United States
Environmental Protection
Agency
Industrial Environmental Research EPA 600 2-80 159
Laboratory June 1980
Cincinnati OH 45268
Research and Development
&EPA
Overview of Foreign
Nonferrous Smelter
Technology
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RESEARCH REPORTING SERIES
Research reports of the Office of Research and Development, U.S. Environmental
Protection Agency, have been grouped into nine series. These nine broad cate-
gories were established to facilitate further development and application of en-
vironmental technology. Elimination of traditional grouping was consciously
planned to foster technology transfer and a maximum interface in related fields.
The nine series are:
1. Environmental Health Effects Research
2. Environmental Protection Technology
3. Ecological Research
4. Environmental Monitoring
5. Socioeconomic Environmental Studies
6. Scientific and Technical Assessment Reports (STAR)
7. Interagency Energy-Environment Research and Development
8. "Special" Reports
9. Miscellaneous Reports
This report has been assigned to the ENVIRONMENTAL PROTECTION TECH-
NOLOGY series. This series describes research performed to develop and dem-
onstrate instrumentation, equipment, and methodology to repair or prevent en-
vironmental 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.
This document is available to the public through the National Technical Informa-
tion Service, Springfield, Virginia 22161.
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EPA-600/2-80-159
June 1980
OVERVIEW OF FOREIGN NONFERROUS
SMELTER TECHNOLOGY
by
A. Christian Worrell III and Mary A. Taft
PEDCo Environmental, Inc.
Cincinnati, Ohio 45246
Contract No. 68-03-2577
Project Officer
John O. Burckle
Energy Pollution Control Division
Industrial Environmental Research Laboratory
Cincinnati, Ohio 45268
INDUSTRIAL ENVIRONMENTAL RESEARCH LABORATORY
OFFICE OF RESEARCH AND DEVELOPMENT
U.S. ENVIRONMENTAL PROTECTION AGENCY
CINCINNATI, OHIO 45268
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DISCLAIMER
This report has been reviewed by the Industrial Environmen-
tal Research Laboratory, U.S. Environmental Protection Agency,
and approved for publication. Approval does not signify that the
contents necessarily reflect the views and policies of the U.S.
Environmental Protection Agency, nor does mention of trade names
or commercial products constitute endorsement or recommendation
for use.
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FOREWORD
When energy and material resources are extracted, processed,
converted, and used, the related pollutional impacts on our en-
vironment and even on our health often require that new and in-
creasingly more efficient pollution control methods be used. The
Industrial Environmental Research Laboratory-Cincinnati (lERL-Ci)
assists in developing and demonstrating new and improved method-
ologies that will meet these needs both efficiently and economi-
cally.
This report presents a brief overview of numerous production
and pollution control processes that are in use or under develop-
ment abroad for the production of nonferrous metals. This report
is not intended to imply that any specific technique is applica-
ble to domestic practice, but rather to inform the reader of new
technological developments so that he may pursue those items of
interest. For additional information, the reader is referred to
the bibliographies following each process description.
David G. Stephan
Director
Industrial Environmental Research Laboratory
Cincinnati
111
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ABSTRACT
Numerous production and pollution control processes that
are not used in the United States are in use or under development
by foreign nonfferrous metal producers. Although some do not
apply to U.S. conditions, others can reduce pollution, increase
production, or lower costs. Many 'of. these foreign processes are
described~in this report.
The descriptions are divided into five categories: pyro-
metallurgical processes, hydrometallurgical processes, electrolytic
processes, air pollution control processes, and water pollution
control processes. If data were available, each process descrip-
tion includes a discussion of econcftnic, environmental, and
energy considerations, as well as a discussion of the basic
operating principles. A detailed analysis of each process is not
attempted in this report. For additional information, the reader
is referred to the list of references and bibliography following
each process description.
Data for this report were obtained from foreign and domestic
journal articles, patents, books, symposium proceedings, and
industry literature. Technology i*i the copper, lead, and zinc
industries is emphasized.
IV
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CONTENT?
Foreword
Abstract
Figures
Acknowledgment
1. Introduction 1
2. Pyrometallurgical Processes 2
Mitsubishi process 2
Britcosmaco process 6
Torco process 8
WORCRA process 10
Continuous fire refining furnace 13
Vacuum refining 13
Extended Arc Flash -Reactor 16
Inco oxygen flash smelting 18
Queneau-Schuhmann process 21
Kivcet process " 23
Cyclone furnace sme.lting 27
Top blown rotary converter 29
Imperial Smelting process 33
Boliden direct reduction process 36
Bergs^e whole battery smelting process 38
Oliforno whole battery smelting process 41
Pyrometallurgical slag treatment 41
Sea nodule processing 43
Automatic tuyere punching 46
3. Hydrometallurgical Processes 49
Minemet process 49
Sherritt-Cominco copper process 51
Acetonitrile extraction process 55
Lurgi-Mitterburg process 56
Pressure-leach process for zinc recovery 57
Electrolytic Zinc prdcess 59
Other techniques for ..treating oxidized
zinc ores 62
Hydrochloric acid leaching of fabric filter dust 63
v
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CONTENTS, (continued)
4. Electrolytic Processes 65
Fluidized-bed electrolysis 65
Dextec process 67
Cathode sheet stripping 68
5. Air Pollution Control Processes 71
Dowa basic aluminum sulfate process 71
Mitsubishi lime/limestone process 73
Magnesium oxide process 77
Boliden cold seawater process 78
Flakt-Boliden citrate process 81
Showa Denko process 84
Zinc oxide process 85
Flash agglomeration 87
Boliden dry selenium filter 90
Boliden wet selenium scrubbing process 91
Boliden activated carbon process 92
Mercuric chloride scrubbing process 93
Outokumpu sulfuric acid scrubbing process 95
\ Centre Nacional de Investigaciones process 98
DeMarc D process 99
Boliden process 100
Mitsui process 101
Gortdrum Mines process 101
Swingaway converter hoods 103
Accordion door secondary hooding 104
Enclosed two-position secondary hooding 106
Air curtain hooding 108
6. Water Pollution Control Processes
Boliden sulfide precipitation process 111
Ferrite precipitation process " 114
VI
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FIGURES
Number Page
1 Schematic view of the Mitsubishi semicommercial
process 3
2 Britcosmaco process 7
3 Torco process 9
4 WORCRA furnace 11
5 Continuous fire refining furnace 14
6 Possible applications of vacuum refining to
the copper smelting process 15
7 Schematic view of a laboratory unit showing
components of the Extended Arc Flash Reactor 17
8 Inco process 19
9 Schematic view of the Q-S process 22
10 Simplified flowsheet of the Kivcet smelting
process 25
11 Top blown rotary converter 30
12 Imperial Smelting furnace 34
13 Slag fuming process 42
14 Automatic tuyere punching system 47
15 Minemet process for copper 50
16 Simplified flowsheet of the S-C process 53
t
17 Zinc recovery by pressure-leach autoclave,
combined with a conventional plant at Trail,
B.C. 58
VII
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FIGURES (continued)
Number Page
18 Simplified flowsheet of the EZ process as
piloted by the New Jersey Zinc Co. 60
19 Schematic view of an FEE cell 66
20 Cathode sheet stripping 69
21 Dowa basic aluminum sulfate process 72
22 Mitsubishi lime/limestone process used at
Onahama 7 5
23 Boliden cold seawater process 79
24 Flakt-Boliden citrate process 82
25 Zinc oxide process 86
26 Flash agglomeration furnace 88
27 Mercuric chloride scrubbing process 94
28 Outokumpu sulfuric acid scrubbing process 96
29 Flowsheet of solids for Gortdrum Mines
process 102
30 Flowsheet of gases for Gortdrum Mines
process 102
31 Accordion door secondary hooding at Onahama 105
32 Enclosed two-position secondary hooding 107
33 Air curtain hooding 110
34 Simplified flowsheet of the Boliden sulfide
precipitation process 112
35 Ferrite precipitation process 115
Vlll
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ACKNOWLEDGMENT
This report was prepared by PEDCo Environmental, Inc., under
the direction of Mr. Timothy W. Devitt. The PEDCo project manager
was Mr. Thomas K. Corwin. Principal authors were Mr. A. Christian
Worrell, III and Ms. Mary A. Taft.
Project officer for the Industrial Environmental Research
Laboratory of the U.S. Environmental Protection Agency was Mr.
John 0. Burckle.
The authors appreciate the efforts and cooperation of every-
one who participated in the preparation of this report.
IX
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SECTION 1
INTRODUCTION
This report is an overview of foreign nonferrous metallur-
gical and pollution control processes not used in the United
States on a commercial or pilot scale. Information on the
status and operating principles of these processes was obtained
from available literature.
Although emerging technology is stressed, not all processes
described represent the latest developments in nonferrous metal-
lurgy; some are older techniques that may not be readily appli-
cable to conditions in the United States (e.g., Imperial Smelting
is not applicable to most U.S. ores). Emphasis, however, is
placed on metallurgical and pollution control techniques that
have been demonstrated on at least a pilot scale and may be of
use to the U.S. copper, lead, and zinc industries. The energy
consumption and environmental effects of a process are addressed
when possible.
For the convenience of readers interested in obtaining
additional information about the processes, references are given
by subsection. When appropriate, a bibliography is also included
at the end of each description.
A special system of referencing is used in this report. One
or more reference numbers directly after a subsection title
indicate that all the material in the subsection is drawn from
the reference(s) cited. A reference number after a sentence
within a paragraph signifies that only the information in that
sentence comes from the reference cited. One or more references
after the last sentence of a paragraph show that all the informa-
tion in the paragraph is based on the reference(s) cited.
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SECTION 2
PYROMETALLURGICAL PROCESSES
2.1 MITSUBISHI PROCESS
In 1961, Mitsubishi Metals Corporation set out to develop a
process of continuous, pollution-free copper smelting. After
pilot-scale operations at the Onahama smelter proved the techni-
cal feasibility of the process, Mitsubishi built a commercial-
scale plant with a capacity of 48,000 Mg/yr of blister copper at
the Naoshima smelter. A second commercial-scale plant is under
construction by Texas Gulf at Timmins in Ontario, Canada. The
capacity of the Canadian plant is projected to be 59,000 Mg/yr,
1-4
and later to be expanded to 118,000 Mg/yr.
As shown in Figure 1, the Mitsubishi process is composed of
three metallurgical stages, each of which is carried out in a
separate furnace. Concentrates are smelted in the first, slag is
cleaned in the second, and matte is converted to blister copper
in the third. Intermediate products in the molten state move
continuously between the furnaces.
Raw materials are charged through the top of the smelting
furnace by vertically installed lances.' Use of lances for
charging and blowing increases rates of smelting and oxidation
and simplifies design and maintenance. Top blowing offers the
advantage of introducing oxygen or fuel oil"through the lances
when necessary.
Matte and slag produced in the smelting unit continuously
flow to the slag cleaning furnace as an emulsion. Cleaned
slag is skimmed continuously and granulated, and the matte is
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RETURNING BY MOVING BUCKET SYSTEM OR AIR LIFTING
COOLANT AIR
u>
SMELTING FURNACE
SLAG CLEANING
FURNACE
CONVERTING FURNACE
SLAG
GRANULATION
i
Figure 1. Schematic view of the Mitsubishi semi commercial process.'
-------
sent to the converting furnace. Average copper loss in this
waste slag is about 0.4 percent.
With lances that introduce blow air, the converting furnace
oxidizes the matte to blister copper, which is continuously
tapped from the furnace. Slag formed during the converting
stage is recycled to the smelting furnace by a moving bucket
system.
The average sulfur dioxide (S02) content of the combined
off-gases from the three furnaces is greater than 10 percent.
When the smelting furnace is operated with 25 percent oxygen-
enriched air, SO2 concentration is fairly constant and permits
economic recovery of sulfur as sulfuric acid or elemental sul-
fur. Dusts generated by the system can be collected and
treated for minor element recovery by conventional techniques
before recycle to the smelting unit.
The Mitsubishi process should not significantly harm water
quality. A settling pond is not needed because slag is treated
in an electric furnace. As with conventional smelting, granu-
lated slag is disposed of easily; depending on local markets, it
may be used for cement manufacture or road beds.
A plant using the Mitsubishi process is estimated to require
only 70 to 80 percent of the capital investment needed to build
a conventional smelter. This low capital investment is at-
tributed to relatively simple engineering design involving the
continuous gravity flow of molten intermediate and final prod-
ucts and to higher output per volume of smelting unit because of
an increased smelting rate compared to reverberatory smelters.
The Mitsubishi process requires only 2.7 x 10 kcal/Mg to
produce a product, whereas conventional smelting requires 5.2 x
6 8
10 kcal/Mg. This low fuel requirement is attributable to the
compactness of the furnace and the maximum use of the reaction
heat of the iron and sulfur content in concentrates by the
production of high-grade matte in the smelting furnace.
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References
1. Amsden, M.P., et al. Selection and Design of Texas Gulf
Canada's Copper Smelter and Refinery. Journal of Metals,
30(7):16-26, July 1978.
2. Suzuki, T. and T. Nagano. Development of a New Continuous
Copper Smelting Process. Presented at the Joint Meeting of
the Mining and Metal Institute of Japan and the American
Institute of Mining, Metallurgical, and Petroleum Engineers,
Tokyo, Japan, May 24-27, 1972.
3. New Copper Processes Move to Commercialization. Engineering
and Mining Journal, pp. 135-136, June 1975.
4. More Details Told on Copper Process. American Metal Market,
May 26, 1972.
5. Price, F.C. Copper Technology on the Move. Engineering and
Mining Journal, pp. RR-WW, April 1973.
6. Mitsubishi's Continuous Copper Smelting Process Goes on the
Stream. Engineering and Mining Journal, 173(8): 66-68,
August 1972.
7. Nagano, T., et al. Commercial Operation of Mitsubishi
Continuous Copper Smelting and Converting Process. Ex-
tractive Metallurgy of Copper. Vol. 1. Metallurgical
Society of AIME, 1976, New York. pp 439-457.
8. Arthur D. Little, Inc. Assessment of the Adequacy of
Pollution Control Technology for Energy Conserving Manufac-
turing Process Options. Cambridge, Massachusetts, December
1975.
Bibliography
Coleman, R.T., Jr. Emerging Technology in the Primary Copper
Industry. Draft prepared by Radian Corp. for the U.S.
Environmental Protection Agency under Contract No. 68-02-
2608. Austin, Texas, August 31, 1978.
Cullom, J.T. Selection of a Suitable Copper Smelting System.
Mining Congress Journal, 64 (4):45-48, April 1978.
Herbert, J.C., et al. Review of "New Copper Extraction Processes."
Journal of Metals, 26(8):16-24, August 1974.
Price, F.C. Copper Technology on the Move. Chemical Engineering,
80(9):RR-BBB,DDD,FFF-HHH, April 16, 1973.
-------
Rutledge, P. Mitsubishi Metal Previews Its Promising New
Continuous Copper Smelting Process. Engineering and Mining
Journal, 176(12):88-89, December 1979.
Texas Gulf Copper Plant at Timmins Will Use Mitsubishi Process.
Engineering and Mining Journal, 176 (5):50,54, May 1975.
Themelis, N.J. The Impact of Energy and Environmental Con-
straints on Copper Smelting Technology. Mining Engineering,
28(l):42-46, January 1976.
Verney, L.R. Pyrometallurgy. Journal of Metals, 39(3):16-18,
March 1977.
2.2 BRITCOSMACO PROCESS1
The Britcosmaco process, developed in Australia by consul-
ting engineer G.F. Brittingham, is intended to combine the best
features currently available in pyrometallurgical treatment of
sulfide ore concentrates. As shown in Figure 2, dry concentrate
and flux are fed into the main smelting shaft with sufficient
preheated or oxygen-enriched air for autogenous smelting. An
enriched white metal and a slag are produced and collect on the
hearth in two layers.
The slag increases in volume and flows along the phase
reaction section of the hearth; contact with low-grade matte
causes more copper to be rejected. As the slag flows toward the
tap hole, it is reduced to a greater extent immediately under the
secondary smelting shaft. Between the shaft and the tap hole,
fine particles of matte disperse through the slag and reduce the
oxygen potential even further, causing additional rejection of
copper.
As it falls down the main shaft, copper is oxidized and works
through the slag layer. It dissolves at the top of the enriched
metal layer and precipitates from the bottom as metal. The
copper is removed by a bottom tapping siphon.
According to Brittingham, the Britcosmaco process is advan-
tageous because copper recovery is higher than for conventional
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GASES AND SUSPENDED
MOLTEN PRODUCTS
FROM SECONDARY
SMELTING SHAFT
GASES AND SUSPENDED
MOLTEN PRODUCTS
FROM MAIN
SMELTING SHAFT
SECONDARY
SMELTING
SHAFT
MAIN
SMELTING
SHAFT
GAS UPTAKE TO
WASTE HEAT
BOILER
PHASE REACTION
HEARTH
GASES AND SOME
SUSPENDED MOLTEN
PRODUCTS FROM
SECONDARY
SHAFT
ENRICHED
WHITE METAL
COPPER TO
FURTHER
TREATMENT
AS REQUIRED
Figure 2. Britcosmaco process.
1
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pyrometallurgical treatment, energy requirements are lower, total
gas volume is smaller, and SO_ concentration in the off-gas is
greater.
Reference
1. Price, E.G. Copper Technology on the Move. Engineering and
Mining Journal, pp. RR-WW, April 1973.
2.3 TORCO PROCESS1'2
The Torco process has been used in Mauritania and Zambia to
recover copper from silicate ores such as dioptase and chryso-
colla, which cannot be extracted from the gangue rock by basic
flotation methods. The process is based on the discovery in 1923
that copper silicates form growths of elemental copper when they
are crushed, mixed with carbon and sodium chloride (common
salt), and smelted in an inert atmosphere. Figure 3 is a basic
flowsheet for a Torco plant. Coarse ore from the mine is crushed,-
dried, and ground. Ground ore is mixed with coal for the reac-
tion and heated in a fluidized bed reactor. The preheated mate-
rial from the first stage reactor overflows into a second fluid-
ized bed, where salt and additional coal are added. Proper
temperature and retention time result in the formation of copper
granules. The mixture is quenched in water and ground, and the
r
copper is separated by flotation. The low-grade copper product
is mixed with feed to a conventional smelter.
Exhaust gases from the fluidized bed reactor pass through
cyclones. After further particulate removal, the gases are dis-
charged to the atmosphere, and solids trapped by the gas cleaning
equipment are fed back to the reaction vessel.
References
1. Treilhard, D.G. Copper—State of the Art. Engineering and
Mining Journal, pp. P-Z, April 1973.
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COARSE ORE FROM MINE
ITORCO REACTOR
CRUSHER
r\
DRYER
\
B
CIRCUIT
v
Q
auF
DRY GRINDING MILL
TAILINGS
TO DAM
FILTER
\
CONCENTRATE
TO FILTER
FINISHED CONCENTRATE
TO SMELTER
FLOTATION PLANT
DRYER
Figure 3. Torco process
1
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2. Opie, W., et al. Selective Recovery of Copper From Copper-
Nickel Sulfide Concentrates by Applying Segregation Tech-
nology. Metallurgical Transactions B. Vol. 10B. American
Society for Metals and the Metallurgical Society of AIME,
New York, March 1979.
2.4 WORCRA PROCESS
The WORCRA process, developed by Conzinc Riotinto of Aus-
tralia, Ltd., is a continuous direct smelting process for copper
concentrates. Experimental evaluation of WORCRA began in early
1963 in a pilot plant located in Cockle Creek, New South Wales,
Australia. Encouraging results lead to the construction of a
larger test furnace in Port Kembla, New South Wales. Trial tests
indicate that the process can work on a larger scale. „
Copper concentrates are smelted to matte, matte is converted
to metal, and slag is cleaned in separate zones of an elongated
hearth-type furnace. Concentrates and flux are added in the
mildly oxidizing smelting zone at an angle which ensures particle
penetration into the melt and aids in the continuous circulation
of matte and slag. As illustrated in Figure 4, concentrates are
occasionally added in the converter zone to help control mag-
netite formation.
Slag generally moves countercurrently to the matte, and
iron and other unwanted materials are continuously transferred to
the slag after oxidation. Copper in the slag reverts to the
matte phase by interaction with iron sulfides in the matte.
Moving slowly through the smelting and converting zones,
matte is lanced with air (or oxygen-enriched air), causing con-
version to white metal and then to copper. If oxygen is not
used, lance air and combustion air preheaters are required, as
are a larger furnace, boiler, and electrostatic precipitator
(ESP). In the converting zone, the hearth slopes downward, and
copper passes continuously to the "copper well," which overflows
10
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BURNER
COPPER
t
AIR SILICEOUS
LANCES FLUX
CONCENTRATE
AND FLUX
CONCENTRATE REDUCING
OR PYRITES /FLAME
CONVERTING ZONE
(OXIDIZING) SMELTING
ZONE
MILDLY
(OXIDIZING)
SLAG
CLEANING
ZONE
(REDUCING)
SLAG
MATTE TAPPED
INFREQUENTLY
Figure 4. WORCRA furnace.
1
11
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with low-grade blister copper product. Extensive fire refining
or converter processing may be required before casting.
The WORCRA furnace yields a constant stream of gas with an
average sulfur dioxide (SO,-) content of 10 percent; use ,of oxygen-
enriched air increases the SO2 content. Furnace gases can be
used for waste heat recovery, and dust can be recovered before
the gases are sent to a sulfuric acid plant or otherwise treated.
Capital costs for a WORCRA plant are expected to be 20 to
30 percent less than for a reverberatory furnace/converter plant
with a similar capacity. Lower operating costs are also pre-
dicted, depending upon local variations in costs for fuel, power,
and labor. Fuel consumption by a WORCRA plant should be 50 to
60 percent of the operating costs of a conventional reverberatory
furnace/converter plant of equivalent production rating.
References
1. New Copper Processes Move to Commercialization. Engineering
and Mining Journal, pp. 135-136, June 1975.
2. Reynolds, J.O. WORCRA Copper and Nickel Smelting. Presented
at the Joint meeting of the Mining and Metal Institute of
Japan and the American Institute of Mining, Metallurgical,
and Petroleum Engineers, Tokyo, Japan, May 24-27, 1972.
\
3. Worner, H.K. WORCRA Metallurgy Looks Promising for Pollu-
tion Control in Copper Plants. Engineering and Mining
Journal, 172:64-68, August 1971.
4. Semrav, K.T. Control of Sulfur Oxide Emissions from Primary
Copper, Lead and Zinc Smelters: A Review. Presented at the
63rd Annual Meeting of the Air Pollution Control Association,
St. Louis, Missouri, June 1970.
Bibliography
Pulling, D.H., et al. Application of a Continous Technique to
Secondary Copper Smelting. Canadian Institute of Mining and
Metallurgy Bulletin, 70(788):122-134, December 1977.
Treilhard, D.G. Copper—State of the Art. Engineering and
Mining Journal, pp. P-Z, April 1973.
Verney, L.R. Pyrometallurgy. Journal of Metals, 29(3):16-18,
March 1977.
12
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2.5 CONTINUOUS FIRE REFINING FURNACE1
A new process of continuous metal refining that involves
supersonic injection of oxygen has reportedly been developed by
KHD Industrieanlagen A.G. of West Germany. After preliminary
experiments, KHD began construction of a pilot plant. At last
report, the plant had been completed, but had not achieved
totally satisfactory results.
Blister copper is introduced to the furnace and passes into
a settling hearth. The melt overflows from the settling hearth
to the oxidation hearth, where pure oxygen is injected through
lances at 1800 km/h countercurrently to the metal flow. Slag
from this stage is drawn off continuously. After oxidation, the
melt underflows to the reduction hearth, where propane is in-
jected for final refining. The metal, as illustrated in Figure
5, exits the triple hearth furnace as fire-refined anode copper.
Reference
1. Continuous Metal Refining With Supersonic Top Blowing.
World Mining, 31(9):17, August 1978.
2.6 VACUUM REFINING1
Vacuum refining has been used in the steelmaking process
since 1952, and Japanese researchers have recently attempted to
apply the technique to production of nonferrous metals, espe-
cially copper. Literature on the process is sparse, but a brief
discussion follows.
Figure 6 shows four possible applications of vacuum refining
to the copper smelting process. The method can be applied to the
refining of blister copper and matte. Ladle refining can be used
to replace a conventional fire refining furnace, although the
sulfur and oxygen content of the blister copper must be carefully
controlled. The continuous treatment of copper is also a pos-
sible application of vacuum refining. Subsequent electrolytic
refining is required in any application.
13
-------
EXHAUST
GASES
BLISTER
COPPER
Figure 5. Continuous fire refining furnace
14
-------
VACUUM
LADLE REFINING
SHAFT
VACUUM
I I UP-TAKE
r
SETTLER REFINING
VACUUM
—Ir
WELL REFINING
VACUUM
CONTINUOUS REFINING
\|y
Figure 6. Possible applications of vacuum refining
to the copper smelting processJ
15
-------
Tests have shown that treatment of blister copper by vacuum
refining rapidly removes sulfur and oxygen as sulfur dioxide.
Adequate recirculation of the molten copper can remove 60 to 80
percent of the lead, 30 to 50 percent of the zinc, and 5 to 20
percent of the arsenic and antimony.
Reference
1. Kametani,. H., and C. Yamauchi. Vacuum Lift Refining in
Copper Smelting. Transaction of'the National Research
Institute for Metals, 20(l):22-59, January 1978.
2.7 EXTENDED ARC FLASH REACTOR
Canadian scientists have developed a new electric furnace
for the direct carbothermic reduction of powdered ore and dusts.
Although particularly applicable to the production of iron,
the furnace can be applied to the smelting of nonferrous metals.
As an outgrowth of work to improve the performance of electric
arc furnaces, the Extended Arc Flash Reactor is offered by Tibur
Metals, Ltd., of Hamilton, Ontario, Canada.
As shown in Figure 7, the furnace consists of four sections:
(1) a rotary preheater, (2) a column or flash reactor, (3) an
extended arc zone, and (4) the hearth. The rotary preheater
brings incoming charge (concentrate and coal) into contact with
hot exiting flue gas containing carbon monoxide. The charge is
dried and heated to approximately 800°C. The feed then enters a
vertical shaft leading to the hearth, where 50 to 60 percent of
the reduction occurs and the temperature increases from 800° to
1500°C.1
Hollow electrodes are used to inject argon gas into the
electric arc. The gas is ionized and forms an extremely hot
plasma', which can reach temperatures of 5000°C. This plasma
extends the arc by which heat can be transferred to the charge as
it passes through this zone. High temperatures within the plasma
ensure reducing conditions in the flash zone, the plasma itself,
and the hearth where final reduction occurs.
16
-------
(.1) ROTARY PREHEATER
(3) EXTENDED
ARC ZONE
2) COLUMN OR
FLASH REACTOR
(4) HEARTH
MOLTEN CHARGE
SPOUT
Figure 7. Schematic view of a laboratory unit showing
components of the Extended Arc Flash Reactor."
17
-------
The advantages of a stabilized extended arc include reduced
electrode consumption, reduced refractory wear, improved power
factor and control, improved heat transfer, and reduced acoustical
2
and electric noise.
References
1. New Unit Smelts Oxide Dusts & Ores. Canadian Chemical
Processing, 61(4):32-34, April 1977.
'.
2. U.S. Patent Office. Patent No. 4,037,043, issued to R.S.
Segsworth. Extended Arc Furnace and Process for Melting
Particulate Charge Therein. July 19, 1977.
2.8 INCO OXYGEN FLASH SMELTING
The International Nickel Company (Inco) of Canada developed
the oxygen flash smelting process in the 1940's to treat copper
and nickel concentrates. By the early 1950's, Inco had success-
fully demonstrated the method on a commercial scale at the Copper
Cliff plant. During this same period, Outokumpu designed and
demonstrated a similar flash smelting furnace, which has been
successfully marketed in many countries, including the United
States.1'2
The Inco oxygen flash smelting furnace is much simpler in
design than the Outokumpu furnace and uses pure oxygen for the
oxidizing medium, whereas the Outokumpu furnace uses preheated or
oxygen-enriched air. The Inco furnace is essentially a reverbera-
tory furnace with an uptake shaft extending the length of the
furnace roof.
Figure 8 illustrates the Inco process. A wet concentrate-
flux mixture is introduced through a rotary sealing feeder to the
fluidized bed dryer. Solid particles are dried while suspended in
the upward stream of hot combustion gas. Moist gases and dry
solids are separated in the two product fabric filters after they
are drawn out through the dryer roof. The product fabric filters
are reported to have a collection efficiency greater than 99.9
2
percent.
18
-------
SAND
CONCENTRATE
AIR
GAS AND AIR
n
FABRIC
FILTER
FLUID BED
DRYER
•EXHAUST
Ifi
FABRIC
FILTER
FROM SECOND
FLUIDIZED
BED DRYER
f .t. t
DRY FEED
STORAGE
WATER
I JACKET
•EXHAUST
t. t. t
OXYGEN *-»
DRY FEED
STORAGE
FLASH
FURNACE
OXYGEN
SLAGrj—^DUMP
POT 1^1
COPPER ^ r—, MA | ft
CONVERTERS \J LADLE
Figure 8. Inco process.
1
-------
The solids are withdrawn from feed bins and transferred by
screw conveyors and gravity to the two burners located at each
end of the furnace. Suspended in a horizontal flow of oxygen,
the feed is injected into the furnace. Oxygen combines with some
of the sulfur and iron to form SO and iron oxides. A fayalite
slag is formed by silica, which is contained in the flux and
concentrate, reacting with the iron oxides. The remaining sulfur
and metal fractions collect in the matte, which is transferred to
one of three Fierce-Smith converters for production of blister
2
copper. Almost all the oxygen is consumed during the process.
Exiting furnace off-gases contain 83 percent S0_, 13 percent
nitrogen, 1.5 percent carbon dioxide, 2.5 percent argon, and
<0.02 percent oxygen. About 2 to 3 percent of the feed exits as
dust during smelting operations. The furnace gas cleaning system
includes a settling and radiation chamber, a spray tower, three
venturi scrubbers, and a wet electrostatic precipitator. After
the gas stream is cooled and cleaned of particulates, the SO -
laden gas is piped to an adjacent liquefaction plant. The Inco
oxygen flash furnace emits less SO and fewer impurities than a
reverberatory furnace handling a typical concentrate for North
America. In addition, the extremely small volume and high
strength of the flash furnace off-gases blended with converter
2
gases allow for a greater degree of control. Other advantages
of this process include simple engineering design, easy process
control, and relatively low energy and capital requirements. The
process, however, is not as amenable to concentrates with high
impurity levels, because such levels can cause unacceptably high
concentrations of impurities in the blister copper. '
References
1. Solar, M.Y., et al. Smelting Nickel Concentrates in Inco's
Oxygen Flash Furnace. Journal of Metals, January 1979.
20
-------
2. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper industry. Prepared by Radian Corporation for the
U.S. Environmental Protection Agency under Contract No.
68-02-2608. Austin, Texas, August 1978.
2.9 QUENEAU-SCHUHMANN PROCESS^
Professors P.E. Queneau and R. Schuhmann, Jr., have invented
an environmentally clean process for the continuous production
of nonferrous metals from sulfide concentrates. The Queneau-
Schuhmann (Q-S) process is used in a pilot plant at Duisburg,
West Germany. Galena concentrates are first oxidized and then
reduced in a molten state in a single, sealed reactor. Rapid
oxidation is accomplished by submerged injection of oxygen;
reduction and total desulfurization can be obtained by submerged
injection of a small amount of pulverized coal. As shown in
Figure 9, bullion leaves the reactor at one end, and slag leaves
at the other.
The capital and operating costs of a commercial Q-S instal-
lation are estimated to be considerably less than those of a
conventional smelter with sinter plant and shaft furnace.
Energy can be saved because the large amount of coke required
for a conventional facility is replaced by a small amount of
low-grade coal.
A Q-S unit also offers environmental advantages over a con-
ventional smelter. The only process gases are those from the
reactor, which contain about 20 percent SO,, by volume. This
£* \
concentration readily lends itself to sulfuric acid or liquid
S0_ production. Because a Q-S unit is compact and operates
continuously, the volume of fugitive gases is relatively small,
and they can be treated in two-stage precipitators, with re-
sulting minor emissions of particulates.
Reference
1. Environmentally Clean Lead Bullion Production. Journal of
Metals, 30(12):11-12, December 1978.
21
-------
LEAD
CONCENTRATES
FLUXES AND
PLANT RESIDUES
TO SULFURIC
ACID PLANT
FLUE DUST
DUST
PRECIPITATOR
PELLETIZING
WASTE HEAT
BOILER
MIXING
CHAMBER
POWDERED
COAL
LEAD
BULLION
Q-S REACTOR
OXYGEN AND
SHIELD GAS
REFINERY
REDUCTANT:
POWDERED COAL
SLAG
TO DUMP
Figure 9. Schematic view of the Q-S process.
22
-------
Bibliography
Davey, T.R. and G.M. Willis. Lead, Zinc, and Tin. Journal of
Metals, 29(3):24-39, March 1977.
Matyas, A.G., et al. Metallurgy of the Direct Smelting of Lead.
TMS No. A75-80. The Metallurgical Society of AIMS, New
York, 1975.
Queneau, P.E., and R. Schuhmann, Jr. The Q-S Oxygen Process.
Journal of Metals, 16:14-16, August 1974.
Ryan, J.P., et al. Lead, 1977. MCP 9. Bureau of Mines, U.S.
Department of the Interior, Washington, D.C., December
1977.
2.10 KIVCET PROCESS
The Kivcet process involves continuous smelting of complex
sulfide concentrates with simultaneous production of lead, zinc,
copper, nickel, and minor metals such as cadmium and tin. The
process has been under research and development since 1963.
After extensive tests, a 45-Mg/day pilot plant was constructed in
the metal mining region of the Uzbekistan Republic in the Soviet
Union. The first 272-Mg/day commercial-scale plant for smelting
complex sulfide concentrates was built at Tashkent, and a 600-
Mg/day plant is under construction.
Sulfide concentrate, containing at least 25 percent sulfur,
and oxygen are fed into a cyclone furnace (see Subsection 2.11),
where the concentrate is autogenously roasted and flash-smelted
in a suspended state. The products leaving the bottom of the
cyclone reactor pass into a distribution chamber, where off-gases
and melt are separated. The off-gases are cleaned of particulates
and sent to the acid plant; the melt is routed to the electric
- 1-4
furnace.
The electric furnace is similar in design to a conventional
furnace and is provided with water-cooled bottom and side walls.
Settling takes place in the electric furnace in a reducing atmos-
phere, and addition of a reducing material such as hydrogen is
23
-------
necessary. Matte containing copper, nickel, cobalt, and precious
metals is periodically tapped and further processed by conven-
tional methods. Zinc is volatilized in the electric furnace and
may be condensed to metal or oxidized to produce high-grade zinc
oxide. Figure 10 is a simplified flowsheet of the Kivcet
process.
Off-gases from the distribution chamber are very small in
volume and high in SO- content. Approximately 80 percent of the
sulfur in the concentrate passes into the off-gases, which con-
tain up to 90 percent S02. The high-strength off-gases may
easily be converted to elemental sulfur, sulfuric acid, or lique-
fied sulfur dioxide.
Slag from the electric furnace contains very little entrained
matte and can be discarded immediately without cleaning. There-
fore, expensive and space consuming cleaning equipment is not
1-3
necessary.
Overall capital costs of a smelter using the Kivcet process
are claimed to compare favorably with other high-technology
processes. Elimination of agglomerating, sintering, and slag
cleaning installations and simplification of gas cleaning and
SO,, capture systems help reduce costs. ~
Before improvements in the electric furnace, energy con-
sumption seemed relatively high because of excess heat loss from
the water-cooled bottom and side walls. After modifications, the
energy consumption of a Kivcet smelter with an annual capacity of
73,000 Mg of crude bullion is estimated to be 600 kWh/Mg of
crude bullion.
The recently developed CS variant of the Kivcet process uses
a flash smelting shaft furnace instead of a cyclone furnace.
This variant has been developed in the Soviet Union for indus-
trial application in single-step autogenous smelting of lead
sulfide concentrate with oxygen. '
Significantly less heat is liberated when smelting lead
sulfide than when smelting a copper sulfide. Therefore, the
24
-------
CONCENTRATE
FEED AND OXYGEN
CYCLONE
REACTOR
I
HIGH-S02 GAS
DISTRIBUTOR
MELT
VOLATILIZED METALS
SLAG
TO WASTE -*-
ELECTRIC
FURNACE
CONDENSER
COPPER MATTE
TO PROCESSING
CRUDE ZINC
Figure 10. Simplified flowsheet of the Kivcet smelting process.
1
25
-------
great heat dissipation characteristic of the cyclone makes
autogenous smelting of lead sulfide very difficult in a cyclone
furnace. The electrothermal section of the variant CS process
and the facilities for zinc recovery and gas treatment are iden-
tical to the standard Kivcet process; only the roasting/smelting
c
section differs.
References
1. Kivcet Process for Complex Ores. World Mining, pp. 26-27.
June 1974.
2. Soviet Continuous Smelting Process Licensed in West. En-
gineering and Mining Journal, 175(7):25, July 1974.
3. Kivcet pamphlet. Licensintorg, Moscow, Soviet Union. Dis-
tributed by Southwire Company, Carrolton, Georgia.
4. Humboldt Wedag's Cyclone - Furnace Smelting Recovers Non-
ferrous Metals. Engineering and Mining Journal,
178(10):45,49, October 1977.
5. Chaudhuri, K.B., et al. How Kivcet Lead Smelting Compares
With Other Direct Reduction Processes for Lead. Engineering
and Mining Journal, 197(4):88-91,118, April 1978.
6. ; Miiller, E. How Kivcet CS Shaft Furnace Simultaneously
Smelts Pb-Zn. World Mining, April 1977.
Bibliography
Davey, T.R., and G.M. Willis. Lead, Zinc, and Tin. Journal of
Metals, 29(3):24-39, March 1977.
MacKey, P.J., et al. Pyrometallurgy—Annual Review. Journal of
Metals, pp. 36-43, April 1978.
Matyas, A.G., et al. Metallurgy of the Direct Smelting of Lead
TMS No. A75-80. The Metallurgical Society of AIME, New
York, 1975.
Price, F.C. Copper Technology on the Move. Engineering and
Mining Journal, pp. RR-ZZ, April 1973.
Ryan, J.P. et al. Lead, 1977. MCP 9. Bureau of Mines, U.S.
Department of the Interior, Washington, D.C., December
1977.
Verney, L.R. Pyrometallurgy. Journal of Metals, 29(3):l6-18,
March 1977.
26
-------
2.11 CYCLONE FURNACE SMELTING1
Humboldt Wedag of Germany, in cooperation with the Metal
Research Institute of the Soviet Union, has developed the cyclone
furnace smelting reactor utilized in the Kivcet Process. This
reactor is suitable for recovering metal from nonferrous ores,
oxide or sulfide concentrates, and metallurgical slags and
residues. Reports indicate that metals can be volatilized in
the elemental form or as compounds in compact, continuously
operating, cyclone furnace smelters, which can fire a variety of
fuels. Metals that cannot be volatilized can be upgraded to
matte.
The key to this smelting technique is a cylindrical cyclone
reactor. Feed smaller than 1 mm in diameter is charged to the
reactor from the top, and a tangentially admitted fuel-air mix-
ture moves the charge downward at high speed. The rapidly
heated particles melt and deposit as droplets on the reactor
walls by centrifugal force, but volatile matter escapes prior to
this deposition. Gases and molten products are discharged from
a central outlet at the bottom of the reactor; the gases then
pass to a secondary combustion chamber, designed as a cooler,
where carbon monoxide, volatilized metals, and metal compounds
are subjected to secondary oxidation by admitting air. This
chamber may also be used as a waste heat boiler. Waste gases
are treated in an electrostatic precipitator and/or fabric
filter to recover metal oxides.
Molten products collect in a settling furnace, where matte
and slag are separated. The matte and slag can be tapped into
ladles or separated continuously in a centrifuge designed for
high-temperature operation.
The current maximum capacity of a cyclone furnace is approx-
imately 275 Mg/day, but the capacity of an entire facility can
be easily increased by use of several furnaces. These furnaces
can be operated in a parallel fashion and require only one waste
gas treatment system and one electric furnace.
27
-------
A cyclone furnace smelter in Bolivia treats antimony sulfide
to produce antimony trioxide. The sulfide is volatilized, and
the antimony trioxide is captured in a fabric filter following an
afterburning operation. The cyclone furnace is 0.85 m in diam-
eter and 1.25 m long and can handle 36 Mg/day of ore, concen-
trate, and fluxes. Recovery efficiencies as high as 96 percent
have been reported. A second plant is under construction in
Bolivia and will be used to upgrade low-tin concentrates. The
capacity of this furnace will be 100 Mg/day.
The processing of tin, lead-bearing shaft furnace slags,
residues from neutral zinc leaching, or lead-zinc intermediate
products from lead and zinc residues does not require an electric
furnace to be used in conjunction with a cyclone furnace. When
residues to be treated contain copper, silver, or gold, an elec-
tric furnace is required to separate slag from matte and concen-
trate these elements in the matte.
Bulk flotation concentrates containing copper,,lead, and
zinc c^n be upgraded by volatilizing the lead and zinc, which are
recovered in the oxide form, and separating the enriched copper
matte from slag in an electric furnace. Tests indicate that lead
and zinc can be volatilized up to 90 percent, yielding a 70
percent metal oxide.
The major advantages of cyclone furnace smelting are the low
capital and operating costs.
Reference
1. Humboldt Wedag's Cyclone - Furnace Smelting Recovers Non-
ferrous Metals. Engineering and Mining Journal, 178(10):45,
49, October 1977.
Bibliography
Budon, V.P. Cyclone-Electrothermal Methods of Processing Copper
and Polymetallic Materials: Effect of Some Factors on
Removal of Material by Gases in Cyclone Smelting. Trans-
lated from Russian by the U.S. Environmental Protection
Agency. EPA-TR73-436C, 1968.
28
-------
Kozhakhmetov, S.M., et al. Cyclone-Electrothermal Methods of
Processing Copper and Polymetallic Materials: Behavior of
Volatile Metals Under Conditions of Cyclone Smelting of
Polymetallic Sulfide Ore. Translated from Russian by the
U.S. Environmental Protection Agency. EPA-TR-436D, 1968.
Onaev, I.A., et al. Cyclone-Electrothermal Methods of Processing
Copper and Polymetallic Materials: The Basic Constructional
and Thermal Characteristics of the Cyclone Smelting Units.
Translated from Russian by the U.S. Environmental Protection
Agency. EPA-TR-73-436A, 1968.
Shaukenbayeua, Z.T., et al. Cyclone-Electrothermal Methods of
Processing Copper and Polymetallic Materials: Effect of
Partial Pressure of SO2 on the Transfixes of Copper Into
Slag in the Copper-Slag-SO2 system. Translated from Russian
by the U.S. Environmental Protection Agency. EPA-TR-73-436A,
1968.
Tonkonogiy, A.V. , _et al. Cyclone-Electrothermal Methods of
Processing Copper and Polymetallic Materials. Translated
from Russian by the U.S. Environmental Protection Agency.
EPA-TR-73-436B, 1968.
2.12 TOP BLOWN ROTARY CONVERTER
The International Nickel Co. of Canada has adapted steel
process technology (the Kaldo process) to nonferrous metallurgy
in developing the top blown rotary converter (TBRC), which can
smelt a variety of concentrates or byproducts. Boliden a.-b. of
Sweden has carried out additional development of the TBRC, and
several commercial-scale plants are in use, including two in
Sweden and one each in Ontario and British Columbia, Canada.
Five TBRC production processes have been investigated and
tested by metallurgists in laboratory, pilot-scale, and commer-
cial-scale plants. These processes have been tested with dust
containing lead-zinc, copper matte or concentrate, copper con-
verter slag, lead concentrate, and nickel concentrate.
Figure 11 shows a TBRC, which is claimed to increase opera-
tional flexibility through control of both temperature and
turbulence and maintain high thermal efficiency. After complex
29
-------
TILT
SWINGABLE
EXHAUST HOOD
WATER COOLED
OXYGEN LANCE
U)
o
EXPANSION
BEARING
PEDESTAL
FURNACE
ROTATION DRIVE
TAPPING SIDE
OPERATING
FLOOR LEVEL
FURNACE SUPPORT
FRAME
CHARGING SIDE
FURNACE TILT-
DRIVE TRAIN
Figure 11. Top blown rotary converter.'
-------
copper concentrates are dried and agglomerated in. a briquetter,
they are fed to the TBRC, which contains a molten sulfide bath.
The vessel atmosphere is controlled by injecting natural gas and
oxygen-enriched air into the molten mixture through a water-
cooled lance. White metal is then sent to a second TBRC for
further oxidation and elimination of impurities.1"3
The furnace rotates constantly, providing thorough contact
between the gas and furnace contents and ensuring even distribu-
tion of heat. Oxygen efficiency greater than 95 percent can be
obtained by adjusting rotational speed and angle of the injection
lance. The molten metal can be removed from the furnace and
separated by electrolytic refining. Slag is left in the vessel
for recovery of valuable metals upon addition of new concentrate.
In the Boliden lead smelter, the TBRC is used to treat
pelletized lead bearing dusts and residues and reduce high-
4 5
grade lead concentrates. ' At Ronnskar, Sweden, the TBRC is
housed in a separate building, which is completely enclosed by a
ventilation casing; the casing incorporates a bag filter. With
oxygen enrichment, the exhaust gases contain up to 50 percent S02
and little particulate matter. The Afton mine/smelter complex in
British Columbia sends waste gas from the TBRC through a combina-
tion of electrostatic precipitators for particulate removal and a
dual-alkali scrubber for SCU removal. A close-fitting exhaust
hood fits over the mouth of the converter during loading, smelt-
ing, and converting operations.
The advantages of the TBRC include lower particulate load-
ings, decreased fugitive emissions, and greater S02 concentra-
tions. The batch-type operation, however, does not produce a
continuous stream of gas. Work at the Inco Copper Cliff Nickel
refinery in Ontario indicates that particulate emissions are
directly related to the number of converter revolutions per minute,
Decreases from 25 to 5 rpm during charging and blowing and from
30 to 15 rpm during reduction reduce stack loss of metal values
7
by half and improve refractory performance.
31
-------
Data indicate that the capital cost of the TBRC process is
considerably less than that of conventional smelting techniques.
The operating cost is also less because of lower maintenance,
2 3
labor, and energy requirements. '
References
1. Boliden Adopts TBRC for Pb and Now Cu Smelter. World Mining,
31(5):44-48, May 1978.
2. Daniele, R.A., et al. TBRC - A New Smelting Technique.
TMS No. A72-101. The Metallurgical Society of AIME, New
York, 1972.
3. Afton, New Canadian Copper Mine on Stream. World Mining,
pp. 42-44, April 1978.
4. Peterson, S., A. Norro, and S. Ericksson. Treatment of
Lead-Zinc Containing Dust in a TBRC. TMS Paper No. All-12.
The Metallurgical Society of the American Institute of
Mining, Metallurgical, and Petroleum Engineers, New York,
1977.
5. Peterson, S., et al. Autogenous Smelting of Lead Concentrate
in TBRC. Paper No. A77-11. The Metallurgical Society of
AIME, New York, 1977.
6. Afton's Copper Smelter Proves Economic at 27,000 Tons Yearly.
Canadian Chemical Processing, 63(2): 22-24, March 21, 1979.
7. Thoburn, W.J., and P.M. Tyroler. Optimization of TBRC
Operation and Control at Inco's Copper Cliff Nickel Refinery.
Presented at the 18th Annual CIM Conference of Metallurgists,
Sudbury, Ontario, August 19-23, 1979.
Bibliography
Halliburton, D., and W.A. Lemmon. An Overview of Controls in
Primary Lead and Zinc. Air Pollution Control Directorate,
Environment Canada, Ottawa, Ontario. 1979.
Herbert, J.C., et al. Review of New Copper Extraction Processes.
Journal of Metals, 26(8):16-24, August 1974.
Mackey, P.J., et al. Pyrometallurgy - Annual Review. Journal of
Metals, pp. 36-43, April 1978.
32
-------
Price, F.C. Copper Technology on the Move. Chemical Engineering.
80(9):RR-BBB,DDD,FFF-HHH, April 16, 1973.
Treilhard, D.G. Copper—State of the Art. Engineering and
Mining Journal, pp. P-Z. April 1973.
U.S. Patent Office. Patent No. 4,032,327. Pyrometallurgical
Recovery of Copper from Slag Material. 1977.
2.13 IMPERIAL SMELTING PROCESS
The Imperial Smelting process has long been a competitive
means of producing lead and zinc and should continue to be eco-
nomically viable because it allows smelting of complex lead-zinc
ores. Although the process is used in 10 countries, it does
not appear applicable to many U.S. concentrates.
The process can vary, but the vertical shaft blast furnace
is the main item of any Imperial Smelting facility. The Societe
Miniere and Metallurgique Pennarroya facility in Noyelles-Godault,
France, first operated an Imperial furnace in 1962 as a replace-
ment for a horizontal retort. The furnace can produce 90,000
Mg of zinc each year. Figure 12 shows the furnace design and
process flow.
At the Noyelles-Godault plant, zinc, lead, and zinc-lead
concentrates are sintered in a Dwight-Llbyd machine to remove
sulfur. A sulfuric acid plant treats the off-gas, and vola-
tilized cadmium is treated separately. The sinter is fed with
fluxes to the top of the furnace, and metallurgical-grade coke is
charged on top. The blast enters the furnace through 15 steel
tuyeres, each 90 mm in diameter. The blast temperature of 900°C
2
is obtained with two cowper stoves that fire furnace gas. The
lead and zinc oxides are reduced by coke combustion gases; lead
metal is produced in the upper portion of the furnace shaft and
sinks toward the bottom. The zinc oxide is reduced in higher-
temperature zones lower in the furnace, and the volatilized
metal, which leaves the furnace with the off-gas; is captured in
33
-------
HOT COKE AND SINTER FEED
BY SKIP HOIST BUCKET
DOUBLE BELL
CHARGE HOPPERS
LEAD PUMP FROM CONDENSERS
TO WATER COOLED LAUNDER
WATER SPRAYS<
BLAST FURNACE GAS
TO MAIN
DISINTEGRATOR
MOISTURE
SEPARATOR
BLUE POWDER
CONDENSER
HOT BLAST MAINS
WATER SEAL
BLUE POWDER
DREDGE TANK
OVERFLOW LIQUOR
TO DORR THICKENER .
BLOWER
TO GAS WASHING PLANT
LEAD RETURN TROUGH
UNDERFLOW BAFFLE
ZINC HOLDING
BATH
^
ZINC
SLAG
AIR PREHEATER
Figure 12. Imperial Smelting furnace.'
-------
condensers. The slag melts in a zone above the tuyeres and sinks
with the lead. Minor elements (e.g., cadmium) either vaporize
and exit with the zinc in the off-gas or sink with the lead and
. 2
slag.
Under the tuyeres is a hearth which receives the bullion,
slag, matte, and speiss. These materials are tapped into a
forehearth approximately every 90 minutes. Because of different
specific gravities, the bullion and slag separate in the fore-
hearth. The bullion is tapped into a ladle and cast into 3 Mg
ingots, and the slag overflows into a granulation system. The
matte and speiss layer builds up until it is tapped; this occurs
2
once every 5 or 6 times the bullion is tapped.
Lead is the washing fluid used in the condensers to capture
the zinc. Gases meet a countercurrent lead spray, which is pro-
jected by rotor blades in the condensers. The zinc is carried
out of the condensers by the lead, and the mixture is separated
in cooling launders.
The Imperial Smelting process is very flexible. The ratio
of lead to zinc in the sinter can vary considerably, and oxide or
sulfide concentrates or ores can be used. Reducing conditions in
the furnace can be varied, as can the operating conditions in the
condensers.
The Imperial Smelting process offers no environmental ad-
vantages over other techniques, and lead pollution remains a
concern. The process does, however, allow treatment of complex
ores and use of coke rather than electricity.
References
1. Walsh, T. Zinc via Imperial Smelting Process Wins Support.
American Metal Market, 85(135):7, July 15, 1977.
2. Bonnemaison, Jean, et al. Imperial Smelting Furnace of
Perarroya. In: AIME World Symposium on Mining and Metal-
lurgy of Lead and Zinc. Vol. 2. Port City Press, Inc.,
Baltimore, Maryland, 1970.
35
-------
Bibliography
Biala, N., et al. Radioactive Isotopes in an Imperial Smelting
Furnace. Journal of Metals, 25(8):22-30, August 1973.
Binetti, G., et al. Combined Zinc-Lead Smelting: Recent Prac-
tice and Developments. Journal of Metals, pp. 4-11, Sep-
tember 1975.
Imperial Smelter Commissioned in Yugoslavia. Engineering and
Mining Journal, 174(8):127-128, August 1973.
Lee, R.W., et al. Cleaning Fume From Zinc-Lead Smelting. Fil-
tration and Separation, 15(3):197-198, 200-203, May/June
1978.
Morgan, S.W., and D.A. Temple. The Place of the Imperial Smelt-
ing Process in Nonferrous Metallurgy. Journal of Metals,
pp. 23-29, August 1967.
Sellwood, R.M. No. 4 I.S.F. Smelter Complex of Imperial Smelt-
ing Corporation, Ltd. In: AIME World Symposium on Mining
and Metallurgy of Lead and Zinc. Vol. 2. Port City Press,
Inc., Baltimore, Maryland, 1970.
2.14 BOLIDEN DIRECT REDUCTION PROCESS
Boliden a.-b. has developed a process for the direct smelt-
ing of lead concentrates and used it for some time at Ronnskar,
Sweden. Several process variations, including the use of an
electric furnace and TBRC's, have been tried. The capacity of
1 9
the Ronnskar facility is about 150 Mg of lead per day. '
Dried, unsintered lead concentrates (72 percent lead) and
recycled flue dust are blended and charged to an electric smelt-
ing furnace by variable-speed screw conveyors through pipes into
1 2
openings in the roof of the furnace. '
Nozzles for combustion air are inserted through the charging
openings in the top of the furnace. Horizontal jets of air are
aimed at the incoming material, and a cyclone-like suspension is
formed. This action provides sufficient time for burning the
bulk of the sulfur and oxidizing the iron and zinc. Recently,
36
-------
oxygen enrichment (10 percent) has been instituted.2 The mois-
ture content of the concentrate, however, must be below 2
percent for proper reaction. Most of the lead sulfide is reduced
to metallic lead and sulfur dioxide. Because sufficient heat is
not generated to sustain this reaction, additional electrical
1-3
energy must be supplied.
The solid and molten products are collected in the slag,
where the reactions are completed, and the slag and lead are then
separated. Coke breeze is applied to the surface of the slag
bath to reduce the lead oxide content. Dross and matte from
downstream converters are crushed and charged with limestone and
dust from the waste heat boiler to the slag bath. The slag is
then electrically heated to about 1350°C, and lead is tapped
directly to one of two converters in 70-Mg batches. The con-
verters also receive recycled dross from the refinery.
After a blow lasting about 40 minutes at a rate of 85
m /min, the lead is allowed to cool in the converter for 7 to 9
hours. During this period, molten lead-copper matte and dross
are separated from the crude lead. The lead is then transferred
to the refinery in ladles, and the dross is crushed with most of
the matte and recycled to the furnace.
Furnace slag is tapped in 18-Mg batches and transferred by
ladle to a fuming furnace. Gases leave the furnace at around
1050°C and are cooled in a waste heat boiler and by air cooling.
After passing through a dry electrostatic precipitator, the S02~
laden gases enter a sulfuric acid plant. Dust is continuously
recycled.
The direct reduction process eliminates some problems asso-
ciated with traditional lead smelting techniques, such as high
energy consumption and large volumes of gas with low S02 concen-
trations. Whereas traditional processes can require 2000 kWh/Mg
lead bullion smelted, the Boliden direct reduction process needs
only 1050 kWh/Mg.2 The lead bullion produced, however, contains
about 3 percent sulfur, which is removed in the converters.
37
-------
A new direct reduction process recently developed by Boliden
uses a TBRC (see Subsection 2.12). The main feature of the new
process is autogenous smelting of dried concentrates (72 percent
lead) with highly enriched air (60 percent oxygen) in a TBRC.
The energy consumption for the oxygen supply (150 m /Mg lead) is
about 70 kWh; therefore, energy demand for the smelting step is
decreased to about 70 percent of the energy demand of the old
2 4
Boliden reduction process. '
References
1. Elvander, H.I. The Boliden Lead Process. In: Symposium
on Pyrometallurgical Process in Nonferrous Metallurgy, Vol.
39. Gordon and Breach Science Publishers, New York, pp.
225-245. 1967.
2. Peterson, S., et al. Autogenous Smelting of Lead Concentrate
in TBRC. Paper No. A77-11. The Metallurgical Society of
AIME, New York, 1977.
3. Boliden Adopts TBRC for Pb and Now Cu Smelter. World
Mining, 31(5):44-48, May 1978.
4. Matyas, A.G., et al. Metallurgy of the Direct Smelting of
Lead. TMS Paper No. A75-80. The Metallurgical Society of
AIME, New York, 1975.
Bibliography
Halliburton, D-, and W.A. Lemmon. An Overview of Controls in
Primary Lead and Zinc. Air Pollution Control Directorate,
Environment Canada, Ottawa, Ontario. 1979.
2.15 BERGS0E WHOLE BATTERY SMELTING PROCESS
In 1975, Paul BergsjzJe and Son A/S began smelting whole
storage batteries at Golstrup, Denmark. This process eliminates
the dangerous job of breaking battery cases. Bergs^e now plans
to construct a $14-million facility using this technology in St.
2
Helens, Oregon.
The BergsjzJe process, which is used in Denmark and Sweden,
involves feeding whole batteries, battery scrap, lead dross,
38
-------
sludge, coke, iron, flux, and agglomerated dust to the top of a
water-jacketed shaft furnace similar to blast furnaces at a
primary lead smelter. Batteries are cracked, spent acid is
drained and neutralized, and soluble metals are precipitated.1'3'4
The furnace is equipped with two rows of tuyeres, one row on
either side, through which preheated (500°C) oxygen-enriched air
is injected. Preheated air has not previously been used in this
manner in secondary lead production. Special covers are fitted
over the tuyeres to minimize fugitive emissions during punching.
Four hooded slag taps and one hooded lead well are located at the
bottom of the furnace. The lead is cast into 1-Mg ingots in
pneumatic equipment. Under proper conditions, the furnace can
smelt continuously for many months.
The furnace off-gases are treated in an afterburner, com-
bined with cooling air, and routed to a fabric filter of Swedish
design that uses a felted polyester cloth. The afterburner
eliminates the need for a tall stack. Collected dusts are con-
veyed in an enclosed system to two flash agglomeration furnaces
before recycle. (Construction and operation of the flash ag-
glomeration furnaces are described in Subsection 5.8.) Gases
collected by the tap hole and lead well hoods are routed to the
fabric filter. The stack emissions have been reported to contain
' • 3 3
200 ppm SO,,, 5 to 10 mg particulates/m gas, 3 to 5 mg lead/m
1 3-5
gas, and trace quantities of chlorine. '
The whole smelting process is electronically controlled and
emissions are monitored. The smelter area is periodically rinsed
and the collected rinse water is treated by precipitation with
soda ash. The effluent is treated at a municipal wastewater
^ 1
facility, and the sludge is recycled to the shaft furnace.
The smelter produces a slag with a lead content of 0.6 to
1.0 percent lead; its disposition is not specified. The matte
contains 25 percent sulfur, 8 percent lead, and some copper and
can be sold to a lead smelter equipped with a roaster.
39
-------
Smelting whole batteries requires about 50 percent less
coke than conventional techniques because of the energy derived
from the cases. Use of preheated, oxygen-enriched air also
reduces coke consumption, but the preheater requires oil.
Additionally, the flash agglomeration furnace consumes about 25
liters of oil per hour.3 Electrical consumption is about 80 kWh
per megagram of charge.
The capital cost of the Golstrup facility is reported to
have been $731,452 in 1975, and the operating cost equals $37.95
per megagram of output from the shaft furnace.
References
1. MacKey, T.S., et al. Smelting of Unbroken Batteries.
Presented at the 106th Annual Meeting of the American
Institute of Mining, Metallurgical, and Petroleum Engi-
neers, Atlanta, Georgia, March 9, 1977.
2. U.S. Department of the Interior. Danish Company Plans
Scrap Lead Plant in Oregon. Minerals and Materials: A
Monthly Survey. Bureau of Mines, Washington, D.C., Decem-
ber 1978.
3. MacKey, T.S., et al. Flash Agglomeration of Flue Dust.
Journal of Metals, 39 (11):12-15, November 1977.
4. Coleman, R.T., et al. Trip Report to Paul BergsjzJe and Son
A/S, Boliden Aktiebolag, and Outokumpu Oy. Prepared by
Radian Corp. for the U.S. Environmental Protection Agency
under Contract No. 68-02-2608. Austin, Texas, November 28,
1977.
5. Coleman, R.T. Emerging Technology in the Secondary Lead
Industry. Prepared by Radian Corp. for the U.S. Environ-
mental Protection Agency under Contract No. 68-02-2608.
Austin, Texas, 1978.
Bibliography
Herbert, I.C., and J.F. Castle. Extractive Metallurgy. Mining
Annual Review (United Kingdom), 1978.
40
-------
2.16 OLIFORNO WHOLE BATTERY SMELTING PROCESS1
Accumulatoren Fabrik of Oerlikon, Switzerland, operates a
secondary lead facility that treats whole batteries in a kiln.
The metal/slag product, which can be handled hygienically, is
granulated and can be smelted at a high throughput rate in a
short rotary furnace.
The Oliforno process produces more flue dust than does the
Bergsjrfe process, and much of the fine sulfur in the feed passes
into the off-gases. The gases are routed through an afterburner
and indirect cooler before they are treated in an ESP.
Reference
1. Herbert, I.e. and J.F. Castle. Extractive Metallurgy.
Mining Annual Review (United Kingdom), 1978.
2.17 PYROMETALLURGICAL SLAG TREATMENT1'2
Russian scientists have developed a pyrometallurgical method
of treating molten slag for the extraction of zinc, lead, tin,
and other nonferrous metals. This process is similar to the slag
fuming technology employed at many U.S. lead smelters. This
process, however, uses combustion products from natural gas
burned outside the furnace as a reducing agent, rather than
pulverized coal injected directly into the furnace.
Molten slag is fed into the slag fuming furnace, and combus-
tion products are blown over the melt by means of a special
device developed for gas burning and feeding. The resulting gas
stream is cooled in a waste-heat boiler, where coarse metal
sublimates are collected. The off-gas is then sent to dust
collecting and gas cleaning equipment. Figure 13 illustrates the
slag fuming process. Metal values recovered can be further
concentrated and processed by known methods.
Two plants in commercial operation in the Soviet Union treat
slags containing nonferrous metals.
41
-------
WASTE HEAT
BOILER
SLAG FUMING
FURNACE
DUST CLEANING AND
COLLECTING EQUIPMENT
COARSE METAL
SUBLIMATES
COMBUSTIBLE-GAS
FIRING DEVICES
Figure 13. Slag fuming process.'
-------
References
1. Ageey, I.P.^et al. Pyrometallurgical Method of Treating
Slags, Containing Nonferrous and Rare Metals. Canadian
Patent No. 832,378, Canadian Patent Office, Ottawa, Canada.
2. Pyrometallurgical Method of Slag Processing. Russian
Technology No. 1-765. Licensintorg, Moscow, Soviet Union.
Distributed by Southwire Company, Carrollton, Georgia.
2.18 SEA NODULE PROCESSING
Processing technology for the recovery and extraction of
deep-sea ferromanganese nodules is beginning to unfold. Several
processes have been developed on a laboratory scale; others are
still on the drawing board. The process technologies ultimately
used will depend on which metals can be economically recovered.
Nickel, cobalt, copper, and manganese are all present in quanti-
ties that make recovery attractive. On the average, Pacific
Ocean nodules are composed of 27 percent manganese, 6 percent
iron, 1.5 percent nickel, 1.3 percent copper, and 0.25 percent
cobalt. Most of the manganese will be discarded to avoid glut-
2
ting the market. Several large-scale processing experiments
are planned in the next several years, and these marine resources
should be developed within the next 10 years. A discussion of
process technology already under development follows.
The Institute of Metallurgy and Electrometallurgy at Aachen
University, West Germany, has been investigating a process for
the elemental separation of manganese nodules. Manganese nod-
ules, ground to particle size from 60 to 120 iam, are treated in
a plasma reduction furnace with hollow electrodes. The reaction
particles are obtained as agglomerated particles or as metal and
slag phases.3 This process has the following advantages:
Weight reduction is carried out near the actual location of
the nodule deposits.
Reaction rates are faster.
The size of the particles is greater.
43
-------
The resulting products can be processed further in hydrocyclones
and by selective flotation.
A hybrid process in operation in Canada to recover copper,
nickel, and cobalt from sea nodules includes a pyrometallurgical
reduction kiln followed by smelting in an electric furnace, and
then hydrometallurgical leaching in an oxygen atmosphere to
extract the metals. Although this highly specialized process is
intended primarily for nickel recovery, it also recovers copper
equivalent to 60 percent of the weight of the nickel. The esti-
4
mated capital cost of a commercial-scale plant is $540 million.
Another process, developed in Hawaii, recovers copper and
nickel by hydrometallurgical techniques. The ferromanganese
nodules react with an aqueous solution of oxalic acid, which
results in the reduction of Mn (IV) to Mn (II) and the subsequent
evolution of CO- and the solubilization of all the transition
metals. Nickel and copper are selectively extracted from the
solution with a commercially available reagent, LIX 64N.
Kennecott Copper Corporation has developed a low-temperature
ammonical leach process to extract metal values from manganese
nodules. Nodules are wet-ground to a fine mesh, and carbon
monoxide is added to the slurry as a reducing agent. A counter-
current decantation wash, carried out in a series of thickeners,
further solubilizes the metal values. The decant solution con-
tains all the copper, nickel, and cobalt, while the tailings
4
contain the manganese. The estimated capital cost of a com-
2
mercial-scale plant is $340 million.
References
1. Burzminski, M.J., et al. Extraction of Copper and Nickel
from Deep Sea Ferromanganese Nodules. Analytical Chemistry,
50(8):1177-1180, July 1978.
2. Cashing in on the Ocean. Public Television Broadcast -
Nova, WGBH, Boston. Boston, Massachusetts. 1979.
3. Marnette, W., et al. Plasma Furnace for Processing Manganese
Nodules. Mining Magazine, 136(5):417, May 1977.
44
-------
4. Processes to Refine Ocean Nodules Assessed. Chemical and
Engineering News, 56(12):22-23, March 20, 1978.
Bibliography
Agarwal, Dr. J.C., et al. A New Fix on Metal Recovery from Sea
Nodules. Engineering and Mining Journal, 177 (12):74-78,
December 1976.
Deep Ocean Floor Nodule Mining - First Generation Techniques are
Here. Mining Engineering, 27 (4):47-52, April 1975.
Environmental Impact of Deep Sea Mining. Journal of Metals,
29(11):6, November 1977.
Frank, R.A. the Promise of Deep Sea Mining. Mining Congress
Journal, 64(2):73-74, 121-122, February, 1978.
Howard-GoIdsmith, R.C. New Techniques in Copper Mining. Mining
Magazine, 138(2):111,113,115,117,118.
Minerals-Hungry U.S. Is Fishing for Oddball Ores. Chemical
Week, May 30, 1970. •
Moncrieff, A.G., et al. The Economics of First Generation Man-
ganese Nodule Operations. Mining Congress Journal, Vol.
60: 46-52, December 1974.
Seabed Mining - Background and Current Outlook: Systems, Methods.
World Mining, 30(13):54-58, 80, December 1977.
Sisselman, R. Ocean Miners Take Soundings on Legal Problems,
Development Alternatives. Engineering and Mining Journal,
176(4):75-86, April 1975.
Thiry, H.B., et al. French Exploration Seeks to Define Mineable
Nodule Tonnages on Pacific Floor. Engineering and Mining
Journal, 178(7):86-87, 171, July 1977.
Tinsley, C.R. Nodule Miners Ready for Prototype Testing.
Engineering and Mining Journal, 178 (1):80-101, January
1977.
Tinsley, C.R. Processing - No Longer a Problem. Mining Engi-
neering, 27(4):53-55, April 1975.
Wadsworth, M.E. Hydrometallurgy. Journal of Metals, 29(3):8-13,
March 1977.
Welling, C.G. Next Step in Ocean Mining - Large Scale Testing.
Mining Congress Journal, 62 (12):46-51, December 1976.
45
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2.19 AUTOMATIC TUYERE PUNCHING1
As part of an effort to boost copper production and increase
sulfur recovery at its Legnica smelter, the Kombinat Gorniczo-
Hutniczy Miedzi. (KGHM) of Poland has developed an automatic,
pneumatic-converter tuyere puncher. The compact punchers are
installed on 33- and 45-Mg Hoboken converters at Legnica and an
80-Mg Hoboken converter at Glogow. They also can be installed on
Peirce-Smith units, as some domestic smelters practice automatic
punching.
The manual removal of slag accretions in tuyeres is a slow,
hazardous task requiring extra manpower. Mechanical punchers are
available, but their use is restricted by the limited space on
the tuyere side of the converter, they require an additional
operator, and they are quite expensive. The KGHM automatic
tuyere puncher eliminates these problems.
The automatic puncher is composed of three assemblies: a
frame assembly that allows the puncher to be moved from one
tuyere xto another, a punching assembly consisting of a punching
rod and a pneumatic drive, and a power supply and automatic
electric/pneumatic control unit. Depending on such local condi-
tions as the space on the converter tuyere side, a puncher weighs
1.5 to 2.5 Mg and is about 2 to 3.5 m long.
The unit can be operated in three different modes. In the
fully automatic mode, the signal activating the puncher reduces
the airflow through the tuyeres below a preselected value. In
this mode, any - tuyere can be eliminated from the punching cycle.
If, after one cycle the quantity of air actually blown into the
converter remains below the preselected value, the cycle is
repeated. Figure 14 illustrates this mode of operation.
In the semiautomatic mode, the converter operator simply
presses a button and a complete punching cycle is performed. If
desired, individual tuyeres can also be eliminated from the cycle
in this mode.
46
-------
AIR.
ORIFICE
PLATE
TRANSDUCER
MANUAL AIR
FLOW SETTING
REMOTE
AIR FLOW
SETTING
MEASURED VALUE
OF AIR VOLUME
'CONTROLLER
TRANSDUCER
CONVERTER
LJLJLJLJLJLJI-II-JLJ
L
AUXILIARY
CONTROL PANEL
CONTROL ROOM
J
PUNCHER
Figure 14. Automatic tuyere punching system.
1
-------
In the fully manual mode, the converter operator controls
the unit with push buttons that move it to the left or right and
activate the puncher.
It requires about 3 to 4 seconds to clean one tuyere, and
move the puncher to the next tuyere. Therefore, the punching
cycle at a 45-Mg Hoboken-type converter equipped with 28 tuyeres
lasts less than 2 minutes, and no special operator is required.
Reference
1. Demidowicz, L., and W. Kozminski. Automatic Tuyere Punch-
ing. Engineering and Mining Journal, February 1979.
48
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SECTION 3
HYDROMETALLURGICAL PROCESSES
3.1 MINEMET PROCESS1
Minemet Recherche has patented a hydro-metallurgical process
for the recovery of metals from sulfide ores and concentrates.
The process, developed on a laboratory scale, is designed to
recover copper from chalcocite, covellite, and chalcopyrite. In
this process, which is illustrated in Figure 15, copper concen-
trate is first roasted to remove impurities, and then selectively
leached by cupric chloride dissolved in a suitable solution, such
as an alkali or alkaline earth chloride, ammonium chloride, or
ferrous chloride. During this leach, sulfur associated with the
ore is converted to the elemental state and remains in the leach
residue for later recovery.
After the leach residue has been filtered out, the leach
solution containing cuprous chloride and ferrous chloride is
separated into two streams. The iron is precipitated as goethite
by simple oxidation of the leach solution by air. The dissolved
copper is recovered by solvent extraction and electrowinning.
After the copper has been extracted the remainder of the solution
is bled off for the recovery of other metals.
Elemental sulfur remaining in the leach residue may be
recovered easily, by flotation or fusion, for internal use or for
sale. After the sulfur is removed, the leach residue still
contains pyrite, which could be an unstable solid waste.
The purported advantages of the Minemet process are high
copper recovery, recovery of iron in the form of geothite and
elemental sulfur (both marketable), low capital and operating
49
-------
COPPER SULFIDE ORES
OR CONCENTRATES
CuCl,
LEACHING
AIR
V V
GOETHITE
PRECIPITA-
•\TION
FeOOH
CuCl2
^GANGUE
r rl AND SULFUR
CuCl I
FeCU AIR
I'
1 .
Cu
SOLVENT
EXTRACTION
STRIPPING
2so4
Cu ELECTRO-
WINNING
I
^•i
«
/BLEED
iri RFrn
OF (
ME!
^
VERY
)THER
•ALS
-
Cu CATHODES
Zn, Pb,
Cu
Figure 15. Minemet process for copper.
50
-------
costs, and the ability to meet antipollution regulations. This
process still must be tested on a pilot-plant and semicommercial
scale, however, before becoming a viable technique.
This process, like most that include electrowinning, con-
sumes more energy than pyrometallurgical methods.1
Reference
1. Demarthe, J.M., et al. A New Hydrometallurgical Process for
Copper. Presented at the 105th Annual Meeting of the AIME,
Las Vegas, Nevada, February 22-26, 1976.
3.2 SHERRITT-COMINCO COPPER PROCESS
In conjunction with the Canadian government, Sherritt
Gordon Mines, Ltd., and Cominco, Ltd., have developed a hydro-
metallurgical technique for producing copper from sulfide con-
centrates. The Sherritt-Cominco (S-C) process is an adaptation
of Sherritt Gordon's pressure leaching system for processing
nickel concentrates. An integrated 8-Mg/day pilot plant de-
veloped for testing the S-C copper process was constructed at
Fort Saskatchewan in 1975 and operated in 1976. Results from
these tests proved so satisfactory that feasibility studies were
conducted on the basic design parameters of a plant in western
Canada and one in Arizona. Both would produce 68,000 Mg of
copper rod a year from chalcopyrite concentrates. The studies
indicate the S-C process could compete with other pyrometallur-
2 3
gical and hydrometallurgical processes. '
Copper concentrate is pelletized, dried, and fed to Herre-
shoff-type multiple-hearth roasters. The 690°C temperature in
the upper hearths drives off some sulfur from the pellets. The
preheated pellets mix with counter-flowing hydrogen gas when they
reach the bottom of the roaster, and additional sulfur is removed
from the feed as hydrogen sulfide is formed. Some of the sulfur
and hydrogen sulfide burns the height of the vessel when air is
51
-------
introduced to the roaster. As shown in Figure 16, the combustion
gases are sent to an acid plant for the production of sulfuric
acid.
After leaving the roaster, the activated concentrate is fed
to a series of rubber-lined leach tanks containing sulfuric acid.
Iron sulfide dissolves in the acid and hydrogen sulfide is
liberated. Most of the hydrogen sulfide gas is sent to a small
4
Glaus plant to be recovered as elemental sulfur.
The copper-free leach liquor is separated from the solids in
a settling tank and undergoes oxidation in a pressurized auto-
clave. Iron precipitates as jarosite, and sulfuric acid is re-
4
generated.
Residue from the acid leach is subject to high-pressure
leaching in a neutral copper sulfate solution. Much of the zinc
and remaining iron dissolve and are replaced by copper. After
liquid-solid separation, the zinc-rich solution is treated with
hydrogen sulfide, which causes zinc sulfide to precipitate. The
4
zinc sulfide is then collected and sold as a byproduct.
The neutral leach residue is treated with recycled acid and
oxygen in a two-stage oxidation leach in which copper, residual
iron, and zinc are dissolved and elemental sulfur is formed.
Over 98 percent of the copper is extracted in the first and
4
second stages of oxidation leaching.
Solid residue from the oxidation leach contains precious
metals, molybdenum sulfate (if present in original feed), ele-
mental sulfur, and gangue. Gangue is rejected in a small flota-
tion unit, and sulfur is removed by filtration and solvent
extraction. The small quantity of precious metal concentrate
4
left can be sold or processed on site.
Copper solution from oxidation leaching requires further
purification before it undergoes electrolysis. Because selenium,
tellurium, iron, arsenic, and bismuth would contaminate the final
copper or lower efficiency of the electrolysis, these impurities
52
-------
COPPER
CONCENTRATE
I
ROASTER
so2
I
*j
"1
^
T ^
CLAUS
PLANT
cm run
AIR, H2,
NATURAL GAS
i±fr&^ LEACH SOLUTION
(CONTAINS IRON)
NEUTRAL LEACH
OXIDATION LEACH
JL JL JL
SILVER/GOLD
RECOVERY
ELECTROPLATING
SOLUTION
PURIFICATION
LEACH SOLUTION
(CONTAINS ZINC, IRON)
SPENT
ELECTROLYTE
Figure 16. Simplified flowsheet of the S-C process
53
-------
are precipitated in a series of high-temperature autoclaves.
After pressure filtration, the pure copper solution passes to the
electrowinning plant.
The electrowinning plant incorporates many recent design
advances, one of which is the use of strippable titanium cathodes
instead of starting sheets. Copper is deposited on the cathodes,
carefully washed, stripped, and then melted and processed to
rod.4
Most of the sulfur found in complex sulfide concentrates is
processed into elemental sulfur rather than sulfuric acid. This
minimizes problems such as meeting emission regulations, trans-
porting acid, and selling the material. Arsenic is removed
during purification as stable and insoluble ferric arsenate.
Iron precipitated as jarosite is impounded in a lined pond. One
problem area may be the production of hydrogen sulfide gas during
acid leaching. Another may be the presence of residual elemental
4
sulfur in the waste gangue.
Like other hydrometallurgical copper processes, the S-C
copper process consumes more energy than pyrometallurgical proc-
esses. The S-C process consumes 92 percent more-fossil fuel than
a facility utilizing a flash smelting/electrorefining/acid plant
process stream; however, in the future, pyrometallurgical
facilities may be required to invest more money and energy to
dispose of S02 as sulfur or neutralized sulfate.
The capital cost of a plant employing the S-C process is
purported to be substantially less than that of a comparable
flash smelting/electrorefinery, but operating costs are estimated
to be slightly higher. Should the flash smelting/refinery be
required to add neutralization facilities, however, operating
costs of the S-C process would compare favorably.
References
1. Filmer, A.O., A.J. Parker, M. Ruane, and L.G.B. Wadley.
Metal and Sulfur Recovery from the Copper Sulfide Precipi-
tate in the Sherritt Gordon Process. In: Proceedings
of the Australian Institute on Mining and Metallurgy, No.
268, December 1978.
54
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2. Swinkels, G.M., et al. The Sherritt-Cominco Copper Process.
CIM Bulletin, 71 (790) :105-139, February 1978.
3. Maschmeyer, D.E-G., et al. Application of Sherritt-Cominco
Copper Process to Arizona Concentrates. Journal of Metals,
pp. 27-31, July 1978.
4. Selling Technology Adds to S-G Profits. Canadian Chemical
Processing, 62(3):33-36, March 1978.
5. Rosenzweig, M. Copper Makers Look to Sulfide Hydrometal-
lurgy. Chemical Engineering, pp. 79-81, January 5, 1976.
Bibliography
Copper Makers Look to Sulfide Metallurgy. Chemical Engineering.
January 5, 1976.
Peters, E. Direct Leaching of Sulfides: Chemistry and Applica-
tions. The 1976 Extractive Metallurgy Lecture, the Metal-
lurgical Society of AIME, Metallurgical Transactions B,
Volume 7B, December 1976.
3.3 ACETONITRILE EXTRACTION PROCESS1
A technique developed in Australia promises energy savings
and decreased sulfur dioxide emissions. Original work on the
acetonitrile extraction process began at Australian National
University; final investigations were made by scientists of the
Mineral Chemistry Research Unit at Murdoch University in Perth,
Western Australia. The process is capable of recycling scrap
copper and refining concentrates faster than conventional methods.
Laboratory experiments have shown that copper can be ex-
tracted efficiently by leaching cuprous sulfate solutions from a
roasted concentrate by use of the organic chemical acetonitrile.
Distillation of the organic produces pure metallic copper powder,
thereby eliminating the electrowinning step. Because acetoni-
trile is decomposed, however, it requires continuous replacement.
Little or no sulfur dioxide and particulates are released
into the atmosphere; however, acetonitrile is a poisonous chem-
ical, and widespread use of this process may cause hazardous
secondary pollution.
55
-------
Energy requirements for this process are less than 60
percent of those for conventional processing. The process uses
a low-grade waste steam rather than electricity or oil.
Reference
1. Propose Copper Leaching Process. Industrial Research, 18:
38-39, August 1976.
3.4 LURGI-MITTERBURG PROCESS1
In a joint effort, Lurgi Chemi and Huttentechnik, G.m.b.H.,
of Frankfurt/Main, West Germany; Kupferbau-Mitterburg of Miihlbach,
Austria; and the Technical University of West Berlin have de-
veloped a pressure acid-leaching system for the recovery of
copper values. The process looks particularly attractive for
smaller plants and for handling concentrates containing antimony
and arsenic, which can cause problems in high temperature proces-
sing. A pilot plant treating 3 Mg of concentrate per day in
Miihlbach began operating in the spring of 1974 and continued
through the first quarter of 1976. The next step will be the
construction of a full-scale plant.
Copper concentrates are fed into a vibrational mill for
grinding to reduce their size. The grinding also causes a high
degree of distortion in crystal lattices, which enhances leach-
ing. The concentrate then is slurried with spent sulfuric-acid
electrolyte and pumped into an autoclave, where (at 115°C and an
oxygen partial pressure of 10 to 20 atmospheres) the copper
dissolves and forms copper sulfate. Following depressurization
and discharge from the autoclave, the leach residue, which
contains all the iron and sulfur, is thickened and filtered. The
filtrate is sent to rubber-lined electrolytic cells for electro-
winning of the copper; spent electrolyte is recycled.
Reference
1. Rosenzweig, M. Copper Makers Look to Sulfide Metallurgy.
Chemical Engineering, pp. 70-81, January 5, 1976.
56
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3.5 PRESSURE-LEACH PROCESS FOR ZINC RECOVERY
Sherritt Gordon Mines, Ltd., and Cominco, Ltd., of Canada
are cooperating in the development of an alternative to conven-
tional roast-leach methods of zinc recovery. The new process
involves use of a pressure-leach reactor to produce zinc sulfate
directly. In 1977, a pilot plant was operated in Fort Saskatch-
ewan, Alberta, and Cominco plans to construct a 64,000-Mg/yr, $20
million commercial unit at its Trail, British Columbia, facil-
ity.1'2
As illustrated in Figure 17, the feed concentrate at Trail
will be split; part will go to the new system and part to the
existing roasters. That portion routed to the pressure-leach
autoclave will be ground in a wet ball mill to less than 325 mesh
before it is combined with recycled sulfuric acid electrolyte in
the autoclave. The first autoclave to operate at Trail will be
3.7 m in diameter and 9.1 m long and will consist of a mild steel
1 2
shell lined with lead and acid brick. '
Reactions in the autoclave will take place at 150°C and 10.5
2
kg/cm . Zinc sulfide, lead sulfide, and iron sulfide will react
to form simple sulfates, elemental sulfur, and water; iron will
precipitate as complex sulfate. The slurry will be flashed as it
leaves the autoclave and then treated in a flotation unit for
sulfur separation. The process stream will then be joined with
the conventionally treated concentrate at the acid leaching step
for combined treatment in the thickeners and filters.
The pressure-leach system offers certain environmental ad-
vantages in that it does not produce S02, but rather recovers
sulfur in the elemental form. Also, pressure leaching does not
form zinc containing residues that require additional treatment.
Cominco estimates that a grass-roots pressure-leach plant
offers a 20 to 25 percent capital investment savings, compared
with a grass-roots roast-leach plant. This can constitute a $35
million savings for each 64,000-Mg/yr plant. Operating costs
should prove to be similar for conventional and pressure-leach
operations.
57
-------
BALL MILL
Ih;
ZINC CONCENTRATE
STEAM TO
PREHEAT
FLOTATION
RECYCLED
ELECTROLYTE
PRESSURE LEACH
AUTOCLAVE
FLASH
TANK
ELEMENTAL
SULFUR
STEAM
00
i
^f
^so2
WASTE HEAT
BOILERS
ACID PLANTS
\
ROASTERS
ZINC CONCENTRATE
CALCINE
ACID THICKENER
LEACHING
1
NEU7
r
FILTERS
\'\'
RECYCLED
ELECTROLYTE
ZINC DUST
FROM SMELTING-i
1 * I-*
l-U-j ^-*I
i
ELECTROWINNING
THICKENER
ZINC
METAL
RESIDUES TO SMELTER
Figure 17. Zinc recovery by pressure-leach autoclave, combined with
a r-nnwoirMnnal nlant at Tva i 1 . R . C. '
a conventional plant at Trail, B.C.
-------
References
1. Whiteside, D. Simplified Zinc Process Does Not Generate
S02. Chemical Engineering, pp. 104-106, August 13, 1979.
2. Cominco Building Worlds First Zinc Pressure Leaching Plant
of 70,000 TPY Capacity. The Northern Miner, April 19, 1978.
3.6 ELECTROLYTIC ZINC PROCESS
The Electrolytic Zinc Company of Australia (EZ) has devel-
oped a method for treating oxidized zinc ores, which the New
Jersey Zinc Company has modified and piloted with the intention
of utilizing it in Thailand. The EZ process is advantageous
because oxidized zinc ore is relatively abundant.1"3
Figure 18 shows the process flow as piloted by New Jersey
Zinc. The ore is ground and then leached with a sulfuric acid/
zinc sulfate solution in a series of three large agitated tanks;
leaching is continuous. Efficiency of zinc extraction ranges
from 90 to 98 percent, depending largely upon the pH in the last
leach tank. Neutralization also takes place in a series of three
agitated tanks; in this step, dissolved silica and other impuri-
ties (e.g., arsenic and iron) are precipitated. Limestone, zinc
calcine, or zinc oxide can be used as the neutralizing agent.
After they are neutralized, solids are separated from the
zinc solution. This step involves use of a filter of a type yet
to be determined; rotary-drum vacuum filters, horizontal-belt
vacuum filters, and pressure filters have been tested. Because
the leaching of low-grade oxidized ores yields four to five times
as much solid residue per unit of zinc as is generated in a
conventional electrolytic zinc plant, quick and efficient de-
watering of residue is required.
The presence of certain impurities in the neutral filtrate
can result in difficulties during electrolysis (as in conven-
tional electrolytic refining). Arsenic, antimony, germanium,
nickel, iron, cobalt, copper, and selenium all must be removed
59
-------
NEUTRAL
SOLUTION
GRINDING AND
CLASSIFYING
ZINC
ORE
MAKEUP
SULFURIC ACID
LEACHING
NEUTRALIZATION
PRIMARY
WASH
WATER
LIMESTONE
BASIC ZINC
LIMESTONE
SIN cflTf
5ULt-«it
SULFATE
PRECIPITATION
FILTRATION
ZINC
DUST
RESIDUE
RESIDUE
WASHING
PURIFICATION
CELL
ACID
RESIDUE FOR
DISPOSAL
PURGE OF
IMPURITIES
r~t
RESIDUE TO
CADMIUM PLANT
EFFLUENT
TREATMENT
ELECTROLYSIS
RESIDUE
STORAGE
CATHODE ZINC TO
MELTING AND CASTING
Figure 18. Simplified flowsheet of the EZ process
as piloted by the New Jersey Zinc CoJ
60
-------
from the solution. Iron and silica are removed during neutrali-
zation, but the others require an additional purification step.
Cementation with zinc powder in conjunction with other reagents
(such as arsenic, antimony, or copper) is a widely applied tech-
nique and the probable method of choice for the EZ process. The
purified solution is routed to electrolytic cells for conven-
tional zinc recovery.
As a result of the successful pilot tests, New Jersey Zinc
plans a 60,000-Mg/yr plant for Tak, Thailand. This facility is
expected to require 4500 kwh electricity per ton of product zinc
and an estimated fixed capital (1977) investment of $800 to $1000
per ton of annual zinc capacity ($48 to $60 million).
References
1. Wood, J.T. , et al. Elecrolytic Recovery of Zinc from Oxi-
dized Ores. Journal of Metals, 29(11):7, November 1977.
2. Wadsworth, M. E. Hydrometallurgy - Annual Review. Journal
of Metals, pp. 32-39, April 1978.
3. Gordon, A.R. Improved Use of Raw Materials, Human and
Energy Resources in the Extraction of Zinc. Electrolytic
Zinc Company of Australasia, Ltd., Melbourne.
Bibliography
Duby, P., et al. Electrometallurgy - Annual Review. Journal of
Metals, April 1978.
Matthew, I.G., and D. Eisner. The Processing of Zinc Silicate
Ores - A Review. Metallurgical Transactions B. Vol. 8B.
Metallurgical Society of the AIME, New York, 1975.
Matthew, I.G., et al. The Hydrometallurgical Treatment of Zinc
Silicate Ores. Metallurgical Society of the AIME, New York,
1975.
U.S. Patent Office. Patent No- 3,656,941. Hydrometallurgical
Treatment of Siliceous Zinc Ores. 1972.
61
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3.7 OTHER TECHNIQUES FOR TREATING OXIDIZED ZINC ORES
Oxidized zinc ores have been used in the production of zinc
metal since the industry's development. At one time they were
the major component of zinc smelter feed, however, their use has
declined because of the inability of electrolytic plants to
process them satisfactorily. Oxidized ores contain significant
amounts of zinc silicates, which dissolve in sulfuric acid to
form silicic acids and colloidal silica. These compounds gel and
interfere in the separation of zinc liquor from residue solids.
Nevertheless, the large amount of oxidized ore throughout the
world and the ease with which it can be mined have encouraged the
development of several processing techniques.
The Radino process was patented in the United States in 1959
as a method for treating a Brazilian zinc silicate ore. This
process utilizes a lengthy leaching cycle and numerous additions
of aluminum sulfate to flocculate the silica and render it fil-
terable.1"3
The "Italian" process (developed by F. Sciacca and patented
in Brazil in 1963) is another process for treating Brazilian zinc
silicate ores. In this process, leaching is carried out in
autoclaves at elevated temperatures to form a granular colloidal
silica.
The Vielle-Montagne process incorporates a series of leach-
ing vessels in which the acidity is progressively increased at
temperatures between 70° and 90°C. The operation lasts 8 to 10
hours and results in the silica precipitating in a crystalline
form. The process was patented in the United States in 1976.4
References
1. Wood, J.T., et al. ELectrolytic Recovery of Zinc from Oxi-
dized Ores. Journal of Metals, 29(11):7, November 1977.
2. Perry, W. Refining Zinc Silicate Ore by Special Leaching
Technique. Chemical Engineering, October 10, 1966.
62
-------
3. U.S. Patent Office. Patent No. 2,874,041. Process of Zinc
Extraction From Ores Formed by or Containing zinc Silicate
or Other Soluble Silicates by Means of Hydrometallurgv.
1959.
4. U.S. Patent Office. Patent No. 3,954,937. Process for the
Treatment of Material Containing zinc and Silica for Recov-
ering Zinc by Hydrometallurgical Way. 1976.
3.8 HYDROCHLORIC ACID LEACHING OF FABRIC FILTER DUST1
Australian researchers have developed a procedure for re-
covering metal values from copper smelter fume. The Electrolytic
Refining and Smelting Company of Australia Pty., Ltd., sold all
of its Port Kembla fabric filter dust to Hardman chemical Pty.,
Ltd., for processing. The dust, valued in excess of $1 million
1977 Australian dollars a year, was leached with sulfuric acid to
obtain an easily purified zinc sulfate solution, and the leach
residue was resold. Declining profitability of the operation
eventually induced Hardman, along with the Mineral Chemistry
Research Unit of Murdoch University, to investigate the pos-
sibility of further treating the residue. Results of their
investigation led to the hydrometallurgical technique described
below.
After the copper smelter fabric filter dust is leached with
excess sulfuric acid and about 65 percent of the zinc is re-
covered as zinc sulfate, the solid residue is leached twice with
hydrochloric acid. This brings into solution more than 90
percent of the major metals, excluding lead. Hot-water washing
of the residue following the acid leach precipitates about 75
percent of the lead as lead chloride.
By proceeding in a stepwise progression, the technique
precipitates most of the metal fractions contained in solution.
The end products include zinc sulfate, zinc chloride, cadmium,
copper, tin, bismuth, and lead chloride. This dissolution,
classification, and recovery process is similar to methods used
in the processing of electrolytic zinc.
63
-------
Reference
1. Giles, D.E., and A. Boden. Hydrometallurgical Treatment of
Port Kembla Copper-Smelter Fume. Proceedings of the Aus-
tralasian Institute of Mining and Metallurgy, No. 262.
pp. 39-47, June 1977.
64
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SECTION 4
ELECTROLYTIC PROCESSES
4.1 FLUIDIZED-BED ELECTROLYSIS1
Akzo Zout Chemie of the Netherlands has developed an elec-
trolytic process for the removal of low concentrations of metals
from process or waste streams with high flow rates. The tech-
nique makes use of a closed fluidized-bed electrolytic (FEE)
cell. Possible applications include separation and winning of
metals in hydrometallurgical processes, treatment of liquid waste
streams before discharge, and electrolyte purification at an
electrolytic zinc plant.
As shown in Figure 19, an FEE cell consists of two compart-
ments separated by a diaphragm; these contain the anode and
cathode. In the cathode compartment, metal particles are flu-
idized by the waste or process stream, and additional metal par-
ticles (about 0.5 mm in diameter) are added. As the electrolysis
results in deposition of the waste or process metal upon the
added metal particles, they increase in size to about 1 mm and
sink to the bottom of the cell, where they are recovered.
The system can be easily expanded by the installation of
many diaphragms in one cathode compartment; thus only one cath-
olyte supply is required. The anodes are placed in the center of
the diaphragms. Because the diaphragms are relatively imperme-
able, the anolyte and catholyte streams remain segregated. This
segregation can be ensured by pressurizing the anode compartment
to prevent inward seepage by the catholyte. Recirculation of the
catholyte is necessary for maintenance of fluidization velocity
even when the feed supply is variable.
65
-------
DIAPHRAGM
ANOLYTE
ANODE
-
h
i
0 o 0 0
0 o 0
0 °
ooo
O o O
0 °
0 0 0
o
0 ° 0°
0 0 °
0 o 0
0 0 °
o o
k
o
o
o
o
o
o
o
o
o
o
o
o
o
CATHODE
(CURRENT FEEDER)
* LAI HULY 1 L
o°0° SUPPLY OF
^- ° SMALL ME!
PARTICLE!
• Fl IITnT7Fn RFD
ruu iu i Z.L.U DL.U
^_ n0 DISCHARC
^^oO OF GROW
0 METAL
PARTICLE
i r
ANOLYTE CATHOLYTE
Figure 19. Schematic view of an FEE cell.
1
66
-------
An Akzo test of FEE in the treatment of copper-containing
wastewater lowered the copper content from 100 mg/liter to less
than 1 mg/liter in one pass. In another test, involving the
removal of mercury from the brine stream of a mercury-cell,
chlor-alkali electrolysis plant, the mercury content of the brine
was reduced from 5 mg/liter to about 0.05 mg/liter. Ionic and
metallic mercury deposited on copper particles, which were re-
covered by distillation of the amalgamated particles.
Other applications have included separation of copper and
nickel in a nickel electrolyte and removal of copper and cadmium
from a zinc electrolyte. Use of FEE in the latter application
has been reported to be very attractive economically.
A cost comparison between FEE and zinc dust cementation for
electrolyte treatment at a 136,000-Mg/yr electrolytic zinc plant
reportedly indicates that the net operating profit achievable
with FEE is $2.8 million per year higher as a result of an extra
yield of 4500 Mg/yr of pure copper and 14 Mg/yr of cadmium.
Capital investment should be $2 million lower for FEE, but energy
requirements and thus operating costs are expected to be higher.
Reference^
1. Fluidized Bed Electrolysis for Metal Recovery at Low Con-
centrations. Mining Magazine, 139(3):313-317, September
1978.
i
4.2 DEXTEC PROCESS1
Dextec Metallurgical Pty., Ltd., of Australia is joining
with China to develop a new copper production process called
Dextec. Reportedly, this process is capable of producing copper
metal from concentrates without the use of a smelter and at a
cost 25 percent below those involved with conventional tech-
niques. The method converts ore to metal in one step, at atmos-
pheric pressure, through use of an electrolytic diaphragm cell.
It also recovers elemental sulfur, thus avoiding S02 air emis-
sions. The technique may prove applicable to other metals as
well.
67
-------
The Peking Institute of Mining and Metallurgy will build a
pilot plant to test the technique at commercial levels. Should
these tests prove successful, the Chinese will be granted free
use of the process, although Dextec will retain the patents and
marketing rights.
References
1. Ore to Metal Without Smelting? Mining Journal, 292(7482):
30, January 12, 1979.
4.3 CATHODE SHEET STRIPPING1
Outokumpu Oy has developed a mechanized method of stripping
starter sheets for electrolytic refining from copper mother
blanks (starter blanks). Mechanized stripping operations were
begun at Outokumpu"s Pori, Finland, refinery in the spring of
1974 in an effort to reduce labor and operating costs.
The system consists of three conveyors and a stripping
2
machine and occupies about 85 m . The stripping machine is
capable of stripping one pair of sheets in 8 seconds and can
handle 900 starter sheets per hour at peak capacity. Although
designed for copper, the machine could easily be modified for
stripping zinc and nickel.
As shown in Figure 20, mother blanks are fed into Conveyor
1 and transferred to the washing unit. The blanks then connect
with Conveyor 2 and pass in vertical single file through the
stripping station.
A mother blank is locked into position during stripping so
that blades may open the upper edges of the sheet and pneumat-
ically operated jaws on both sides of the sheet may grasp the
opened upper edges. Copper sheets are peeled from the mother
blanks and placed on adjustable hydraulic tables. Stripped
mother blanks are then coated with a chemical and collected on
Conveyor 3, where they are positioned with the required spacing
for the stripper cells.
68
-------
CTi
GENERAL LAYOUT OF STRIPPING MACHINE
FOR COPPER CATHODE STARTER SHEETS
9.9m
STARTING
SHEET PILES.
STARTING (3
SHEET
PILES
MOTHER
BLANKS IN
MOTHER
BLANKS OUT
VERTICAL SECTION OF STRIPPING SECTION
T
A. MOTHER BLANK
B. STARTING SHEET
1. STRIPPING CONVEYOR
2. STRIPPING BLADE
3. GRIPPING JAW
4. ADJUSTABLE HYDRAULIC TABLE
5. PALLET
6. PIE STRAIGHTEMER
7. GUIDE BEAM
8. CONTROL DESK
1. CONVEYOR 1
2. CONVEYOR 2
3. CONVEYOR 3
4. WASHING UNIT
5. STRIPPING STATION
6. MAINTENANCE TRACK
7. SPRAY COATING UNIT
8. CONTROL DESK
Figure 20. Cathode sheet stripping.
1
-------
References
1. Automatic Starter Sheet Stripping at Outokumpu. Engineering
and Mining Journal, p. 106, June 1975.
70
-------
SECTION 5
AIR POLLUTION CONTROL PROCESSES
5.1 DOWA BASIC ALUMINUM SULFATE PROCESS
Dowa Mining of Japan has developed a process for weak-
stream S02 removal that utilizes a basic aluminum sulfate solu-
tion as the absorbent. Four Japanese nonferrous facilities use
this process to treat acid plant tail gases; its application at
Dowa's double-contact sulfuric acid plant in Okayama is de-
1-4
scribed.
At 80°C, tail gas from the acid plant contains 600 ppm S09
3
and 50 ppm SO3 and is produced at a rate of 140,000 Nm /h. As
shown in Figure 21 the gas is routed to the lower end of an
absorbing tower, where it is contacted countercurrently by a
downward spray of basic aluminum sulfate absorbent, which
absorbs the S0?. The solution then flows to an oxidizing tower,
where fine air bubbles are injected to oxidize the sulfite to
sulfate. After a brief retention time, most of the absorbent is
recycled to the absorption tower, but a small portion is sent to
a tank for neutralization with limestone. The neutralized
slurry overflows to a thickener, where gypsum crystals precipi-
tate; thickener underflow is routed to a centrifuge where gypsum
is filtered, washed, and discharged. Thickener overflow, fil-
1-3
trate, and wash water are recycled to the absorption tower.
Treated gas passes through a mist eliminator and then exits
through a stack. Concentration of S02 is reportedly 10 to 15
ppm.
Absorption of S02 is increased with higher aluminum concen-
trations and basicity and lower temperatures. In commercial
operation, basicity should be lower than 40 percent to prevent
71
-------
GAS EXIT
ELIMIN,
WASHING
II HATER
' CENTRIFUGE
Figure 21. Dowa basic aluminum sulfate process.
1
72
-------
aluminum precipitation and plugging in the absorption tower.
Basicity is defined as the degree of neutralization of aluminum
sulfate solution and may be quantified as follows:
A12(SO4)3 Basicity = 0%
A12(S04)3*A12°3 Basicity = 50%
Al(OH)3 Basicity = 100%
Because of incomplete washing, some aluminum is lost — about 0.5
kg per metric ton of gypsum.
Direct operating costs (1976) are estimated to be $3.00 per
10,000 -Nm of input gas; this figure excludes labor costs.
Typical operating requirements are 0.81 Mg CaCO /h, 15 kg
A12(S04)3 (A1203 8%)/h, 7 H20/h, and 65 kwh electricity/10,000
tail gas. 1-3
References
1. PEDCo Environmental, Inc. Pollution Control in the Jap-
anese Primary Nonferrous Metals Industry. Prepared for the
U.S. Environmental Protection Agency under Contract No.
68-02-1375, Task No. 36. Cincinnati, Ohio, March 1978.
2. The Dowa Basic Aluminum Sulfate and Gypsum Process. Dowa
Mining Co., Ltd., Dowa Engineering Co., Ltd.
3. Flue Gas Desulfurization Process Using Basic Aluminum Sul-
phate as the Absorbent (Dowa Process). Dowa Mining Co.,
Ltd., Dow Engineering Co., Ltd.
4. Campbell, I.E. The State of the Art of Flue Gas Desulfuri-
zation Technology. Presented at the 1978 International
Mining Show of the American Mining Congress, Las Vegas,
Nevada, October 9-12, 1978.
5.2 MITSUBISHI LIME/LIMESTONE PROCESS
Mitsubishi Metals Company of Japan has developed a system
of SO control utilizing lime as a nonregenerative absorbent.
The development stage lasted about 20 years, and the process is
now in operation at Mitsubishi's Onahama smelter.
1-5
73
-------
The reverberatory furnace gas stream contains an average of
2.4 percent S02. At full capacity, the furnaces discharge a gas
stream of about 3120 Nm3/min. It is reported that the FGD
system is capable of lowering the SO- content of the gas stream
to below 100 ppm before it is discharged.
Flue gas from the Onahama reverberatory furnace passes
through waste heat boilers and Cottrells, as shown in Figure 22.
It then enters a horizontal wash tower. The sprays and internal
baffles in this cylindrical tower lower the gas temperature to
60°C. Gas temperature is reduced further in the five heat
exchangers immediately downstream of the wash tower. ' '
At this point, the gas stream can be split, with part going
to an MgO system and part to the lime system. Currently, the
poor demand for sulfuric acid dictates full flow to the lime/
limestone system, where SO- is removed by two absorber towers set
in series. A lime slurry is sprayed from the top of the towers
and the absorbent flow rate through the first absorber is about
half that in the second absorber. Approximately half the SO^
is absorbed in each tower.
Because some absorbent is entrained in the gas stream, a
mist eliminator is located downstream of the towers. The gas
stream passes through a blower and an electrostatic precipitator
before it is emitted through the stack.
The pH of the calcium sulfite slurry at the bottom of the
first absorber tower is between 4 and 5; the slurry is first
pumped to a pH adjuster, where the pH is lowered to between 3 and
4 with sulfuric acid. It then passes (at a rate of about 200
m /h) to three oxidation towers, where it is mixed with com-
pressed air. Most of the gypsum slurry flows from the oxidizing
towers to a thickener, but about 20 percent goes to the lime tank
to supply seed crystals. The crystals later enter the absorption
system in the second tower to help prevent deposition on absorber
surfaces.
74
-------
-J
Ul
FLUE WASHING
GAS I T(>««
ZOO Nm /min
WATER LEAKS
1.0 Mq/h IN ATOMIZER
WATER FROM
SPILL GAS TCA WATER FRO»
2600
BALI
MILL
^LIQUID/
CYCLONE
-K
)ER
«
•«
I
COOLING
TOWER
Figure 22. Mitsubishi lime/limestone process used at Onahama,
-------
Underflow from the thickener is a 25 percent slurry. This
slurry is routed to a bank of 14 centrifuges where solid gypsum
is separated at a rate of 450 Mg/day. The H2O content of the
slurry is 8 to 10 percent. The liquid is recycled to the
thickener. Thickener overflow is routed to the slaking system,
where it is used for diluting the slaked lime. Excess water
from the thickener overflow is neutralized and discharged.
The slaked lime passes to a ball mill and liquid cyclone.
The heavy material from the liquid cyclone is recycled to the
ball mill, and the light material (calcium hydroxide or lime
milk) is sent to the milk holder and dilution tank. From the
dilution tank, the milk goes to the lime feed tank and back to
the second absorber.
Versions of this process are in use on coal-fired gen-
erators in the United States, but these domestic applications do
not make use of compressed air in the oxidation phase.
References
1. Weisenberg, I.J., et al. Appendixes: S02 Control for the
Primary Copper Smelter Reverberatory Furnace. Prepared for
IERL, U.S. Environmental Protection Agency, Cincinnati,
Ohio, under EPA Contract No. 68-03-2398. August 1977.
2. PEDCo Environmental, Inc. Pollution Control in the Jap-
anese Primary Nonferrous Metals Industry (draft). Prepared
for U.S. Environmental Protection Agency under Contract No.
68-02-1375. Cincinnati, Ohio, March 1978.
3. Matthews, J.C., et al. S02 Control Processes for Non-
ferrous Smelters. EPA 600/2-76-008, January 1976.
4. Scrubbers Remove S02 from Copper Smelter Off Gases. Wet
Scrubber Newsletter, 56:4, February 28, 1979.
5. Arthur G. McKee and Company. Systems Study for Control of
Emissions Primary Nonferrous Smelting Industry, Vol. I,
Vol. II, and Vol. III. PH 86-65-85, San Francisco, Cali-
fornia, June 1969.
6. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper Industry. Prepared for the U.S. Environmental
Protection Agency under Contract No. 68-02-2608. Radian
Corporation, Austin, Texas, August 1978.
76
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7. Onahama Smelting and Refining Co. Double Expansion of
Onahama Smelter and Refinery. Presented at the 103rd AIME
Annual Meeting, Dallas, Texas, February 1974.
Bibliography
Pacific Environmental Services. Pollution Control in Japan at
the Onahama and Naoshima Primary Copper Smelters With
Particular Emphasis on Reverberatory Furnaces. Prepared
for the U.S. Environmental Protection Agency under Contract
No. 68-03-2398. 1977.
5.3 MAGNESIUM OXIDE PROCESS1'2
The magnesium oxide SO2 concentration system was developed
by Mitsubishi Metals in conjunction with Tsukishima Kikai
Company over a period of about 3 years (1969 to 1971). Full-
scale operation at Mitsubishi's Onahama smelter began in 1973.
Initial treatment of the furnace off-gas is the same as for
the lime/limestone process, with the two systems sharing the
waste heat boilers, Cottrells, horizontal wash tower, and five
heat exchangers. During this treatment gas temperatures are
reduced to about 45°C. The gases routed to the MgO system are
then passed through a turbulent contact absorber (TCA), where
they are contacted by a countercurrent flow of magnesium hy-
droxide slurry. Magnesium sulfite and some magnesium sulfate
are formed in this stage.
The absorbent containing the magnesium sulfate is passed to
a liquid cyclone, where it is rough cut to a slurry of 40 to 50
percent solids. Overflow from the cyclone is filtered and
combined with the recovered MgO; the high-solids underflow is
treated in centrifuges for further removal of liquid. Collected
liquid is filtered and recycled.
The centrifuged magnesium sulfite is sent to two indirect
rotary steam-heated dryers, and is then combined with coke and
treated in a rotary calciner. The calciner breaks down magne-
sium sulfite to S00 and MgO. The S09 is routed to an acid
2 ^
plant. The MgO is sent to a slaker, where it is mixed with
77
-------
overflow from the centrifuge and cyclone systems; the resulting
magnesium hydroxide is then either further processed by a tube or
ball mill or returned directly to a storage tank before it is
recycled. The storage tank also receives makeup magnesium
hydroxide.
In 1977, the capital cost of the 88,000-Nm /h system at
Onahama was estimated to be $2.5 million. A breakdown of annual
operating costs indicates the following: MgO, $500,000; bunker C
fuel, $400,000;-coke, $10,000; steam, $200,000; water, $100,000;
and labor, $400,000. Total operating cost is probably around $2
million.
References
1. Weisenberg, 'I.J., G.M. Meisel, and J.O. Burckle. Magnesium
Oxide S02 Concentration System for Reverberatory Furnace
Off-Gases at the Onahama Copper Smelter. TMS Paper A-79-2.
The Metallurgical Society of AIME, Warrendale, Pennsylvania,
1979.
2. Pacific Environmental Services. Pollution Control in Japan
at, the OnahamaTand Naoshima Primary Copper Smelters With
Particular Emphasis on Reverberatory Furnaces. Prepared for
the U?S. Environmental Protection Agency under Contract No.
68-.03-2398. 1977.
5.4 BOLIDEN COLD SEAWATER PROCESS
Since 1970, Boliden a.-b. of Sweden has been utilizing cold
seawater to scrub SO? from the combined gas stream at its Ronnskar
works. The combined stream is a blend of gases from an electric
copper furnace, multiple hearth furnaces, converters, and a lead
smelter. Two commercial-scale plants are in operation at Ronnskar;
the older plant (1970) has an SO0 capacity of 5 Mg/h, and the new
TO
plant (1976) has an SO2 capacity of 15 Mg/h. '
The combined gas stream has a flow rate of 60,000 to 70,000
m /h at about 150°C; S02 concentration ranges from 0.5 to 5
percent and averages 3 percent. As shown in Figure 23, after
78
-------
SEA
WATER,
ABSORPTION
TOWER
STEAM
10 HEAT STEAM
EXCHANGERS
PRODUCT 0°-10°C
STORAGE —
LIQUID S09
REFRIGERATION
UNIT
DESORPTION
TOWER
n2JU4
DRYING
TOWER
Figure 23. Boliden cold seawater process.'
79
-------
the gas has been cleaned in wet electrostatic precipitators and
treated for mercury removal, it is routed to the absorption tower
where the SO2 is absorbed in the cold (4° to 11°C) seawater. At
this operating temperature, the solubility of SO2 in water is
about 12 g/liter. The treated gas contains 0.1 percent S02 or
less.1'4
Seawater from the absorber is heated to 60 °C by injection of
low-pressure steam; it then enters a vacuum desorption tower that
operates at a pressure of 20 kPa absolute. The S02 is released
in the desorption tower and is cooled to between 12° and 15 °C by
a slipstream from the absorption tower. The stripped seawater is
2 4
cooled and discharged to the sea. '
After cooling, the SO- gas is dried by contact with sulfuric
acid, and the pressure is raised slightly to avoid freezing in
the refrigeration unit. The gas is further cooled and compressed;
this results in its liquef ication. The liquid SO0 is stored at a
4 *•
temperature of between 0° and 10°C.
This process offers advantages in that a wide range of SO-
concentrations can be treated with high removal efficiencies and
without scaling problems. The availability of cold seawater is
extremely limited in geographic range, however, and discharged
4
water may contain high levels of sulfite.
References
1. Weisenberg, I.J., et al. Appendixes: S02 Control for the
Primary Copper Smelter Reverberatory Furnace. EPA Contract
No. 68-03-2398, IERL, U.S. Environmental Protection Agency,
Cincinnati, Ohio, August 1977.
2. New Gas Treatment and Dust Precipitation Systems. World
Mining, 31(5) :48, May 1978.
3. Devitt, T.W. Visit to Boliden Smelter, Skellef teahamn,
Sweden. PEDCo Environmental, Inc., Cincinnati, Ohio, June
6, 1977.
4. Coleman, R.T. , Jr. Emerging Technology in the Primary
Copper Industry (draft). Prepared for U.S. Environmental
Protection Agency under Contract No. 68-02-2608. Radian
Corporation, Austin, Texas, August 31, 1978.
80
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5.5 FLAKT-BOLIDEN CITRATE PROCESS
Boliden a.-b. of Sweden, in conjunction with the Norwegian
Technical Institute (SINTEF) and Svenska Flaktfabriken (Flakt),
has developed a variation of the cold seawater S02 scrubbing
process that can be used where cold seawater is not available.
The success of the process results from the discovery that the
addition of weak bases having good buffering properties to water
increases the SO2 absorption capacity. In the pilot-scale plant
at Boliden's Ronnskar works, a combination of citric acid and
sodium citrate is used because it has a low vapor pressure and
is chemically stable. The absorption section of the process is
similar to the Bureau of Mines citrate process.
The pilot plant, which began operation in September 1976,
cleans approximately 5000 Nm /h of off-gases from various smelt-
ing processes. The SO_ concentration varies from 0.2 to 6 per-
cent; however, typical operation is 3 percent S0_. For a 3
percent SO9 off-gas stream, the steam consumption is 2 kg steam/kg
£t
S0_ removed, and the oxidation rate is 0.1 percent of the ab-
£t i
sorbed SO,,. The gas stream is cleaned of particulate, cooled by
direct water injection, and passed through a mist eliminator
before it enters the absorption tower. The gas temperature at
the absorber inlet is between 45° and 65°C. The absorber
operates at atmospheric pressure and gas/absorbent contact is
countercurrent; absorption efficiency is between 90 and 99
percent. The cleaned gas is passed through a mist eliminator
1-3 4
and exits through a stack to the atmosphere. '
As shown in Figure 24, the absorbent is pumped from the
bottom of the absorber tower to the top of the stripping tower,
where stripping is accomplished by countercurrent steam treat-
ment. Low-pressure steam is introduced at the bottom of the
stripping tower, and the tower is kept under vacuum. This
treatment produces a mixture of S02 and water vapor. The
absorbent is recirculated to the absorber tower. '
81
-------
CLEAN GAS
TO MIST
ELIMINATOR
CONDENSER
ABSORBER
GAS FROM
PARTICULATE
REMOVAL
MAKEUP
STEAM
SODIUM
SULFATE
REGENERATION
ALTERNATIVE
PLANTS
Figure 24. Flakt-Boliden citrate process.'
LIQUID
S02
.SULFURIC
ACID
ELEMENTAL
SULFUR
82
-------
The mixture of S02 and water from the stripping column is
cooled in a condenser, which separates most of the water from the
mixture. The condensate, which contains only a small quantity of
S02, is returned to the stripping tower. The concentrated SO
stream can be routed directly to a Glaus plant for production of
elemental sulfur or to a contact plant for sulfuric acid pro-
duction, or it can be condensed by refrigeration to liquid SO .
£»
If liquid S02 is produced, the water remaining in the gas first
must be removed by contact with concentrated sulfuric acid in a
1 4
small packed tower. '
After drying, the S0_ gas still contains some inert gas.
The dried SO2 gas is conveyed to a Freon cooler for condensation,
and the liquid S0~ is stored under pressure. The tail gas is re-
14
turned to the absorber. '
Some sodium sulfate may form in the absorbent and must be
removed. Sulfate levels can be kept acceptable by regeneration
of an intermittent or continuous bleed-off from the stripper of
about 1 percent of the stripped absorbent. Seed crystals and a
cooling unit are used to recover sodium citrate and remove sodium
2
sulfate.
Application to a 1.5 percent SO2 gas stream of 157,700 Nm /h
would require approximately 325 kW and about 30,000 kg/h of low-
grade steam.
References
1. Weisenberg, I.J., et al. Appendixes: SO Control for the
Primary Copper Smelter Reverberatory Furnace. EPA Contract
No. 68-03-2398, IERL, U.S. Environmental Protection Agency,
Cincinnati, Ohio, August 1977.
2. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper Industry. Prepared for the U.S. Environmental Pro-
tection Agency under Contract No. 68-02-2608. Radian
Corporation, Austin Texas, August 1978.
3. Matthews, J.C., et al. S02 Control Processes for Nonferrous
Smelters. EPA 600/2-76-008, January 1976.
4. The Flakt-Boliden Process for S02 Recovery. AB Svenska
Flaktfabriken, 1977.
83
-------
5.6 SHOWA DENKO PROCESS
Showa Denko K.K. of Japan has developed a double alkali
process for S0? removal, which can be applied to acid plant tail
gases and other weak SO? streams. At the time of development,
declining demand for sodium sulfite and increasing marketability
of gypsum in Japan combined to make this process economically
advantageous in comparison with sodium processes. The process is
in use at Nippon Mining's copper-lead smelter at Saganoseki,
Japan.
At Saganoseki, exhaust gases from the single-contact sul-
furic acid plants, which contain 1500 to 2000 ppm S0«, are
1
treated in the Showa Denko scrubber.
actions involved in this process are:
Principal chemical re-
(1)
(2)
(3)
(4)
NaS03 4
2NaHSO
2NaHS0
CaCO
CaS0
CaS0
1/2
1/2
1/2
2NaHSO
2CaSO
4
(5) 2NaOH + S02 -»• Na2SO3 + H2O
The weak S02 stream enters a venturi and flows upward into a
separator. The absorbent, consisting of Na?SO_, -NaHSO_, and
NaHSO., is injected into the venturi throat and is entrained in
the upward gas flow. When the gas flow rate declines in the
separator section, the absorbent drops out of suspension.
Following absorption of the S02 (Reaction 1) and deentrain-
ment, part of the absorbent solution is neutralized with lime-
stone or lime (Reaction 2) and forms a calcium sulfate slurry.
This slurry is filtered, and the filter cake is reslurried.
Sulfuric acid is added to reduce the pH of the filter cake slurry
(Reaction 4) . The slurry is then oxidized with air to form
calcium sulfate, which crystallizes as gypsum. The gypsum
crystals are dehydrated in a centrifuge.
84
-------
The addition of sulfuric acid to the filter cake slurry
promotes gypsum precipitation and minimizes loss of sodium. The
pH of the absorbent is maintained with caustic and filtrate (Re-
action 5) . (Minimizing sodium loss lessens the need for caus-
tic.) Concentration of SO2 in gases leaving the Showa Denko
system is between 50 and 100 ppm.^
Versions of this process are in use in the United States,
but not for treatment of weak streams at nonferrous smelters.2
References
1. PEDCo Environmental, Inc. Pollution Control in the Japanese
Primary Nonferrous Metals Industry. Prepared for the U.S.
Environmental Protection Agency under Contract No. 68-02-
1375. Cincinnati, Ohio, March 1978.
2. Matthews, J.C., et al. SC>2 Control Processes for Nonferrous
Smelters. EPA 600/2-76-008, January 1976.
5.7 ZINC OXIDE PROCESS
Mitsui Mining and Smelting Company, Ltd., of Japan has
developed a zinc oxide process for S09 removal from acid plant
12
tail gas at its Hikoshima zinc smelter. '
Tail gas from the single-contact acid plants contains 0.15
to 0.18 percent SO2- This gas is routed to an absorber tower
where it contacts a countercurrent stream of zinc oxide and
water. Zinc calcine or crude zinc oxide is slurried with acid
plant drainage and some other industrial effluents to make up the
absorbent solution.
Over 97 percent of the SO2 is absorbed; the absorbent is
then sent to a tank for decomposition with either heat or sulfu-
ric acid. The Hikoshima smelter is reportedly using acid decom-
position at present; temperature is 85° to 90°C and zinc sulfate
and water are produced. The SO2 content at discharge is 30 ppm
or less and the gas is cleaned further in a wet ESP. '
A gas volume of 480 Nm3/min is treated. Figure 25 is a
simplified flow diagram.
85
-------
EXHAUST
SOo-LADEN
TAIL GAS
FROM ACID
PLANT
S02 TO ACID PLANT
ZINC SULFATE
SOLUTION
ABSORBING
TOWER
>
ZnO AND H20
11 i
TANK 1
Figure 25. Zinc oxide process.
1
86
-------
This process is advantageous in that the zinc sulfate
solution may be treated in electrolytic cells for zinc recovery.
A weak stream of sulfuric acid or soluble sulfates is produced,
however, and therefore constitutes a potential source of water
pollution. '
References
1. Mitsui Mining and Smelting Co., Ltd. zinc Oxide-Based SO?
Scrubbing System.
2. PEDCo Environmental, Inc. Pollution Control in the Japanese
Primary Nonferrous Metals Industry (draft). Prepared for
the U.S. Environmental Protection Agency under Contract No.
68-02-1375. Cincinnati, Ohio, March 1978.
5.8 FLASH AGGLOMERATION
The smelting of lead-containing scrap materials such as
storage batteries generates a dust rich in lead oxides; however,
this dust also consists of oxides of other metals together with
chlorides, sulfides, and sulfates. At most secondary lead
smelters, this flue dust is directly recycled to the blast
furnace. This practice leads to dust losses; degradation of the
dust also occurs because the lead chloride present in the dust is
more volatile than the other material in the furnace, and recir-
culation of the dust tends to "distill" the chloride. Agglomera-
tion of the captured dust and the addition of fluxes allow the
melting point of the dust to be adjusted so it melts lower in the
furnace shaft; the chloride is then fixed in the slag. Paul
Bergs0e and Son A/S of Denmark have devised a system of accom-
plishing this agglomeration in an efficient manner.
As indicated in Figure 26, dust from fabric filter hoppers
is fed via screw conveyor to an agglomeration furnace. A flux
such as sodium carbonate or borax may be added to the dust while
it is on the conveyor. Upon landing on a sloped hearth within
the furnace, the dust is melted by an impinging flame. The
molten material flows down the hearth, out a tap hole, and into a
87
-------
BURNER
LIMESTONE
OR FeO
ADDITION
PROCESS VENT
FABRIC FILTER
DUST HOPPER
AGGLOMERATION
FURNACE
MOLTEN
DUST
COOLING/TRANSPORTATION
CRUCIBLE
Figure 26. Flash agglomeration furnace.
1
88
-------
waiting cooling/transportation crucible. The solidified material
is later broken into suitably sized chunks, mixed with coke and
flux, and charged to the blast furnace along with batteries or
scrap.
Flue dust agglomeration decreases the volume of material
being returned to the blast furnace by about 80 percent; thus
additional batteries or scrap may be charged. Since the recycle
stream is reduced from 10 percent of the charge to about 2 per-
cent, an increased furnace throughput of 8 percent is theoret-
ically possible. For a 23,000-Mg/yr smelter, realization of one
half this increase in throughput has been calculated to constitute
a $66,120 per year (1977) increase in gross revenues. The cost
of the flash agglomeration has been estimated to be less than
$100,000 (1977) installed. In addition, both stack and fugitive
dust emissions are reduced.
The agglomeration furnace is oil fired and consumes 24.6
liters of oil per hour. '
References
1. Coleman, R.T. Emerging Technology in the Secondary Lead
Industry. Prepared for the U.S. Environmental Protection
Agency under Contract No. 68-02-2608. Radian Corporation,
Austin, Texas, 1978.
2. Mackey, T.S., et al. Flash Agglomeration of Flue Dust.
J. Metals, 29(11):12-15, November 1977.
3. U.S. Patent Office. Method for Treating Flue Dust Con-
taining Lead. Patent No. 4,013,456 issued to Svend Bergsjzfe,
March 22, 1977.
4.- Mackey, T.S. Flash Agglomeration of Flue Dust. Australian
Mining, 69(6):52-54, June 1977.
5. Mackey, T.S., et al. Smelting of Unbroken Batteries.
American Institute of Mining, Metallurgical, and Petroleum
Engineers, 106th Annual Meeting, Atlanta, Georgia, March 9,
1977.
6. Coleman, R.T., et al. Trip Report to Paul Bergs^e and Son
A/S, Boliden Aktiebolag, and Outo kumpu Oy. Performed for
the U.S. Environmental Protection Agency under Contract No.
68-02-2608. Radian Corporation, Austin, Texas, November 28,
1977.
89
-------
5.9, BOLIDEN DRY SELENIUM FILTER
Boliden a.-b. of Sweden has developed a process that uses a
selenium-impregnated filter to remove mercury as mercuric sel-
1-4
enide (HgSe) from roaster gases at nonferrous smelters.
The gases are cleaned of particulate and dried prior to
entering an absorption reactor tower. The tower is closed at the
top so that gases entering the bottom diffuse radially through a
cylindrical filter consisting of an inert porous material that
1-4
has been soaked in selenious acid. Red amorphous selenium is
precipitated when the gases pass through the filter, as shown in
the following reaction:
H2Se03 + H20 + 2S02 -* Se + 2H2SO4
The filter medium is 0.5 m thick and the outer 0.3 m is impreg-
nated with the precipitated selenium. The total filter area of
2
about 100 m is divided into two parallel units. The gas passes
through the filter at a velocity of about 0.17 m/s, resulting in
a contact time of 1 to 2 seconds. Condensation of moisture on
the filter must be avoided to prevent deactivating the selenium.
The filter element can continue to absorb mercury until its
content reaches 10 to 15 percent. At that time the filter is
treated to recover mercury and regenerate the selenium. ~
This type of selenium filter has been in operation for
several years at Boliden"s Ronnskar works and has effectively
removed 90 percent of the mercury from a gas stream of 60,000
3 3
Nm /h flow rate containing 4 to 5 percent S0_ and 0.85 mg/m
mercury. Occasionally, collection efficiencies of up to 99.9
percent have been achieved. Particulate mercury has been col-
lected with approximately the same efficiency.
References
1. Sundstrom, 0. Mercury in Sulphuric Acid: Boliden Processes
Can Control Hg Levels During or After Manufacture. Sulphur
No. 116, pp. 37-43, The British Sulfur Corp., January-
February 1975.
90
-------
2. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
3.. Habashi, F. Metallurgical Plants: How Mercury Pollution Is
Abated. Environmental Science and Technology, 12(13):1372-
1376, December 1978.
4. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper Industry (draft). Prepared for the U.S. Environ-
mental Protection Agency under Contract No. 68-02-2608.
Radian Corporation, Austin, Texas, August 31, 1978.
5.10 BOLIDEN WET SELENIUM SCRUBBING PROCESS
Boliden a.-b. has been using the wet selenium scrubbing
technique since 1972 to remove mercury from roaster gases at its
Helsingborg, Sweden, zinc facilities. The process utilizes
amorphous solid selenium in suspension in a wet scrubber solu-
tion; the solution contains between 20 and 40 percent sulfuric
acid. The acid concentration must be kept within these limits
because complex and highly soluble selenium-sulfur compounds are
formed at low H2SO. concentrations. At high concentrations, the
acid's oxidizing power becomes noticeable and selenium dioxide or
1-4
selenite is formed.
When the sulfide ore being roasted contains sufficient
selenium, it may not be necessary to add more selenium to the
scrubbing system; however, the activated selenium slurry must be
recirculated at an adequate rate.
Treatment of roaster gases containing 6 mg/m mercury
reportedly results in reduction of mercury concentrations to
about 0.1 1
References
1. Sundstrom, O. Mercury in Sulphuric Acid: Boliden Processes
Can Control Hg Levels During or After Manufacture. Sulphur
No. 116, pp. 37-43, The British Sulphur Corp., January-
February 1975.
2. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
91
about 0.1 to 0.5 mg/m .
-------
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
3. Habashi, F. Metallurgical Plants: How Mercury Pollution Is
Abated. Environmental Science and Technology, 12(13):1372-
1376, December 1978.
4. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper Industry. Prepared for the U.S. Environmental
Protection Agency under Contract No. 68-02-2608. Radian
Corporation, Austin, Texas, August 1978.
5.11 BOLIDEN ACTIVATED CARBON PROCESS
The activated carbon process is used at Boliden's Ronnskar
operation in Sweden. The process is based on the ability of
activated carbon to adsorb 10 to 12 percent of its own weight in
mercury vapor. Gas streams containing low levels of mercury (0.1
ig/m'
1-4
to 0.2 mg/m ) are particularly well suited to application of this
process,
The Boliden unit consists of three towers connected in
2
parallel, each with a capacity of 4.0,000 Nm /h of gas and con-
taining 23 m of activated carbon. During normal operation,
mercury in the gas stream enters the filter at a rate of 10 g/h
and leaves at 0.9 g/h, an efficiency of 90 percent. The gas
input normally contains 3 to 6 percent SO-, but the carbon is
activated initially with a 100 percent SO- stream. For proper
operation, the gas must be thoroughly dried and the temperature
1-4
must not exceed 50°C.
2
Initial cost of the Boliden unit in 1971 was about $210,000.
References
1. Sundstrom, 0. Mercury in Sulphuric Acid: Boliden Processes
Can Control Hg Levels During or After Manufacture. Sulphur
No. 116, pp. 37-43, The British Sulphur Corp. January-
February 1975.
2. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
92
-------
3. Habashi, F. Metallurgical Plants: How Mercury Pollution Is
Abated. Environmental Science and Technology 12(13)•1372-
1376, December 1978.
4. Coleman, R.T., Jr. Emerging Technology in the Primary
Copper Industry. Prepared for the U.S. Environmental
Protection Agency under Contract No. 68-02-2608. Radian
Corporation, Austin, Texas, August 1978.
5.12 MERCURIC CHLORIDE SCRUBBING PROCESS1
Boliden a.-b. and Det Norske Zink Kompani have constructed a
full-scale chloride scrubbing facility for removal of mercury
from roaster gases at a plant in Eitrheim, Odda, Norway. The
chloride scrubber is part of the gas cleaning system in an acid
plant treating zinc roaster off-gases.
As illustrated in Figure 27, a chloride scrubber consists of
a tower in which the chloride solution is sprayed downward,
meeting a countercurrent flow of roaster gases; the tower is
equipped with a mist eliminator, sludge separator, and recycle
pump. Part of the wash solution is directly recycled and part is
first treated in the sludge separator. Underflow is treated for
mercury recovery.
Over a 2-year period, the process helped produce sulfuric
acid with a mercury content of 0.5 ppm from the roasting of zinc
concentrates with mercury levels of up to 150 ppm.
Initial cost of the scrubber, which was built in 1972-73,
was about $310,000; operating costs have been estimated at less
than 0.10/Mg of monohydrate. Power requirements for pumps and
fans are 8 kWh/Mg of sulfuric acid.
Reference
1. Sundstrom, 0. Mercury in Sulphuric Acid: Boliden Processes
Can Control Hg Levels During or After Manufacture. Sulphur
No. 116:37-43, The British Sulphur Corp., January-February
1975.
93
-------
CLEANED GASES
TO ACID PLANT
ATvAA
SLUDGE
SEPARATOR
•MIST ELIMINATOR
SCRUBBER
-ROASTER
OFF-GASES
TO MERCURY
RECOVERY
HgCT2 SOLUTION
RECYCLE PUMP
Figure 27. Mercuric chloride scrubbing process.
1
94
-------
5.13 OUTOKUMPU SULFURIC ACID SCRUBBING PROCESS
The Outokumpu process was developed to remove mercury from
S02-laden roaster gases at the Kokkola electrolytic zinc plant;
the process began operation in 1970. The same process has been
used in Japan at the Onahama plant of Toho Zinc Co. and the Miike
smelter of Mitsui Mining and Smelting Co., Ltd.1'2
At Kokkola, the gas from the fluidized bed roaster contains
9 to 11 percent SO2 at 950°C. After the gas passes through a
waste heat boiler and a series of electrostatic precipitators and
cyclones, the temperature drops to about 350°C. Gas leaving the
electrostatic precipitators contains about 40 to 80 mg Hg/m3 and,
as shown in Figure 28, enters a sulfatizer, which consists of a
brick-lined tower packed with ceramic shapes. In the sulfatizer,
the gas contacts a countercurrent flow of concentrated (over 80
percent) sulfuric acid. Selenium and mercury are scrubbed from
1 2
the gas along with any zinc or iron present. '
The temperature of the gas is controlled by an external heat
exchanger and by the rate of acid circulation to prevent dilution
of the sulfuric acid by moisture in the gas. Thus, the vapor
pressure of the water in the acid corresponds to the water
partial pressure of the gas. At Kokkola, the acid inlet and
1 2
outlet temperatures are 40° and 180°C. '
The metal containing acid flows from the tower bottom to a
strong tank, from which a portion is pumped through a heat ex-
changer and returned to the tower. Another portion is routed to
, . 1,2
a clarifier for removal of the mercury, selenium, and zinc.
Gas exits the sulfatizer at 180°C with about 0.2 mg ele-
mental mercury/Nm3 and is then washed with weak (about 30 per-
cent) sulfuric acid in the normal scrubber of the sulfuric acid
plant. The temperature is decreased to about 60° to 70°C, and
the chloride content is reduced to a level suitable for sulfuric
acid production. The gas is tested in an electrostatic pre-
cipitator before being routed to the acid plant. '
95
-------
HEAT
EXCHANGER HOT
ELECTROSTATIC
PRECIPITATOR
GAS
FROM ELECTROSTATIC
PREC1PITATORS BLOWER
ov
Hg-FREE
"*" S02 GAS
Hg-FREE
DILUTE H2S04
MERCURY
UASTERUATER
TO FURNACE
RESIDUE TO
Se RECOVERY
Figure 28. Outokumpu sulfun'c acid scrubbing process.'
-------
Concentrated acid carryover, residual mercury, and selenium
are removed in this weak scrub. These materials are separated
from the weak acid in the clarifier, and the underflow is com-
bined with the underflow from the sulfatizer's clarifier. The
clarifier overflow is used to make up concentrated acid loss and
is returned to the sulfatizer.1'2
Precipitates from underflow of the two clarifiers are
washed with water to dissolve the sulfuric acid, iron, and zinc
salts; the residue contains about 50 percent insoluble mercury
and 10 percent selenium compounds. Washwater is fed to the
roasting furnace to ensure complete metal recovery, and the
residue is mixed with a fixed proportion of lime and fed to a
kiln, where it is heated to 650° to 700°C. The mercury compounds
decompose, and mercury vaporizes, exits the kiln, and is re-
covered in a condenser. Sulfur and selenium combine with the
lime to form stable compounds; this material serves as feedstock
for production of metallic selenium. Gas from the kiln, which
contains mercury not recovered in the condenser, is combined with
1 2
roaster gases ahead of the sulfatizer. '
The Outokumpu process is working successfully with concen-
trates containing up to 250 ppm mercury. The cleaned gas stream
entering the acid plant usually contains between 0.1 and 0.2 mg
mercury/Nm3, and the product acid contains less than 1 mg mer-
cury/liter.
One disadvantage of the Outokumpu process is that the high
temperatures involved can cause corrosion. Another disadvantage
is that the dust removed by the hot electrostatic precipitator
preceding the sulfatizer can scatter and react with the acid to
produce other sulfates, which may deposit within the pipes when
the acid is cooled.
The installed cost of the sulfatizer and associated equip-
ment is between $13 and $16 per annual megagram of sulfuric acid
capacity. Retrofit of a facility such as this is more expensive.
The most significant operating cost appears to be maintenance,
which is about 10 percent of the yearly capital costs.
97
-------
References
1. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, 1976.
2. Kangas, J. , et al. Smelter Gwses Yield Mercury. Chemical
Engineering, pp. 55-57, September 6, 1971.
Bibliography
Outokumpu Oy, Minimization of Dust and Gas Emissions in the
Nonferrous Metal Industry. Helsinki, Finland, 1978.
Flak's, N. Seminar on the Control of Emissions From the Non-
ferrous Metallurgical Industries. Economic Commission for
Europe, The United Nations. Dubrovnik, Yugoslavia, November
19-24, 1973.
5.14 CENTRO NACIONAL DE INVESTIGACIONES PROCESS1
Centre Nacional de Investigaciones (CENIM) in Spain has
developed a process for the removal of mercury from the stack
gases at its Almaden mercury plant.
Mercury ore is heated on rotary kilns to volatilize mercury:
HgS + 02 -> Hg + S02
The gas is cooled, and the mercury is condensed; gases leaving
the condenser contain about 70 ppm mercury. The gases are washed
with sodium thiocyanate to reclaim this mercury. About two-
thirds of the mercury is dissolved as the sodium thiocyanate
complex, and one- third precipitates as mercuric sulfide as
indicated below:
3Hg +• SNaSCN + ^E^O^ + S02 •»• 2Na2 [Na(SCN)4l + HgS
Sodium sulfate is added to the solution to achieve complete
precipitation of mercury as HgS and concurrently to regenerate
the sodium thiocyanate solution:
Na2[Hg(SCN)4] + Na2S •*• + 4NaSCN
i
98
-------
The HgS is filtered out, and the solution is recycled; however,
the formation of sodium sulfate eventually reaches the point
where it crystallizes and contaminates the HgS. The sodium
thiocyanate and sodium sulfate are separated by fractional
crystallization to avoid this problem.
References
1. Habashi, F. Metallurgical Plants: How Mercury Pollution is
Abated. Environmental Science and Technology, 12 (13)-1372-
1376, December 1978.
5.15 DEMARC D PROCESS
1,2
Showa Denko K.K. of Japan has developed the DeMarc D process
for removal of mercury from smelter gases containing sulfur
dioxide. Three Japanese smelters treating copper and zinc con-
centrates have piloted this patented process.
Smelter gases are cleaned in a boiler/electrostatic precipi-
tator system and are then cooled to 130° to 150°C; they are next
passed through towers packed with activated carbon. The carbon
is impregnated with an inexpensive reagent. Steam is used to
1 2
regenerate the carbon and recover the mercury. '
The pilot systems have reportedly operated successfully on
3
gas streams containing 8 percent S02 and 5 mg mercury/Nm . Over
99 percent of the mercury is removed from the gas, and the
product sulfuric acid contains less than 0.1 ppm mercury.
The capital cost of a pilot-scale facility (50,000 m /h) in
1976 was reported as $520,000, and the estimated direct operating
cost for a gas containing 5 mg mercury/m is $2/Mg acid. Energy
consumption for 1000 m3 of gas is 2.26 liters of fuel oil and 6.7
kWh of electricity. Smelters with large gas volumes can ex-
perience difficulties with the DeMarc D process. Use of a filter
technique can be expensive, and a large amount of gas cooling is
required.
99
-------
References
1. Showa Denko K.K. DeMarc-Process for Removal and Recovery of
Mercury from Gas Streams. Tokyo, Japan.
2. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates, Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
5.16 BOLIDEN PROCESS1
The process developed by Boliden a.-b. of Sweden for removal
of mercury from sulfuric acid involves the addition of sulfur as
sodium thiosulfate to the acid. The sodium thiosulfate decom-
poses to form elemental sulfur in a fine colloidal form and
oxides of mercury. This process is only applicable to acid of 85
percent strength or less because higher concentrations oxidize
the sulfur to sulfur dioxide.
A dosage of 0.5 kg of sodium thiosulfate per cubic meter of
cooled, diluted acid is enough to reduce the mercury content from
15 ppm to less than 0.5 ppm in about 1 hour, after which the acid
is filter-pressed.
References
1. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates, Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
Bibliography
Habashi, F. Metallurgical Plants: How Mercury Pollution Is
Abated. Environmental Science and Technology, 12(13):1372-
1376, December 1978.
100
-------
Sundstrom, 0. Mercury in Sulphuric Acid - Boliden Processes Can
Control Hg Levels During or After Manufacture. Sulphur, No
116, pp. 37-43, January-February 1975.
5.17 MITSUI PROCESS1'2
Mitsui Mining and Smelting Co. uses powdered aluminum to
remove mercury from sulfuric acid produced at a lead-zinc smelter
in Kamioka, Japan. In addition, Akita Zinc, Ltd., reportedly
uses the Mitsui process at an electrolytic zinc plant in lijima,
Japan.
The process involves adding powdered aluminum to acid and
allowing 24 hours for a reaction in water-cooled vessels. The
acid solution is then pressure-filtered to produce sulfuric acid
containing about 5 ppm mercury and a filter cake of about 10
percent mercury. The filter cake can also contain selenium,
lead, and other impurities.
References
1. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates, Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
2. PEDCo Environmental, Inc. Pollution Control in the Japanese
Primary Nonferrous Metals Industry. Prepared for the U.S.
Environmental Protection Agency under Contract No. 68-02-1375,
Task No. 36. Cincinnati, Ohio, March 1978.
5.18 GORTDRUM MINES PROCESS
Gortdrum Mines, Ltd., of Ireland has developed an unusual
technique for the treatment of mercury-laden copper concentrates.
The technique is applied to copper-arsenic-antimony sulfide
concentrates containing 0.05 to 0.9 percent mercury before ship-
1 2
ment to a smelter. '
As shown in Figure 29, the concentrate is fed to a multiple-
hearth roaster, where it is heated in a neutral atmosphere (1 to
2 percent O2) to a temperature of 650°C at the final hearth.
101
-------
DRYING
AIR AND OIL ZONE
^BURNER
FABRIC
FILTER
ELECTROSTATIC
FILTER
iJsrv
DUST
CYCLONE
HOT GAS
FAN
DISCHARGE
STACK
AIR
AND CASTING MULTIPLE-
OIL ZONE H£ARTH
FURNACE
SCRUBBER
REDWOOD TANKS
V
CONDENSERS
A /
SOOT THICKENERL_,__I
Figure 29. Flowsheet of solids for Gortdrum Mines process.
FILTER STOCKPILE
CAKE CONCENTRATE
MERCURY SOOT
2
1
SCREW ^^S
CONVEYOR
JL
ELEVATOR
ROTARY i/Jlf
nnvrn ti~*i s-\ xx •
SCREW ^
CONVEYOR ^
RFI T f
DLL \ ^
CONVEYOR A f«-
MM
^«
SCREW
V CONVEYOR
P\ ^-v rf-^ -^J
f
1
2
3
4
5
6
1 1
r\ r* o j«t j\ 3
P^^ ,.'.r..T.t wninFiTTF
MULTIPLE -
HEARTH
FURNACE
STOCK PILE
PADDLE COOLER
MIXER
Figure 30. Flowsheet of gases for Gortdrum Mines process.'
102
-------
Oxygen content is limited to minimize physical and chemical
changes in the concentrate and inhibit formation of sulfur
dioxide and arsenic and antimony trioxides. Approximately 90
percent of the mercury is volatilized in the roaster; this mer-
cury is removed from the roaster off-gas by cooling the gas from
160° to 22°C in water spray condensers. The condensate, which
contains about 50 percent mercury and 2 percent copper, is col-
lected in a water seal and treated for recovery of the mercury.
Figure 30 shows the final wet scrubber, which follows the con-
denser and cleans the gas of any remaining condensed mercury
before discharge to the atmosphere. Mercury concentration in the
tail gas has not been reported.
Because this process is operated as a separate facility, the
capital and operating costs should be high. A unit capable of
treating 12,000 Mg/yr of concentrate was estimated to cost over
$2,000,000 (Canadian) in 1976. Operating costs in 1972 were
given as $23.87(Canadian)/Mg; fuel oil and electricity accounted
2
for about 17 percent ($4.02) of this amount.
References
1. Stuart, M., et al. Mercury Removal From Copper Concentrate.
Transactions of the Society of Mining Engineers, AIME, Vol.
256, September 1974.
2. Reimers, J.H., et al. A Review of Process Technology for
Mercury Control in Gases in the Nonferrous Metallurgical
Industry for the Air Pollution Control Directorate. Jan H.
Reimers and Associates, Limited, Metallurgical Consulting
Engineers, Oakville, Ontario, Canada, October 1976.
5.19 SWINGAWAY CONVERTER HOODS
Peirce-Smith converters are a major source of fugitive emis-
sions at primary copper smelters. Various types of collection
systems are used, either singly or in combination, to control
some of these emissions.
Nippon Mining Company of Japan uses a swingaway hood with a
retractable secondary hood above. During a blister copper pour,
emissions are deflected into the secondary retractable hood.
103
-------
Emissions entering the hooding system are then routed to a col-
lection system. During blowing, slagging, and pouring of blister
copper, this system can be very efficient.
The swingaway hood can be a steel shell with a castable
refractory lining, pillar mounting, and motorized mechanism. In
the retracted position, a swingaway hood must be clear of the
converter aisle and slightly behind the converter to eliminate
interference with the crane.
References
1. Niimura, M., et al. Control of Emissions at Onahama Copper
Smelter. Presented at the Joint Meeting of the Mining and
Metallurgical Institute of Japan and the American Institute
of Mining, Metallurgical, and Petroleum Engineers, Tokyo,
Japan, May 24-27, 1972.
2. Sharma, S.N., et al. Control of Secondary Emissions From
Copper Converters in Copper and Nickel Converters. In:
AIME Symposium Proceedings, New Orleans, Louisiana, February
19-21, 1979. pp. 312-335.
3. PEDCo Environmental, Inc. Pollution Control in the Japanese
Primary Nonferrous Metals Industry. Prepared for the U.S.
Environmental Protection Agency under Contract No. 68-02-
1375, Task No. 36. Cincinnati, Ohio, March 1978.
Bibliography
Symposium Proceedings: Control of Particulate Emissions in the
Primary Nonferrous Metals Industries. U.S. Environmental
Protection Agency and the Southern Research Institute,
Monterey, California, March 18-21, 1979. Monterey, Cali-
fornia.
5.20 ACCORDION DOOR SECONDARY HOODING1"3
Accordion door secondary hooding for copper converters is in
use at the Onahama Smelting and Refining Co., Ltd., Onahama,
Japan. This involves placing a completely enclosed hood over the
primary hood, as shown in Figure 31. During blowing and non-
blowing, fugitive gases are collected through a separate duct-fan
system, cleaned in a wet ESP, and discharged through a 75-meter
stack.
104
-------
TO FABRIC
FILTER
TO SCRUBBER
REVERBERATORY
TO WET ELECTRO-
STATIC PRECIPITATOR
Figure 31. Accordion door secondary hooding at Onahama.'
105
-------
The converter building at Onahama is completely sealed, and
the ventilation air and gas escaping the secondary hoods are
drawn out by a fan and routed through a fabric filter before
discharge.
References
1. Niimura, M., et al. Control of Emissions at Onahama Copper
Smelter. Presented at the Joint Meeting of the Mining and
Metallurgical Institute of Japan and the American Institute
of Mining, Metallurgical, and Petroleum Engineers, Tokyo,
Japan, May 24-27, 1972.
2. Sharma, S.N., et al. Control of Secondary Emissions from
Copper Converters in Copper and Nickel Converters. In:
AIME Symposium Proceedings, New Orleans, Louisiana, February
19-21, 1979. pp. 312-335.
3. PEDCo Environmental, Inc. Pollution Control in the Japanese
Primary Nonferrous Metals Industry. Prepared for the U.S.
Environmental Protection Agency under Contract No. 68-02-
1375, Task No. 36., Cincinnati, Ohio, March 1978.
Bibliography . •.'''.
•' ' n '.-
Coleman, R.T. Trip Report: Mitsubishi Metals Corporation,
Onahama, Japan, July 9-10, 1979. Radian Corporation,
Austin, Texas, 1979.
Symposium Proceedings: Control of Particulate Emissions in the
Primary Nonferrous'Metals Industries. U.S. Environmental
Protection Agency and the Southern Research Institute,
Monterey, California, March 18-21, 1979.
5.21 ENCLOSED TWO-POSITION SECONDARY HOODING1
Boliden a.-b. of Sweden uses an enclosed, two-position gate
design in its secondary hooding of.copper converters. As shown
in Figure 32, this secondary hood is located over the front of
the primary hood and supports a trackway for a single rolling
door that covers only 50 percent of the hood opening. One-half
of the hood is always open to capture fugitive gas.
The rolling door must be down for charging and up to allow
access to the ladle, but at other times the door position, and
thus the ventilated zone, is adjustable at the discretion of the
106
-------
EXISTING EXHAUST SYSTEM
CENTRIFUGAL FAN
CONVERTER,
FLUE
SECONDARY HOOD DOOR
IN MAX. UP POSITION
SECONDARY HOOD
PRIMARY HOOD
FLUX FEEDER
LADLE IN FILLING
POSITION
LADLE IN FILLING
POSITION
SIDE ELEVATION
FRONT ELEVATION
Figure 32. Enclosed two-position secondary hooding.
107
-------
converter operator. Access to the converter for observation,
sampling, cleaning, and repairing is provided by openings in the
sides of the secondary hood.
2
The secondary hood, which is about 34 m in area, clears the
whole primary hood door mechanism and cooling system. The
exhaust system operates continuously.
Reference
1. Sharma, S.N., et al. Control of Secondary Emissions From
Copper Converters in Copper and Nickel Converters. In:
AIME Symposium Proceedings, New Orleans, Louisiana, February
19-21, 1979. pp. 312-335.
Bibliography
Symposium Proceedings: Control of Particulate Emissions in the
Primary Nonferrous Metals Industries. U.S. Environmental
Protection Agency and the Southern Research Institute,
Monterey, California, March 18-21, 1979.
5.22 AIR CURTAIN HOODING1
Mitsubishi Metal Corporation uses primary and secondary
hooding in conjunction with an air curtain to control fugitive
emissions from Peirce-Smith converters at its primary copper
smelter in Onahama, Japan.
The primary hoods fit tightly on each converter (with a gap
of only 5 to 8 cm), almost completely capturing emissions during
the blowing cycle. During charging and pouring the primary hood
draft is reduced by about 65 to 80 percent, capturing about 30 to
50 percent of fugitive emissions. Control of the remainder of
the fugitives is provided by the secondary hoods, air curtains,
and building exhaust ventilation.
The secondary hoods are sheet metal panels on either side of
the converter that rise about 14 meters from the floor. At the
top of these panels, three fans are fitted into the hood, cre-
ating an air curtain across the slot in the top of the hood. The
slot is large enough to allow crane cables to maneuver during
108
-------
charging and pouring. On the other side of the slot is a duct
that collects the exhaust from the three air curtain fans. The
exhaust duct for the secondary hood is atop the panels on the
side farthest from the aisle. Figure 33 shows this configura-
tion. The secondary hood operates all the time; it has been
observed to capture an estimated 30 to 50 percent of the gases
escaping the primary hood during charging and pouring, and
nearly all fugitives during blowing and standby.
The unique feature of this system, however, is the air cur-
tain across the slot in the secondary hood. Three fans create
the stream of air that passes above the slot and enters the
capture duct on the opposite side of the slot. This "push/pull"
technique is extremely effective in collecting fugitive gases
that would normally escape through the slot. The gases collected
by the air curtain contain 100 to 200 ppm SO ; they are sent
through a fabric filter prior to discharge through a stack.
Reference
1. Coleman, R.T. Trip Report: Mitsubishi Metals Corporation,
Onahama, Japan, July 9-10, 1979. Radian Corporation, Austin,
Texas, 1979.
109
-------
14 m „—^
PRIMARY HOOD
(OPEN)
AIR CURTAIN
SLOT FOR
CRANE CABLE
SECONDARY
HOOD EXHAUST
TO MgO SCRUBBER
AIR CURTAIN
EXHAUST
TO FABRIC
FILTER
FEED HOPPER
TUYERE AIR
Figure 33. Air curtain hooding.
1
11.0
-------
SECTION 6
WATER POLLUTION CONTROL PROCESSES
6.1 BOLIDEN SULFIDE PRECIPITATION PROCESS
Boliden a.-b. of Sweden has recently installed a wastewater
treatment facility at its Ronnskar works that makes use of sul-
fide precipitation for control of heavy metals. This new facil-
ity treats both process waters and plant runoff from the primary
copper and lead smelter.1'2
The Ronnskar plant is designed to handle an average waste-
water flow of about 200 m /h. A flow diagram for the plant is
presented in Figure 34. Process water with a pH of approximately
2.0 enters the plant and is treated with sodium hydroxide to
raise the pH to about 3.0. This initial pH adjustment helps pre-
vent corrosion of the equipment. Pretreatment of runoff consists
of equalization and grit removal. Following pretreatment, the
process and runoff streams are combined for subsequent treatment
in a series of three flash-mix reactors.
The first reactor has a capacity of 50 m and an average
residence time (RT) of about 15 minutes. Sodium hydroxide is
added to the wastewater in this reactor. A pH of 4.0 to 5.0 is
attained; this pH range appears to be optimum for sulfide precip-
3
itation.
The second reactor has a capacity of 30 m with an average
residence time of about 9 minutes; sodium sulfide is added to the
stream. The third reactor is auxiliary to the second, providing
additional residence time for treatment with sodium sulfide if
required. The addition of the sodium sulfide produces the fol-
lowing reactions.
Ill
-------
NaOH
NaOH
Na2S
PROCESS
MATER^ ^
100 m3/h
pH = 2.0 P
PH
SURFACE
RUNOFF
100 m3/h
GF
PLING KEY:
II
H ADJUSTMENT
= 1.3 h
= 3.0
(
(IT REMOVAL
. DAILY 24-hou
COMPOSITES
A r~ r\r~ f i Hi f- Aur
1 JL JL
A •• AM A«> i A
RT «= 15 min Rl = y irnn Rl= 15 min 1
pH = 4-5 ^
CLARIFIER 75 (
DIAMETER = 15 m L- ~
RT = 3 h
SLUDGE
HOLDU
TANK
ci M rtnc" ci i inpr ^M • A>
oULrlUt oLUUbt — w
TO ROASTER
FILTRATE .
FILTER
r BACKWASH
i re ST< . i
A
*
-^ POLYME
« SOLIDS,
~ll
IG
< i
^
VACUUM F
AREA = 3
FILTER YIE
200 kg/h
TAKEN AS REQUIRED
TO HOLDING POND
MULTIMEDIA
FILTERS
FLUORIDE PRECIPITATOR
RT = 40 min.
DISCHARGE = 3 mg/liter solids
Figure 34. Simplified flowsheet of the Boliden sulfide
precipitation process.^
-------
Na2S + Na+ + s = M++ + S =
The dissolved metal cations form metal sulfide precipitates.
Excess sulfide must be avoided to eliminate the potential for-
mation of hydrogen sulfide.3
Clarification and sludge thickening are accomplished in a
circular clarifier having a diameter of 15 m and an average flow
rate of 200 m /h; detention time is about 3 hours. The clarifier
is equipped to feed polymer coagulants to the stream.3
Overflow from the clarifier has been estimated to contain
less than 50 mg/liter of solids, and bench-scale tests by Boliden
indicate that levels as low as 15 mg/liter may be possible. Two
trimedia filters, however, are utilized at the plant to achieve a
level of 3 mg/liter solids. The filters are each 3 m in diameter
and 1.3 m deep; they are reportedly backwashed every 24 hours,
with the backwash returned to the pH adjustment tanks.
Sludge from the clarifier has been estimated to contain 5 to
6 percent solids; it is first routed to a holding tank where a
polymer is added to aid in subsequent dewatering. Dewatering is
accomplished in a vacuum filter operating 8 to 10 hours per day,
2
7 days per week. The filter has a fabric area of 38 m and pro-
duces an estimated 200 kg dry solids per hour of sludge with a
solids content of 25 percent. This sludge is charged to the cop-
per roasters and the filtrate is returned to the pH adjustment
tanks.
The process water initially contains a relatively high con-
centration of fluorides, and evaporation during processing results
in an even higher level. Consequently, a separate unit is in-
stalled in the wastewater treatment plant for the control of
fluorides. This unit consists of a flash mix/flocculator to
which hydrated lime is added.
Domestic research into the effectiveness of sulfide precipi-
tation indicates its apparent superiority to conventional hydrox-
ide (lime) precipitation in the removal of heavy metals from both
synthetic mixtures and actual scrubber waste from a copper smelt-
4
xng operation.
113
-------
References
1. Coleman, R.T., et al. Test Program for Characterizing
Boliden Aktiebolag's Sodium Sulfide Precipitation Process.
Prepared for U.S. Environmental Protection Agency under Con-
tract No. 38-02-2608. Radian Corporation, Austin, Texas,
March 6, 1978.
2. Devitt, T.W. Visit to Boliden Smelter, Skellefteahamn,
Sweden. PEDCo Environmental, Inc., Cincinnati, June 6,
1977.
3. Coleman, R.T. Emerging Technology in the Primary Copper
Industry (draft). Prepared for U.S. Environmental Pro-
tection Agency by Radian Corporation under Contract No.
68-02-2608. Radian Corporation, Austin, Texas, August 31,
1978.
4. Bhattacharyya, D., A.B. Jumawan, Jr., and R. B. Grieves.
Separation of Toxic Heavy Metals by Sulfide Precipitation.
Separation Science and Technology, 14(5):441-452, 1979.
Bibliography
Coleman, R.T., et al. Treatment Methods for Acidic Wastewater
Containing Potentially Toxic Metal Compounds. Prepared for
U.S. Environmental Protection Agency under Contract No.
68-02-2608. Radian Corporation, Austin, Texas.
Robinson, A.K. Sulfide-vs-Hydroxide Precipitation of Heavy
Metals from Industrial Wastewater. In: First Annual
Conference of Advanced Pollution Control for the Metal
Finishing Industry. American Electroplaters Society and
U.S. Environmental Protection Agency, Lake Buena Vista,
Florida, 1978. pp. 59-65.
Schlauch, R.M., et al. Treatment of Metal Finishing Wastes by
Sulfide Precipitation. EPA 600/2-77-049, February 1977.
6.2 FERRITE PRECIPITATION PROCESS1
Fuji Kasui Engineering Co., Ltd., of Japan, has developed a
process for removing heavy metals from wastewater by converting
them to ferrite precipitates (patent).
In this process, as shown in Figure 35, iron compounds such
as ferrous sulfate are used as precipitating agents in quantities
that depend upon the type of metal to be removed. For example,
experimental work indicates that zinc, manganese, and copper can
114
-------
FERROUS SULFATE
WASTEWATER
CONTAINING
HEAVY METALS
ALKALI
AIR
NEUTRAL-
IZATION
OX I DA
TION
II
FERROMAGNETIC
PRECIPITATES
CONTAINING
HEAVY METALS
NEUTRAL AQUEOUS
SOLUTION
Figure ,35; Ferrite precipitation process.
1
115
-------
easily be converted to ferrite. The addition of ferrous salt in
a"quantity twice the molar ratio of those metals is sufficient to
accomplish this conversion. Metals such as tin and lead, however,
do not easily form solid ferrites; they require the addition of
large quantities of ferrous salt.
Posttreatment metal concentrations in the aqueous solutions
used in experimental trials of the ferrite precipitation process
were significantly lowered: for example, copper from 9500 ppm to
0.5 ppm and cadmium from 1800 ppm to 0.1 ppm. The heavy metal
ions were converted to ferrite precipitates and separated magnet-
ically.
This process has several advantages, which are described be-
low, over conventional techniques.
Precipitates formed by conventional processes may redissolve
after disposal; those formed by this new process have very
little risk of redissolution.
Precipitates formed by conventional processes are so fine
that they cannot be filtered, while precipitates formed by
the ferrite process have an average particle size suitable
for filtration. In addition, they can be separated and
removed magnetically because they are ferromagnetic mate-
rials.
Iron compounds used in this process, such as ferrous sul-
fate, can be obtained in large quantities as industrial
wastes.
The most suitable condition for precipitation in conven-
tional processes varies according to the type of heavy
metal. Consequently, each heavy metal must be treated
individually. In the ferrite process, wastewaters that
contain various types of heavy metals can be treated at the
same time once the most suitable condition for precipitating
ferrite is established.
Unlike conventional processes, this one can remove hexavalent
chromium ions Cr (VI). This occurs by reduction of the
hexavalent chromium ions followed by precipitation.1
Reference
1. U.S. Patent Office. Ferrite Process: New Technology for
Removing Heavy Metals from Wastewaters. Patent No.
3,822,210 issued to Fuji Engineering Company, Ltd.
116
-------
TECHNICAL REPORT DATA
(Please read Instructions on the reverse before completing]
. REPORT NO.
EPA-600/2-80-159
3. RECIPIENT'S ACCESSION NO.
ITLE AND SUBTITLE
verview of Foreign Nonferrous Smelter Technology
5. REPORT DATE
JUNE 1980 ISSUING DATE.
6. PERFORMING ORGANIZATION CODE
AUTHOR(S)
A. Christian
Vtary A. Taft
Worrell III
8. PERFORMING ORGANIZATION REPORT NO.
i. PERFORMING ORGANIZATION NAME AND ADDRESS
EDCo Environmental, Inc.
1499 Chester Road
Cincinnati, Ohio 45246
10. PROGRAM ELEMENT NO.
1AB604
11. CONTRACT/GRANT NO.
68-03-2577
12. SPONSORING AGENCY NAME AND ADDRESS
Industrial Environmental Research Laboratory
}ffice of Research and Development
J. S. Environmental Protection Agency
Cincinnati, Ohio 45268
13. TYPE OF REPORT AND PERIOD COVERED
One of Series
14. SPONSORING AGENCY CODE
EPA/600/12
15. SUPPLEMENTARY NOTES
roject Officer: John 0. Burckle
16. ABSTRACT
Numerous production and pollution control processes that are not used in the United
states are in use or under development by foreign nonferrous metal producers. Although
some do not apply to U.S. conditions, others can reduce pollution, increase production,
>r lower costs. Many of these foreign processes are described in this report.
The descriptions are divided into five categories: pyrometallurgical processes,
lydrometallurgical processes, electrolytic processes, air pollution control processes,
md water pollution control processes. If data were available, each process description
Includes a discussion of economic, environmental, and energy considerations, as well as
i discussion of the basic operating principles. A detailed analysis of each process is
lot attempted in this report. For additional information, the reader is referred to
the list of references and bioliography following each process description.
17.
KEY WORDS AND DOCUMENT ANALYSIS
a.
DESCRIPTORS
b.lDENTIFIERS/OPEN ENDED TERMS
Air Pollution Control
Smelting Process
Pyrometallurgy
Hydrometallurgy
Copper
Lead
Zinc
Water Pollution Control
COS AT I Field/Group
Sxhaust Emissions
imelting
Trace Elements
Dilution
13B
8. DISTRIBUTION STATEMENT
glease to Public
19. SECURITY CLASS (ThisReport)
Unclassified
127
20. SECURITY CLASS (Thispage)
Unclassified
22. PRICE
EPA Form 2220-1 (Rev. 4-77)
PREVIOUS EDITION IS OBSOLETE
117
U.S. GOVERNMENT PRINTING OFFICE: 1980--657-165/0018
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