?2 SB)
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 Grease
COD
A*
Cd
Cu
Mn
Mo
Pb
Zn
fm
Total Cyanida
Fluonda
CONCENTRATION
(mat I )
IN WASTE WATER
500.000
2398
2.0
135
0.01
074
51
56.6
5.3
98
76.9
1.305
002
62
TOTAL WASTE
kg/day
14.000.000
42.000
32
2.200
016
12
820
900
85
160
1.200
21,000
032
99
Ib/day
32.000.000
92.000
70
4.800
0.35
26
14OO
2.000
190
350
2.600
46.000
070
220
RAW WASTE LOAD
par unit ore milled
kg/ metric ton
995
30
OO023
016
0000012
0.00086
0059
0064
0.0061
0011
0086
15
0000023
00071
Ib/short ton
1390
60
00046
032
0000023
00017
Oil
0.13
0012
0023
017
30
0000046
0014
par unit concentrate produced
kg/ metric ton
610.000
1330
1 4
96
0.0070
0.52
36
39
37
70
524
915
0014
43
Ib/short ton
1.200.000
3.670
28
190
0014
10
72
79
74
140
105
-1.830
0028
87
                    V-122
                   DRAFT

-------
                            DRAFT
Mining is both underground and open-pit, with underground
operations which began approximately 67 years ago,  and the
first open-pit production in 1973.  Recovery of molybdenite
is by flotation in five stages, yielding a final molybdenite
concentrate containing more than 93 percent MoS2^.  Tungsten
and tin concentrates are produced by gravity and magnetic
separation, with additional flotation steps used to remove
pyrite and monazite.  Recovered pyrite is sold as possible
(currently, about 20 percent of production), with the balance
delivered to tails.  The monazite float product also reports
to the tailing pond, since recovery of monazite is not
profitable for this operation at this time.

The mill operation is located on the continental divide at
over 3,353 meters (11,000 feet) above MSL.  The local terrain
is mountainous.  Climate and topography have a major impact
on water-management and tailing-disposal practices, with a
heavy snow-melt runoff and the presence of major drainages
above tailing-pond areas posing problems.

Mill Description.   Figure V-26 presents a greatly simplified
diagram of the flow of ore through the mill.  Following crush-
ing and grinding, roughing and scavenging flotation are used
to extract molybdenite from the ore.  Nearly 97 percent of
the incoming material—currently, about 39,000 metric tons
(43,000 short tons) per day—is thereby rejected and sent
directly to the byproduct recovery plant.  The flotation
concentrate, averaging about 10 percent MoS2^ is fed to four
stages of further flotation.  Reagents used in the primary
flotation step are summarized in Table V-41.  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 Moj32 content.
Reagent use in the cleaner grinding and flotation circuit is
summarized in Table V-A2.

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 temp-
erature raised to 70 degrees Celsius (158 degrees Fahrenheit),
and monazite is removed by flotation.  The tailings from this
                             V-1123

                             DRAFT

-------
                                 DRAFT
        Figure V-26. SIMPLIFIED MILL FLOW DIAGRAM FOR MILL 6102
    TO
TAILINGS
CRUSHING
(3 STAGES)
28% + 3 MESH
GRINDING IN
BALL MILLS
1
36% * 100 MESH
»|
1
FLOTATION



* I
CONCENTRATE |^ 	 1 FLOTATION
96% OF MILL FEED
CLEANER
FLOTATION 	 +- T° .
(4 STAGES) TA"
1

« LIGHTS GRAVITY SEPARATION DRYING
~~ "~ (HUMPHREY'S SPIRALS) 1
*
PYRITE
FLOTATION
1
TAILS
© |


1
^ MONAZITE «- 	 MONAZITE
~~ CONCENTRATE " FLOTATION
"*VV
MAGNETIC
SEPARATION
* *
f
CONCENTRATE
(93% + MoS2)

^_/Xx\ INDICATES I
V..X SAMPLING POINT \
„ 1

NONMAGNETIC MAGNETIC
TIN TUNGSTEN
CONCENTRATE CONCENTRATE
                                 V-12A


                                DRAFT

-------
                       DRAFT
   TABLE V-41. REAGENT USE FOR ROUGHER AND SCAVENGER
             FLOTATION AT IV ILL 6102
REAGENT
Pine oil
Vapor oil
Syntax
Lima (Calcium oxide)
Sodium silicate
Nokes reagent
PURPOSE
Frother
Collector
Surfactant and Frother
Adjustment of pH to 8.0
Slime Dispersant
Lead Depressant
CONSUMPTION
kg/metric ton
ore milled
0.18
0.34
0.017
0.15
0.25
0.015
Ib/short ton
ore milled
0.35
0.67
0.034
0.30
0.50
0.03
TABLE V-42. REAGENT USE FOR CLEANER FLOTATION AT MILL 6102
REAGENT
Vapor oil
Sodium cyanide
Nokes reagent
Dowf roth 250
Valco 1801
PURPOSE
Collector
Pyrite and Chalco-
pyrite Depressant
Lead Depressant
Frother
Flocculant
CONSUMPTION
kg/metric ton
ore milled
0.45
0.13
0.45
0.015
0.003
Ib/short ton
ore milled
0.90
0.25
0.90
0.03
0.006
                        V-125


                       DRAFT

-------
                            DRAFT
flotation step are dewatered, dried, and fed to magnetic
separators, which yield separate tin (cassiterite) and tungsten
(wolframite) concentrates.  Reagent use in the flotation of
pyrite and monazite is summarized in Table V-43.

Effluent samples were taken at three points in mill 6102
due to the complexity of the process.  A combined tailing
sample was taken representative of the total plant effluent,
and, in addition, effluents were sampled from two points
in the process (marked 19 and 20 on the flowsheet, Figure
V-26).  Although flows at these points are very small compared
to the total process flow, they were considered important
because of the acid conditions prevailing in monazite flota-
tion.  Concentrations and total loadings in the mill effluent,
and concentrations in the effluents from pyrite flotation and
monazite flotation, are presented in Tables V-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 signifi-
cantly 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.
                              V-126

                             DRAFT

-------
                         DRAFT
  TABLE V-43. REAGENT USE AT BYPRODUCT PLANT OF MILL 6102
            (Based on total byproduct plant feed)
REAGENT
I
PURPOSE
I
CONSUMPTION
kg/metric ton
ore milled
Ib/short ton
ore milled
PYRITE FLOTATION
Sulfuric acid
Z-3 Xanthate
Dowfroth 250

ARMAC C
Starch
Sulfuric acid
IpH Regulation
Collector
Frother
0.018
0.0005
0.0005
0.036
0.001
0.001
MONAZITE FLOTATION
Collector
WO2 Depressant
pH Regulation
0.0005
0.0005
0.0005
0.001
0.001
0.001
TABLE V-44. MILL 6102 EFFLUENT CHEMICAL CHARACTERISTICS
          (COMBINED-TAILINGS SAMPLE)
PARAMETER
TSS
TDS
Oil and Grease
COD
As
Cd
Cu
Mn
Mo
Pb
Zn
Fe
Fluoride
Total Cyanide
CONCENTRATION
img/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
2.400
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/shot ton
1996
9.3
0.016
0.098
O.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
                         V-L27


                        DRAFT

-------
                             DRAFT
   TABLE V-45. CHEMICAL CHARACTERISTICS OF ACID-FLOTATION STEP
PARAMETER
PH
Cd
Cu
Fe
Mn
Mo
Pb
CONCENTRATION (mg/Jl ) AT INDICATED POINTS OF FIGURE V-26
PYRITE FLOAT (19)
45*
0.01
02
4.2
4.0
3.0
0.3
MONAZITE FLOAT (20)
1.5
0.042
0.5
490
53.3
4.0
1.34
•Value in pH units
                             V-128




                            DRAFT

-------
                           DRAFT
Mine/Mill 6104

This complex operation combines mining,  beneficiation, and
chemical processing to produce a pure ammonium paratungstate
product as well as molybdenum and copper concentrates.  A
total of 10,000 cubic meters (2.9 million gallons) of water
are used each day in processing 2,200 metric tons (2,425
short tons) of ore.  The bulk of this water is derived from
the 47,000 cubic meters (13 million gallons) of water pumped
from the mine each day.

The mill process is illustrated in Figures V-27 and V-28,
which also show water flow rates.  After crushing and grinding,
sulfides of copper and molybdenum are floated from the ore,
employing xanthate collectors and soda ash for pH modifica-
tion.  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 concen-
tration, ammonium paratungstate is crystallized out of
solution and dried.

Effluent streams from parts of the operation specifically
concerned with beneficiation were sampled and analyzed*
along with the combined discharge to tails for the complete
mill.  Mine water was also sampled, and analyses have been
reported in Table V-38.  Data for a composite effluent from
beneficiation operations, several individual beneficiation
effluents, and the combined plant discharge are presented
in Tables V-46, V-47, V-48, V-49, and V-50.

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.
                            V-129

                           DRAFT

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

WET
ORE



WATER
FROM
CREEK





121
(32

m3Ml
0000
MTen
1
<" CRUt
II— *• Al
GRIN
C781000gpd)
1
" nan
HINQ O14JC
WHO
40Sn3Aky
(131MO gpd)
^ 1
180 0*1 BULK MILFIDS
"^ FLOTATION
4KB ml/d«y 1121.000 0d)

WATER
FROM
MINE





'

f
IOMETER
OO-FTI
THICKENER
i/wzni-/an
MILOOOgpd

(29X100 gpd)



COPPER/
MOLYBDENUM
SEPARATION




. 1*082 PRODUCT
1 OJOSB m*jtov
1 (lOgpdl
TO STOCKPILE
110 IB3J^taV
4 1
1071
^ OCMEELITE O7'°
^ THICKENER
STI
(127400 gpd)





BLOW
46«il
I121J
AM ATMO
ERI
K7WN
23m3Aky
(eunogpd
t
,
»3/*w
00 gpdl


TO ^
IPHERE
B4B m /dw
(144.000 gpd)
ITIOM
'

DRVINQ
ROASTINO .- _
*1
L

toO, PRODUCT
TO
STOCKPILE







~|Mn3M«
|(700gpd) .

«.
SCRUBBER



,

U3Sm3Ahv
(360000 vdl

FLOTATION
1
UNDEI

4.380 m*/*y
lt.1Bl.000 B>d)

IFLOW
r
402 «3Atay (130400 gpdl


CONCEI

PER
ITRATE
ENER
96n3AHv
OSJOOgpdl

Cu CONCENTRATE PRODUCT
TO STOCKPILE
114 m3Aby
(30.000 gpd)
FILTI
WAS!


299m3/
(79.000
RINO
ID
KINO
290m3/
179,000
MOLYBDENUM
r SEPARATION



m.tae^ioogpd)
TO
TAILING POND
chv
gpd)
1.106 m'/dly
(292400 gpd)
dw
gpd)
400M3Mw
dOBjOOOtpd)


TO
' S> SOLVENT
EXTRACTION
(FIGURE V-2BI
                        V-130


                       DRAFT

-------
                               DRAFT
         Figure V-28. INTERNAL WATER FLOW FOR MILL 6104
                    FOLLOWING MOLYBDENUM SEPARATION
                                                TO ATMOSPHERE
   FROM MOLYBDENUM
     SEPARATION
     (FIGURE V 27)
TO BOILERS
                                V-131



                                DRAFT

-------
                              DRAFT
    TABLE V-46. COMPOSITE WASTE CHARACTERISTICS FOR BENEFICIATION
              AT MILL 6104 (SAMPLES 6, 8, 9, AND 11)

PARAMETER

PH
COO
Oil and Create
At
Cd
Cu
Mn
Mo
Pb
Zn
Fluoride
Cyanide
CONCENTRATION
(mg/e) IN
WASTEWATER

10'
238
11 4
<0.07
004
49
22.5
190
0.22
6.3
4.8
02

TOTAL WASTE
kg/day

1.100
55
<034
0 19
24
110
91
1 1
30
23
096
It/day

2.400
120
<075
042
53
240
200
24
66
51
21
RAW WASTE LOAD
per unit ore processed
kg/metric ton

0.50
0.025
< 0.0002
0000086
0011
0.050
0041
000050
0014
0010
000044
Ib/short ton

10
0.050
<00003
000017
0022
010
0083
00010
0027
0021
000088
per unit total
concentrate produced
kg/metric ton

81
041
<0003
00014
0.18
081
067
00081
023
016
00072
Ib/short ton

16
0.81
<0007
0.0028
036
16
1 3
0016
046
0.32
0014
•Value in pH units
TABLE V-47. WASTE CHARACTERISTICS FROM COPPER-THICKENER OVERFLOW
           FOR MILL 6104 (SAMPLE 5)
PARAMETER
PH
CO
Cu
Mn
Mo
Pb
Fe
CONCENTRATION
(mg/JJ) IN
WASTEWATER
IT
0.26

-------
                            DRAFT
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 mil led
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/i) 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
                           V-133
                           DRAFT

-------
                     DRAFT
TABLE V-50. WASTE CHARACTERISTICS OF COMBINED-TAILING
          DISCHARGE FOR MILL 6104 (SAMPLES 15,16, AND 17)
PARAMETER
TDS
Oil and Grease
COD
As
Cd
Cr
Cu
Mn
Mo
Pb
V
Total Cyanide
CONCENTRATION
(mg/i) 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
tb/thort ton
340
2.2
26
<0.01
0.0046
0.0046
0.078
7.4
0.32
< 0.003
<007
< 0.002
                       V-134
                      DRAFT

-------
                              DRAFT
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 aquifer producing mine
water, which must be intermittently pumped out (for approxi-
mately % hour every 12 hours).  Total effluent volume is
less than 4 cubic meters (1,000 gallons) per day.  Samples
of this effluent were not obtained because of inactivity
during the site visit.  It is expected to be essentially the
same as the mill water-source well, which drains the same
aquifer and which was sampled.

Mine/Mill 6106

Ferronickel is produced at this site by direct smelting of
a silicate ore (garnierlte) 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.  Beneficia-
tion 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.
                              V-135

                              DRAFT

-------
                               DRAFT
  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/£) IN
WASTEWATER
1232
1
39.7
1.4
<0.07
<0.01
0.02
0.52
0.19
0.5
0.02
<0.5
<0.02
0.44
6.9
<0.01
TOTAL WASTE
kg/day
64
0.052
2.1
0.073
< 0.004
< 0.0005
0.0010
0.027
0.0099
0.026
0.0010
<0.03
< 0.001
0.023
0.36
< 0.0005
Ib/day
140
0.11
4.6
0.16
<0.01
< 0.001
0.0022
0.059
0.022
0.057
0.0022
<0.07
< 0.002
0.051
0.79
< 0.001
RAW WASTE LOAD
per unit ore processed
kg/ metric ton
1.4
0.0011
0.046
0.0015
C0.0001
<0.00001
0.000022
0.00058
0.00022
0.00057
0.000022
<0.0007
<0.00002
0.00050
0.0078
<0.00001
Ib/short ton
2.8
0.0022
0.092
0.0030
<0.0002
<0.00002
0.000045
0.0012
0.00043
0.0011
0.000045
<0.001
<0.00004
0.0010
0.016
< 0.00002
per unit total
concentrate produced
kg/metric ton
130
0.10
4.2
0.14
< 0.009
< 0.0009
0.002
0.053
0.020
0.052
0.0020
<0.06
< 0.002
0.045
0.71
< 0.0009
Ib/short ton
250
0.20
8.4
0.28
<0.02
< 0.002
0.010
0.11
0.040
0.10
0.010
<0.13
< 0.004
0.091
1.4
< 0.002
  TABLE V-52. CHEMICAL COMPOSITION OF WASTEWATER, TOTAL WASTE, AND
            RAW WASTE LOADING FROM MILLING AND SMELTER EFFLUENT
            FOR MILL 6106
PARAMETER
PH
TSS
TDS
Oil and grease
As
Cd
Cu
Ml
Mo
Pb
Zn
Fe
Nl
CONCENTRATION
(mall)
IN WASTEWATER
8.6*
226.9
212
3.4
< 0.07
< 0.005
< 0.03
0.53
0.5
<0.1
0.06
24
0.4
TOTAL WASTE
kg/day
3.600
3.300
54
<1
<0.08

-------
                             DRAFT
Mine/Mill 6107

At this operation, vanadium pentoxide, V205_, is produced from
an open-pit mine by a complex hydrometallurgical process
involving roasting, leaching, solvent extraction, and precipi-
tation.  The process is illustrated in Figure 111-21 and also
in Figure V-29 (which shows system water flows).  In the
mill, a total of 7,600 cubic meters (1.9 million gallons)
of water are used in processing 1,140 metric tons (1,250
short tons) of ore, including scrubber and cooling wastes and
domestic use.

Ore from the mine is ground, mixed with salt, and palletized.
Following roasting at 850 degrees Celsius (1562 degrees
Fahrenheit) to convert the vanadium values to soluble sodium
vanadate, the ore is leached and the solutions acidified to
a pH of 2.5 to 3.5.  The resulting sodium decavanadate
(Ha6yi002_8) is concentrated by solvent extraction, and ammonia
Is added to precipitate ammonium vanadate, which Is dried
and calcined to yield a V205_ product.

The most significant effluent streams are from leaching and
solvent extraction, from wet scrubbers on roasters, and from
ore dryers.  Together, these sources account for nearly 70
percent of the effluent stream, and essentially all of its
pollutant content.  Analyses for these waste streams are
summarized in Tables V-53, V-54, and V-55.  Effluents from
the solvent-extraction and leaching processes are currently
segregated from the roaster/scrubber effluent, although they
are both discharged at the same point, to avoid the genera-
tion of voluminous calcium sulfate precipitates from the
extremely high sulfate level in the SX stream and the high
calcium level in the scrubber bleed.  Both of these waste
streams exhibit extremely high dissolved-solid concentra-
tions (over 20,000 mg/1) and are diluted approximately 10:1
immediately prior to discharge.

Mercury Ores

Water flow and the sources, nature, and quantity of the wastes
dissolved in the water during the processes of mercury-ore
mining and beneficlation are described in this section.
                               V-137


                              DRAFT

-------
                                Figure V-29. WATER USE AND WASTE SOURCES FOR VANADIUM MILL 6107
O
3)
f
     oo


CI ""^>
.J*.
UjOOOgpT)
©1
t


4^20
MJO
?*)

i
' T
)«n)
f
*
SOLVENT
EXTRACTION
COOLING
1

U14£/mm 378tMn 114t/min 189 I/mm 871 C/mui IBSt/min
<400gpm) (IMgpml (30 gpm) (Mgpm) 1230 gpm) T(J (50 gpm)
1 i i I i AT"E |
WASHING. n_v.
WATER LEACH. AND ai™
SOLVENT EXTRACTION SCHUBI
SOLVENT , ,
ksKIM X
PONO^
1 RAIN 1 , (*)
114 I/mm AMMONIA
(30 gpm) TREATMENT
^U_lr —
^ — *->v
f 189^50.000-1(50.000.000^11) \
( WEST )
X, 	 EFFLUENT POND ^^/
I 	 ZJMUmm
1610 gpm)
, 3fi
NOTE "
RUNOFF FROM RAIN
IS NOT CONSIDERED
EXCEPT WHERE IT
ENTERS THE PROCESS
R SANITARV MISC ROASTER
IER USES USES OFF-GAS
SCRUBBER
*
SEWAGE
©TREATMENT
___^^
,__JL:
©
189
(SO
J MISC
j EVAPORATION UStS
t
®"
	 	 	
«P"I FROM S£~r
TAILING f "/.
DAM I **J
/"3.7aB.OOO-t(1.000.000«l)\ Xl1^66.000-t(3.000.000^ir\ 1 ^^~
l HOLDING J ( SCRUBBER ) 303 (/mm
X^^^ POND ^^r ^^^aLttD fOND^^S ISO gpm)
rt»rr — T "^ — i

j SOLIDS | j SOLIDS |
(M)
1
1

^~ 	
f DIVERSION
•^N
B30l/mn
(140gpcn)
2^70t/min T 5301/mm
(800 gpm) 1140 gpm)
1 RECYCLE
"SJ,1" WATER
MISC USES TOWER
I 1.136 (/n*,
1300 gpm)
1 	 — '
_^ 1.13B(/min
AN>\ '*" ;•""'
FER W-l
wJ
CD) VERSION
J.OND__^
;
jT jT (^^.ff^^v— a^-n^
RECYCLE WA^E X^J^UENT POND^X ' 	 '
^ ,- U49(/nitn
J* * (330 gpm)
40 gpm)
F®


O
3)
                  (»«) - SAMPLE NUMBER


                  SAMPLES (TOAND ftt^ARE
                                MINE WATER SAMPLES

-------
                               DRAFT
TABLE V-53. WASTE CHARACTERIZATION AND RAW WASTE LOAD FOR MILL 6107
          LEACH AND SOLVENT-EXTRACTION EFFLUENT (SAMPLE 80)
PARAMETER
pH
TOS
Oil ind green
COO
NH3
Ai
Cd
Cr
Cu
Mn
Mo
Pb
V
Zn
Fa
Ca
Chloride
Fluoride
Sulfate
CONCENTRATION
(mo/ III
INWASTEWATER
3.5-
39.350
94
475
0.16
035
0.037
1.15
016
54
< 0.1
< 0.05
31
052
0.26
206
7.900
4.6
26.000
TOTAL WASTE
kg/day
-
83.000
200
1.000
0.34
074
0.078
2.4
032
110
<0.2
< 0.1
65
1 1
055
430
17.000
9.7
56.000
Ib/day
-
180.000
440
2.200
0.75
1.6
0.17
5.3
07
240
< 0.4
< 0.2
140
24
1.2
950
37.000
21
120.000
RAW WASTE LOAD
par unit ore milled
ke> metric ton
-
73
0.18
0.88
0.0003
0.00065
0.000068
0.0021
000028
O.OS6
< 00002
< 0.0001
0057
0.00096
00006
0.38
15
0.0085
48
Ib/ihoft ton
-
146
0.35
176
00006
0.0013
000014
0.0042
000056
0.19
< 0.0004
< 00002
Oil
0.0019
0001
0.75
30
0017
96
per unit concentrate produced
ho/metric ton
-
6.570
16
79
0.027
0.059
0.0061
0.19
0.025
8.6
<002
<0.01
5.1
0.086
0045
34
1.350
077
4.320
Ib/thon ton
-
13.100
32
160
0054
0.12
0.012
038
005
17
<0.04
<0.02
10
0.17
009
68
2.700
1 5
8.640
  •Value in pH umti
                               V-139
                               DRAFT

-------
                               DRAFT
   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/U
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
C0.1
<0.05
33
0.37
31
130
4,800
1.5
290
RAW WASTE LOAD
per unit ore milled
kg/metric ton
_
_
3.5
0.007
0.027
0.0009
<0.00003
<0.000002
0.00011
0.00003
0.0018
<0.00004
<0.00002
0.013
0.00015
0.012
0.054
1.9
0.0006
0.12
Ib/short ton

^
7.0
0.014
0.054
0.0018
< 0.00006
< 0.000004
0.00023
0.00006
0.0037
< 0.00009
< 0.00004
0.026
0.00030
0.025
0.11
3.9
0.0012
0.23
•Value in pH units
                              V-140
                              DRAFT

-------
                              DRAFT
  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
Gd
Cr
Cu
Mn
Mo
Pb
V
Zn
Ca
Chloride
Fluoride
Sutfate
CONCENTRATION
(mg/£)
IN WASTEWATER
2.3*
2,000
80.768
5
1,844
0.04
0.08
< 0.006
0.9
<0.03
5.5
-
<0.05
-
< 0.003
78,400
59.500
7.5
780
TOTAL WASTE
kg/day
-
2,400
97,000
6.0
2,200
0.05
0.096
< 0.006
1.1
<0.04
6.6
—
<0.06
—
< 0.004
94,000
65,000
9.0
940
Ib/day
—
5,200
210,000
13
4300
0.11
0.21
<0.01
2.4
<0.09
15
-

-------
                            DRAFT
 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  classi-
 fication for sizing and feed into the concentration process.

 Water  is not used in mercury mining operations and is  dis-
 charged,  where it collects, only as an indirect  result of a
 mining operation.  This water normally results from ground-
 water  seepage 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 containing pollutants:
 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 seepage into the pit.
 Ground-water seepage 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.
                            V-142

                           DRAFT

-------
                                    DRAFT
      Figure V-30. WATER FLOW IN MERCURY MILLS 9101 AND 9102
        (NO DISCHARGE)

164m*Atav
(4.320 gpd)
GRAVITY-
MILL
t
^f TAILIMR
^V POMD
x^_
                             1.649 m3 diy (432.000 gpd)


                               (a) MINE/MILL 9201
       (NO DISCHARGE!
                                          3.8 m'/min 11.000 jpml

                            •DUE TO BEGIN OPERATION IN 1975


                               (b> MINE/MILL 9202
MILL
(NO DISCHARGE  WATER NOT USED
9ENEFICIATION LIMITED TO
CRUSHING AND/OR GRINDING TO
PROVIDE FURNACE FEED.)
                       (e) OTHER MERCURY OPERATIONS
                                    V-143


                                    DRAFT

-------
                             DRAFT
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 opera-
tion, due to the exposure of sulfide minerals and their
subsequent oxidization to sulfuric acid.  At acid pH, the
potential for solubilization of most heavy metals is greatly
increased.

Wastewater emanating  from mercury mills consists almost
entirely of process water.   High suspended-solid loadings
are the most characteristic  waste constituent of a mercury
mill waste stream.  This is  primarily due to the necessity
for fine grinding of  the ore to make it amenable to a parti-
cular beneflciation 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 pro-
cessed.  These heavy  metals,  the suspended solids, and process
reagents present are  the principal waste constituents of a
mill waste stream.  In' addition, depending on the process
conditions, the  waste stream may also have a high or low pH.
The pH is of concern, not only because of its potential
toxiclty, 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.
                              V-144

                             DRAFT

-------
                           DRAFT
 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

 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
                           V-145


                           DRAFT

-------
                            DRAFT
 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 mag-
 nesium 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.

 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.
                            V-146


                            DRAFT

-------
                        DRAFT
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
020
                         V-147


                         DRAFT

-------
                         DRAFT
TABLE V-57. WASTE CHARACTERISTICS AND RAW WASTE LOADINGS
          AT MILLS 9201 AND 9202
MILL
9201
9202
IPUot Operation!
MILL
9201
9202
IPUoiOpornion)
MILL
9201
9202
(Pilot OporMionl
MINE
9201
9202
(Pilot Operation)
jnHi
6.6
-
Hg
WASTE LOAD
CONCEN.
r"AJIOM In kg/1000 mnric ton
Img/ZI (rb/IOOOehortMral
of conamreie produad
-
00072 11
122)
ki kg/1000 rn.tr* Mm
lfe/1 000 draft Mm)
of oramUled
-
0084
10.188)
Sb
CONCEN-
TRATION
<06
0.03
WASTE LOAD
In kg/1000 metric Mm
db/IOOOriurtMra)
of ooncMifraw produced
< e.900
K13JOO)
BO
1100)
In kg/1000 metric 1om
(ft/1000 ehontoml
oforenHIMd
<006
K010I
04
(08)
Te
CONCEN
TRATION
<0.08

WASTE LOAD
in kg/1000 metric torn
llb/1 000 riiort torn)
of eonantrete produced
< 1.100
(C2.200)
-
m kg/1000 metric torn
IB/I 000 dran torn)
of ore milled
<0008
K001W
:
Zn
CONCEN-
TRATION
014

WASTE LOAD
ui kg/1000 metric Mm
Ub/IOMdioriMinl
ol eonamrele produad
1.930
13.8601
-
m kg/1000 metric ton
iB/IODOfhorttam)
at ore milled
0014
(00281
-
Fi
CONCEN-
TRATION
<06
OM
WASTE LOAD
m kg/1000 metric Mm
lB/1000ihorttom)
of oonoMitratB produced
103300)
80
1160)
HI kg/1000 metric torn
(Ib/IOOOctontom)
olonmUled

CONCEN-
TRATION
InVtl
002
0.3S
WASTE LOAD
hi kg/1000 metric Mm
ID/1000 dan lam)
of oonamrete produced
270
(540)
600
11.200)
in kg/ 1000 metric torn
(lb/1000 dwn torn)
of ore millid
0002
10004)
5
(10)
Mn
CONCEN-
TRATION
(mg/t)
500
0.06
WASTE LOAD
m kg/1000 metric Mm
(fc/1 000 dwn lorn)
of conantrpto produad
688.000
11.376,0001
79
lisa)
in kg/1000 metric torn
lib/1000 short tontl
ol ore mill*)
5
110)
065
(1301
SULFIOE
CONCEN
TRATION
Img/t)
,06
~
WASTE LOAD
» kg/1000 metric torn
IB/1000 diort lorn)
of concerrlrele prodiiad
< 6000
KI3400)
-
m kg/1000 metric torn
lib/1000 dwn tontl
ol ore rrultod
<005
K010I
-
                         V-148


                         DRAFT

-------
                            DRAFT
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) at concentrations exceeding 0.05 percent
as defined by the Nuclear Regulatory Commission, and which
are subject to NRG 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 water-
use 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-sltu 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 economi-
cal 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 seepage
from aquifers.  All known mills in this industry segment
use a hydrometallurgical process.

The quantity of water used in milling is approximately equal
in weight to that of the ore processed.  The quantity of
makeup water is equal to or less than this in the presence
of recycling.  From these considerations, it is apparent that
the entire uranium milling industry uses a fairly small quan-
tity of water, about 30,000 cubic meters (7.9 million gallons)
per day.
                             V-149

                            DRAFT

-------
                           DRAFT
          Figure V-31. TYPICAL WATER-USE PATTERNS
 F   IN-SITU LEACH  ~~\
I    ~~
MINE

H

TREATMENT



MILL
i
»f IMPOUNDME
^^^^^ ^—
i

                   (a) WET MINE/MILL COMPLEX
TREATMENT


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


TREATMENT
                DISCHARGE
                     (e) SEPARATED WET MINE
                            V-150

                           DRAFT

-------
                              DRAFT
Waste Constituents

Radioactive Waste Constituents.   Radium is the most potent
cumulative poison known, and its dissemination has been care-
fully supervised since the 1920's.  The chemistry of radium
is similar to that of calcium, barium, and strontium.  The
amount of radium that may be dissolved in water which can
be used for human consumption has been set by the International
Commission on Radiological Protection (ICRP) at three picocuries
(or picograms) per liter—i.e., about six orders of magnitude
below other typical pollutant concentrations.

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 could be established
between the rate of decay of uranium into radium and the
rate of decay of radium into its daughters.  Once this equili-
brium is established, the ratio of uranium to radium equals
the ratio of the half-lives—i.e., 2.7 million.  An equili-
brated ore with a typical grade of 0.22 percent uranium would
contain 0.82 microgram of radium per kilogram.  If the milling
process were to permit all of this radium to go into solution
in the water used to leach (or otherwise treat) the ore, and
if water were used in the typical ratio of one ton to each
ton of ore, then, without recycling, the concentration of
radium in tailings would also be 0.82 microgram per liter.
Although it is one of the least soluble substances, radium
sulfate is soluble to 20 micrograms per liter, and the above
concentration could be accommodated as a solute.  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 range from 0.000001 to 0.0021 microgram
per liter (1 picogram to 2,100 picograms per liter).  These
concentrations are often quoted in curies (Ci)—i.e., 1 to
2,100 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.)
                               V-151

                              DRAFT

-------
                              DRAFT
The radium content of tailing-pond solutions must, therefore,
be reduced by a factor of up to 600 to achieve ICRP drinking
water quality of 3 p'Cl/1.

Thorium.   There are other radioactive species that result
from the decay of uranium.  Only thorium 230, with a half-
life of 80,000 years, Is usually considered In addition to
radium 226.  It Is observed In concentrations of 1 to 100,000
pCi/1 In tailing-pond solutions.  A maximum concentration
for Th 230 of 2,000 pCl/1 Is recommended In an early radiological-
exposure standard (10CFR 20) that also sets a Ra 226 limit
of 30 pCi/1  (which is ten times the IRCP potable-water limit).

Chemical and Physical Waste Constituents.   Chemical contami-
nants of milling wastewaters derive from compounds Introduced
in milling operations or are dissolved from ore in leaching.
The common physical pollutants—primarily, suspended solids—
figure prominently in discharges from wet mines, and In the
management of deep-well disposal and recycle systems.
Seventy percent or more of radium 226 usually ends up in
solid tailings, including the suspended solids of effluents.

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 quanti-
ties of surfactants—presumably, by saponlfication (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
                               V-152


                              DRAFT

-------
                           DRAFT
  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)
  Alky I 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
                             V-153
                            DRAFT

-------
                              DRAFT
       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 n Ammonia, Phosphate.
Total Solids. Sulfata, Surfactants
Pb
Acidity. COD. Color. Dissolved Solids. Phosphate.
Total Solids
Ag. B. Ba. Hg. Zn
Color. Dissolved Solids. Fluoride. Sulfate. Total
Solids. Turbidity
Chloride. Color. Dissolved Solids. Total Solids.
Turbidity
Ag. Hg. K. Mg. Na
Color. Conductivity. Fecal Coliforrn. Hardness.
Phosphate. Suspended Solids. Total Solids.
Turbidity
Al. As. B. Ba. Be. Ca. Cd. Cr. Cu. Fe. Hg. Mg, Mo.
Ni. Pb, Sb. Se. Zn
Ammonia. Chloride. Hardness, Nitrete. 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. Sulfata, Suspended Solids
Al. As. Cr. Fe, Mn. Ni, Pb. Ti, V. Zn
Chloride, COD. Dissolved Solids. Kieldahl Nitrogen,
Nitrate. Volatile Solids
Co, K. Mn. Na
(None among the analyzed items)
•"Background" is defined in text.
                               V-154

                              DRAFT

-------
                            DRAFT
of more than ten.   Comparison of Tables V-58 and V-59
illustrates that more, rather than fewer, pollutants are
observed to be "added" by the operation than are predicted
from process chemistry and ore characteristics.  Observed
pollutant increases in conjunction with toxicant lists
were, therefore, used to select the parameters on which field
sampling programs were to concentrate.  (See also Section VI.)
Table V-59 also illustrates some specific differences among
the subcategories of SIC 1094 that are further explored in
the following discussion.

Constituents Introduced in Acid Leaching.  Acid leaching
(discussed in Section III) dissolves numerous ore constituents,
between one and three 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, molybdlc 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 opera-
tion 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 condi-
tion incoming ores by reaction with acid-consuming gangue.
Although this step aids in controlling pH of raw wastes, it
does not reduce the amount of sulfates therein.

Oxidants are added to the acid leach liquor following Initial
contact with ore and after reducing gases, such as hydrogen
and 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
(NaC103) and manganese dioxide (Mn02) 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.
                             V-155


                             DRAFT

-------
                             DRAFT
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 sulfldes, 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 wastewater of an alkaline leach  mill is largely derived
from two  secondary processes (Figure V-32):  tailing repulping,
and purification  (or sodium removal).   The leach itself is
recycled  via  the  recarbonatlon loop.   The wastes discarded
to tailings often contain organic compounds derived from
the ores.  Oxidizing agents are  used in leaching, but air
and oxygen gas under pressure have been found to serve as
well as more  expensive oxldants  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 sodium-
free 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,  a  needed fertilizer, 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 contri-
butions 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
                             V-156

                            DRAFT

-------
                DRAFT
Figure V-32. ALKALINE-LEACH WATER FLOW
         GROUND ORE
                                   TO ATMOSPHERE

FRESH WATER
OR
TAILING-SOLUTION
RECYCLE


ALKALINE
LEACH
t
FILTERING
i



uatu
STACK

, OMO
LEACH ,
RECYCLE 1
1 1
RECARBONATION
          PREGNANT
           LEACH
        PRECIPITATION
           CRUDE
          PRODUCT
                                                 STACK
                                                'GAS
t
                                  BARREN
                                   LEACH
FRESH
WATER


PURIFICATION
(SODIUM REMOVAL)
                            REPULPED
                            'TAILINGS
                             WASTE
                             WATER'
                                           . TO TAILING
                                           POND
            END
          PRODUCT
             I
             TO
          STOCKPILE
                V-157


                DRAFT

-------
                      DRAFT
Figure V-33. AMMONIUM CARBONATE LEACHING PROCESS
      MINING
                                TO ATMOSPHERE
        ORE
*
I
                                  WASTE GAS

AIR

WATER
—&•
—*•
GRINDING
1
PRESSURE LEACHING
1
COUNTERCURRENT
DECANTATION
^ LEACH
^SOLUTION"
PREGNANT
SOLUTION
AMMONIA AND
CARBON DIOXIDE
ABSORPTION TOWERS
t

' t
1
I
STEAM
1
1
1
GAS
1
1
STEAM S
PRECIP
i
1
TRIPPING
RANIUM
TATION

      SLURRY
                         •FILTRATE-
       TO TAILING
         POND
                                                SLURRY
                                              FILTRATION
                                               PRODUCT
                                                 TO
                                              STOCKPILE
                      V-158


                      DRAFT

-------
                           DRAFT
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)
of resin 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 contribution as total organic carbon (TOC)
due to other sources (for example, organic ore constituents).
Most of this contribution is in suspended solids; this
is illustrated by the fact that TOC is only about 6 mg/1
in the supernatant of the raw waste stream discussed
above.

Solvents are lost at a rate of up to 1/2000 of the water
usage in the SX circuit.  This ratio is set by the solubil-
ities of utilized solvents, which range from 5 to 25 mg/1,
and by the fact that inadequate slime separation can
lead to additional loss to tailing solids.  TOC of the
raw waste supernatant at mills using SX was found to
be 20 to 24 mg/1.  It is, again, impossible to determine
what part of this measurement should be ascribed to SX
solvents—particularly, in view of highly carbonaceous
ores.

The most objectionable constituents present in mill effluents
may be the very small amounts (usually, less than 6 ppm)
of the tertiary amines or alkyl phosphates employed in
solvent extraction.  In some cases, these compounds have
been found to be toxic to fish.  Dilutions of up to 1,500:
1 have been used (Reference 28) to lower the organic
concentrations of effluents of this type to acceptable values.
An analytic procedure for the entire class of these mate-
rials 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 waste-
water.
                            V-159

                           DRAFT

-------
                             DRAFT
Process Descriptions. Water Use, and Waste Characteristics for
Uranium. Radium, and Vanadium Ore Mining and Milling

Four mine/mill complexes in the licensed segment of the SIC
1094 category were visited to collect data on the utilization
of water and the characteristics of raw and treated wastes.
Water use in the mines and mills is listed in Table V-60, and
treatment systems employed are listed in Table V-61.

The consumption of water is seen to vary from 0.75 to 4.3
cubic meters per metric ton (180 to 1,030 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 illus-
trated 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 seepage and evaporation is probably not quite
equal to the raw waste input from the plant, and expansion of
the tailing-pond area may be necessary.  Similarly, it proved
difficult to account for the rain water entering the open-
pit mine of the operation in Figure V-38.  If and when it rains
into this mine, some water evaporates immediately from the
surface, while the rest runs into a central depression or
seeps into underground aquifiers.  The first and last effects
in combination are clearly dominant; less than ten percent of
the calculable water input is seen to evaporate from the
central depression (Figure V-38).
                             V-160

                             DRAFT

-------
                                     DRAFT
   Figure V-34. WATER  FLOW IN MILLS 9401, 9402, 9403, AND 9404
                                •t*<** «• n*ir
 MINE  I      .    »|  IONEXCHANOE  I I*ni100yfdl „ I DISCHARGE I
	,	1 8J3»m3Miy   I      ,   	1          H	I
       lao;ioo-di         \^7^
                         (1.320.000 gpd)
              1.636 oi3Miy
              (431^00 gpd)
                                   (a) MILL 9401
                                                       . 4.3J8 m3M«
                                                       ' (1.140.000 gpd)
        3.800 m:._.
        (061.000 gpd)
,3««1
'-'I	I
                                                      « 6.307 m3Miy
                                                      I < (1.400.000 gpd)
                                                                  -»4   LOSS
                                   (b) MILL 9402
                                                                           TO .
                                                                     ATMOSPHERE A
                                                                         EVAPORATION

r— ""^ M80
(i.80o.t
B.400 m3Miy
11.430.000 gpd)


.3«W
DO gpd)
DISCHARGE
MILL
f 620 m3Miy
1 (137.400 gpdl
I— — I RECARBONATION
I_

1
P
_ s^^.

(174.400 gpdl
1


3POND%-

ITATION
                                   (c) MILL 9403
               gpd)
  OPEMOT
    M1Mi  '  1J81mW
                                              TO ATMOSPHERE
                                                 TAILING
                                                  POMD   -X CAPACITY OF
                                                           1.636 m3Mn
                                                           1432/WOgpdl
                                  (d) MILL 9404
                                       V-161


                                      DRAFT

-------
                                 DRAFT
      TABLE V-60. WATER USE AND FLOWS AT MINE/MILLS 9401, 9402, 9403,
                 AND 9404

WATER CATEGORY
WATER USED
MINE/MILL 9401
m3/day
gpd
MINE/MILL 9402
m3/day
gpd
MINE/MILL 9403
m3/day
gpd
MINE/MILL 9404
m3/day
gpd
MINE PORTION
Water Supply
Discharge
Supplied to Mill
Recycled to Mill
Loss (Evaporation, etc.)
8,339
3,339
0
6.000
estO
2,203,000
882.100
0
1.321.000
estO
11.552
4,325
S.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
est 1.530
0
0
0
est 1,630
est404.200
0
0
0
est 404,200
MILL PORTION
Makeup Water
Water in Circuit
Discharge
Evaporation and
Seepage
2.700
3.200
0
2.700
713.300
845.400
0
713,300
6,307
8,900
0
6.307
1,402,000
2.351.000
0
1,402.000
6.060
6^80
5,400
660
1.601,000
1.738.000
1,427.000
174,400
5300
5,300
0
6,300
1.400.000
1.400,000
0
1,400.000
 N/A - Not available
       TABLE V-61. WATER TREATMENT INVOLVED IN U/Ra/V OPERATIONS
FEATURE

Smling Baun
Evaporating Pond
Ion-Exchange Plant

Tailing Pondlil
Ion-Exchange Plant
Recarbonitar
OacpWall
Utitiution of

Or« Handling
PARAMETER
S MINE/MILL
9401
9402
MINE PORTION
Area in hecterai lecrei )
Retention Time in houn
Area in hectares lacret)
U3 Og Concentration in mg/l
U3 OB Removal in *
OJ (0 74)
eft 20
N/A
25
96
O.J 11.7)
en 80
N/A
2 to 12
98
9403

N/A
N/A
N/A
N/A
N/A
MILL PORTION
Area in hectare* (acres)
Number term-connected
Daily Water UM in metric ton (inort tons)
Daily Water UM in metric torn (ihori lont)
Capacity in metric com Miort torn) water per day
Send/Slime Separator*
Decant Facilities
Filter.
Coprecipitation
21 (SI 8)
1
490 (540)
1.635(1 .8021
0
YM
V«
100(247)
6
N/A
N/A
0
Vei
TOTAL OPERATION
Capacity in metric ton hhon torn) per day || 3.200(3.527)
6.40017.055)
24 (59.3)
3
N/A
520(573)
0
Yea

1.400(1343)
9404

N/A
N/A
2 14.91
N/A
N/A

107 (264)
1
N/A
N/A
1.635(1.602)
Ves
Ves
Yes

2.70012.976)
N/A - Not available
                                V-162


                                DRAFT

-------
                             DRAFT
               Figure V-35. FLOWCHART OF MILL 9401
                                                         TO ATMOSPHERE
                                                             L
TO ATMOSPHERE
       TO STOCKPILE
                              V-163


                             DRAFT

-------
                             DRAFT
             Figure V-36. FLOW CHART FOR MILL 9402
                     MINING
    5.300 m3/day
   (1,400,000 gpd)
                      ORE
                     CRUSHING
                       AND
                     GRINDING
2S°4
Nad0
   3  \.
NaCI
NH,
                    LEACHING
                   THICKENERS
                    SOLVENT
                  EXTRACTION
                   PRECIPITATION
                  YELLOW CAKE


                  TO STOCKPILE
                                                      EVAPORATION
                                                      AND SEEPAGE
                                                           5,300 m3/day
                                                          (1.400.000 gpd)
                              V-164


                            DRAFT

-------
                           DRAFT
       Figure  V-37. FLOW CHART OF  MILL 9403
                     0 8 m'Atey
                      Ulgpml
                FLOAT TAILS
                ALTERNATIVE
                                                           WASH EQUIVALENT TO MOISTURE
                                                           RETAINED IN CAKE (ELIMINATED
                                                                   IN DRVER)
                     FROM ALKALINE
                       TAILING SUMP
                                                         6B4m3Miv
                                                         (120 opm)
  FROM ACID
TAILING SUMP
                                                 2jM3iMnc Ion/Ay
                                                 (3.134 ikon um/diy)
                                                     LIQUID
                                                               44-HECTARE 1108-ACRE)
                                                            TAILING POND WITH 23-HECTARE
                                                               (SB-ACRE) LIQUID POOL
TO STOCKPILE
                           V-165



                          DRAFT

-------
             DRAFT
Figure V-38. FLOW CHART OF MILL 9404
MILL WATER SUPPLY = 5,300
RAIN
1,530 m3/day
< i (404,200 gpd)
m3 (1.400.000 gal) per day
TO ATMOSPHERE
EVAPORATION
.137 m3/day
(36.200 gpd)
C320-hectare (790-acre) /2-hectare (4.9-acreKA
OPEN-PIT MINE I EVAPORATION ) )
Vy^POND^X^/


TOTAL LOSS OF
5,300 m3/day
(1,400,000 gpd) F
ii L


( TAILING ^
V POND A
\ FILTER ]
CAPACITY OF
1,635 m3/day
(431,900 gpd)
i i
( DEEP WELL


( WELL 2 ^

REPULP ] « SANDS
i i
)
BARREN
SOLUTION

V-166
DRAFT

MINING |
ORE

GRINDING •*— |
1
ACID LEACH
BIN
"1 ••
HYDROCYCLONES j
SLIMES
RESIN-IN-PULP
ION EXCHANGE ^_,
1
PREGNANT
SOLUTION
1ELLUE

SE
NT
(HIGH CD
PRECIPITATION
AND
FILTERING

~I}_ 	
— WASHING

-LOW c rJ


-------
                             DRAFT
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
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.
                             V-167

                            DRAFT

-------
                             DRAFT
Radioactive Nuclides.    Decay  products of uranium include
Isotopes of uranium,  thorium,  proactinium, radium, radon,
actinium, polonium,  bismuth, and  lead.   (Fission products,
generated by bombardment with  cosmic or manmade neutrons,
are undoubtedly also  present but  are too rare to deserve
further consideration.)  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.  A fair fraction
of the most-toxic isotope, radium 226, remains with solid
tailings and sediment  in mine-water settling basins.  Raw
waste, including these solids,  shows concentrations of
radium and uranium that  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 alarming, although 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 and seepage pond
illustrate 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, occasion-
ally, are distinguishable as oils or greases, surfactants, or
phenol.  The small amounts of  organlcs that are observed are
reviewed in Table V-63.

Inorganic Anions.   These may  be  distinguished into two classes:
(1) Sulfldes, cyanides,  and fluorides, for which technically
and economically feasible treatments (e.g., oxidation and lime
precipitation) are readily available; and (2) Chlorides, sul-
fates, nitrates, and phosphates,  which are present in fairly
large concentrations in mill wastes and cannot be removed
economically.  Distillation and reverse osmosis, while tech-
nically feasible, raise  the cost  of recovered water to the
level of $5 per cubic  meter ($20/1000 gal) and require a
large energy expenditure.  Impoundment, in effect, results
in distillation in regions like the southwestern states.
                              V-168


                             DRAFT

-------
                            DRAFT
    TABLE V-62. RADIONUCLIDES IN RAW WASTEWATERS FROM
                URANIUM/RADIUM/VANADIUM MINES AND MILLS
RADIONUCLIDE
and units of measurement
RADIUM 226
in picocuries/£
THORIUM
in mg/£
URANIUM
in mg/£
CONCENTRATION
MINES
200 to 3,200
<0.1
4 to 25
ACID MILLS
200 to 700
(4.100)*
(1.1)"
30 to 40
ALKALINE MILLS
100 to 19,000
N/A
4 to 45
•Parentheses denote values measured in wastewater concentrated by evaporation
N/A = Not available
 TABLE V-63. ORGANIC CONSTITUENTS IN U/Ra/V RAW WASTE WATER
PARAMETER
Total Organic Carbon (TOG)
Oil and Grease
MBAS Surfactants
Phenol
CONCENTRATION (mg/'l
MINES
16 to 45
3 to 4
0.001 to 7
<0.2
ACID MILLS
6 to 24
1
0.5
<0.2
ALKALINE MILLS
1 to 450
3
0.02
<0.2
                             V-169

                             DRAFT

-------
                             DRAFT
Other anlons are grouped  together in conjunction with the
light-metal cations as  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 wastewaters are  subject to inclusion in the cate-
gory of total dissolved solids.  Calcium, titanium, magnesium,
and aluminum respond to some treatments  (e.g., lime neutrali-
zation) and are shown separately.  Table V-65 shows concen-
trations of aluminum, beryllium, calcium, magnesium, and
titanium found in wastewater effluents of mines and mills
covered in this ore category.

Heavy Metals.  The leach  processes in the uranium/vanadium
industry involve highly oxidizing conditions that leave a
number of ore metals—specifically, arsenic, chromium, moly-
bdenum, uranium, and vanadium—in their most oxidized states,
often as arsenates, chromates, molybdates, uranates and vana-
dates.  These anionic species are, typically, much more soluble
than cations of these metals that precipitate as hydroxides
or sulfldes 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,
copreclpitate 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 catlonic heavy metals that had been expected to occur
from data on ores and processes include lead, manganese,
iron, and copper.  Field  sampling results added nickel, sil-
ver, strontium, and zinc  to this list.  The observed concen-
trations of these metals  are shown in Table V-67.  Cadmium
was found in a concentration above the lower detection limit
(20 micrograms per liter) at one alkaline mill discharge.

Other Pollutants.  Acid leach mills discharge a portion of
the acid leach; alkaline  leach mills discharge sodium car-
bonate; and mine water  is found to be well buffered with
measurable acidity and  alkalinity.  Chemical oxygen demand
is occasionally high, and raw wastes, reslurried only to
the extent needed for transport to tailings, carry a high
load of total solids.   These factors are reflected in the
data shown in Table V-68.   These measures indicate the need
                              V-170


                             DRAFT

-------
                        DRAFT
 TABLE V-64. INORGANIC ANIONS IN U/Ra/V RAW WASTEWATER
PARAMETER
Sulfide
Cyanide
Fluoride
Total Dissolved Solids (TDS)
CONCENTRATION (mg/l)
MINES
<0.5
<0.01
0.45
1.400 to 2,000
ACID MILLS
<0.5
<0.01
< 0.01
15.000 to 36.000
ALKALINE MILLS
< 0.5
< 0.01 to .04
1.4 to 2.1
5,000 to 13.000
TABLE V-65. LIGHT-METAL CONCENTRATIONS OBSERVED IN U/Ra/V
          RAW WASTEWATER

PARAMETER
Aluminum
Beryllium
Calcium
Magnesium
Titanium
CONCENTRATION (mg/l]
MINES
0.4 to 0.5
0.01
90 to 120
35 to 45
0.8 to 1.1
ACID MILLS
700 to 1.600
0.08
220
550
7
ALKALINE MILLS
0.2 to 20
0.006 to 0.3
5 to 3,200
10 to 200
2 to 15
  TABLE V-66. CONCENTRATIONS OF HEAVY METALS FORMING
            ANIONIC SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Arsenic
Chromium
Molybdenum
Uranium
Vanadium
CONCENTRATION (rnqfZ }
MINES
0.01 to 0.03
<0.02
0.5 to 1.2
2 to 25
as to 2.1
ACID MILLS
0.1 to 2.5
2 to 9
0.3 to 16
30 to 180
120
ALKALINE MILLS
0.3 to 1.5
<0.02
<0.3
4 to 50
0.5 to 17
                       V-171

                       D«AFT

-------
                        DRAFT
  TABLE V-67. CONCENTRATIONS OF HEAVY METALS FORMING
            CATION 1C SPECIES IN U/Ra/V RAW WASTEWATER
PARAMETER
Silver
Copper
Iron
Manganese
Nickel
Lead
Zinc
CONCENTRATION (mg/£ 1
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£to1
0.9 to 1.6
< 0.2 to 40
05
<0.5 to 0.7
OA
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
                         V-172


                        DRAFT

-------
                            DRAFT
 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.

 Waste Loads  in Terms of_ Production.  The loads of those
 pollutants  that indicated conditions warranting  treatment
 at the exemplary plants were related to ore production to
 yield relative waste loads.  The data for three  subcategories
 of the SIC  1094 segment are presented in Table V-69  (mines)
 and  Tables  V-70, V-71, V-72, and V-73 (mills).

 Occasional  large ratios between  the parameters observed at
 differing operations are believed to be due to ore quality.
 The  point is illustrated by TOC at mills 9401 and 9403:
 The  operators of mill 9401 had contracted to run an ore
 belonging to mine 9404 on a toll basis.  The ore carried a
 high carbonaceous material content that caused water at
 the  9401 mill to turn brown and may have adversely affected
 the  concentration process at mill 9404.  Mill 9403, in con-
 trast, was concentrating its own, much cleaner, ore.  The
 ratio of 200:1 in TOC is, therefore, expected.

 Radium, uranium, and thorium will be control parameters in
 this segment on the basis of the code of Federal Regulation
 10,  part 20  (10CFR20) and 10CFR25, regardless of their con-
 centrations in effluents.  This decision is dictated by the
 subcategorization and the assumption that concentrations
 of source material (uranium and thorium) in the ore exceed
 0.05 percent.

Metal Ores - Not Elsewhere Classified (SIC 1099)

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 character-
istics 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.
                            V-173

                           DRAFT

-------
                                   DRAFT
 TABLE V-69. CHEMICAL COMPOSITION OF WASTE WATER AND RAW WASTE LOAD
             FOR URANIUM MINES 9401 AND 9402
PARAMETER
TSS
COD
TOC
Alkalinity
Ca
MB
Fe
Mo
V
Ra
Th
U
MINE 9401
CONCENTRATION
(mO/ 8.)
IN WASTEWATER
_
242
16.8
224.4
93
46
0.47
0.6
1.0
3,190*
-
12.1
RAW WASTE LOAD
kg/day
—
2.300
150
2.100
860
420
4
6
9
29,700t
-
113
to/day
—
5.200
320
4.600
1.900
920
10
11
20
-
-
248
MINE 9402
CONCENTRATION
(mg/£)
IN WASTEWATER
299
600
25
-
117
36
0.23
0.53
<0£
2.710*
<0.1
11.6
RAW WASTE LOAD
kg/day
640
7 ,OOO
290
-
1.300
410
3
6
<6
31.100*
Owt
134
to/day
1,400
15.000
640
-
3.000
910
6
13
<13
—
<2.5
294
 •Valua In plcoeiirtes/£
  Value in picocurln/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
(rug/ >
IN WASTEWATER
294.000
65.6
460
12.200

-------
                                      DRAFT
     TABLE V-71. CHEMICAL COMPOSITION OF WASTEWATER AND RAW WASTE

                 LOAD FOR MILL 9402 (ACID- OR COMBINED ACID/ALKALINE-

                 MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Acidity
Al
Cu
Mn
Pb
At
Cr
Mo
V
Ra
U
COMCENTHATION
img/i )
IN WASTEWATER
525.000
63.5
24.0
35.000
1.594
2.7
105
2.1
23
9.0
16.0
125
234'
31.1
TOTAL WASTE
kg/day
4.100.000
337
127
185.700
8.460
14
557
11
12
48
85
663
1.240*
165
Ib/day
9.000.000
743
281
409.500
18.600
32
1.228
25
27
105
187
1.462
i ..
364
RAW WASTE LOAD
per unit ore milled
kg/metric ton
1.000
0.082
0.031
45
21
0003
0.14
0003
0.003
0.012
0.021
0.16
0.30"
0.040
Ib/short ton
2,000
0 16
0.062
91
4.1
0.007
0.27
0.005
0006
0023
0041
0.32
0.27"'
0.080
per unit concentrate produced*
kg/metric ton
450.000
37
14
20.400
930
1.6
61
1.2
1.3
5.2
9.3
73
136tr
18
Ib/short ton
900.000
74
28
40,800
1.860
31
122
24
27
10
187
146
124***
36
 •On the basis of 1973 production of 98.2% U.O- and 1 8% MO,
 A                          J O        3

  Value in picocunes/2


 "Value in microcuriei/day
 tt
  Value in microcuries/metnc Ion

•"Value in microcuries/shorl ton
                                       V-175
                                     DRAFT

-------
                                    DRAFT
  TABLE V-72. CHEMICAL COMPOSITION OF WASTE WATER AND RAW WASTE
               LOAD FOR MILL 9403 (ALKALINE-MILL SUBCATEGORY)
PARAMETER
TSS
COO
TOC
Alkalinity
Ca
MB
Ti
Al
Cu
ft
Mn
Ni
Pb
Zn
Al
Mo
V
Ra
Th
U
Fluoride
CONCENTRATION
(mg/ 1 1
IN WASTE WATER
111.000
278
<1
1.160.6
3.200
100
0.396
18
1 1
1 6
38
052
0.69
< 05
1 4
< 0.3
<05
111'
<0.1
39
1 4
TOTAL WASTE
kg/day
1.4OO.OOO
146
< 5.2
5.980
16.640
990
2.1
94
57
8.3
198
2.7
3.6
< 2.6
73
< 1 6
< 2.6
580"
•.05
20
7 3
Ib/day
3.100.000
319
< 11
13.190
36.680
2.180
45
206
13
18
436
6
79
< 57
16
< 34
< 57
-
<1
45
16
RAW WASTE LOAD
par unii on millad
kg/ metric ton
1.000
01
< 0.0037
43
12
071
00015
0.07
00041
0.0059
014
0.0019
00026
< 0.0019
00052
< 00011
< 0.0019
0.41"
< 00004
0032
00052
Ib/ihort ton
2.000
02
< 00074
85
24
14
00029
013
00081
0.012
028
0.0039
00051
< 00037
001
< 00022
< 0.0037
037"'
•.00008
0.064
001
par unit eoncantrata produced*
kg/matnc ton
1.050.000
109
39
4300
1.3
743
1.5
70
4.3
63
149
20
27
2
6.6
< 12
2
431 "
C0.4
34
55
Ib/ahort ton
2.100.000
217
78
9.000
25
1.486
31
141
86
13
297
41
5.4
4
11
< 23
4
392'"

-------
                                   DRAFT
    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
C*
M»
Ti
Al
Cu
Fe
Mn
Ni
R»
Zn
At
Cr
Mo
V
Ra
U
CONCENTRATION

-------
                              DRAFT
Water Uses.    The  primary  use  of water  in each of these
Industries  is  in the beneficiation  process, where it is
required  for the operating conditions of the process.  Water
is a primary material in the flotation  of antimony, titanium,
and rare-earth minerals; in the leaching of beryllium ore;
in the concentration of  titanium, zirconium, and rare-earth
minerals  (monazite)  from beach-sand deposits; and in
the extraction of  platinum metals from  placers by gravity
methods.  No primary tin ore deposits of any commercial
significance are currently being mined  in the U.S.  However,
a small amount of  tin is recovered  as a byproduct of a
molybdenum  operation through the use of flotation and magnetic
methods.

Water is  Introduced  into flotation  processes at the ore grinding
stage to  produce a slurry  which is  amenable to pumping,
sluicing, or classification for sizing  and feed into the
flotation circuit.   In leaching processes, water is the solvent
extraction medium.   Water  also serves as the medium for gravity
separation  of  heavy  minerals.

In underground mining of antimony ore and in open-pit mining
of titanium and  beryllium  ores, water is not used directly
but, rather, is  present  (if at all)  only as an Indirect con-
sequence  of these  mining operations.  The mining of sand placer
deposits  for titanium, zirconium, and rare-earth minerals is
done by dredging,  in which a pond is required for flotation
of the barge.  In  mining a placer for platinum-group minerals,
a barge may be floated either  in the stream or on an on-shore
pond, depending  on the location of  the  ore.

Water flows of the antimony, beryllium, platinum, rare-earth
titanium, and  zirconium  mineral operations visited are presented
in Figures V-39, V-AO, and V-41.

Sources of Wastes.    There are two  basic sources of effluents
containing pollutants:   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 seepage 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
                              V-178

                              DRAFT

-------
                                 DRAFT
Figure V-39. WATER FLOWS AND USAGE FOR MINE/MILLS 9901 (ANTIMONY) AND
          9902 (BERYLLIUM)
           INO DISCHARGE)

m TO ma ™s««
MjOOO TO 100.000 gpdl
FLOTATION
MILL

288 TO 343 m3/d«y
TAILING-
POND
IMPOUNDMENT


EVAPORATION
AND
SEEPAGE

                                                                  (NO
                                                                  DISCHARGE)
                        (a) ANTIMONY MINE/MILL 9901
                        (b) BERYLLIUM MINE/MILL 9902
                                  V-179


                                  DRAFT

-------
                                   DRAFT
Figure V-40. WATER FLOWS AND USAGE FOR MINE/MILLS 9903 (RARE EARTHS)
           AND 9904 (PLATINUM)
                                                         TO ATMOSPHERE
to08m3/nUnim A
(20gpm) I
IAROE)
0.08 m3/mlnuti
(Zlgprnl
0 M m3/mlniiU
(06 gpm)

LEACH CONCENTRATION

t


f
1 6 m'/minuti
(412 gpm)

EVAPORATION
-./EVAPORATION M
EVAPORATION
I
0.08m3/ml«rt."V TOND J
(20 gpm) ^- 	
^1 TAILING \
1.88 m'/minut.
(618 gpm)
RECLAIM
TANK

^V POND )

0.2 m3/ffllnuu
61 gpm)

               NOTE. FOR BYPRODUCT RECOVERY. SEE PART (b) OF FIGURE V-41 (MINE/MILL 89061
                         (a) RARE-EARTH MINE/MILL 9903


1

O.r •
2«.730 tn3/dty V^
— (6.480.000 gpd) ^^
24.730 m3/d«y
(8.480.000 gpd)
>REDQE
POND


J ^
^S *9-BO° "i3***
*"^ (12^80XXW gpd)
49^00 m3Miv
(12.960.000 gpd)
DREDG
BENEFI


EWITH
CIATION


                        (b)  PLATINUM MINE/MILL 9904
                                   V-180


                                  DRAFT

-------
                                     DRAFT
Figure V-41. WATER FLOWS AND USAGE FOR TITANIUM MINE/MILLS 9905 AND 9906
  OPEN-PIT
    MINE
   2.668 m3/day
   (699,000 gpd)
DISCHARGE
   TO
  RIVER
      FLOTATION AND
  MAGNETIC-SEPARATION
          MILL
             36,069 m3/day
             (9,460.000 gpd)
                                                   INTERMITTENT
                                                DISCHARGE (SEASONAL)
                              35.19 m3/day
                              (9.220,000 gpd)
                                     PUMP
                                     BASIN
                                            o
                                        878 m°/day
                                        (230,000 gpd)
                            (a) TITANIUM MINE/MILL 9905
                                                                   TO ATMOSPHERE
                        WET
                        MILL
                      (GRAVITY
                    SEPARATION)
  1—OVERFLOW •


12.099 m3/day
(3,170.000 gpd)
                            BULK
                       'CONCENTRATE"
        DRY
        MILL
   (ELECTROSTATIC
        AND
 MAGNETIC METHODS)
                          12,099 m3/day
                                                      7.633 m3/day
                                                     (2.000.000 gpd)
                                                                    EVAPORATION
fuar >
)gpd)
12.595
(3.300,0
^^ OTO
1
m3/day
00 gpd)

17,175 m3/rfay
(4,500.000 gpd) I
RAIN AND TO
RUNOFF STREAM
                   (b) TITANIUM/ZIRCONIUM/MONAZITE MINE/MILL 9906
                                       V-181


                                      DRAFT

-------
                              DRAFT
 at  this  time.   Effluents from beach-sand dredging  operations
 orglnate as precipitation,  runoff,  and groundwater seepage.
 In  addition, effluents result from the fresh water used  in
 wet mill gravity beneficiation of the sands  and, subsequently,
 are usually discharged into dredge ponds.

 The waste* constituents present in a mine or  mill discharge
 are functions  of the mineralogy of the ores  exploited  and of
 the milling or extraction processes and reagents employed.
 Acid conditions prevailing  at a mine site also  affect  the
 waste components by Influencing the solubility  of  many
 metallic components.

 Wastewater from a placer or sand mining operation  is primarily
 water that was used in a primary or secondary gravity  separation
 process.   Also, where a placer does not occur in a stream,
 water is often used to fill a pond on which  the barge  is
 floated.   The  process water is generally discharged into either
 this pond or an on-shore settling pond.   Effluents of  the
 settling pond  usually are combined with the  dredge-pond dis-
 charge,  and this comprises  the final discharge.  The principal
 wastewater constituents from these operations are  high suspended-
 solid loadings and coloring due to  high concentrations of
 humlc acids and tannic acid from the decay of organic  matter
 Incorporated into former beach sands and gravels being mined.

 Wastewater emanating from mills processing lode ores consists
 almost entirely of process  water.   High suspended-solid loadings
 are  the  most characteristic waste constituent of a mill waste
 stream.   This  is primarily  due to the necessity for fine
 grinding of the ore to make it amenable to a  particular benefi-
 ciation  process.   In addition,  the  increased  surface area of
 the  ground ore enhances the possibility for  solubillzation of
 the  ore  minerals and gangue.   Although the total dissolved-
 solld 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 toxlcity,  but also because of its
effect on  the  solubility of  the waste constituents.

Wastewater  emanating from a  beach-sand dredging pond consists
of water in excess of that needed to  maintain the pond at the
                               V-182


                              DRAFT

-------
                              DRAFT
 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 tannlc acids  which are Indigenous to  the  ore
 body and which result  in coloring of the water.

 Description o£ Character and Quantity of_ Wastes

 The quantity of wastes resulting from mining and milling
 activities is discussed below individually for each of  the SIC
 1099 metals.

 Antimony

 Process Description -  Antimony Mining.  Currently,  only one mine
 exists  which is operated solely  for the recovery of antimony
 ore (mine 9901).   This ore  is mined from an underground mine
 by drifting (following the  vein).

 As indicated in Figure V-39, no  discharge currently exists
 from the mine.

 Process Description -  Antimony Milling.   Only one  mill is
 operating for the  recovery  of antimony ore as the primary
 product.  This mill (9901)  employs the froth flotation
 process to concentrate the  antimony sulfide mineral, stibnite
 (Figure 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:  solubilizatlon 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 commer-
 cially valuable ore body of mine 9901 are:
                              V-183

                             DRAFT

-------
                         DRAFT
TABLE V-74. REAGENT USE AT ANTIMONY-ORE FLOTATION MILL 9901
REAGENT
Dowfroth 250 (Polypropylene glyool
methyl ethers)
Aerofloat 242 (Essentially Aryl
dithiophosphorie acids)
Ume (Calcium oxide)
Lead nitrate
PURPOSE
Frother
Collector
Depressant
Activating
Agent
CONSUMPTION
kg/metric ton
ore milled
0.4
0.1
2 to 3
05
Ib/short ton
ore milled
0.8
0.2
4 to 6
1.0
                         V-184





                         DRAFT

-------
                            DRAFT
           Stibnite            (Sb2S3)
           Pyrite              (FeS2)
           Arsenopyrite        (FeAsS)
           Sphalerite          (ZnS)
           Argentite           (Ag2S)
           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 North-
 west Basin.   Waste loadings for  important constituents of
 wastewaters from mill 9901 are listed in  Table V-76.

 Beryllium

 Process  Description - Beryllium  Mining.    Beryllium ore is
 mined on a large scale at  only one  domestic operation.   At
 mine 9902,  bertrandite (H2Be4Si209)  is  recovered by open-pit
 methods.   A small amount of beryl  is  also mined  in  the  U.S.
 by  crude open-cut and hand-picking  methods.  As  indicated
 in  Figure V-39,  no discharge  currently  exists  at mine  9902.

 Process  Description - Beryllium Milling.    Currently, only
 one  domestic  beryllium operation uses water  in a beneficia-
 tion process.  This  operation  is identified  as mill 9902 and
 employs  a proprietary  acid  leach process  to  concentrate
 beryllium oxide  from the ore.

 Quantities of Wastes.   As  indicated in Figure V-31, approxi-
mately 3,061 cubic meters  (802,000 gallons) per day of waste-
water are discharged from mill 9902.  Waste constituents
                            V-185

                           DRAFT

-------
                                     DRAFT
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
Be
B
Cd
Ca
Cr
Cu
Total Fa
Pb
Mg
Total Mn
CONCENTRATION (mg/£l
BJ*
8.6
11.0
113*
170
149
68
40
14
43
7JB
6.2
0.23
<0402

-------
                         DRAFT
TABLE V-76. MAJOR WASTE CONSTITUENTS AND RAW WASTE LOAD
           AT ANTIMONY MILL 9901
PARAMETER
PH
TSS
COD
TOC
Fa
Pb
Sb
Zn
Cu
Mn
Mo
CONCENTRATION
(mg/e) IN
WASTEWATER
8.3»
997
43
7.8
18.8
0.13
64.0
4.35
0.12
0.40
<0.2
RAW WASTE LOAD
per unit concentrate produced
kg/metric ton
—
74.78
3.22
0.585
1.41
0.0097
4.8
0.366
0.009
0.03
< 0.01 5
Ib/thort 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
                          V-187


                         DRAFT

-------
                           DRAFT
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          8102^
      Feldspar        Al silicates with Ca, K, and Na
      Fluorite        CaF2_
      Carbonates
      Iron Oxide Minerals
      Tourmaline      (XY3A16_(B03)3_(Si6018) (OH)4)
           where      X = Na, CaT 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 rela-
tively high concentration, as are aluminum (552 mg/1),
beryllium (36 mg/1),  and zinc (19 mg/1).  Raw waste loads
are not presented  because of the proprietary nature of the
process and production and because no effluent results at
this unique operation.

Rare Earths

Process Description - Rare-Earth Metals Mining.  The rare-
earth mineral monazite (Ce, La, Th, Y)P04_) is recovered
predominantly as a byproduct from sand placers mined by
dredging—primarily,  for their titanium mineral content.
(Refer to information on mill 9906, as described for titanium.)
The rare-earth mineral bastnaesite Is also currently recovered,
as the primary product, by an operation mining the ore from
an open-pit mine (mine 9903).

As indicated in Figure V-40, no discharge currently exists
at mine 9903.
                            V-188

                           DRAFT

-------
                          DRAFT
  TABLE V-77. CHEMICAL COMPOSITION OF RAW WASTEWATER FROM
            BERYLLIUM MILL 9902 (NO DISCHARGE FROM TREATMENT)
PARAMETER
Conductivity
Color
Turbidity MTU)
TDS
Acidity
Alkalinity
Hardnm
COO
TOC
OllandGrean
MBAS Surfactant*
Al
As
Be
Ba
B
Cd
Ca
Cr
Cu
Total Fe
Pb
Mg
CONCENTRATION (mg/£)
17.000*
88'
1.3
18,380
3J03B
0
4.000
22
55
<1
0.78
552
0.16
36.0
<6X>
OJK
0.047
43.0
0.20
QJ07

-------
                             DRAFT
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.   These rare-earth minerals,
bastnaesite  and some monazite,  are 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 con-
centrate that  averages 57 to 65 percent REO,  depending upon
the heads.   This concentrate is leached with  hydrochloric
acid to  remove calcium and strontium carbonates, increasing
the REO  content in the leached  concentrate by as much as 5
to 10 percent.   This  concentrate is  processed In a  solvent
extraction plant to produce high-purity europium and yttrium
oxides;  a cerium hydrate product;  a  concentrate of  lanthanum,
praesodymium and neodymium;  and a concentrate of samarium and
gadolinium (Figure V-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 decanta-
tlon circuit.   The primary thickener overflow containing the
chlorides is fed into the europium circuit, while the leached
solids from  the countercurrent  decantatlon 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 tempera-
ture of  60 degrees Celsius (140 degrees Fahrenheit) prior to
countercurrent  extraction of europium with  organic solvent
(90 percent  kerosene  and  10  percent  ethyl/hexyl  phosphoric
acid) .  The  raffinate from the  extraction  circuit makes up the
feed for  the lanthanum circuit, which is discussed later.
                              V-190


                             DRAFT

-------
                                    DRAFT
Figure V-42. BENEFICIATION OF BERTRANDITE, MINED FROM A LODE DEPOSIT,
            BY FLOTATION (MILL 9903)
         FRESH
         WATER
               OM «3/mfe  1M mj/mtmm
                         g1S»»il  f
        EQUIPMENT
                                CRUSHING
GRINDING
                                        CYCLONE
                                        UNDERFLOW




































t














_ I ,
1 'G
JHJOH4NTEKSITY







CLONE!



FIRST C
1 FLOTATI
T5
'
SECOND. THIRD
FLOTATIC
UNDEI
tCAVE
FLOTATIO
i
-FROTH— '
i








—FROTH 1 FIRS1



»— FROTH 	


:LEANER
ON CELLS
TH
'
. AND FOURTH «
HER — w
N CELL*
IFLOW
NOER
N CELLS




LASBIFICATION AND MOLYBDENUM

I
CONDITIONING P* 1 CONDITIONING l*~1 O
1 *C
L«»»FI

I '
AMD SECOND ROUOHEB .. ,
FLOTATION CELLS
UNDERFLOW
* ^~-
iF^TATlSNTErLh0"0"'10— K. ™



[HYDROCHLORIC 1
ACID |

|0™ 	 irL-.H 	
MCbNIHAlb I"™* AmTATQRS
1
ALTERNATIVE
1
I
f

1

i • f SODIUM 1
IORZAN | | CARBONATE |

t » t
»NO STAGE OF 1 ^ 1 FIRST STAGE OF 1
DNDITIONING f*~~| CONDITIONING |
1 0»C i
||1SO»FI J

'
i 1


	 v
ING >
NO J
—--^

1M n'/mlniti
(SlSgpml

1 fc LEACHED CONCENTRATE _».TO
THICKENER WAS
1 DRYER 1 1
LJL-J ,
ALTERNATIVE
1


                                     FINAL
                                    PRODUCT
                                                              SEPARATION OF RARE
                                                               EARTH METALS BY
                                                              SOLVENT EXTRACTION
                                                               (SEE FIGURE V43)
                                   TO STOCKPILE
                                     V-191


                                    DRAFT

-------
o
a
>
•n
VO
tVJ
                      Figure V-43. BENEFICIATION OF RARE-EARTH FLOTATION CONCENTRATE BY
                                 SOLVENT EXTRACTION (MILL 9903}
ILT
4
"t
!»« OVERFLOW
FILTER *~l THICKENER 1
1 1
i
, f
DRYER | ""'SiV'0" •*— | AMMONIA ]
JAR
Fl
NU
YDS

STW
	 . I 	 1
IE OR f
ANUM CARBON
«« F'IT" CUROP.UM
J PRODUCT


LE
| THICKENER «,« 	 1

1

SODIUM HVDROSULFIDE
AND AMMONIA
*
PRECIPITATION ^ 	 R«flNAT[
FILTRATE
DRUM
FILTER
t
I


GA

SODIUM CARBON
1
GADOIIMIUM/SAMA R IUM
PRECIPITATION TANK

EU
PURI
J
'
tOPIUM
•(CATION
J

-------
                            DRAFT
After loading the organic with europium, the europium is
stripped in the solvent extraction strip circuit with AN hydro-
chloric acid.  The pregnant strip solution contains iron,
which is removed in precipitation tanks by the addition of soda
ash to lower the pH to 3.0 to 3.5.  This causes ferric hydrox-
ide 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-40, raw wastes
are discharged at a rate of 1.96 cubic meters (518 gallons)
per minute from the flotation circuit and at a rate of 0.08
cubic meter (21 gallons) per minute from the leach/solvent
extraction plant.  These waste streams are not combined, and
both are characterized in Table V-78.  These data are based
upon the analysis of raw-waste samples collected during the
mill visit.  Table V-79 presents the results of chemical
analyses for the rare-earth metals.

Reagents used in the flotation, leach, and solvent extraction
processes of mill 9903 are identified below.
                            V-193

                            DRAFT

-------
                               DRAFT
        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
s.o2
Al
As
Be
B
Cd
Ca
Cr
Cu
Total Fe
CONCENTRATION (mg/JU
FLOTATION
9.02"
-
m
•
14/476
380.000
-
-
3,100
.
-
•
-
-
.
-
-
0.35
.
-
LEACH/
SOLVENT
EXTRACTION
8-23*
345
2.125
80t
52
76.162
786
7,220
>1.500
47
<1
212
126
<0.1
0.01
0.009
<0.01
< 0.005
2.910
0.04
<0.03
0.03
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Chloride
Fluoride
Sulfate
Nitrate
Phosphsto
Cyanide
Phenol
CONCENTRATION (mg/£)
FLOTATION
.
•
0.6
-
.
< 0.3
-
.
-
-
-
-
•
-
-
.
365
.
-
-
-
-
LEACH/
SOLVENT
EXTRACTION
<0.05
6.6
3jO
0.85
<0.1
<03
94
0.015
OJ09
650
4.6
3.36
7 JO

-------
                    DRAFT
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/ £ )
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
027
0.51
0.041
< 0.001
0.006
< 0.001
                    V-195

                    DRAFT

-------
                               DRAFT
Flotation Circuit

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

Leach Circuit

Leaching Agent       Hydrochloric Acid

Solvent-Extraction Circuit

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

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
(CeFCOS., with La, Nd, Pr, Sm, Gd, and Eu); barite (BaS04_);
calcite (CaC03^; and strontianite (SrC03).

The dissolved-solid content of the leach/solvent-extraction
waste stream are extremely high (76,162 mg/1) and are due
largely to chlorides (54,000  mg/1).  The metals present at
highest concentrations are those which would be expected on
the basis of known mineralization and use in the process.
These are strontium (4.5 mg/1) and barium (less than 10 mg/1).
The high concentration of tellurium (3.36 mg/1) is unexplained
on the basis of known mineralization, but mineralization is
assumed to be the source of this element.  Waste characteristics
and raw waste loading for the rare-earth flotation and concentrate
leaching/solvent extraction processes are given in Table V-80.

Platinum-Group Metals

Process Description - Platinum Mining.   Production of platinum-
group metals is largely as a  byproduct of gold and copper
refining, and primary ore mining is limited to a single dredging
operation (mine 9904) , which  is recovering platinum-metal
alloys and minerals from a placer deposit.
                               V-196

                               DRAFT

-------
                                    DRAFT
TABLE V-80. CHEMICAL COMPOSITION AND RAW WASTE LOAD FROM RARE-EARTH
            MILL 9903
PARAMETER
CONCENTRATION
{mg/£) IN
WASTEWATER
RAW WASTE LOAD f
per unit of concentrate
kg/metric ton
Ib/thort ton
per unit ore milled
kg/metric ton | Ib/short ton
(a) Flotation Mill
pH
TSS
TOC
Cr
Mn
V
Fluoride
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
< 0.001 6
1.89
(b) Leach/Solrant-Exchange Mill
pH
TSS
TOC
Si02
Cr
Mn
V
Te
Ni
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
C0.0003
0.003
0.001
_
1.67
0.094
0.00250
0.00008
0.006
< 0.0006
0.006
0.002
_
—
—
—
—
—
—
—
—
_
—
—
—
—
—
—
—
-
       Value in pH units

       Based upon maximum production achievable (part a) or estimated amount of flotation concentrate
       produced (part b)
                                    V-197


                                    DRAFT

-------
                            DRAFT
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 sep-
arated from the platinum-group metal alloys and minerals by
magnetic separation.  The final platinum-group metal concen-
trate is produced from  the magnetic-separation product by dry
screening and passing the resultant material through a blower
to remove the remaining lights.

Quantities of Wastes  .   Wastes resulting from the mining and
milling activities of this operation cannot be considered
separately, since the wet mill discharges to the dredge pond.
No reagents are required in the milling process, and, as a
result, the principal waste constituent from this operation
is suspended solids  (30 mg/1).  Table V-81 lists the chemical
composition of the wastewater and waste loads from mine/mill
9904.

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 (FeCr2_04)
      Ferroplatinum  (Fe, Ft, Ir, Os, Ru, Rh, Pd, Cu, Ni) alloy
      Iridium/ruthenium/osmium alloy
      Taurite (Ru, Ir,  Os)S2_
      Unnamed mineral (Ir, Rh, Pd)S
      Mertieite (Pt^(Sb, As)2_)
      Sperrylite (PtAs2_)
      Gold (Au)

Tin

Tin is recovered in the U.S. only as a byproduct of a molyb-
denum operation.  At this mine (6102), the ore is mined by
glory-hole methods, in  which the sides of an open hole are
                             V-198

                            DRAFT

-------
                             DRAFT
TABLE V-81. CHEMICAL COMPOSITION AND LOADING FOR PRINCIPAL WASTE
          CONSTITUENTS RESULTING FROM PLATINUM MINE/MILL 9904
          (INDUSTRY DATA)
PARAMETER
Alkalinity
Conductivity
Hardness
COD
BOD
TS
TDS
TSS
(N) NH3
Kjaldahl Nitrogen
Al
Cd
Cr
Cu
Total Fe
Pb
Zn
Chloride
Fluoride
Nitrate
Sulfate
Sulfide
CONCENTRATION
(mg/U
WASTEWATER
83
109*
35.6
7.6
35
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
                              V-199
                             DRAFT

-------
                              DRAFT
 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 to tin mining and
 milling.

 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  opera-
 tions  by  dredging methods.  In these operations,  a pond is
 constructed above the ore body,  and  a dredge is floated on
 the pond.   The dredges currently used normally are equipped
 with suction head cutters to mine the mineral sands.  Wastes
 from dredge ponds and wet mills  are  combined;  therefore,
 these  operations are discussed under one heading:   Dredging
 Operations.

 Quantities  of_ Wastes:  Mine 9905.    This is the only existing
 mine from which titanium lode  ore is mined.   Water is discharged
 from this open pit at a rate of  2,668 cubic meters (699,000
 gallons)  per day.   The chemical  composition of this waste is
 presented in Table V-82.   As these data  show,  oils and  grease
 (3.0 mg/1),  fluorides (3.20 mg/1), total KJeldahl  nitrogen
 (2.24 mg/1),  and nitrates (15.52 mg/1) are present at relatively
 high concentrations.   The oils and greases undoubtably  result
 from the heavy equipment  used  in the mining operations, and
 the  fluorides  are  probably indigenous to the ore body.  How-
 ever,  the reason for the  high  concentrations of nitrogen and
 nitrates may be explained in part  by the use of nitrate-based
 blasting agents.

 Process Description - Titanium Milling;  Mill  9905.   Ore
 brought to  this  mill is beneficiated by  a  combination of
 the magnetic-separation and  flotation processes  (Figure V-44).

The ore is  initially crushed and  then screened.  Both the
undersize and  the  oversize screened  ores are magnetically
cobbed to remove the  nonmagnetic rock, which is discarded.
Oversize magnetic  rock undergoes further crushing and
screening, while undersize material  is fed  into the grinding
                              V-200


                              DRAFT

-------
                          DRAFT
    TABLE V-82. CHEMICAL COMPOSITION OF RAW WASTE WATER
               FROM TITANIUM MINE 9905
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
TSS
Acidity
Alkalinity
Hardnest
COD
TOG
Oil and Create
MBAS Surfactants
Total Kjaldahl N
Al
A»
Be
Ba
a
Cd
Ca
Cr
Cu
Total Fa
CONCENTRATION (mg/£)
1.000-
11.3*
037
1.240
14
BA
1382
6464
6.4
10J
3.0
0.32
2.24
0.1
0.1
0.003
<1
0.01
< 0.002
94.6

-------
                          DRAFT
Figure V-44. BENEFICIATION AND WASTE WATER FLOW OF ILMENITE
          MINE/MILL 9905 (ROCK DEPOSIT)
MAKEUP ^V _
WAIfcH J 878n,3/dav*
' 	 (232.000 gpd)
36.069 m3/doy
(9.450.000 gpd)
i— MAGN
MAGNETITE
, * ,
DEWATERER
•-FILTRA1

c
WATER
"RETURN TO MILL"
35.191 m3/day
(8 220 000 and) l"~


MINING

ORE
CRUSHING


\
GRINDING

\
CLASSIFICATION


1
MAGNETIC
SEPARATION
ETICS 	 Z-NONMA

re-J
^

TAILINGS
1

SNETICS-i
ILMENITE
AND GANGUE
1
FLOTATION
CIRCUIT
1
THICKENER
-1-. \
TAILING A
POND )
^^ ^^^
FILTER
-=T 1
i
*
DRIER
SEASONAL I -1
DISCHARGE 1

CONCENTRATE
TO SHIPPING
                         V-202


                        DRAFT

-------
                            DRAFT
 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 iloenite  concentrate  is  filtered and dried
 prior to shipping.

 Quantities of_  Wastes:  Mill 9905.   Wastes  are discharged
 from this  mill at a rate of 35,191 cubic meters  (9,220,000
 gallons) per day.  The results of  a chemical analysis  of this
 wastewater are presented in Table  V-83.  These data are  based
 on  analysis of raw waste samples collected  during the  mill
 visit.

 Reagents consumed in  the flotation circuit  of mill 9905  are
 identified in  Table V-84.   The principal associated minerals
 in  the ore body of mine 9905 are listed in  Table V-85.   These
 reagents and constituents of the ore body  comprise the princi-
 pal 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  compila-
 tion of the concentrations  and the raw waste loading of  the
 principal  constituents of raw wastewater from mill 9905.

 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
                            V-203


                           DRAFT

-------
                           DRAFT
    TABLE V-83. CHEMICAL COMPOSITION OF RAW WASTE WATER
               FROM TITANIUM MILL 9905
PARAMETER
Conductivity
Color
Turblditv (JTU)
TOS
TSS
Acidity
Alkalinity
Hardnen
COD
TOC
Oil and Greats
IMBAS Surfactants
Total Kjeldahl N
Al
At
Ba
B
Cd
Ca
Cr
Cu
Total Fa
CONCENTRATION (mull)
650*
18.0*
13.
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
360
0.58
0.43
500
PARAMETER
Pb
Mg
Total Mn
Ni
Tl
V
K
Se
Ag
Na
Sr
Te
Ti
Zn
Mo
Co
Phenol
Chlorida
Fluoride
Sulfata
Nitrate
Phosphate
CONCENTRATION (mg/£)
< 0.05
187.5
5.9
1.19
<0.1
2.0
23.7
0.132
0.015
41
0.29
< OJ06
2.08
7.6
< 0.1
< 0.1
< 041
19.1
325
213
ojea
< 0.05
•Value in mlcromhot/cm

* Value in cobalt units
 TABLE V-84. REAGENT USE IN FLOTATION CIRCUIT OF MILL 9905
REAGENT
Tall oil
Fuel oil
Methyl amyl alcohol
Sodium bifluoride
Sulfuric acid
PURPOSE
Frother
Frother
Frother
Depressant
pH Modifier
CONSUMPTION
kg/metric ton
ore milled
1.33
0.90
0.008
0.76
1.775
Ib/short ton
ore milled
2.66
1.80
0.016
1.52
3.55
                            V-204

                           DRAFT

-------
                            DRAFT
         TABLE V-85. PRINCIPAL MINERALS ASSOCIATED WITH
                   ORE OF MINE 9905
MINERAL
llmenite
Magnetite
Pyroxene
Feldspar
COMPOSITION
FeTiOa
FeaOA
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
(mB/K) IN
WASTEWATER
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
0.005
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
                             V-205


                             DRAFT

-------
                                    DRAFT
Figure V-45. BENEFICIATION OF HEAVY-MINERAL BEACH SANDS (RUTILE, ILMENITE,
           ZIRCON, AND MONAZITE) AT MILL 9906
                                                                     TO
                                                                 ATMOSPHERE
    ORE + WATER
         t
20,100 m/day
(5,310.000 gpd)
                 ORE FED
               FROM DREDGE
                                                                    i
                  EVAPORATION
       POND
      WATER
     RECYCLE
                   H2O

                  _*_
                VIBRATING
                 SCREENS
              7,570 mj/day
             (2,000,000 gpd)
—OVERSIZE-
             WASH
            'WATER'
            SPIRALS OR LAMINAR
      FLOWS (ROUGHERS AND CLEANERS)
           11.000 m3/day
           (3.170,000 gpd)
                   T
                                                     —TAILINGSn
      12,000 m3/day
      (3.170,000 gpd)
                                 TO DRY MILL
                                 (FIGURE 111-30)
                                                           rWASTE (DREDGE)]
                                                                 POND
                                                       15.615 m3/day
                                                      (4,500,000 gpd)
                                                              DISCHARGE
                                    V-206
                                    DRAFT

-------
                             DRAFT
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 opera-
tions 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 loads of principal wastewater constituents discharged
from the milling operations at mills 9906 and 9907 are given
in Table V-88.

Zirconium

Zirconium is recovered as a byproduct of the mining and milling
of sand placer deposits, which have been described under Waste
Characteristics of Titanium Ores.  No operations for zirconium
alone are known in the United States.  The waste characteristics
and water uses accompanying mining and milling to obtain
zircon concentrate are, therefore, identical to those of
the previously described operations.
                             V-207


                            DRAFT

-------
                             DRAFT
      TABLE V-87. CHEMICAL COMPOSITION OF RAW WASTE WATER
                 AT MILLS 9906 AND 9907
PARAMETER
Conductivity
Color
Turbidity (JTU)
TDS
T8S
Acidity
Alkalinity
COD
TOC
Total KMdahl N
Oil and Greaw
MBAS Surfactant*
Al
Aa
Be
Be
B
Cd
Ca
Cr
Cu
CONCENTRATION (ma/Jll
MILL 9906
200*
51,400*
<0.1
1,644
11,000
47.2
47.6
1^38
972
0.66
400
<0.01
69.0
0.06
< 0.002
<0.6
0.10
< 0:002
0.10
QJ03
1
<0.1

-------
                            DRAFT
TABLE V-88. RAW WASTE LOADS FOR PRINCIPAL WASTEWATER CONSTITUENTS
          FROM SAND PLACER MILLS 9906 AND 9907
PARAMETER
TSS
TOC
COD
Oil and Grwu
Ti
Fe
Mn
Cr
Phaplun
MILL 0906
CONCENTRATION
(mo/ 11
IN WASTEWATER
11.000
872
1.337
400
<0.2
4.9
0.36
0.03
036
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
330
28.2
40.13
12
< 0.006
0.15
0.0011
00009
0.011
In/then ton
660
68.4
80.26
24
< 0.012
0.30
0.0022
0.0018
0022
MILL 9907
CONCENTRATION
(mg/ei
IN WASTEWATER
209
321
361.6
40
0.4
0.93
<0.01
<0.01
0.4
RAW WASTE LOAD
(per unit total concentrate produced)
kg/metric ton
6.01
7.71
8.68
036
0.01
0.022
< 0.0024
< 0.0024
0X11
Ih/ihortton
10.02
15.42
17.36
1.92
0.02
0.044
< 0.0048

-------
                            DRAFT
                          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 wastewater
constituents finally selected as being of pollution signifi-
cance for the ore mining and dressing industry are based upon
(1) those parameters which have been Identified as known con-
stituents of the ore-bearing deposits and overburden, (2)
chemicals used in processing or extracting the desired
metal(s), and (3) parameters which have been identified as
present in significant quantities in the untreated wastewater
from each subcategory of this study.  The wastewater constit-
uents are further divided into (a) those that have been
selected as pollutants of significance (with the rationale
for their selection), and (b) those that are not deemed
significant (with the rationale for their rejection).  This
Section is concluded with a summary list of the pollution
parameters selected for each category.

GUIDELINE PARAMETER-SELECTION CRITERIA-

Selection of parameters for use in developing effluent limita-
tion 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 undesir-
           able tastes and odors in water supplies.
                            VI-1

                           DRAFT

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

£H, 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.

Haters 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, sine, 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 pB close to 7.  This is very
significant for  providing safe drinking water.

Extremes of pH or rapid pB changes can exert stress condi-
tions or kill aquatic life outright.  Dead fish, associated
algal blooms, and foul atrenches are aesthetic liabilities
of any waterway.  Even moderate changes from "acceptable"
criteria limits  of pB are deleterious to some species.
The relative toxicity to aquatic life of many materials
is increased by  changes In the water pB.  Hetalocyanide
complexes can increase a thousand-fold la toxicity with a
drop of 1.5 pB units.  The availability of many nutrient
substances varies with the alkalinity and acidity.  Ammonia
is more lethal with a higher pB.
                            VX-2


                          DRAFT

-------
                           DRAFT
The lacrimal fluid of the human eye has a pH of approxi-
mately 7.0, and a deviation of 0.1 pH unit from the norm
may result in eye irritation for the swimmer.  Appreciable
irritation will cause severe pain.

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; laun-
dries; 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
                             VI-3

                           DRAFT

-------
                             DRAFT
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 those benthlc
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 Inexhaust-
ible food source for sludgeworms and associated organisms.

Turbidity is principally a measure of the light-absorbing
properties of suspended solids.  It is frequently used as
a substitute method of quickly estimating the total suspended
solids when the concentration is relatively low.

High suspended-solid concentrations are contributed as part
of the mining process, as well as the crushing, grinding,
and other processes commonly used in the milling industry
for most milling operations.  High suspended-solid concen-
trations are also characteristic of dredge-mining and gravity-
separation operations.

Oil and Grease

Oil and grease exhibit an oxygen demand.  Oil emulsions may
adhere to the gills of fish or coat and destroy algae or
other plankton.  Deposition of oil in the bottom sediments
can serve to exhibit normal benthic growths, thus Interrupt-
ing the aquatic food chain.  Soluble and emulsified material
ingested by fish may taint the flavor of the fish flesh.
Water-soluble components may exert toxic action on fish.
Floating oil may reduce the re-aeration of the water surface
and, in conjunction with emulsified oil, may interfere with
photosynthesis.  Water-insoluble 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.
                              VI-4


                              DRAFT

-------
                             DRAFT
Levels of oil and grease which are toxic to aquatic organisms
vary greatly, depending on the type and the species suscepti-
bility.  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 lubri-
cants and spillage of fuels, as well as the usage of reagents
in many milling processes.

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

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
waterwquality 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 con-
centrations that makes them less competitive and able to
sustain their species within the aquatic environment.  For
example, reduced DO concentrations have been shown to inter-
fere with fish populations through delayed hatching of eggs,
reduced size and vigor of embryos, production of deformities
in young, interference with food digestion, acceleration of
blood clotting, decreased tolerance to certain toxicants,
reduced food efficiency and growth rate, and reduced maximum
sustained swimming speed.  Fish food organisms are likewise
affected adversely in conditions with suppressed DO.  Since
all aerobic aquatic organisms need a certain amount of oxygen,
the total lack of dissolved oxygen due to a high COD can kill
all inhabitants of the affected area.
                              VI-5

                             DRAFT

-------
                            DRAFT
 The total organic  carbon (TOG)  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 contamina-
 tion.   Forms of carbon analyzed by this test, among others,
 are:   soluble,  nonvolatile organic carbon; insoluble, par-
 tially 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
 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 environ-
 mental fates and effects  are  largely  unknown.

 Cyanide

 Cyanides in  water  derive  their  toxiclty primarily from undis-
 sociated 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.
                              VI-6

                             DRAFT

-------
                           DRAFT
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.

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 non-
ionized 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 pres-
ence of dissolved oxygen, is converted to nitrate (N03^ by
nitrifying bacteria.  Nitrite (N02), 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 irri-
tation 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
                            VI-7

                           DRAFT

-------
                              DRAFT
concentration limits, it has been widely recommended that
water containing more than 10 mg/1 of nitrate nitrogen
(N03_-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 undlsso-
ciated 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 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-Leute'
ammonium carbamate process, for precipitation of metal salts,
or for pH control.   In the ore mining and dressing industry,
high levels at selected locations may thus be encountered.

Aluminum

Aluminum is one of the most abundant elements on the face of
the earth.  It occurs in many rocks and ores, but never as
a 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, how-
                              VI-8


                              DRAFT

-------
                              DRAFT
 ever,  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 con-
 centrations of aluminum.   Where  soils are quite acidic  (pH
 below  5.0), aluminum toxicity to plants becomes very  signifi-
 cant.   Aluminum presence is primarily observed in wastewaters
 from the bauxite-ore mining industry.  At pH 4.5,  1.0 mg/1
 of aluminum reduces  the yield of corn 25 percent;  at concentrations
 of 2.28 and 4.56 mg/1 of aluminum, yields are decreased by
 39 percent  and 59  percent,  respectively.

 Antimony

 Antimony is rarely found pure in nature, its common forms
 being  the sulfide, stibnite (Sb2S3)  and the  oxides  cervantite
 (Sb2_04)  and valentinite (Sb203).  Any antimony discharged to
 natural waters has a  strong tendency to precipitate and be
 removed by  sedimentation and/or  adsorption.

 Antimony compounds are  toxic  to  man  and are  classified  as acutely
 moderate or chronically severe.   A dose of 97.2 mg of antimony
 has  reportedly been  lethal  to an  adult.  Antimony potassium
 tartrate, once in  use medically  to treat certain parasitic
 diseases, is no longer  recommended because of the frequency
 and  severity of toxic reactions,   including cardiac disturbances.

 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 by-
product recovery.
                               VI-9

                              DRAFT

-------
                             DRAFT
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 arsenopyrlte  (FeS2_.FeAs2_).

Arsenic is normally present in sea water at concentrations
of 2 to 3 micrograms per liter and tends to be accumulated
by oysters and other shellfish.  Concentrations of 100 mg/kg
have been reported in certain shellfish.  Arsenic is a cumu-
lative poison with long-term chronic effects on both aquatic
organisms and mammalian species, and a succession of small
doses may add up to a final lethal dose.  It is moderately
toxic to plants and highly toxic to animals—especially, as
arsine (AsH3) .

Arsenic trioxide, which also is exceedingly toxic, was studied
in concentrations of 1.96 to 40 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 As2_03_ 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 intes-
tinal walls from the air, water, and food.  Arsenic is a normal
constituent of most soils, with concentrations ranging up to
500 mg/kg.  Although very low concentrations of arsenates may
actually stimulate plant growth, the presence of excessive
soluble arsenic in irrigation waters will reduce the yield of
crops, the main effect appearing to be the destruction of
chlorophyll in the foliage.  Plants grown in water containing
one mg/1 of arsenic trioxides show a blackening of the vascular
bundles in the leaves.  Beans and cucumbers are very sensitive,
while turnips, cereals, and grasses are relatively resistant.
Old orchard soils in Washington that contain 4 to 12 mg/kg
of arsenic trioxide in the topsoll 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
solubillzed in mining and milling by oxidation of the ore and
appear in the effluent stream.
                              VI-10

                              DRAFT

-------
                              DRAFT
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 hydrox-
ide at the normal pH range of such waters.

Absorption of beryllium from the alimentary tract is slight,
and excretion is fairly rapid.  However, as an air pollutant,
it is responsible for causing skin and lung diseases of variable
severity.

Concentrations of beryllium sulfate complexed with sodium
tartrate up to 28.5 mg/1 are not toxic to goldfish, minnows,
or snails.  The 96-hour minimum toxic level of beryllium sulfate
for fathead minnows has been found to be 0.2 mg/1 in soft water
and 11 mg/1 in hard water.  The corresponding level for beryllium
chloride is 0.15 mg/1 in soft water and 15 mg/1 in hard water.

In nutrient solution, at acid pH values, beryllium is highly
toxic to plants.  Solutions containing 15 to 20 mg/1 of beryllium
delay germination and retard the growth of cress and mustard
seeds in solution culture.  The presence of beryllium in
wastewaters was detected only in raw-waste effluents from the
mining and milling of bertrandite.

Cadmium

Cadmium in drinking water supplies is extremely hazardous to
humans, and conventional treatment, as practiced in the United
States, does not remove it.  Cadmium is cumulative in the liver,
kidney, pancreas, and thyroid of humans and other animals.
A severe bone and kidney syndrome  in Japan has been associated
with the ingestion of as little as 600 micrograms per day of
cadmium.

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.
                              VI-11


                              DRAFT

-------
                             DRAFT
Cadmium acts synergistically with other metals.  Copper and
zinc substantially increase its toxicity.  Cadmium is concen-
trated 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.

Chromium
Chromium, in its  various  valence  states, is hazardous to man.
It can produce  lung  tumors when inhaled and induces skin sensi-
tizations.  Large doses of chromates have corrosive effects
on the intestinal tract and  can cause  inflammation of the
kidneys.  Levels  of  chromate ions that have no effect on man
appear to be so low  as to prohibit determination to date.

The  toxicity of chromium  salts toward  aquatic life varies
widely with the species,  temperature,  pH, valence of the
chromium, and synergistic or antagonistic effects—especially,
that of hardness. Fish are  relatively tolerant 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 docu-
mented .
                               VI-12


                              DRAFT

-------
                             DRAFT
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 precipi-
tation of copper carbonate or other insoluble compounds.  The
sulfates of copper and zinc, and of copper and cadmium, are
synergistic in their toxic effect on fish.

Copper concentrations less than 1 mg/1 have been reported to
be toxic—particularly, in soft water—to many kinds of fish,
crustaceans, mollusks, insects, phytoplankton, and zooplankton.
Concentrations of copper, for example, are detrimental to some
oysters above 0.1 ppm.  Oysters cultured in  sea water con-
taining 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.
                             VI-13


                             DRAFT

-------
                             DRAFT
Fluorides

As the most reactive non-metal, fluorine is never found free
in nature, but rather occurs as a constituent of fluorite or
fluorspar (calcium fluoride) in sedimentary rocks and also as
cryolite (sodium aluminum fluoride) in igneous rocks.  Owing
to their origin only in certain types of rocks and only in a
few regions, fluorides in high concentrations are not a common
constituent of natural surface waters, but they may occur
in detrimental concentrations in ground waters.

Fluorides are used as insecticides, for disinfecting brewery
apparatus, as a flux in the manufacture of steel, for preserving
wood and mucilages, for the manufacture of glass and enamels,
in chemical industries, for water treatment, and for other
uses.

Fluorides in sufficient quantity are toxic to humans, with
doses of 250 to 450 mg giving severe symptoms or causing death.

There are numerous articles describing the effects of fluoride-
bearing waters on dental enamel of children; these studies
lead to the generalization that water containing less than 0.9
to 1.0 mg/1 of fluoride will seldom cause mottled enamel in
children; for adults, concentrations less than 3 or 4
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.  Concentra-
tions of 30 to 50 mg/1 of fluoride in the total ration of
dairy cows are considered the upper safe limit.  Fluoride
from waters, apparently, does not accumulate in soft tissue
to a significant degree, and it is transferred to a very small
extent into milk and, to a somewhat greater degree, into eggs.
Data for fresh water indicate that 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 aquifers or from water contact from
                              VI-14


                              DRAFT

-------
                              DRAFT
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 tnicrograms per liter,
depending on geographic location.  When the pH is low, however,
appreciable amounts of iron may remain in solution.

Standards for drinking water are not set for health reasons.
Indeed, some iron is essential for nutrition, and larger quantities
of iron are taken for therapeutic reasons.  The drinking-water
standards are set for esthetic reasons.

In general, very little iron remains in solution; but, if the
water is strongly buffered and a large enough dose is supplied,
the addition of a soluble iron salt may lower the pH of the
water to a toxic level.  In addition, a fish's respiratory
channel may become irritated and blacked by depositions of
iron hydroxides on the gills.  Finally, heavy precipitates of
ferric hydroxide may smother fish eggs.

The threshold concentration for lethality to several types
of fish has been reported as 0.2 mg/1 of iron.  Concentrations
of 1 to 2 mg/1 of iron are indicative of acid pollution and
other conditions unfavorable to fish.  The upper limit for
fish life has been estimated at 50 mg/1.  At concentrations of
iron above 0.2 mg/1, trouble has been experienced with populations
of the iron bacterium Crenothrix.

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 and dressing
industry also contributes elevated iron levels through mine
drainage.
                               VI-15

                              DRAFT

-------
                             DRAFT
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
tog/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 solu-
tion.

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 cumu-
lative '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 rag/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 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 48-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.
                              VI-16

                              DRAFT

-------
                            DRAFT
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, 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
Polycelis 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.

On the basis of the literature surveyed, it appears that the
concentrations of manganese listed below are deleterious to
the stated beneficial uses.
                             VI-17

                            DRAFT

-------
                             DRAFT
     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

Elevated manganese concentrations are found in the effluents
of iron-ore, lead, and zinc mining and milling operationsand
would be expected from any future operations exploiting mangan-
ese 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 organie
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 parti-
cularly 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 synthe-
size 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 factors
                              VI-18

                             DRAFT

-------
                             DRAFT
of 10,000 or more have been reported from water to brook trout.
The chronic effects of mercury on aquatic organisms are not
well-known.  The lowest reported levels which have resulted
in the death of fish are 0.2 micrograms per liter of mercury,
which killed fathead minnows exposed for six weeks.  Levels
of 0.1 microgram per liter decrease photosynthesis and growth
of marine algae and some freshwater phytoplankton.

Mercury has been observed in significant quantities in the
wastewater in operations associated with sulfide mineralization,
including mercury ores, lead and zinc ores, and copper ores,
as well as precious-metal operations of gold and silver.
It may be liberated in mine waters as well as in effluents
of flotation concentration and acid-leaching extraction.

Molybdenum

Molybdenum and its salts are not normally considered serious
pollutants, but the metal is biologically active.  Although
the element occurs in some minerals, it is not widely distri-
buted in nature.  The mean level of molybdenum in the U.S.
has been reported to be 68 micrograms per liter.

The 96-hour minimum toxic level of fathead minnows for molybdic
anhydride  (MoOjl) was found to be 70 mg/1 in soft water and
370 mg/1 in hard water.  The threshold concentration for dele-
terious effects upon the alga Scenedesmus occurs at 54 mg/1.
I-i 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.

Nickel

Elemental nickel seldom occurs in nature, but nickel compounds
are found  in many ores and minerals.  As a pure metal, it is
                              VI-19

                              DRAFT

-------
                             DRAFT
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 experi-
ments with plants in solution cultures have shown that nickel
at 0.5 to 1.0 mg/1 is inhibitory to growth.

Nickel salts can kill fish at very low concentrations.   Data
for the fathead minnow show death occurring in the range of 5
to 43 mg, depending on the alkalinity of the water.

Nickel is present in coastal and open ocean concentrations in
the range of 0.1 to 6.0 micrograms per liter,  although the
most common values are 2 to 33 micrograms per liter.  Marine
animals contain up to 400 micrograms per 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).  Concen-
trations 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 wastewater in the titanium, rare-earth, mercury, uranium,
and, occasionally, in the iron-ore mining and milling industries.

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
                              VI-20


                              DRAFT

-------
                             DRAFT
the water and the specific vanadium compound present.   The
common bluegill can be killed by about 6 mg/1 in soft  water
and 55 mg/1 in hard water when the vanadium is expressed as
vanadyl sulfate.  Other fish are similarly affected.

Limitation and control of vanadium levels appear to be necessary
in the effluents from operations employing leaching methods to
extract vanadium as a primary product or byproduct.  As treated
here, it can be expected to be contributed by the ferroalloy
industry, where high vanadium levels were observed both in
barren solutions from a solvent extraction circuit and in
scrubber waters from ore roasting units.  High vanadium values
are also found associated with uranium operations, where
vanadium is also obtained as a byproduct.

Zinc

Occurring abundantly in rocks and ores, zinc is readily refined
into a stable pure metal and is used extensively for galvanizing,
in alloys, for electrical purposes, in printing plates, for
dye manufacture and for dyeing processes, and for many other
industrial purposes.  Zinc salts are used in paint pigments,
cosmetics, pharamaceuticals, dyes, insecticides, and other
products too numerous to list herein.  Many of these salts
(e.g., zinc chloride and zinc sulfate) are highly soluble in
water; hence,- it is to be expected that zinc might occur in
many industrial wastes.  On the other hand, some zinc salts
(zinc carbonate, zinc oxide, and zinc sulfide) are insoluble
in water; consequently, it is to be expected that some zinc
will precipitate in and be removed readily  from most natural
waters.

In zinc-mining areas, zinc has been found in waters in concen-
trations 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.
                              VI-21 •

                              DRAFT

-------
                             DRAFT
In soft water, concentrations of zinc ranging from 0.1 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 species,
age, and conditions, as well as with the physical and chemi-
cal 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 long-
term 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 opera-
tions; and at most  ferroalloy-ore mining and milling  sites.

Radioactivity

Ionizing radiation, when  absorbed  in  living  tissue in quan-
tities substantially  above that  of natural background levels,
is  recognized as  injurious.  It  is necessary, therefore,  to
prevent excessive  levels  of  radiation from reaching any  liv-
ing organism:  humans,  fishes, or  invertebrates.  Beyond the
obvious fact  that  radioactive wastes  emit ionizing radiation,
such wastes are also  similar  in many  respects to  other chemi-
cal wastes.   Man's  senses cannot detect  radiation unless it
is  present  in massive amounts.
                              VI-22


                              DRAFT

-------
                              DRAFT
Plants and animals, to be of any significance in the cycling
of radionuclides in the aquatic environment, must accumulate
the radionuclide, retain it, be eaten by another organism,
and be digestible.  However, even if an organism accumulates
and retains a radionuclide and is not eaten before it dies,
the radionuclide will enter the "biological cycle" through
organisms that decompose the dead organic material into its
elemental components.  Plants and animals that become radio-
active in this biological cycle can thus pose a health hazard
when eaten by man.

Aquatic life may receive radiation from radionuclides present
in the water and substrata and also from radionuclides that
may accumulate within their tissues.  Humans can acquire
radionuclides through many different pathways.  Among the
most important are drinking contaminated water and eating
fish and shellfish that have concentrated nuclides from the
water.  Where fish or other fresh or marine products that
have accumulated radioactive materials are used as food by
humans, the concentrations of the nuclides in the water must
be further restricted, to provide assurance that the total
intake of radionuclides from all sources will not exceed
the recommended levels.

To prevent unacceptable doses of radiation from reaching
humans, fish, and other Important organisms, the concentra-
tions of radionuclides in water, both fresh and marine, must
be restricted.

Radium

Radium is a natural product of the disintegration of uranium.
It undergoes spontaneous disintegration with the formation of
radon (Rn 222), one gram of radium producing about 10 exp (-14)
ml of radon per day.  The mean radium concentration in
raw waters has been measured at 0.049 x 10 exp(-9) mg/1,
based on 42 sources serving approximately one-fifth of the
population of the U.S.

Radium is the most hazardous radioelement of the 14 radio-
active isotopes that occur in nature with uranium as their
parent.  The bulk of the radium discarded as waste is gener-
ally retained undlssolved in tailing ponds;  however, the
remaining dissolved portion can constitute a significant
stream-pollution problem, affecting domestic and industrial
                              VI-23

                              DRAFT

-------
                            DRAFT
water supplies, irrigation, stock watering, and aquatic life.
In addition, dissolved radium can also escape to streams and
build up in bottom muds.  When first introduced into natural
waters, a large part of the materials present in radioactive
wastes, including radium, becomes associated with suspended
organic and inorganic particulates that settle to the bottom,
and many radloisotopes are eventually bound chemically to the
sediments.

Radium, like the other radioisotopes, is passed through the
various trophic levels of the food chain and are either bio-
concentrated or released.  Possible effects to an individual
organism exposed to radium may include death, inhibition
or stimulation of growth, physiological change, behavioral
changes, developmental abnormalities, or shortening of life
span.

Radium is present in mine and process waters of the uranium
industry because of the radioactive decay of uranium isotopes.

Thorium

Thorium is a grayish-white, lustrous metal which occurs
naturally in several mineral forms, such as the silicate
(thorite), the oxide (thortanite) , and the phosphate (mona-
zite) .  The toxicity of thorium, based on radiation effects,
is estimated to be three times that of uranium.

Besides radioactive toxicity, thorium minerals also exhibit
chemical toxicity.  The median threshold effect of thorium
nitrate on Scenedesmus . E^. coll. and a protozoan (Mlcroregna)
was found to be 0.4 to 0.8 mg/1 of thorium.  The lethal con-
centration of thorium chloride for three mature, small fresh-
water fish exposed for 24 hours was reported to be about 18
Thorium is present in mine and process waters of the uranium
mining and processing industry.

Uranium

Uranium is present in wastes from uranium mines and mills
and nuclear fuel processing plants, and the uranyl ion may
                            VI-24


                            DRAFT

-------
                             DRAFT
naturally occur in drainage waters from uranium-bearing ore
deposits.

Many of the salts of uranium are present in sea water,
which has an average concentration of the metal of about 3
micrograms per liter.  Uranium is stabilized by hydrolysis,
which tends to prevent its physical and chemical interaction
and, thus, prevents its removal from sea water.  The uranium
salts are considered to be four times as germicidal as phenol
to aquatic organisms.

Natural uranium (U 238) is concentrated by the algae Ochro-
monas by a factor of 330 in 48 hours.  The threshold of
uranyl nitrate, expressed as uranium, was found to be 28 mg/1
for a protozoan (Microregma), 1.7 to 2.2 mg/1 for Escherichia
coli, 22 mg/1 for the alga Scenedesmus, and 13 mg/1 for Daphnia.
The nitrate, sulfate, and acetate salts of uranium have been
found to be more toxic to fathead minnows, Pimephales promelas,
in soft water than in hard water, with the 96-hour minimum
toxic level for uranyl sulfate being 2.8 mg/1 in soft water
and 135 mg/1 in hard water.

A factor of 0.01 has been recommended for application to
marine 96-hour minimum toxic-level data for sensitive organisms.

The radiation hazard from uranium is considered to be much
less than that of radium; in addition, uranium has a much
shorter "biological" half-life and affects less-sensitive
tissues (gastrointestinal tract).  The chemical toxicity is
believed to exceed the radiation hazard from ingestion.

Uranium concentrations in raw mine/mill wastes are often
below 40 mg/1, and discharge levels below 1 mg/1 are routinely
achieved.

Flotation Reagents

The toxicity of organic floation agents—particularly, collec-
tors and their decomposition products—is an area of consider-
able uncertainty, particularly in the complex chemical environ-
ment 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.
                            VI-25


                            DRAFT

-------
                             DRAFT
Data available on  the  fates and potential toxicities of many
of the reagents  indicate  that only a broad range of tolerance
values is known.   Table VI-1 is a list of some of the more
common flotation reagents and their known toxicities as judged
from organism tolerance information.

Asbestos

"Asbestos"  is a  generic term for a number of fire-resistant
hydrated silicates that,  when crushed or processed, separate
into flexible fibers made up of fibrils noted for their great
tensile strength.   The asbestos minerals differ in their
metallic elemental content, range of fiber diameters, flexi-
bility, hardness,  tensile strength, surface properties, and
other attributes which may affect their respirability, deposi-
tion, 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 asbes-
tos fibers within  the  human body, including their potential
entry through the  gastrointestinal tract.  There is evidence
that bundles of  fibrils may be broken down within the body
to individual fibrils.  Asbestos has the possibility of being
a hazard when waterborne  in large concentrations; however, it
is insoluble in  water.

To date, there is  little  data on the concentrations of asbestos
in ore mining and  milling water discharges.  Knowledge of the
concentrations in  water that pose health problems is poorly
defined.  Currently, this area is being investigated by many
researchers  concerning themselves with health, movement, and
analytical techniques.

Because of public  reports concerning the presence of asbestos
in wastewater from an  iron-ore beneficiation operation, a
reconaissance analysis for asbestos was performed on samples
collected as part  of site visits to four discharging iron-ore
beneficiation operations.  The raw wastewater and effluent
of tailing ponds at each  facility were examined for the pre-
sence or absence of asbestos or asbestos-like fibers.  The
method of analysis used for detection was one based upon pub-
lished literature  and  employed scanning electron microscopy.
                             VI-26

                            DRAFT

-------
                                         DRAFT
  TABLE Vl-l. KNOWN TOXIC ITY OF SOME COMMON FLOTATION REAGENTS
                USED IN ORE MINING AND MILLING INDUSTRY
TRADE NAME
Aerofloat 26
Aerofloat 31
Aarof loot 238
Aeroftaat242
AerofrothW
AerofrothTI
AflVO ^TOHIOlVf
404
3477
AROSURF
MO-MA
Dowf roth 260
DowZ-6
OowZ-11
DowZ-200
Jaguar
M.I.B.C.
-
Suparfloc 16
CHEMICAL COMPOSITION
FIT Mitt laHv onrf ffltlilnirfimntwirir arlri
Etnntlally «ryl dithiophoaphorlc acid
Sodium dl-taeondary butyl
dithlophosphate
essentially aryl dithlophosphorlc MM
PMydJyeol type compound
Mixture of tulf hydryl type compounds
Unknown
Unknown
Chromium tills (ammonium, potassium.
•nd todlum chromota and ammonium,
potassium, and codlum dkhromata)
Copper ulfata
Cresylic acid
Polypropylene dlycol methyl •than
Potassium amyl xanthata
Sodium bopropyl xanthata
liopropyl ethylthionocarfaamate
Basad on guar gum
Uma (calcium oxide)
Mathvllwbutylcarblnol
Pinaoll
Potaolum farrleyanida
Sodium farrocyanlda
Sodium hydroxide
Sodium olaata
Sodium illicata
Sodium aulflda
Sulfuric acid
Polyaerylamida
FUNCTION
Collertor/Promoter
Collector/Promoter
Collector/Promoter
Frothor
Frothar
Collect or/Promoter
Collect or/Promotar
11 m
Oepreolng agent
Activetlng agent
Frothar
Frother
Collector/Promoter
Collector/Promoter
Collector/Promoter
Floeculant
pH modlfiar and
floeculant
Frother
Frother
Dapiming agent
DepraBlng agent
pH (nooificr
Frother
Deprenlng agent
Activating agent
pH modlfiar and
floeculant
Flocculant
KNOWN TOXIC
RANGE imafi)
1000 to 10^00
10 to 1000
1000 to \OJOOO
>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
Oi to 2.0
10 to 100
10 to 1000
>1000
1 to 100
0.25 to 2.6
1 to 1600
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
Modorata
Moderate
Low
Moderate
Moderate to High
Moderate
Moderate
Moderate
Moderate
Moderate
Moderate
Low
Texterty   Tolerance Level

 High       <1.0 mg/l
Moderate    I.Oto lOOOmg/Jt
 Low
           >1000mg/£
NOTE:  Toxic range » a function of orginlim tailed and water quality. Including hardneai
       and pH. Therefore, toxiclty data presented in this table aro only generally indica-
       tive of reagent toxiclty. Although the toxiclty ration presented here are based on
       many different organism*, much of the data are presented in relation to almon,
       fathead minnows, sticklebacks, and Daphnia.
                                          VI-27


                                         DRAFT

-------
                           DRAFT
Fibers were not detected in any of the samples with the excep-
tion of the influent to the tailing pond from Mill 1107.
Energy-dispersive x-ray analysis indicated, however,  that
the fiber was not of an asbestos type.  Both raw and  treated
wastewaters from mills 1107, 1108, 1109, and 1110 were examined,
and no asbestos or asbestos-like minerals were found.

While the results of the survey indicate the absence  of asbes-
tos 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, how-
ever, that a reconalssance evaluation for asbestos be performed
at each iron-ore mining and beneficiation operation to deter-
mine whether possible asbestos levels of concern are  present.
                             VI-28

                            DRAFT

-------
                           DRAFT
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.

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.
                            VI-29

                           DRAFT

-------
                             DRAFT
Nitrate and Nitrite

There are Insufficient data for dissolved nitrates and nitrites
to justify their consideration as harmful pollutants, although
nitrogen and nitrate contributions are known to stimulate
plant and algal growth.  There is no treatment available to
practically reduce these ions.

Selenium

The levels of selenium observed in the wastewaters from mines
and mills are not sufficiently high for selenium to be con-
sidered as a harmful pollutant.

Silicates

Silicates may be present in the wastewaters from the ore min-
ing 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 signi-
ficant levels of zirconium are present in the industry to be
classed as harmful.

Total Dissolved Solids

High dissolved-solid concentrations are often caused by acid
conditions or by the presence of easily dissolved minerals
in the ore.  Since economic methods of dissolved-solid reduction
do not exist, effluent limitations have not been proposed
for this parameter.
                             VI-30

                            DRAFT

-------
                            DRAFT
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 differ-
ent 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.
                              VI-31

                           DRAFT

-------
                                      DRAFT
      TABLE VI-2. SUMMARY OF PARAMETERS SELECTED FOR EFFLUENT

                   LIMITATION BY METAL CATEGORY
      PARAMETERS
                                PARAMETERS SELECTED FOR EFFLUENT LIMITATIONS
Ore
Ores
Uranium, Radium,
and Vanadium Ores
                                                          Metal Ores, Not Elsewhere Classified
Anti

Ores
Plati

Ores
pH (Acidity/ Alkalinity)
Total Suspended Solids (TSS)
Oil and Grease
*
Chemical Oxygen Demand (COO)
Lit I
   n
       I
i
i
Total Organic Carbon (TOC)
Cyanide
Ammonia
Aluminum
Antimony
Arsenic
Beryllium
Cadmium
Chromium
Copper
Fluoride
Lead
Manganese
Mercury
Molybdenum
Nickel
Vanadium
Zinc
Radium
Uranium
i
                      i
              i
                        IE' H 1C 11 If  11 !
                      i
                      i
                      i
t
i
i
                            i
                                                       a
                                                       i
                            i
                                                   ^.
                                                   o
                                                   1
                                                       i
                                                       i
                                       VI-32




                                      DRAFT

-------
                              DRAFT
                         SECTION VII

               CONTROL AND TREATMENT TECHNOLOGY
INTRODUCTION

Waterborne wastes from the mining of metal-ore minerals consist
primarily of suspended solids and metals in solution.   The
mineralogy of the ore and associated overburden and the chemical
character of percolating mine waters influence the metal con-
tent of mine wastewater, while solids suspended in the wastewater
are influenced by the methods of mining as well as the physical
nature and general geologic characteristics of the ore.  Sep-
arate treatment of mine water by methods other than pH control,
Impoundment for solids removal, or combined treatment  with mill
effluents is not common.

The wastewaters from ore milling and beneflciation operations
are characterized by high suspended-solid loads, heavy metals
in solution, dissolved solids, and process reagents added
during the concentration process.  Impoundment and settling-
pond facilities, primarily for suspended-solid removal, are in
widespread use in the treatment of mill effluents, and this
treatment technology is effective in removal of other  waste-
water components as well.  Space requirements and location
often affect the utilization of this widespread treatment
technology and dictate the economics of the operations.
Other treatment technologies for removal of dissolved  com-
ponents are, for the most part, well-known but are often
limited in usage by the large volumes of wastewater to be
treated and the costs of such large-scale operations.

The control and treatment of the waterborne wastes found in
the mining and beneficiation of metal-ore minerals are influ-
enced by several factors:

      (1)  Large volumes of mine water and wastewater  from
           ore-concentrating operations to be controlled and
           treated.

      (2)  Seasonal, as well as daily, variations in the
           amount and chemical characteristics of mine water
           influenced by precipitation, runoff, and
           underground-water contributions.
                              VII-1

                              DRAFT

-------
                            DRAFT
      (3)  Differences in wastewater composition caused by
           ore mineralogy and processing techniques and
           reagents.

      (4)  Geographic location and climatic conditions.
           (Treatment and control technology selection and
           economics are influenced by the amount of water
           to be handled.)

CONTROL PRACTICES  AND TECHNOLOGY

Control technology, as discussed in this report, includes
techniques and practices employed before, during, and after
the actual mining or milling operation to reduce or eliminate
adverse environmental effects resulting from the discharge
of mine or mill wastewater.  Effective pollution-control
planning can reduce pollutant contributions from active
mining and milling sites and can also minimize post-operational
pollution potential.  Because pollution potential may not cease
with closure of a mine or mill, control measures also refer to
methods practiced after an operation has terminated production
of ore or concentrated product.  The presence of pits, storage
areas for spoil (non-ore material, or waste), tailing ponds,
disturbed areas, and other results or effects of mining or
milling operations necessitates integrated plans for reclamation,
stabilization, and control to return the affected areas to a
condition at least fully capable of supporting the uses which it
was capable of supporting prior to any mining and to achieve a
stability not posing any threat of water diminution, or
pollution and to minimize potential hazards associated with
closed operations.

Mining Techniques

Mining techniques can effectively reduce amounts of pollutants
coming from a mine area by containment within the mine area or
by reducing their formation.  These techniques can be combined
with careful reclamation planning and Implementation to provide
maximum at-source pollution control.

Several techniques have been implemented to reduce environmental
degradation during strip-mining operations.  Utilization of the
box-cut technique in moderate- and shallow-slope contour mining
                             VII-2

                             DRAFT

-------
                               DRAFT
has increased recently because more stringent environmental
controls are being Implemented.

A box cut is simply a contour strip mine in which a low-wall
barrier is maintained.  Spoil may be piled on the low wall side.
This technique significantly reduces the amount of water dis-
charged from a pit area, since that water is prevented from
seeping through spoil banks.  The problems of preventing slides,
spoil erosion, and resulting stream sedimentation are still
present, however.

Block-cut mining was developed to facilitate regrading, mini-
mize overburden handling, and contain spoil within mining
areas.  In block-cut mining, contour stripping is typically
accomplished by throwing spoil from the bench onto downslope
areas.  This downslope material can slump or rapidly erode
and must be moved upslope to the mine site if contour regrading
is desired.  The land area affected by contour strip mining is
substantially larger than the area from which the ores are
extracted.  When using block-cut mining, only material from
the first cut Is deposited in adjacent low areas.  Remaining
spoil is then placed in mined portions of the bench.  Spoil
handling is restricted to the actual pit area for all areas
but the first cut, which significantly reduces the area dis-
turbed .

Pollution-control technology in underground mining is largely
restricted to at-source methods of reducing water influx into
mine workings.  Infiltration from strata surrounding the
workings is the primary source of water, and this water reacts
with air and sulfide minerals within the mines to create acid-
pH conditions and, thus, to Increase the potential for solubili-
zatlon of metals.  Underground mines are, therefore, faced
with problems of water handling and mine-drainage treatment.
Open-pit mines, on the other hand, receive both direct rainfall
and runoff contributions, as well as infiltrated water from
intercepted strata.

Infiltration in underground mines generally results from rainfall
recharge of a ground-water reservoir.  Rock fracture zones,
joints, and faults have a strong influence on ground-water
flow patterns since they can collect and convey large volumes
of water.   These zones and faults can intersect any portion
of an underground mine and permit easy access of ground water.
In some mines, infiltration can result in huge volumes of
                               VI I-3

                              DRAFT

-------
                             DRAFT
 water that must be handled and treated.   Pumping can be a
 major part of the mining operation In terms  of equipment and
 expense—particularly,  In mines which do not discharge by
 gravity.

 Water-Infiltration control techniques, designed to reduce
 the amount of water entering the workings, are extremely
 Important In underground mines located In or adjacent to
 water-bearing strata.   These techniques  are  often employed
 in  such mines to decrease the volume  of  water requiring handling
 and treatment, to make  the mine workable, and to control
 energy costs associated with dewatering.  The techniques
 Include pressure grouting of fissures which  are- entry points
 for water Into the mine.   New polymer-based  grouting materials
 have been developed which should improve the effectiveness
 of  such grouting procedures.   In severe  cases,  pilot holes
 can be drilled ahead of actual mining areas  to determine
 if  excessive water is likely to be encountered.   When water
 is  encountered,  a small pilot hole can be easily filled by
 pressure  grouting,  and  mining activity may be directed toward
 non-water-contributing  areas in the formation.   The feasi-
 bility of such control  is a function  of  the  structure of the
 ore body,  the type of surrounding rock,  and  the characteristics
 of  ground water in the  area.

 Decreased water volume,  however,  does not necessarily mean
 that  wastewater pollutant loading will also  decrease.  In
 underground mines,  oxygen,  In the presence of humidity,
 interacts  with minerals on  the mine walls and floor to permit
 pollutant  formation e.g., acid mine water, while water flowing
 through the mine transports pollutants to the outside.  If the
 volume of  this water is decreased but the volume of pollutants
 remains unchanged,  the  resultant  smaller discharge will contain
 increased  pollutant concentrations, but approximately the same
 pollutant  load.   Rapid  pumpout of the mine can,  however, reduce
 the contact time and significantly reduce the  formation of
 pollutants.

 Reduction  of mine discharge volume can reduce water handling
 costs.  In cases of acid  mine drainage, for  example, the same
 amounts of neutralizing agents  will be required  because pollu-
 tant  loads  will  remain unchanged.  The volume of mine water
 to be  treated, however, will  be reduced significantly, together
with  the size  of the necessary  treatment and  settling facilities.
 This cost  reduction, along with cost  savings which can be
                              VII-4

                             DRAFT

-------
                             DRAFT
attributed to decreased pumping volumes (hence, smaller pumps,
lower energy requirements, and smaller treatment facilities),
makes use of water infiltration-control techniques highly
desirable.

Water entering underground mines may pass vertically through
the mine roof from rock formation above.  These rock units
may have well-developed joint systems (fractures along which
no movement occurs), which tend to facilitate vertical flow.
Roof collapses can also cause widespread fracturing in over-
lying rocks, as well as joint separation far above the mine
roof.  Opened joints may channel flow from overlying aquifers
(water-bearing rocks), a flooded mine above, or even from
the surface.

Fracturing of overlying strata is reduced by employing any
or all of several methods:  (1) Increasing pillar size; (2)
Increasing support of the roof; (3) Limiting the number of
mine entries and reducing mine entry widths; (4) Backfilling
of the mined areas with waste material.

Surface mines are often responsible for collecting and
conveying large quantities of surface water to adjacent or
underlying underground mines.  Ungraded surface mines often
collect water in open pits when no surface discharge point
is available.  That water may subsequently enter the ground-
water system and then percolate into an underground mine.
The influx of water to underground mines from either active
or abandoned surface mines can be significantly reduced
through implementation of a well-designed reclamation plan.

The only actual underground mining technique developed
specifically for pollution control is preplanned flooding.
This technique is primarily one of mine design, in which a
mine is planned from its inception for post-operation
flooding or zero discharge.  In drift mines and shallow slope
or shaft mines, this is generally achieved by working the
mine with the dip of the rock (inclination of the rock to
the horizontal) and pumping out the water which collects
in the shafts.  Upon completion of mining activities, the
mine is allowed to flood naturally, eliminating the possibility
of acid formation caused by the contact between sulfide
minerals and oxygen.  Discharges, if any, from a flooded
mine should contain a much lower pollutant concentration.
A flooded mine may also be sealed.
                             VII-5

                            DRAFT

-------
                             DRAFT
 Surface-Water Control

 Pollution-control technology related to mining areas,  ore-
 beneficiation facilites, and waste-disposal sites  is generally
 designed for prevention of pollution of surface waters (i.e.,
 streams, impoundments, and surface runoff).  Prior planning
 for waste disposal is a prime control method.   Disposal sites
 should be Isolated from surface flows and impoundments to
 prevent or minimize pollution potential.  In addition,  several
 techniques are practiced to prevent water pollution:

      (1)  Construction of a clay or other type of  liner
           beneath the planned waste disposal area  to prevent
           infiltration of surface water (precipitation)  or
           water contained in the waste into the ground-water
           system.

      (2)  Compaction of waste material to reduce infiltration.

      (3)  Maintenance of uniformly sized refuse to enhance
           good compaction (which may require additional
           crushing).

      (4)  Construction of a clay liner over the material
           to  minimize infiltration.   This is usually succeeded
           by  placement of topsoll and seeding  to establish
           a vegetative cover for erosion protection and
           runoff  control.

      (5)  Excavation of diversion ditches surrounding the
           refuse  disposal site to exclude surface runoff
           from the area.   These ditches can also be used
           to  collect seepage from refuse piles,  with subse-
           quent treatment,  if necessary.

Surface  runoff in the immediate area of beneficiation facili-
ties presents  another potential pollution problem.   Runoff
from haul  roads,  areas near conveyors,  and  ore  storage piles
is a potential source of  pollutant loading  to nearby surface
waters.  Several  current  industry practices  to  control this
pollution  are:
                              VII-6


                             DRAFT

-------
                             DRAFT
      (1)  Construction of ditches surrounding storage areas
           to divert surface runoff and collect seepage that
           does occur.

      (2)  Establishment of a vegetative cover of grasses in
           areas of potential sheet wash and erosion to
           stabilize the material, to control erosion
           and sedimentation, and to Improve the aesthetic
           aspects of the area.

      (3)  Installation of hard surfaces on haul roads,
           beneath conveyors, etc., with proper slopes to
           direct drainage to a sump.  Collected waters
           may be pumped to an existing treatment facility
           for treatment.

Another potential problem associated with construction of
tailing-pond treatment systems is the use of existing valleys
and natural drainage areas for impoundment of mine water or
mill process wastewater.  The capacity of these Impoundment
systems frequently is not large enough to prevent high
discharge flow rates—particularly, during the late winter
and early spring months.  The use of ditches, flumes, pipes,
trench drains, and dikes will assist in preventing runoff
caused by snowmelt, rainfall, or streams from entering
Impoundments.  Very often, this runoff flow is the only
factor preventing attainment of zero discharge.  Diversion of
natural runoff from impoundment treatment systems, or con-
struction of these facilities in locations which do not
obstruct natural drainage, is therefore, desirable.

Ditches may be constructed upslope from the impoundment to
prevent water from entering it.  These ditches also convey
water away and reduce the total volume of water which must be
treated.  This may result in decreased treatment costs, which
could offset the costs of diversion.

Segregation or Combination £f_ Mine and Mill Wastewaters

A widely adopted control practice in the ore mining and dressing
industry is the use of mine water as a source of process water.
In many areas, this is a highly desirable practice, because
it serves as a water-conservation measure.  Waste constituents
may thus be concentrated into one waste stream for treatment.
                              VII-7

                              DRAFT

-------
                              DRAFT
In other  cases,  however,  this practice  results In the necessity
for discharge  from a mill-water  impoundment system because,
even with recycle  of part of the process water, a net positive
water balance  results.

At several sites visited  as part of  this study, degradation
of the mine water  quality is caused  by  combining the waste-
water streams  for  treatment at one location.  A negative
efxect results because water with low pollutant loading serves
to dilute water  of higher pollutant  loading.  This often
results in decreased water-treatment efficiency because con-
centrated waste  streams can often be treated more effectively
than dilute waste  streams.  The  mine water in these cases may
be treated by  relatively  simple  methods; while the volume of
wastewater treated in the mill impoundment system will be
reduced,  this  water will  be treated with increased efficiency.

There are also locations  where the use  of mine water as process
water has resulted in an  improvement in the ultimate effluent.
Choice of the  options to  segregate or combine wastewater
treatment for  mines and mills must be made on an individual
basis, taking  into account the character of the wastewater
to be treated  (at  both the mine  and the mill), the water
balance in the mine/mill  system,  local  climate, and topography.
The ability of a particular operation to meet zero or reduced
effluent  levels  may be dependent  upon this decision at each
location.

Regradlnfl

Surface mining may often  require  removal of large amounts
of overburden  to expose the ores  to be  exploited.  Regrading
Involves mass  movement of material following ore extraction
to achieve  a more  desirable land  configuration.  Reasons for
regrading  strip  mined land are:

      (1)   aesthetic Improvement  of land surface
      (2)   returning usefulness  to land
      (3)   providing a suitable base for revegetation
      (4)   burying  pollution-forming materials, e.g.  heavy metals
      (5)   reducing erosion and subsequent sedimentation
      (6)   eliminating landslidlng
      (7)  encouraging natural drainage
      (8)   eliminating ponding
      (9)  eliminating hazards such as high cliffs and deep pits
     (10)  controlling water pollution
                              VII-8

                              DRAFT

-------
                            DRAFT
Contour regrading is currently the required reclamation
technique for many of the nations's active contour and area
surface mines.  This technique involves regrading a mine to
approximate original land contour.  It is generally one of
the most favored and aesthetically pleasing regrading tech-
niques because the land is returned to its approximate pre-
mined state.  This technique is also favored because nearly
all spoil is placed back in the pit, eliminating oversteepened
downslope spoil banks and reducing the size of erodable re-
claimed area.  Contour regrading facilitates deep burial of
pollution-forming materials and minimizes contact time
between regraded spoil and surface runoff, thereby reducing
erosion and pollution formation.

However, there are also several disadvantages to contour
regrading that must be considered.  In area and contour
stripping, there may be other forms of reclamation that provide
land configurations and slopes better suited to the intended
uses of the land.  This can be particularly true with steep-
slope contour strips, where large, high walls and steep final
spoil slopes limit application of contour regrading.  Mining
is, therefore, frequently prohibited in such areas, although
there may be other regrading techniques that could be
effectively utilized.  In addition, where extremely thick
ore bodies are mined beneath shallow overburden, there may
not be sufficient spoil material remaining to return the land
to the original contour.

There are several other reclamation techniques of varying
effectiveness which have been utilized in both active and
abandoned mines.  These techniques include terrace, swale,
swallow-tail, and Georgia V-ditch, several of which are quite
similar in nature.  In employing these techniques, the upper
high-wall portion is frequently left exposed or backfilled
at a steep angle, with the spoil outslope remaining somewhat
steeper than the original contour.  In all cases, a terrace of
some form remains where the original bench was located, and
there are provisions for rapidly channeling runoff from the
spoil area.  Such terraces may permit more effective utiliza-
tion of surface-mined land in many cases.

Disposal of excess spoil material is frequently a problem
where contour backfilling is not practiced.  However, the
                             VII-9

                            DRAFT

-------
                              DRAFT
same problem can also occur, although less commonly, where
contour regrading  is in use.  Some types of overburden rock—
particularly, tightly packed sandstones—substantially expand
in volume when they are blasted and moved.  As a result,
there may be a large volume of spoil material that cannot be
returned to the pit area, even when contour backfilling is
employed.  To solve this problem, head-of-hollow fill has been
used for overburden storage.  The extra overburden is placed
in narrow, steep-sided hollows in compacted layers 1.2 to 2.4
meters (4 to 8 feet) thick and graded to control surface drainage.

In this regrading  and spoil storage technique, natural ground
is cleared of woody vegetation, and rock drains are constructed
where natural drains exist, except in areas where Inundation
has occurred,  This permits ground water and natural percola-
tion to leave fill areas without saturating the fill, thereby
reducing potential landslide and erosion problems.  Normally,
the face of the fill is terrace graded to minimize erosion of
the steep outslope area.

This technique of  fill or spoil material deposition has been
limited to relatively narrow, steep-sided ravines that can be
adequately filled  and graded.  Design considerations include
the total number of acres in the watershed above a proposed
head-of-hollow fill, as well as the drainage, slope stability,
and prospective land use.  Revegetation usually proceeds as
soon as erosion and siltation protection have been completed.
This technique is  avoided in areas where under-drainage materials
contain high concentrations of pollutants, since the resultant
drainage would require treatment to meet pollution-control
requirements.

Erosion Control

Although regrading is the most essential part of surface-mine
reclamation, it cannot be considered a total reclamation tech-
nique.  There are  many other facets of surface-mine reclamation
that are equally important in achieving successful reclamation.
The effectivenesses of regrading and other control techniques
are Interdependent.  Failure of any phase could severly reduce
the effectiveness  of an entire reclamation project.

The most Important auxiliary reclamation procedures employed
at regraded surface mines or refuse areas are water diversion
and erosion and runoff control.  Water diversion involves
                              VII-10

                              DRAFT

-------
                             DRAFT
collection of water before It enters a mine area and conveyance
of that water around the mine site, as discussed previously.
This procedure decreases erosion and pollution formation.
Ditches are usually excavated upslope from a mine site to
collect and convey water.  Flumes and pipes are used to carry
water down steep slopes or across regraded areas.  Riprap and
dumped rock are sometimes used to reduce water velocity in the
conveyance system.

Diversion and conveyance systems are designed to accommodate
predicted water volumes and velocities.  If the capacity of
a ditch is exceeded, water erodes the sides and renders the
ditch ineffective.

Water diversion is also employed as an actual part of the
mining procedure.  Drainways at the bases of high walls intercept
and divert discharging ground water prior to its contact with
pollution-forming materials.  In some instances, ground water
above the mine site is pumped out before it enters the mine
area, where it would become polluted and require treatment.
Soil erosion is significantly reduced on regraded areas by
controlling the course of surface-water runoff, using inter-
ception channels constructed on the regraded surface.

Water that reaches a mine site, such as direct rainfall, can
cause serious erosion, sedimentation, and pollution problems.
Runoff-control techniques are available to effectively deal
with this water, but these techniques may conflict with
pollution-control measures.  Control of chemical pollutants
forming at a mine frequently involves reduction of water
infiltration, while runoff controls to prevent erosion usually
increase infiltration, which can subsequently increase
pollutant formation.

There are a large number of techniques in use for controlling
runoff, with highly variable costs and degrees of effectiveness.
Mulching is sometimes used as a temporary measure which protects
the runoff surface from raindrop impacts and reduces the velo-
city of surface runoff.

Velocity reduction is a critical facet of runoff control.
This is accomplished through slope reduction by terracing or
grading;  revegetatlon; or use of flow impediments such as
dikes, contour plowing, and dumped rock.  Surface stabilizers
have been utilized on the surface to temporarily reduce eroda-
bility of the material itself,  but expense has restricted use
of such materials in the past.
                             VII-11

                             DRAFT

-------
                             DRAFT
Revegetation

Establishment of good vegetative  cover  on a mine area Is
probably the most effective  method  of controlling runoff and
erosion.   A critical factor  In  mine revegetatlon Is the quality
of  the  soil or spoil material on  the surface of a regraded mine.
There are several methods  by which  the  nature of this material
has been controlled.  Topsoll segregation during stripping Is
mandatory In many states.  This permits topsoll to be replaced
on  a regraded surface prior  to  revegetatlon.  However, In
many forested,  steep-sloped  areas,  there Is little or no top-
soil on the undisturbed  land surface.   In such areas, overburden
material Is segregated In  a  manner  that will allow the most
toxic materials to be placed at the base of the regraded mine,
and the best spoil material  Is  placed on the mine surface.

Vegetative cover provides  effective erosion control; contri-
butes significantly to chemical pollution control; results
in  aesthetic improvement;  and can return land to agricultural,
recreational,  or silvicultural  usefulness.  A dense ground
cover stabilizes the surface (with  its  root system), reduces
velocity of surface runoff,  helps build humus on the surface,
and can virtually eliminate  erosion.  A soil profile begins
to  form,  followed by a complete soil ecosystem.  This soil
profile acts as an oxygen  barrier,  reducing the amount of
oxygen  reaching underlying materials.   This, in turn, reduces
oxidation,  which is a major  contributing factor to pollutant
formation.

The soil  profile also tends  to  act  as a sponge that retains
water near  the  surface,  as opposed  to the original loose spoil
(which  allowed  rapid infiltration).  This water evaporates
from the mine  surface, cooling  it and enhancing vegetative
growth.   Evaporated water  also  bypasses toxic materials under-
lying the  soil,  decreasing pollution production..  The vegetation
itself  also  utilizes large quantities of water in its life
processes and  transpires it  back  to  the atmosphere, again
reducing  the amount of water reaching underlying materials.

Establishment of an adequate vegetative cover at a mine site
is  dependent on a number of  related  factors.  The regraded
surface of many spoils cannot support a good vegetative cover
without supplemental treatment.   The surface texture is often
                             VII-12


                             DRAFT

-------
                             DRAFT
too irregular, requiring the use of raking to remove as much
rock as possible and to decrease the average grain size of
the remaining material.  Materials toxic to plant life, usually
buried during regrading, generally do not appear on or near
the final graded surface.  If the surface is compacted, it is
usually loosened by discing, plowing, or roto-tilling prior
to seeding in order to enhance plant growth.

Soil supplements are often required to establish a good vege-
tative cover on surface-mined lands and refuse piles, which are
generally deficient in nutrients.  Mine spoils are often acidic,
and lime must be added to adjust the pH to the tolerance range
of the species to be planted.  It may be necessary to apply
additional neutralizing material to revegetated areas for some
time to offset continued pollutant generation.

Several potentially effective soil supplements are currently
undergoing research and experimentation.  Flyash is a waste
product of coal-fired boilers and resembles soil with respect
to certain physical and chemical properties.  Flyash is often
alkaline, contains some plant nutrients, and possesses moisture-
retaining and soil-conditioning capabilities.  Its main function
is that of an alkalinity source and a soil conditioner, although
it must usually be augmented with lime and fertilizers.  How-
ever, flyash can vary drastically in quality—particularly,
with respect to pH—and'may contain leachable materials capable
of producing water pollution.  Future research, demonstration,
and monitoring of flyash supplements will probably develop the
potential use of such materials.

Limestone screenings are also an effective long-term neutra-
lizing agent for acidic spoils.  Such spoils generally continue
to produce acidity as oxidation continues.  Use of lime for
direct planting upon these surfaces is effective, but it
provides only short-term alkalinity.  The lime is usually
consumed after several years, and the spoil may return to its
acidic condition.  Limestone screenings are of larger particle
size and should continue to produce alkalinity on a decreasing
scale for many years, after which a vegetative cover should be
well-established.  Use of large quantities of limestone should
also add alkalinity to receiving streams.  These screenings
are often cheaper than lime, providing larger quantities of
alkalinity for the same cost.  Such applications of limestone
are currently being demonstrated in several areas.
                             VII-13

                             DRAFT

-------
                              DRAFT
Use of digested  sewage sludge  as  a  soil  supplement also has
good possibilities  for replacing  fertilizer and  simultaneously
alleviating  the  problem of  sludge disposal.   Sewage sludge
is currently being  utilized for revegetation  in  strip-mined
areas of Ohio.   Besides supplying various nutrients, sewage
sludge can reduce acidity or alkalinity  and effectively increase
soil absorption  and moisture-retention capabilities.  Digested
se./age sludge can be applied in liquid or dry form and must
be incorporated  into the spoil surface.  Liquid  sludge applica-
tions require large holding ponds or  tank trucks, from which
sludge is pumped and sprayed over the ground,  allowed to dry,
and disced into  the underlying material.  Dry sludge application
requires dryspreadlng machinery and must be followed by discing.

Limestone, digested sewage  sludge,  and flyash are all limited
by their availabilities and chemical  compositions.  Unlike
commercial fertilizers, the chemical  compositions of these
materials may vary  greatly,  depending on how  and where they
are produced.  Therefore, a nearby  supply of  these supplements
may be useless if it does not  contain the nutrients or pH
adjusters that are  deficient in the area of intended application.
Flyash, digested sewage sludge, and limestone screenings are
all waste products  of other processes and are, therefore, usually
inexpensive.   The major expense related  to utilization of any
of these wastes  is  the cost of transporting and applying the
material to  the  mine area.   Application may be quite costly
and must be  uniform to effect  complete and even revegetation.

When such large  amounts of  certain  chemical nutrients are
utilized, it  may also be necessary  to institute controls to
prevent chemical pollution  of  adjacent waterways.  Nutrient
controls may  consist of preselection  of vegetation to absorb
certain chemicals,  or of construction of berms and retention
basins in which  runoff can  be  collected and sampled, after
which it can  be  discharged  or  pumped  back to  the spoil.
The specific  soil supplements  and application rates employed
are selected  to  provide the  best  possible conditions for the
vegetative species  that are  to be planted.

Careful consideration should be given to species selection
in surface-mine  reclamation.   Species are selected according
to some land-use  plan,  based upon the degree of pollution con-
trol to be achieved  and the  site environment.   A dense ground
                              VII-14


                              DRAFT

-------
                             DRAFT
cover of grasses and legumes is generally planted,  in addition
to tree seedlings, to rapidly check erosion and siltation.
Trees are frequently planted in areas of poor slope stability
to help control landsliding.  Intended future use of the land
is an important consideration with respect to species selection.
Reclaimed surface-mined lands are occasionally returned to
high-use categories, such as agriculture, If the land has
potential for growing crops.  However, when toxic spoils are
encountered, agricultural potential is greatly reduced, and
only a few species will grow.

Environmental conditions—particularly, climate—are important
in species selection.  Usually, species are planted that are
native to an area—particularly, species that have been success-
fully established on nearby mine areas with similar climate
and spoil conditions.

Revegetation of arid and semi-arid areas involves special
consideration because of the extreme difficulty of establishing
vegetation.  Lack of rainfall and effects of surface distur-
bance create hostile growth conditions.  Because mining in
arid regions has only recently been initiated on a large scale,
there is no standard revegetation technology.  Experimentation
and demonstration projects exploring two general revegetation
techniques—moisture retention and irrigation—are currently
being conducted to solve this problem.

Moisture retention utilizes entrapment, concentration, and
preservation of water within a soil structure to support vege-
tation.  This may be obtained utilizing snow fences, mulches,
pits, and other methods.

Irrigation can be achieved by pumping or by gravity, through
either pipes or ditches.  This technique can be extremely
expensive, and acquisition of water rights may present a major
problem.  Use of these arid-climate revegetation techniques in
conjunction with careful overburden segregation and regrading
should permit return of arid mined areas to their natural states.

Exploration, Development, and Pilot-Scale Operations

Exploration activities commonly employ drilling, blasting,
excavation, tunneling, and other techniques to discover,
locate, or define the extent of an ore body.  These activities
                             VII-15

                             DRAFT

-------
                              DRAFT
vary  from small-scale (such as a single drill hole)  to large-
scale (such as  excavation of an open pit or outcrop  face).
Such  activities frequently contribute to the pollutant loading
in wastewater emanating from the site.   Since available  facili-
ties  (such as power sources) and ready accessibility of  special
equipment and supplies often are limited,  sophisticated  treat-
ment  is  often not  possible.   In cases where exploration  activity
is being carried out, the scale of such operations is such that
primary  water-quality problems Involve the presence  of increased
suspended-solid loads and potentially severe pH changes.  Ponds
should be provided for settling and retention of wastewater,
drilling fluids, or runoff from the site.  Simple, accurate
field tests for pH can be made,  with subsequent pH adjustment
by addition of  lime (or other neutralizing agents).

Protection of receiving waters will thus be accomplished, with
the possible additional benefits of removal of metals from
solution—either in connection with solids removal or by precipi-
tation from solution.

Development operations frequently are large-scale, compared to
exploration activities,  because they are intended to extend
already  known or currently exploited resources.  Because these
operations are  associated with facilities  and equipment  already
in existence, it is necessary to plan development activities to
minimize pollution potential,  and to use existing mine or mill
treatment and control methods and facilities.  These operations
should,  therefore,  be subject to limitations equivalent  to
existing operations with respect to effluent treatment and
control.

Pilot-scale operations often involve small to relatively large
mining and beneficiation facilities even though they may not
be currently operating at full capacity or are in the process
of development  to  full-scale.   Planning of such operations
should be undertaken with treatment and control of wastewater
in mind  to ensure  that effluent  limitation guidelines and
standards of performance for the category or subcategory will
be met.   Although  total  loadings from such operations and
facilites are not  at the levels  expected from normal operating
conditions,  the  compositions of  wastes  and the concentrations
of wastewater parameters are likely to  be similar.  Therefore,
Implementation of  recommended  treatment  and control technologies
must be  accomplished.
                             VII-16

                             DRAFT

-------
                             DRAFT
Mine and Mill Closure

Mine Closure (Underground).   Unless well-planned and well-
designed abatement techniques are implemented, an underground
mine can be a permanent source of water pollution.

Responsibility for the prevention of any adverse environmental
impacts from the temporary or permanent closure of a deep mine
should rest solely and permanently with the mine operator.
This constitutes a substantial burden; therefore, it behooves
the operator to make use of the best technology available for
dealing with pollution problems associated with mine closure.
The two techniques most frequently utilized in deep-mine pollu-
tion abatement are treatment and mine sealing.  Treatment tech-
nology is well defined and is generally capable of producing
acceptable mine effluent quality.  If the mine operator chooses
this course, he is faced with the prospect of costly permanent
treatment of each mine discharge.

Mine sealing is an attractive alternative to the prospects of
perpetual treatment.  Mine sealing requires the mine operator
to consider barrier and ceiling-support design from the perspec-
tives of strength, mine safety, their ability to withstand
high water pressure, and their utility for retarding ground-
water seepage.  In the case of new mines, these considerations
should be included in the mine design to cover the eventual
mine closure.  In the case of existing mines, these considera-
tions should be evaluated for existing mine barriers and ceiling
supports, and the future mine plan should be adjusted to include
these considerations if mine sealing is to be employed at mine
closure.

Sealing eliminates the mine discharge and inundates the mine
workings, thereby reducing or terminating the production of
pollutants.  However, the possibility of the failure of mine
seals or outcrop barriers increases with time as the sealed
mine workings gradually became inundated by ground water and
the hydraulic head increases.  Depending upon the rate of
ground-water influx and the size of the mined area, complete
inundation of a sealed mine may require several decades.
Consequently, the maximum anticipated hydraulic head on the
mine seals may not be realized for that length of time.  In
                             VII-17

                             DRAFT

-------
                             DRAFT
addition, seepage through, or failure of, the barrier or mine
seal could occur at any time.  Therefore,  the mine operator
should be required to permanently maintain the seals, or to
provide treatment in the event of seepage or failure.

Mine Closure (Surface).   The objectives of proper reclamation
management of closed surface mines and associated workings are
tc (1) restore the affected lands to a condition at least fully
capable of supporting the uses which they were capable of
supporting prior to any mining, and (2) achieve a stability
which does not pose any threat to public health, safety, or
water pollution.  With proper planning and management during
mining activities, it is often possible to minimize the amount
of land disturbed or excavated at any one time.  In preparation
for the day the operation may cease, a reclamation schedule
for restoration of existing affected areas, as well as those
which will be affected, should be specified.  The use of a planned
methodology such as this will return the workings to their
premined condition at a faster rate, as well as possibly reduce
the ultimate costs to the operator.

To accomplish the objectives of the desired reclamation goals,
it Is mandatory that the surface-mine operator regrade and
revegetate the disturbed area during, or upon completion of,
mining.  The final regraded surface configuration is dependent
upon the ultimate land use of the specific site, and control
practices described in this report can be incorporated into
the regradlng plan to minimize erosion and sedimentation.
The operator should establish a diverse and permanent vegeta-
tive cover and a plant succession at least equal in extent of
cover to the natural vegetation of the area.  To assure compli-
ance with these requirements and permanence of vegetative cover,
the operator should be held responsible for successful revege-
tatlon and effluent water quality for a period of five full
years after the last year of augmented seeding.  In areas of
the country where the annual average precipitation is 64 cm
(26 in.) or less, the operator's assumption of responsibility
and liability should extend for a period of ten full years
after the last year of augmented seeding, fertilization, irri-
gation, or effluent treatment.

Mill Closure.   As with closed mines, a beneficiation faci-
lity's potential contributions to water pollution do not cease
upon shutdown of the facility.  Tailing ponds, waste or refuse
                             VII-18

                             DRAFT

-------
                             DRAFT
piles, haulage areas, workings, dumps,  storage areas, and
processing and shipping areas often present serious problems
with respect to contributions to water  pollution.   Among the
most important are tailing ponds, waste piles, and dump areas.
Failure of tailing ponds can have catastrophic consequences,
with respect to both immediate safety and water quality.

To protect against catastrophic occurrences, tailing ponds
should be designed to accommodate, without overflow, an abnormal
storm which is observed every 25 years.  Since no  wastewater
is contributed from the processing of ores (the facility being
closed), the ponds will gradually become dewatered by evapora-
tion or by percolation into the subsurface.  The structural
integrity of .the tailing-pond walls should be periodically
examined and, if necessary, repairs made.  Seeding and vegetation
can assist in stabilizing the walls, prevent erosion and sedi-
mentation, lessen the probability of structural failure, and
improve the aesthetics of the area.

Refuse, waste, and tailing piles should be recontoured and
revegetated to return the topography as near as possible to
the condition it was in before the activity.  Techniques
employed in surface-mine regrading and  revegetatlon should be
utilized.  Where mills are located adjacent to mine workings,
the mines can be refilled with tailings.  Care should be taken
to minimize disruption of local drainage and to ensure that
erosion and sedimentation will not result.  Maintenance of
such refuse or waste piles and tailing-disposal areas should
be performed for at least five years after the last year of
regrading and augmented seeding.  In areas of the  country
where the annual average precipitation is 64 cm (26 in.)
or less, the operator's assumption of responsibility should
extend for a period of ten full years after the last year of
augmented seeding, fertilization, irrigation, or effluent
treatment.

TREATMENT TECHNOLOGY

Each of the techniques currently employed in the ore mining
and dressing industry, as well as advanced waste treatment
technology which might be employed in present or future opera-
tions, is discussed in this section.

The treatment technologies currently practiced in  the ore
mining and dressing industry encompass a wide variety of tech-
niques ranging from the very simple to  the highly  sophisticated.
                            VII-19


                              DRAFT

-------
                                DRAFT
While  a  limited number of basic treatment  practices are standard
 (settling  or tailing ponds,  pH control,  etc.)  and  employed
at almost  all operations, individual operations  have approached
specific pollution problems  in many different  ways.

Impoundment  Systems

This group of systems utilizes treatment technology which is
primarily  designed to deal with suspended  solids,  but which is
frequently used with such other techniques as  pH control, to
accomplish removal of dissolved constituents as  well.

Tailing  Ponds.    This type of  treatment  is the most common
treatment  technique used in  the ore mining and dressing industry
today.   The  design of a tailing pond is  primarily  for suspended-
solid  removal and retention.   Such  a pond  must be  large enough
to provide sufficient retention time and quiescent conditions
conducive  to settling.   If properly designed,  and  if retention
time and surface area are sufficient, a  tailing  pond may also
effect to  some degree the stabilization  of oxidizable consti-
tuents as  well as the balancing of  influent quality and quantity
fluctuations and the storage of storm water.

Tailing  ponds are often situated to capitalize upon natural
terrain  factors in order to minimize the requirements for dam
construction.   The containment dam  is often constructed of
available  earth and rock materials,  as well as tailings.  In
other cases,  concrete basins may be constructed.   Because of
natural  terrain conditions, they may be  constructed using one,
two, three,  or  even four walls.  The containment dam must be
raised periodically to  accommodate  the rising  level of contained
tailings and water.   In most cases,  the  basin  provides perpetual
storage  for  any materials settled out of the water treated.
Typically, a concrete basin is periodically dredged and the
solids stored in a waste-disposal area.  Retention time in
ponds has  been  reported to vary from as  little as  four hours
to as much as several months at average  flow conditions (for
discharging  systems).

Water leaves  a  tailing  pond by decantatlon, evaporation,
seepage  through- the dam or to  underlying materials, or by discharge.
Decanted water  may be recycled for  use in  the  mill, discharged,
or treated further.   In some operations, in arid or semi-arid
areas, evaporation from the tailing-pond surface may equal the
rate of  input,  allowing zero-discharge operation of the pond
without  recycle of water.
                               VII-20

                               DRAFT

-------
                               DRAFT
Seepage losses from tailing ponds may flow into permeable
underlying strata and enter ground water, or may flow through
the containment dam and result in surface flows of water.
Where dam construction has made use of tailings, some seepage
will almost always be observed.  Resulting seepage waters are
often collected in ditches and pumped back into the tailing
pond.  Seepage may also be limited by the use of pond liners
of various materials (clay, asphalt, plastic, etc.).

Low-cost, relatively simple construction and the ability to
perform multiple functions simultaneously have led to the wide
acceptance of tailing ponds as a prime treatment and tailing-
disposal method utilized by the ore mining and dressing industry.
There are a number of problems associated with the utilization
of tailing ponds as treatment facilities, however.  Improper
design of inlet and discharge locations, Insufficient size
and number, and insufficient retention time are the most common
problems.  Algal growths in tailing ponds are quite common
during warm months, a factor which may influence such effluent
water-quality parameters as TOG, COD, and BOD.  A minimum
retention time of 30 days and the added capability of retaining
runoff associated with a storm likely to occur once in 20 years
are recommended by one source (Reference 29).

The relative advantages and disadvantages of a tailing pond as
a treatment system are listed below.
Advantages
Disadvantages
Performs large number of
treatment processes—parti-
cularly, suspended-solid
removal.

Can achieve high treatment
efficiency and often pro-
duce acceptable effluent
quality.
Often, only practical means
of long-term solids disposal
     Lacks responsive means of control;
     difficult to optimize large number
     of processes performed.
     Covers large surface area—may
     contribute high net precipitation
     to overall water balance; land
     availability and topography Influ-
     ence location.

     Creates potentially severe rehabili-
     tation problem if tailings contain
     sulfide minerals.
                             VII-21

                               DRAFT

-------
                             DRAFT
                               Disadvantages

                               Often difficult to isolate from
                               contributing drainage areas—
                               storm water influences retention.

                               Subject to climatic variations—
                               particularly, thermal skimming and
                               seasonal variation in bio-oxidation
                               efficiency.

                               Often difficult to ensure good
                               flow distribution.
                               Requires careful control of
                               seepage through dams.

                               Installation expensive in some
                               situations, due to high cost of
                               retaining structures.
Advantages

Large retention has a balan-
cing effect on effluent
quality.

Large surface area aids
oxidation and evaporation.
Can often be constructed
using mining equipment
and materials.

Little operating expertise
normally required.

Commonly used treatment
method, familiar to
industry.

Clear supernatant water may
serve as a reservoir  for
reuse.
Tailing ponds in the ore mining and dressing industry range
from pits to large, engineered structures of 1000 acres with
massive retaining dams.  For large tailing dams, wall heights
of 200 feet or more have been reached by building up the dams
over a period of time.

Routinely reached levels of suspended-solid concentrations in
treated effluent range from 10 to 30 mg/1 at mines and mills
visited or surveyed as part of this study.  In tailing ponds
with decant structures for recycle of water, levels in excess
of 50 mg/1 of suspended solids were rarely observed.

Settling Ponds.   Settling ponds differ from tailing ponds
primarily in size and in the concentrations of influent solids
treated.  In general, relatively low initial solid loads are
removed, necessitating only occasional dredging to maintain
adequate settling volume behind the dam.  Suspended-solid
removal to very low levels is often possible when initial
concentrations of suspended solids are low.  Settling ponds
                             VII-22

                             DRAFT

-------
                              DRAFT
find their greatest usefulness in association with mines having
low wastewater solids loads.   Effluent levels of A.3 mg/1,
6.2 mg/1, and 17 mg/1, for example, were observed from three
different settling ponds.

Such ponds may serve a variety of purposes in addition to
removal of suspended solids,  including COD reduction and cooling.
As basins for a variety of chemical treatments,  they can
provide sufficient retention time for completion of reactions,
for pH control, for chemical precipitation, and  for the removal
of solids produced.

Secondary Settling Ponds.    Settling ponds or tailing ponds are
frequently used in a multiple arrangement.  The  purpose of
this scheme is to further reduce suspended-solid loading in
the sequential ponds and to allow the subsequent use of precipi-
tation or pH control before discharge or recycle.  The ponds
enable further reduction in suspended solids and in dissolved
parameters.  An excellent example is the use of  secondary
settling ponds (sometimes called polishing ponds) in the copre-
cipitation of radium with barium.  Removal of radium could be car-
ried out in the tailing pond, but the pollutant  radium would be
distributed over a large area, presenting a hazard to workers
and a potential future hazard from the use of abandoned tailings
in construction (a practical use in some areas).  By using a
small secondary settling pond to coprecipitate radium, the
pollutant is confined to an area to which access can be res-
tricted, even after the operation is closed.  Similar considera-
tions would suggest the use of such ponds for the precipitation
of heavy metals by lime or sulflde treatments.

Clarifiers and Thickeners

A method of removing large amounts of suspended  solids from
wastewater is the use of clarifiers or thickeners, which are
essentially large tanks with directing and segregating systems.
The design of these devices provides for concentration and
removal of suspended and settleable solids in one effluent
stream and a clarified liquid in the other.  Clarified waters
may be produced which have extremely low solids  content through
proper design and application.

Clarifiers are not generally capable of handling tailing-solid
levels above about 50 percent, due to the necessity for rake
                              VII-23

                               DRAFT

-------
                               DRAFT
operation and hydraulic transport of suspended solids from
the device.  The concentration from a mine-water clarifier
at one site, for example, is 3 mg/1 suspended solids.

Clarifiers may range in design from simple units to more complex
systems involving sludge blanket pulsing or sludge recycle to
improve settling and increase the density of the sludge.
Settled solids from clarifiers are removed periodically or
continuously for either disposal or recovery of contained
values.  Thickeners are used when the main purpose is to produce
a clarified overflow with a concentrated tailing effluent in
the underflow.

Thickeners have a number of distinct advantages over settling
or tailing ponds:

      (1)  Less land space Is required.  Area-for-area, these
           devices are much more efficient in settling capacity
           than ponds.

      (2)  Influences of rainfall are reduced compared to ponds.
           If desired, the clarifiers and thickeners can be
           covered.

      (3)  Since the external construction of clarifiers and
           thickeners consists of concrete or steel (in the
           form of tanks), ground-seepage and rain-water runoff
           influences do not exist.

      (4)  Thickeners can generally be placed adjacent to a
           mill, making reclaim water available nearby with
           minimal pumping requirements.

The use of clarifiers and thickeners, together with tailing
or settling ponds, may improve treatment efficiency; reduce
the area needed for tailing ponds; and facilitate the reuse
or recycle of water in the milling operation.  The use of
flocculants to enhance the performance of thickeners and
clarifiers is common practice.

Clarifiers and thickeners also suffer some distinct disadvan-
tages compared to ponds:

      (1)  They have mechanical parts and, thus, require
           maintenance.
                              VII-24

                               DRAFT

-------
                              DRAFT
      (2)  They have limited storage capacity for either
           clarified water or settled solids.

      (3)  The internal sweeps and agitators in thickeners
           and clarifiers require more power and energy for
           operation than ponds.

Flocculation

This treatment process consists basically of adding reagents
to the treated waste stream to promote settling of suspended
solids.  The solids may be deposited in tailing ponds (where
high suspended solids are involved) or in clarifier tanks (in
cases of lower solids loads).

Flocculating agents Increase the efficiency of settling facili-
ties and are of several general types:  ferric compounds, lime,
aluminum sulfate, and cationic or anionic polyelectrolytes.
Causticized wheat and corn starch have also been used.   The ionic
types, such as alum, ferrous sulfate, lime, and ferric  chloride,
function by destroying the repelling double-layer ionic charges
around the suspended particles and thereby allowing the particles
to attract each other and agglomerate.  Polymeric types function
by forming physical bridges from one particle to another and
thereby agglomerating the particles.  Recyclable magnesium
carbonate has also been proposed as a flocculant in domestic
water treatment.

Flocculating agents are added to the water to be treated under
controlled conditions of concentration, pH, mixing time, and
temperature.  They act to upset the stability of the colloidal
suspension by charge neutralization and flocculation of sus-
pended solids, thus Increasing the effective diameter of these
solids and increasing their subsequent settling rate.

Flocculating agents are most commonly used after the larger,
more readily settled particles (and loads) have been removed
by a settling pond, hydrocyclone, or other treatment.  Agglomer-
ation, or flocculation, can then be achieved with less  reagent,
and with less settling load on the polishing pond or clarifier.

Flocculation agents can be used with minor modifications and
additions to existing treatment systems, but the costs  for
                              VII-25


                               DRAFT

-------
                              DRAFT
the flocculating chemicals are often significant.  Ionic types
are used in concentrations of 10 to 100 mg/1 in the wastewater,
while the highest-priced polymeric types are effective in
concentrations of 2 to 20 mg/1.

The effectiveness and performance of individual flocculating
systems may vary over a substantial range with respect to
suspended-solid removal, accessory removal of soluble com-
ponents by adsorptive phenomena, and operating characteristics
and costs.  Specific system performance must be analyzed and
optimized with respect to mixing time, flocculant addition
level, settling (detection) time, thermal and wind-induced
mixing, and other factors.

Centrifugation

Centrifugation, which may be considered as a form of forced
or assisted settling, may be feasible  in specific control
applications.  With the volume of gross wastewater flows
at most mine/mill complexes, it is probable that centrifuga-
tion may be more applicable to component in-process waste
streams.  The presence of abrasive components or significant
amounts of solid material smaller than approximately 5
micrometers in diameter in the treated water would tend to
disqualify Centrifugation as a solid-removal option.

Hydrocyclones

While hydrocyclones are widely used  in the separation, classifi-
cation, and recovery operations involved in mineral processing,
they are used only Infrequently for  wastewater treatment.
Even the smallest-diameter units available (stream-velocity
and centrifugal-separation forces both increase  as the diameter
decreases) are  ineffective when particle size is less than
25 to 50 micrometers.  Larger particle sizes are relatively
easy to settle  by means of small ponds, thickeners or clarifiers,
or other gravity-principle settling  devices.  It is the smaller
suspended particles  that are  the most  difficult  to remove, and
it is these  that cannot be removed by  hydrocyclones but may
be handled by ponds  or other  settling  technology.  Also, hydro-
cyclones are of doubtful effectiveness when flocculating agents
are used to  increase  settling rates.   This method is generally
most effective  in  the  25-  to  200-micrometer size range for
particles.
                               VII-26

                               DRAFT

-------
                               DRAFT
Filtration

Filtration is accomplished by passing the wastewater stream
through solid-retaining screens or cloths or particulate
materials such as sand, gravel, coal, or diatomaceous earth
using gravity, pressure, or vacuum as the driving force.
Filtration is a versatile method in that it can be used to
remove a wide range of suspended particle sizes.

Filtration is not generally useful on mill waste streams, due
to the high volumes of material involved.  The large volumes
to be treated would require large filters.  The cost of these
units, and their relative complexity compared to settling ponds,
has restricted their use.

A variety of filtration techniques, including disc and drum
units, find process applications and may be applicable to some
waste streams—particularly, where segregated waste streams
require special treatment.

Likely applications of filtration include pretreatment of input streams
using reverse-osmosis and ion-exchange units (discussed later).

High values contained in suspended solids may, in some cases,
offset the capital and operating expenses of filtering systems.
The use of filtration as a normal unit process in treating
uranium-mill tailings for value recovery through countercurrent
washing is indicative of the possible use of filtration in
tailing treatment.  In this instance, the final washed tail
filter cake is reslurried for transport to the tailing pond.
In situations where biological treatment of component or
combined waste streams is required to reduce BOD, COD, or
bacterial loads, trickling filters may be required, but their
application as primary treatment for the bulk mine or mill
effluent is considered unlikely.

The specific applicability and size specifications for filter
modules must be evaluated on a case-by-case basis, taking into
account the process stream characteristics, solids filter-
ability, desired dryness of filter cake, and other parameters.

Ultimate clarification of filtered water will be a function of
particle size, filter-media porosity, filtration rate, and
other variables.  In general, for the majority of mine or
mill waste waters subjected to this treatment, post-treatment
                               VII-27


                               DRAFT

-------
                             DRAFT
suspended-solid levels of less Chan 20 percent of influent
loadings are anticipated.  Thus, If used after primary floccu-
lation and settling, suspended solids levels of 20 mg/1 should
be obtainable.

Neutralization

Adjustment of pH is the simplest and most common treatment
practiced in the mining and milling Industry today.  The addi-
tion of either acidic or basic constituents to a wastewater
stream to achieve neutralization generally Influences the
behavior of both suspended and dissolved components.  In most
instances of interest in mining and milling activities, waste-
waters are treated by base addition to achieve pH conditions
in the range of 6 to 9.

Acid waste streams (considerably more common than highly basic
effluents) may be neutralized by addition of a variety of
basic reagents, including lime (calcium oxide), limestone,
dolomite (CaMg(C03)2J, magnesite (MgC03), sodium hydroxide,
soda ash (sodium carbonate), ammonium hydroxide, and others
to raise the pH of treated waste streams to the desired level.
Lime is most often used because it is inexpensive and easy to
apply.  Soda ash and caustic soda are commonly used to supply
alkalinity in leaching and hydrometallurgical processes, where
the formation of calcium precipitates would be objectionable,
but the cost advantages of using lime generally preclude the
use of soda ash and caustic soda in large-scale waste treatment.

Ammonia neutralization is most frequently a processing techni-
que, where ammonia affords a strong advantage in being volatile
in the final product, allowing the recovery of nearly pure
oxides.  In waste treatment, its volatility is a disadvantage
because of the COD it presents, its toxicity, and the produc-
tion of undesirable nitrites and nitrates as oxidation products.
Its use is not widespread, although ammonia neutralization of
a wastewater stream is practiced at one site in the ferroalloy-
ore mining and milling category.

Excessively basic waste streams are not common but may be
neutralized by addition of an acid—most commonly, sulfuric.
Since many heavy metals form insoluble hydroxides in highly
basic solutions, sedimentation prior to neutralization may
prevent the resolubillzation of these materials and may
simplify subsequent waste-treatment requirements.  Carbon
dioxide has also been used to adjust the pH of effluent
waters to acceptable levels prior to discharge (recarbona-
tion).
                             VII-28

                             DRAFT

-------
                               DRAFT
Essentially any wastewater stream may be treated to a final pH
within the range of 6 to 9.  Generally, the stream will be
sufficiently uniform to allow adequate pH control based only
on the volume of flow and predetermined dosage rates, with
periodic adjustments based on effluent pH.  Automated systems
which monitor and continously adjust the concentration
of reagents added to the wastewater are also currently available.

As discussed previously, pH control is often used to control
solubility (also discussed under Chemical Precipitation Pro-
cesses) .  Examples of pH control being used for precipitating
undesired pollutants are:

                (1)  Fe(+3) + 30H(-)-*Fe(OH)3_

                (2)  Mn(+2) + 20H(-)->Mn(OH)2_+ 2H(+) + 4e(-)

                (3)  Zn(+2) + OH(-)-»Zn(OH)2_

                (4)  Pb(+2) + 20H--»Pb(OH)2_

                (5)  Cu + 20H(-) -»Cu(OH)2_

Reaction (1) is used for removal of iron contaminants.  Reaction
(2) is used for removal of manganese from manganese-containing
wastewater.  Reactions  (3), (4), and (5) are used on wastewater
containing copper, lead, and zinc salts.  The use of lime to
attain a pH of 7 will theoretically reduce heavy metals to these
levels (Reference 30):

                Metal          Concentration (mg/1 at pH 7)

                Cu(+2)         0.2 to 0.3

                Zn(+2)         1.0 to 2.5

                Cd(+2)         1.0

                Ni(+2)         1.0

                Cr(+2)         0.4

The careful control of  pH, therefore, has other ancillary bene-
fits, as illustrated above.  The use of pH and solubility rela-
tionships to Improve removal of wastewater contaminants is
further developed below.
                               VII-29

                               DRAFT

-------
                               DRAFT
Chemical Precipitation Processes

The removal of materials from solution by the addition of
chemicals which form  insoluble  (or sparingly soluble) compounds
with them is a common practice  in hydrometallurgical ore bene-
ficiation and in waste treatment in the ore mining and dressing
industry.  It is especially useful for the removal of heavy
metals from mine effluents and  process wastes.

To be successful, direct precipitation depends primarily upon
two factors:

     (1)  Achievement of a sufficient excess of the added
          ion to drive the precipitation reaction to completion.

     (2)  Removal of  the resulting solids from the waste stream.

If the first requirement is not met, a portion of the pollutant(s)
will be removed from  solution,  and desired effluent levels may
not be achieved.  Failure to remove the precipitates formed
prior to discharge  is likely to lead to redissolutlon, since
ionic equilibria in the receiving stream will not, in general,
be those created in treatment.  Effective sedimentation or
filtration is, thus,  a vital component of a precipitation
treatment system and  frequently limits the overall removal
efficiency.  Sedimentation may  be effected in the tailing basin
Itself, in secondary  or auxllliary settling ponds, or in
clarifiers.  Industry experience has shown the value of treat-
ment of wastes prior  to delivery to the tailing impoundment.
Benefits derived include:  improved settling of precipitates
due to interaction  with tailings; simplified disposal of sludges;
and, generally, suppressed solubility of materials in tailing
solids.

The use of precipitation for wastewater treatment varies from
lime treatments  (to precipitate sulfates, fluorides, hydroxides,
and carbonates) to  sodium sulfide precipitation of copper, lead,
and other toxic heavy metals.   The following equations are
examples of precipitation reactions used  for wastewater treat-
ment:

     (1)  Fe(+3) +  Ca(OH)2_ 	» Ca(+2) + Fe(OH)3_

     (2)  Mn(+2) +  Ca(OH)2_ —> Ca(+2) + Mn(OH)2_

     (3)  Zn(+2) +  Na2C03_   —> Na(+) +  ZnC03_
                              .VII-30

                               DRAFT

-------
                               DRAFT
      (4)  S04/-2) + Ca(OH)£ 	>  CaSOA. + 20H(-)

      (5)  2F(-) + Ca(OH)2  	> CaF2_ + 20H(-)

One drawback of the precipitation reactions is that the varying
solubility of metal species and the possibility of widely diver-
gent formation and precipitation rates limit the ability of
this treatment to deal with all waste constituents.

Lime Precipitation.   The use of lime to cause chemical preci-
pitation has gained widespread use in the ore mining and dressing
Industry because of its ease of handling, because of its economy,
and because of its effectiveness in treatment of a great variety
of dissolved materials.  The use of other bases is, of course,
possible, as previously discussed.  However, the use of lime
as a treatment reagent Is probably the best-known and best-
studied method.

A typical lime neutralization/precipitation system is illustrated
in Figure VII-1,  Generally, water is pumped or discharged to
a holding or settling pond, where suspended-solid levels are
reduced.  Either in conjunction with the primary pond itself
or in a mixing basin or tank, a slurry of lime and water is
delivered for mixing with the wastewater stream.  Secondary
settling ponds are then used to collect the usually high volumes
of sludges which may be recovered.  These impoundments may
be dredged periodically to remove sludges, or the sides of the
basin may be built up.  Discharge of the water then usually
takes place.

The treatment conditions, dosages, and final pH must be optimized
for any given waste stream, but, In general, attainment of a
pH of at least 9 is necessary to ensure removal of heavy metals.
To attain desired levels of control for many heavy metals, it
is necessary to attain a pH of 10 to 12 in many instances.

The levels of concentration attainable In an actual operating
system may vary from the limits predicted on the basis of purely
theoretical considerations, but extremely low levels of metals
discharged have been reached by the use of this treatment method.
Figure VII-2 Illustrates the theoretical solubilities of several
metal ions as a function of pH.  The minimum pH value for complete
precipitation of metal ions as hydroxides is shown in Figure VII-3.
                              VI1-31

                               DRAFT

-------
                             DRAFT
    Figure VIM. LIME NEUTRALIZATION AND PRECIPITATION PROCESS FOR
             TREATMENT OF MINE WATER PRIOR TO DISCHARGE
 FROM MINE OR MILL
                                                SLUDGE
                                               REQUIRING
                                                DISPOSAL
   TO
DISCHARGE
                       SOURCE:  Reference 31
                             VII-32


                             DRAFT

-------
                                DRAFT
Figure VII-2. THEORETICAL SOLUBILITIES OF METAL IONS AS A FUNCTION OF pH
    0.0001
         6        7


    SOURCE: Reference 32
                                VII-33


                                DRAFT

-------
                                    DRAFT
Figure VI1-3. MINIMUM pH VALUE FOR COMPLETE PRECIPITATION OF METAL IONS AS
           HYDROXIDES
  PH
in n
9.0
o n
7.0
6.0
5.0
4.0
3.0
2.0
1.0
0.0
•
i
I
7.2
^^m
(
5.2
4.2 '









1.3
••

















12
IM





















i.4









1.3

«









1.5














>.7



	 1
i










D.i
MM
1

	







           Sn+2    Fe+3   AI+3   Rb+2   Cu+2   Zn+2   Nj+2   Fe+2  Cd+2  Mn+2
                           LIME
                      NEUTRALIZATION
                                   LIME PRECIPITATION
            SOURCE:  Reference 31
                                    VII-34


                                    DRAFT

-------
                               DRAFT
An example of the performance of lime precipitation at elevated
pH is given for Fe, Fb, Zn, Cd, Hg, and F in Figure VII-4.
These data are taken from a representative mill, where removal
efficiency is plotted against pH.  The curves are not always
complete for lack of data; it is not advisable to extrapolate
them without further measurements, because chemical changes
may occur that reverse an apparent consistent trend.

Purely theoretical considerations of metal-hydroxide solubility
relationships suggest that the metal levels tabulated below
are attainable (Reference 29).

                Final Concentration
                (microgram per liter)    p_H

                   1 to 8                9.5

                   10 to 60              10

      Pb             1                    8

      Fe(total)      1                    8 (if totally Ferric)

Many factors, such as the effects of widely differing solu-
bility products, mixed-metal hydroxide complexing, and metal
chelation, render these levels of only limited value when
assessing attainable concentrations in a treatment system.

Among the metals effectively removed at basic pH are:  As, Cd,
Cu, Cr(-f3), Fe, Mn, Hi, Pb, and Zn.  Based upon published
sources, industry data, and analysis of samples, it appears
that the concentrations given in the tabulation below may be
routinely and reliably attained by hydroxide precipitation
in the ferroalloy-ore mining and milling industry.

Metal      Concentration       Metal     Concentration
               (mg/1)                        (mg/1)
 As             0.50            Mn            1.0
 Cd             0.50            Ni            0.05
 Cu             0.05            Pb            0.10
 Cr(+3)         0.05            Zn            0.15
 Fe             1.0
                              VII-35

                               DRAFT

-------
                             DRAFT
  Figure VI1-4. HEAVY-METAL PRECIPITATION vs pH FOR TAILING-POND
            EFFLUENT pH ADJUSTMENTS BY LIME ADDITION
    50
    60
#
H

o
iu

S   70

    an
    w
    90
   100
                                              11
13
                             pH
                       SOURCE: Reference 33
                            VII-36



                            DRAFT

-------
                               DRAFT
Some metallic pollutants of interest in the uranium-ore mining
and milling industry, together with results produced by lime
precipitation in conjunction with a rise in pH from 6.7 to
12.7, are shown below:

      Metal               Concentration (mg/1)
                          pH°6.7            pH=12.7

      Cd                  1.3            less than 0.02

      Fe                  6.0            less than 0.1

      Ni                  0.13           less than 0.05

      Cu                  5.3                      0.05

      Zn                 31.25                     0.11

      Mn                 26.5                      0.04

The use of lime or limestone will also provide some removal of
fluoride ion, by precipitation of calcium fluoride.  The
theoretical limit of fluoride-ion solubility in a solution
containing 10 mg/1 of calcium, for example, is less than 4 mg/1.

Other examples of the efficiency of lime precipitation as a
treatment method are discussed by ore category later in this
section.  An Important point is Illustrated in the data pre-
viously presented here, however.  All metals do not remain In
solution at elevated pH.  Examples of that phenomenon are
the variations in solubilities of lead and zinc, which are
precipitated at approximately pH 9.  Above pH 9, these metals
rapidly resolubllize.  Lime precipitation has not been demon-
strated as being effective for removal of tellurium, molybdenum,
arsenic, and mercury at high pH, because these metals become
more soluble with increasing alkalinity.

Sulfide Precipitation.   The use of sulfide ion as a precipi-
tant for removal of heavy metals accomplishes more complete
removal than the use of hydroxide for precipitation.  Sulfide
precipitation is currently being used In wastewater treatment
to reduce mercury levels to extremely low levels (Reference
34).  Highly effective removal of Cd, Cu,  Co, Fe, Hg, Mn,
Ni, Pb,  Zn, and other metals from mine and mill wastes can be
                             VII-37

                              DRAFT

-------
                               DRAFT
accomplished by treatment with either sodium sulfide or hydrogen
sulfide.  The use of this method depends somewhat on the avail-
ability of methods for effectively removing precipitated solids
from the waste stream, and on removal of the solids to an
environment where reoxidation is unlikely.

Several steps enter into the process of sulfide precipitation:

      (1)  Preparation of sodium sulfide.  Although this product
           is often in oversupply from byproduct sources, it
           can also be made by the reduction of sodium sulfate,
           a waste product of acid-leach milling.  The process
           Involves an energy loss in the partial oxidation of
           carbon (such as that contained in coal).

           Na2S04. + 4C - >  Na2S + 4CO )gas)

      (2)  Precipitation of the pollutant metal (M) in the
           waste stream by an excess of sodium sulfide:

           Na2S + MS04_  — >  MS (precipitate) + Na2_S04_

      (3)  Physical separation of the metal sulfide in thickeners
           or clarlfiers, with reducing conditions maintained
           by excess sulfide ion.

      (4)  Oxidation of excess sulfide by aeration:

           Na2S + 202  — >
           This process  usually  involves  iron as an intermediary
           and is  seen to  regenerate unused sodium sulfate.

On the whole, sulfide precipitation removes both heavy metals
and some  sulfur from waste streams but  requires some energy
expenditure.

In practice,  sulfide precipitation can  be applied only when
the pH is sufficiently high (greater than about 8) to assure
generation of sulfide ion  rather than bisulfide or hydrogen
sulfide gas.  It is  then possible to add  just enough sulfide,
in the form of sodium sulfide, to precipitate the heavy metals
present as cations;  alternatively, the  process can be continued
until dissolved oxygen in  the effluent  is reduced to sulfate
and aerobic conditions are obtained.  Under these conditions,
                               VII-38

                                DRAFT

-------
                             DRAFT
some reduction and precipitation of molybdates, uranates,
chromates, and vanadates may occur, but ion exchange seems more
appropriate for the removal of these anions.

Due to the toxiclty of sulfide ion, and of hydrogen sulfide
gas, the use of sulfide precipitation may require both pre-
and post-treatment and close control of reagent additions.
Pretreatment involves raising the pH of the waste stream to
minimize evolution of H2S, which would pose a safety hazard
to personnel.  If desirable, this may be accomplished at
essentially the same point as the sulfide treatment, or by
addition of a solution containing both sodium sulfide and a
strong base (such as caustic soda).  The sulfides of many
heavy metals, such as copper and mercury, are sufficiently
Insoluble to allow essentially complete removal with extremely
low residual sulfide levels.  Treatment for these metals with
close control on sulfide concentrations could be accomplished
without the need for additional treatment.  Adequate aeration
should be provided to yield an effluent saturated with oxygen.

Coprecipitation.   In coprecipitation, materials which cannot
be removed from solution effectively by direct precipitation
are removed by incorporating them into particles of another
precipitate, which is separated by settling, filtration, or
another technique such as flotation.  Current practice is
exemplified by the use of barium chloride addition for radium
control in the uranium industry.

Radium sulfate, one of the least soluble substances, is soluble
to 20 micrograms per liter, while allowable concentrations in
drinking water are about 6 million times less.  The process of
coprecipitation for radium separation was perfected by M.S.
Curie and has been used extensively in radiochemistry.  The
carrier for radium is barium, usually added as barium chloride
(BaC12) in a concentration of about 10 mg/1 and In the presence
of more sulfate ion than is necessary to precipitate barium
sulfate (BaSOA).  Almost all RaSOA^ that is present is copreci-
pitated, and removal to a level of about 1 picocurie (1 pc/1)
or 1 plcogram per liter, is current practice.  The results of
tests on the addition of BaC12^ BaS04^, and BaC03_ to neutral
and acidic effluents are shown in Table VII-1.

The importance of coprecipitation in the ferroalloy industry
has been demonstrated by extensive experiments (References
35 and 36).   In that work, molybdenum, which appears
                             VII-39

                             DRAFT

-------
                    DRAFT
TABLE VIM. RESULTS OF COPRECIPITATION REMOVAL OF
          RADIUM FROM WASTEWATER
EFFLUENT pH
Neutral
Acidic
REAGENT
BaS04
BaCOj
BaCI2
BaC03
BaC12
REAGENT
ADDITION
(rngfi.}
300
1000
100
200
30
60
100
200
100
200
300
100
PRE- AND POST-PRECIPITATION
RADIUM CONCENTRATIONS
(pc/£)
BEFORE
100
300
470
490
800
440
400
430
160
150
150
150
AFTER
30
70
30
40
20
6
2
2
18
20
30
6 to 15
X RADIUM
REMOVED
70
77
94
92
97
99
99
99
88
87
80
90 to 97
                    VII-40


                    DRAFT

-------
                             DRAFT
in effluents from many mines and mills as the molybdate (Mo04_-)
anion (which is not removed effectively by hydroxide or sulfide
precipitation), is removed by incorporation into ferric hydrox-
ide precipitates formed at acid pH (4.5 optimum) by the addition
of ferric sulfate or ferric chloride (at levels of about 100
mg/1).  Removal of resulting precipitates by filtration and
flotation has been reported to yield effluents containing 0.2
mg/1 for mill waters initially containing 4.9 mg/1 of molybdenum
(Reference 37).  In a pilot-plant study using ferric sulfate
and flotation recovery of precipitates, removal of more than
95 percent of influent molybdenum, to levels of 0.02 to 0.1
mg/1, has been obtained.

Since the process used for molybdenum removal is performed at
acid pH, it is necessary to acidify the (typically, alkaline)
mill waste stream after separation of solids in the tailing
pond effects the molybdenum removal.  A base is then added to
neutralize the effluent prior to discharge.  For large waste
stream flow, reagent costs may be an important consideration.
Although molybdenum values are concentrated to about 5 percent
in the precipitates removed, they do not appear to represent
a marketable product at this time.

Other Precipitation Systems.   Other types of precipitation
systems have been employed, such as those used for the preci-
pitation of sulfate (Reference 38), fluoride (as calcium
fluoride), or others (Reference 39).  Starch-xanthate com-
plexes have recently been reported to be effective in aiding
precipitation of a variety of metals, Including Cd, Cr, Cu,
Pb, Hg, Ni, Ag, and Zn (Reference 40).   Scavenging or
coprecipitation studies have been conducted on municpal waste-
waters (Reference 41).  In specialized cases, precipitation
may be induced by oxidation, which produces a less soluble
heavy-metal product.  The chlorine oxidation of Co(+2) to Co
(+3) at a pH of approximately 5 produces the insoluble Co^3_
(xH20).  Oxidation of Fe(+2) to Fe(+3) results in  the precipi-
tation of hydrous ferric oxide, even at relatively low pH.
Oxidation of As(+3) to As(+4) improves precipitation removal
(Reference 40).  The use of oxidation is further discussed
later in this section.
                             VII-41

                             DRAFT

-------
                               DRAFT
Reduction

Reduction techniques have particular applicability to the removal
of hexavalent chromium and copper from waste streams in the
ferroalloy-ore mining and milling industry.  Copper is often
recovered in current practice by reduction of the metal and
subsequent deposition on scrap iron in the waste stream (cementa-
tion) .  Since the  effluent levels resulting from cementation
are still high,  generally 10 mg/1 or more, it is necessary to
follow use of this process with another removal step, such as
hydroxide precipitation.

Reduction of chromates to trivalent chromium, with subsequent
precipitation of the chromium as the hydroxide, is a standard
waste-treatment  practice in a number of industries and may find
application in the ore mining and dressing industry, where
leaching practices give rise to wastewater contaminated with
chromates.  Commonly used reducing agents include sulfur
dioxide and ferrous salts of iron.  With sulfur dioxide and a
pH of 2.5, chrornate may be reduced rapidly and completely.
Removal of the Cr(OH)3^ precipitate formed in treatment of the
relatively dilute  wastes to be expected in mill effluents may
prove difficult, necessitating careful management of the
treatment system and the use of flocculants such as Fe(OH)3^
to aid in settling.  Effluent levels of 0.5 mg/1 of total chromium
and 0.05 mg/1 of hexavalent chromium may be reliably attained
by the treatment (Reference 42).

Sodium borohydrlde reduction has been applied to reducing
soluble mercury  levels in chlor-alkali and mercury processing
plants and to reducing lead levels in wastes arising in the
tetra-alkyllead  manufacturing process (U.S. Patents 3,736,253,
3,764,528, and 3,770,423).  Stannous  (tin) compounds have been
used  for the reductive deposition of palladium during electro-
plating processes. Electroreduction of metals is widely practiced
in electrowlnning  and electrorefining systems  for copper, nickel,
cobalt, and other  metals.

Treatment in the ore mining and dressing industry differs from
the above techniques, chiefly because of the lower concentra-
tions of soluble,  reducible species and because of the presence
of numerous other  reducible species in the wastewater.  Unless
preconditioning  of treated waters is employed, excessive reducing-
agent consumption  may occur.  Secondary recovery systems  (settling,
filters, etc.) may be necessary to permit removal of reduced
                              VII-42

                               DRAFT

-------
                              DRAFT
components.   The recovery of values from waste residues is a
potential option with this treatment method.   In some instances,
application  of this process option to internal streams prior
to discharge and/or combination with other waste streams may
offer substantial enhancement of value recovery from treatment
products.

Oxidation, Aeration, and Air Stripping

A number of the waste components resulting from mining and
milling may be removed or rendered less harmful by oxidation
or removal to the atmosphere.  Among these are cyanide, sulfide,
ammonia, and a variety of materials presenting high COD levels.
The simplest approach to effecting these processes is aeration
of the waste stream, which occurs naturally in pumping it and
in distributing it at the tailing pond.  More elaborate imple-
mentation achieves more complete and rapid results in air strippers,
and by controlled introduction of stronger oxidants, such as
chlorine or ozone.

Cyanide  (CN-) is removed by oxidation to cyanate (CNO-) and,
ultimately, to C02^ and N2^.  This is accomplished in  standard
practice by rapid chlorination at alkaline pH  (about 10.5)
using caustic soda.  The probable reaction with excess chlorine
has been expressed as:

2NaCN +  5C1^ + 12NaOH  	> N£ + 2Na2C03, + lONaCl +  6H20

A pH of  10  to 11 is  recommended for operating  conditions.
This process may be  performed  on either  a batch or continuous
process.  Approximately  2.72 kg (6  Ib) each of caustic  soda
and chlorine are normally  required  to oxidize  0.45 kg  (1  Ib)  of
cyanide.  If metal-cyanide  complexes are present, extended
chlorination  for several hours may  be necessary.

In treatment of mill effluent  in  the gold milling industry,
some cyanide  is  lost in  the process and  is present  in  the mill
tailings.   Some  of  the cyanide decomposes  in  the  tailing  pond,
and  it  appears  that  a high level  of removal is generally  effected
by naturally  occurring oxidation  in tailing ponds.   Except
where cyanide  is used as a leaching reagent,  high concentrations
of cyanide  are  not  normally encountered.  The use of cyanide as
a depressant  in the flotation  process  is an additional source of
cyanide in  wastewater.   Effluent  levels  characteristically
encountered are less than 0.05 mg/1 total cyanide.
                               VII-43

                               DRAFT

-------
                              DRAFT
Effective and proper use of chlorlnation or ozonation should
result in complete destruction of cyanide in mill treatment
systems.  At locations where very low levels are encountered
in wastewater streams, aeration devices, auxiliary ponds, or
long retention times may provide removal to below acceptable
levels.

Ammonia used in a solvent extraction and precipitation opera-
tion at one milling site is removed from the mill waste stream
by air stripping.  The countercurrent-flow air stripper used
at this plant operates with a pH of 11 to 11.7 and an air /liquid
flow ratio of 0.83 cubic meter of air per liter water (110
cubic feet of air per gallon of water).  Seventy-five percent
removal of ammonia is achieved by reducing total nitrogen
levels for the mill effluent to less 5 mg/1, 2 mg/1 of which
is in the form of nitrates.  Ammonia may also be removed from
waste streams through oxidation to nitrate by aeration—or, more
rapidly, by ozonation—or use of chemical oxidants, although
these procedures are less desirable due to the impact of
nitrates on the receiving water.

The removal of a variety of COD-producing pollutants from effluent
streams by oxidation in the tailing ponds and/or delivery lines
is evident in data from visited sites.  Where high reagent
dosages or other process factors lead to elevated effluent
COD levels, aeration or the use of stronger oxidants may be of
value.  In general, the use of strong oxidants in the tailing
pond will be highly undesirable, since the oxidation of sulfide
minerals in the tails can lead to increased acid production
and greater solubility of ore constituents, including heavy
metals.  Aeration will be best practiced in other impoundments
also.

Adsorption

Activated carbon is a sorptive material characterized by high
surface area within its internal pore system.  Pores generally
range from 10 to 100 Angstrom units  (0.001 to 0.01 micrometer),
and surface areas of up to 1000 square meters/gram are considered
normal  for carbons of this type.  Due to the dimensions of the
pores,  to the highly convoluted internal surface (and, thus, very
high surface area), and to the residual organic contents of
carboxyic, carbonyl, and hydroxyl compounds, activated carbon
                               VII-44

                               DRAFT

-------
                               DRAFT
exhibits adsorptive, absorptive,  and slight residual ion-
exchange capabilities.  In contrast to alumina,  silica gel,
and other adsorbents, however,  activated carbon exhibits a
relatively low affinity for water.   Compounds which are readily
removed by activated carbon include aromatica, phenolics,
chlorinated hydrocarbons, surfactants, organic dyes, organic
acids, higher-molecular-weight  alcohols, and amines.  Current
applications of this material also  center around the control
and removal of color, taste, and  odor components in water.

Activated carbon has been shown to  significantly reduce concen-
trations of a variety of inorganic  salts, including most heavy
metals.  Lead concentrations have been reduced from 100 mg/1
to 0.5 mg/1 (Reference 43).  Reports of Hg, V, Cr, Pb, Ni,
Cd, Zn, Fe, Mn, Ca, Al, Bi, Ge, As, Ba, Se, and Cu removal have
appeared in the literature—most  often, as results of laboratory-
scale treatment (References 44 and  40).

In addition to use in tertiary sewage treatment, activated carbon
has found a variety of industrial-waste applications.  At one
facility, phenols are removed from 600 cubic meters (150,000
gallons) per day of chemical plant  wastewater containing 62,000
mg/1 of total dissolved solids (Reference 45).  Influent
and effluent levels for this treatment facility are 100 mg/1
and less than 1 mg/1 of phenol, respectively.  As in this operation,
carbon may be regenerated in a furnace with approximately
95-percent carbon recovery to reduce materials cost for the
operation.

In addition to the economics of operation dictating regenerative
processes, recovery of metal values using the principles of this
treatment is possible.  Some indication of the economic success
of this approach may be gained from the reported viability of
the "resin-in-pulp" or "carbon-in-pulp" process employed at mill
4105 in the gold-recovery circuit.   In this case, cyano-complexes
of gold (and, probably, other metals) are reversibly adsorbed
from alkaline solution by activated carbon.  Activated-carbon
treatment of acid mine water has been used for iron (+2) removal
(Reference 46).

The application of carbon adsorption, or adsorption by other
materials  (such as peat), to mining and milling wastewater is
more likely to be limited by cost than by technical feasibility.
Removal of flotation or solvent-extraction reagents from waste
streams may be practical in some operations, If waste streams
                              VII-45

                               DRAFT

-------
                               DRAFT
are segregated.  Carbon adsorption could be an important factor
in achieving a high degree of water recycle in flotation mills
where reagents or decomposition products in the feed water
would interfere with processing.

Other Adsorption Methods.   While activated carbon is one specific
adsorbent used for wastewater treatment, there are many addi-
tional materials which show varying adsorptive capacities for
wastawater constituents.  Many of these candidate sorbing media
have been evaluated only in a preliminary fashion under full-
scale conditions, and few of these have been evaluated with
reference to behavior in actual mine/mill effluents.

Reported adsorbing species include tailing materials (Reference
47), waste wool  (Reference 48), silica gel, alumina,
hydrous zirconium oxide  (Reference 49), peat moss (Refer-
ence 50)f, hydrous manganese oxides (Reference 51), and
others.  The sorptive capacity of various soils is currently
under study in conjunction with increased utilization of spray
irrigation as a method of wastewater  disposal (Reference 52).

To date, little  experience in large-scale wastewater disposal
involving waters similar to mine/mill effluents has been reported
for land disposal by  spray irrigation.  Capital costs,  operating
costs, and performance experience with municipal, food-industry,
and paper-industry waste disposal, however, suggest the potential
desirability of  this  procedure  (Reference 53).  Any spray-
irrigation disposal of mine/mill wastes must be preceded by
settling systems or other  treatments  to reduce the suspended-
solid load.

Ion Exchange

Ion exchange  is  basically  a  process  for removal of various
ionic species  in or  on fixed  surfaces.  During the  fixing
process,  ions  in the  matrix  are exchanged  for soluble  ionic
species.   Cationic,  anionic,  and  chelating  ion exchangers
are available  and  may be either solid or liquid.  Solid ion
exchangers  are generally available in granular, membrane, and
bead  forms  (ion-exchange resins)  and may be employed in upflow
or downflow beds or  columns,  in agitated baskets, or in co-
current-  or countercurrent-flow modes.  Liquid ion  exchangers
are usually employed  in equipment similar  to  that employed
                               VII-46


                                DRAFT

-------
                              DRAFT
in solvent-extraction operations (pulsed columns), mixed
settlers, rotating-disc columns, etc.)-  In practice,  solid
resins are probably more likely candidates for end-of-pipe
wastewater treatment, while either liquid or solid ion exchang-
ers may be utilized in internal process streams.

Individual ion-exchange systems do not generally exhibit equal
affinity or capacity for all ionic species (cationic or
anlonic) and, so, may not be suited for broad-spectrum removal
schemes in wastewater treatment.  Their behavior and perfor-
mance are usually dependent upon pH, temperature, and concen-
tration, and the highest removal efficiencies are generally
observed for polyvalent ions.  In wastewater treatment, some
pretreatment or preconditioning of wastes to adjust suspended-
solid concentrations and other parameters is likely to be
necessary.

Progress in the development of specific ion-exchange resins and
techniques for their application has made the process attrac-
tive for a wide variety of industrial applications in addition
to water softening and deionization.  It has been used exten-
sively in hydrometallurgy—particularly, in the uranium indus-
try—and in wastewater treatment (where it often has the
advantage of allowing recovery of marketable products).  This
is facilitated by the requirement for periodic stripping or
regeneration of ionic exchangers.  If regeneration produces
a solution waste, its subsequent treatment must be considered.

Table VII-2 shows different types of ion-exchange resins and
the range of conditions and variety of purposes for which
they are employed.

Disadvantages of using ion exchange in treatment of mining
and milling wastewater are relatively high costs, somewhat
limited resin capacity, and insufficient specificity—
especially, in cationic exchange resins for some applica-
tions.

Although it is suitable for complete deionization of water,
ion exchange Is generally limited in this application, by
economics and resin capacity, to the treatment of water con-
taining 500 mg/1 or less of total dissolved solids.  Since
IDS levels in mining and milling effluents are often higher
than this level, application of ion exchange to the economic
reduction of total dissolved solids at high flow rates must
be evaluated.
                              VII-47

                              DRAFT

-------
                     DRAFT
TABLE VI1-2.  PROPERTIES OF  ION EXCHANGERS FOR
            METALLURGICAL APPLICATIONS
DESIRED
CHARACTERISTIC
CHEMICAL
STABILITY TO:
PHYSICAL STABILITY FOR:
Acidi
Alkalies
Oxidation
Temperature
Organic Solvents
Removal of weak
acids
Removal of strong
acids
High regeneration
efficiency
High capacity
High porosity
Hydrogen exchange
at low pH
Salt splitting
pH range (operating)
GENERALLY RECOMMENDED APPLICATION
CATION EXCHANGERS
Inorganic
Zeo-Dur




•







6.2 to
8.7
SI
I




•



•
•


6.9 to
7.9
Organic
Sulfonated
Coal
Zeo-Karb
•






•

•
•

0 to
11
Resins
Permutit Q
•
•
•
•
•



•

•

0 to
13
Car-
boxylic
Resin
Permutit H-70
•
•
•
•



•
•



3.5 to
12
ANION EXCHANGERS
Weakly
Basic
Gran-
ular
De-Acidite
•
•




•
•
•
•


0 to
12
Bead
Permutit W
•
•
•
•
•

•





0 to
13.9
Strongly
Basic
Gran-
ular
Permutit A
•
•



•
•

•


•
0 to
13.9
Bead
Permutit S
•
•
•
•
•
•


•


•
0 to
13.9
           SOURCE: Reference 54
                     VII-48
                     DRAFT

-------
                            DRAFT
For recovery of specific ions or groups of ions (e.g.,  dival-
ent heavy-metal cations, or metal anions such as molybdate,
vanadate, and chromate), ion exchange is applicable to a much
broader range of solutions.  This use is typified by the
recovery of uranium from ore leaching solutions using strongly
basic anion-exchange resin.  As additional examples, one may
consider the commercial reclamation of chromate plating and
anodizing solutions, and the recovery of copper and zinc
from rayon-production wastewaters (Reference 54).
Chromate plating and anodizing wastes have been purified and
reclaimed by ion exchange on a commercial scale for some time,
yielding economic as well as environmental benefits.  In
tests, chromate solutions containing levels in excess of
10 mg/1 chromate, treated by ion exchange at practical resin-
loading values over a large number of loading elution cycles,
consistently produced an effluent containing no more than
0.03 mg/1 of chromate.

High concentrations of ions other than those to be recovered
may interfere with practical removal.  Calcium ions, for
example, are generally collected along with the divalent
heavy-metal cations of copper, zinc, lead, etc.  High calcium-
ion concentrations, therefore, may make ion-exchange removal
of divalent heavy-metal ions impractical by causing rapid
loading of resins and necessitating unmanageably large resin
inventories and/or very frequent elution steps.  Less diffi-
culty of this type is experienced with anion exchange.  Avail-
able resins have fairly high selectivity against the common
anions,  such as Cl(-) and  SOji(-2).  Anions adsorbed along
with uranium include vanadate, molybdate, ferric sulfate
anionlc  complexes, chlorate, cobalticyanide, and polythionate
anions.  Some solutions containing molybdate prove difficult
to elute and have caused problems.

Ion-exchange resin beds may be  fouled by  particulates, pre-
cipitation within the beds, oils and greases,  and biological
growth.  Pretreatment of water, as discussed earlier,  is
therefore, commonly required for successful operation.  Gen-
erally,  feed water  is  required  to be treated by coagulation
and  filtration  for  removal of  iron and manganese, C02_, H^S,
bacteria and algae, and hardness.  Since  there is some lati-
tude  in  selection of  the ions  that are  exchanged for  the  con-
taminants  that  are  removed,  post-treatment may or may  not  be
required.
                             VII-49

                             DRAFT

-------
                            DRAFT
Since, in many cases, calcium is present in ore mining and
milling wastewater in appreciably greater concentrations than
are the heavy-metal cations whose removal to low levels is
sought, use of ion exchange in that mode may have several
disadvantages, since costs would be high, and little advan-
tage would be offered over lime or sulfide precipitation.
For the removal of anions, however, the relatively high
costs of ion-exchange equipment and resins may be offset par-
tially or totally by the recovery of a marketable product.
This has been demonstrated in the removal of uranium from
mine water, and the removal of molybdate anions is now under
investigation in pilot-plant studies at two operations,
although results are not yet available.  The application of
this technique will depend upon a complex set of factors,
including resin loading achieved, pretreatment required, and
the complexity of processing needed to produce a marketable
product from eluent streams.

The practicality of the ion-exchange process will be enhanced
by practices such as waste segregation, recycle, etc., which
allow the treatment of smaller volumes of more concentrated
solutions.  Similar factors apply to the treatment of mining
and milling waste streams bearing vanadate and chromate anions,
although prior experience in ion-exchange recovery of these
materials should aid the development of treatment schemes for
such wastes.

Modified Desal Process.  A demonstration plant for generating
potable water from acid coal-mine drainage, in operation since
early 1973, treats 3,028 cubic meters  (800,000 gallons) per day
of water which contains pollutant loadings similar to
those of acid mine drainage (Reference 55).  The plant
was originally designed for a capacity of 1,893 cubic meters
(500,000 gallons) per day, but it is expected that the plant's
capacity can be further increased to 3,785 cubic meters
(1,000,000 gallons) per day through use of improved operating
techniques.

The Modified Desal Process portrayed in Figure VII-5 is a
variation of a system originally developed to produce potable
water from brackish supplies by means of cation- and anion-
exchange resins.  The primary purpose of ion exchange in treat-
ing acid mine water, however, is to remove sulfate, so only
                             VII-50

                             DRAFT

-------
                                       DRAFT
                Figure VI1-5. DIAGRAM OF MODIFIED DESAL PROCESS
FROM

MINE
                                                         I
                                                         I
SETTLING
BASINS






GRAVITY
FILTERS


PRODUCT
WATER
                                                  LEGEND
                                                MAIN PROCESS

                                               • ADDITIONS OR LOSSES

                                               • REGENERATION PROCESS
                                     SOURCE: Reference 66
                                       VII-51





                                       DRAFT

-------
                              DRAFT
an anion-exchange resin is necessary.  The process uses a weak-
base anion resin in the bicarbonate form to replace sulfate or
other anions.  The solution of metal bicarbonates is aerated
to oxidize ferrous iron to the ferric form and to purge the
carbon dioxide gas.  The increase in pH causes iron, aluminum,
and manganese to precipitate as insoluble hydrous oxides.
Some calcium and magnesium carbonates also precipitate.  To
produce improved quality water, well within potable limits,
lime treatment precipitates more calcium and magnesium by
converting the bicarbonates into less soluble carbonates.

The exhausted resin is regenerated with ammonium hydroxide,
which converts the resin to the free-base form.  Introduction
of carbon dioxide converts the resin back to the bicarbonate
form, and the regenerated solution of ammonium sulfate is pro-
cessed to recover the ammonia through lime addition.  The
resultant calcium sulfate is transported to mine pits for dis-
posal.  Regeneration occurs after about 18 hours of operation,
and the plant currently utilizes the original ion-exchange
resin.

Operating data for the plant are shown in Table VII-3.  It
is felt that this system, or a modification thereof, might
provide effective removal of sulfate and dissolved solids in
the ore mining and dressing industry.

Present operating costs  for water produced at the Phillipsburg,
Pennsylvania, plant are  $0.40 to 0.50 per 3.79 cubic meters
(1,000 gallons)  of water.  However, a considerable reduction
in cost might be achieved  for the mining  industry for two
reasons.  The first  is  that the demonstration plant contains
much  instrumentation  and many features that would be unnec-
essary in a  facility  designed merely  for  production.  Secondly,
integration  of the  ion-exchange system with presently existing
lime-neutralization  plants  could eliminate the necessity  for
many  features of the  Modified Desal  Process system.

Although  the cost  for treating  3.79  cubic meters  (1,000  gallons)
of raw mine  drainage  appears  favorable, volumes  in  excess  of
57,000 cubic meters  (15,000,000 gallons)  of drainage generated
daily at many  facilities  require a  substantial  total invest-
ment  in  time, material  resources, and energy.  Also,
individual  treatment  plants with design capacities  of
up  to 34,065 cubic meters  (9,000,000 gallons)  per day would
necessitate  the  installation  of multiple  ion-exchange  units
                               VII-52

                               DRAFT

-------
                       DRAFT
TABLE VII-3. ANALYTICAL DATA FOR MODIFIED DESAL PROCESS
PARAMETER
pH
Total hardness (CaCO3)
TDS
Calcium (CaCO3)
Magnesium (CaCOg)
Iron
Sulfate
CONCENTRATION (mg/£ )
RAW WASTEWATER
3.7»
395
1,084
295
100
101
648
EFFLUENT WATER
9.5"
184
284
85
99
0.2
192
 •Value in pH units
                        VII-53





                        DRAFT

-------
                              DRAFT
at most discharge outfalls.  This configuration would greatly
decrease cost effectiveness for a treatment aimed specifically
at removing sulfate and dissolved solids.

Ultrafiltration and Reverse Osmosis

Ultrafiltration and reverse osmosis are similar processes in
which pressure is used to  force water through membranes which
do not allow passage of contaminants.  They differ in the
scale of contaminants passed and in the pressures required.
Ultrafiltration generally  retains particulates and materials
with a molecular weight greater than 500, while reverse-
osmosis membranes generally pass only materials with
a molecular weight below 100  (Sodium chloride, although below
a molecular weight of 100, is  retained, allowing application
to desalinization) .  Pressures used in Ultrafiltration gen-
erally range from 259 to 517 cm of Hg (50 to 100 psi), while
reverse osmosis is run at  pressures ranging from 2,068 to
9,306 cm of Hg  (400 to 1,800 psi).

Ultrafiltration has been applied on a significant commercial
scale to the removal of oil from oil emulsion, yielding a
highly purified water effluent and an oil residue sufficiently
concentrated to allow reuse,  reclamation, or combustion.
Equipment  is readily available, and present-day membranes are
tolerant of a broad pH range.  Application of Ultrafiltration
to mining  and milling waste streams, where high dosages of
oils are used in  flotation—as at a formerly operated man-
ganese mill—may  provide a practical technique for removing
these waste components, possibly allowing reuse as well.

Reverse osmosis (RO)  is conceptually similar to ultrafiltra-
tion.  It  also  involves  the application  of an external pressure
to a solution  in  contact with a  semipermeable membrane to
force water through  the membrane while  excluding both soluble
and  insoluble  solution  constituents.   In its rejection of
soluble constituents,  reverse osmosis  performs a water-treat-
ment  function not fulfilled by Ultrafiltration systems under
simple operating  conditions.

Reverse osmosis is considerably  less  tolerant of  input-stream
variations in  conditions  and  requires,  in general, considerable
pretreatment.   Concentration  of  wastes  is generally  limited
by saturation  of  solutions and the formation of  precipitates,
                               VI I-54

                               DRAFT

-------
                             DRAFT
which can decrease the effectiveness of the apparatus.   As a
result, residual volumes of waste in the mining and milling
industry would, in many cases, be unmanageably large.   A
pilot-plant operation has been run on mine drainage streams,
and production of a high-quality water effluent has been
shown to be technically feasible.  Pretreatment requirements,
costs, and the problems of disposal of residual wastes make
the practicality and economic achievability important  con-
siderations.

Reverse osmosis has been demonstrated capable of rejecting
heavy-metal species from purified water streams with a high
degree of efficiency (Table VII-4).  Reverse-osmosis systems
have been evaluated for acid mine water treatment (References
57 and 58).  Related studies have been conducted with metal-
finishing effluents (Reference 59).  In most instances,
pretreatment of water, and conditioning with respect to pH,
temperature, and suspended-solid levels, is necessary for
reverse-osmosis module use.  Membrane lifetime and constancy
of efficiency are both adversely affected by inadequate treat-
ment of waters prior to membrane contact.  In general, labora-
tory performance of reverse-osmosis systems has shown some-
what higher purification efficiencies than have been observed
in pilot-plant operations (Reference 40).  The present
state-of-the-art with regard to RO technology indicates that
details of extrapolation of laboratory and current pilot-plant
data to full-scale operation need to be worked out.  Data on
membrane lifetime, operating efficiency, rejection specificity,
and other factors remain to be more fully quantified.

High-Density-Sludge Acid Neutralization

The conventional lime neutralization of acid or mine wastes
usually leads to the formation of low-density sludges which
are difficult to dewater (floes).  The use of ground lime-
stone avoids this problem but does not allow for the attain-
ment of pH levels necessary to effectively remove such metals
as zinc and cadmium.  A process which utilizes extensive
recycle of the previously precipitated sludge allows the
attainment of sludges of much higher density, thus allowing
more rapid sedimentation of the sludges ultimately produced
and easing solid-disposal problems.
                              VII-55

                             DRAFT

-------
                    DRAFT
TABLE VI1-4. REJECTION OF METAL SALTS BY REVERSE-
           OSMOSIS MEMBRANES
PARAMETER
Iron
Magnesium
Copper
Nickel
Chromium (hexavalent)
Strontium
Cadmium
Silver
Aluminum
TYPICAL REJECTION PERCENT
99
98
99
99.2
97.8
99
98
96
99
 SOURCE:  Reference 40
                      VII-56


                     DRAFT

-------
                              DRAFT
Solvent Extraction

Solvent extraction is a widely utilized technique for the
separation and/or concentration of metallic and nonmetallic
species in the mineral processing industry.  It has been
applied to commercial processing of uranium, vanadium, tung-
sten, thorium, rhenium, rare earths, beryllium, columbium,
copper, zirconium, molybdenum, nickel,  boron, phosphoric acid,
and others (References 60 and 61).  Reagent-processing
equipment for this technique is highly developed and generally
available (Reference 62).  It is anticipated that such
equipment would require modification to be applicable to treat-
ing the low levels of soluble metals in most waste streams.
Pretreatment and post-treatment of waters treated by this
technique would probably be required to control influent pH,
suspended solids, and other parameters, as well as effluent
organic levels.  It is likely that this treatment strategy  may
be most applicable in internal process streams or as an add-on
for the recovery of values from waste-concentration streams
such as distillate or freeze residues,  reverse-osmosis brines,
etc.

Because of the speculative nature of solvent extraction as
applied to wastewater treatment, the unknown costs of rea-
gents, and possible pretreatment/post-treatment demands,
accurate treatment or capital costs for this option do not
appear readily derivable at this time.

Evaporation and Distillation

Evaporation may be employed as a wastewater-treatment tech-
nique in a variety of ways:

      (1)  Total evaporation of wastewater may produce solid
           residues and eliminate effluent water discharge.

      (2)  Concentration of wastewater by evaporation may
           balance dilution by makeup and infiltration water
           and allow for an approach to total recycle, thus
           minimizing discharge volume.  The buildup of detri-
           mental species upon evaporation will normally
           require a bleed stream from the evaporation system,
           thus precluding total water recycle.  A bleed stream,
           of course, might be handled by total evaporation,
           rather than by discharge to a waterway.
                             VII-57

                             DRAFT

-------
                               DRAFT
      (3)   Concentration by evaporation may  allow subsequent
           removal of concentrated wastewater  components  to
           acceptable levels for smaller-volume  discharge or
           reuse.

      (4)   Ultimately, complete distillation of  wastewater
           may allow the almost total reuse  or recycle of
           contained water, while rendering  discharge unnec-
           essary  and allowing potential recovery of values
           from nonvolatile residues.   In the  absence of
           recoverable values,  disposal of sludge resulting
           from distillation might become a  problem of sub-
           stantial magnitude.   The presence of  volatile
           wastes  in the effluent may require  additional
           treatment of distillate to achieve  adequate quality
           for some uses.

Energy  sources for evaporation may be artificial (steam, hot
gases,  and electricity) or natural (solar,  geothermal, etc.).
In present practice,  many  of the mining and milling operations
in the  Western and Southwestern United States employ solar
evaporation as a  principal means of water treatment.  Evapora-
tive  losses of water at some installations  may  exceed 7,572
cubic meters  (2,000,000 gallons)  per year for each 0.4 hectare
(1 acre) of evaporative surface;  with adequate  surface acreage,
this  loss  may allow for zero-effluent-discharge  operation.
At present, this  evaporated water is  not  collected for reuse
at these operations.

A multistage  flash-distillation process has been applied to
treat acid mine drainage  (from a coal mine) in a pilot plant
(Reference 63  ) •   The process is mechanically complex but
results in a  solid residue and essentially  pure  water, suit-
able  for human consumption.  This approach  to pollution con-
trol  involves  the use of considerable energy  associated with
vaporizing vast volumes of water.   Its technical applicability
to treating mine  water has been demonstrated, but it is not
clear that organic wastes  potentially present in mill effluents
would be successfully controlled  by such  a  process.

Techniques for  Reduction of_ Wastewater Volume

Pollutant  discharges  from  mining  and  milling  sites may be
reduced by limiting the total  volume  of_ discharge, as well
                              VII-58

                              DRAFT

-------
                                 DRAFT
as by reducing pollutant concentrations In the waste stream.
Volumes of mine discharges are not, In general, amenable to
control,  except Insofar as the mine water may be used as Input
to the milling process In place of water from other sources.
Techniques for reducing discharges of mill wastewater include
limiting  water use, excluding incidental water from the waste
stream, recycle of process water, and impoundment  (with water
lost to evaporation or seepage).

In most of the industry, water use should be reduced to the
extent practical, because of the existing incentives for doing
so (i.e., the high costs of pumping the high volumes of water
required, limited water availability, and the cost of water-
treatment facilities).  Incidental water enters the waste
stream primarily through precipitation directly and through
the resulting runoff  influents to tailing and settling ponds.
By their  very nature, the water-treatment facilities are sub-
ject to precipitation inputs which, due to large areas, may
amount to substantial volumes of water.  Runoff influxes are
often many times larger, however, and may be controlled to a
great extent by diversion ditches and (where appropriate)
conduits.  Runoff diversion exists at many sites and is under
development at others.

Recycle of process water is currently practiced primarily where
it is necessary due to water shortage, or where it is economi-
cally advantageous because of high water costs.  Recycle to
some degree is accomplished at many ore mills, either by
reclamation of water at the mill or by the return of decant
water to  the mill from the tailing pond or secondary impound-
ments.  Recycle is becoming, and will continue to become, a
more frequent practice.  The benefits of recycle in pollution
abatement are manifold and frequently are economic as well as
environmental.  By reducing the volume of discharge, recycle
not only reduces the gross pollutant load,  but also allows
the employment of abatement practices which would be uneconomic
on the full waste stream.   Further, by allowing concentrations
to increase,  the chances for recovery of waste components to
offset treatment cost—or, even,  achieve profitability—are
substantially improved.  In addition,  costs of pretreatment of
process water—and,  in some instances, reagent use—may be
reduced.

Recycle of mill water almost always requires some treatment of
water prior to its reuse.   In many instances,  however,  this may
                               VII-59

                                 DRAFT

-------
                             DRAFT
entail only the removal of solids In a thickener or tailing
basin.  This is the case for physical processing mills, where
chemical water quality is of minor importance, and the practice
of recycle is always technically feasible for such operations.
In flotation mills, chemical interactions play an important
part in recovery, and recycled water can, in some instances,
pose problems.  The cause of these problems, manifested as
decreased recoveries or decreased product purity, varies and
is not, in general, well-known, being attributed at various
sices and times to circulating-reagent buildup, inorganic salts
in recycled water, or reagent decomposition products.  Exper-
ience in arid locations, however, has shown that such problems
are rarely insurmountable.  In general, plants practicing bulk
flotation on sulfide ores can achieve a high degree of recycle
of process waters with minimal difficulty or process modifica-
tion.  Complex selective flotation schemes can pose more
difficulty, and a fair amount of work may be necessary to
achieve high recovery with extensive recycle in such a circuit.
Numerous examples where this has been achieved may be cited
(Reference    64   ).  Problems of achieving successful recycle
operation in such a mill may be substantially alleviated by
the recycle of specific process streams within the mill, thus
minimizing reagent crossover and degradation.  The flotation
of non-sulfide ores  (such as scheelite) and various oxide
ores using fatty acids, etc., has been found to be quite
sensitive to input water quality.  Attempts at water recycle
in such operations have posed severe problems, and successful
operation may require a high degree of treatment of recycle
water.  In many cases, economic advantage may still exist over
treatment to levels which are acceptable for discharge, and
examples exist in current practice where little or no treat-
ment of recycle water has been required.

Technical limitations on recycle in ore leaching operations
center on inorganic  salts.  The deliberate solubilization of
ore components, most of which are not to be recovered, under
recycle operations can lead to rapid buildup of salt loads
incompatible with subsequent recovery steps (such as solvent
extraction or ion exchange).  In addition, problems of corro-
sion or sealing and  fouling may become unmanageable at some
points in the process.  The use of scrubbers for air-pollution
control on roasting  ovens provides another substantial source
of water where recycle is limited.  At leaching mills, roasting
will be practiced to increase solubility of the product material.
Dusts and fumes from the roasting ovens may be expected to
                              VII-60

                              DRAFT

-------
                             DRAFT
contain appreciable quantities of soluble salts.   The buildup
of salts in recycled scrubber water may lead to plugging of
spray nozzles, corrosion of equipment, and decreased removal
effectiveness as salts crystallizing out of evaporating scrubber
water add to particulate emissions.

Impoundment is a technique practiced at many mining and milling
operations in arid regions to reduce point discharges to, or
nearly to, zero.  Its successful employment depends on favorable
climatic conditions (generally, less precipitation than evapo-
ration, although a slight excess may be balanced by process
losses and retention in tailings) and on availability of land
consistent with process-water requirements and seasonal or storm
precipitation influxes.  In some instances where impoundment
is not practical on the full process stream, impoundment and
treatment of smaller, highly contaminated streams from specific
processes may afford significant advantages.  Many operations
currently practicing Impoundment achieve a sufficiently favor-
able water balance to allow zero surface discharge only through
major seepage losses to ground water.  The development of
restrictions on seepage may have substantial impact on the
practice of impoundment.

Electrodialysis

Electrodialysis is fundamentally similar to both reverse osmosis
and ultrafiltration to the extent that it employs semipermeable
membranes to allow separation of soluble cationic and anionic
impurities from water.  An imposed electrical field is used
to provide a driving force for ion migration, in analogy to
either osmotic or external pressure in reverse-osmosis, dialytic,
or ultrafiltration systems.

Electrodialysis is generally employed in the treatment of waters
containing less than 5,000 to 10,000 mg/1 of dissolved solids
to achieve final levels of less than 500 mg/1 (Reference   39   ).
Applications have been reported in desalinization of seawater
involving feed water containing 38,000 mg/1 chloride and producing
a product water containing 500 mg/1 chloride (Reference    49  ).

To date, electrodialysls has not been employed in large-scale
operations within the mining/milling industry segments reviewed
and studied In this program.  The potential for isolation and
recovery of byproduct or waste values exists but has not been
confirmed.
                             VII-61


                             DRAFT

-------
                               DRAFT
 Freezing

 This process depends on the formation of pure ice  crystals
 from the contaminated solution being treated.   Results of
 freezing experiments on acid mine-drainage samples (from a coal
 mine)  indicates that suspended solids act as  condensation
 nuclei and,  if present, are entrained with the "pure" ice
 obtained.   Once solids have been removed, of  course, the mine
 drainage may still contain other contaminants.

 Experimentally, agitation and slow freezing rates  have allowed
 reductions  in dissolved materials in the range of  35 to 90
 percent (Reference  40    ).

 This process results in a concentrated stream,  which still
 requires treatment.   It has a theoretical advantage over
 distillation because only about one-sixth of  the energy should
 be required.   Laboratory-scale experiments indicate it may be
 a feasible  treatment technique for mine and mill water treat-
 ment,  but it has not been fully tested.

 Biological Treatment

 The  ability  of various biota—both flora and  fauna—to assimilate
 soluble constituents from contacting waters is  being documented
 with increasing frequency.   In general,  these  studies have
 considered the undesirability of  such assimilations, rather
 than viewing  them from the standpoint of potential water-
 treatment options or systems.   If  trace  or toxic constituents
 can  be  metabolized,  detoxified, or fixed by various organisms,
 the  periodic  removal of organisms  containing concentrates of
 these materials may  be a viable removal  mechanism.

 The  use  of this technique at  one  facility visited involves a
 combination of sedimentation  ponds and biological treatment
 in the  form of  meanders.   The meander system is an artificial
 system  designed to contain—and,  thereby,  control—excessive
algal growth  and  the associated heavy metals which are trapped
and  assimilated by the algae  (Reference    65    ).  The algal
growth occurs  naturally and was a  problem associated with
the discharge  prior  to installation  of the present system.
The system was  designed as a  series  of broad,  shallow,  rapidly
flowing meanders, which increase the  length of  the treatment
section and encourage  the  growth of algae before discharge,
                              VII-62

                              DRAFT

-------
                              DRAFT
while simultaneously trapping any suspended  heavy metals.
To prevent the algae and the associated  heavy metals  from
escaping the system, an additional final sedimentation  pond
is placed at the end of the system.

The system can be effective if sufficient land  is available
to allow the construction of an adequate meander system, and
if the climate is such that algae growth is  not precluded during
parts of the year.   These conditions  effectively prevent wide-
spread application of this treatment  technique.
                              VII-63

                              DRAFT

-------
                             DRAFT
EXEMPLARY TREATMENT OPERATIONS BY ORE CATEGORY

The manner  in which ore mine and mill operators have approached
the design  and  construction of treatment and control facilities
varies from quite  simple  to somewhat sophisticated (utilizing
recycling,  zero-discharge operations).  To attain extensive
recycling or zero  discharge, extensive process changes and/or
redesign have often been  necessary.  Performance of the many
vdired operations  used In each ore category varies with the
operating characteristics of the facility, the ore mineralogy,
and other factors.   Descriptions, by ore category, of the treat-
ment and control processes used in the ore mining and dressing
industry and the consequent treatment levels attained are
Included here to provide  a more complete explanation and examina-
tion of the control and treatment technology currently in use.

Iron Ore

This discussion includes  examples of mines that have discharges
(Subcategory I), mills which employ physical or chemical
benefIciation (Subcategory II), and mills using magnetic- and
physical-separation methods to extract iron from Mesabi Range
ifon formations.

Mining Operations.   Mine 1105 is an open-pit operation that
accumulates water.   Water is pumped directly from the pit to a
settling pond of sufficient volume to remove suspended solids
prior to discharge.  No chemical coagulants are used, because
the suspended-solid concentration generally is less than 10
mg/1.  Because  this operation produces low levels of dissolved
components, dissolved-solid treatment is unnecessary.  Suspended-
solid concentrations after treatment have been observed to
remain low, but historical data obtained during periods of high
rainfall and high  pumping rates are lacking.

Table VII-5 is  a compilation of data measured in this study and
by the operators.   It can be observed that many of the parameters
measured appear to  increase in the effluent stream after treatment,
Measurements made  during  this study were confirmed by duplicate
Industry sample analysis.  Conditions existing at the mine
settling pond should be noted, however.  At the mine discharge,
an extremely low flow was encountered,  and only intermittent
pumping of  the mine was being employed.  At the settling-pond
discharge, however,  flow  conditions were adequate for sampling.
                            VII-64

                             DRAFT

-------
                             DRAFT
TABLE VI1-5. CHEMICAL CHARACTERISTICS OF SETTLING-POND DISCHARGE AT
          MINE 1105
PARAMETER
PH
TSS
TDS
COD
Oil and Grease
Total Fe
Dissolved Fe
Mn
Sulfate
AVERAGE
MINE-DISCHARGE
CONCENTRATION (mQ/Jt)
This Study
7.4*
10
225
9.7
<1
<0.02
< 0.02
0.04
24
Industry
7.9*
6
243
4.5
< 5
—
^0.1
^ 0.1
—
AVERAGE
SETTLING-POND
DISCHARGE
CONCENTRATION (mg/£)
This Study
7.4*
25
283
13.7
<1
0.1
<0.02
<0.02
35
Industry
8.0*
as
291
15
<5
-
^0.1
^ 0.1
—
AVERAGE
SETTLING-POND
DISCHARGE
CONCENTRATION
(mg/£)
8.0*
3.4
-
-
(<10)
—
-
-
—
 Value in pH units
                             VII-65


                              DRAFT

-------
                             DRAFT
Historical data  obtained  at  this  location for nine months during
1974  show that a range  of 1  to  9  (average of 3.4) mg/1 of
TSS was encountered  after settling.

Mills Employing  Physical  and/or Chemical Separations.   Iron-
beneflciation plant  1109  uses magnetic  separation, coupled
with a froth-flotation  sequence that removes undesired silica
in the iron concentrate.   The processing circuit uses 587
cubic meters (155,000 gallons)  of water per minute, with a
recycle rate of  568  cubic meters  (150,000 gallons) per minute.
Thickeners, located  adjacent to the concentrator, are used to
reclaim water close  to  the site of reuse so as to minimize
pumping requirements.   Superfloe  16, an anionic polyacrylamide,
is added to the  thickeners at a rate of 2.5 grams per metric
ton (0.0049 pound per short  ton)  of mill feed to aid in clarifi-
cation of the water  in  the thickeners.  The thickener underflow
is pumped to a 850-hectare (2,100-acre) tailing basin for the
sedimentation of the solids.  Mine water is also pumped to the
basin.  The effluent leaves  the basin after sufficient retention
and flows into a creek  at an average rate of 22.3 cubic meters
(5,900,000 gallons)  per day.  Chemical  analysis of the waste-
water to the tailing pond (mine and mill water) in comparison
to the effluent  water quality and waste loading is given in
Table VII-6.

Mills Employing  Magnetic  and Physical Separations.   Mill 1105
is located in the Mesabl  Range  of Minnesota and is processing
ore of the Biwabik formation.   Crude magnetic taconite is
milled to produce a  fine  magnetite.  The mill's water system
is a closed loop having no point-source discharges to the
environment.  The plant processes use 20.4 cubic meters (54,000
gallons) per minute, with 189 cubic meters (50,000 gallons)
per minute returned  from  the tailing-thickener overflow and
15.1 cubic meters (4,000  gallons) per minute returned from the
tailing pond or  basin.  The tailing thickener accumulates all
the milling-process  wastewater  containing the tailings.  A
nontoxic polyacrylamide flocculant (SuperFloc 16) is added to
the thickener to assist the settling out of solids.  Tailing-
thickener underflow  is  pumped to  a tailing basin of 470 hectares
(1,160 acres), where the  solids are settled and the clear water
is recycled back into the  plant water-use system.  A simplified
water-use sequence is shown in  Figure VII-6.
                             VII-66

                             DRAFT

-------
                                 DRAFT
       TABLE VII-6. CHEMICAL COMPOSITIONS OF RAW AND TREATED
                   WASTELOADING AT MINE/MILL 1109
PARAMETER
PH
TSS
TDS
COD
Total Ft
Oinolnd Ft
Mn
Sultal*
Alkalinity
MINE EFFLUENT
CONCENTRATION
(main
8.3"
12
308
27.8
0.30
0.02
0.66
37
181
MILL EFFLUENT
CONCENTRATION
(mo/£>
8.6"
(66%)
360
138
0.04
0.04
-
20.7
238
WASTE 1
PER UNIT F
kg/rmtrle ton

1.346
0.88
0.033
0.0001
0.0001
-
0.06
0.68
-OAD
RODUCT
Ib/ihort ton

2.890
178
0.066
0.0002
0.0002
-
010
1 16
FINAL DISCHARGE
CONCENTRATION
Img/ 1)
8J"
10
222
18X1
0.76
OM
<0.02
3£
120
WASTE I
PER UNIT P
kg/mttrle ton

0.02
0.48
0.039
0.0018
00010
< 0.00004
00078
0.26
.OAD
RODUCT
Ib/thort ton

0.04
0.86
0.078
0.0032
0.0020
0.00008
00162
0.52
HISTORICAL
CONCENTRATION*
(mg/t>
7.7"
34
_
_
_
0.60
0.06
_
-
 Anrige of nlno raluii (August through October 1974)
•
 Vilu* In pH unln.
                                 VII-67


                                 DRAFT

-------
                        DRAFT
Figure VI1-6. MILL 1105 WATER-USE SYSTEM (ZERO DISCHARGE)

WATER

1 '
^._.., ..... p

PROCESS
PRODUCT
\ >
THICKENING
1
OVERFLOW UNDERFLOW
	 t
FILTRATION
1 1
CAKE FILTRATE
1 1
\
TO FINAL
PROCESSING
ROCESS PLANT i
1
PROCESS
TAILING
1


THICKENING
1
UNDERFLOW OVERFLOW
/""SEDIMENTATIONS
\_^ BASIN^./
SETTLED CLARIFIED
SOLIDS EFFLUENT
|

                             TO WASTE
                         VII-68





                        DRAFT

-------
                              DRAFT
Copper Ores

The discussion that follows describes treatment and control
technology in current use in the five subcategories of the
copper-ore mining and dressing Industry.

Mining Operations.   Mine water generated from natural drain-
age is reused in mining, leaching, and milling operations wherever
possible in the copper mining industry.  Because of an excess
of precipitation in certain areas of the country, a location
which is not proximate to a milling facility, or an inability
to reuse the entire amount of mine wastewater at a particular
mill, a discharge may result.  The amounts of precipitation
and evaporation thus have an Important influence on the
presence or absence of mine-water discharge.

To avoid discharge, mine effluent .may be reused in dump, heap,
or in-situ leaching as makeup water.  As a leach solution, it
is acidified (if necessary), percolated through the waste dump,
sent through an iron-precipitation facility, and recycled to
the dump (Figure VII-7).

Large quantities of water are usually needed in the copper
flotation process.  Mine-water effluent is used at many
facilities as mill process makeup water.  The mine water may
pass through the process first, or it may be conveyed to the
tailing pond, from which it is used for mill flotation with
recycled process water (Figure VII-8).  The practice of com-
bining mine water with mill water can create water-balance
difficulties unless the mill circuit is capable of handling
the water volumes generated without a discharge resulting.
The discharge of mine water into a mill process system which
creates an excess water balance and subsequent discharge may
have a detrimental effect on the mine water because of contam-
ination by mill flotation reagents and residual wastes.

Acid mine water is encountered in the copper mining
industry, and methods of neutralization usually employed
Include the addition of lime and limestone.

Lime precipitation is also often used to enable the removal of
heavy metals from wastewater by precipitation as hydroxides.
Tables VII-7 and VII-8 show examples of the use of lime
precipitation for treatment of mine water at two locations of
mine 2120.  The use of this treatment technology yields reduc-
                              VII-69

                              DRAFT

-------
                               DRAFT
Figure VII-7. CONTROL OF EFFLUENT BY REUSE OF MINE WATER IN LEACHING
           (MINE 2122)
                                 EVAPORATION
                                 AND SEEPAGE
        MINE
— EFFLUENT
  3270 m3/day
  (864,000 gpd)
 STORAGE
RESERVOIR
EVAPORATION
AND SEEPAGE


DUMP LEACH
BED
1
PREGNANT
SOLUTION
                                                   RECYCLED
                                                    BARREN
                                                   SOLUTION
                                     IRON
                                 PRECIPITATION
                                    PLANT
                                      I
                                    CEMENT
                                    COPPER
                                      TO
                                   STOCKPILE
                                 VII-70
                                DRAFT

-------
                          DRAFT
Figure VI1-8. CONTROL OF MINE-WATER EFFLUENT BY REUSE IN THE
          CONCENTRATOR (MINE/MILL 2119)
                                           MINE
                MILL/
            CONCENTRATOR
                                  37,100 m3/day
                                     (9,792,000
                                         gpd)
               TAILING
             THICKENERS
 RECYCLED
'OVERFLOW
                                   RECYCLED
                                     POND
                                     WATER
            (NO DISCHARGE)
                          VII-71


                           DRAFT

-------
                               DRAFT
TABLE VI1-7. CONCENTRATION OF PARAMETERS PRESENT IN RAW WASTEWATER
           AND EFFLUENT FOLLOWING LIME PRECIPITATION AT MINE 2120B
PARAMETER
PH
TDS
TSS
Oil and Grease
TOC
COD
B
Cu
Co
As
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
CONCENTRATION (mg/£)
RAW WASTEWATER
6.1 •
2.200
40
<1
3.2
<10
0.04
5.3
0.1
< 0.07
31.25
<0.5
6.0
26.5
0.175
0.13
< 0.5
1.55
0.0005
< 0.1
TREATED WASTEWATER
12.7*
3.000
34
<1
1.2
<10
< 0.01
0.05
<0.04
<0.07
0.11
<0.5
<0.1
0.04
< 0.005
<0.05
<0.5
0.85
0.0002
<0.1
EFFICIENCY OF TREATMENT
IN REMOVAL OF POLLUTANTS
(% REMOVAL)
INCREASED
INCREASED
15%
-
63%
-
>75%
99%
>60%
-
99.7%
-
>98%
99.9%
>71%
62%
-
45%
60%
^™
 •Value in pH units
                               VII-72


                               DRAFT

-------
                                 DRAFT
TABLE VI1-8. CONCENTRATION OF PARAMETERS PRESENT IN RAW WASTEWATER
            AND EFFLUENT FOLLOWING LIME PRECIPITATION AT MINE 2120C
PARAMETER
PH
TDS
TSS
Oil & Grease
TOC
COD
S04
Cu
Co
Ai
Zn
Sb
Fe
Mn
Cd
Ni
Mo
Sr
Hg
Pb
CONCENTRATION (mg/«)
RAW WASTEWATER*
4.7"
450
35
17
2.3
<10
300
6.2
0.06
<0.07
6.2
<0.5
8.6
1.42
0.03
<0.05
<0.05
0.09
0.0005
<0.1
TREATED WASTEWATER*
7.8"
—
3
—
—
.-
220
0.25
—
0.004
0.45
—
0.5
—
0.01
—
—
—
0.0005
0.01
EFFICIENCY OF TREATMENT
IN REMOVAL OF POLLUTANTS
(% REMOVAL)
INCREASES
—
91%
—
—
-
27%
96%
-
—
93%
—
94%
—
67%
—
—
—
-
—
    Data obtained from sampling and analysis.
    Data obtained from plant monitoring records.
    Value in pH units.
                                 VII-73

                                 DRAFT

-------
                            DRAFT
tions approaching 100 percent for several heavy metals of
interest.

Various techniques are employed to augment the use of lime
neutralization.  Among these are secondary settling ponds,
clarifier tanks, or the addition of flocculating agents (such
as polyelectrolytes) to enhance removal of solids and sludge
before discharge.  Often, readjustment of the pH is necessary
after lime treatment.  This can be accomplished by addition
of sulfuric acid or by recarbonation.  The use of sulfide
precipitation may be necessary in some instances for further
removal of metals such as cadmium and mercury.

Mine Employing Hydrometallurgical Process.   Acid solutions
employed in dump, heap, and in-situ leaching are recycled in
this subcategory of the copper industry, allowing the recovery
of copper in the iron precipitation plant.  Water is added to
replace losses due to evaporation and seepage.  Acid is added
to control pH.  Table VII-9 lists the operations surveyed and
their control of acid solutions.  Only one operation surveyed
discharges a small amount of "bleed water" to surface waters.

Control of seepage and collection of acid-leach solution are
sometimes aided by the construction of specially prepared
surfaces, upon which heaped ores are placed for leaching.
These surfaces may be constructed of asphalt, concrete, or
plastic.

One facility currently bleeds the acid-leach solution and
treats the bleed by neutralization and precipitation with
alkaline (limed) tailings from the mill.  The treated water
flows into the tailing pond for settling and is subsequently
recycled with the decant water to the mill.

Treatment of the leach solutions used in this subcategory is
sometimes necessary for control of dissolved solids, which
build up during recycling.  Holding ponds are constructed to
retain leach solutions for a sufficient time to allow the
iron salts to precipitate from solution and settle, before
the solution is recycled to leach beds.  In conjunction
with, or in place of holding ponds, pH control aids in pre-
venting iron salts from precipitating in pipes or In the
leach dump.
                             VII-74

                             DRAFT

-------
                                DRAFT
 TABLE VI1-9. DUMP, HEAP, AND IN-SITU LEACH-SOLUTION CONTROL
              AND TREATMENT PRACTICE (1973)
PLANT
2101
2102
2103
2110
2116
2118
2123
2107
2108
2122
2124
2125
2104
2120
CONTROL



Zero discharge


Zero discharge

Zero discharge
99.4% recycle
98.7% recycle
TREATMENT



Recycle without treatment


20% to evaporation ponds
•
All effluent circulated through
holding ponds or reservoirs
None
Bleed is limed and settled in
tailing pond
DISCHARGE



None


None

None
654m3/day(avg)*
2551 m3/day (avg)"
to tailing pond (not
discharged)
 •Inadequate pumps. Operation required to attain zero discharge by State Regulations in 1977.
••The treated bleed is recycled to the mill with the decant.
                                VII-75


                               DRAFT

-------
                              DRAFT
Evaporation ponds are also employed to accomplish zero dis-
charge of acid-leach bleed solutions.

Mill Employing Vat Leaching for Extraction.   Zero discharge
has been reached by all facilities studied (Table VII-10).
Makeup water is required to replace evaporative losses and
the moisture which remains in the discarded, leached ores.

Complete recycling of barren leach and wash solutions is
usually practiced.  However, one facility presently reuses
its spent vat-leach solution in a smelter process to achieve
zero discharge.

Mill Employing Concentration by_ Froth Flotation.   Mills
employing froth flotation constitute two subcategories of the
copper-ore mining and dressing Industry.  The two subcategories
are divided on the basis of climatic conditions as: (1) mills
located in areas where net evaporation is less than 76.2 cm
(30 in.); and (2) mills located in areas where net evaporation
equals or exceeds 76.2 cm (30 in.).  All facilities currently in
operation in subcategory (2) discharge no wastewater effluent.

Process water from froth flotation contains large amounts of
suspended solids, which are normally directed to a large lagoon
to effect settling of these solids.  Surface runoff, such as
that resulting from snow melt, heavy-rainfall events, streams,
and drainage, should be conveyed around the tailing pond, thus
preventing runoff water from contacting process effluents.
In this manner, the volume of water which must be treated or
impounded is reduced.

Mill tailing-pond water may be decanted after sufficient
retention time.  One alternative to discharge, and an aid to
reducing the amount of effluent, is to reuse the water in
other facilities as either makeup water or full process water.
Usually, some treatment is required for reuse of this decanted
water.  Figure VII-9 illustrates the control of effluent by
reuse, as practiced at mill 2124.

The volume of water to be treated in flotation mills can be
effectively reduced, and the quality of the discharge often
                               VII-76

                               DRAFT

-------
                            DRAFT
TABLE VI1-10. SOLUTION-CONTROL PRACTICE IN VAT LEACHING OF COPPER ORE
MILL
2102
2116
2124
2126
CONTROL
100% recycle
100% recycle
100% recycle
Zero discharge
RECYCLE TREATMENT
None
None
None
Spent acid sent to acid plant for
• reuse
                            VII-77





                            DRAFT

-------
                            DRAFT
Figure VI1-9. CONTROL OF EFFLUENT THROUGH REUSE OF MILL FLOTATION-
          PROCESS WATER IN OTHER FACILITIES (MINE/MILL 2124)
                            VII-78

                            DRAFT

-------
                               DRAFT
substantially improved,  by the separation of mine water,  sewage,
smelter drainage, refinery wastes,  and leach bleed solution
from the tailing-pond circuit.  It  has been observed that
separation of mine water, with subsequent treatment and dis-
charge of the mine water only, can  allow mill tailing decant
water to be recycled fully.  As a result, lower total pollu-
tant loads may be discharged to the environment.  Using mine/mill
2121 as an example, Figure VII-10 was constructed to illustrate
current practice, as well as alternative future practice which
would result in a reduction of the  waste loads discharged.

Separation of mine water and other  wastes from contact with
mill process water is suggested in  all cases where pollutant
load and water volume are factors.   Not only do these waste
waters contribute to the pollutants present in the tailing-
pond water, but they may dilute the water to be treated or cause
excess water-volume conditions to result which cannot be
handled by recycling.

If sewage plant overflow contributes to the tailing-pond water
volume to the extent that it cannot be accommodated in recycling,
this water should be properly treated and handled separately.

Smelter and refinery wastes often contribute a heavy load of
dissolved metals to tailing ponds.   These wastes can affect
the quality of the decant water, as well as effluent volumes.
It may be necessary to handle wastes from these sources separately,
and/or as recommended under the appropriate conditions for the
Effluent Limitation Guidelines for the Copper Smelting and
Refining Industry.

The most efficient control of the volume and pollutant dis-
charge of mill flotation-process water is to recycle the
excess water which would overflow from the tailing-pond decant
area.  Of the 27 major copper mills surveyed, 24 are known to
be recycling all or a portion of their process water.  The
impetus for recycling has often been the lack of an adequate
water supply.  However, the feasibility of recycling process
water appears to have been considered at all facilities.

Through the use of diversion ditching, evaporation (when
available), reservoirs, and separation of other process water,
the volume of water to be recycled can be adjusted to allow
reuse.  Treatment of the recycled water is usually required
and may include secondary settling, phosphate or lime addition
(for softening), pH adjustment, or aeration.
                               VII-79

                               DRAFT

-------
                                  DRAFT
Figure VII-10. REDUCTION IN WASTE POLLUTANT LOAD IN DISCHARGE BY SEPARATION
           OF MINEWATER FROM TAILING  POND FOR SEPARATE TREATMENT
           (MILL 2121)
CURRENT












OTHER
WASTES

MILL
PROCESS
WATER
(LIMED)




1 MINE 1
1 '
EFFLUENT
1
- 1 1
^ X^~TAILIr\H5X
^v.^ rotto^J
IT

1 EFFLUENT \(t

1 1^-

TOTAL WASTE LOAD DISCHARGED
Par 24 noun in kg/day (Ib/day)



Flaw



pH
TSS
Oil
Cu
A.
Zn
Fa
Cd
Ni
Hg
Pb
and Greata





















LIME I



x
)

AT®

102.000 m3/day (27.000.000 gpd)

8.4'
020 (1,384)
41S (913)
27 (59.4)
<8 (<17.6)
6.2 (11.41
103 (227)
< 2 « 4 4)
<5.2 K11.4)
<001 (<0022)
C103 (<227)












ALTERNATIVE
1 1







1 	 »H MILL
-Tj^j
PROCESS
WATER
X 	 POND_ 	 /
RECYCLE T
1


1 MINE 1
1 '
T
LIME
TREATMENT
t
SETTLING |
{

DISCHARGE







K§)


ESTIMATED TOTAL WASTE LOAD DISCHARGED, USING LIME
PRECIPITATION/ AT ®
Par 24 noun In kg/day (Ib/day)

Flaw

pH
TSS
Oil and Gresu
Cu
At
Zn
Fa
Cd
Ni
Hg
Pb
Raw (No Trutmant)
3,800 m3/day

(1.000,000 gpd)
7.4'
267 (587)
<4 K8.8)
4(8.8)
<03 «066)
108 (23.8)
<04 «0.88)
<007 1< 0.1 541
<0.2 «044)









< 0.0005 « 0.001 10)
<0.4 «088)

After Treatment
3.800 m3/day
(1,000,000 gpd)
12.7'

129 (284)
<4 «8.8)
0.2 (0.441
<0.3 «066)
0.4 (038)
<04 (<0.88)
<002
« 0.044)
<0.2 I < 0.44)
0.0004 (040088)
<0.4 (
<0£8)
    Value in pH units.
                                 VII-80

                                  DRAFT

-------
                                DRAFT
The majority of copper mills currently operating recycle their
mill process water.  Of the remaining facilities that currently
discharge, half are recycling at least 35 percent of their
process water.  Treatment of discharged water consists of
settling alkaline wastewater in a tailing pond.  A variety of
treatment approaches are currently used in this subcategory,
including:

     (1)  Settling Only
     (2)  Lime Precipitation and Settling
     (3)  Lime Precipitation, Settling, Use of Polyelectrolytes,
               and Secondary Settling

One operation is currently building a treatment facility which
will include lime precipitation, settling, and aeration.

Table VII-11 shows the reduction of pollutant concentrations
attained in six mills under different conditions of recycling,
lime addition, and settling.  A wide variation in practice
is used to obtain varying degrees of concentration for waste
constituents present in treated wastewater.

Additional treatment of wastewater by polyelectrolyte addition,
to reduce suspended solids in tailing-pond discharge, is also
practiced at one mill.  Secondary settling ponds are used
to settle the treated solids prior to discharge.

The effectiveness of the use of coagulants (polymers) is
demonstrated in Table VII-12.  These data, obtained from a
pilot operation, indicate effective reductions of copper,
iron, and cobalt, with substantial reductions of aluminum
and manganese.

Recycling of process water from the tailing pond has not been
difficult for most copper mills surveyed employing this tech-
nique.  However, treatment of the pond water has been necessary
for selected problems encountered.  Potential problem areas
present at these operations include buildup of scale deposits,
pH changes in the tailing pond or in makeup water, and presence
of flotation reagents in the recycled water.  Effective
methods of treatment to alleviate these conditions are phos-
phate treatment (softening) for scale control, adjustment of
pH by liming, and the use of aeration or secondary settling
ponds to assist in degradation of flotation reagents.
                                VII-81

                                DRAFT

-------
                            DRAFT
  TABLE VII-11. REDUCTION OF POLLUTANTS IN CONCENTRATOR TAILS
             BY SETTLING AT VARIOUS pH LEVELS
PARAMETER
PH
TSS
Al
Ai
Cd
Cr
Cu
Fa
Pb
Mn
Ho
Ni
Se
Zn
Sb
Co
Mo
Comments:

PARAMETER
pH
TSS
Al
At
Cd
Cr
Cu
Fa
Pb
Mn
Hg
Ni
SB
Zn
Sb
Co
Mo
Conwitsfits*
CONCENTRATION (mg/ 1)
MILL 21 19
BEFORE
SETTLING
11.6-
705,000
< 1.0
< 0.07
< 0.05
< 0.05
0.1 B
0.8
< 0.5
< 0.05
0.0002
< 0.1
0.02
< 0.05
< 0.2
< 0.05
< 0.2
AFTER
SETTLING
7.7*
10
< IX)
< 0.07
< 0.05
< 0.05
0.05
0X18
< 0*5
0.3
< 0.0001
<0.1
0.08
< 0.05
< 0.2
< 0.05
<0.2
lima added after mill
water recycled
MILL 2120
BEFORE
SETTLING
11.1"
282.000
1.6
0.6
< 0.02
< 0X15
0.8
52
< 0.1
OX»7
0.0008
< 0.05
—
0.1
< 0.5
< 0.04
< 0.5
AFTER
SETTLING
9j6*
8
<0£
< OX>7
< OXW6
< 0X15
048
< 0.1
<0.1
OX>3
0X1011
< 0X15
0X14
< 0X15
< OS
< 0.04
< OS
lime added after mill
water recycled
MILL 2121
BEFORE
SETTLING
10J»
166X»0
lOXi
< OJ07
< 0X12
< 0.06
3.5
18.5
02
0.35
0X1098
< 0X15
0X12
0.9
< OX>
< OXM
< OS
AFTER
SETTLING
8.4*
6
<0.5
< 0.07
< 0.02
< OXS
0.3
<0.1
<0.1
OXM
< 0X1001
6
OXW9
< 0.05
0.02
< 0.05

-------
                             DRAFT
   TABLE VII-12. EFFICIENCY OF COAGULATION TREATMENT TO REDUCE
               POLLUTANT LOADS IN COMBINED WASTE (INCLUDING
               MILL WASTE) PRIOR TO DISCHARGE (PILOT PLANT)
POLLUTANT
PARAMETER
Flow
pH
TDS
TSS
Al
Ai
Cd
Cu
Fo
Pb
Mn
Hg
Ni
Co
Zn
WASTE LOAD IN INFLUENT TO PROCESS
kg/1000 metric toni
76.134 m3/day
7.6«
3.600
10
2.3
0.2
<0.05
B.8
120
3.3
0.4
a 0001
< ai
as
<0.05
Ib/IOOOgal
ie.860,400gpd
7.6»
6
0.02
0.004
0.0003
< 000009
0.02
0.21
0.008
0.0007
0.0000001
< 0.0002
0.02
< 0.00009
WASTE LOAD IN EFFLUENT TO DISCHARGE
kg/1000 metric tons
76.198 m3/dev
9.0*
3.900
14
< 1
0.9
< 0.05
0.9
0.7
2.8
0.1
0.0003
<0.1
0.9
57%
-
-
90%
>99%
15%
71%
-
-
90%
-
•Value in pH units
                             VII-83
                             DRAFT

-------
                               DRAFT
Lead and Zinc Ores

A discussion of the treatment and control technologies currently
employed in the lead and zinc ore mining and dressing industry
is included in this section.  Three subcategories are repre-
sented:  Mines with alkaline drainage not exhibiting solubili-
zatlon of waste constituents, mines with acid or alkaline drainage
exhibiting extensive solubilization of metals, and lead and/or
zinc mills.

Mines With Alkaline Drainage Not Exhibiting Solubilization of
Metals.  The operations represented by this subcategory gen-
erally employ treatment by impoundment in tailing or sedimen-
tation ponds.  Mine 3105 (producing lead/zinc/copper concentrates)
is located in Missouri.   The mine recovers galena
(FbS) , sphalerite  (ZnS), and chalcopyrite (CuFeS).  Production
began in 1973, and the operation was expected to produce 997,700
metric tons (1,100,000 short tons) of ore in 1974.

The water pumped from this mine is treated by sedimentation
in an 11.7-hectare (29-acre) pond.  The average mine drainage
flow rate is 8,300 cubic meters (2,190,000 gallons) per day.
The effluent from this basin flows to a nearby surface stream.
The chemical characteristics of the wastewater before and
after treatment are presented in Table VII-13, together with
data for nine months of 1974.  The treatment sequence is as
follows:  mine pumping, followed by clarification basin, followed
by discharge (8,300 cubic meters (2,190,000 gallons) per day).
Relatively simple treatment employed for mine waters exhibiting
chemical characteristics similar to mine 3105 can result in
attainment of low discharge levels for most constituents.
Reduction of parameters such as total dissolved solids, oil
and grease, chloride, sulfate, lead, and zinc—as well as excel-
lent reduction of total suspended-solid concentrations—is obtained
by this treatment method.

Mine Drainage (Acid or Alkaline) Exhibiting Extensive Solubili-
zation £f_ Metals.  The characteristics of wastewater from
mines in this subcategory are such that treatment must be
applied to prevent the discharge of soluble metals, as well
as suspended solids.  The treatment practice, as currently
                               VII-84

                               DRAFT

-------
                                 DRAFT
TABLE VII-13. CHEMICAL COMPOSITIONS OF RAW AND TREATED MINEWATERS
            FROM MINE 3105 (HISTORICAL DATA PRESENTED FOR COMPARISON)
PARAMETER
pH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
Ammonia
Hg
Pb
Zn
Cu
Cd
Cr
Mn
Fe
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£)
RAW MINE
DRAINAGE*
7.4»*
196.0
330.4
138
326
<10
< 1.0
29.0
0.030
<0.05
0.0001
0.3
0.03
<0.02
< 0.002
<0.02
<0.02
<0.02
63.5
57
1.2
DISCHARGE*
8.1 **
16^0
173.2
< 2
204
<10
3.0
17.0
0.032
<0.05
< 0.0001
0.1
<0.02
<0.02
0.005
<0.02
0.35
0.11
45.5
44.5
1.0
DISCHARGE (HISTORICAL)*
AVERAGE
7.8**
—
-
3.4
-
—
-
1.9
-
-
-
0.050
0.032
< 0.005
< 0.005
-
-
0.086
-
-
-
RANGE
7.4**to8.1**
—
-
<1 to 9
-
-
-
< 1 to 5
-
-
-
0.011 to 0.01 2
0.008 to 0.11
<0.050 to 0.070
(< 0.005)
—
-
0.033 to 0.21
-
-
-
     •Analysis of single 4-hour composite sample
     tMonthly analysis over January 1974 through September 1974
    ••Value in pH units
                                  VII-85

                                  DRAFT

-------
                             DRAFT
employed,  involves  chemical  (often,  lime)  precipitation and
sedimentation.

Mine wastewaters  in this subcategory are often  treated by
discharge  into  a  pond  or basin in which the  pH  is controlled.
An approach often used is to discharge the mine wastewater
into a mill tailing pond, where wastewater is treated at a
pH range which  causes  the precipitation of the  heavy metals
as insoluble hydroxides.   The presence of  residual solids from
the milling process is thought to provide  nucleation sites
for the precipitation  of the hydroxides.   In cases where
ferrous iron is present,  it  is desirable to  cause the oxidation
to the ferric form  and,  thus,  to  avoid the potential for acid
formation  by processes similar to the reactions forming acid
mine drainage.  Vigorous aeration of the wastewater can accom-
plish oxidation,  usually after addition of the  pH-adjusting
agent.

The treatment process  described is similar to the type of pH
control, and subsequent  physical  treatment,  usually associated
with froth-flotation recovery of  sulfides  of lead, zinc, and
copper (which is  followed by sedimentation of the tailings).
The milling process itself is,  therefore,  an analog for a
process of treating mine  wastes in this subcategory.

Mine 3107  is an underground  lead/zinc mine located in Idaho.
Galena and sphalerite  are mined,  with approximately 544,200
metric tons (600,000 short tons)  of  ore mined per year.  The
mine has been in  operation most of this century.

Mine water pumped from lower levels  of the mine, as well as
water from upper  levels  (which flows  by gravity), exits the
mine tunnel and is  piped  to  a  central impoundment, 48.5 hectares
(120 acres) in  area.   The average mine flow  is  16,500 cubic
meters (4,360,000 gallons) per day.  Waste streams, including
the tailings from the  concentrator,  a smelter,  and an electrolytic
zinc plant, also  flow  to  the central  impoundment area.   The
overflow from this  impoundment  area,  29,000  cubic meters
(7,700,000 gallons)  per  day,  is treated in a high-density,
sludge-type chemical-precipitation plant.  The  characteristics
of the raw mine waste, the overflow  from the central impound-
ment area, and  the  final  effluent from the treatment process
are presented in  Table VII-14.
                             VII-86

                             DRAFT

-------
                                  DRAFT
TABLE VII-14. CHEMICAL COMPOSITIONS OF RAW AND TREATED WASTEWATERS
             FROM MINE 3107 (HISTORICAL DATA PRESENTED FOR COMPARISON)
PARAMETER
pH
Alkalinity
Hardness
TSS
TDS
COD
TOC
Oil and Grease
P
Ammonia
Mercury
Lead
Zinc
Copper
Cadmium
Chromium
Manganese
Iron
Sulfate
Chloride
Fluoride
CONCENTRATION (mg/£)
RAW
MINE WATER
3.2'
14.6
671
<2
1.722
47.6
2.3
3.0
<0.02
1.8
0.0001
0.3
38.0
0.04
0.055
0.17
57.2
2.5
750
<001
0.063
OVERFLOW FROM
CENTRAL POND
2.0t
0.0
2.356
<2
2.254
39.7
4.3
<1
0.08
1.6
0.0468
3.1
180.0
0.52
140
0.67
41.0
59.0
1.862
1.2
19
TREATED
EFFLUENT
8.5*
3.2
1.242
<2
2.030
43.6
4.0
17
<0.02
0.80
0.0007
<0.1
5.1
0.04
0.048
0.50
0.32
0.85
1.744
15
21
HISTORICAL DATA
AVERAGE
7.4*
-
-
-
-
-
-
-
-
-
0.002
0.093
1.43
0.020
0.044
_
-
-
-
-
-
RANGE
6.9* to 7.6*
-
-
-
-
-
-
-
-
-
< 0.001 to 0.005
0.057 to 0.153
0.79 to 2.08
0.010 to 0.043
0.032 to 0.058
_
-
_
-
-
-
    Average foi months includes 10-24 hour composite samples.
    Value in pH units.
                                  VII-87


                                   DRAFT

-------
                             DRAFT
The treatment process is shown schematically in Figure VII-11.
Provision has been made for pumping the recovered sludge back
to the mill, should recovery of metal values prove practical.
At present, the sludge is disposed of at a solid-waste disposal
site.

Mine 3101 is an underground mine, located in Maine.  The
mine recovers sphalerite and the byproducts chalcopyrite,
galena, and pyrite which are present in the formation.
The mine began production 1972 and produced 208,610 metric
tons (230,000 short tons) of ore in 1973.

The water pumped from the mine, 950 cubic meters (250,000
gallons) per day, is treated by mixing it with mill tailing
discharge, plus additional lime as required for pH control,
in a reservoir with a capacity of 37.85 cubic meters (10,000
gallons).  The combined waste is then pumped to a 25-hectare
(62-acre) tailing pond.  The discharge from the tailing pond
is sent to an auxiliary pond.  The combined retention time in
the two ponds is 35 days at maximum flow.  Water is recycled
for the process from the auxiliary pond, and the excess is
discharged.  The chemical characteristics of the mine water
and the final discharge, treated in the above manner, are given
in Table VII-15.

Lead and/or Zinc Mills.  As discussed in Section V, the
wastewater from lead/zinc flotation mills differs from mine water
in that a number of reagents are added to effect the separation
of the desired mineral or minerals from the host rock.  These
waste streams also contain finely ground rock, as well as
minerals, as a result of grinding to allow liberation of the
desired minerals during the froth-flotation process.

The most common treatment method in use in the lead/zinc-
milling industry is the tailing or sedimentation pond.  Often
considered a simple method of treatment, properly designed
tailing ponds perform a number of important functions simultan-
eously.  Some of these functions include removal of tailing
solids by sedimentation, formation of metal precipitates,
long-term retention of settled tailings and precipitates,
stabilization of oxidizable constituents, and balancing of
influent-water quality and quantity of flow.
                             VII-88

                             DRAFT

-------
                                     DRAFT
Figure VII-11. SCHEMATIC DIAGRAM OF TREATMENT FACILITIES AT MINE 3107
    22.7 m /mm
   (6,000 gpm) max
            49-hectare
            (120-acre)
            CENTRAL
              POND
                                              56.8-m3 (15,000-gal)
                                             RECYCLE MIX TANK
                                            (5-minute retention time)

                                                                   RECYCLE
                                                                   SLUDGE
            1,442-m
           (381,000-gal)
         AERATION BASIN
       (41-minute retention time)
  333-m
FLOCCULATION
                                                  16,653-m
                                                 (4,400,000-gal)
                                                 THICKENER
                                              (8-hour retention time)
                                                                    SLUDGE
                                                                    BLEED
                                                   MILL
                                                RESERVOIR
                                    VII-89


                                     DRAFT

-------
                       DRAFT
TABLE VIMS. CHEMICAL COMPOSITIONS OF RAW AND
             TREATED MINE WATERS FROM MINE 3101
PARAMETER
pH
TSS
TDS
COD
Pb
Zn
Cu
Cd
Cr
Mn
Fe
CONCENTRATION (mg/£>*
RAW MINE
WATER
6.9t
-
—
-
0.035
2.608
0.012
0.004
< 0.010
0.996
0.359
TREATED
DISCHARGE
8.7*
7.2
595
25
< 0.024
0.096
0.016
< 0.002
< 0.010
0.055
0.303
         •Average for year of 1974 as reported for NPDES permit

         *Value in pH units
                       VII-90


                       DRAFT

-------
                             DRAFT
In the lead/zinc-ore milling industry, a biological treatment
method, used in conjunction with stream meanders, was observed
at one location.  This treatment method has been described in
the previous discussion in this section.

The ability to recycle the water in lead/zinc flotation mills
is affected by the buildup of complex chemical compounds
(which may hinder extraction metallurgy) and sulfates (which
may cause operating problems associated with gypsum deposits).
One solution to these problems is a cascade pond system.
There, the reclaimed water from thickeners, filters, and
tailing ponds may be matched with the requirements for each
point of the circuit (Reference 66).

In another study (Reference 67), the many operational
problems associated with the recycling of mill water are
described in detail.  The researchers have observed that
recycling at the operations studied had not caused any
unsolvable metallurgical problems and, in fact, indicate that
there are some economic benefits to be gained through decreasing
the amounts of flotation reagents required.

Mill 3103 is located in Missouri and recovered galena,
sphalerite, and chalcopyrite from 846,000 metric tons (934,000
short tons) of ore in 1973.

The mill utilizes both mine water and water recycled from the
tailing pond as feed water.  The concentrator discharges
9,500 cubic meters (215,000,000 gallons) per day of tailing
slurry to its treatment facility.  The treatment facility
utilizes a 42.5-hectare (105-acre) tailing pond with esti-
mated retention of 72 days, a small stilling pond at the
base of the tailing-pond dam, and a shallow 6.1-hectare (15-
acre) polishing pond before discharge.  A schematic diagram of
average daily water flows for the facility is given in Figure
VII-12.  Effluent chemical composition and waste load discharged
at mill 3103 using the above treatment are given in Table VII-16.

Mill 3102 is located in Missouri.  This mill processed
approximately 1,450,000 metric tons (1,600,000 short tons)
of ore in 1973.  Galena and sphalerite are recovered as concen-
trates at this operation.
                             VII-91

                             DRAFT

-------
                                    DRAFT
   Figure VII-12. SCHEMATIC DIAGRAM OF WATER FLOWS AND TREATMENT
                FACILITIES AT MILL 3103
                                 7,570 m3/day
                                 (2,000,000 gpd)
         RECYCLE
         WATER
  3,785 m3/day
(1,000,000 gpd)
 EVAPORATION
      AND
    SEEPAGE
      I
                       15,150-m3 (4,000,000-gal)
                           RESERVOIR
          •WATER-
       est 1,160 m3/day
       test 300,000 gpd)
TO
SMELTER
MILL
i


9,500 m3/day
(2,500.000 gpd)
i
CONCENTRATES
37.9 m3/day
(10,000 gpd)
                             TO
                             STOCKPILE
                                             1,515 m3/day
                                             (400,000 gpd)
                        ( STILLING POND
DND)
                                 10.100 m3/day
                                 (2.600.000 gpd)
                            DISCHARGE
                                                   est 3,785 m3/day
                                                   (est 1,000.000 gpd)
                                    VII-92


                                    DRAFT

-------
                             TABLE VII-16. CHEMICAL COMPOSITIONS AND WASTE LOADS FOR RAW AND
                                           TREATED MILL WASTE WATERS AT MILL 3103
PARAMETER
PH
TSS
COD
Oil and Gr«M
Cyanide
Ho
Pb
Zn
Cu
Cd
Cr
Mn
Total Fa
MILL RAWWASTEWATER
CONCENTRATION (mg/fj
THIS PROGRAM*
7.9"
464.000
til
3.0
<0.01
< 0.0001
03
0.12
0.36
0.011
<0.02
003
0.05
HISTORICAL*
7.9* •
1.7
-
-
-
-
0.107
0.288
0014
0.001
0.002
0.169
0.03
RAW WASTE LOAD par unit me millad
kg/1000 matnc tana
1490.000"
400
12
< 0.024
0.00024
0.480
0.288
0.865
0.026
0.048
0.072
0.12
lb/1000 chart lam
2.180.000*'
800
24
< 0.048
< 0.00048
0.860
0.576
1.730
0.052
< 04)96
0.144
0.24
FINAL TREATED DISCHARGE
CONCENTRATION (mg/l)
THIS PROGRAM*
13"
16
726
30
<0.01
< 00001
0.1
0.07
<0.02
< 0.002
<0.02
0.05
0.09
HISTORICAL*
J3"
1.4
-
-
N/A
-
0.028
04)45
0.006
< 0.001
0.001
0074
0.032
EFFLUENT WASTE LOAD par unit on mllad
kg/1000 matric torn
40
1.700
7
0024
000024
0.24
0.168
< 04)48
< 0.005
< 0.048
0.12
0.282
lb/1000 abort tons
80
3,400
14
0.048
< 040048
048
0336

-------
                              DRAFT
The mill utilizes mine water exclusively as feed.  It discharges
15,150 cubic meters  (4,000,000 gallons) per day of tailing slurry
to a large tailing pond.  This pond also receives about 3,785
cubic meters (1,000,000 gallons) per day of excess mine water
and another 3,785 cubic meters (1,000,000 gallons) per day of
surface-drainage water.  This tailing pond presently occupies
32.4 hectares  (80 acres) and will occupy 162 hectares (400 acres)
when completed to design.  The tailing-pond decant water is
discharged to a small stilling pool and then enters a meander
system, where biological treatment occurs.  An additional
sedimentation basin of approximately 6.1 hectares (15 acres),
for removal by sedimentation of any algae which breaks loose
from the meander system, has been constructed near the end of
the meander system for use Just before final discharge.  A
schematic diagram of the mill operation and the treatment facility
is presented in Figure VII-13.

Water characteristics for the effluent from the mill, the overflow
from the tailing pond, and the final discharge treated utilizing
the above technology are presented in Table VII-17.

Mill 3105 is located in Missouri and recovered galena, spha-
lerite, and chalcopyrite from an estimated 997,000 metric
tons (1,100,000 short tons) of ore in 1974.

This mill utilizes water recycled from its tailing-pond system
and makeup water from its mine as feed water.  The mill dis-
charges 7,910  cubic meters (2,090,000 gallons) per day of wastes
to a 11,8-hectare (29-acre) tailing pond.  The decant from this
pond is pumped to an 7.3-hectare (18-acre) reservoir, which also
receives the required makeup water from the mine.  The mill
draws all its  feed water from this reservoir.  No discharge
occurs from the mill.

A schematic diagram  of the water flows and treatment facilities
is presented in Figure VII-14.

Mill 3101 is located in Maine and recovered sphalerite and
chalcopyrite from 208,000 metric tons  (230,000 short tons)
of ore in 1973.

This mill utilizes only water recycled from its treatment
facilities as  feed water.  The mill discharges to a mixing tank,
                              VII-94

                              DRAFT

-------
                             DRAFT
Figure VII-13. SCHEMATIC DIAGRAM OF WATER FLOW AND TREATMENT
            FACILITIES AT MILL 3102 (TAILING POND/STILLING POND/
            BIOLOGICAL TREATMENT/POLISHING POND)
 TO ATMOSPHERE
                            15.150 m3/dav
                            14,000,000 gpd)
                            22,300 m3/day
                            (5.900,000 gpd}
I
34.100 m
<9,000,00
r
DISCHARGE
                                         7,560 mj/day
                                         (2,000,000 gpd)
                                        9,100ms/day
                                        (2,400.000 gpd)
                                                      NATURAL
                                                       SPRING
                              VII-95


                              DRAFT

-------
                                   DRAFT
  TABLE VI1-17. CHEMICAL COMPOSITION AND WASTE LOADING FOR RAW AND
               TREATED MILL WASTEWATER MILL 3102
PARAMETER
pH
TSS
COD
Oil ind Orana
Cyanide
MB
Pb
Zn
Cu
Cd
Cr
Mn
Total ft

PARAMETER
pH
TSS
COO
Oil and GIHM
Cyanida
H«
Pb
Zn
Cu
Cd
Ci
Mn
Total Fa
MILL RAW WASTEWATER
CONCENTRATION
(mg/ei*
8.8"
248.000
488.1
0
0.03
< 00001
1.9
046
<0.02
0.006
<0.02
0.08
053
RAW WASTE LOAD
par unit era mllM
kg/1000 mettle lorn
_
900.000
1.400
0
04187
< 0.0003
6.5
1.33
< 0.0068
0.014
< 0.0068
OJ32
164
lb/1000 riion torn
_
1,800.000
2,800
0
0174
< 00006
11
266
< 0X1116
0028
<00116
0464
308
TAILING-POND DECANT
CONCENTRATION
Img/ll'
7.8"
18
663.8
8X1
1
-
-
0.003
WASTE LOAD
par unit ora millad
kg/1000 matnc tana
-
86
88
26
0082
< 0.0003
0.26
0.1

-------
                                   DRAFT
  Figure VM-14. SCHEMATIC DIAGRAM OF WATER FLOW AND TREATMENT
               FACILITIES AT MILL 3105
           MINE
10,900 m3/day
(2,880,000 gpd)
 8.300 m3/day
(2,190,000 gpd)
                  2,615 m3/day
                  (690,000 gpd)
        7.3-hectare
         (18-acre)
        RESERVOIR
       MINE-WATER
       TREATMENT
  8,300 m3/day
 (2,190,000 gpd)
        DISCHARGE
 7,900 m3/day
(2,090,000 gpd)
                                          MILL
                                           I
                                       TAILINGS
                                      (35% SOLIDS)
                                5.510 m3/day
                               (1.460,000 gpd)
                                7,900 m3/day
                               (2.090,000 gpd)
                        2,380 m3/day
                       (630,000 gpd)
                                        RECYCLE

                                5.300 m3/day
                                (1,400,000 gpd)|
                                   VII-97
                                   DRAFT

-------
                              DRAFT
where nine water Is treated by chemical precipitation that
is achieved by combustion with the tailing slurry and liming
as required.  This combined waste is introduced into a tailing
pond, which discharges to an auxiliary pond.  The combined
retention time in the two ponds is 35 days at maximum flow.
A schematic diagram of the mill-water circuit is shown in
Figure VII-15.  The separate treatment of mine water and surface
runoff would allow this operation to achieve total recycle.
Discharge data for this mine/mill complex were presented as
mine discharge for mine 3101 earlier in this section.

Gold Ores

The discussion that follows describes treatment and control
technology in current use in the gold-ore mining and dressing
industry.  Aspects of treatment and control which are unique
to the gold-ore category are described, in addition.

Mining Operations.   Wastewater treatment at mining operations
in the gold-ore mining industry consists of three options as
currently practiced in the U.S.:  (1) Direct discharge without
treatment; (2) Incorporation of mine water into a mill process-
water circuit; and (3) Impoundment and discharge.  Impoundment
of mine water without discharge may be currently practiced at
locations in arid regions, due to evaporation.  Direct discharge
of mine waters with high suspended-solid content is one po-
tential hazard associated with direct discharge—particulary,
with respect to placer, dredging, or hydraulic mining opera-
tions.  Current best practice in this segment of the industry is
use of the dredge pond or a sedimentation basin for settling,
and the use of tailing gravel and sands for filtration of the
discharge stream.  Levels of suspended solids attained routinely
with this method can be approximately 30 mg/1, or less if an
adequate residence time for the wastewater in the impoundment
can be obtained.  Few, if any, placer mines currently provide
treatment for the wastes generated from their stripping and
sluicing operations.  Settling ponds, however, have been widely
shown to be effective in improving water quality by reducing
turbidity.

Techniques used for the control of suspended solids discharged
from placer mining operations, regardless of size, are not
being employed on a major scale at present.  The termination
of mining operations, even with treatment facilities, does
not eliminate water-quality degradation, however, because most
                              VII-98

                              DRAFT

-------
                                  DRAFT
Figure VII-15. SCHEMATIC DIAGRAM OF TREATMENT FACILITIES AT MILL 3101
     2.15 m3/min (569 gpm)
            0.19 m3/min
              (50 gpm)
   0.38 m3/min
    (100 gpm)
                                               0.03 m3/min (8 gpm)
                                 SHIPPED WITH
                                 CONCENTRATE
                                PROCESS WATER
                                      I
                                VACUUM PUMP
  2.01 m3/min (531 gpm)
                                         0.11 m3/min (30 gpm)
                                  ESTIMATED
                                 MINE WATER
                                                0.47 m3/min
                                                 (125 gpm)
                                 DRILL WATER
0.19 m /min
 (50 gpm)
                                   RUNOFF
                                  FROM RAIN
                COOLING WATER
                 AND UTILITIES
    NEUTRALIZATION
     AS REQUIRED
                        0.33 to 1.88 m3/min (87 to 497 gpm)
                        (MONTHLY AVERAGES)

                        0.99 m3/min (262 gpm)
                        (YEARLY AVERAGE)
                  DISCHARGE
            0.16 m3/min
            (42 gpm)
                                                      EVAPORATED
                                  VII-99


                                  DRAFT

-------
                               DRAFT
operations which use impoundment usually construct the settling
or tailing pond adjacent  to  the stream being worked.  With
erosion taking place continuously, these facilities are seldom
permanent.

Mining operations exploiting lode ores which discharge mine
water from open-pit or underground operations commonly dis-
charge directly to a receiving stream, provide process water
for a mill circuit, or discharge wastewater to a mill tailing
pond.  An example of the  effectiveness of settling on water
quality is discussed under Gold-Ore Milling Operations.  Mill
tailing ponds have demonstrated effective treatment, primarily
for suspended-solids removal, but secondarily for heavy-metal
removal.

Open-pit gold mining operations in arid regions often have
little or no mine discharge, whereas underground mines typically
discharge water from the  mines.

Milling Operations.   In-plant control techniques and processes
used by the gold milling  industry are processes which were
designed essentially for  reagent conservation.  These processes
are the reagent circuits  indicated in the process diagrams of
Figures III-9 and 111-10.

In the cyanidation process used at mills 4101, 4104, and 4105,
gold is precipitated from pregnant cyanide-leach solutions
with zinc dust.  The precipitate Is collected in a filter
press, and the weak, gold-barren cyanide solution which remains
is recycled back to the leaching circuit.  This solution may
be used as a final weak leach, or the solution may be returned
to its initial concentration with the addition of fresh cyanide
and used as a strong leach.  In these processes, recycling of
cyanide reagent effects an estimated 33- to 63-percent saving
of this reagent.  Loss of cyanide from the mill circuit is
primarily through retention  in the mill tailings.  Recycling
of cyanide reduces the quantity of cyanide used and also reduces
the amount of reagent present in effluent from discharging
mills.

In a similar manner, mercury is typically recycled in amalga-
mation processes.  Currently, amalgamation is practiced at
only one milling operation (mill 4102) .  This mill uses a
barrel amalgamation process  to recover gold.  At this mill,
                              VII-100

                              DRAFT

-------
                             DRAFT
the gold is separated from the amalgam in a high-pressure
press, and the mercury is returned to the amalgamator for
reuse.  Some mercury is lost from this circuit—primarily,
through retention in the mill tailings.

Ultimate recovery or removal of mercury from the waste stream
of a mill presents an extremely difficult task.  To do so
requires removing a small concentration of mercury, usually
from a large volume of water.  Advanced waste treatment methods,
such as ion exchange, might achieve as much as 99 percent removal,
but the expense for treating large volumes of water would be high.
Primarily as a result of this, and in light of recent stringent
regulation of mercury in effluents, the gold milling industry
has been, taking advantage of the process flexibility available
to it and has, for the most part, replaced amalgamation with
cyanidation processes for gold recovery.  This process flexi-
bility is the best control currently being practiced by the
industry for minimizing or eliminating mercury waste loading.

The primary wastes emanating from a gold mill are the slurried
ore solids.  For this reason, mill effluents are typically
treated  in tailing ponds, which are designed primarily to
provide  for the settling and collection of the suspended  solids
in the mill tailings.  In most cases,  these operations dis-
charge from tailing ponds, and the usual practice is to decant
the water  from the top of the pond at  a point where maximum
clarification has been attained.  In some  facilities, two or
more  ponds are connected in  series, and wastewater is decanted
from  one to another before  final discharge.

Although the  structure, design, and methods of ponding may
vary  somewhat  in accordance  with local topography and volume
of wastewater, the desired  goal is the same—to achieve
maximum  settling and  retention of solids.

To  illustrate the effectiveness of settling ponds  as  treatment
systems  in the gold-ore milling  industry,  the  discussion
which follows outlines an operation which  recovers gold  and
other metals  and  treats wastewater by use  of  a tailing pond.

Mill  4102  is  located  in  Colorado.  This mill  beneficiates
ore containing sulfides  of  lead,  zinc, and copper, in
addition to native  gold  and silver.   During 1973,  163,260
metric tons  (180,000 short  tons)  of  ore were  milled  to
produce lead/copper  and  zinc concentrates  by  flotation
and gold by amalgamation.
                               VII-101


                               DRAFT

-------
                               DRAFT
Makeup water for the mill circuit is drawn from a nearby creek.
This water is introduced into the grinding circuit for trans-
portation and flotation of the ore.  Prior to entering the
flotation circuit, the ground ore is jigged to produce a
gravity concentrate.  This concentrate contains most of the
gold, which is recovered by amalgamation.  After amalgama-
tion, the jig concentrate is fed into the flotation circuit,
because some lead is contained in the material.

Mill tailings are discharged to a tailing pond at a rate of
2,290 cubic meters (600,000 gallons) per day.  Decant from
this pond flows to a smaller polishing pond prior to final
discharge to a stream.  The tailing pond and the polishing
pond have a total area of 18.2 hectares (45 acres).

Table VII-18 presents the chemical composition of mill water
and raw and treated waste load for mill 4102, which practices
amalgamation for gold and froth flotation for sulfide minerals.
These data indicate that removal of selected metals is achieved
to a degree; however, the treatment is most efficient in the
removal of suspended solids.

Mill 4101 is located in Nevada.  This mill recovers gold
occurring as native gold in a siltstone host rock which
is mined from an open pit.  Schuetteite (HgS04^2HgO)
also occurs in the ore body, and mercury is recovered
as a byproduct during furnacing of the gold concen-
trate.  Ore milled during 1973 totaled 750,089 metric tons
(827,000 short tons).  This figure normally is 770,950 metric
tons (850,000 short tons) but was lower than usual due to a
20-day labor strike.

This mill employs complete recycle of the tailing-pond decant.
However, due to consumptive losses, some makeup water is re-
quired, and this water is pumped to the mill from a well.
Water is introduced into the grinding circuit for transportation
and processing of the ore by the agitation/cyanidation-leach
method.

Mill tailings are discharged at a rate of 2,305 cubic meters
(603,840 gallons) per day to a 37-hectare (92-acre) tailing
pond.  Approximately 1,227 cubic meters (321,500 gallons)
per day of tailing-pond decant are pumped back to the mill
                             VII-102

                               DRAFT

-------
                                 DRAFT
TABLE VII-18. WASTE COMPOSITIONS AND RAW AND TREATED WASTE LOADS
             ACHIEVED AT MILL 4102 BY TAILING-POND TREATMENT
PARAMETER
pH
TSS
COD
Oil ind OrMM
Cd
Cr
Cu
Toltl F.
Pb
Totil Mn
MS
Zn
MILL WASTEWATER
CONCENTRATION
(mo/ HI
9.1-'
496.000
1142
1
<002
<0.02
0.03
1.0
<01
626
00014
1.3
RAW WASTE LOAD
par unit or* milled
kg/1000 imtrie torn
-
2.871.000
66
6.8
<0.12
< 012
0.17
6
<0.6
48
0008
76
Ib/IOOOilwttom
-
5.742.000
132
116
<0.24
<024
0.34
12
<12
86
0016
1B.O
TAILING-POND EFFLUENT
CONCENTRATION
(me/11
10A*
4
2235
1
<0.02
0.06
1.2
1.8
<0.1
637
00011
0X15
TREATED WASTE LOAD
kg/1000 nntric ton
-
20
130
6
<01
OJ
7
8
<0.6
40
0.006
OJ
lb/1 000 ihort torn
-
40
260
12

-------
                              DRAFT
from a reclaim sump.  No discharge from this operation results.
Potential slime problems in the mill circuit are controlled
through adjustment of the pH to 11.7 and by use of Separan
flocculant in the circuit.

Table VII-19 gives the results of chemical analysis of mill
effluent and tailing-pond decant water after treatment.  No
waste loadings are given, since no discharge results.  Samples
were obtained from this facility to determine the effectiveness
of treatment, even though the mill has no discharge.  Note,
however, that this mill has an alkaline-chlorination unit avail-
able for use in cyanide destruction should emergency conditions
require a discharge.

Data from both mills indicate that dissolved heavy metals
are removed to some degree in the tailing pond, but more
effective technology is required for removal of these waste
constituents.  Although such technology is not currently used
in the gold mining and milling Industry, it is currently
available.  This technology also has special application to
mine discharges, as they usually contain relatively high
dissolved-metal loads.  This technology will also be appli-
cable to those situations where sufficient reduction of
metals and cyanide in tailing-pond effluents is not being
achieved.

Conventional treatment available for dissolved heavy metals
generally involves:

     (1)  Coagulation and sedimentation employing alum, iron
          salts, polyelectrolytes, and others.

     (2)  Precipitation with lime, soda ash, or sulfides.

These treatment technologies have been previously discussed
in this section.  Treatment by these methods is not normally
practiced in this industry category.  However, where metal
mining wastes are treated, the most common means used is to
discharge to a tailing pond, in which an alkaline pH is
maintained by lime or other reagents.  Heavy-metal ions are
precipitated at elevated pH; these ions are then settled out,
together with suspended solids, and maintained in tailing
ponds.
                               VII-104

                               DRAFT

-------
                          DRAFT
TABLE VII-19. CHEMICAL COMPOSITIONS OF MILL WASTE WATER AND
           TAILING-POND DECANT WATER AT MILL 4101 (NO
           RESULTANT DISCHARGE)
PARAMETER
pH
TSS
Turbidity (JTU)
TDS
COD
Oil and Grease
Cyanide
As
Cd
Cr
Cu
Total Fe
Pb
Total Mn
Hg
Zn
CONCENTRATION (mg/ £ )
MILLWASTEWATER
12.26*
545,000
6.70
4,536
43
<1
5.06
0.05
0.10
0.06
0.17
<0.5
< 0.1
0.02
-
3.1
TAILING-POND DECANT
11.29*
12
1.0
4,194
43
<1
5.50
0.04
0.02
0.03
0.13
< OS
< 0.1
0.90
0.152
2.5
    •Value in pH units.
                           VII-105

                           DRAFT

-------
                             DRAFT
Mercury presents a  special  problem  for control, due to its
potential for conversion  in the  environment to its highly
toxic methyl-mercury  form.   The  amalgamation process still
finds some use  in the gold  milling  industry, and, in addition,
this metal sometimes  occurs with gold in nature.  Although
mercury will precipitate  as the  hydroxide, the sulfide is
much more insoluble.   It  is expected that, where dissolved
mercury occurs  in mine or mill wastes, it will be treated for
removal by sulfide  addition.  This  reaction requires alkaline
conditions to prevent the loss of sulfide ion from solution
as H2S.  Theoretical  considerations of solubility product and
dissociation equilibria suggest  that, at a pH of 8 to 9,
mercury ion will be precipitated from solution to a concentra-
tion of less than 10  exp(-41) g/1.  In practice, it is not
likely that this level can  be achieved.  However, by optimi-
zing conditions for sulfide precipitation, mercury should be
removed to a concentration  of less  than 0.1 microgram/liter
(0.1 ppb).

The conditions under  which  lime  precipitation of heavy metals
is achieved must take into  consideration auxiliary factors.
As indicated, the most important  of these factors is pH.
The minimum solubility of each metal hydroxide occurs at a
specific pH; therefore, optimum  precipitation of particular
metals dictates regulation  and control of pH.  When more than
one metal is to be  precipitated,  the pH must necessarily be
compromised to obtain the maximum coprecipitation achievable
for the given metals.

Another factor which  must be considered is the oxidation
state(s) of the metal or  metals  to  be treated.  For example,
As(+5) is much more amenable to  chemical treatment than is
As(+3).  In addition,  cyano-metallic or organo-metallic com-
plexes are generally  much more difficult to remove by chemical
treatment than are  free metal ions.  Where these factors
impede chemical treatment,  prior  oxidation of the waste stream
can be employed to  destroy  the metal complexes and oxidize
metal ions to a form  more amenable  to chemical treatment.
This oxidation may  be achieved by aeration of the waste stream
or by the addition  of chlorine or ozone.

To achieve high clarification and removal of solids and chemi-
cally treated metals,  it  is essential to provide good sedimenta-
tion conditions in  the  tailing pond.  Typically,  this is done
in the industry by  designing tailing ponds to provide ade-
                            VII-106

                            DRAFT

-------
                             DRAFT
quate retention time for the settling of solids and metal
precipitates.  Specification of a recommended retention time
for traditional tailing-pond design is problematical, because
the influence of pond geometry, inlet/outlet details, and other
factors that ensure even distribution and an absence of
short-circuiting are of greater importance than the theoretical
retention provided.  A design retention time of 30 days, based
on the average flow to be treated, is often specified and is
appropriate if short-circuiting due to turbulence or stratifi-
cation does not occur.  The use of a two-cell pond is recom-
mended to increase control and reliability of the sedimentation
process.

In some cases, suspended solids or metal precipitates may
retain surface charges or colloidal properties and resist
settling.  These solids and colloids can be treated for
removal by the addition of coagulating agents, which either
flocculate or act to neutralize or Insulate surface charges
and cause the suspended solids and colloids to coagulate and
settle.  These agents may be such flocculants as alum
(A12_(S04_)j3) or iron salts, or such coagulants as clays,
silica, or polyelectrolytes.

Cyanide destruction has been previously discussed in this
section.  The technology for oxidation and destruction of
cyanide is well-known and currently available.  Where dis-
charges of cyanide have the potential to enter the environment,
complete destruction prior to discharge is recommended.

Technology For Achieving Np_ Discharge £f_ Pollutants.  Elimin-
ation of point discharges is currently being achieved in the
industry by two slightly different technologies:  impoundment
and recycle.  Where impoundment is used, the mill tailings
are simply discharged to a pond and retained there.  Recycling
exists where tailing-pond water is decanted and returned to
the mill for reuse.  A mill or mine/mill complex is potentially
capable of employing either of these technologies, whereas a
mine alone may only be able to make use of impoundment.

The feasibility of impoundment is dependent on the overall
water balance of the location of the mine/mill's mine or mill.
In arid regions, the impoundment of tailings is a feasible
alternative to discharging and is, in fact, being practiced.
                             VII-107


                              DRAFT

-------
                              DRAFT
Where recycle systems are employed, the design must also
take water balance into consideration.  In those areas where
precipitation exceeds evaporation during all or part of the
year, some system to divert runoff away from the tailing
pond is required to keep excess water in the pond to a minimum.
Also, where heavy rainfalls periodically occur, tailing ponds
must he designed to hold the excess water accumulated during
these periods.  A mine/mill complex may find it necessary to
segregate the mine and mill effluents to further relieve the
recycle system of excess water.  In such cases, it is expected
that the mine effluents will be treated by the chemical methods
discussed previously and then discharged.

To some extent, a mill may depend on inherent loss of water
from the system to maintain a balanced recycle system.  These
losses include any or all of the following:

     (1)  Consumptive losses in the milling process (i.e.,
          retention of mositure in the concentrate, etc.);

     (2)  Retention of moisture by the tailing solids in the
          tailing pond;

     (3)  Evaporation;

     (4)  Seepage and percolation of water from the tailing pond.

The extent of these losses is dependent on a number of factors,
namely:

     (1)  Milling process employed;

     (2)  Evaporation rate (function of climate and topography);

     (3)  Type of material used to construct the tailing pond;

     (4)  Characteristics of tailing solids;

     (5)  Characteristics of soil underlying the tailing pond;

     (6)  Use of liners, diversion ditches, and other methods.
                              VII-108

                              DRAFT

-------
                             DRAFT
Given the present state of technology available and the
demonstrated status of recycle within the gold milling industry,
the maintenance of a balanced recycle system is technologically
feasible.

The feasibility of a recycle system must also consider the
effects of the reclaim water upon the mill circuit.  For example,
it has been indicated previously that reclaiming cyanidation-
process water could result in a loss of gold should this water
be introduced at the ore-grinding stage.

In the Province of Ontario, it has been found that the level of
cyanide in the tailing-pond decant from active mine/mill oper-
ations approximates 0.02 to 0.5 percent of total cyanide mill
additions (Reference    59   ).  However, data indicate that
the concentration of cyanide in tailing-pond decant may build
up if the decant is being reclaimed.  If this occurs, the
alkaline-chlorination method can be used for cyanide destruction.
Complete destruction of cyanide can be achieved by excess
addition (8.5:1) of chlorine.  On this basis, the recycling
of cyanidation-process water is considered technologically
feasible.

Recycling and zero discharge are currently being accomplished
at mill 4101, which is milling gold by the cyanide/agitation-
leach process (Figure 111-10).  The overall water balance
for this mill has been presented in Figure V-22.  Treatment
efficiency data for this mill, presented in Table VII-19,
indicate a buildup of dissolved solids and cyanide in the
reclaim water.  However, no loss in percent recovery as a
result of recycling has been reported by this mill.  In addi-
tion, the recovery rate for this mill does not differ from
that of cyanidation mill 4105, which does not recycle process
water.

Silver Ores

The discussion which follows describes treatment and control
technology currently employed in the silver-ore mining and
dressing industry.  Aspects of treatment and control pertaining
to the silver-ore category are described.

Mining Operations.   Wastewater treatment at silver mining
operations primarily consists of discharge of wastewater to
a mill tailing pond, or direct discharge without treatment.
                              VII-109

                              DRAFT

-------
                              DRAFT
Mining of silver ores primarily exploits the sulfide minerals
tetrahedrite  ((Cu, Fe, Zn, Ag)12Sb4S13) and argentite (Ag£S)
and native silver.  Native silver often occurs with gold,
copper, lead, and zinc minerals.  Little water use is encountered
in silver-ore mining, with the exception of dredging, where
silver is recovered as a minor byproduct.

Separate treatment of mine water per ee_ is not typically
practiced in this industry; however, where practiced, treatment
is performed in conjunction with treatment of mill wastewater
in a tailing pond.

Milling Operations.   As discussed in Section V, milling
processes currently employed in the silver industry are froth
flotation (about 99 percent of U.S. mill production), cyanidation
of gold ores, and amalgamation.  Cyanidation and amalgamation
recovery of silver currently constitute approximately 1 percent
of U.S. silver production by milling.  The occurrence of silver,
either with gold in a free state or as a natural alloy with
gold, has also resulted in production of silver at refineries.
Silver is often recovered also as a byproduct of the smelting
and refining of copper, lead, and zinc concentrates.

Cyanidation for gold and silver is currently being practiced
at mill 4105 (gold category), but wastewater treatment tech-
nology as currently practiced consists of a sand reclaimer pond
for removal of coarse solids only.  Amalgamation for gold
and silver is currently limited to one known site.  Waste-
water treatment at this facility has been described previously
for mill 4102.

Mill 4105, which recovers both gold and silver, currently
practices in-plant recycling of reagents, as indicated in
Section III for Gold Ores.  This results in economies of
both cost and reagent use, as well as prevention of the dis-
charge of cyanide for treatment or into the environment.
In-plant control practices common to silver flotation mills
are based on good housekeeping measures, employed to prevent
spills of flotation reagents.  The feed of these reagents into
a circuit Is carefully controlled, because a sudden increase
or decrease of some reagents could have adverse effects on
recovery from the flotation circuit.
                              VII-110

                              DRAFT

-------
                              DRAFT
Wastes resulting from silver milling are typically treated
in tailing ponds.  These ponds function primarily to facilitate
the settling and retention of solids.  Except in the case of
total impoundment, the clarified tailing-pond water is currently
discharged.  At mill 4401, a further reduction of waste loading
is achieved by partial recycle of the tailing-pond decant water
(approximately 60- to 75-percent recycle).  Mill 4402 has
achieved zero discharge through total recycle of tailing-
pond decant water.  Flotation is the predominant method
currently used to concentrate silver ore.  Flotation circuits
are commonly run under alkaline conditions.  For example,
soda ash, caustic soda, and hydrated lime are added to the
circuit of mill 4402, and lime is added to the circuit of
mill 4401.  These reagents are added to the mill circuits
to act as depressants and pH modifers and consequently make
the tailing pond alkaline.  This facilitates the removal of
metals as hydroxides in the tailing pond.  However, note that
the reagents producing an alkaline pH in the tailing pond are
added in the mill to control the process conditions there, and
a high degree of control over the pH in the tailing pond is
not currently practiced in the industry.  To facilitate opti-
mum precipitation of metal hydroxides in the tailing pond, a
higher degree of control over the pH may be required in some
cases.  Highly alkaline conditions (pH range of 10 to 11)
may be required to effect greater removal efficiency in treat-
ment facilities.

The removal of antimony from silver-milling wastewater presents
a special problem, as the hydroxide may be very unstable
and is not reported to exist.  The presence of antimony in
wastewater has been noted, because it is closely associated
with silver in some ore bodies—especially, those of the
Coeur d'Alene District of Idaho.  The sulfide of antimony
is highly insoluble; therefore, treatment for antimony removal
will involve sulfide precipitation.  Although Na2_S is itself
toxic at high concentrations, the amount required to treat
the levels of antimony found in mine and mill wastewater
(approximately 2 to 3 mg/1) is small (approximately 1 mg/1)
and will be consumed in the precipitation reaction.  Sulfide
precipitation must be carried out under alkaline conditions
to prevent the removal of sulfide ion from solution as H2J5 gas.

Cyanide is used as a pyrite depressant at mill 4401.
                             VII-111


                               DRAFT

-------
                               DRAFT
This mill is also recycling its process water with no apparent
adverse affects from this reagent.  However, should the destruction
of cyanide become necessary for process control or as a safety
measure in treating accidental leaks from the treatment system,
the alkaline-chlorination method is an effective treatment
for the destruction of cyanide.  This process has been discussed
previously in this section.  An example of tailing-pond
treatment as practiced at mill 4401 is described below.

Mill 4401 is located in Idaho.  Ore is brought to the
mill from an underground mine.  Valuable minerals in
the ore body are primarily tetrahedrite, but chal-
copyrite and galena also occur.  During 1973, 182,226 metric
tons (200,911 short tons) were milled to produce a copper/
silver concentrate.

Water used at the mill consists of both reclaim water and
makeup water, pumped from a nearby creek.  This water is Intro-
duced into the grinding circuit for the transportation and
flotation of the ground ore.  Mill tailings are discharged
at a rate of 3,188 cubic meters (835,200 gallons) per day to
the tailing-pond system.  This system is composed of three
tailing ponds and a clarification pond.  Two of the tailing
ponds are inoperative, due to extensive damage resulting from
a recent flood.  Prior to this flood, tailings were distributed
to the three ponds, and their decant was pumped to the clarifi-
cation pond.  This system covers a total area of 4.5 hectares
(10.9 acres).  Presently, water is both discharged and recycled
back to the mill from the clarification pond.  Approximately
1,649 cubic meters (432,000 gallons) per day are recycled,
while 1,141 cubic meters (299,000 gallons) per day are discharged,
Mine water is also discharged to this pond system at a rate of
553 cubic meters (145,000 gallons) per day.

A new tailing pond is under construction and is expected to be
in use soon.  This pond will have an area of 6.9 hectares
(17.0 acres).

Table VII-20 gives the chemical composition of raw and treated
waste loads from mill 4401, which uses tailing pond
treatment.  Decreases in several parameters, in addition
to suspended-solid removal, are noted.  COD, cyanide, copper,
mercury, and nickel are all reduced significantly, while
important reductions of TOC and antimony are also observed.
                             VII-112

                               DRAFT

-------
                               DRAFT
 TABLE VII-20. WASTE COMPOSITIONS AND RAW AND TREATED WASTE LOADS
            AT MILL 4401 (USING TAILING-POND TREATMENT AND
            PARTIAL RECYCLE)
PARAMETER
PH
TSS
Turbidity IJTUI
COD
TOC
Oil ind Gram
Cyinid*
At
Cd
Cr
Cu
Totil ft
Pta
Mn
Hg
Ni
Ag
Zn
Sb
MILL WASTEWATER
CONCENTRATION
(mg/D
-
555.000
20
596
220
7
DOS
<007
<002

-------
                              DRAFT
Control and Treatment Technology To Achieve No Discharge.
Currently, two silver mills are recycling their process water.
Mill 4402 reclaims all  of  its  tailing-pond decant, while mill
4401 presently reclaims approximately 60 percent of its
tailings-pond decant.   However, operation of mill 4401 with
complete recycle could  be  achieved, and would be, were it not
currently less expensive to use fresh water pumped from a
nearby well, rather  than recycled process water from an impound-
ment as makeup water.

The feasibility of recycle entails consideration of the overall
water balance at a given mill  and possible interferences in
the mill circuit caused by the recycling of process reagents
and/or buildup of dissolved solids.  Water-balance considerations
and recycling of cyanide reagent have been discussed previously
in Section VII.

Silver ores are concentrated primarily by the froth flotation
process, and it has  been noted previously that recycled flota-
tion reagents might  interfere  with the mill circuit.  However,
no published data exist which  would support this position.
Recycling successfully  being carried on at mill 4402 (total
recycle—no discharge)  and mill 4401 (partial recycle) demon-
strates the feasibility of achieving total recycle and zero
discharge.  It is expected that unwanted quantities of a
particular frother appearing in a recycle stream (from a
tailing area, etc.)  can probably be reduced or eliminated by:

     (1)  increasing the retention time of the frother-con-
          taining wastes to facilitate increased oxidation or
          blodegradation before recycle to the mill; or

     (2)  oxidation  of  the frothers through application of a
          degree of  mechanical aeration, etc., to the waste
          stream; or

     (3)  selecting  another frother with superior breakdown
          properties for use in the mill.

A further degree of  control of the recycle system can be
gained by use of a two-cell pond.  In this system, clarified
water from the primary  pond would be decanted to the second
                             VII-114

                              DRAFT

-------
                             DRAFT
pond, which would be used as a surge basin for the reclaim
water.  This system would lend itself to increased control
over the slime content of reclaim water.  This is desirable,
since these slimes have been thought to inhibit differential-
flotation processes in some mills.  In addition, the second
pond would provide a site for the implementation of mechanical
aeration, should this treatment become necessary.

Segregation of Waste Streams.   At certain mine/mill complexes,
for the mill to achieve a balanced recycle system, it may be
necessary to segregate the mine and mill waste streams.
In such cases, it is expected that, prior to discharge,  the
mine effluents would be chemically treated for the removal
of metals and suspended solids in settling ponds.  As pre-
viously discussed, this treatment would normally involve
precipitation of metals using lime and/or sulfides.

The discussion which follows describes a silver milling
operation currently operating with recycle and zero discharge.

Mill 4402 is located in Colorado.  Ore is brought to
the mill from an underground mine.  Valuable minerals
in the ore body include sulfide of silver—primarily,
argentite, galena, and free or native silver.  During 1973,
75,005 metric tons (82,696 short tons) of this ore were milled
to produce a lead/silver concentrate.

Process water is recycled at this mill.  However, makeup water
is required, and this water is pumped from a well.  Water is
introduced into the grinding circuit to facilitate trans-
portation and flotation of the ground ore.  Mill tailings are
sent through two stages of cyclones to remove sands, which are
used for backfilling stopes in the mine.  Cyclone overflow
is discharged to a 1.6-hectare (4-acre) tailing pond at a rate
of 1,511 cubic meters (396,000 gallons) per day.  Clarified
pond water is recycled back to the mill at a rate of 962
cubic meters (252,000 gallons) per day.

A new tailing pond is being built at this mill.  This pond will
have an area of 6 hectares (15 acres).

Table VII-21 demonstrates the treatment efficiency achieved
in the mill tailing pond and compares mill raw-wastewater
                            VII-115

                             DRAFT

-------
                          DRAFT
TABLE VI1-21. CHEMICAL COMPOSITIONS OF MILL RAW WASTE WATER
           AND TAILING-POND DECANT WATER AT MILL 4402
PARAMETER
TSS
Turbidity (JTU)
COD
TOC
Oil and Grease
Cyanide
As
Cd
Cr
Cu
Total Fe
Pb
Total Mn
Hg
Ni
Ag
Zn
Sb
CONCENTRATION (ing/ 1 )
MILL RAWWASTEWATER
90,000
1.05
22.70
29.0
2
< 0.01
0.07
< 0.02
< 0.1
0.22
1.80
0.56
1.75
0.149
0.10
< 0.02
0.37
< 0.2
TAILING-POND DECANT
2
0.575
22.70
17.5
2
< 0.01
< 0.07
< 0.02
< 0.1
< 0.02
1.59
0.10
1.80
0.002
0.11
< 0.02
2.3
< 02
                         vri-ii6
                         DRAFT

-------
                              DRAFT
input to tailing-pond decant water recycled to the mill.
No waste loads are presented, because no discharge results.

Bauxite Ore

As discussed in Section IV, Industry Categorization,  two
bauxite mines currently operating in the U.S. extract bauxite
ores from open-pit and underground mines.  The characteristics
of pollutants encountered in wastewaters from these operations
are discussed in Section V.  The current treatment technology
and industry practice for treatment of bauxite-mine drainage
are described below.

Lime neutralization is the only treatment method presently being
employed by the two domestic bauxite producers to treat mine
water.  Both acidic and alkaline waters are treated by this
technique, but, due to the relatively small amount of alkaline
water that is treated daily  (83 cubic meters, or 22,000 gallons,
per day), only acid mine-water neutralization is discussed in
detail here.

Generally, mine water and surface drainage destined for treat-
ment undergo settling in a number of natural depressions,
sumps, and settling ponds before reaching the lime-neutraliza-
tion facility; thus, suspended-solids loadings are reduced.

The addition of lime to raw mine drainage to reach elevated
pH causes precipitation of heavy metals  as insoluble or
slightly soluble hydroxides.  Formation  of specific metal
hydroxides is controlled by  pH, and removal of the suspended
hydroxides is accomplished by settling.  The discussion of
this treatment technique is  presented in the early portion
of Section VII under Chemical Precipitation.

Two variations of  lime storage at bauxite-minewater-treatment
facilities are employed, and both systems achieve slightly
different efficiencies of  pollutant removal.  It  is doubtful
that treatment efficiency  is dependent upon  lime  storage
technique, but the pH and  pH control of  the  limed solution
are probably  the dominant  factors in determining  concentration
levels attained in settling  ponds.

Figure VII-16 is a schematic flowsheet of  the lime-neutraliza-
tion facility at open-pit  mine 5101, which processes approxi-
mately  7,165  cubic meters  (1,900,000 gallons) per day  of  raw
mine drainage.
                             VII-117

                               DRAFT

-------
                               DRAFT
Mine 5101.   Open-pit mine complex 5101 is located in
Arkansas and produces about 2,594 metric tons (2,860
short tons) of high-silica bauxite daily.  There are several
pits associated with the water-treatment facility, and acid
waters collected from the pits, spoils-storage areas, and
disturbed areas are directed to the treatment plant.

Mine 5101 treats the major portion of its open-pit mine drainage
through the treatment plant, as shown in Figure VII-16.  Other
open-pit drainages which require intermittent pumping for
discharge will be treated by a mobile lime-treatment plant in
the near future.  At this operation, about 0.45 kg (approxi-
mately 1 pound) of slurried lime is used to neutralize 3.79
cubic meters (1000 gallons) of acid mine water.  This facility
has a controlling pH probe, located in the overflow from the
detention tank, which activates the automatic plant and pump
cutoffs at a high point of pH 9.0 and a low point of pH
6.0.  The operating pH generally ranges from 7.5 to 8.0,
and the pH of the effluent discharge ranges from 6.3 to 7.3.

Table VII-22 lists analytical data for raw mine water (silt-
pond overflow) and treated effluent (as the discharge leaves
the overflow weir at the sludge pond).

Mine 5102.   Open-pit mine 5102 is also located in
Arkansas and mines a high-silica-content bauxite deposit.
Contaminated surface drainage from outlying areas and ground-
water accumulation in the holding pond produce about 14,140
cubic meters (4,000,000 gallons) of raw drainage daily.
Surface drainage collects from an area of approximately 662
hectares  (1,635 acres) of disturbed and undisturbed land.

An experimental lime-neutralization plant has been operated
at mine 5102 and processes approximately 2,650 cubic meters
(2,700,000 gallons) per day of acid mine drainage.

This mining operation presently treats less than 10 percent
of its total raw mine drainage, but full-scale operation of
a treatment plant having a capacity of 11,355 cubic meters
(3,000,000 gallons) per day is expected in mid-1975.  The
new plant will operate similarly to the present plant, but
an enlarged system of settling lagoons and sludge drying
beds should provide adequate treatment efficiency.
                              VII-118

                               DRAFT

-------
                            DRAFT
Figure VII-16. LIME-NEUTRALIZATION PLANT FOR OPEN-PIT MINE 5102
    LIME-SLURRY
     STORAGE
      TANKS
                                      SLUDGE
                                     SETTLING
                                       POND
RAW-WATER
 HOLDING
  POND
CLEAR-WATER
  SETTLING
   POND
1.84 m°/day
(486 gpm)
                                                 1.84 m3/day
                                                 (486 gpm)
                                                      DISCHARGE
                            VII-119
                             DRAFT

-------
                             DRAFT
       TABLE VII-22. CHEMICAL COMPOSITIONS OF RAW AND
                     TREATED MINE WATERS AT MINE 5101
PARAMETER
PH
Acidity
Alkalinity
Conductivity tf
TDS
TSS
Total Fe
Total Mn
Al
Ni
Zn
Sr"
Fluoride
Sulfate
CONCENTRATION (mg/£)
RAW MINE DRAINAGE
RANGE
2.8 to 3.0f
250 to 397
0
1000
560 to 617
2
7.2 to 21 .8
3.2 to 3.5
23.8 to 18.6
0.3 to 0.31
0.82 to 1.19
0.5
0.048 to 0.290
490 to 500
AVERAGE*
2.9*
324
0
1000
589
2
14.5
3.4
21.2
0.3
1.01
0.5
0.17
495
TREATED EFFLUENT
RANGE
6.0 to 6.8f
OtolO
6 to 13
1000
807 to 838
1.2 to 4.0
0.14 to 0.20
2.25 to 3.37
0.33 to 0.80
0.18 to 0.19
0.07 to 0.09
1.74
0.03 to 0.67
500 to 581
AVERAGE*
6.4*
0.5
10
1000
823
3
0.2
2.8
0.6
0.2
0.08
1.74
0.35
541
 •Values based on two grab samples
 f Value in pH units
"Value in mieromhos/cm
"value* based on one grab sample
                             VII-120

                              DRAFT

-------
                           DRAFT
The treatment used at mine 5102 involves slurried storage of
lime in large agitator tanks for eventual mixing with mine
water in the confines of a pipeline.   About 0.83 kg (1.82 Ib)
of hydrated lime is used to neutralize 3.79 cubic meters
(1000 gallons) of raw mine water.  This lime rate maintains
the influent to the sludge pond at a pH of 9.0 to 11.0, and
efficient from the clear-water settling pond varies from a pH
of 6.0 to 8.0.

Table VII-23 lists the chemical composition of both raw mine
water (influent to the treatment plant) and the treated efflu-
ent (discharge from clear-water settling pond).

Ferroalloy Ores

The ferroalloy-ore mining and dressing category includes,
for purposes of treatment here, operations mining and bene-
ficiating ores of cobalt, chromium, columbium and tantalum,
manganese, molybdenum, nickel, tungsten, and vanadium (one
operation extracting non-radioactive vanadium).  Vanadium
obtained from milling of uranium, vanadium, and radium ores
under NRC licensing is covered as part of the uranium-ore
category.  Since the subcategorization of this category is
not based upon end product recovered, but rather upon the
process used, representative mines and mills are used to
illustrate wastewater treatment and control as practiced in
ferroalloy-ore subcategories.

Currently, there are no operations mining or beneficiating
ores of chromium, cobalt, columbium , and tantalum. A mangan-
iferous ore is currently being mined at one location in the
U.S., but no wastewater results, and no milling activities
are carried on.  A second manganiferous ore mine and mill
is expected to reopen in late 1975 or 1976.  Consequently,
treatment and control technology currently employed in the
molybdenum, nickel, tungsten, and vanadium industries will
be used as examples here to represent treatment used in
subcategories of this category.

Mining Operations.   Mining of ferroalloy ores is by both
underground and open-pit methods.  Mine wastewater is
characterized by high and variable flow and dissolved heavy
metals, and is often acidic.  At open-pit mines, seasonal
                          VII-121

                           DRAFT

-------
                      DRAFT
TABLE VI1-23. CHEMICAL COMPOSITIONS OF RAW AND
              TREATED MINE WATERS AT MINE 5102
PARAMETER
pH"
Acidity*
Alkalinity
Conductivity *
TDS
TSS»
Total Fe«
Total Mn
Al»
Ni
Zn
Sr
Fluoride
Sulfate"
CONCENTRATION (mg/l)
RAW MINE
DRAINAGE
2.91"
240
0
2.21 2««
468
45
49.0
1.56
14.8
0.05
0.24
0.1
0.59
432
TREATED
EFFLUENT
7.2'
0
30
897"
630
6.6
0.29
<0.02
0.12
<0.02
<0.02
—
0.56
343
       •Values based on industry samples and represent the
        average of eight or more grab samples taken in 1974.
       * Value in pH units
      ••Value in micromhos/cm
                       VII-122

                        DRAFT

-------
                              DRAFT
fluctuations in mine water may be extreme.  At such opera-
tions, acidic streams from sulfides in mine waste dumps cdd
to the waste load of the wastewater requiring treatment.

Mine water is often used as mill process water at underground
mines.  At open-pit operations, seasonal variability generally
makes mine water an unacceptable source of process water.
Treatment for suspended-solid removal is almost universally
practiced in the ferroalloy-ore mining industry.  Both treatment
in tailing ponds with mill wastewater and use of separate
treatment systems such as settling ponds and clariflocculators
(variants of mechanical clarifiers in which mixing is provided
for flocculant distribution) are used.  Where waste streams
are acidic, neutralization is generally practiced.  Where open-
pi