EPA-R2-73-247a
September 1973
Environmental Protection Technology Series
PROTECTION
AGENCY
Water Pollution Control III
Primarv Noiiferrous -
Metals Industry - Vol. t
Copper, Zinc, And Lead
I
55
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Office of Research and Development
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RESEARCH REPORTING SERIES
Research reports of the Office of Research and
Monitoring, Environmental Protection Agency, have
been grouped into five series. These five broad
categories were established to facilitate further
development and application of environmental
technology. Elimination of traditional grouping
was consciously planned to foster technology
transfer and a maximum interface in related
fields. The five series are:
1. Environmental Health Effects Research
2. Environmental Protection Technology
3. Ecological Research
U. Environmental Monitoring
5. Socioeconomic Environmental studies
This report has been assigned to the ENVIRONMENTAL
PROTECTION TECHNOLOGY series. This series
describes research performed to develop and
demonstrate instrumentation, equipment and
methodology to repair or prevent environmental
degradation from point and non-point sources of
pollution. This work provides the new or improved
technology required for the control and treatment
of pollution sources to meet environmental quality
standards.
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EPA-R2-73-247a
September 1973
WATER-POLLUTION CONTROL IN THE
PRIMARY NONFERROUS-METALS INDUSTRY
VOLUME I. COPPER, ZINC, AND LEAD INDUSTRIES
By
J. B. Hallowell
J. F. Shea
G. R. Smithson, Jr.
A. B. Tripler
B. W. Gonser
Contract No. 14-12-870
Project 12010 FPK
Project Officer:
John ,Ciancia
Edison Water Quality Research Laboratory4 NERC
Edison, New Jersey 08817
Prepared for:
OFFICE OF RESEARCH AND MONITORING
U.S. ENVIRONMENTAL PROTECTION AGENCY
WASHINGTON, D.C. 20460
For sale by the Superintendent of Documents, U.S. Government Printing Office
Washington, D.C. 20402 - Price $1.90
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EPA Review Notice
This report has been reviewed by the Environmental
Protection Agency and approved for publication.
Approval does not signify that the contents neces-
sarily reflect the views and policies of the
Environmental Protection Agency, nor does mention
of trade names or commercial products constitute
endorsement or recommendation for use.
ii
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ABSTRACT
In a review of water-pollution-control practices in the primary non-
ferrous metallurgy industries, unpublished data were obtained on waste-
water-control practices from 28 companies operating 75 separate pro-
ducing plants.
Volume I of this study is concerned with the processes and water prac-
tices in the copper, lead, and zinc industries. Data supplied by 50
copper-, lead-, and zinc-producing plants showed that waste-water-
control practices are influenced by climate and production processes.
No waste water is discharged from many operations in the Southwestern
desert area. About two-thirds of the waste waters discharged by the
balance of the plants were from tailings ponds and exhibited neutral
to alkaline pH. In some cases, these waste waters contained "trace"
amounts of heavy metals.
The most pressing needs and problem areas identified were related to
smelter-refinery-type waste waters, which were generally acid and
carried metals or metalloids in the range <1 to 100 ppm. Other areas
where improvements are needed include the treatment of plant sanitary
wastes, accurate and economical analyses of low levels of waste-water
components, expanded and improved scientific bases for setting hazard
limits of heavy metals in water, and development and cost-evaluation
of alternative or novel water treatment and metal and by-product pro-
duction processes.
This report was submitted in fulfillment of Contract No. 14-12-870
under the sponsorship of the Office of Research and Monitoring,
Environmental Protection Agency.
ill
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CONTENTS
Section
I Conclusions
II Recommendations
III Introduction
IV Procedures
V The Primary Copper Industry
VI The Selenium and Tellurium Industry
VII Water Usage in the Copper Industry
VIII The Primary Lead and Zinc Industries
IX The Cadmium Industry
X By-Product Gold and Silver
XI The Arsenic Industry
XII The Primary Antimony Industry
XIII The Primary Bismuth Industry
XIV Water Usage in the Lead-Zinc Industry
XV Discussion
XVI Acknowledgment
XVII References
1
3
5
7
c
51
55
71
113
123
127
129
135
137
145
165
167
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FIGURES
1 Geographical Distribution of Primary Copper Production 11
Facilities in the United States
2 Past Production and Estimated Growth of the Primary 19
Copper Industry
3 Basic StepsCopper Ore to Finished Product 23
4 Diagram of Operations in Copper Production 24
5 Flowsheet of Typical Flotation Process 29
6 Reverberatory Matting Furnace for Copper Ores 34
7 Flowsheet of Selenium-Tellurium Recovery Process 53
8 Generalized Water Flow Diagram for Mine-Concentrator 61
Operation
9 Generalized Water-Flow Diagram for Smelter-Type Operation 62
10 Generalized Water Flow Diagram for Electrolytic Refinery 63
11 Geographical Distribution of Lead and Zinc Mines in the 80
United States
12 Diagram of Operations in Lead and Zinc Production 83
13 Process Steps in Differential Flotation of Lead-Zinc Ores 88
14 Generalized Flowsheet of Lead Smelter 91
15 Diagrams of the Operation of Sintering Machines 93
16 Diagrammatic Sketch of a Typical Lead Blast Furnace 95
17 Diagram of Alternative Lead-Softening Procedures 99
18 Diagram of Steps in Fire Refining of Lead 100
19 Illustrations of the Elements of the Horizontal Retort 106
Operation
20 Schematic Drawing of the Electrothermic Zinc Reduction 109
Furnace
vi
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FIGURES (continued)
Page
21 Typical Flowsheet of an Electrolytic Zinc Plant 111
22 The Pachuca Tank 112
23 Cadmium Recovery in the Zinc Retort Process 115
24 Cadmium Recovery From Zinc Metal 118
25 Cadmium Recovery From Lead, Zinc, or Copper Smelter 119
Dusts
26 Cadmium Recovery From Purification Sludge of Zinc 121
Electrowinning Operations
27 Diagram Showing Steps in the Recovery of By-Product 124
Gold and Silver
28 Steps in the Concentration and Purification of Arsenic 128
Trioxide
29 Diagram of Process Steps in the Production of Antimony 134
at Sunshine Mining Operation
VII
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TABLES
No.
1 Mine Production of Copper in the United States 12
2 Major Copper-Producing Mines in the United States 13
3 Copper Smelters and Refineries 16
4 Selected Economic Statistics for the Primary Copper 17
Industry
5 Copper Minerals Important in U.S. Production 20
6 United States Producers of Selenium and Tellurium 51
Metals and Compounds
7 Sample of Copper-Producing Operations 56
8 Water Data for Copper-Producing Facilities 57
9 Water Uses and Recirculation Practice in Copper 59
Producing Mines and Plants
10 Compositions of Waste Waters From Mine and Concentrator 66
Operations
11 Water Analyses Associated With Copper Smelting 67
Operations
12 Composition of Waste Streams From Copper Smelters and 68
Refineries
13 Waste Water Treatment Practices in the Copper Industry 69
14 Production of Lead and Zinc in the United States in 72
1969, by State and Class of Ore, From Old Tailings, Etc.,
in Terms of Recoverable Metals
15 Lead and Zinc Mines Reported by 1968 Industry Survey 74
16 Zinc-Producing Plants in the United States 78
17 Lead Smelters and Refineries in the United States 79
18 Selected Economic Statistics for the Primary Zinc and 81
Lead Industries
19 Consumption Pattern of Lead and Zinc in 1968 81
viii
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TABLES (continued)
No.
20 Ore Minerals of Lead and Zinc 82
21 Approximate Grade and Association of Lead-Zinc Ores 84
in the United States
22 Ranges of Compositions of Concentrates in Lead and Zinc 89
Metallurgy
23 Chemical Specifications for Commercial Pig Lead 102
24 Producers of Cadmium Metal 113
25 Salinet Antimony Statistics 130
26 Industrial Consumption of Primary Antimony in the 130
United States, by Class or Material Produced
27 Consumption of Bismuth in the United States in 1969 135
28 Sample of Lead- and Zinc-Producing Operations Responding 138
to Survey
29 Water Use Data for Lead and Zinc Mines and Plants 139
30 Composition of Waste Waters and Receiving Streams in 142
Lead-Zinc Metallurgical Processes
31 Waste Water Treatment Practices in Lead-Zinc Smelting 144
and Refining Operations
32 Summary of Reported Waste Treatment Practice 147
33 Summary of Waste Water Treatment Practices 146
34 Reported Total Water Costs 149
35 Water Treatment Methods and Costs 150
36 Waste Water Treatment Methods and Costs 151
37 Summary of Plants and Recommended Areas of Future 155
Development Work
38 Measures Under Study or in Effect 157
39 Assessment of Information Obtained in Terms of Unit 159
Processes and Future Needs
ix
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SECTION I
CONCLUSIONS
1. Fourteen of the approximately fifty plants reviewed during the
study reported zero discharge of waste water as a consequence of their
location in the Southwestern desert.
2. Approximately 38 billion of the 56-1/2 billion gallons of water
discharged from the balance of the 50 operations may be categorized as
tailings-pond effluent ranging from neutral to alkaline and containing
low levels of metals as the critical impurities.
3. The balance of discharged water is attributable to smelter-refinery
operations and is highly variable in characteristics.
4. The principal area of research needs identified in this study was
for treatment methods to deal with acidic waste waters containing <1
to 100 ppm of metals or metalloids, generated primarily by electrolytic
refineries and by smelters. Other research needs identified included
the treatment of plant sanitary wastes, accurate and economic analysis
of waste water components, scientific bases for hazardous limits of
heavy metals, and development of new metallurgical processes to decrease
water usage or discharge.
5. The current plans of industry to improve waste water control in-
clude the approaches of total recycle, treat and discharge, and combi-
nations of these, i.e., partial recycle with reduced discharge.
6. In only two instances were treatment plans being considered which
would result in increased recovery of products.
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SECTION II
RECOMMENDATIONS
Based on the findings of this study, it is recommended that:
1. Programs be carried out as rapidly as possible to provide cost-
effectiveness evaluations of methods of treating acid, metal-bearing
wastes and waste water, preferably using samples of specific operating
plant wastes as exampled in this report.
2. There be developed new methods for the treatment of plant sanitary
wastes, in the context of the size, surrounding resources, and water
requirements of the metallurgical or industrial plant.
3. Diverse programs be undertaken to provide methods, manpower, equip-
ment, and facilities to achieve routinely accurate, economical analyses
of water down to the parts-per-billion range.
4. Sufficient study be performed of the published literature, plant
practice, and/or mechanistic behavior to establish the practical limits
of water quality which can be achieved by tailings-pond type treatments.
5. Long-term programs be supported to establish what levels of heavy
metals in water are hazardous to man and aquatic life.
6. New developments in metallurgical processing be assessed in terms
of water requirements and waste water generation and that new process
developments be supported which use less water or produce recyclable
or zero discharges of waste water.
7. A more intensive study be conducted to obtain data on individual
unit process waste sources. Such a study should include some provision
for sampling and analysis. This work should be aimed at defining point
sources, characteristics, volumes, unit waste loads and the effective-
ness of treatment practices in more detail than the current study has
shown to be presently available.
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SECTION III
INTRODUCTION
The purposes of this study are to assess current water-pollution-con-
trol practices and problems and to provide a basis for recommending
specific directions to develop water-pollution-control methods in the
nonferrous metals industry. The approach used in this study and in
this report has been to deal with specific practices and problems
rather than with generalities or overall considerations. This study,
in attempting to be specific, has thus dealt with the variegated and
individualistic nature of the metal-producing operations which con-
stitute the total industry.
This report is organized to present the data obtained in groupings
which are determined by the relationships between production operations.
Thus, Volume I deals with the production processes and water-use prac-
tices of the copper, lead, and zinc industries. These major metal-
plant operations include the manufacture of the following by-products:
cadmium; "by-product" gold, silver, platinum, palladium, and rhodium;
arsenic; bismuth; thallium, indium, selenium, tellurium, and molybde-
num concentrates or oxides; antimony; and the metal compounds arsenic
trioxide, copper sulfate, nickel sulfate, zinc oxide, zinc sulfate,
and sulfuric acid. Volume II deals with operations producing alumina
and aluminum, mercury, primary gold and silver, and molybdenum and
tungsten.
In preparing this report, it was found necessary to deal with manufac-
turing processes in a somewhat intensive manner in order to identify
the process steps which produce characteristic components of waste
water; i.e., in some cases the primary metal product is so efficiently
extracted that little waste is generated, while in other cases, margi-
nal amounts of material from by-product operations or nominally minor
components of raw materials appear in waste streams.
The method used in the study was to assemble the often fragmentary in-
formation available in the literature and, more importantly, to obtain
information directly from the industry. The letters of inquiry sent
to selected companies requested information on current practices of
water usage and on the sources, characteristics, and amounts of waste
water, as well as the industry's own statements of problem areas and
treatment needs. Two constraints were placed on the data supplied
by industry: that such information be supplied voluntarily and at the
expense of industry, and that the specific sources of each lot of in-
formation remain anonymous.
In reporting the results of this study, separate sections have been
devoted to each metal or metal grouping. Each section discusses the
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size and distribution of the particular industry, its economic char-
acteristics, the technology employed, its water-usage characteristics,
its waste-collection, -treatment, and -disposal practices, and antici-
pated future treatment requirements.
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SECTION IV
PROCEDURES
The sequence of tasks in this program consisted of:
1. Compilation of a list of the companies and plant facilities in the
United States which produce nonferrous metals
2. Contacting the identified companies by telephone, letter, or per-
sonal visits to obtain the desired information
3. Compilation and analysis of the data obtained to prepare this
report
4. Use of documentary sources wherever available to obtain required
associated information such as industry structure, overall economic
aspects, process technology, prior studies on water usage, and other
supplementary information.
The open literature was found to contain little specific information on
water problems or treatment in the nonferrous-metals industry. This
study, therefore, placed its main reliance for specific details on the
voluntary contributions of unpublished data from industry. Further,
it was found that the degree of precise knowledge of plant water usage
is still highly variable from plant to plant.
Out of 56 companies approached, about 50 percent responded by submit-
ting data in varying degrees of detail. Their replies covered 78
facilities ranging from single mine operations to complexes of mines,
mills, smelters, refineries, etc. The responses received provided in-
formation on 35 mines, 33 concentrators, 8 copper leach-precipitation
operations, 27 copper, lead, and zinc smelters, 20 refineries, 8 sul-
furic acid plants, and 8 power plants.
In view of the extreme complexity of the nonferrous-metals industry
and recognizing that many who use this report may be unfamiliar with
the industry, it was decided that each section should include a more
or less detailed description of the industrial segment, its structure,
and technology as a convenient aid to understanding and interpreting
its water- and waste-treatment problems and practices.
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SECTION V
THE PRIMARY COPPER INDUSTRY
Major Characteristics
The primary copper industry is a vast network of mines and treatment
facilities spread over 19 states in the country, producing annually
about 1.7 million tons of refined copper. The unit operations re-
quired to produce this copper, aside from exploration or development,
are mining, physical concentration of ores, smelting, and refining.
In the case of some ores not amenable to physical concentration, either
technically or economically, auxiliary processes such as leaching,
cementation, solvent extraction, and electrowinning are also employed.
Although these processes will be described in detail later in this re-
port, brief definitions of them are presented at this time to assist
the reader in understanding the intervening tables and texts.
Exploration
Development
Mining
Concentration
Smelting
Refineries
Leaching
-- Locating, outlining, and evaluating ore bodies.
-- Preparing the ore body for mining, as by removing
overburden, sinking shafts, driving tunnels, etc.
-- Removal of the ore from open-pit or underground mines.
-- Copper ores typically containing from about 0.6 to
about 1.5 to 2.0 percent copper yield over 90 percent
of the copper produced. These are concentrated by
the relatively cheap physical concentration process of
flotation, to eliminate the valueless material and
produce a "concentrate" containing 15 to 35 percent
copper which can then be further processed.
-- A high-temperature series of furnacing operations in
which copper concentrates produced from concentrating
ores, together with relatively small quantities of
higher grade ore and other copper products, are con-
verted to a partially refined copper.
-- There are two types of refineries--electrolytic and
"fire" refineries. The purpose of refining is to pro-
duce copper of acceptable physical, chemical, and
electrical properties for subsequent fabrication into
finished products. Most copper is refined electro-
lytically.
-- Some ores are too low in grade or of such mineralogical
composition that they cannot be economically treated
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even by the relatively cheap flotation process.
These are "leached", that is, subjected to treatment
by dilute sulfuric acid solution, whereby copper is
dissolved for subsequent recovery by specialized
methods such as cementation, solvent extraction,
electrowinning, etc.
Geographical Distribution
The distribution of copper-producing facilities throughout the United
States is shown in Figure 1.
Mines
Table 1 is a state-by-state breakdown of the various types of mines
producing copper in the United States for the year 1969 and the amount
produced in that year.^ ' >^) For comparison, preliminary figures for
1970 are included. The table shows, for each of the states producing
copper, the number of mines in operation and the tonnages of copper
they produced. The table also shows the number of leaching installa-
tions in the various states and the tonnage produced by them in 1969.
As shown in the table, the "copper" mines, i.e., those operated pri-
marily for their copper value, produced, in 1969, about 1.3 million
tons of recoverable copper out of a total production for that year
of about 1.55 million tons, or more than 80 percent. Except for one
large mine in Michigan and five relatively small mines in Tennessee,
the "copper" mines are located in the West and Southwest in the states
of Arizona, Montana, Nevada, New Mexico, and Utah.
The next major contributors to copper production are the various leach-
ing cleanup and recovery operations which accounted, in 1969, for
about 170,000 tons of copper or about 10 percent of total production.
Many of these and most of the tonnage they produced are associated
with copper-mining operations, which customarily leach low-grade and
oxidic ores.
Copper production from the by-product mines (copper-lead-zinc, tung-
sten, iron, and precious metals), while economically important, is
relatively small.
Major copper mines, their location, and approximate production data
are shown in Table 2.
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Concentrators
Concentrating facilities shown in Figure 1 refer only to the copper
mines and not to the mines in which copper is a by-product. As ex-
plained previously, water problems associated with the by-product
mines are properly chargeable to other segments of the nonferrous-
metal-production industries.
It will be noted that the number of concentrators shown for the copper
mines in the major producing states are considerably less than the num-
ber of copper mines (Montana, Nevada, Utah, Arizona, New Mexico). This
is because one concentrator may serve a number of mines.
Smelters
The location by states and capacities of smelters and refineries are
shown in Table 3.
Contribution to the United States Economy
Up-to-date data to illustrate the direct importance of the primary
copper industry to the United States economy were not available at the
time of this study. The latest information, for 1967 which was unfor-
tunately a strike year, will provide some idea of the magnitude of
this contribution. These data, selected from the 1967 Census of
Manufactures^) ; are shown in Table 4.
Future Development of the Industry
Projections of the growth rate of the copper industry have been pub-
lished in various terms. One such projection predicts an increase in
consumption of 4-1/2 to 5 percent during the 1970 to 1975 period. *> '
This prediction, in terms of total consumption, gives a rate higher
than the predicted rate of expansion of the general economy. However,
the fact that the prediction is in terms of consumption makes it hard
to apply to considerations of domestic primary-plant operations (i.e.,
water use) since consumption includes imports of raw material and fin-
ished products, and production from secondary materials.
A prediction published by the Bureau of Mines is given in terms of
metal content of domestically mined ores and would be considered a
better indicator of the possible expansion of plants within the United
States. This prediction is based on 1968 production data, and gives
15
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TABLE 4. SELECTED ECONOMIC STATISTICS FOR THE PRIMARY
COPPER INDUSTRY (1967/ '
Copper Smelting and
Characteristics Ores Refining
Number of establishments
Total 157 32
With 20 employees or more 61 32
Number of employees 21,000 11,600
Payroll, millions of dollars 170.9 80.6
Value of shipments, millions of
dollars 675.9 1184.1
Capital expenditures, millions of
dollars 123.6 51.7
Materials consumed:
Ores, concentrates, precipitates,
short tons -- 3,601,500
Other forms of copper metal, short
tons -- 435,700
Delivered cost, millions of dollars 827.6
Cost of all other materials and
supplies, millions of dollars 77.2
Cost of purchased fuels and electrical
energy, millions of dollars 21.5
Product made by primary copper refiners,
short tons 1,307,900
17
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an estimate up to the year 2000 with a range of possible growth rates
from a low of 3.7 percent to a high of 5.2 percent.(8) The production
portion of this estimate is shown in Figure 2, along with past produc-
tion figures. The estimated expansion of primary mine production is
shown as related to past mine production. The figure also illustrates
that primary refinery production, while related to domestic mine pro-
duction, shows the inclusion of domestic and/or imported raw materials
and scrap in primary refinery production. The estimated primary
domestic mine production for the year 2000, ranging from 3.8 to 6.1
million tons of copper, is coupled, in the published prediction, with
an estimated demand for primary copper of 4.9 to 7.8 million tons, with
the difference again being attributable to imports.
Raw Materials
The major raw material in the primary copper industry is, of course,
the ore. In 1970, for example, over 300 million tons of ore were
mined to produce 1.7 million tons of copper. Other raw or auxiliary
materials going into the production of refined copper are as follows:
Mining
-- blasting materials, fuels, timber
Transportation -- fuels, lubricants
Concentration
SmeIting
various conditioning chemicals (such as zinc sulfate,
lime, sodium sulfide, cyanide); flotation reagents
(such as xanthates, phosphates, oils, and frothing
agents); chemicals for improving flocculating,
settling, and filterability (such as various pro-
prietary compounds, lime)
fluxing materials such as quartz rock, silicate ores,
lime rock, fuels (coal, coke, natural gas, oil)
Fire Refining -- various fuels, purification gases such as air, oxy-
gen, and natural gas
Electrolytic -- sulfuric acid, various addition agents, such as glue,
Refining etc., to improve deposition
Machinery -- a high capital investment is made in specialized
machinery such as drilling and earth-moving machines
and other plant equipment, e.g., furnaces.
There are about 160 known copper minerals in nature. Only about a
dozen of these are commercially important and 5 or 6 account for over
90 percent of the copper produced. The minerals of importance in
United States ores are shown in Table 5.
18
-------
go suox jo suof[-[jj4
19
-------
TABLE 5. COPPER MINERALS IMPORTANT IN U.S. PRODUCTION^)
Mineral
Composition
Copper Content
of the Mineral Form,
percent
Native copper metal
Chalcopyrite
Chalcocite
Covellite
Bornite
Enargite
Cuprite
Malachite
Azurite
Chrysocolla
Cu
Sulfide Ores
CuFeS2
Cu2S
CuS
Oxide Ores
CuC03-Cu(OH)2
2CuC03-Cu(OH)2
CuSi03- 2H20
100
35
80
66
63
48
89
57
55
36
These minerals occur in a variety of modes to make up what is called
"ore", that is, a deposit which can be mined at a profit. Rarely in
the United States are they found in any quantity as massive deposits
of the pure mineral that can be mined selectively and smelted, but are
practically always dilute mixtures with other rocks of little or no
value. The most important copper-ore type in the United States is the
so-called "porphyry" copper deposits. These are extensive masses of
rock throughout which crystals of various copper minerals are more or
less uniformly disseminated, and which, although low grade, may pro-
fitably be mined on a massive nonselective scale. The copper minerals
generally associated with the porphyries are the various oxides, such
as cuprite and malachite which have been converted to these oxide forms
by weathering processes and, lower in the deposit, various sulfide
minerals such as chalcocite, covellite, and chalcopyrite. The por-
phyry copper deposits account for the major portion of copper produc-
tion. The average grade of ore for all copper production in 1968 was
0.6 percent. The copper content of the porphyry ores, therefore, must
lie around 0.6 percent. Porphyry ores are mined by open-pit methods.
Other major types of ore are the vein, pipe, and bedded deposits,
which generally yield higher grade ores ranging from 3 to about 10
percent copper, and which are usually mined by underground methods.
The copper minerals in such deposits are sulfidic, frequently
20
-------
associated with sulfides of other metals such as pyrite or pyrrhotite
and include chalcopyrite, bornite, chalcocite, and covellite. A few
of the deep United States copper deposits contain some copper-arsenic
minerals such as enargite or tennantite.
Native copper may occur in unimportant quantities in oxidized ore de-
posits, but in one place in the United States (Michigan) it is a sig-
nificant ore mineral associated with covellite, a copper sulfide.
Minor, though economically important, concentrations of copper, gener-
ally as sulfide but occasionally as an oxidic material, also are found
with the ores of other metals (lead, zinc, silver, iron).
The copper ores of the United States may be classified as follows:
(a) Concentrating-grade ores. As mined, these will contain from
about 0.6 to about 1 or more percent copper mostly as sulfide. They
account for about 75 to 80 percent of primary copper production in the
country. The so-called porphyry coppers of the west, mined by open-
pit methods, are the most important typical sulfide ores of concen-
trating grade. Underground mines also produce such ores. These ores
may also contain oxidic copper in minor proportions.
(b) Low-grade (or leaching grade) ores. These are the ores contain-
ing significant quantities of copper as sulfide and/or oxide and which
cannot be economically concentrated by the basic process. They may
contain from a few tenths to about 0.5 percent copper. In mining,
such ores are segregated and leached. They occur near the top and to
some extent throughout copper deposits, and in the course of mining
must be removed so that the higher grade ore can be mined systematic-
ally and economically.
(c) High-grade sulfide ores. These are generally obtained from under-
ground operations and occur in vein, pipe, and bedded deposits. They
may range from about 3 to 10 percent copper in the United States.
Their tonnage is not great.
(d) Mixed sulfide-oxide ores. These are ores in which the sulfides
and oxide minerals are present in approximately equal amounts, and
which require special treatment to yield an economic level of recovery.
Such ores may range in grade from about 0.6 to 2 percent copper, with
the bulk of them being below 1 percent.
(e) Native copper ores. The only important deposits containing
native copper ores occur in Michigan. As mined, they contain about 1
percent of copper, generally associated with some silver.
(f) In a few locations, and generally on a small scale, copper may
occur largely or altogether in oxide minerals, such as chrysocolla,
malachite, or azurite. Such ores are generally treated by leaching
methods.
21
-------
(g) Ores of other metals, notably lead and zinc, but also silver,
molybdenum, tungsten, and iron, may yield copper as a by-product. As
discussed previously, little emphasis will be placed on such ores or
their treatment in this section of the report.
Most of the copper sulfide ores mined in the United States contain
gold and silver in small but economically significant concentrations.
These are recoverable as such in the conventional smelting-electrolytic
refining process, but not in most processes involving leaching or fire
refining. The sulfide copper ores often contain molybdenum in impor-
tant amounts which can be readily recovered in normal concentration
operations.
Major Process Operations
Figure 3 is a graphic description of the basic copper process which
accounts for virtually all of the copper produced in the United States.
About 80 percent of the copper ores mined in the country go through all
stages of the process shown, from mining and concentration through
electrolytic refining. Amost all of the remaining 20 percent of the
copper in ores treated by other processes is merely preconcentrated by
these other processes and at some stage or other is introduced into the
basic process for eventual conversion into refined copper. Minor
amounts of primary copper are produced strictly by hydrometallurgical
methods.
Figure 4 is a flowsheet of the primary copper industry from ore to re-
fined copper. The figure shows the various types of ores and the pro-
cessing or handling steps that are involved in treating the variety of
ore grades and types encountered and illustrates the relationship of
these auxiliary processes to the basic process.
In the figure, the materials and processing steps are numbered. These
numbers are keyed in with the subsequent text in which pertinent ex-
planations and details are presented.
As shown in Figure 4, there are seven indicated routes (Roman numerals')
to copper production, depending largely on the grade and type of ore.
Route I, which accounts for the production of over 90 percent of the
primary copper produced in the United States, is shown at the extreme
left side of the figure. This is the basic process of concentration,
smelting, and refining depicted in Figure 3. In most of the other
routes, some product, at some stage in the process, will be sent to
some stage of the basic process.
22
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MINING
-J^
r$$i^
i Ut>*ilT
Blasting
The ore body is broken up by
blasting.
MILLING
>kr«*"
Looding
The ore. averaging about 1 per-
cent copper, is loaded into ore
cars by electric shovels
Hauling
The cars ol ore are hauled to
the mill
on-
^/
'^M^
Crushing
The ore is crushed to pieces
the size ol walnuts
Grinding
The crushed ore is ground to
a powder
SMELTING
^.jsgj.^--.
V CONC(N'»»'II
Concentrating
The mineral bearing particles
in the powdered ore are con-
centrated
The copper concentrates (av
eragmg about 30 percent
copper) are roasted to remove
sulfur
REFINING
_^I 6USH8
Reverberator/ Furnace
The roasted concentrate is
smelted and a matte, contain
mg 32 42 percent copper, is
produced
SUG
Converter
The matte is converted into
blister copper with a purity of
about 99 percent
r-~^
Refining Furnace
Blister copper is treated in a
refining furnace.'
FABRICATING
Electrolytic Refining
Copper requiring further treat-
ment is sent to the electrolytic
refinery."
iff lire rtttieij cooper meeij ihe speci
! 01 llbi««tois, i! 'I bSfd without turret
urt^t ((fined (ectiolft'Ci'lr when
Ihe x*cul (nopel'ei ot e>c'-:'jt'( coo«( tit
ifQutfd (j »n*n the ces;e' ii lo te uied lor
eleclMcil co'iiuclars, iidiO' «he" P'eticus
rtljli nt cveyf-t in s^'ke-t QU^!I!I« to
HIFIHfD COPPER','
Rolling
Drawing
fire refined or electrolytic copper and/or brass
13 mixture of copper and zmci is made into
sheets, tubes rods and wire
Exiruding
Sheets, tubes, rods and wire are further fabricated
into the copper articles you see :n everyday use
FIGURE 3. BASIC STEPSCOPPER ORE TO FINISHED PRODUCT
(A)
23
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Mining
There are three principal mining methods used in the copper industry:
open pit, caving^ and stoping.
The choice of mining method depends on many factors, mostly associated
one way or the other with economics. They include such considerations
as size, configuration, and grade of the ore body, its depth below the
surface, topography, climate, accessibility, etc.
Open Pit. Open-pit mining, as its name implies, involves the removal
of ore from deposits at or near the surface, by a cycle of operations
consisting of drilling blast holes, blasting the ore, loading the
broken ore onto railroad cars or trucks, and transporting it to the
concentrators. In a few cases, blasting is not required and the ore
is "ripped" by bulldozers prior to loading. Before a given deposit
can be worked by open-pit methods, barren surface rock or earth must
be removed to uncover the ore body. This can be a major operation.
In a mine now being developed in Arizona, as much as 500 feet or more
of overburden was removed before systematic mining could be begun.
In addition to the barren overburden or waste rock, it is also neces-
sary, in open-pit mining, to remove low-grade copper ores from the top
or throughout the ore body so that mining of the concentrating-grade
ore can proceed systematically and with optimum efficiency. As a rule
of thumb, concentrating-grade ore in large open-pit mines contains
about 0.6 to 1 percent of copper in a form recoverable by concentration.
Significant proportions of low-grade ore, containing from 0.2 to about
0.5 percent of copper are encountered in open-pit mining. Such low-
grade ore, containing oxide and/or sulfide copper minerals, cannot be
economically treated in a concentrator. Where in earlier years this
low-grade ore was discarded to waste, it is now hauled from the mine,
piled in large heaps, and subjected to leaching.
The open-pit mines in the United States account for about 80 percent of
the country's copper production. Most are quite large, yielding from
about 15,000 to over 100,000 tons of ore per day.
Haulage from open-pit mines to concentrators is done by truck or rail.
In some cases, conveyor belts are used to transport the ore from the
pit to the concentrator. Tractor-drawn scrapers are occasionally used.
Underground Mining Methods. About: 20 percent of the copper ore mined
in the United States comes from some form of underground operation.
The most widely practiced of these is the "caving" technique in which
an ore body is undercut, and the ore under natural stressess fractures.
caves, and falls thrcugh vertical or nearly vertical tunnels called
"raises" to loading chutes on a still lower level in the mine. Block
caving is an efficient, low-cost operation, but requires homogeneous,
rather weak ore bodies of more or less regular configuration for
25
-------
optimum results. Several large copper mines in the United States use
this form of underground mining.
Another important underground mining method is the so-called "top
slicing". In this method a cut is made on the top section of the ore
body and the ore removed. During mining the floor of the "slice" is
covered x^ith a timber "mat". After the "slice" is mined out, the roof
(overburden) is blasted and falls on the timbered floor. As successive
slices are mined and caved, the overburden subsides, filling the spaces
formerly occupied by ore. The mat of timbers effectively separates ore
from the waste overburden and results in a less diluted ore. Although
more expensive than block caving, it is preferred for the higher grade
underground ore bodies.
Stoping, another method of underground mining, is the working or ex-
cavating of a deposit by methods which do not involve natural caving
or caving induced by blasting. In this method, the ore body is exca-
vated directly and the void space left after removal of the ore'is
filled by waste rock, tailings, etc., or the walls and roof are sup-
ported by timbering or bolting.
There are a number of stoping methods, each best adapted to the size,
grade, and configuration of the ore deposit, the type and strength of
the rock enclosing the deposit, etc.
The Basic Copper Production Process
(Items 3-16 in Figure 4, Route 1. See also Figure 3)
As previously pointed out, about 80 percent of the copper produced in
United States mines is treated by the basic concentration-smeltirig-
refining route. This involves processing copper ore containing less
than 1 percent of copper which is concentrated by flotation to produce
a copper concentrate containing 15 to 25 percent or perhaps more copper.
These copper concentrates are smelted and refined to produce various
shapes and grades of intermediate copper products.
The lineup of operations and materials showu in Column I of Figure 4
illustrates this basic process. The following sections, numerically
keyed to the diagram in Figure 4, discuss these (Nos. 3 through 16).
Concentrating-Grade Ore (3)
Concentrating-grade ore is ore containing from about 0.6 to as much as
2 percent of copper present as sulfide minerals which can be concen-
trated by the process later to be described. This type of ore is pro-
duced mainly by the large open-pit copper mines in the west and south-
west. It consists largely of such copper minerals as chalcopyrite,
26
-------
chalcocite, covellite, and bornite in various combinations. Some con-
centrating- grade sulfide ores also may contain sulfoarsenide minerals
such as enargite and tennantite. Many of the concentrating-grade
sulfide ores also contain molybdenum, gold, silver, selenium, and
tellurium as economically important by-products.
The nonore minerals associated with these ores, virtually valueless,
cover a gamut, depending on their geological origin, and include
quartz (silicon dioxide), various igneous and sedimentary silicates
(feldspars, altered feldspars, clay, shale), limestone, and pyrite.
These nonmetallic minerals will amount in most cases to over 95 per-
cent by weight of the ores.
Concentration (4)
In the United States today, concentration of copper is done by flota-
tion. The concentration of copper ores is carried out in what are
commonly called "mills". The purpose of "milling" is to recover most
of the copper in the ore in a concentrate form. The basis of the flo-
tation process by which this "concentrating" is done is as follows:
Copper sulfide minerals (and other sulfide minerals as well) possess
certain surface characteristics that are quite different from those
of the other minerals (silicates, quartz, etc.) present in the ore.
When added to a suspension of the finely ground ore in water in small
quantities, certain reagents can be made to attach themselves selec-
tively to the surfaces of sulfide minerals. These reagents are of
such a nature that, after the attachment, the surfaces of the sulfides
are nonwettable and are said to be "hydrophobic" and "aerophilic". If
a mixture of such sulfide and oxide minerals in water is properly
treated with such reagents, and small bubbles of air are passed through
the mixture, these nonwettable or aerophilic particles attach them-
selves to the air bubbles and, even though much denser than water, are
levitated or "ballooned" to the surface.
The flotation process for copper has been developed to take full ad-
vantage of this phenomenon. It involves these basic steps.
(a) Crushing and grinding of the ore to "liberate" the copper minerals,
which, in the ore itself as mined, may be completely or partially en-
capsulated by noncopper minerals. Crushing is done, in stages, by
passing the ore through large jaw crushers and finally through gyra-
tory or cone crushers. As mined the ore may contain boulders up to 6
feet in size. Crushing produces a final size of about 1 to 2 inches.
The crushed product is then wet ground in ball mills or rod mills.
These are huge cylinders containing steel balls or rods, which rotate
on a horizontal axis. The crushed ore is fed with water into one end
of the rotating mill, where the action of the tumbling balls or rods
27
-------
breaks up or "grinds" the crushed ore to the particle size required to
"liberate" and/or reduce the size of the copper minerals.
The required size for flotation is always less than 35 mesh (0.016-
inch diameter) which is about the maximum that can subsequently be
floated effectively. Grinding to finer sizes is frequently required
to obtain adequate liberation of minerals from gangue rock.
(b) Flotation. The ground mixture of ore and water is then treated
with the required chemicals and passed continuously through a series of
flotation cells. In copper metallurgy the required chemicals are those
necessary to maintain proper pH conditions, the flotation reagents
themselves which will selectively coat the copper sulfide particles,
and a frothing reagent which will stabilize the froth which is formed
by the passage of air bubbles to the surface. The flotation cells are
rectangular tanks of various sizes and design. They are fitted with
agitators which serve the dual purpose of keeping the ore particles in
suspension and providing a constant supply of fine air bubbles to the
cell.
As the mixture of ore and water, properly dosed, passes through a
series of cells, the air bubbles collect the selectively coated copper
sulfide minerals and transfer them to the froth. The froth which is
being constantly produced is continuously scraped by mechanically oper-
ated paddles, and constitutes the "concentrate". The nonfloating
material in the slurry constitutes the "tailings" and is discharged
from the bottom of the cell. The above is a simplification to illus-
trate the manner in which one flotation cell would work.
In actual practices, copper mills employ literally scores, if not
hundreds, of flotation cells with a wide variety of flow patterns.
Figure 5, the flowsheet of an Arizona copper mill, illustrates the
complexity often encountered in flotation processing.
According to this flowsheet, the crushed ore is fed to rod mills which
effect partial grinding and then to ball mills where the required de-
gree of fineness is achieved. The ball-mill circuit is closed with
wet cyclones. The cyclones are so designed that adequately ground ore
is removed from the grinding circuit and sent on to flotation and in-
sufficiently ground ore is returned to the ball mills. This avoids
overgrinding and its consequent deleterious effect on the selectivity
of flotation and concentrate grade. After grinding, the slurry of ore
and water goes to the first stage, or rougher, flotation. Here the
objective is not to obtain a high-grade finished concentrate but
rather to obtain a high recovery of copper in a relatively low-grade
concentrate. As shown in the diagram, the rougher flotation section
in this example consists of 16 banks with 12 cells per bank. The
usual practice is to feed the ore slurry into the first cell in a bank
and to discharge tailings from the last cell. By this arrangement, the
underflow from the first to the eleventh cell progresses to the next
28
-------
Ore Irom m
-------
cell in line, where additional time and reagent addition results in an
additional copper recovery.
The combined rougher concentrates produced by scraping the froth from
the rougher cells are sent to hydroseparators which effect partial
separation of solids, and then to a large thickener which completes the
solid separation and produces a virtually clear overflow. The solids
from the hydroseparators are treated in cyclones to remove any oversize
particles that may be present. These are reground in "regrind ball
mills". The overflow from the cyclones in this regrind circuit joins
the underflow from the thickeners and constitutes all of the rougher
concentrate containing all of the recoverable copper. The rougher
concentrate, which may contain about 81 percent copper, is then re-
treated in the "cleaner" cells. In the "cleaning" operation, generally
no additional flotation reagent is required, enough remaining on the
sulfide particles to ensure that they will remain floatable. The ob-
jective of the cleaning operation is to drop out the noncopper minerals
by passing the pulp through banks of flotation cells similar to those
used in the rougher operation. Concentrates from the cleaner cells are
recleaned in a smaller but similar "recleaner" circuit. Since in re-
cleaning some concentrate may be lost with the underflow, provisions
are made to return this to the middlings thickener.
In the flowsheet shown in Figure 5, the tailings from the "cleaner"
circuit are "scavenged", i.e., treated with additional reagent to re-
cover floatable copper that escaped in the cleaning operation. Since
much of the copper recovered in the scavenging operation will not be
sufficiently liberated from gangue material, the scavenger tailings
are returned to regrinding and subsequent cleaning and recleaning.
The copper concentrates produced are thickened and filtered, loaded in-
to railroad cars, and sent to the smelter. The tailings produced in
the operation, consisting of the underflows from the rougher and sca-
venging circuits, are thickened. Clear overflow from the thickener
flows to a reservoir for recirculation. The thickened tailings are
ponded and clear supernatant liquor from the pond is recirculated
through the reservoir. Also recirculated for reuse through the reser-
voir is the overflow from the middlings thickener.
The flowsheet in Figure 5 should be regarded only as typical. Other
concentrators employ other types of equipment and different flow pat-
terns. Some will use spiral classifiers in place of cyclones to "close"
grinding circuits. Some may not employ recleaning or scavenging. Ore
tonnages, grade, and flotation characteristics of the minerals will
dictate the number, size, and hookup of flotation cells, thickeners,
etc. In regions where water is plentiful, recirculation of water will
not necessarily be practiced.
30
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Copper Concentrates (5)
The concentrates produced by milling operations are always in the form
of filter cake containing about 30 percent by weight of water and, on
a dry basis, from about 20 to 35 percent copper. The copper content
and chemical composition of concentrates will vary from mine to mine.
The following tabulation lists some typical compositions of concen-
trates produced in Arizona mills.
Typical Copper Concentrate Compositions,
percent
Cu
24.8
31.3
23.3
15.4
Fe
23.0
22.7
24.2
31.0
S
33.0
29.8
31.0
36.9
SiO?
12.9
14.5
9.4
11.0
CaO
_ _
--
2.4
--
Al2_°3
2.5
0.1
2.4
4.3
Molybdenum Recovery (6)
Many of the copper ores in the West contain molybdenum in important
quantity and this valuable by-product is often recovered in concentra-
tors. The recovery of molybdenum is described in Volume II of this re-
port. Molybdenum in copper ores occurs as molybdenite, which is a
molybdenum sulfide (MoS2). This mineral is what is known as a "natural"
floater and, in pure form, is so water repellant that it can be made to
float in the flotation process without the use of reagents. It accom-
panies the copper minerals during flotation, and up to about 80 percent
of the molybdenum present in the ore will report with the copper concen-
trates under normal conditions. In one method of molybdenum recovery,
the copper concentrates are treated with lime and steam to destroy the
flotation reagents which still adhere to them. The concentrates are
then treated by flotation without reagents. The naturally floating
molybdenum is collected, along with some copper, in a rougher concen-
trate, which is subsequently cleaned and recleaned. In another pro-
cess, the molybdenum is separated from the copper rougher concentrates
by selective flotation, cleaning, and recleaning.
Smelting (7-8-9-10)
There are relatively few smelters in the United States and only about
a dozen large mills are located close to smelters. The larger copper
companies which operate mines, concentrators, and smelters ship con-
centrates to their own facilities. These smelters are more or less
centrally located in the area of the company's mining interests and
only relatively short hauls by rail or truck are necessary. Companies
31
-------
engaged only in mining and concentrating ship their concentrates to
smelters owned by other companies.
In most cases the transportation is intrastate, but in some instances,
long rail or truck hauls may be involved. Some Arizona copper mines
have shipped concentrates to El Paso, Texas.
The overall purpose of smelting is to separate copper from the iron,
sulfur, and gangue materials that are present in copper concentrates.
The steps required include roasting, matte smelting, and converting.
Roasting (7)
Formerly, most copper concentrates were roasted prior to smelting in
the reverberatory furnace. The purposes of roasting are to dry the
concentrates, eliminate sulfur to a controlled degree, and remove cer-
tain impurities such as arsenic, antimony, and selenium. At present
the roasting step is frequently bypassed and new smelters have been
erected which do not use roasters.
Where roasters are bypassed, drying may be accomplished by air drying
or the use of mechanical equipment such as kilns. In some plants the
roasting furnaces have been converted to simple driers. The elimina-
tion of sulfur in roasting is employed to control the sulfur and hence
the copper concentration in the matte. Some modern concentrators are
now able to control the sulfur concentration in concentrates, so that
roasting is unnecessary. Many ores do not contain arsenic and antimony
in quantities that must be removed prior to smelting.
In current practice, roasting is carried out either in multiple hearth
furnaces or in fluidized-bed roasters. The former is a large cylindri-
cal device up to 40 feet in height and about 20 feet in diameter con-
taining 8 to 12 circular hearths, one above the other. A shaft carry-
ing "rabble" arms for each hearth extends the length of the axis. As
the shaft is made to rotate, "shoes" affixed to the rabble arms rake
or move the concentrates over the surface of the hearth to ports dis-
charging to the next lower hearth. These discharge ports are alter-
nately near the circumference and at the center of the hearth, and the
shoes or rakes are so set that the concentrates are moved toward the
discharge port. In operating such roasters, the concentrate is charged
on the top hearth and progresses downward from hearth to hearth. When
additional heat is required, fuel is injected into the bottom hearth
countercurrent to the flow of hot gases. Heat is provided by igniting
the concentrates and passing air through the furnace. When the concen-
trates contain sufficient sulfur (25 percent plus), no additional fuel
is required. Temperatures in this type of roaster will range from
about 300 to 400 F on the top hearth to as much as 1600 to 1700 F on
the bottom hearth.
32
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Matte Smelting (8)
Matte smelting is universally done in reverberatory furnaces. These
are large roomlike structures, up to about 130 feet long and 38 feet
wide, with capacities for treating up to 1500 tons of charge per day,
on a continuous basis.
Figure 6 is a diagrammatic sketch illustrating the general structural
and operational features of a reverberatory furnace.
The main objective of treatment in a reverberatory furnace or "matte"
smelting is to collect virtually all of the copper in a molten copper-
iron-sulfide material, called "matte", suitable for subsequent treat-
ment in converters.
The feed material to the reverberatory furnace consists of the follow-
ing:
(a) Copper concentrates either roasted or unroasted
(b} Possibly some high-grade copper ore
(c) Possibly some cement copper
(d) Converter slag
(e) Reagglomerated dust from smelting operations
(f) Fluxing materials, such as limestone, etc., required to form a
fluid slag from silicates that may be introduced with the charge.
This feed material is added to the furnace at one end through charging
ports on the roof near the side walls. As the charge enters the fur-
nace, it collects in piles alongside the furnace walls still in a solid
state. Heat is supplied by the burning of powdered coal, oil, or gas
through burners placed in the end wall near the charging end of the
furnace. The entire interior of the furnace is heated to temperatures
ranging from about 2000 to 3000 F. Radiation of this heat from the
roof and walls of the furnace causes the charge to melt and form two
separate liquid phases on the bottom a matte containing almost all of
the recoverable copper as a sulfide and on top a liquid slag composed
of silicates of iron and other metals introduced with the charge
material. A small proportion of the total copper in the charge is also
present in the slag.
The molten matte and slag flow to the discharge end of the furnace and
are tapped off into ladles. The slag is discarded to waste, and the
matte, still liquid, is charged into converters.
33
-------
-* [; E o
cp ~H
.u
' ,«?>1
<; vf
\ HI
\\ '
T n r?
A- y
TT^
j
t?
v; csj
I %
S "»-, 0>.
-------
Typical compositions of matte and
tabulation^):
;lag are shown in the following
Cu, percent
Fe, percent
S, percent
SiC>2, percent
CaO, percent
A^O^, percent
1
Ma 11 e s
4
45.0 43.4 31.6 27.2
23.0 22.7 35.0 40.7
24.0 25.0 25.0 24.0
Slags
0.6 0.43 0.40
42,5 34.5 35.0
4
0.34
40.7
34.5 39.2 37.0 37.6
5.8 6.8 5.0 3.5
6.3 4.8 7.2 9.6
Converting (9)
In this step, the matte produced in the reverberatory furnace is con-
verted to a relatively impure form of copper called blister copper by
an oxidation process involving the blowing of thin streams of air
through the molten material. The operation is done in two stages. In
the first, the iron sulfide component of the matte is oxidized to sul-
fur dioxide and iron oxide, leaving nearly all of the copper as molten
copper sulfide or "white metal". Sulfur dioxide leaves the converter
as a gas and may be processed to sulfuric acid. Fumes and dusts which
may contain lead, bismuth, arsenic, etc., produced by converting are
collected and sent to lead smelters. The iron oxide which is formed
during coaverting reacts with silica, added as a flux to the converting
operation, to form a molten iron silicate slag. The reaction occurring
in the first stage of converting is approximately
CuFeS2 + 0? + SiO ->
Matte Added Silica
as air added
for
fluxing
"-' o T
Sulfur
dioxide
to s u 1 -
furic acid
manufacture
Iron
silicate
slag
removed
and
Cuprous
sulfide
(wbIte
metal)
recirculated
When this reaction is completed, the slag is removed from the converter.
Because converter slag normally contains significant concentrations of
copper, it is recirculated to the matte-smeItine step in the reverber-
atory furnace.W After removal of the converter slag, the blowing
with air is continued until virtually all of the remaining sulfur is
oxidized and removed as sulfur dioxide, in accordance with the overall
equation
Cu2S + 0 > Cu + SO .
The converted copper is still relatively impure, containing varying
35
-------
amounts of heavy metals, arsenic, and sulfur. It also contains virtu-
ally all of the precious metals originally present in the concentrates.
This "converted" copper may be transferred while still molten to re-
fining furnaces. Some, however, may be cast into pigs prior to subse-
quent refining. When cast it exhibits a typically rough upper surface
upon solidification, owing to the expulsion of gases, mainly air and
sulfur dioxide. This rough surface accounts for the name "blister"
copper.
Converting is done i~ large horizontal cylindrical furnaces up to 30
feet in length and about 12 feet in diameter, with a centrally located
aperture for charging, unloading, and exit of gases. The furnaces are
also fitted with numerous tuyeres along their length near the bottom
through which the air necessary for converting is blown. The converters
are mounted on trunnions so that they may be tilted. During the con-
verting operation the furnaces are run in an upright position with the
aperture directly beneath a hood through which the converting gases are
exhausted. They are tilted or tipped for charging, discharging, and
inspection.
Typical analyses of "blister" copper produced by converting are
shown in the following tabulation:
S am-
ple
1
2
Percent
Cu
98.4
98.8
As
0.02
0.10
Sb
0. 178
0.04
Pb
0.001
0.15
Ni
0.005
0.05
Zn
0.003
0.12
Fe
0.13
0.25
S
0.20
0.17
A "
Au
0.295
0.31
Ag
111.
30.
V
9
25
3 99.5 0.035 0.015 0.001 0.04 0.002 0.030 0.06 0.02 2.50
* Ounces per ton.
Fire Refining (11)
Fire refining is a high-temperature furnacing process, involving oxida-
tion, slagging, and reduction, by which "blister" copper is further
purified to produce either a final product--"fire-refined" copperor
anodes for subsequent electrolytic refining. "Fire-refined" copper,
i.e., copper which has been refined by this furnace process, constitutes
only a relatively minor percentage of the total production of refined
copper. Most of the primary U, S. copper is further purified by elec-
trolytic refining.
Fire refining to produce electrodes for subsequent electrolytic refining
is referred to as "anode furnace refining". The purpose of anode fur-
nace refining is to purify and degas the blister copper so that the
anodes produced will be physically and chemically acceptable for elec-
trolysis. This refining removes the large amounts of cuprous oxide and
other impurities, which, if allowed to remain in the anodes, would
segregate and weaken their structure and result in rough surfaces. As
36
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is discussed under Electrolytic Refining (13), weak, nonhomogeneous
anodes with nonflat surfaces complicate that operation.
Anode furnace refining is done in two kinds of furnaces:
(1) Reverberatory furnaces somewhat similar to, but considerably
smaller than, the reverberatory furnaces used in matte smelting. Such
furnaces are used for refining solidified blister copper that has been
previously cast into pigs and remelting high-grade scrap.
(2) Cylindrical furnaces similar to but smaller than the previously
described converters. These furnaces can be tilted on trunnions in
somewhat the same fashion as converters and are designed to accept
molten charges of blister copper directly from converters. Refining
furnaces are fueled by coal, natural gas, or oil.
The operation is carried out by introducing air beneath the molten
metal surface to oxidize part of the copper to cuprous oxide. The
cuprous oxide, soluble in molten copper, reacts with zinc, tin, or
iron present in the copper to form sulfur dioxide and metal oxides.
Sulfur dioxide passes off as a gas and the oxides of iron, zinc, and
tin, with added silica, form a slag that can be removed by skimming.
Other impurities, such as lead, arsenic, and antimony, are only parti-
ally removed by this method. To remove large amounts of these impuri-
ties it is necessary to employ basic fluxes such as lime or soda in-
stead of silica. Other metals that may be present in the copper, such
as gold, silver, and nickel, and some of the metalloids, selenium and
tellurium, resist this treatment and remain with the copper.
After oxidation and slag removal are accomplished, the copper will con-
tain considerable oxygen as cuprous oxide which must be removed before
casting anodes. Deoxidation is done traditionally by the addition of
coke and by "poling". Poling amounts to inserting rather large poles
of green hardwood beneath the surface of the copper and allowing them
to decompose into reducing gases. These gases reduce the cuprous
oxide to copper metal. An alternative for reducing the copper oxide
is to pass reformed natural gas through the bath of molten copper.
After reduction, if electrolytic refining is to follow, the fire-re-
refined copper is cast into anodes. If fire-refined copper is the de-
sired end product, it is cast into various "refinery" shapes, such as
wire bar, billets, cakes, and ingots.
Fire-Refined Copper (lla)
Fire-refined copper must conform to fairly stringent requirements. The
wire bar, cakes, and billets to meet specifications must contain a min-
imum of 99.5 to 99.85 percent copper, depending on their subsequent use,
Any silver present may be counted as copper. Metal and metalloid
37
-------
impurities, such as arsenic, nickel, bismuth, lead, etc., are also
rigidly specified.
Anodes, which represent the major tonnage of fire-refined copper, are
cast into rectangular shapes of various sizes and weights, depending
on specific refinery practice. A typical anode will be about 36 to 39
inches wide, 35 to 38 inches long, and about 1.25 to 1.5 inches thick,
weigh from about 500 to 700 pounds, and contain about 99.5 percent of
copper.
Electrolytic Refining (13)
The purpose of electrolytic refining is to produce a very high purity
copper, and, at the same time, to separate from the copper and recover
the valuable metal and metalloid impurities that are present in the
"fire-refined", or more properly, "anode-furnace-refined" copper.
Typical compositional ranges of anodes are shown in the following tabu-
lation:
Percent
Cu 98.5 to 99.6
Pb 0.01 to 0.32
Fe 0.01 to 0.06
Bi 0.0001 to 0.009
As 0.004 to 0.32
Sb 0.002 to 0.24
Ni 0.002 to 0.60
Se 0.01 to 0.20
Te 0.006 to 0.08
Au 0.01 to 3.5 oz/ton
Ag 2.4 to 9.1 oz/ton
Oxygen 0.1 to 0.3
Electrolytic refining is carried out in the following manner: the
electrolytic tanks are rectangular vessels about 10 to 15 feet long,
about 3 to 3-1/2 feet wide, and about 3-1/2 to 4 feet deep. They are
usually constructed of concrete lined with lead or antimonial lead.
Each tank is fitted on the top, along each side, with heavy copper
supporting bars for carrying the anodes (and cathodes) and through
which the electric current flows. The anodes, cast with lugs or hooks,
are hung on these bars at regulated spaces of about 3.5 to 4.5 inches
apart. The cathodes are thin sheets of electrolytically refined copper
1 to 2 inches longer and wider than the anodes, but only about 0.025-
inch thick and weighing approximately 8 to 10 pounds. These cathodes,
called "starting" sheets, are made in a separate section of the re-
finery. The starting sheets are fitted with loops of copper at their
38
-------
tops, through which copper bars are inserted. The copper bars serve
the same function as the lugs on the cathode. When loaded, the elec-
trolytic tank will contain from 30 to 40 cathodes and anodes, depending
on the length of the cell and the spacing.
These anodes and cathodes are immersed for virtually all of their height
in the electrolyte, a dilute solution of sulfuric acid and copper sul-
fate. Typical electrolytes contain about 40 grams per liter of copper
and up to about 200 grams per liter of free sulfuric acid. When loaded,
the cell is returned to operation in the refining circuit. The entire
refining circuit may consist of hundreds of such tanks, arranged in
optimum configurations for the circulation of the electrolyte from tank
to tank, the electrical current flow, and the systematic operation of
the refinery.
When electricity is flowing, the current first traverses the anodes,
where copper and impurities such as nickel, arsenic, and iron are elec-
trolytically oxidized and put into solution. Gold, silver, selenium,
and tellurium do not oxidize and go into solution, but enter the elec-
trolyte as finely divided solids called slimes which may settle to the
bottom of the tank.
The current then traverses the electrolyte to the cathode where reduc-
tion or deposition of the copper takes place. This is a highly selec-
tive operation wherein soluble impurities such as nickel, arsenic,
bismuth, and antimony are not deposited with the copper so long as
they are kept below certain concentrations in solution. Operating con-
ditions and schedules for preventing or minimizing the effect of such
interferers have been thoroughly developed.
As the copper is deposited, the current traverses the cathode and en-
ters a so-called support bar which connects to the next tank in line
and which becomes the anode bar for that tank. Here, the same process
is repeated and rerepeated through the entire circuit.
After operation for about 2 weeks, by which time the cathode of pure
copper has grown to a weight of about 40 percent of that of the origi-
nal anode, the cathode is removed and washed and new starting sheets
are inserted on each side of the anode. The electrolysis is resumed
for another period of about 2 weeks to produce a second cathode of
approximately the same weight. The anode, now 80 to 90 percent con-
sumed, is removed from the tank, and sent back to the anode refining
section of the plant where it is remelted to be recast into full-sized
anodes. Complete or nearly complete dissolution of anodes in the elec-
trolytic process is impractical, and would lead to mechanical failures
of the metal, with attendant short circuiting.
Periodically, the slimes which have accumulated in the tanks and which
contain precious metals and other values are removed and are sent to
refining operations for the recovery of these values.
39
-------
During electrolytic refining, there is a continuing buildup of soluble
impurities such as nickel, arsenic, antimony, and bismuth in solution.
Control of the concentration of these impurities to tolerable limits
is achieved by either continuously or intermittently withdrawing a por-
tion of the electrolyte from the circuit and replacing it with fresh
solution.
Various methods are employed for the recovery of copper and other
values in this withdrawn or "bleed-off" solution. In the most widely
practiced method, the copper in the bleed-off solution is electrolytic-
ally stripped by passing the solution through cells, similar to the
refining cells but employing insoluble (e.g., iron) anodes in place of
copper anodes. In this operation, where there is no copper anode to
continuously supply copper to the solution, the effect of the electro-
lysis is to drive the copper out of solution as copper metal onto a
cathode. This stripping, considerably less electrically efficient than
the refining electrolysis, may be done in stages, with the first stage
recovering most of the copper as a refined grade, the second as an in-
termediate grade suitable for casting or alloying, and the final as a
low-grade copper-arsenic sludge which requires subsequent treatment.
After such stripping, the soluble impurities remaining are crystallized
by evaporation, the salts (such as nickel sulfate) recovered, and the
acid returned to the refining circuit.
By-Product Slimes (13a)
An important by-product of electrolytic refining is the slimes, or
anode mud, which accumulate in the system and which may amount to as
much as 1 to 15 percent by weight of the original anode. This mud con-
tains virtually all of the lead, selenium, tellurium, gold, and silver
originally present in the anode, a major portion of the bismuth and
antimony, and a lesser proportion of the arsenic.
Typical ranges of anode mud analyses are shown in the following tabu-
Copper
Percent
10-55
Gold
Silver
Lead
Antimony
Sulfur
Selenium
Te llurium
Iron
Arsenic
Nickel
--
--
< .1-15
< .1-16
< .1-2
< .1-8
1-12
0.1-0.2
0.1-7.3
0.1-0.9
Oz/Ton
0.1-500
25-10,000
Remarks
Can be largely removed by screen-
ing (to about 1 to 2%)
40
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Cathode Copper (14)
The following tabulation shows the compositional range of electrolytic
cathodes:
Element Percent
Copper 99.95
Arsenic 0.0001 to 0.001
Antimony 0.002 to 0.001
Bismuth 0.00001 to 0.0002
Lead 0.0002 to 0.001
Nickel 0.0001 to 0.002
Selenium 0.0003 to 0.001
Tellurium 0.0001 to 0.0009
Gold 0.002 to 0.01 oz/ton
Silver 0.05 to 0.5 oz/ton
In addition, hydrogen, resulting from the electrolysis, may be present
in small amount.
Some electrolytic cathodes may be sold as such, but by far the bulk of
the cathodes produced are melted, given a form of fire refining, and
cast into various refinery shapes.
Melting and Casting (15)
This operation is generally referred to as cathode refining and casting,
Strictly speaking, it is not refining because, in many cases, the re-
sultant product contains slightly less copper than the cathodes them-
selves. Its main purpose is to produce the conventional cast refinery
shapes of refined copper. In melting the cathodes to accomplish this,
sulfur and oxygen are introduced from air, coke, etc., with which the
copper comes in contact. Preparatory to casting, these must be re-
moved and the oxygen controlled to produce forms that are physically
and chemically acceptable by established commercial standards.
The operation is carried out in either reverberatory furnaces of the
type used in fire refining (11) or in electric furnaces. When electric
furnaces are used, the refining operations to remove sulfur and control
oxygen are not required.
Refinery Shapes (16)
Electrolytic copper after melting is cast into wire bar, billets, and
cakes as is fire-refined copper, to meet established commercial
41
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standards for physical, chemical, and electrical properties.
Leaching-Grade Ores (17) (Route II, Figure 4)
The major tonnage of leaching-grade ores fall into two categories:
(1) Ores containing sulfide minerals which are so fractured or shat-
tered in texture that air and water can be made to percolate through
them as they exist in the ground. Ore bodies of this type occur at
several places in Arizona and Tennessee. As will be discussed later,
such ores may be profitably leached "in place" without being mined.
(2) Ores containing either sulfides or mixtures of sulfides and oxide
minerals which are too low in grade to justify the cost of milling.
Such ores are generally encountered in large quantities in open-pit
mining in the process of uncovering concentrating-grade ores. These
ores, which must be removed in any event to permit systematic mining
of concentrating-grade ores, are generally treated by heap leaching.
Other types of leaching-grade ores exist which are most profitably
treated by specialized methods. These are discussed in later sections
of this report.
Heap Leaching and Leaching in Place (18)
Heap leaching consists in piling or dumping low-grade sulfide or mixed
sulfide ore rocks in large piles, in natural basins or specially pre-
pared areas with good drainage characteristics, and circulating dilute
acidic solutions through them. In the most improved form of heap leach-
ing, the site of the heap is carefully prepared to insure that the
leach solution, after it has traversed the pile, will not be lost by
permeating into the ground, but will all be contained in the system for
subsequent copper recovery. In most cases, the ore as mined, which may
range from boulder size to a few inches, constitutes the feed to the
heap. In some cases, preliminary crushing or secondary blasting may
be used to reduce large boulders so as to permit better contact of the
copper minerals with the leach solution. When a heap or "dump" is
prepared, a leach solution, essentially dilute sulfuric acid, is dis-
tributed evenly over the top of the pile and allowed to trickle through
the pile. Such heaps may occupy scores of hundreds of acres and be
scores of feet in depth. After traversing the pile, the leach solution
drains to prepared reservoirs at the base of the pile, from where it is
recirculated by pumping back to the distributing system on top of the
pile and to subsequent copper-recovery operations.
In heap leaching, the oxide copper minerals dissolve directly in the
leach solution according to the equation
42
-------
CuO + H0SO. -> CuSO. + H00.
24 42
The dissolution of sulfide minerals, however, is slower and indirect
and requires oxidation.
CuS + SO >
One method of achieving such oxidation is to sectionalize the heap and
allow selected sections to drain completely and to aspirate air, while
other sections are still being leached. After a period during which
the air has oxidized the copper sulfides to copper sulfate, leaching in
the section is resumed, with the consequent dissolution of copper
sulfate.
The presence of ferric sulfate in the leach solution also promotes the
oxidation of copper sulfides, by a reaction similar to
4Fe_(SO,)0 + CuS + 4H00 -> SFeSO. + CuSO. + 4H_SO. .
243 2 4 424
In recent years, patents have been issued covering the employment of
bacteria as leaching aides. These bacteria are said to produce both
ferric sulfate and sulfuric acid, which are the effective leaching
media.
Pregnant solutions from the heap- leaching operations, containing up to
several grams per liter of copper as copper sulfate, are then treated
by various copper- recovery processes such as cementation or solvent
extraction. These are discussed later.
Leaching in place is generally done on depleted ore bodies or on ore
bodies too low in grade to justify conventional mining. An essential
to leaching- in- place methods is a wel 1- fractured deposit, which will
permit the permeation of leach solution and air. Leaching in place
is done by the intermittent circulation of leach solution and air
through such deposits, the air being required for oxidation of sulfide
minerals. The pregnant solution from in-place leaching drains into
tunnels cut beneath the deposit and is pumped to the surface for copper-
recovery treatment, generally by cementation.
Percolation leaching and agitation leaching, applicable mainly to large
deposits of oxide ores, are discussed later in the report.
Copper Solutions From Leaching (19)
The pregnant solutions derived from leaching normally contain up to
several grams per liter of copper as copper sulfate. The method does
not recover gold, silver, selenium, and tellurium that are valuable
by-products of the smelting operation. Solutions from this type of
leaching are too dilute for the recovery of copper by direct means,
43
-------
such as electrolysis, and must be treated by an intermediate concentra-
tion process such as cementation or solvent extraction.
Cementation (20)
Cementation is a process whereby copper in solution is precipitated or
"cemented" out of solution as a finely divided metallic product by re-
placement with some less noble (that is, more active) metal. In the
copper industry, scrap iron is the cementing metal used. The scrap
iron may be in the form of salvaged tin cans, factory scrap such as
punch plate, or drillings, automobile scrap, baling wire, etc. Massive
scrap, such as rails, may be used at the head end of & penetration
launder to reduce any ferric sulfate in solution. Tin cans are pre-
ferred.
Cementation is carried out by directing the flow of copper solution
through launders (or channels) loaded with scrap iron. As the solution
flows through the bed of iron, the reaction
Fe + CuSO. -> Cu + FeSO.
4 4
solution precipitated solution
occurs, the copper depositing in loosely adherent form on the iron sur-
faces from which it may be easily detached by washing. Cementation
operations are so designed and scheduled that when a launder or section
of launder is loaded with copper, it may be taken out of the circuit and
the copper washed out of the system through screens into settling basins
by means of high-pressure water sprays. Washing techniques vary. In
some cases, they are highly mechanized, employing machines called
"sweepers" which traverse the length of the launder directing high-
velocity jets of water into the scrap and driving the fine cement cop-
per into settling basins. In others, high-pressure, manually operated
hoses are employed. Powerful electromagnets are used to remove iron
from the launders to expose cement copper to the action of the wash
waters.
In the cementation process, iron, which replaces copper in solution, is
consumed. The theoretical consumption is about 0.9 pound of iron per
pound of copper. In actual practice, iron consumption is generally
much higher, owing to side reactions with free acid and ferric sulfate
in the leach solutions
H-SO. + Fe -> FeSO, + H0
24 42
3 + Fe -> SFeSO^
Cementation recovers virtually all of the copper from the leach solu-
tion, leaving a barren or stripped solution containing ferrous sulfate
44
-------
and some free acid. Such solutions are returned to leaching operations
with or without the addition of more sulfuric acid.
Cement Copper (21)
The product of cementation is called cement copper. It is a relatively
impure material containing from about 70 to 90 percent copper and con-
taminated with such materials as shards of scrap iron which escaped
screening; dirt, waste, rust, etc., introduced with the scrap iron;
basic copper and iron sulfates originating from occlusion or adsorption
of sulfates from the leach solution; and considerable oxygen, the re-
sult of the atmospheric oxidation of the highly reactive fine copper
particles. Most cement copper produced in the United States is sent to
smelters where it generally is fed to the reverberatory matte furnaces.
In a few cases, it may be beneficiated by physical and/or chemical pro-
cesses to an acceptable grade of copper powder for use in powder-
metallurgy applications.
Solvent Extraction (22)
Solvent extraction is a new technique for recovering copper from leach
solutions, adopted recently by several western companies to supplant
cementation. In this process, the dilute leach solutions are contacted
with relatively small volumes of an organic solvent which selectively
extracts copper into the organic phase.
The organic solution containing the copper is then contacted with a
relatively small volume of aqueous "stripping" solution under condi-
tions of acidity that cause the copper to leave the organic phase and
enter the aqueous stripping phase. The resulting solution, containing
upwards of 30 grams per liter of copper, is now sufficiently concen-
trated and pure for subsequent treatment by electrolysis to produce
cathode copper.
Electrowinning (23)
Electrowinning is the term given the process of electrolyzing or de-
positing copper from leach solutions in the form of cathodes, of a
purity comparable to electrolytically refined copper. The overall
principles, operation, and results associated with electrowinning are
similar to those of electrolytic refining (16) with these exceptions:
(1) Copper anodes which supply the feed of copper to electrolytic re-
fining cells are not used in electrowinning. Insoluble anodes (mainly
antimonial lead) are substituted and the copper feed to electrowinning
45
-------
cells is in the form of solution, either from the solvent extraction
process (25) or from specialized leaching operations (34-35) which are
discussed later (see Figure 4). Cell construction, electrical hook-
ups, starting sheets, etc., are similar to those used in electrolytic
refining.
(2) Power requirements for electrowinning are 8 to 10 times greater
than for electrolytic refining, but this factor is offset to some ex-
tent by lower labor costs.
(3) Because long-life, insoluble anodes are used in electrowinning,
some of the major cost items of electrorefining are minimized. Cathodes
produced by electrowinning operations are melted and cast into the con-
ventional refinery shapes.
High-Grade Sulfide Ores (24) (Route III, Figure 4)
High-grade ores are almost exclusively the product of underground
mining. They may range in grade from about 3 to 10 percent of copper.
They constitute a minor portion of the copper production. As shown in
Figure 4, they may be smelted directly or sent to concentrators, de-
pending on economic considerations.
Mixed Sulfide-Oxide Ores (25) (Route IV, Figure 5)
These are the ores which are concentration grade (about 0.8 to 1.0 per-
cent copper), but in which the copper values occur as both sulfide and
oxide minerals. Sulfide minerals are, as has been explained earlier,
amenable to concentration. Oxide minerals are only partially so. It
has been stated that the loss of oxide minerals in flotation may amount
to as much as 50 percent. Special recovery methods are therefore
applied to these mixed ores.
Two such processes are
(1) Preliminary percolation or "vat" leaching to recover oxide copper,
followed by treatment of the leach residue in concentrators to recover
the sulfide minerals
(2) The so-called leach-precipitation-flotation procedure.
Vat LeachingConcentration (30)
In this process the ore--a mixture of sulfide and oxide mineralsis
crushed to about 3/8 inch in size and first leached by the percolation
46
-------
or "vat leaching" technique described in detail later in the report.
This extracts upwards of 90 to 95 percent of the oxide copper present
and recovers it in concentrated solutions which can be electrolyzed to
produce cathode copper (23) and/or dilute solutions which must be
treated by some concentration process such as cementation (20).
After removal of the oxide copper, the residue, containing practically
all of the sulfide copper originally present, is excavated from the
vats and transferred to a concentrator where the sulfide copper is re-
covered by flotation (4).
Leach-Precipitation-Flotation (27)
This is an alternative process for treating mixed sulfide--oxide ores.
In this process, the ore is leached in dilute sulfuric acid to solubi-
lize oxide copper minerals. The ore is first crushed and ground and
separated into sand (relatively coarse granules) and slime (fine pow-
der) fractions. The sands, which account for about 80 percent of the
weight of the original ore, are leached in dilute sulfuric acid to dis-
solve the oxide copper from the oxide minerals. After washing, they are
further ground and treated by conventional flotation processes to re-
cover copper sulfides in the form of flotation concentrates. The preg-
nant solution from the leaching of the sands is then used to leach cop-
per from the slimes in agitated tanks. After this secondary leaching,
sponge iron (i.e., fine powder) is added to the mixture of copper solu-
tion and slimes. This precipitates the dissolved copper as metallic
"cement" copper. This "cement" copper, which responds to flotation
much as do copper sulfide minerals, is recovered by a modified proce-
dure as a flotation concentrate which is combined with the flotation
concentrate from the treatment of the sands.
Native Copper Ores (28) (Route V, Figure 4)
Native copper ores occur predominantly in Michigan in the Lake Superior
District. In the mines now being worked in that state, native copper
constitutes only a small percentage of the value, with chalcocite being
the major copper mineral. In times past, when native copper was an im-
portant constituent of the Michigan ores, specialized concentration
procedures were employed to recover it. These included stamp-milling,
jigging, tabling, etc. At the present time, the Michigan ores, all
underground, are treated by more or less conventional concentration and
refining methods.
47
-------
Oxide Copper Ores (29) (Route VI, Figure 4)
Some ores, notably in Arizona and Nevada, are predominantly oxide ores
containing chryso'colla, azurite, and malachite. If sufficiently high
grade, which is a rare occurrence, they may be sent directly to smelt-
ers. If low grade (less than about 1.5 to 2 percent), these oxide
ores are treated by various leaching techniques. Very low-grade oxide
ores were discussed earlier and may be heap leached (18). Those con-
taining upwards of about 0.8 percent copper are generally leached bv
percolation or vat-leaching processes.
Vat Leaching (30)
In vat or percolation leaching processes, the ore is first crushed
to about 3/8 to 1/2 inch in size and loaded into vats. These vats are
walled concrete structures of various sizes, ranging in capacity from a
few hundred tons up to 20,000 tons of ore. The floors of leaching vats
are, in effect, filters which will permit the upflow or downflow of
leach solutions and wash solutions. Vat floors are constructed in
various ways as with timbers, graded gravels, sand, fiber matting, etc.
The vats are usually arranged in banks or groups, with a common wall
between adjacent vats. The number of vats in a bank may range from
6 to more than 12. Auxiliary piping, drains, sumps, pumps, tanks, etc.,
are arranged to accommodate the required flow pattern of leaching and
washing solutions through the system. Conveyor-belt systems provide
the most efficient and satisfactory methods for loading ore into vats,
although smaller installations may employ such means as front-end
loaders, drag scrapers, etc. Unloading or excavating after leaching,
in the larger plants, is done by traveling clam shells.
Leaching in vats is a carefully designed and executed process aimed at
extracting upwards of 90 percent of the copper in the ore. It is accom-
plished by passing dilute sulfuric (about 2 to 5 percent) acid solution
through the crushed ore mass in the vats by upflow percolation, downflow
percolation under flooding conditions, or downflow nonflooded percola-
tion (i.e., by spraying or otherwise distributing leach solution onto
the top surface of the ore and permitting it to trickle down through
the ore).
The pattern of leaching varies from installation to installation, de-
pending on the reactivity of the copper minerals in the ore, the number
of vats provided, acid strength, components of the gangue, solution con-
centrations desired, etc. Commonly, the countercurrent pattern is
followed in which leach solution from one vat is advanced to the next
containing a new charge.
Complete dissolution of the recoverable copper to sulfate may require
a week or more. After conversion of the copper oxides to soluble
48
-------
copper sulfate is accomplished, it is necessary to wash the ore free of
soluble copper. This is done by a countercurrent process in which
fresh wash water is added to the vat next to be discharged and passed
then to the adjacent vat. Storage tanks for wash waters representative
of various washing stages are provided in the larger installations.
Pregnant solutions of various strengths may be prepared, depending on
the leaching pattern. In some installations the pregnant solution may
comprise both the leach and wash solutions. In some of the larger
plants, leach solutions containing up to 35 grams per liter of copper
are withdrawn and sent to electrowinning processes. The wash solutions
generated in these cases will be considerably diluted, and are treated
by cementation or solvent extraction.
By-Product Copper Ores (31) (Route VII, Figure 4)
Much of the by-product copper comes from lead-zinc, molybdenum, or iron
ores. Most of it is received and separated from the circuit by selec-
tive floation processes developed for the treatment of these ores. In
one case, where copper is the by-product of iron-ore beneficiation, the
copper is recovered by a combination of leaching and cementation.
49
-------
SECTION VI
THE SELENIUM AND TELLURIUM INDUSTRY
Size and Characteristics
There are no selenium and tellurium ores; hence there is no selenium
or tellurium mining or concentration. The principal, if not the only,
raw material for the recovery and refining of these metalloids in the
United States is the electrolytic slimes, a by-product of copper re-
fining. A minor contributor to selenium-te1lurium production may be
the fumes and dusts from lead-smelting operations.
In terms of mass, the industry is small. In 1969, for example, total
selenium production in the United States was about 600 tons, while
tellurium production was about 120 tons. This production was distri-
buted among the seven companies shown in Table 6.(*-)
TABLE 6. UNITED STATES PRODUCERS OF SELENIUM AND
TELLURIUM METALS AND COMPOUNDS
American Metal Climax, Carteret, N.J. X X
American Smelting and Refining Co., X X
Baltimore, Md.
International Smelting and Refining X X
Co., Perth Amboy, N.J.
Kennecott Copper Corp., Garfield, X
Utah
Kawecki-Berylco Industries Inc., X
Boyertown, Pa.
Phelps Dodge Refining Corp., Maspeth, X X
N.Y.
U.S. Smelting Lead Refinery, Inc., X
East Chicago, 111.
Raw Materials and Processes
Anode slimes, as discussed previously in the section of this report
dealing with copper, are complex mixtures of base metals, precious
metals, and selenium and tellurium. The precious metals, silver and
gold, are by far the major values in slimes. Consequently, the
51
-------
processes which are used to recover and refine selenium and tellurium
are designed also to recover precious metals.
Several processes, all of them intricate and requiring considerable
care and personal attention by operators, are available.(22,23)
Figure 7 is a flow diagram showing one of the approaches. As shown in
this figure, the slimes--previously decopperized by leaching with sul-
furic acid (as described under copper)--are smelted in a small reverber-
atory furnace, called a Dore furnace. The first slag which forms is
removed; sodium carbonate and sodium nitrate in the weight ratio of 2:1
is then added. The soda slag formed contains selenium and tellurium.
During smelting, some selenium and tellurium are lost from the Dore
furnace by volatilization. These are recovered by scrubbing the fumes
with water.
The soda slag is ground in a rod mill and leached in hot water to dis-
solve selenium and tellurium; the slurry is filtered and the filtrate
containing most of the selenium and tellurium is combined with the
scrubber water which contains the selenium and tellurium lost from the
Dore furnace. Although not shown in Figure 7, gold and silver are re-
covered as Dore alloy from the Dore' furnace.
The combined filtrate and scrubber water are neutralized with sulfuric
acid to a pH of about 5.5. At this pH, tellurium precipitates and is
filtered off and sent to a series of refining steps. The solution con-
taining selenium is acidified with sulfuric acid and treated with sul-
fur dioxide gas. This treatment precipitates selenium as metal which
is filtered, dried, and packaged for shipment. The finished selenium
product has a minimum purity of 99.5 percent.
The filter cake containing tellurium is treated with sodium hydroxide-
sodium sulfide solution to redissolve tellurium. Precious metals that
may be present at this stage remain in the insoluble residue, are sep-
arated by filtration, and are returned to the Dore' furnace. The solu-
tion containing tellurium as sodium tellurite is neutralized with sul-
furic acid to a pH of 5.5 whereupon tellurium reprecipitates as puri-
fied tellurium dioxide (Te02). This may be dried, ground, and sold as
such, or reduced to metal by carbon in a crucible and cast into ingots.
Numerous variations of this process are practiced by other companies.
This process and the variations of it are the result of much research
and process development work undertaken by the companies to determine
optimum conditions for treating their particular feed materials.
52
-------
Dore'
Ju.noce
si ag
Sodo slog
i
Rod mill
I
.Leoch lonk
- Fillc. press
1
Flue gas
y
Scrubber
Evopora
Neuirclizing tank
Se-Te separation
. »2SO<
stofotje tonk
r^
Dor< lurnace
;
lie. c
T= beo,,n,
solution
Punfied
Te02
Punlird lellunum
G.md,
pflck, ond
ship
Se beo.mg solul
i
S«p,ec,p,,o,,on<, H2»(
lonk S02
Fill., boi
Selenium:
D,y.
g,md.
screen,
peck, and
ship
FIGURE 7. FLOWSHEET OF SELENIUM-TELLURIUM RECOVERY PROCESS
(23)
(Courtesy, United States Metals Refining Co.,
Division of American Metal Climax, Inc.)
53
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SECTION VII
WATER USAGE IN THE COPPER INDUSTRY
Data on water usage were obtained for about 30 copper-producing facili-
ties in the country. Some of the data were gathered from the open lit-
erature, but the major and most meaningful data were obtained from in-
formation supplied by copper-producing companies.
Table 7 is a general picture of the coverage achieved in the study.
The table shows the gross intake and discharge volumes reported by re-
spondents for the indicated copper-producing facilities. The first
line of the table, for example, shows that a copper mining and leaching
operation used 1,066 million gallons per year of water and discharged
45.4 million gallons per year of this as waste. The data also indicate
that a number of mining and milling operations (Lines 3-10, e.g.) which
use large quantities of water produce no waste effluent at all. Such
facilities, located in water-scarce sections of the country, recircu-
late as much water as possible to their processing. Water consumption
in these instances is due to evaporation and seepage into the ground.
Table 8 is an extension of these data to show total water use (includ-
ing recirculated water) and the intake, use, and consumption of water
per ton of metal produced for various associations of facilities (mining
and leaching, mining and concentrating, etc.).
The values shown in Table 8 indicate the wide variation in water-utili-
zation practices. The category of "mine and concentrator" operations
should represent the simplest case, with minimal variations in process
technology or operations. Yet even here, the water use in gallons per
ton of contained metal, a term which should approach expression of the
common features of the process, varies by more than a factor of ten,
i.e., from 13,000 to over 190,000 gallons per ton of contained metal.
The explanation for this variation would require a specific study of
the operational details not disclosed in this investigation.
Copper mining and leaching operations, represented by three examples in
the table, show about the same tenfold variation in water usage per ton
of metal (15,000 vs 147,000). Two of these show the high recycle prac-
tice that is characteristic of the industry as a whole. The operation
showing the lowest of the three recirculation rates (773 MGY) may need
to decrease, in the future, losses to seepage, but additional data
would be required to confirm this.
In the category of "Mine, Concentrator, and Smelter", the water usage
and consumption per ton of metal and the recirculation rate fall into a
reasonably closely grouped pattern. Specific studies based on addi-
tional information would be necessary to explain the variations in water
55
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TABLE 8, WATER DATA FOR COPPER-PRODUCING FACILITIES
Total
Water
Intake ,
MGY(a)
2,170
2,560
1,980
949
795
363
1,070
630
4,760
3,960
2,310
437
4,200
3,510
32,600
4,230
6,740
610
540
5,820
456
656
804
210
297
1,930
6,875
Total Use Gallons
Discharge to New and Intake per
Receiving Recirculated , Ton of
Waters, MGY(a) MGY*-a) Metal
0
0
0
0
0
0
45.4
120
0
0
5.1
0
0
0
19,500-28,
2,780
625
473(b)
0
1,130
2.4
606
734
75.2
297
1,720
6,600
Mine and Concentrator
6,410 28,900
3,720 43,700
9,300 41,200
2,740 36,000
2,7/0 26,500
480 9,070
Mine and Leaching Operations
7,670 20,500
773 12,600
25,000 16,300
Mine, Concentrator, and Smelter
16,300 30,400
32,600
21,900
17,500 41,000
31,500 28,400
Concentrator and Power Plant
500 144,000 130,000
Concentrator and Smelter
6,850 56,100
7,830 136,000
Smelter
980 6,400
5,280 3,650
19,900 17,900
Refinery
3,730 1,070
2,390
3,320 3,470
256 763
297 1,600
Smelter and Refinery
9,110 9,670
8,510 42,800
Gallons
Used
per Ton
of Metal
85,400
63,200
193,000
104,000
92,300
13,300
147,000
15,500
85,500
125,000
--
175,000
254,000
573,000
91,000
156,000
10,300
39,300
61,400
8,770
--
14,300
927
1,600
45,500
52,900
Consumption
(Intake Minus
Discharge) ,
gallons per
ton of metal
28,900
43,700
41,200
36,000
26,400
9,100
19,600
10,200
16,300
30,400
31,100
41,000
28,400
16,500-52,300
19,300
124,000
1,100
3,650
16,100
1,070
182
304
490
0
1,100
1,710
(a) Million gallons pet year.
(b) Receiving "water'1 is a watercourse which is normally dry; it carries
"natural" runoff during seasonal rains.
57
-------
usage consumption and recirculation calculated for the groupings "Con-
centrator and Smelter", "Smelter", and "Smelter and Refinery". These
variations are probably associated with such factors as size, location,
technical and economic considerations, etc.
The data in Table 9, also derived from information supplied by pro-
ducers, shows the allocation of new water by percentage to various
phases of the operation (process, cooling, boiler feed, air-pollution
control, sanitary, and miscellaneous) for various associations of
facilities. The table also shows the degree of recycling practiced in
the entries under "Use Ratio".
Waste-Water Sources, Characteristics, and Amounts
The ways in which water is used and waste is generated in the copper
industry in metallurgical operations are indicated in Figures 8 through
10. These figures are generalized flowsheets based on information sup-
plied to this survey. Water sources for mines and concentrators in-
clude wells, surface waters, or mine water itself. The waters may be
treated in any number of ways before use, but the most common are indi-
cated: potable (sanitary) water is usually chlorinated and, if neces-
sary, clarified by flocculation, filtering, etc. It may also be soft-
ened, depending on the quality of the source. The potable water may be
obtained from a nearby municipal supply. The sanitary water circuit
includes drinking water, change-room showers, toilets, etc. In the
case of isolated operations, the circuit includes the nearby townsite
in its entirety. All plants either treat the sanitary wastes or dis-
charge to a municipal sewer (which may, as indicated above, be corpor-
ate operated). Plant sewage-treatment effluents may go to ground in
isolated, desert operations, to surface or estuarine receivers, or be
recirculated to the metallurgical process. Another common use of water
in mining operations (as shown in Figure 8) is for dust control, in
areas of underground or open-pit mining, ore-haulage roads, drill
"cooling", or at dusty process areas (primary crushers, conveyor trans-
fer points, etc.). Most of such water remains in the ore as moisture
and follows a recirculation path with the ore to the concentrator. The
remainder of such water is lost to ground or to evaporation.
Flotation requires large amounts of water, which is recirculated to the
process after dewatering tailings and concentrates in thickeners, clas-
sifiers, filters, etc. The thickening operation produces an overflow
of substantially clear water, which is immediately recirculated. The
underflow of the thickener is tailings slurry which is pumped to a tail-
ing disposal site. Here the solids settle slowly and the clear water is
either discharged or recycled.
Most mining and milling operations require steam generators, with the
steam being used for electrical-power generation or general plant use.
Cooling water may be recirculated through a cooling tower or spray
58
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pond loop and reused.
A water loop characteristic of the copper industry is the leaching
operation where the large quantities of water, really dilute acid solu-
tion, are recirculated between an ore dump and a copper-precipitation
plant. Losses to evaporation and ground seepage must be made up with
intake water, which may range in quality down to sewage-plant effluent.
Spent leach liquor no longer usable owing to the buildup of various
salts may be discharged, evaporated, or sent to a tailings pond. Mine-
concentrator operations, particularly in the arid sections of the west,
often employ complete recirculation with no discharge to surface waters.
Figure 9 is a generalized flowsheet showing the use of water in smel-
ters. These may be associated with mine and concentrator and utilize
the same water sources, sanitary circuit, and the tailings pond for
discharge, or they may be a separate installation. The common features
of the water circuits of a smelter are the manifold cooling functions:
spray cooling of calcine in the roasting or sintering operations, where
water is mostly lost to evaporation or stays in the material; the cool-
ing of furnace components (doors, shells, etc.); and the cooling of the
cast metal product and the casting molds. Furnace cooling requires no
contact of the water with the material being processed, and cooling of
solidified metal and molds generally results only in the pickup of mold-
dressing materials (oils or silica-base compounds). This cooling water
may be settled to remove suspended solids and recycled or discharged.
In some plants, molten slag from the furnaces is granulated with water
jets, the slurry dewatered, and the water settled and either recycled
or discharged.
Smelting operations may include sulfuric acid plants. Water is used in
these in a number of ways. Water scrubbers are used to remove dusts
and fumes from sulfur dioxide gas streams from roasters and converters
before admission to the sulfuric acid plant. Water is used to dilute
the sulfuric acid product to the desired strength. It is used to cool
reactors and machinery. Acid plants produce discharges of cooling-
tower blowdown and "acid plant waters" from the sumps of the gas clean-
ing or conditioning steps. The locations of smelters vary from deserts
to seacoasts and thus sources of supply and types of receivers may in-
clude independent systems, municipal systems, or estuarine waters, or
combinations thereof. 'In some installations, slags are granulated by
shock cooling with water before disposal or recirculation.
The pattern of water usage and waste treatment or disposal in electro-
lytic refineries is depicted in Figure 10. Make-up water is used for
the preparation of electrolytes for both the primary product and, in
many cases, precious or other metal by-products as well as the prepara-
tion of solutions for various leaching and chemical operations. Wet
gas scrubbers may use water for either or both of the functions of air-
pollution control and by-product recovery. Virtually all electrolytic
refineries employ melting, refining, and casting operations to produce
salable forms of the primary and by-product metals, and require
64
-------
use of cooling water for furnace components, cast metal, molds, etc.
When these are located near seacoasts, cooling is accomplished with
salt water, usually pumped from and returned to estuaries. Cooling is
also required for the characteristic electrical equipment such as trans-
formers, rectifiers, generators, etc. Steam generation is also re-
quired for controlling the temperature of the electrolytic baths. The
discharges from electrolytic refineries are many and varied, as are the
disposal techniques and types of receivers.
Specific Sources of Waste Water
Specific sources of waste water in the copper industry are as follows:
1. Water-treatment wastes, including filter backwash, sludges from
primary settling and ion-exchange regeneration solutions.
2. Sanitary wastes.
3. Indirect and direct cooling water from the cooling of furnaces,
machinery, casting operations, etc. Generally, such water is recircu-
lated to cooling towers or spray ponds for reuse, and is eventually
discharged as blowdown. Some plants, notably smelters and refineries
located near large bodies of water such as the sea, may use once-through
cooling water.
4. Process wastes including mine drainage, flotation-plant discharges,
discarded leaching solutions, scrubber water, smelter waste water, dis-
carded electrolyte from refineries, sanitary waste, clean-up water, etc.
Often these may be combined in tailings ponds for treatment and recir-
culation.
5. Boiler and power plant wastes including boiler blowdown and ash-
pit overflow.
Data on the quantity of these individual waste streams were not avail-
able to this study. Some data on the characteristics of such wastes
were obtained however. Table 10 shows the composition of raw waste to
tailings ponds and of tailings-ponds effluents for several mine-concen-
trator combinations. Table 11 gives similar data for several opera-
tions involving mines, concentrators, and smelters. Table 12 shows
compositions of waste streams from copper smelters and refineries.
The methods of waste treatment used in the copper industry as revealed
by this study are shown in Table 13.
65
-------
TABLE 10. COMPOSITIONS OF WASTE WATERS FROM MINE AND
CONCENTRATOR OPERATIONS
(All analyses in mg/1)
(a)
Components
As
BOD
Cd
CN
Cu
F
Fe
Mn
Pb
pH
TDS
TSS
Zn
COD
Colif orm
Metal Sulfides
Raw Waste
Discharge from
Concentrator
0.01
5.8
0.01
0.06
0.10
2.9
0.12
0.08
0.005
7.3
3412
48
0.03
--
--
~ ~
to Tailings
Plant Drain
A
0.09
164
--
--
0.33
2.1
--
--
--
7.0
2310
549
--
343
211,000
0.23
Ponds
Plant Drain
B
0.03
50
--
--
0.28
1.3
--
0.
0.
6.7 7.
3950
53
2.
106
20,000
0.18
Tailings
Pond Effluent
0.006
--
nil
0.01
0.05
8 nil
18 0.02
64
--
--
0
--
--
"* "~
(a) BOD = Biological oxygen demand
CN = Free cyanide
TDS = Total dissolved solids
TSS = Total suspended solids
COD = Chemical oxygen demand
Coliform = Bacteria per 100 ml.
66
-------
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67
-------
TABLE 12. COMPOSITION OF WASTE STREAMS FROM COPPER SMELTERS
AND REFINERIES
(Analyses in tng/1)
Components
As
BOD
Cd
COD
Cu
Metal
Sulfides
Pb
pH
Se
soA
TSS
TDS
Zn
Fe
Ni
Cr
Hg
Oil
F
Smelter
Acid Combined Electrolytic Refineries
Plant Plant I II III IV
Water Waste Intake Discharge Combined
<5 22.1 -^ -- -- o:55
211 10 70
0.6 <0.1
457
<9 7.4 0.07 0.15 50-100 0.35 2.8
11-7
<5 7.2 " 1 0.05
2.5-3.0 1.9 7.4 7.7 1-2 6.1 7.6
0.7 " 26,000 0.7
~* 1716 26,000
167 __ 4g
2-30,000 1024
<6 7.3 0.3 0.4
0.2 1.00 50
<.l <0.1 10-20
<0.03
<0.1
10.5
0.6
(a) BOD = Biological oxygen demand
CN = Free cyanide
TDS - Total dissolved solids
TSS = Total suspended solids
COS = Chemical oxygen demand.
68
-------
TABLE 13. WASTE WATER TREATMENT PRACTICES IN THE COPPER INDUSTRY
Type of Waste Treatment
Mine and Concentrator Recirculatton with no discharge in arid cli-
mate; sedimentation in tailings ponds; neu-
tralization, flocculation if required
Smelter Combined with mine and concentrator wastes
where possible, recirculation in arid climate
Electrolytic Refinery Practice variable; neutralization, evapora-
tion, precipitation as required
Leaching Plants Recirculation with discarded leach liquor
routed to tailings ponds where acid is neu-
tralized
Future Waste-Water-Treatment Methods in the Copper Industry
Cooperating respondents in the copper industry indicated that the
following steps were either being taken or being seriously considered
to improve the quality of waste discharged:
1. Increased recirculation of waste water with the ideal objective of
100 percent reuse
2. Neutralization of waste discharges from tailings ponds when indi-
cated
3. The use of polyelectrolyte flocculants in place of lime to obtain
better settling
4. Increased size of settling ponds
5. Neutralization and precipitation of leach solutions now being dis-
carded and reuse of the water in leaching
6. Installation of settling ponds in smelters with provisions for pH
and temperature control so that water can be completely recirculated
7. Rerouting of sewer lines in smelters so that wastes with suspended
solids burden can be segregated and sent to thickeners or lagoons
8. Isolation and reuse of "clean" water now being discharged to waste
by smelters
69
-------
9. Reduction in the quantity of cooling water now being used in
smelters to cool converter hoods by installation of a new system
10. Neutralization of scrubber waters from the gas cleaning step in
sulfuric acid plants
11. Reduction of the amount and strength of the waste from electroly-
tic refineries by recirculating cathode wash water, spills, etc.
12. Investigation of the possibility of deep-well injection in con-
junction with some recycling for smelter and refinery wastes.
70
-------
SECTION VIII
THE PRIMARY LEAD AND ZINC INDUSTRIES
Size and Characteristics
The lead and zinc industries are closely associated. Lead and zinc
minerals occur in many major ore bodies in such intimate mixtures that
they must be mined together. Most lead ores contain significant quan-
tities of zinc. Similarly, many zinc ores contain economically signi-
ficant amounts of lead. Even when lead or zinc minerals predominate in
a given ore, neither is entirely free of the other and, at some stage
of processing, separations must be made. In addition, lead and zinc
ores commonly contain other metals, such as copper, bismuth, antimony,
gold, and silver, which for technological or economical reasons must
be separated from lead and zinc. The processes for treating lead ores,
zinc ores, and lead-zinc ores are highly complicated operations inter-
woven with each other as well as with processes for copper, precious
metals, cadmium, etc.
As an example of the interconnection of the lead-zinc industries, Table
14 shows the amounts of lead and zinc recovered from so-called lead
mines, zinc mines, lead-zinc mines, and other sources in the country.
This table also indicates the geographical distribution of the mining
section of the industry. As shown in the table, virtually all of the
so-called lead mines produce zinc in an amount averaging, the country
over, to about 13 percent of the lead production. The table also
shows that most of the designated zinc deposits yield some lead, with
two notable exceptions in Tennessee and Pennsylvania. In the desig-
nated lead-zinc mines, the recovery of each metal is about equal. In
complex ore bodies associated with copper, the ratio of lead to zinc
recovery is about 1 to 2.2 and, in the case of the miscellaneous
sources, zinc recovery preponderates over lead recovery by more than 2
to 1.
The total number of mine sources of zinc and lead in the United States
has been placed at 300. These mines are mostly in the Rocky Mountain
and Missouri-Tennessee areas, but minable deposits ,are scattered in
many of the states throughout the country, excluding the Gulf Coast
states. Smelters and refineries are similarly widely located. Zinc
aqd lead production in the United States ranks fourth and fifth, re-
spectively, in tonnage, after iron, aluminum, and copper. The industry
consists of a large number of individual mining operations, with an
associated concentrating facility (usually a flotation process), and a
lesser number of smelting and refining operations.
71
-------
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Geographic Distribution
The names, ownerships, locations, and ore-production figures of se-
lected major lead and zinc producing mines, as gathered in a survey by
the lead and zinc industry in 1968 and published in 1970(^-2) ; are
shown in Table 15. Tables 16 and 17 show similar data for zinc-produc-
ing plants and lead smelters and refineries, respectively.
The geographic distribution of the mining sources and of the smelters
and refineries is shown on the map in Figure 11. Table 17, previously
referred to, lists the lead and zinc production by states, for 1969,
from various ore sources. As may be gathered from the table, the
states east of the Mississippi River produced about 60 percent of the
zinc, while western states produced 36 percent, and the Kansas-
Oklahoma area produced 4 percent.
In contrast, mines west of the Mississippi produced 96 percent of the
lead ores, with the most important lead-producing areas being the Summit
Valley region in Montana, the Metaline Falls area in Washington, the
Upper San Miguel region in Colorado, and the rapidly developing New
Lead Belt in southeastern Missouri.
Table 18 lists salient economic statistics for that phase of the lead-
zinc industry concerned with smelting and refining for the year 1967.'°)
Similar data for the mining and milling phases of the industry were not
available to this study. It has been stated, however, that production
personnel engaged in mining and milling of lead and zinc in 1968
amounted to about 10,000. Table 19 lists the major uses of lead and
zinc.
Projected Growth
In spite of current fluctuations in price, plant closings, and uncer-
tainty of the future market for tetraethyl lead, the Bureau of Mines
prediction for the annual rate of growth of the lead and zinc industries
ranges from 1.2 to 3.4 percent for both metals.'^' The 1968 base pro-
duction figures and predicted high and low figures for the year 2000
were given as follows:
Zinc Lead
1968 Primary Production, short tons 529,000 359,000
Year 2000 Primary Production Range,
short tons:
High -- 1,120,000
Low 786,000 520,000
1968-2000 Growth Rate, percent
High 3.3 3.4
Low 1.2 1.8
73
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-------
TABLE 18. SELECTED ECONOMIC STATISTICS FOR THE PRIMARY
ZINC AND LEAD INDUSTRIES
Number of Employees
Payroll, millions of dollars
Capital Expenditures, millions of dollars
Materials Consumed
Ores, Concentrates, short tons
Delivered Cost, millions of dollars
Cost of All Other Materials and Supplies,
millions of dollars
Cost of Purchased Fuel and Electrical Energy,
millions of dollars
Product Primary Refined Metal, short tons
Value, millions of dollars
Zinc
8,100
57.8
25.8
1,757,300
147.5
15.9
22.9
1,037,700
339.1
Lead
2,700
18.9
18.5
1,002,600
152.8
109.8
7.1
379,894
106.37
(a) For smelting and refining only; does not include mining.
(6)
TABLE 19. CONSUMPTION PATTERN OF LEAD AND ZINC IN
(Short Tons)
Construction
Metal Products
Transportation
Storage Batteries
Electrical Equipment
Pigments
Plumbing and Heating
Chemicals
Machinery
Miscel laneous
Pigments and Compounds
Unclassified
Total
Lead
--
364,250
--
513,703
--
107,258
--
262,526
--
23,106
--
17,924
1,288,767
Zinc
340,000
--
400,000
50,000
210,000
220,000
240,000
--
160,000
141,000
220,000
-
1,761,000
81
-------
The 1968 production figures used here are metal contents of domestic
mined ore, which are supplemented by imports or ore and metal and most
importantly by recovery of metal values from secondary (scrap) metal.
Raw Materials and Processes
As an aid in understanding the somewhat involved nature of the pro-
cesses discussed below, the following discussion is numerically keyed
to the diagram in Figure 12.
The ore minerals of lead and zinc are listed in Table 20.
TABLE 20. ORE MINERALS OF LEAD AND ZINC
(13)
Mineral
Chemical Composition
Remarks
Lead
Galena
Cerussite
Anglesite
Sphalerite
Franklinite
PbS
PbCO,
PbSO
ZnS
4
Zinc
(Zn,
Principal ore mineral
Results of weathering of galena;
may occur in minor amount in
galena deposits
Ditto
Principal ore mineral in U.S.
Important ore mineral in New
Jersey; minor in rest of U.S.
Zincite
Willemite
Gaslarite
Calamine
ZnO
Zn2Si04
ZnSO. -7H-0
4 2
Zn4Si207(OH)2.H20
Unimportant
Unimportant
Ditto
"
These minerals occur in a wide variety of ore types, grades, and assoc-
iations with each other and with the ore minerals of copper, silver,
gold, and manganese, and nonmetallics such as fluorite. Table 21 ex-
emplifies the ranges encountered in United States ores, the associated
ore minerals, and typical gangue rock.
82
-------
I/I ta
11 X
i
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-------
- a.
ffl O
c u
£X - OJ
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' 3 "O
I -= O
84
-------
The ores of lead and zinc may be arbitrarily classified according to
their principal mineral values. The so-called lead (1) ores, in which
lead is a major component and zinc minor, are best exemplified by the
ores of southeastern Missouri in which lead is almost exclusively the
major metal value. The so-called "lead ores" also occur in many of the
Western states and production data for "lead ores", as opposed to zinc
ores, lead-zinc ores, etc., are annually published by the Bureau of
Mines. These lead ores customarily contain economically significant
quantities of zinc, copper, gold, silver, etc.
The so-called zinc ores (2) occur mostly east of the Mississippi
(Virginia, Wisconsin, New York, Kentucky, Illinois). Some are found in
states west of the Mississippi (Colorado, Missouri, Kansas, Oklahoma).
The zinc ores almost always contain significant quantities of lead,
with zinc greatly preponderating over lead.
The lead-zinc ores (3) in which lead and zinc both occur in significant
quantity and in which copper is not the major mineral value, with
neither lead nor zinc greatly preponderating over the other, are found
in the West and Southwest and usually contain copper and precious
metals.
The so-called copper-lead-zinc ores (4) are also limited to western de-
posits and are considered to be those ores which yield economically im-
portant quantities of copper.
Lead and zinc are also recovered from other ores (5) and materials such
as gold and silver ores, tailings piles, etc. These are exclusively
western sources.
The above classifications are used by the United States Bureau of Mines
in reporting annual data of the lead-zinc industry in the Minerals
Year Book.
Mining (6)
In the United States, almost all lead and zinc ores are mined by under-
ground techniques in contrast to the major open-pit operations of the
copper industry. Limited mining of lead and zinc by small open-pit
operations has been used in the Tri-state region of Missouri and in
Washington. Some zinc mines in the early stages of their development
are also mined by open-pit methods. Underground methods most commonly
used for lead and zinc mining include block-caving, cut and fill (de-
scribed in the section of this report on copper), room and pillar, and
various stoping techniques.
85
-------
Concentration (7)
Ores of lead and zinc are always concentrated. Two general methods of
concentration are employed.
In the one, called gravity concentration, advantage is taken of the
high specific gravity of the lead minerals galena (specific gravity
7) and of the zinc mineral sphalerite (specific gravity 4.7) to separ-
ate them from the lighter gangue or nonore materials. Gravity separa-
tion is done either by jigging or by float-sink methods in so-called
heavy media. Each route requires that the ore first be crushed to a
particle size at which the lead and zinc minerals are liberated or
"broken free" of gangue or waste material.
In jigging, the ore is crushed to the desired particle size, which may
range from 48 mesh up to 3 inches, and transported by water flow through
a machine, called a jig, in which the water-ore medium is pulsated ver-
tically as by the action of a diaphragm. Upward pulsations cause the
ore particles to rise. After the upward impulse, they settle again,
with the heavier minerals such as galena and sphalerite settling faster
than the light gangue or waste materials. The operation is so regulated
with respect to the frequency and intensity of the upward pulsations
that the lead and zinc minerals accumulate in the bottom and may be
drawn off as a concentrate. Lighter materials overflow the top. Jig-
ging is an effective means of concentration if the minerals have a
specific gravity-difference of 0.5 unit or more, although special
treatment in jigs may permit separation of materials exhibiting speci-
fic-gravity differences as little as 0.25. Accordingly, if liberation
is obtained, it is possible to produce, by jigging, separate lead and
zinc concentrates.
The float-sink or heavy-media method consists of feeding the crushed
ore to a suspension of ferrosilicon in water, which has an apparent
specific gravity of about 3.4. In this operation, heavy minerals, such
as galena, sphalerite, pyrite, etc., sink and the lighter gangue min-
erals remain afloat. The concentrates of heavy minerals are drawn off
the bottom of the heavy-media vessel, washed free of ferrosilicon par-
ticles, and sent to further processing. Similarly the floating tail-
ings are separated from ferrosilicon by washing and discarded. The
ferrosilicon, washed from the concentrates and tailings, is recovered
and recirculated to the heavy-media vessel.
Concentrates obtained by jigging or float-sink methods may be treated
directly by smelting practices. More often, however, middling products
result. Middling products are those in which the ore mineral is still
locked or attached to gangue minerals. Middling products are generally
ground finer and sent to flotation.
These gravity methods are often used in lead-zinc metallurgy as precon-
centration steps to eliminate gangue constituents of the ore.
86
-------
Gravity methods are particularly suitable for simple lead or lead-zinc
ores associated with gangue or waste minerals of low specific gravity.
Ores of this type are common in the mines of the Mississippi Valley
and the Eastern United States.
Concentration of lead and zinc ores by flotation is the general prac-
tice, however. As discussed in the section of this report on copper,
flotation is the process whereby mineral particles in a slurry of
water are treated with various chemical reagents to control their sur-
face properties so that they are either wettable or nonwettable, i.e.,
hydrophilic or aerophilic. As with copper, lead and zinc sulfide min-
erals can, by treatment with small quantities of flotation reagents,
be made aerophilic, so that, in a flotation cell system, they will
attach themselves to rising bubbles of air and be transported or "bal-
looned" to a froth on top of the slurry. This froth is mechanically
removed and constitutes a "flotation concentrate". The flotation pro-
cess with lead and zinc ores is quite complex and exhibits many varia-
tions depending on the type of ore, the association of minerals within
the ore, and the desired grade and recovery levels of lead or zinc con-
centrate required. By the use of relatively small quantities of condi-
tioning chemicals it is possible to manipulate the surface properties
of the different sulfide materials so that high-grade concentrates of
lead and of zinc can be produced from even the most complex ores of
lead, zinc, copper, etc.
The flotation process is used not only exclusively on many ores, but is
also often employed to affect additional recovery and separation of
lead and zinc from the products of the gravity separation methods dis-
cussed previously.
Table 22 shows typical analyses of lead and zinc concentrates produced
by gravity and flotation procedures (8, 9, 10, lOa).
A generalized flowsheet of the flotation process is shown in Figure 13.
Lead Production
High-Purity Lead Concentrates (11)
Some of the ores of southeast Missouri, which are almost exclusively
lead ores, yield extremely high-grade lead concentrates. Such concen-
trates may contain 70 percent or more lead, less than 2 percent silica,
less than 4 percent iron plus zinc sulfides, and insignificant quanti-
ties of precious metals, arsenic, antimony, bismuth, etc. These are
suitable for treatment in the ore hearth furnace.
87
-------
Ore
To Tailings
pond
Primary
Crushing
Operations
Thickener f >. TO Tailines
Pond
Filter ->
Zinc Concentrates
FIGURE 13. PROCESS STEPS IN DIFFERENTIAL FLOTATION OF LEAD-ZINC ORES
(Overflows/froths and underflows/tails indicated diagrnmatically)
-------
TABLE 22. RANGES OF COMPOSITIONS OF CONCENTRATES IN LEAD
AND ZINC METALLURGY (12)
Percent
Constituent Lead Concentrates Zinc Concentrates
?b
Zn
Au
Ag
Cu
As
Sb
Fe
Insolubles
CaO
S
Bi
Cd
45-60
0-15
0-a few oz/ton
0-50 oz/ton
0-3
0.01-4.0
0.01-2.0
1.0-8.0
0.5-4.0
tr(a)-3.0
10-30
tr^-O.l
0.85-2.4
49.0-53.6
__
--
0.35
--
--
5.5-13.0
3.4
--
30.7-32.0
--
0.24
(a) tr = trace.
Ore Hearth Smelting (12)
The modern ore hearth furnace is a mechanically rabbled, shallow basin
about 10 feet long, 2 feet wide, and 1 foot deep set under fume hoods
and fitted with a cluster of tuyeres for the admission of air across
its length. The high-grade ore, mixed with coke and limestone, is
spread over the hearth in successive thin layers. Roasting and reduc-
tion occur on the hearth. A travelling rake progresses slowly from one
end of the basin to the other to keep the charge loose and to break up
the slightly caked surface. Melted lead settles to the bottom and is
removed through a siphon arrangement to a pot. The slag is removed
mechanically and quenched in water. Fumes and dusts are collected for
subsequent recovery of lead.
The molten lead product collected in the pot may be cast directly into
marketable lead shapes, or may be treated by dressing before casting.
Dressing, as is discussed later, consists of cooling the molten lead to
a temperature just above its melting point and holding it at that
temperature for several hours under an oxidizing atmosphere. As a re-
sult, impurities such as copper and antimony collect in the lead oxide
layer on the surface and are skimmed off. The high-purity ores of
southeast Missouri are said not to require such dressing. Slag, fumes,
and dust from the ore hearth furnace may contain from 15 to 40 percent
of the lead in the original charge, and are generally re-treated in the
lead blast furnace.
89
-------
The chemical reactions occurring in the ore hearth furnace are:
2PbS + 30
Galena (air)
2PbO
Lead oxide
Sulfur dioxide
PbS + 2PbO
Galena Litharge
3Pb
Product
Sulfur dioxide
Refined lead from the ore hearth furnace typically contains about 99.93
percent lead, less than 0.002 percent silver, 0.04 percent copper,
0.015 percent zinc, and traces of iron and bismuth.
The ore hearth furnace was previously employed in some Kansas smelters
to produce high-purity lead oxide as well as metallic lead.^ '
Normal-Grade Lead Concentrates (14)
Most lead-zinc ores will not yield concentrates of the grade required
for ore hearth smelting. In addition, these concentrates may contain
significant quantities of iron, copper, arsenic, antimony, silver, gold,
bismuth, and cadmium, which either must be removed to produce a suit-
able quality of lead, or which in themselves are worth recovering.
For these reasons the blast-furnace path is the most widely used smelt-
ing method in the lead industry.
A typical composition of lead concentrate fed to a blast furnace is
shown below:
Pb
Zn
S
Fe
Ag
Cd
Insoluble
Percent
32
5
10
9
0.05
5
Oz/Ton
30-150
Smelting and Refining Processes (15 through 25)
The smelting and refining processes include charge preparation (blend-
ing of the concentrate with flux, return products, etc.) (15), sinter-
ing (16), blast-furnace smelting (17), and subsequent refining opera-
tions to remove and perhaps recover metallic impurities. The unit pro-
cesses are indicated in Figure 14.
90
-------
Lead
Concentrates
Flux
^
< ' Du s t
"\ Fume ! Collector
Sinter^
\ fume ^
_s^~ Uust
Collector
x Solids
1 *" Gra
Stack
A
^ Stack
Recycle
to
F Blast
urnace
Granulation
Settler
( Water v
I To Waste
Partial
Recycle
Settler
lag- _____ |
*
Cooling
Water
Lead
Islag
Dross
Copper
Matte
Dressing
Plant
Gold-Silver-
Zinc-Lead
Zinc Fuming
Furnace
D ,.
Refjnery
Antimony, Arsenic
and other Impurities
Water
i >- Fume to Baghouse
J ^ Slag to Waste
Pure
Lead
FIGURE 14. GENERALIZED F1.0WS1IEET OF LEAD SMELTER
91
-------
Charge Preparation. General practice in charge preparation (15) may
involve the blending of lead concentrates with fluxes and a variety of
loose-end products, such as fumes, slags, etc., which contain recover-
able lead and other metals. This blended mixture is pelletized after
the addition of moisture (up to 10 percent) by rolling in rotating
drums or on inclined disks to form spherical pellets about 1/2 inch or
more in diameter. The pelletized concentrates are then sintered.
Sintering. Sintering is done on a "sintering machine" which, in
essence, is a travelling-grate furnace. Figure 15 illustrates the
general structure and function of a downdraft type of sintering
machine. The lead industry in the United States uses both this type
and also "updraft" sintering procedures, wherein a positive pressure
of air is supplied from below the travelling grate of the sintering
machine and slightly reduced pressure is maintained above the bed. In
the operation, a layer of pelletized concentrates are laid down on the
bed of the sintering machine and are ignited by burners above the bed.
The balance of the charge (up to about 1 foot deep) is spread on the
burning layer, and the travelling grate then enters the windbox section.
Under the effect of the applied updraft, the bed burns from the bottom
up. The usual United States practice relies on the sintering reaction
to remove nearly all sulfur from the concentrates. The sulfur burns
off as 502 in the first portion of the sintering operation. Most
plants divide the exhaust gases from sintering into two portions. The
gases from the initial portion of sintering contain 4 to 8 percent S02>
are passed through gas-cleaning systems (usually wet scrubbers) and
then vented to the atmosphere through a stack. These wet gas scrubbers
reportedly operate with closed-circuit scrubbing-water systems, in
which the dust- and fume-laden scrubber water is clarified by settling
and reused in the scrubbers.
The objectives of the sintering operation are not only to remove sulfur
as S02 and S03 and to eliminate, by volatilization, much of the unde-
sirable impurities such as arsenic and antimony, but, equally as impor-
tant, to produce "sinter" of suitable size distribution and strength
for subsequent treatment in the blast-furnace process.
The product of the sinter machine is screened and fine material is re-
circulated to the sintering process. The sinter product is next pro-
cessed in the lead blast furnace with additions of coke and fluxing
agents.
The sintered product fed to the blast furnace will vary in composition
from plant to plant. The following tabulation is an example of a
finished-sinter composition.
92
-------
a Pallets
b Ore
c Ignition furnace
d Wind box
e. Suction box
f. Stack
g Feed hopper
h Clean out doors
i Direction of travel
D^ight-Lloyd Sintering Machine
J L
f 5,33
b
B. Igniter
b Suction box
c. Suction pipe
d Moving ore bed
Tr^-.^^-^f'/l , J .
\ ,
0 /
/
e. Feed hopper
f. Sinter
g Ore
h. Zone of reaction
.:."!',>. /X
I ^ '
^ /
/
i
\
\
Section Through Dwight-Lloyd Sinter Bed
FIGURE 15. DIAGRAMS OF THE OPERATION OF SINTERING MACHINES
(14)
93
-------
Percent Oz/Ton
Cu 3
Pb 36
Fe 10
CaO 11
Insol 10
S 1.4
Zn 10
Cd 0.04
Silver -- 30-150
Blast Furnace Smelting. The lead blast furnace itself is a water-
cooled vertical furnace, rectangular in shape, 22 to 28 feet in length
and 15 to 20 feet in height, and may range in width from 5 to 10 feet,
sometimes being tapered from 10 feet wide at the top to a minimum
width of 5 feet. Associated facilities include charging devices, an
exhaust-gas-handling system, an air-supply system of bustle pipes and
tuyeres for the introduction of air to the charge at several levels,
and a refractory-crucible structure at the bottom with provisions for
continuous or intermittent tapping of lead bullion and slag. Figure 16
is a diagrammatic sketch of a typical lead blast furnace.
The charge to the blast furnace always includes sinter, coke, and flux-
ing or slagging additions such as silica and limestone, and generally
includes recycled slag from associated operations, cadmium residues, re-
finery dross, and fume from dust-collecting equipment.
The chemical reactions occurring in the blast-furnace process are com-
plex and numerous :
CO + C -> 2CO
PbO + CO -» Pb + CO
Cu 0 + CO -> 2Cu + CO
Sb 0 + SCO + 2Sb + 3C02
As 0 + 3CO -> 2As + 3C02
CO + 2Fe3°4 + C02
Fe00. + CO + 3FeO + C00
34 2
FeO + CO -> Fe + CO
PbS + Fe -> Pb + FeS
-------
FRONT VIEW
SIDE VIEW
21
a. Hearth or crucible
b. Tuyeres
c. Bustle pipe
d. Lead well
e. Shaft
f. Water jackets
g. Charge floor j. Slag spout
h. Arents siphon tap k. Relief valve
i. Charge doors
FIGURE 16. DIAGRAMMATIC SKETCH OF A TYPICAL LEAD BLAST FURNACE
95
-------
PbO + Fe Pb -I- FeO
2PbO + PbS 3Pb + SO
PbSO. + PbS 2Pb 4
4 L
The products of the blast furnace are:
(a) Base bullion (18), which normally may contain quantities of copper,
arsenic, antimony, or bismuth, which must be removed by further process-
ing to produce an acceptable lead. The lead bullion also may contain
precious metals in quantities worth recovering. The composition of the
base bullion will vary from plant to plant, but will in general contain
95 to 99 percent lead with impurities ranging as follows:
Copper -- up to 2.5 percent
Zinc -- negligible
Antimony -- up to 2 percent
Arsenic -- up to 1 percent
Bismuth -- up to 0.03 percent.
(b) Slag, consisting of iron, calcium, and magnesium silicates, sub-
stantially all of the zinc in the original charge (up to about 15 to
20 percent zinc), small quantities of arsenic and antimony, and vari-
able amounts of lead (1.5 to about 4 percent). If the slag contains
more than about 6 percent of zinc, it is generally treated for zinc re-
covery.
In this case, the slag, usually while still molten, is charged to a zinc
fuming furnace, commonly a reverberatory-type furnace, with or without
additions of other zinc-bearing materials (other recycled drosses,
dusts, etc.). The charge is heated to a high temperature through addi-
tion of fuel (coal) and air blown into the molten slag, and the zinc is
boiled off and oxidized to zinc oxide dust particles which are collected
in dust-collecting equipment such as cyclones, precipitators, dust cham-
bers, and baghouses.
Lead blast-furnace slags also may be granulated by impacting a stream
of molten slag with a high-pressure water jet. The granulated slag may
be dewatered and either recycled as part of the charge materials to the
sinter process, or, depending on slag composition and plant facilities,
may be totally discarded.
The slag-granulating water circuit is usually closed to allow recovery
of the slag in settling tanks with the clarified water recirculated.
Water loss by evaporation during granulating may be made up with waste
water from other plant operations such as cooling-tower blowdown or neu-
tralized effluent from an associated acid plant.
(c) Matte. In some blast-furnace practice, a matte phase consisting
of copper and iron sulfides may be formed as a discrete liquid layer
96
-------
between bullion and slag and may be isolated. This material is usually
sent to copper smelters for further treatment.
Refining Operations. Impure base bullion, containing copper, antimony,
arsenic, precious metals, and bismuth, requires additional refining.
After production of the base bullion (18) the preliminary refining
steps, dressing (19) and softening (20), are always carried out in se-
quence. Thereafter, two routes to refined lead are available: fire re-
fining (21) and electrolytic refining (22). These operations are dis-
cussed in the following sections. The descriptions given here are sub-
ject to the variations of individual plant practice.
Dressing (19). Dressing is performed in vessels referred to as kettles.
The kettles are generally hemispherical in shape, up to 24 feet in
diameter, and are constructed of welded steel plate up to 1-1/2 inches
thick, holding up to 250 tons. The kettles are gas heated with exter-
nal refractory-brick insulation. Permanent auxiliary equipment for
stirring, skimming, transfer of products, etc., is generally provided.
The major purpose of dressing is to remove copper. The separation of
copper is effected by lowering the temperature of the metal close to,
but slightly above, the melting point of lead. At this temperature
the solubility of copper in lead is minimal and excess copper is re-
jected from the melt to form a crust or head on the melt. This is sep-
arated by skimming from the liquid lead. In some practice, sulfur is
added to the dressing kettle to enhance the removal of copper as copper
sulfide, which also appears in the crust. By dressing, the copper con-
tent of the lead is reduced from as high as several tenths of a percent
to as low as 0.005 percent. The liquid lead is then transferred to a
second kettle, where a second decoppering cycle may be performed. The
dross, which may typically contain about 90 percent lead oxide, 2 per-
cent copper, and 2 percent antimony, with entrained gold and silver, is
re-treated in a by-product reverberatory furnace to recover lead as
bullion and to produce a copper matte for subsequent treatment by cop-
per smelter practice.
Softening (20). After dressing to remove the major portion of copper,
the bullion is then subjected to a "softening" operation for removing
antimony, arsenic, and tin. The term "softening" is used because the
removal of the impurities, principally antimony, produces a product of
lower hardness and strength. In contrast, lead alloyed with antimony
is commonly referred to as "hard lead" or antimonial lead.
The softening may be done in either of two ways, either by air oxida-
tion of the molten bullion in a reverberatory furnace or by oxidative
slagging with a flux of sodium hydroxide and sodium nitrate.
The air-oxidation process consists of treatment of drossed lead in a
reverberatory-type furnace with air introduced into the bath through
pipes or lances. In the air-oxidation method of softening, most of the
impurities are removed in a primary slag which is skimmed off. The
aeration is continued with the formation of a final slag. This two-
97
-------
stage slagging permits the maximum degree of removal of impurities. The
slag produced contains the oxides of copper, arsenic, antimony, and tin
as complex oxides (lead stannate, lead arsenate, lead antimonate) and
entrained metal, and is further treated to recover antimony, antimony
oxide, antimonial (hard) lead, a tin-rich skim (sold to tin-recovery
operations), and sodium arsenate, which is generally discarded.
After softening by the air-oxidation, reverberatory-furnace treatment,
the lead bullion is drained from beneath the slag and treated by fire
or electrolytic refining.
An alternative method of softening is an oxidative slagging technique
in which a sodium hydroxide-sodium nitrate mixture is stirred into the
molten lead to oxidize arsenic, antimony, tin, etc., which enter the
slag as arsenates, antimonates, and stannates of sodium. At least two
versions of the oxidative slagging process are used, the kettle process
and the Harris process. The major difference between them is that, in
the kettle process, the resulting slag is discarded to waste, while in
the Harris process the slag is extensively treated by a hydrometallur-
gical process to recover sodium hydroxide and, where indicated, arsenic,
antimony, and tin products.
Figure 17 illustrates the various modifications of the softening step.
Fire Refining (12). The fire refining of lead has as its objective the
removal of silver, gold, and bismuth, the bulk of which metals are not
removed in dressing and softening. Figure 18 illustrates the fire re-
fining process. It is comprised of the following steps:
(1) Desilvering by the so-called Parke's process in which the softened
lead is treated with zinc metal at about 900 F, with stirring, for sev-
eral hours. The zinc combines preferentially with gold and silver to
form zinc-gold and zinc-silver compounds which are virtually insoluble
in lead and which are lighter than lead. Enough zinc is added to com-
bine with all the sold and silver and to saturate the lead with zinc.
The zinc-precious no I'M alloys accumulate on the surface and are skimmed
off. By conducting the desilvering in two stages as is indicated in
Figure 18, it is possible to isolate high-gold and high-silver products
to facilitate the recovery of these metals in subsequent processing.
(2) After removal of the gold and silver, zinc is removed by a process
called vacuum dezincing, conducted in a separate kettle, using a port-
able bell-shaped vessel whose open bottom is lowered into the liquid
lead to form a seal and allow evacuation of the space above the melt.
The upper portion of the dezincing chamber contains a condenser, a
stirrer which extends down into the melt, and connections to the vacuum
system. As vacuum is applied to the chamber, any zinc in the lead
leaves the melt by vaporization and condenses on the condenser.
Electrolytic Refining (24). Electrolytic refining of lead utilizes
anodes cast from partially refined or softened lead. The process
98
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involves the transfer of lead from the impure anode through a hydro-
fluosilicic acid electrolyte and deposition on the cathode, consisting
of a starting sheet of electrolytical ly refined lead. Impurities or
contained metal values determine the feasibility of electrolytic re-
fining. The by-products include bismuth, gold, silver, selenium, tell-
urium, antimony, and arsenic.
Formerly, the electrolytic process was the only feasible method for pro-
ducing refined lead with an extremely low bismuth content. In late
years, a fire-refining technique for debismuthizing lead has provided
some competition for the electrolytic process in this respect.
The cast anodes are suspended in cells alternating with starting sheets
with approximately 2 inches between electrodes. Anodes weighing about
400 pounds and approximately 3 feet long by 2 feet wide are suspended
alternately between starting sheets in wood or concrete tanks lined
with asphalt. A typical cell is about 2-1/2 feet wide by 4 feet deep
and 8 feet long. The cells are arranged in banks as in copper electro-
lysis (see section on copper). Except for the nature of the metal, the
electrolyte, and electrical conditions, the operation is similar to
copper refining.
The electrolyte used in lead refining is fluosilicic acid, t^SiFg,, and
is usually operated at temperatures of about 100 F. This acid is highly
corrosive and thus rubber or plastic piping is used for its circulation.
The electrolyte is prepared in the plant by the reaction of sulfuric
acid with fluorspar to produce hydrofluoric acid which reacts with
silica to form hydrof luosilicic acid
CaF- + H-SO, + CaSO, + 2HF
224 4
6HF + Si00 -> H0SiF, + 2H00.
226 2
The active compound in the electrolysis is identified as lead fluosili-
cate, which forms when the lead reacts with fluosilicic acid:
2H0SiFr + 2Pb + 0_ -» 2PbSiF£ + 2H O.
26 i D 2.
In operation, the cast anodes and starting sheets are placed in the
tanks and electrolysis is carried out for about 4 days, after which all
electrodes are removed from the cell and washed. The cathodes, refined
lead, are melted and cast into product forms. The anodes are washed
to recover electrolyte adhering to their corroded surfaces and to re-
move slimes, and then returned to the electrolytic cell for production
of a second cathode. Each anode of about 400 pounds will yield two
refined- lead cathodes weighing about 180 pounds each.
The slimes of residual impurities which separate during the electroly-
sis contain arsenic, antimony, gold, bismuth, silver, etc. The slimes
are dewatered by centrifuging or filtering and the solid residues are
treated for the recovery of valuable materials. Slimes treatment is
101
-------
described in other sections of this discussion.
Refined Lead (22.25)
Refined lead produced by fire refining or electrolytic refining is
melted and cast into pigs. Various grades, based on specifications,
are produced. These are shown in Table 23.
TABLE 23. CHEMICAL SPECIFICATIONS FOR COMMERCIAL PIG
Silver _. .minimum percent
IvOpper.. __ maximum percent
Copper , . . __ minimum percent.-
Silver and copper together
maximum percent, .
Arsenic. maximum percent. .
Antimony and tin together
maximum percent
Arsenic, antimony and tin together
maximum percent.
Iron.. _ maximum percent
Btemuth . maximum percent
Lead (by difference) _ . minimum percent
Corroding
lead'
0 0015
0015
. 0025
.0015
.0095
0015
002
05
99 9-1
Chemical
lead"
n 09n
. 002
080
. 040
. 002
001
' 002
00,3
99 90
Add
lead '
n no9
080
. 040
002
001
002
025
99 90
Copper
lead '
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015
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in
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i CnrrodlnR !c;id Is a de-5l(in Hlon that h;is been used Tor many ytars In
th? trade to describe lead wh.ii.ti has been refined to a hlgli (iejr^a o(
purity.
' Chsmloiil lead lias been iiofd for many years In ttw trads to describe
the undPsUveriz«'d iftad produced from southeastern Missouri ore-s.
' Acid lead Is nude by adding copper to fully reflncd lead.
4 Copper lead Is made by adding ropper to fully reftnod lead.
' C
-------
pyrometallurgical and the hydrometallurgical, involving leaching and
electrolysis. Sulfur removal is always required, regardless of the
path followed.
The purpose of roasting zinc concentrates is to eliminate most of the
sulfur in sphalerite and to convert zinc sulfide to zinc oxide or zinc
sulfate. When the pyrometallurgical route is to be followed, as com-
plete removal of sulfur as possible is advantageous. If the hydro-
metallurgical route is followed, the objective of roasting is to con-
vert the acid-insoluble sphalerite to acid-soluble zinc oxide with con-
trolled amounts of zinc sulfate, and to avoid the formation of zinc
ferrite. The main reactions occurring in roasting are:
2ZnS + 302 -> 2ZnO + 2SC>2
2ZnO + 2SO + 2ZnSO,.
These reactions are controlled to produce the desired amount and form
of residual sulfur, depending on whether the roasted product is in-
tended for pyrometallurgical reduction or for electrolytic reduction to
metal.
Currently used methods of roasting are
(1) Multiple-hearth furnace roasting
(2) Flash roasting
(3) Fluidized-bed roasting.
In multiple-hearth furnace roasting, the concentrates are fed through
a large cylindrical furnace with a number (up to 12) of circular
hearths one above another. Rotating arms with attached rakes move the
charge of concentrates inward and outward on alternate hearths so that
the charge falls (through ports) from one hearth to the next, counter-
current to a stream of hot gases. The roasted calcine product is dis-
charged at the bottom of the furnace. Oxidation of the charge supplies
some heat to support the process, but additional fuel in the form of
coal, oil, or gas is usually supplied. Temperature control to avoid
the formation of ferrites and regulate the amount of zinc sulfate is
achieved by introducing cooling air at certain hearth stages.
In flash roasting, the concentrates are showered into a stream of hot
gases in large cylindrical furnaces. As the concentrates descend, oxi-
dation is extremely rapid, much more so than in the multiple-hearth
furnace chamber.
In fluidized-bed roasting, the concentrates are kept in agitated sus-
pension by a stream of upward-flowing combustion gases. The suspended
particles form a "fluid" bed, virtually homogeneous with respect to
temperature and composition. Concentrates are fed to fluidized-bed
103
-------
roasters either dry or in slurry form; fluidizing air and combustion
gas is supplied from a blower to a windbox below the bed. The fluid-
izing gas is distributed by means of a perforated plate or a multitude
of small pipes at the bottom of the bed to ensure constant upward velo-
city through the cross section of the bed. As feeding of concentrates
progresses, the bed overflows through a discharge port. Fluidized-bed
roasters may be up to 20 feet in diameter. Bed depths will range to 6
feet deep when in a fluidized condition. Capacities up to 10 tons per
hour have been obtained with equipment of this size.
One of the major advantages of fluidized-bed roasting is the excellent
control of temperature it affords. This is important in the avoidance
of ferrite formation and the regulation of the sulfur content of the
roasted product.
Equipment auxiliary to roasting operations is (in the usual operating
sequence): waste-heat boilers, dust-catching equipment (usually cy-
clones), electrostatic precipitators, gas-cooling and/or -cleaning
towers, and sulfuric acid plants. Gases produced in the roasting of
zinc concentrates contain from about 9 to 13 percent SO- and these are
sent to sulfuric acid plants.
Sintering (26)
Another method of oxidizing the sulfur from zinc concentrates is by the
Dvight-Lloyd sintering process. The sintering machine has been de-
scribed previously in this section (Figure 15, page 93). In zinc metal-
lurgy, unlike lead metallurgy, sintering is done almost exclusively by
downdraft methods. Roasted or raw sulfide concentrates, mixed with
other charge components (e.g., coal, silica sand, lime), are deposited
on the first portion of the bed through hoppers or nozzles. The charge
is carried on the travelling grate beneath burners which serve to ig-
nite the upper surface of the bed, and then over a windbox which serves
to draw air and combustion gases down through the burning sinter bed.
After the sintering reaction, the sinter product falls from the end of
the returning sinter plates and is sized for use in the following reduc-
tion operations. In the sintering of zinc concentrates, which are apt
to contain cadmium, chlorides in some form (either salt or chloride
solutions from cadmium recovery operations) may be intermixed with the
sinter feed. These chlorides assist in the volatilization and removal
of lead, cadmium, and silver during sintering.
Sintering accomplishes the aim of agglomerating the zinc-bearing mater-
ials in a product whose size and strength are suitable for the pyro-
metallurgical furnacing process. In most plants, a selected part of
the sinter product is recirculated to assure proper removal of lead,
cadmium, and sulfur, and to achieve the desired degree of densification.
104
-------
Retorting (27) and Condensation (28)
Retorting is a reduction-volatilization process in which zinc oxide,
produced by roasting and briquetting, by roasting and sintering, or by
sintering alone, is reacted with coal or coke at high temperature.
The reactions which occur during retorting are:
ZnO + C -> Zn + CO
2ZnO + C -* 2Zn + CO
ZnO + CO -> Zn + CO
CO + C -> 2CO
Fe 0 + 3C -> 2Fe + 3CO
Fe 0 + C -+ 2FeO 4- CO
ZnS + Fe -> Zn + FeS
Zn + CO -> ZnO + CO (blue powder formation)
Zn.SiO, + 2C -> 2Zn + 2CO + Si00
24 2
FeO + SiO? -> xFeO-SiO (slag formation).
At the temperature of retorting (about 1200 C), the reduced zinc is
vaporized and carried out of the retort with carbon monoxide to con-
densing systems. Three methods of retorting are used in the United
States. They are:
(1) The batch horizontal retorting process
(2) The continuous verticc.1 retorting process
(3) The electrothermal retorting process.
The horizontal-retort plant consists of a long narrow furnace into
which fire-clay retorts are inserted crosswise to the length of the
furnace as shown in Figure 19. Up to 600 retorts are arranged in banks
on either side of the heating chamber. The retorts are cylinders up to
9.5 inches in diameter and about 5 feet in length with wall thicknesses
of about 1 inch. They are made of carefully selected mixtures of re-
fractory clays. The butt end of each retort is closed. The open outer
end is "luted" or sealed by fireclay packing to a refractory condenser
which collects the evolved zinc vapor until it is removed. Metal ex-
tensions to the condensers, called prolongs, may be used to collect
vapor that escapes the condensers.
105
-------
Retort for zinc distillation as set in furnace
with condenser and prolong
o Air melrt
b Gas main
c Retort
d Condenser
e Prolong
/ Working door
3 Residue pit
h chuta
\ Rear lining
j Front "
k Tie rods
I Residue car
m Rails for charging rncchlna
ladles, etc.
fit
(/) pn)
TT^H
-;^. i -s.
Horizontal distillation furnace
FIGURE 19. ILLUSTRATIONS OF THE ELEMENTS OF THE HORIZONTAL RETORT OPERATION1'17^
106
-------
The charge to the retorts varies from plant to plant, but as an example,
the charge at one plant consists of roughly 57 percent sinter, 22 per-
cent zinc oxide, 19 percent coal and coke, and small amounts of dross,
salt, and fluorspar. Charging is done mechanically. The retorts are
heated in a controlled cycle to a maximum charge temperature of about
1200 C and then allowed to cool before emptying and recharging. The
cycle from charge-to-charge may take 1 to 2 days. Zinc metal, called
spelter, is periodically removed from the condensers during the cycle.
The spelter, in addition to zinc metal, may contain some lead, and most
of the cadmium in the charge. Part of the lead is removed later in re-
fining. During retorting, some of the zinc is superficially oxidized
by C02 (reaction 8 above) in the condenser, does not coalesce to enter
the liquid phase, and is collected as a dust called blue powder. This
is recirculated to retorting.
Recovery of zinc in horizontal retorting may exceed 90 percent. After
retorting is finished, the residues, amounting to about 25 percent of
the weight of the original charge, and containing from 5 to 15 percent
zinc, are removed. If the residue is sufficiently rich in lead, gold,
and silver, it is usually shipped to a lead smelter for recovery of
these values.
The continuous vertical-retort reduction process is carried out in a
large (roughly 1 by 5 by 25 feet high) retort constructed of silicon
carbide within a combustion chamber, with a refractory-brick outer
covering, heated by suitably positioned gas burners. The charge to the
furnace consists of briquettes containing a compacted mixture of anthra-
cite coal and zinc calcine. The briquettes are prepared by mixing fine-
ly ground coal and calcine with a tar or similar binder, compressing
the mixture into small pillow-shaped forms, and heating these forms in
a furnace to a temperature of about 900 C. The heat-treatment opera-
tion hardens the briquettes and removes volatile materials.
The prepared briquettes are charged at the top of the vertical-retort
furnace through a movable-lid arrangement. As the charge progresses
downward through the retort, the reduction reaction proceeds, liberat-
ing zinc vapor and carbon monoxide. The zinc vapor is led from the
upper portion of the retort through a refractor)? passage to impinge up-
on and condense in an agitated pool of molten zinc. The furnace gas
(largely CO) is drawn through the condensing chamber into a gas scrubber
and recycled to the burners. The liquid zinc product is transferred to
holding furnaces or distilling columns. Recovery of zinc in vertical-
retort furnaces exceeds 90 percent, and blue-powder production amounts
to only about 3 percent of the zinc charged, considerably less than in
horizontal batch retorting. The residue also contains considerably
less zinc than that from batch retorts.
The electrothermic process of the St. Joe Minerals Corporation utilizes
a relatively hard, strong sinter and coke as the charge to the top of
the furnace, which is roughly 8 feet in diameter and 30 to 40 feet in
height. The power input to the furnace consists of electrical power
107
-------
supplied through three pairs (upper and lower) of electrodes which re-
sults in a flow of electrical current directly through the charge, pro-
ducing a resistance-heating effect. The vaporized zinc is collected by
means of an annular ring at about mid-height on the furnace. The zinc
vapor is drawn into a U-shaped condenser filled with molten zinc. The
exit side of the U-shaped tube condenser is connected through a gas
washer to a vacuum system. The furnace gas is essentially carbon mon-
oxide which is used as a fuel gas for other plant operations. The zinc
metal is continuously bled from the condenser and led to purifying dis-
tillation columns.
Figure 20 illustrates the electrothermic retorting equipment.
Auxiliary operations for recovering zinc from miscellaneous materials
such as zinc-bearing lead blast-furnace slags and low-grade materials
(as concentrator tailings, retort residues, plant sweepings, etc.) in-
clude slag fuming and Waelz furnacing.
Refining (29)
Some portion of the output of a zinc smelting furnace may require sub-
sequent refining, depending on such factors as the current market for
various grades of zinc, the concentrates used, the quantities of metals
other than zinc (cadmium, lead, gold, silver, etc.) present in the
spelter, or the conditions of the particular smelting process. Refin-
ing may be accomplished by any of three methods: liquation (30), dis-
tillation (31), or electrolysis (34). Electrolysis is discussed later.
Liquation. Liquation is the term used to denote a process of separat-
ing impurities on the principle of liquid solubilities and freezing
points. In the liquation process, impure molten zinc is cooled to a
temperature just above its freezing point. As the temperature is de-
creased, the solubilities of lead and iron in the zinc decrease. As
the zinc cools, the lead and iron are, to a major degree, rejected
from the liquid zinc. The lead and iron rejected constitute intermedi-
ate products, which may be further treated for the recovery of the
metals or may be transferred or sold to another plant. The purified
zinc is cast into salable form.
Distillation (31). Distillation also is used to purify the spelter
zinc and to separate and recover lead and cadmium as by-products. Dis-
tillation is carried out in vertical fractionating columns. These col-
umns are constructed of silicon carbide and enclosed in gas-fired heat-
ing chambers. The impure spelter zinc, containing lead and cadmium, is
fed into the top of the first distillation column, which is maintained
at a temperature below the boiling point of lead but above the boiling
point of zinc and cadmium. Lead containing small amounts of zinc and
cadmium is drawn from the bottom of the first column. Most of the zinc
and cadmium vaporizes and passes from the first column through a heated
108
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109
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and insulated carryover to a second column which is maintained at a
temperature below the boiling point of zinc and above the boiling point
of cadmium. The zinc condenses as a pure liquid and is discharged from
the bottom of the second column to casting or alloying operations. The
cadmium fraction in the overhead from the second column and the impure
lead-zinc-cadmium bottoms from the first column are treated for the re-
covery of these metals as described in the section dealing with primary
cadmium.
Electrolytic Refining (32 through 35). The electrolytic process pos-
sesses several advantages over fire-refining methods. It is applicable
to low-grade concentrate. Impurities are readily removed and can be re-
covered as valuable products, and zinc of as high purity as desired can
be made.
In the electrolytic reduction of zinc to metal, the calcines are
leached to dissolve the soluble salts of zinc and other metals, the sol-
ution is purified, and pure zinc is plated from the leach liquid. The
spent electrolyte, after electrolysis of the zinc, is returned to the
leaching step. An example of process steps is presented in Figure 21.
The process steps involve first the leaching of the calcine with dilute
sulfuric acid. Leaching may be done continuously or by batch. If batch
leaching is done, tankage capacity in the leaching section must be about
doubled to compensate for time lost in filling and emptying tanks. Con-
tinuous leaching has the disadvantage of requiring a feed of uniform
composition with respect to impurities and, for this reason, batch
leaching is more widely employed; leaching is generally done in Pachuca
tanks (Figure 22).
Two general types of leaching, single and double, are employed.
In single leaching, the calcines are treated with spent electrolyte un-
til all of the zinc is dissolved in a still-slightly-acid medium. Lime
is then added to precipitate all of the iron, silica, antimony, and
arsenic present. Extremely good control is required to avoid the loss
of significant quantities of zinc by coprecipitation. Double leaching
avoids the necessity for such control. In double leaching, the cal-
cines are leached with a deficit of spent electrolyte so that the re-
sulting solution is safely on the basic side, and impurities such as
iron, arsenic, etc., remain undissolved. Clarified leach liquor from
this first stage is sent to the purification step. The residue, still
containing some zinc, is separated and sent to a second stage of leach-
ing where it is treated with a slight excess of acid in spent electro-
lyte. The resulting acidic solution is then combined with additional
spent electrolyte and recirculated to the first stage of leaching. In
either case, the leach liquor is clarified in thickeners and filtered.
The filter cake is generally shipped to a lead smelter for the recovery
of lead and precious metals which do not dissolve in leaching. The
filtered solution is treated with zinc dust in the so-called purifica-
tion step (33) to precipitate cadmium, copper, nickel, and cobalt and
110
-------
Ferric iron solution
Sand
Dorr Classifier
-Overflow Sand
To Smelter
Calcine
*
New and spent electrolyte
Leaching Tank
Sand Cone
~T~
Overflow
Dorr Thickener
J }
Overflow Underflow Wash Water
Zinc Dust
Purification Tanks
Filter
t *
Filtrate Cake
Press
To Re-treatment
Cake Filtrate
1 I
To Re-treatment Electrolytic Cells
t
Cathode
Melting Furnace
I
f t
Slab Zinc Skimmings
To Market To Smelter
FIGURE 21. TYPICAL FLOWSHEET OF AN ELECTROLYTIC ZINC
111
-------
refiltered. This filter cake may be treated to reclaim cadmium or
other metals. The filtrate may be re-treated the same way to reduce
the impurities to even lower levels. The purified solution is sent to
the electrolytic cells (34).
(d)
a Shell
6 Stand pipe
c Solution or water
d Compressed air
e Angle irons
/ Inlet of air lift
g Outlet
Ji Discharge flange
FIGURE 22. THE PACHUCA TANK
(17)
The electrolytic deposition of the zinc from the purified zinc sulfate
solution follows generally the process described previously in this re-
port for the electrowinning of copper. Insoluble anodes (lead alloyed
with about 1 percent silver) and aluminum starting-sheet cathodes are
the electrodes used. The anodes are 2-1/2 x 4 feet and are 3/8-inch
thick. The aluminum starting sheets are approximately the same length
and width and are about 3/16-inch thick. Tanks are usually constructed
of concrete with lead linings. Zinc electrowinning operations are
classified by the current density used in the cell and the acid concen-
tration, with the lower ranges of the combination being 20 to 40 amperes
per square foot and 6 percent acid strength. High ranges of the two
parameters are 100 amperes per square foot and 25 to 30 percent H2S04
acid concentrations.
The zinc deposited on the aluminum starting sheets is periodically
stripped to obtain a slab of zinc metal. These anode slabs are subse-
quently melted and cast into commercial forms of slabs or pigs to form
the product of the primary zinc industry. The melting operation may in-
clude the addition of, for example, 3-1/2 percent aluminum to the pure
zinc to produce pigs or slabs of alloyed zinc for later remelting in
zinc-die-casting shops.
112
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SECTION IX
THE CADMIUM INDUSTRY
Size and Characteristics
There is no discrete primary cadmium industry. Primary cadmium metal
and compounds are by-product materials produced by companies princi-
pally engaged in the recovery of zinc either as a major value or as a
by-product itself. This is because there is no separate ore of cadmium
and because it occurs in significant quantities in nature exclusively
in association with zinc sulfides. Major cadmium producers in the
United States are shown in Table 24. (
TABLE 24. PRODUCERS OF CADMIUM METAL
(1)
American Smelting and Refining Co,
American Zinc Co.
The Anaconda Co.
Blackwell Zinc Co.
The Bunker Hill Co.
Eagle Picher Inds. Inc.
National Zinc Co.
The New Jersey Zinc Co.
St. Joe Minerals Corp.
United Refining and Smelting Co.
Denver, Colorado
Corpus Christi, Texas
East St. Louis, Illinois
Great Falls, Montana
Blackwell, Oklahoma
Kellogg, Idaho
Galena, Kansas
Bartlesville, Oklahoma
Palmerton, Pennsylvania
Monaca, Pennsylvania
Franklin Park, Illinois
In 1969, the production of cadmium from various sources was distributed
as follows^1):
Source
Zinc recovery operations
Furnacing
Electrolytic
Imports, scrap
Total
Pounds
6,800,000
4,000,000
1.846.000
12,646,000
Two and one-half million pounds of cadmium was produced in the forms
of compounds, sulfides, lithopone, sulfoselenide, and cadmium oxide.
The consumption pattern for 1969 was as follows
(1)
113
-------
Total
Consumption,
Form percent Application
Metal 60-70 Plating, Ni-Cd batteries, alloys
Sulfide pigments 12-15 Plastics, paints, enamels, lac-
quers, inks
Other compounds 15-20 Vinyl stabilizer, phosphors in TV
and fluorescent lamps, semi-
conductor compounds, plating
bath additives
The total value of cadmium production in 1969 was about $40,000,000.
There has been and will undoubtedly be more effort on the part of en-
vironmentalists to force substitution of other materials for many uses
of cadmium owing to its acknowledged toxicity.
What effect this will have on future cadmium markets will depend on
whether technical and engineering forces can convince lawmakers that
perfectly safe practices for containing cadmium toxicity have been de-
veloped and are readily applicable.
Raw Materials and Processes
The starting material for cadmium-recovery processes are by-products of
other metal-production operations, all involving zinc. These are
(1) Fumes and dusts from the roasting and sintering of zinc concen-
trates
(2) Zinc metal containing relatively high concentrations of lead and
cadmium
(3) Dusts from the smelting of lead-zinc ores or copper-lead-zinc ores
(4) Purification sludge from zinc electrowinning plants.
Cadmium From Zinc Concentrates
In the recovery of zinc from zinc concentrates by the retort process,
zinc sulfide concentrates are roasted, sintered (see zinc section), and
introduced into retorts with reducing agents such as coal or coke. When
the mixture is heated, the zinc oxide is reduced to zinc metal which is
volatilized, condensed, and cast into ingots. Unless special provisions
are taken, much of the cadmium and lead almost always present in zinc
concentrates would accompany the zinc. Figure 23 outlines a process
used to separate these elements from the zinc and at the same time
114
-------
Zinc Sulfide Concentrates
I
Return Sinter
Roasting (1)
->
Fue 1
Finished Zinc
Sinter (4)
To Retorting (5)
Lead Sulfate (9)
T j T> /lr,\
Lead Recovery (10)
Flue Dust (la)
Sintering (2)
Cadmium (3) and
Lead Fume
t
Dust Collection (6)
odium Chlorate
Sulfuric Acid
Leaching (7)
Filtration (8)
T
Cadmium Sulfate
Solution (11)
i S
Zinc Dust
Precipitation (12)
Filtration (13)
1
Cadmium Sponge (14)
Metallic Cadmium (18)
Casting (19)
Chloride
Solution
(21)
Residue
(20)
FIGURE 23. CADMIUM RECOVERY IN THE ZINC RETORT PROCESS
(18)
115
-------
recover cadmium.
Roasting (1) is done in multiple-hearth furnaces under oxidizing condi-
tions to eliminate most of the sulfur. At the temperature of the roast-
ing operation, cadmium tends to volatilize preferentially and concen-
trates to some extent in the flue dust (la). This is collected in suit-
able equipment such as cyclones, baghouses, and electrostatic precipi-
tators and is returned to join the roasted product. The roasted concen-
trates and flue dust are then charged onto the surface of a sintering
machine (2) (a travelling grate furnace) which transports a bed of the
material about 1 foot deep beneath a burner where the charge is ignited.
Suction boxes beneath the grate draw the hot gases down through the
charge as it travels from the feed to the discharge end.
The objective of sintering is to agglomerate the calcine, remove resi-
dual sulfur, and remove most of the lead and cadmium from the zinc.
This lead-cadmium removal is accomplished by chloridizing the charge
with the zinc chloride solution (21) returned from filtration step (13)
or by the addition of sodium chloride solution. Under proper condi-
tions of temperature and chloride addition, cadmium and lead chloride (3)
volatilize from the charge and are drawn down through the bed and into
the fume-collecting facilities for subsequent treatment. The sintered
zinc oxide product (4) is sent to the retorting furnaces (5) for the re-
covery of zinc metal. In some cases, the lower few inches of the sin-
tered bed of material, which may still contain some lead or cadmium, is
separated and recirculated through the sintering operation (4a).
After collection (6) the fumes and dusts are leached (7) with dilute
sulfuric acid and sodium chlorate. The purpose of the sodium chlorate
is to ensure the complete dissolution of cadmium sulfide which would
otherwise remain undissolved. In the leaching operation, cadmium and
lead are converted to sulfates and/or chlorides. Cadmium sulfate and
chloride remain in solution, but lead is almost completely converted to
insoluble lead sulfate. This compound, together with other insoluble
materials (quartz, silicates, etc.) is filtered out and sent to lead
recovery. The solution of cadmium sulfate (and/or chloride) (11) is
then treated with powdered zinc to precipitate cadmium (12). Precipita-
tion is carried out with a deficit of zinc, so that the resulting cad-
mium will be as zinc free as possible. The slurry is filtered (13).
The solution from filtration, containing practically all of the zinc
added and about 10 percent of the cadmium as chlorides and sulfates, is
returned to the sintering operation to provide the necessary chloride
ion (21) .
The filter cake, called sponge (14), if properly prepared, will contain
about 30 percent moisture and only small amounts of lead and zinc.
This sponge is then steam dried (15), mixed with coal or coke and lime
(16), and transferred to a retort furnace where the cadmium is reduced
to metal, vaporized, recondensed, and cast into required shapes (17, 18,
19). Residue from the reduction furnace or retort containing some
116
-------
cadmium (20) is returned to the sintering operation.
Cadmium From Zinc Metal
Some zinc made by the retort process contains both cadmium and lead in
appreciable quantities. An example is the so-called prime Western zinc
which may contain about 1.5 percent lead and several tenths of a per-
cent cadmium. Such a material can be highly refined by the process
shown in Figure 24. This process involves two fractionation stages,
each at different temperatures. In the first, operated at a tempera-
ture in excess of 1000 to 1150 C, cadmium and zinc are boiled off,
leaving a high lead-zinc alloy. The cadmium-zinc vapors are condensed
and refractionated (2) at about 800 C. In this operation, cadmium is
volatilized and a high-purity zinc product is withdrawn at the bottom
of the fractionation column.
The cadmium product from this process generally requires further treat-
ment by additional volatilization or chemical means.
Cadmium Recovery From Lead Smelter Dusts
Smelter dusts, which customarily contain about 5 percent cadmium or
less, as produced, may be treated in several ways to recover cadmium.
Figure 25 shows three typical routes. In the first route (left side
of Figure 25), the starting material is dusts from the smelter bag-
houses (1). In the case of burned baghouse dusts, the dusts are col-
lected in bag filters and dropped into masonry or refractory chambers
called cellars where they are ignited and allowed to oxidize by burning
or smoldering. This material, after continued burning, will contain
about 15 percent cadmium, 15 percent zinc, and 50 percent or more of
lead. This is sent to a lead smelter where it is refumed (la) in a so-
called "Godfrey" furnace with fuel and limestone at a temperature regu-
lated to more or less selectively volatilize cadmium. The fume col-
lected in a baghouse (2) typically may contain up to 55 percent cadmium
and about 15 percent each of lead and zinc. This material (3) is gener
ally sent to an electrolytic zinc refinery for further treatment.
Another route is shown in the center column of Figure 25. By this route,
the baghouse dust containing about 3 to 4 percent cadmium is refumed in
a reverberatory furnace in several batch stages (5) in order to obtain
a fumed product of optimum grade for subsequent processing (6). This
fume, which may contain 40 to 45 percent cadmium is contacted (7) with
water and oxidizing agents (KMnO^ and NaC103) in a ball mill and leached
by agitation with additional sodium chlorate, ferrous sulfate, and sul-
furic acid. The purpose of the oxidizing agents and the ferrous sul-
fate is to ensure that cadmium is all dissolved and that arsenic is oxi-
dized and precipitated as ferric arsenate. The slurry is filtered (9);
117
-------
Lead-
Zinc -<-
Alloy
Zinc Metal
(Impure)
Melting
Fractionation (1)
V
Zinc-Cadmium
Vapors
J
Condensation
Zinc-Cadmium
Liquid
V
Fractionation (2)
Refined Zinc
99.99% Pure
Cadmium Vapors
I
Condensation
Cadmium-Zinc Metal
or Cadmium Dust
To Purification
FIGURE 24. CADMIUM RECOVERY FROM ZINC METAL
118
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the residue, consisting of lead sulfate and ferric arsenate, is sent
back to lead smelters.
Precipitation with zinc dust (10) under controlled conditions of acid-
ity produces a fairly high-grade "sponge" (11). This is filtered,
washed by repulping with water, refiltered, and dried (12). The filter
cake is then granulated and formed into small briquettes (13). The bri-
quettes are then retorted batchwise in graphite crucibles (14) to vola-
tilize cadmium. The cadmium vapors are condensed and the liquid metal
cast into bars. These are remelted under a layer of sodium hydroxide
flux in electric furnaces. If thallium is present, the sodium hydrox-
ide flux is removed and ammonium chloride is added to remove thallium
from the cadmium (15). The cadmium is then recovered with sodium
hydroxide flux and the purified metal, which may contain 99.95 percent
of cadmium, is cast into bars.
The third route, shown at the right side of Figure 25 is also applicable
to lead and zinc smelter fumes. In this process the fume or dust is
leached by repulping in water (17) and the slurry is treated with chlor-
ine gas. Subsequent steps of filtration (19), precipitation, and iso-
lation of cadmium sponge (20 and 21) are similar to those previously
described. The resultant sponge may be subsequently treated by electro-
lytic or pyrometallurgical methods to produce pure cadmium.
Cadmium Recovery From Purification
Sludge in Zinc Electrowinning
The electrowinning process for zinc is described in the section of this
report dealing with that metal. The successful electrodeposition of
zinc requires extremely pure solutions. Cadmium, among other metals,
such as copper, nickel, cobalt, and iron, must be virtually completely
removed. In zinc metallurgy, both copper and cadmium are eliminated
from solution by treating the zinc electrolyte with powdered zinc.
Generally this is done stagewise. In the first stage, only enough zinc
metal is added to precipitate copper, which, being more noble than cad-
mium, reacts preferentially with the zinc powder. After removal of the
copper precipitate by filtration, the zinc electrolyte is then treated
with additional zinc dust to precipitate the cadmium. This precipitate,
high in zinc, contains virtually all the cadmium originally in the elec-
trolyte and constitutes the so-named purification sludge. Its process-
ing is shown in Figure 26.
This purification sludge (1) is first allowed to oxidize (2) in air,
with or without heating, to render the cadmium readily soluble. The
oxidized material is then leached (2) with spent electrolyte from the
electrolysis step. After filtration (4) to remove insoluble copper
introduced from the zinc electrolyte, the filtrate is stage-treated
with zinc dust (6, 7, 16, 17, 18. 19) to minimize zinc concentration in
the cadmium sponge (8). This sponge containing about 80 percent
120
-------
Spent
Electrolyte
(22)
^<-
Purification Sludge (1)
Y
Oxidation (2)
Leaching
Filtration (4)
->- Copper Cake
Filtrate (5)
I
First Precipitation
Filtration (7)
Filtrate
(16)
Cadmium Sponge (8)
(807o Cd 57o Zn)
Oxidation (9)
Excess
Zinc Dust
.with
excess
zinc
Leaching (10)
Filtrate
(19)
To Zinc Leaching
(20)
Filtration (11)
Filtrate (12)
Electrolysis (13) << Glue, etc.
Cadmium Cathodes (14)
Melting (15)
T
Sha
Cast Shapes
FIGURE 26. CADMIUM RECOVERY FROM PURIFICATION SLUDGE OF ZINC
ELECTROWINNING OPERATIONS(19)
121
-------
cadmium with 5 percent zinc, is again oxidized by steam drying to en-
hance the solubility of cadmium (9). It is leached (10) in spent elec-
trolyte from the electrolysis circuit and filtered (11). The filtrate
(12), containing about 200 grams per liter of cadmium as sulfate, is
then electrolyzed to produce cadmium cathodes.
Electrolysis (13) is carried out in banks of cells similar to zinc
cells, described under that section of this report. These cells are
fitted with lead anodes and aluminum cathodes. In the electrolysis,
cadmium plates onto the aluminum cathodes from which it is periodically
stripped. After passage through the cells, the spent electrolyte is re-
turned to various leaching stages in the circuit (22).
Stripped cathode metal is washed, dried, and melted under a flux and
cast into various shapes, including ingots, balls for use as electro-
plating anodes, etc.
122
-------
SECTION X
BY-PRODUCT GOLD AND SILVER
Recovery of Gold and Silver From Copper Ores ' '
Figure 27 outlines routes for the recovery of silver and gold.
Silver and gold follow copper through all stages of concentration,
smelting, and fire refining (1). When copper is refined electrolytic-
ally, however, these metals together with selenium, tellurium, etc.,
collect as the finely divided solids in the electrolytic tank. These
are called slimes (2) (see the section on copper). These slimes, as
produced, contain high percentages of fine copper and are first leached
in hot dilute sulfuric acid (3) to dissolve excess copper, and some
tellurium and selenium, away from the silver and gold (4). Typically,
such leached slimes will contain about 10,000 ounces of silver per ton,
with varying proportions of gold, lead, selenium, arsenic, antimony,
etc., depending on the characteristics of the original ore (5). The
treated slimes are then smelted in a small reverberatory furnace, called
a Dore' furnace (6), with various fluxes such as limestone, borax, fluor-
ite, silica, etc., selected to produce a fluid slag of the base metals
so that they may be separated from the bulk of the molten silver and
gold. This slag (7) normally contains significant amounts of gold and
silver, and to recover these values it is usually returned to the cop-
per smelter. The gold and silver alloy, now rid of most of its base
metal and metalloid impurities, is called Dore metal (8). Typically,
such an alloy will contain 90 or more percent of silver, gold, platinum,
and a small amount of copper. The molten Dore metal is cast into small
anodes (10). These are electrolyzed in small specialized cells in
nitrate solution (11). Two main types of electrolytic cells are used.
In one type, the Thum cell, carbon cathodes constitute the floor of the
cell, and the impure silver anodes are suspended in a shallow receptacle
with a cloth bottom. On electrolysis in this type of cell, the silver
dissolves anodically and is deposited in small crystals on the bottom
cathode. These crystals are raked either manually or mechanically from
the cell, washed, and dried (13). During the electrolysis, gold and
platinum do not dissolve but collect as slimes on the cloth bottom of
the anode receptacle (12). These are sent to further processing for
the recovery of these metals.
The electrolytically refined silver crystals (13) which will exceed 99.9
percent purity are subsequently melted and cast into bars weighing about
100 pounds. Another type of cell, called the Moebius cell, is also used
in silver refining. In this cell the anodes are enclosed in cloth bags
and suspended alternately between cathodes. During electrolysis in the
nitrate solution, silver is anodically dissolved and deposits as
123
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loosely adherent crystals on the cathodes. These are periodically
mechanically scraped off into a basket on the bottom of the cell. The
gold and platinum slimes in the Moebius cell are retained in the anode
bags.
Recovery of Gold and Silver From Lead Concentrates
The recovery of gold and silver from ores in which lead and zinc are the
major components by weight is less straightforward than the recovery
from copper ores. Silver follows copper through to the electrolytic re-
fining step in copper metallurgy. In lead metallurgy, however, silver
may follow several routes. After lead concentrates are sintered (15)
and smelted in the blast furnace (16), a portion of the silver will
accompany the copper matte that is normally formed. This matte (17) is
returned to copper smelters for the recovery of both silver and copper
as outlined in Column I of Figure 27. Much of the silver, however,
accompanies lead in a bullion (18). Bullion from most ores contains
enough gold and silver to make their extraction profitable. In addi-
tion, it contains undesirable impurities such as copper, zinc, tin,
antimony, and arsenic which must be removed by refining. The process-
ing steps for refining lead will vary with the composition of the im-
pure bullion and the end product desired. The steps shown in Column II
from the dressing step on are typical.
Dressing (19) consists of holding the molten bullion at a temperature
just above the melting point, during which operation copper rises to
the top and is skimmed off. The last traces of copper are removed by
adding sulfur. The copper dross is returned to a copper smelter for
copper recovery and the recovery of whatever silver may have accompanied
the copper dross. Arsenic, antimony, and tin are subsequently oxidized
and are also skimmed from the surface of the lead.
After dressing, two routes are available for the production of refined
lead. One, essentially fire refining, includes a silver- and gold-
recovery procedure involving the addition of zinc metal to the molten
lead (21, 22, 23, 24, 25, 26). When zinc is added in the desilvering
step, the precious metals alloy with the zinc and rise to the surface.
The last traces of zinc in the lead are removed either by vacuum distil-
lation or by fluxing with caustic soda (23). The zinc fraction or zinc
crusts (25) removed from the lead are then "cupelled" (26) in a small
reverberatory furnace under strongly oxidizing conditions to convert
the lead to molten lead oxide which carries off other materials besides
gold and silver. Lead oxide with its burden of impurities such as zinc,
arsenic, antimony, etc., is returned to the lead blast furnace. The
molten gold and silver remains in the cupel furnace as Dore metal, and
is cast into anodes for treatment as shown in Column I of Figure 27
(10,11).
Another path for producing refined lead after the dressing operation is
125
-------
by electrolytic refining (29) which is described in the section of this
report dealing with lead. Electrolytic refining is generally employed
when undesirably high concentrations of bismuth are present. During
the electrolysis of lead, silver and gold, as in copper electrolysis,
separate as slimes. These are collected and subjected to cupellation
(26) and electrolysis (27) as described previously.
Refining of Precious Metals
The impure gold recovered by any of the above processes is refined by
treatment in the so-called Dore furnace. In this operation, the mater-
ial is melted with appropriate fluxes (soda ash, borax, silica, etc.)
under oxidizing conditions to produce a gold-base alloy which may still
contain silver and the platinum-group metals and a slag which will con-
tain impurities such as copper, zinc, etc.
If platinum-group metals are absent or virtually so, the Dore'' metal is
purified, while still molten, by the Miller process, in which chlorine
is bubbled through the charge. This treatment volatilizes base metals
which may be present and converts silver to molten silver chloride salt
which rises to the top of the melt and can be poured or skimmed from
the gold. Refined gold made by this process generally contains about
99.6 percent gold, and is suitable for many purposes.
If platinum metals are present or if higher purity is desired, the Dore
metal is cast into small anodes and electrolyzed in chloride solution by
a miniaturized method analogous to the refining of copper by electroly-
sis. In this process, the Wohlwill process, gold is electrolytically
oxidized as the anode passes into solution and is deposited in pure
form on the cathode. The resulting cathode is melted and cast into
bars of 99.9+ percent purity. In the Wohlwill process, silver is also
electrolytically oxidized at the anode, but is quickly and almost com-
pletely converted to insoluble silver chloride.
126
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SECTION XI
THE ARSENIC INDUSTRY
Size and Characteristics
The production of arsenic in the United States is almost exclusively as
a by-product from the smelting of copper and lead ores. In the roasting
and smelting of such ores, arsenic oxidized to As203, is volatilized,
and is recovered in fumes and flue dust. These contain varying amounts
of the trioxide (up to 30 percent) associated with the oxides of other
metals such as copper, lead, antimony, zinc, or tin, and various other
particulate materials.
The most marketable form of arsenic is the pure trioxide, with little
demand for pure elemental metallic arsenic. In the United States, the
only primary metal producer now making either arsenic trioxide or ar-
senic is the Tacoma smelter of the American Smelting and Refining
Company (the only producer in 1969).(1) In earlier years Anaconda also
produced arsenic in Montana. No production data have been published in
recent years. United States imports in 1969 amounted to 18,171 tons
and had decreased from about 27,000 tons in 1967.
The method of producing the trioxide from smelter flue dusts consists
of sequential volatilization operations, with operating temperature
being the method of controlling the selective volatilization of the
oxide from flue dusts and the purity of the product.
Thus, processing consists of roasting the flue dusts to volatilize the
contained As2C>3, which concentrates in the fume from the roasting oper-
ation. The collected fume is again heated in a reverberatory furnace,
and the fume from this operation is passed through a series of chambers
to collect fractions of various purity from the volatilized furnace out-
put. The purity level is determined by cooling-chamber temperature.
These steps are indicated in Figure 28.
127
-------
Smelter Flue Dusts from Reverberatory Furnaces
(Up to 301
Reprocess
90%
To Market
Collected Fume
90 - 95% As00,
Reverberatory
Furnace
Refining
Fume
Cooling
Chambers
or
"Kitchens"
95% As 03
Collected at Furnace End
99-99.9% As '0,
to Market
FIGURE 28. STEPS IN THE CONCENTRATION AND PURIFICATION
OF ARSENIC TRIOXIDE(21)
128
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SECTION XII
THE PRIMARY ANTIMONY INDUSTRY
Size and Characteristics
The primary antimony industry is relatively small. In 1969, total pri-
mary antimony production in the United States was 13,203 tons, about 17
percent of world production.(I/ Salient production data are given in
Table 25. The consumption and uses of antimony as metal and compounds
are shown in Table 26.
Facilities producing antimony in the United States and their activities
are as follows:
Sunshine Mining Company
Coeur D'Alene, Idaho
Several properties adjacent
to above
Stampede Mine, Kantishna
District, Alaska
Gold Creek Mine
Elko Co., Nevada
Smokey Claims
Humboldt Co., Nevada
Stibnite Mine, Souders Co.,
Nevada
Wells Fargo Mine, Stevens Co.,
Washington
Various lead and lead-zinc mines
with antimony as by-product
National Lead Co. Smelter,
Laredo, Texas
American Smelting and Refining
Co.
Mining, concentrating, and electro-
winning of antimony metal
Mining
Mining and concentration with ex-
port of concentrates
Mining, shipment of ore to National
Lead Smelter, Laredo, Texas
Mining, shipment of ore to National
Lead Smelter, Laredo, Texas
Idle in 1969, but sold to the U.S.
Antimony Corp.
Development work with stockpiling
of antimony silver ore
See section on lead
Pyrometallurgical production of an-
timony metal and antimony oxide
Principal producer of by-product
antimony at Omaha, Neb., and
Perth Amboy, N.J., plant
129
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TABLE 25. SALIENT ANTIMONY STATISTICS*^1)
(Short tons)
United States:
Production:
Primary:
Price: New York, average cents per pound
World: Production - ~
1 Includes primary antimony content of antimonial
1965
845
12.389
24 321
14
14,879
16,919
45.75
69,456
1966
927
1-1,539
24,258
29
19,712
19,681
45.75
67,627
1967
892
12,466
23,664
82
17,419
17,350
45.75
63,565
lead produced at primary lead
1968
856
12.4X9
23 , 699
109
17,343
18,520
45 75
67,737
smelters.
1969
938
13,203
23,840
207
17,032
17,843
57.57
72,059
TABLE 26. INDUSTRIAL CONSUMPTION OF PRIMARY ANTIMONY IN
THE UNITED STATES, BY CLASS OR MATERIAL
PRODUCED(L)
(Short tons, antimony content)
Product
METAL PRODUCTS
Antimonial lead
Solder -- - - -. -
Other - -
Total...
NONMETAL PRODUCTS
Plastics
Other - - -
Total
Grand total.- -
1965
36
6,382
821
68
76
49
104
244
642
214
8,636
16
46
1,971
1,853
W
855
1,469
477
1,596
8,283
16,919
1966
154
6,285
731
164
62
44
107
155
515
219
8,435
27
50
3,188
2,074
832
2,224
870
1,980
11,245
19,681
1967
209
5,539
653
141
54
31
118
184
382
223
7,534
30
43
3,454
1,884
665
1,785
948
1,007
9,816
17,350
1968
156
6,817
755
178
46
50
105
255
423
258
9,043
33
37
2,774
2,037
859
2,318
440
979
9,477
18,520
1969
115
6,723
758
55
33
56
105
242
541
137
8,765
37
30
2,096
2,108
722
2,558
433
1,094
9,078
17,843
W Withheld to avoid disclosing individual company confidential data; included with, "Other."
130
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McGeen Chemical
M and T Chemical Co.
Harshaw Chemical Company
Foote Mineral Co.
Hummel Chemical Co.
Antimony oxide producer
Antimony oxide producers
Antimony oxide producers
Antimony sulfide producers
Antimony sulfide producers
Raw Materials and Processes
The most important antimony mineral is stibnite, Sb2S-:j, although vari-
ous oxide minerals may be associated with stibnite deposits in small
amounts. Antimony ores, containing stibnite and minor amounts of oxide
minerals, occur in a few instances in deposits in which antimony is a
major value, as in Idaho and Nevada. From such deposits, it is possi-
ble to obtain high-grade ore (+40 percent antimony), by selective min-
ing, and concentrates containing 15 to 30 percent antimony. Most of the
domestic antimony production comes from imported ores treated at the
Laredo, Texas, antimony smelter. Other production is based on antimony
derived from the processing of lead ores.
Mining and Concentration
Only a few mines are operated principally for the recovery of antimony
alone and these are generally small. Most of the antimony ore bodies
are not amenable to large-scale mining methods employed in the mining
of copper, zinc, and lead, owing to the size, irregularity, and distri-
bution of the ore veins. These are mined selectively, ore and waste
being removed selectively to prevent dilution. A larger scale, open-
pit gold-tungsten-antimony mine was formerly operated in Idaho, but
this was closed down in 1969. The bulk of the mine production of anti-
mony in the United States is from the Sunshine silver mine in Idaho.
This is a unique operation in which antimony in the silver concentrates
is leached and won electrolytically as high-purity metal. Most of the
primary antimony mined in the United States is a by-product of lead
ores.
Concentration of antimony ores in this country now appears to be chiefly
by selective mining and hand picking. Flotation concentration has been
employed successfully in the past for complex ores in which antimony
was a major coproduct.
131
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Metal-Reduction Processes
The methods selected for the production of antimony metal or oxide are
largely determined by Lhe grade of the ore or concentrates. The lowest
grade ores or concentrates (5 to 25 percent antimony) are generally
treated by volatilization processes; those containing 25 to 40 percent
antimony are smelted in a blast furnace; high-grade ores containing 45
to 60 percent antimony may be treated by a process called liquation, in
which the antimony is reduced, melted, and drained from the gangue mater-
ial; special conceicrat.es, such as the silver-antimony concentrate pro-
duced by the Sunshine Mining Company in Idaho, are treated by leaching
and electrowinning. Condensed oxides from volatilization processes,
liquated antimony sulfide, and some intermediate-grade ores are smelted
in reverberatory furnaces.
Volatilization. Volatilization processes depend on the volatility of
antimony trioxide and are especially suitable for low-grade ores (5 to
25 percent antimony).
In the volatilisation process, the ore is heated with carbonaceous
material under controlled conditions of temperature and aeration to con-
vert stibnite (Sb2S3) to volatile antimony trioxide (85203). The oxide
is condensed and recovered in suitable dust-collecting equipment. After
collection, the oxide is reduced to metal by subsequent furnacing with
coke under reducing conditions. By careful manipulation of conditions,
volatilization processes are capable of producing, high-purity oxides in
a single step.
Liquation. The liquation process requires high-grade ores running 50
to 60 percent antimony. The objective of liquation is to me it, selec-
tively, the stibnite (85283) and to drain it from the gangue materials.
The product, antimony sulfide, may be sold as such or it may be subse-
quently converted to antimony metal. The residues resulting from liqua-
tion still contain significant quantities of antimony and these are
generally treated by volatilization processes for more complete recovery.
Liquation may be carried out on a batch basis in liquating furnaces
which are simply clay crucibles with perforations in their bottoms to
permit outflow of the melted antimony sulfide. Tube furnaces and re-
verberatory furnaces, which permit continuous operation, also may be
used.
Blast Furnace Smelting. The National Lead jmeluing operation at Laredo,
Texas, utilizes this method. The charge to the blast furnace consists
of mixed oxide and sulfide ores from mines in Mexico or the United
States, and by-products of other smelting operations; mattes, sl;igs,
flue dusts, or residues from lead and zinc refineries. The blast-fur-
nace operation is somewhat similar to that of lead blast-furnace prac-
tice. A vertical or snafu furnace with, a rectangular cross section is
used, with low-pressure air being supplied through tuyeres. The ore
132
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and residues, charcoal, and fluxing additions are charged at the top of
the furnace and the products of molten antimony metal and slag are
tapped from the hearth at the bottom of the furnace. The blast-furnace
process is capable of producing extremely pure metal with minimal anti-
mony loss.
Reverberatory Furnace Smelting
Impure oxides produced by volatilization processes, liquated antimony
sulfides, residues, and ores containing antimony and impure metal may
be converted to high-purity metal by reverberatory-furnace smelting,
either batch or continuous. In this operation, the antimony-bearing
charge is heated with coke and such fluxing materials as soda ash and
sodium sulfate. Coke reduces the oxides of antimony to metal and the
soda slag removes gangue minerals and sulfur. If iron is present in
the charge, it reacts with sulfur to form a matte between the metal and
slag. Considerable antimony is volatilized as antimony oxide during
reverberatory-furnace smelting. This is collected in suitable equip-
ment and returned to the furnace.
Leaching-Electro lysis
Currently, primary antimony metal is being produced at the Sunshine
Mining Company in Idaho by a leaching-electrolysis process. The pro-
cess is depicted in Figure 29. In this process, a complex copper-anti-
mony sulfide concentrate is leached with sodium sulfide solution which
dissolves the antimony as sodium thioantimonate (Na^SbS^). The leach
solution is clarified by settling and filtration, and electrolyzed in
diaphragm cells to yield antimony of 93 to 99 percent purity. During
the electrolysis, some sulfide sulfur is oxidized to thiosulfate or
sulfate. With recirculation of the electrolyte, the concentration of
these oxidized forms of sulfur would build up to the point where they
would interfere with the solubilization of antimony in the leaching
step. By treating spent electrolyte with barium sulfide, these sul-
fates are kept at acceptable levels, and the sulfide concentration is
maintained. The barium sulfate and thiosulfate precipitate is filtered
and heated with coal in a rotary kiln to reconstitute barium sulfide
for reuse.
133
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Cu, Sb, Ag Ore
1.
Leaching
Solution
Clear
Solution
Electrolysis
Electrolytic
Antimony
Metal
Residue (Cu, Ag)
Impurity
Precipitation
Electrolyte
To smelters
Residue
BaS
FIGURE 29. DIAGRAM OF PROCESS STEPS IN THE PRODUCTION OF
ANTIMONY AT SUNSHINE MINING OPERATION
Coal
134
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SECTION XIII
THE PRIMARY BISMUTH INDUSTRY
Size and Characteristics
In 1969, three producers of bismuth were reported operating in the
United States. These were the Omaha plant of the American Smelting and
Refining Company; the East Chicago plant of the U.S. Smelting and
Refining Company; and the Texas City Smelter and the Los Nietos
(California) Refinery, both operated by the Gulf Chemical and Metal-
lurgical Corporation.'^' No production figures are available.
The consumption and uses of bismuth are given in Table 27. Its uses in
lowmelting alloys include applications in fuses, temperature-sensitive
devices, and mold and die materials for forming plastics and other
metals. Bismuth is used as an alloying addition to cast irons, alumi-
num alloys, amalgams, and bearing alloys. The largest use is in cos-
metics in the form of bismuth oxychloride.
TABLE 27. CONSUMPTION OF BISMUTH IN THE UNITED
STATES IN 1969 (POUNDS)
Low-Melting Alloys 748,393
Aluminum, Cast Iron, and Other Alloys 523,710
Pharmaceuticals and Cosmetics 1,250,539
Experimental Uses 252
Other 9,065
Total 2,531,959
Raw Materials and Processes
There is no mineable bismuth ore in the United States, and bismuth pro-
duction is tied firmly to the production of copper and lead. Bismuth,
present in the ores of these metals, follows them through concentration
and all the preliminary steps of processing to be finally removed in
lead-refining steps. In copper smelting, for example, bismuth persists
with copper until the converting operation, in which much of it volati-
lizes with lead, antimony, etc., as fumes. These are collected and
sent to lead smelters. The portion of bismuth which escapes volatiliza-
tion during converting is finally removed in the electrolytic copper
135
-------
refinery in the slimes. Treatment of these slimes for the recovery of
gold and silver leaves bismuth associated with lead. Bismuth associ-
ated with lead ore accompanies lead through all steps of processing and
must be removed by special refining processes discussed in the section
of this report dealing with lead.
Refining of bismuth is done by chlorination. When the bismuth has been
removed from lead by the pyrometallurgical procedure, it is collected
in a dross containing calcium, bismuth, magnesium, and lead. This
dross is melted and chlorinated to remove calcium and magnesium as a
molten chloride slag and a bismuth-lead alloy. The alloy is fluxed with
molten sodium hydroxide to remove traces of arsenic and tellurium that
may be present. Residual silver and gold are removed by treatment with
zinc, as described in the section of this report dealing with lead. The
resultant lead-bismuth-zinc alloy is then subjected to chlorination
treatment with chlorine gas under controlled conditions. Zinc and lead
combine with chlorine preferentially over bismuth and are removed as
molten chloride slags. The molten bismuth is finally treated with air
and caustic soda and cast. This method of refining is said to yield
bismuth 99.999 percent pure.
During the electrolytic refining of lead, bismuth enters the slimes
with any arsenic, silver, gold, or antimony that may have been present.
These slimes are filtered, dried, and melted to produce a slag and
metal. The metal phase is blown with air to remove arsenic and anti-
mony. It is then cupelled, i.e., heated under oxidizing conditions, to
produce a lead-bismuth slag and a Dore' alloy of gold and silver. The
lead-bismuth slag is reduced by carbon to form a lead-bismuth alloy
which is refined by chlorination as described above.
136
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SECTION XIV
WATER USAGE IN THE LEAD-ZINC INDUSTRY
Nineteen firms responded to the canvas of the lead-zinc industry for
information on water usage. The data are shown in Table 28, which in-
dicates the types of operations reported on (mine-mill, smelter, etc.)
and the total intake and discharge volumes.
Data on water usage, reelrculation practices, and usage per ton of metal
reported by lead and zinc producers are given in Table 29.
The water-use data for lead and zinc mines permits the following general
conclusions and observations.
(a) Intake and discharge quantities are related to a number of factors.
The principal determinants are plant capacity and the availability (and
presumably the cost) of water.
(b) The overall recycle ratio of total water use to raw water intake is
generally low (1, no recycle, to 1.5, 50 percent recycle), although in
one case (No. 14, Table 29) a recycle ratio of 2.6 was reported.
(c) The volume of water reported used in processing and auxiliary oper-
ations per ton of metal for both lead and zinc mine-concentrator combi-
nations ranged between 10,660 to 21,600 gallons in a reasonably close
pattern. The spread is probably attributable to differences in ore
grade and slight variations in plant practice.
(d) There "; s wide scattering in the data showing the distribution of
intake water to process, cooling, etc. Specific reasons for this are
not clearly discernible from the reports supplied to the study. The
variability i;; associated *7ith plant size and a host of other factors
such as location, ore grade climate conditions, water availability,
cost, etc.
The data for ;inc and lead smelter discharges also reflect the effects
of such variables.
Recycle ratios tend to be high in most of the lead and zinc smelter
operations, reflecting the large percent usage for cooling, the usual
operation of cooling towers and cooling-water circuits, and the trend
to economize on the use of water treated chemically to protect tubes
and piping.
In five of the smelter operations included in the data listed, sulfuric
acid plants were operated in association with the smelters, and the oper-
ation of these plants apparently exerts no consistent recognizable
137
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effects on the external- or internal-water-use patterns. The variations
in individual plant systems, design, and practice have greater effects
on water usage than the presence or absence of an acid plant.
The lead and zinc smelter-refinery combinations reported relatively
lower water requirements. In four out of the five zinc or lead refin-
eries responding, and in the industry in general, smelting and refining
are carried out in the same plant. Three out of the five operations for
which data are listed include sulfuric acid plants. The water-use char-
acteristics of the zinc refining operations (17 and 18) are distinguish-
able one from the other on the basis of climate and type of operation.
The electrolytic zinc refinery (17) is in an arid climate and shows low
intake, low discharge, very high use and consumption, and high recircu-
lation rates. The other zinc refining operation (18), located in an
area of plentiful rainfall, which refines zinc by distillation, shows
much higher intakes and discharges but lower use and consumption.
Data on the lead smelter and refinery operations (10 and 11) show the
effects of the difference of plant design and capacity on water usage.
Both are situated in the same general area of the country and employ
blast-furnace smelting and fire refining. The second plant listed (11)
operates a sulfuric acid plant in conjunction with smelting and refining.
Data on a single lead-refining operation (12) indicated that water was
used only for cooling and sanitary purposes. All water was obtained
from and returned to a municipal supply and sewer system.
Waste-Water Sources and Characteristics
The sources of waste water in the lead-zinc industry are listed in the
following tabulation:
Water Treatment
Clarification sludges
Sand filter backwash
Water-softening backwash or sludges
Ion-exchange regeneration wastes
Sanitary Wastes
Process Water (smelters and refineries)
Slag granulation waste water
Gas scrubbing waste water
Spent electrolyte
Direct Cooling Water
Steam jet ejector condensate
Cooling Waters
Indirect once through cooling water
Cooling tower blowdown
Miscellaneous
Boiler blowdown.
140
-------
The quantity of waste water discharged from lead and zinc plants has
been discussed in the previous section. The compositions of waste
waters and the receiving streams reported in the survey are shown in
Table 30.
Although these data represent only a small segment of the lead-zinc
operations, they provide an insight into the natures and types of wastes
that are generated by the industry.
Effluents from the tailings ponds of mine-concentrator facilities appar-
ently do not pose serious problems in general. There is a suggestion
however that the effluents from some concentrators might contain concen-
trations of cadmium and lead at or near the level of concern.
Smelter and refinery effluents have a much greater pollution potential
than those from concentrators with respect to both cadmium and lead.
Waste-treatment practice in the smelter-refinery segment of the indus-
try ranged from virtually none, through simple settling, to neutraliza-
tion- floe culat ion- settling.
Table 31 summarizes the treatment practices reported by the lead-zinc
contributors to this study.
Water Costs
Data obtained on water costs in the lead-zinc industry were sparse,
nondefinitive, and certainly not representative of the industry. Water
costs for the group of nonferrous industries covered in this section of
the report are discussed in a later section.
Current Plans
Lead-zinc producers in many cases echoed copper producers in reporting
current plans for waste-treatment improvement and water management.
Increased reuse of water by recycling, the substitution of polyelectro-
lytes for lime as a flocculant for tailings, the complete recycle
through settling and cooling ponds for smelter wastes, segregation of
contaminated process streams from clean water streams, neutralization
of sulfuric acid plant wastes, recycling of wash water and electrolyte
spills in lead and zinc refineries, an investigation into the feasibil-
ity of deep-well injection, all these are proposed and in some cases
are being presently pursued.
141
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144
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SECTION XV
DISCUSSION
In the foregoing sections of this report an effort has been made to
group appropriately the findings of the study on water usage; waste
water sources, characteristics and quantities; waste treatment prac-
tices; and research needs.
In the belief that a general overview of the segments of the nonferrous
industry dealt with in this section will provide an additional basis
for insight into the general water and waste problems, this summarizing
discussion section is included. It covers:
(1) Waste water treatment practices
(2) Water costs
(3) Current problems and future research needs
(4) Current plans of the industry.
Waste Water Treatment Practices
The data obtained have shown that the approximately 50 plant operations
discussed in this report currently discharge slightly over 56-1/2 bil-
lion gallons of water per year. The characteristics of these waste
waters have been discussed in previous sections of this report. This
section of the report will deal in a general fashion with how this waste
water is treated.
Of the 50 operations involved, 14 (or 28 percent) reported zero dis-
charge. These operations, all located in dry climates, employ tailings
ponds to close the water circuit, with high evaporation rates and pos-
sibly ground seepage making discharge unnecessary. Included in these
14 facilities with zero discharge of water water were 12 mines and con-
centrators, 1 copper leaching operation, and 1 smelter. In terms of
treatment methods, all 14 of these plants employed evaporation as a
waste treatment method. The application of this method is limited by
geographical/climate considerations.
Tailings ponds also were used by plants in areas of plentiful water for
sedimentation of solids and in some cases as a reservoir for recircula-
tion of water. These discharged about 38 billion gallons of water per
year or 67 percent of the 56-1/2 billion gallons of total discharge.
Thus, the tailings pond treatment practice appears to be the most
145
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predominant practice, serving not only to provide evaporation and clari-
fication but also to permit recycling of water.
A compilation of the waste treatment practices used in these segments of
the nonferrous industry is shown in Table 32. This listing shows the
types of operations, treatment practices and amounts, the receivers of
discharges and some indication of the characteristics of the waste
waters treated by different methods. Table 33 shows the overall statis-
tics on treatment practices. The data as gathered from the responses
of industry show that 77.7 percent of all waters discharged are given
some treatment by the metallurgical operations included in this study.
Treatment in tailings ponds is, as stated before, the predominant
method of treatment. Other forms of settling, in at least one case
aided by coagulating agents, and dilution before release account for
the next largest amounts of water.
TABLE 33. SUMMARY OF WASTE WATER TREATMENT
PRACTICES
Treatment
Method
Tailings Pond
Other Settling
Dilution
Neutralization
No Treatment
Total Discharge
Evaporated
Amount Treated,
MGY
37,984
2,043
2,016
25
12,798
54,866
248
Percent of
Total
69.2
3.7
3.7
0.1
23.3
100.0
Sanitary wastes were common to all of these operations. Of 48 plants,
intake of 2,116 MGY or about 2 percent of the total intake of 112,900
MGY was used for sanitary purposes. Of 36 cases in which this flow
could be traced, 426 MGY per year of sanitary waste water returned to
the following receivers:
5 plants - municipal sewers 179.4 MGY
6 plants - surface waters 196.8 MGY
3 plants - estuaries 49.5 MGY
All of the water discharged to surface waters and estuaries was treated
in some manner before discharge.
The remainder of all intake "sanitary" water was traced to recircula-
tion, evaporation, or ground seepage. The remaining plants used the
following practices:
146
-------
-------
17 - plant operated septic tanks or lagoons
3 - treat and recirculate to leach operation
2 - treat and recirculate to tailings ponds
Only a few examples of cost data for plant operated sewage treatment
were reported in segregated form and are listed below:
Treatment Method
Package treatment plant
Package treatment plant
Clarifiers, digesters,
oxidation ponds
Aeration, settling, perco-
lation, dilution
Septic tank, chlorination,
oil skimming
Package treatment plant
Amount
Treated,
MGY
5.2
21.6
73
13.5
24.5
19.6
Capital
Investment,
$1000
54
Operating and
Maintenance Costs
$0.11/1000 gal
$0.19/1000 gal
53
$2.77/1000 gal
$0.82/1000 gal
$0.21/1000 gal
Dilution, however, was reported by only one respondent.
It should be pointed out that there is some uncertainty about the cases
shown in Table 32 in which treatment is reported as "none". Many of
the nominally untreated wastes consisting of mixed plant waste waters
may have received some treatment as by mixing or by settling and through
misunderstanding; these were not considered treatments by the respondent.
Water Costs
Data obtained were not adequate to relate water costs to a specific
segment of the industry.
The cost data assembled by this review covered various categories: costs
of supply, specific treatment costs for both intake and waste waters,
and total water costs. The reported total water costs are listed in
Table 34. Intakes are rounded to one significant figure to conceal
specific plant identities as are the larger intakes. With only three
exceptions, capital investment in total water systems for the plants
increases with increasing intake, a not unexpected type of behavior.
However, the reported operating and maintenance costs for the various
water systems show little consistency with size of operation. The low-
est water cost, $0.003/1000 gallons, was reported for the uses of salt
water pumped from and returned to an estuary, and used only for cooling.
The highest cost reported, $0.33/1000 gallons, was for a mining and con-
centrating operation in the southwestern desert region. The water costs
of plants purchasing from municipal supply systems ranged from $0.08 to
148
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$0.26 per 1000 gallons. One respondent defined total water costs as
$0.03 per ton of ore processed by flotation concentration and $0.039
per 1000 gallons of total water pumped.
TABLE 34. REPORTED TOTAL WATER COSTS
Intake, Operating and Maintenance
' Capital Investment, $ Costs, $/1000 gal
<5,000
<5,000
5,000
4,000
4,000
2,000
2,000
1,000
1,000
800
700
500
500
500
400
400
300
300
200
100
100
70
27,883,000
3,533,000
2,713,183
--
950,000
600,000
--
250,000
933,000
__
--
95,000
--
150,000
185,152
--
230,000
43,568
_ _
0.03
0.07
0.04
0.01
0.05
0. 08^b)
0.02
0.12
0.33
0.10
0.27
0.003(c)
0.01
0.08
0.13
0.04(b)
0.26(b)
0.09
0.21
0.06
(a) Rounded to one significant figure.
(b) Purchased municipal water, all other are independent systems.
(c) Obtained from estuary.
The reported costs of water treatment before use are listed in Table 35.
Chlorination of potable water was commonly practiced at most plants with
independent water supplies. Only two plants reported chlorination of
all plant intakes. The two examples of costs for chlorination range
from about one to three cents per 1000 gallons, the cost varying, in
only two examples, inversely with amount treated.
The reported costs for softening and other treatment of boiler feed
water varied from about 1-1/2 to 80 cents per 1000 gallons, with again
large variations and no consistent pattern to the variations. Boiler
feed water treatment practices included (1) the use of condensate from
evaporators, (2) ion exchange demineralization, (3) chlorination, and
(4) chemical softening, using such additions as polyphosphates, sulfites,
149
-------
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sodium chloride, dolomitic lime, and magnesium oxide. The treatment
costs associated with boiler feed water were higher than any other
treatment cost except sewage treatment, were applied to relatively
small fractions of total intakes, and were more often recirculated.
The specific cost data reported for waste treatment processes is listed
in Table 36. Treatment in tailings ponds is the cheapest of the three
treatments shown, showing costs from one to five cents per 1000 gallons
of discharge. The costs of other forms of settling, i.e., those used
by operations other than flotation concentrators were reported by only
two respondents. Reports varied widely.
TABLE 36. WASTE WATER TREATMENT METHODS AND COSTS
Amount
Treated,
MGY
300
500
2500
7000
1000
1000
5
20
20
20
Treatment
Method
Tailings pond
Ditto
11
11
Settling*^3)
r i
Sewage Plant
Ditto
tr
it
Capital
Investment ,
$
1,270,000
50,000
208,377
< 10,000,000
363,000
25,000
54,000
--
53,000
Operation and
Maintenance
Costs ,
$/1000 gallons
0.017
0.048
0.020
0.007
0.097
0.005
0.11
0.19
0.21
0.82
(a) Coagulating agents added.
Treatment costs for sanitary wastes were reported in separate form in
four cases, with the cost as shown previously, ranging from about 10 to
80 cents per 1000 gallons treated.
The major conclusion derivable from a consideration of the reported
data is that a better and more reliable report of costs would require
that all operations be costed on a uniform basis, with considerably
more detailed review of internal costs and operations than was possible
in this study. Some respondents merely noted that water costs were dis-
tributed among various operational costs and not broken down for dis-
tinct operation. It also appeared that most operators consider waste
water treatment an integral part of plant operation, the preponderance
of replies noting total water costs with no breakdown of the specific
costs of waste water treatment.
Certainly, more important than cost or method is the effectiveness of
151
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the treatment. In only a few cases was effectiveness of treatment
method rated, in terms other than "satisfactory". Bound up in the term
"effectiveness" is the concept of external standards. Because of the
current circumstances, what has been "satisfactory" in the past may not
be now or in the near future.
Only two examples of quantitative measurement of effectiveness of treat-
ment were submitted in this survey. The first example quoted here is
from a large copper concentrating and smelting complex with high reported
capital investments in waste treatment facilities and very low operating
and maintenance costs:
Operation
Substance
Concentrator
M
11
II
II
Suspended solids
Dispersant
Collector
Promoter
Frother
Flocculant
Treatment
Settling, lime
Actual process
Precipitation,
settling
Sewage treatment Biological oxygen demand Digestion, fil-
tration
Ditto Suspended solids Settling, fil-
tration
" Pathogenic bacteria Chlorination
Reduction,
percent
99.99 (+)
99.20
99.99 (+)
99.99 (+)
99.00 (+)
99.50
78.00
85.00
99.00 (+)
The second example is from a zinc smelting operation reporting high in-
vestment and relatively high operating and maintenance costs:
"Two parallel systems, each comprised of a number of settling
lagoons, are employed for the removal of zinc-bearing materials
and other suspended matter. Coagulant aids are used to promote
settling efficiency. Chlorides and cadmium-bearing wastes are
also routed through these systems to enhance waste water quality.
Effectiveness of treatment is as follows:
Substance
Suspended solids
Zinc
Cadmium
Chlorides
Percent Removal
98 +
98+'
Undefined
Undefined
Current Problems and Future Research Needs
The current and future research needs expressed by industry covered a
broad spectrum, including:
152
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(1) Firm bases for the establishment of discharge water standards with
particular respect to hazard levels of trace metal concentrations
(2) The development of standardized and adequate analytical techniques
(3) The development of new process technology to reduce discharges by
increasing recycling, and
(4) Improved waste treatment processes for both industrial and sanitary
waste with emphasis on metal and acid removal.
The statements from industry are summarized below, grouped according to
major topic.
Limits and Standards
A more scientific basis for setting metal limits in discharges is
needed. This would require more accurate information on the hazard
levels of heavy metal ions than now seems available. The setting of
limits for the discharge of metals to harbors is a special problem.
Some respondents discharging to harbors stated they were not aware of
any serious problem at present, but if told that the pollutants carried
by water effluents were too high they would take steps to reduce the
leveIs.
Analytical Techniques and Programs
Sufficiently precise and accurate procedures for determining trace com-
ponents of the waste are available. These should be adopted to verify
the very low concentrations of metals, etc., in settling pond effluents
and to determine background concentrations in natural water sources.
The supply of capable inorganic analysts and biologists continues to la£
behind the load being imposed by regulatory agencies in setting stand-
ards and limits on specific contaminants. Undergraduate and graduate
level training and research need to be encouraged and supported in this
most critical area. One respondent indicated that an extensive analy-
tical campaign was begun before start-up of a plant, to establish back-
ground levels of critical components in natural waters of the area.
After start-up the impact of the plant's discharge on water quality
will be assessed to determine compliance with the state laws pertaining
to water pollution.
153
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Specific Problems in Industrial Wastes
Some respondents indicated a definite need for the development of low-
cost procedures to remove small to trace quantities of metals, such as
Zn, Pb, Cd, As, and Sb from effluents to ensure full compliance with
state standards. Others anticipated a need for the improved treatment
and disposal of waste acidic solutions and wash waters from tank house
and by-product plants containing ^804, HC1, ^803, Cu, 804, As, Sb,
and trace metals. Improved oil separation was also cited as a need by
some.
Storm water runoff was cited by one respondent from the southwest de-
sert area as a contributant of acid and suspended solids to receiving
streams. Seepage from lagoons into groundwater was stated to be an ill-
defined but possibly significant problem.
Sanitary Wastes
One respondent indicated that the existing concept of conventional,
biological treatment of human wastes needs to be reexamined for the
possibility of reducing costs involved in collection, treatment, pumping,
etc. Another indicated it would be desirable to have available some
method by which the sewage treatment effluent could be recycled as
cooling water for metal-casting operations.
One general response dealt with the need for incentives to encourage
research by industries to improve technology which would minimize the
quantity and improve the quality of discharges. These incentives would
be supplemental to those now in existence for waste treatment systems.
About 60 percent of the returns from industry indicated no research
needs, either by omission of any reply or by an explicit statement to
that effect. (About half of .these, it may be remembered, have zero
discharge). Probably reflecting the basis for the belief that no re-
search was needed is the statement by one respondent operating a con-
centrator .
"The main treatment will always consist of settlement of
suspended solids, which is completely effective now. A
large disposal area with a settling lagoon will always be
needed, so no research or less costly method is indicated."
A summary of the positive responses discussed above is listed in Table
37.
154
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155
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Current Plans
Fourteen (or 28 percent) of the responding operations gave information
regarding planning and/or the work currently under way to reduce the
volume of waste water or to reduce pollutant levels. The comments from
industrial operations are summarized in Table 38.
As shown in the table, only one copper leaching operation (5) and pos-
sibly one smelting operation (9) have plans which offer any possibility
of increased product recovery. All other plans are directed at waste
water control. No definitive costs were indicated in the responses.
Nine of the fourteen plans include increased recycling and associated
decrease of discharges. Four plans include possible total recycle,
i.e., a closed water circuit with decreased intake and zero discharge.
On the other hand, three tailings-ponds operators state the intent or
necessity of continuing discharges and employing additional treatment
methods. Two of the smelter-refinery type operations indicate plans
for extensive waste treatment or disposal (deep well injection) facili-
ties. Other cases indicate various combinations of process change,
waste water reduction and recycling, waste water segregation, and addi-
tional treatment of some specific portions of waste water. The smelter-
refinery type operations show concern for improvements in both sanitary
and industrial waste water discharges.
The general trend in these plans is for increased recycle and decreased
discharge. Although most of the plants reporting are in areas of plen-
tiful rainfall, it is not unlikely that some, by the implementation of
these plans, may achieve extremely low or possibly zero discharge of
effluents, a situation that is now only attained by plants in desert
climates, through evaporation and permeation.
If the trend, as expressed by these responses, is a general one in the
industry, as it well may be, decided reduction in water usage and pol-
lution can be expected.
As part of this program, the information obtained has been reviewed
and assessed in terms of unit processes and associated waste waters.
The numerous unit processes covered are presented in Table 39.
The approach has been to review each metal in terms of all unit pro-
cesses, indicate whether water is used and follow with waste water char-
acteristics and control status, wherever such information was obtained
in the program. Also included in Table 39 are recommendations for the
next action appropriate for each process step indicated.
The data indicate that there are many specialized problems but a brief
discussion may aid in the understanding of the table entries. The
recommendations are in terms of double objectives: (1) identifying
needs for research, development, or demonstration to lend impetus to
water quality improvement, and (2) assessing the approach of the study
156
-------
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to the point where unit process waste loads and effluent standards
could be set. There follow some generalizations which may be drawn
from the table.
The cowion wastessanitary and indirect cooling watershave reached
the situation where past published literature and current data could
be used to set waste loads, treatment costs, and effluent standards.
Indirect, cooling water and boiler blow-down currently are being treated
in some plants for recycle to the same circuits. These wastes are not
unique to the metals industry. However, R&D is recommended for the
development of a process for treating sanitary wastes for reuse in the
context of an industrial plant which ma)7 have resources (waste heat or
chemicals) or water quality demands different from those of municipal
treatment plants. Such a process would decrease industrial water de-
mands and discharges.
One of the general characteristics of smelting operations are acid
plant waters, i.e., liquors from preliminary gas cleaning steps prior
to manufacture of sulfuric acid. The nature of the waste water (re-
flecting individual ore compositions) would indicate complex, high, and
variable waste loads, and the need for treatment processes capable of
removing the complex wastes. Also identified in Table 39 are similar
complex waste loads from by-product (Cd, In, As, Tl) metals recovery
operations using chemical or hydrometallurgical processes. These waste
loads are characteristic of by-product operations in the lead and zinc
industry and of custom smelters handling fume and flue dust from other
industries. The'acid plant liquors (which will probably become more
common as SOX control increases) and the by-product metal operation
wastes appear to be compatible for simultaneous R&D studies of water
treatment processes aimed at achieving satisfactory removal of all the
by-product metal impurities from the complex waste.
Spent electrolytes appear as wastes in the copper, lead, zinc, gold,
and silver industries. More analyses are required to establish waste
loads or effluent standards. The possibility of recycle should be de-
terriined more clearly by a survey of industry, especially in the case
of gold and silver parting plants, where no information was obtained.
Although equipment wash water and electrolyte spills were identified in
this stud\ as contributing to waste water, it is "judged that the most
effective means of control of these would be plant design and waste
segregat ion.
Tv;o other major types uf waste water identified were mine drainage and
ilotation concentration effluent. Both these streams are high in vol-
ume and low in impurity content. The mixed response to this survey
varied from immediate discharge--to treat and discharge (with varying
treatment)--to total recycle, and would indicate that practice is not
uniform throughout the industry. Thus, case studies are recommended of
selected best and worst situations, with the evaluation of the impor-
tance of factors contributing to the most efficient performance of the
163
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better settling pond type operations (e.g., retention time, flocculants,
oxidation promoting factors, or other factors).
At the time of this surveywhich gathered data produced mostly in 1970--
data allowing the calculation of unit waste loads existed for only a few
of the unit processes identified, but industry was progressing in the
accumulation of analytical data. A subsequent study should yield an in-
creased amount of specific data.
164
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SECTION XVI
ACKNOWLEDGMENT
The sections of this report dealing with waste-water sources, charac-
teristics, amounts, and treatment practices are based Largely on the
contributions of data and information of producers of nonferrous metals.
These contributions were made on a purely voluntary basis and at no
cost to the project. The constraint of confidentiality does not allow
the recognition of the contributors to this study on an individual or
corporate basis, thus due acknowledgment must be given to the nonferrous
metals industry in general.
Acknowledgment can be given to the American Mining Congress Ad Hoc
Committee on Wastewater Treatment Practice Survey. This committee
served to establish policies, supply information, and contribute expert
technical review and comment on various sections of this report.
The program covered by this report was conducted by Battelle-Columbus
during the period of June, 1970, through June, 1971. Battelle staff
and consultants participating in this program were A. B. Tripler, Jr.,
J. B. Hallowell, J. F. Shea, R. H. Cherry, Jr., G. R. Smithson, Jr.,
and B. W. Gonser.
The support of this program by the Water Quality Office, Environmental
Protection Agency, and the help provided by Messrs. R. L. Feder, E. L.
Dulaney, and John Ciancia of that office is acknowledged with sincere
thanks.
165
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SECTION XVII
REFERENCES
1. Staff, Bureau of Mines, Minerals Yearbook. 1969, "Volume I-II,
Metals, Minerals, and Fuels", United States Department of the
Interior, Bureau of Mines, U.S. Government Printing Office, Wash-
ington, D.C. (1971).
2. Staff, Bureau of Mines, Minerals Yearbook, 1968. "Volume I-II,
Metals, Minerals, and Fuels", United States Department of the
Interior, Bureau of Mines, U.S. Government Printing Office, Wash-
ington, D. C. (1969).
3. Annual Data 1971; Copper Supply and Consumption. 1951-1970, Copper
Development Association, Inc.
4. McMahon, A. D., "Copper, A Materials Survey", Bureau of Mines
Information Circular 8225, U.S. Department of the Interior, Bureau
of Mines, U.S. Government Printing Office, Washington, D.C.
5. 1970 E/MJ International Directory of Mining and Mineral Processing
Operations, Published by Mining Informational Services, Engineering
and Mining Journal, McGraw-Hill, New York (1970).
6. U.S. Bureau of the Census, Census of Manufactures, 1967, "Industry
Series: Smelting and Refining of Nonferrous Metals and Alloys",
MC67(2)-33C, U.S. Government Printing Office, Washington, D.C.
(1970).
7. U.S. Industrial Outlook 1970, U.S. Department of Commerce, U.S.
Government Printing Office, Washington, D.C. (1969).
8. Bureau of Mines, "Mineral Facts and Problems, 1970 Edition", Bureau
of Mines Bulletin 650, U.S. Department of the Interior, Bureau of
Mines, U.S. Government Printing Office, Washington, D. C. (1970).
9. Hayward, C. R., An Outline of Metallurgical Practice, 3rd Edition,
Van Nostrand, New York (1952).
10. Bray, J. F., Nonferrous Production Metallurgy, Second Edition, Wiley,
New York (1947), p 164.
11. Lanier, H., "Copper", Kirk-Othmer Encyclopedia of Chemical Techno-
logy, Wiley and Sons, New York (1965), Vol 6, pp 131-181.
167
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12. Cotter-ill, C. H. , and Cigan, J. M. (editors), AIME World Symposium
on Mining and Metallurgy of Lead and Zinc,"Volume II, Extractive
Metallurgy of Lead and Zinc", The American Institute of Mining,
Metallurgical, and Petroleum Engineers Inc., Port City Press,
Baltimore, Maryland (1970).
13. The Refining of Non-Ferrous Metals, The Institution of Mining and
Metallurgy, London (1950).
14. Callaway, H. M., "Lead, A Materials Survey", Bureau of Mines In-
formation Circular 8083, U.S. Department of the Interior, Bureau
of Mines, Washington, D. C. (1962).
15. McKay, J. E., "Lead", Kirk-Othmer, op. cit., Vol 12, pp 207-247.
16. Bureau of Mines, "Mineral Facts and Problems, 1965 Edition",
Bureau of Mines, U.S. Department of the Interior, Bulletin 630,
Washington, D. C. (1965).
17. Materials SurveyZinc, compiled for the Materials Office, National
Security Resources Board by the U.S. Department of the Interior,
Bureau of Mines, with the cooperation of the Geological Survey
(March, 1951).
18. Howe, H. E., "Cadmium and Cadmium Alloys", Kirk-Othmer, op. cit.,
Vol 3, pp 884-899.
19. Schlecten, A. W., and Thompson, A. P., "Zinc", Kirk-Othmer, op.
cit., Vol 22, pp 555-603.
20. Butts, A., and Coxe, C. , SilverEconomics, Metallurgy, and Use,
D. Van Nostrand Co., Inc., Princeton, N.J. (1967).
21. Carapella, S. C., Jr., "Arsenic", Kirk-Othmer, op. cit., Vol 2,
pp 711-717
22. Elkin, E. M., and Margrave, J. L., "Selenium and Selenium Com-
pounds", Kirk-Othmer, op. cit., Vol 17, pp 809-833.
23. Lausche, A. M., "Selenium and Tellurium, A Materials Survey",
Bureau of Mines Information Circular 8340, U.S. Department of the
Interior, Bureau of Mines, U.S. Government Printing Office,
Washington, D.C. (1967).
168
ftU.S. GOVERNMENT PRINTING OFFICE:1973 5^6-312'168 1-3
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1
V
,/icccAmcn Number
y
2
Subject Flold & Group
05G
SELECTED WATER RESOURCES ABSTRACTS
INPUT TRANSACTION FORM
Organization
Battelle Columbus Laboratories, Columbus, Ohio
Titlo
Water-Pollution Control in the Primary Nonferrous-Metals Industry Volume I,
Copper, Zinc, and Lead Industries
10
Aut2ior(e)
W 1 1 1 /-.I.TO IT T R
16
Project Designation
EPA, OR&M Contract No.
14-12-870
Shea, J. F.
Stnithson, G. R. , Jr.
Tripler, A, B.
Gonser, B. W.
Noto
22
Citation
Environmental Protection Agency report number,
EPA-R2-73-247a, September 1973.
23
Descriptors (Starred First)
Heavy Metals Water Pollution Control, Mining, Smelting, Refining
Waste Water Treatment Costs, Cu, Pb, Zn, Al, Hg, Au, Ag, Mo, W, Cd, Bi, Sb
25
Identifiers (Starred first)
Mining, Nonferrous Metals, Smelting
Heavy Metals
27
Abstract
The purpose of the program was to identify specific research needs in the area of
water pollution in the primary nonferrous metals industries. This program consisted
of a survey of literature and the acquisition of data from industrial operations.
The contents of the final reports (2 volumes) include: the identification of process
steps using water and/or generating wastewater, the amounts of water used for various
purposes, recirculation rates, amounts of wastewaters, specific or characteristic
substances in wastewaters, the prevalence of wastewater treatment practice, methods,
and costs; current treatment problems, and plans for future practices of recirculation
or wastewater treatment.
The metals reported on included copper, lead, zinc, and associated byproducts (arsenic,
cadmium, silver, gold, selenium, tellurium, sulfuric acid, salts and compounds),
mercury, (primary)gold and silver, aluminum, molybdenum, and tungsten.
The information presented includes detailed processing descriptions and flowsheets,
tabulations of quantities of water intake, quantities used by category, recirculated
water, discharge water quantities and analyses, water treatment costs. Representative
water flowsheets are given.(Hallowell-Battelle-Columbus).
Abstractor
J. B. Hallowell
Battelle-Columbus Laboratories
SEND. WITH COPY Of DOCUMENT, TOi WATtR UK SOURCE."; SCIENTIFIC I N F C f< M A T 1 Cm CLNTTR
U.S. DEPARTMENT OF THE INTERIOR
WASHINGTON, D.C. 20240
CPOI 10VO - 407 -8» J
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