United States Industrial Environmental Research
Environmental Protection Laboratory
Agency Cincinnati OH 45268
Research & Development
Bulletin
Nonferrous Metals
Technical Awareness
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DISCLAIMER
This document has been reviewed by the Office of Research and Development,
U.S. Environmental Protection Agency, and approved for prepublication on
a limited basis prior to preparation of a final report. Approval does not
signify that the contents necessarily reflect the views and policies of
the Environmental Protection Agency nor does mention of commercial
products constitute endorsement or recommendation for use.
This report is intended to provide state-of-the-art information on
nonferrous metals developments. This overview document should be useful
in providing a background perspective to assess environmental conditions
for the industry. The "user" should be aware that some of the technical
aspects may be changed in the final report.
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NON-FERROUS METALS INDUSTRIES
IN THE NEWS (m
UJ
C3
... a publication of the USEPA Metals and Inorganic Chemicals Branch, Office of Research
and Development. It is designed to spotlight selected recent events, concerns, and
technology within the nonferrous metals industry as reported in the open literature, and does
not necessarily reflect the views of the USEPA.
July 1979
HIGHLIGHTS
• Aluminum faacilitieA closed in 1977 by enen.gy coni>tn.ainti> an.e being
reopened in n,et>poni>e to &tn.ong manket de.ma.ndi>. (Item* 3 and 4,
Chemical Engineering and kme.Jiic.an Metal Man.ket. )
9 Inventory redaction, continued &on.eign pJioda.cti.on iJULb, and mon.e
attractive demand* and pnice* ^on. coppun and molybdenum OJKL
ApaSLking renewed activity in the. domestic, copper indu&tsiy. (lte.mt>
S-12, PaydUt and oth&i
• Wo/uwda U> the. winning bidden, who may ie.ap benefit* in taking
the. Papago-owne.d LakeAhotLe. coppeA fiacA-tity H.e.ce.ntly written
by He.cla and El Pa&o. [Ite.m 12, NontheAn Mine.*..)
OSHA and EPA lead Atandandt> continue, to faobteA debate, and fihetofiic
between induAtiy, government, and public interest g>ioupt>, and
appear to be jJoA. fiiom a settled i&Aue. EPA, OUi.ce of, Re&caich and
Development hat> initiated a neAeaich ptiogtum in thu> anea. (Item 29}
{All item& linden, "lead" numcnouA &OUA.CU.}
'Demand and police* continue to encouAage explanation and n.eacti.vation
o& gold and bilvcn. opcnationt>. Small operations, in pan£iculan., an.e
w.olitcnatina. (See "PneciouA M&taW; rh^ Mining ^econ.d an^ otken.
Individually contributed news items may be submitted to Metals and Inorganic Chemicals
Branch, Industrial Environmental Research Laboratory, USEPA, 5555 Ridge Avenue.
Cincinnati, Ohio 45268, 513-684-4491.
Prepared by Battelle's Columbus Laboratories under EPA Grant R805095.
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Aluminum
(1) One of the nation's largest secondary aluminum smelters, Michigan
Standard Alloys, Inc., of Benton Harbor, Michigan, will phase out its alu-
minum operations and its dross mill within 30 days because of problems with
pollution regulations.
(American Metal Market)
(2) In an effort to "break the bauxite barrier", Alumet intends to •
open-pit mine alunite on 13,000 acres and build a processing complex on 5.000
acres of land in Beaver County, Utah. The company has been trying since mid-
1975 to obtain an environmental impact statement (EIS) for the $500-million
project. They paid a $93,250 assessment for preparation of the statement by
the Bureau of Land Management (BLM) but have refused to pay a second assess-
ment for $166,750. After 2 years' litigation, the court enjoined the BLM
from charging Alumet. The Department of Justice and Alumet will argue the
case later this year.
(Engineering and Mining Journal)
(3) Reynolds Metals Company will reopen its primary reduction plant at
Corpus Christi, Texas, because of an increase in demand for aluminum. Pro-
duction from a 57,000-ton/yr potline will begin in May, bringing Reynolds to
94 percent of its primary aluminum capacity. A second potline will be acti-
vated as soon as the first returns to production.
(Chemical Engineering)
(4) Alcoa is also reactivating its Port Comfort, Texas, smelter. Two
potlines are scheduled to be reactivated in May and the third in July. The
company's primary aluminum production will be increased by 90,000 short tons
per year.
(American Metal Market)
(5) Ogden Metals will increase its production capacity for secondary
aluminum ingot by 10 percent to 300 million pounds per year, with a $1.6-
million investment in its Wabash Alloys, Inc., subsidiary in Cleveland. Oh'io.
and in Wabash, Indiana. The Wabash unit will also enter the scrap market for
aluminum and bimetallic cans.
(American Metal Market)
Beryllium
(6) Brush Wellman, Inc.. will expand and upgrade its minerals process-
ing plant at Delta, Utah.A $0.5-million beryllium sludge treatment process
will be installed in addition to a $l-million plant for uranium production.
(USBM Minerals and Materials)
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Cobalt
(7) An investigation is being undertaken into the feasibility of
reopening the Blackbird mine at Cobalt, Idaho. An active producer in the
1950's, new technology will enable the recovery of more cobalt from the ore
in the milling process. There is presently no domestic supply of mined
cobalt, and less than 4 percent of the domestic demand for refined cobalt
is supplied by U.S. producers.
(The Mining Record)
Copper
(8) The Duval Corporation resumed operations at its Esperanza copper-
molybdenum mine near Tucson in Pima County, Arizona, in April. It has been
closed since October, 1977. Production is being resumed because of the
recent firming trend in the copper market and the strong, world-wide demand
for molybdenum. Production capacity of the Esperanza concentrator is 33 mil-
lion pounds of copper and 3 million pounds of molybdenum annually.
(New Mexico Paydirt, Ski 1 lings' Mining Review)
(9) Cypress Mines Corporation will resume ore stripping operations in
May and restart the concentrator early in the third quarter of 1979 at the
Pima mine near Tucson, Arizona, down since September, 1977. Production will
reopen on a limited scale because of the surge in the price of copper and
molybdenum. Molybdenum is a by-product contained in substantial quantities
in the Pima ore. Limited production of copper is expected to total about
24,500 stpy (of contained copper) or one-third the 1977 production level.
(Arizona Paydirt)
(10) Closed since the summer of 1977, Inspiration Consolidated Copper
Company will reopen its Christmas mine near Phoenix, Arizona, in about 10
weeks. Full production capacity is about 6,000 stpd of ore.
(American Metal Market)
(11) The Oracle Ridge property near Tucson, Arizona, is being developed
jointly by Continental Materials Corporation and Union Miniere as an under-
ground copper mine capable of supplying 14,000 st of ore to the concentrator,
weekly. A copper recovery of 89 percent and a concentrate grade in excess
of 32 percent copper were consistently obtained during the test program.
However, the project has been substantially curtailed until completion of a
comprehensive study to more precisely define the ore zones and reassess the
underground mining conditions. The study is expected to be completed during
1979.
(Engineering and Mining Journal, Ski 11 ings' Mining Review)
(12) The Lakeshore copper mine near Casa Grande, Arizona, has been
leased by the Papago tribe to Noranda Exploration, Inc. Noranda will begin
underground mining in 6 months of the 472-mi11 ion-ton deposit averaging 0.75
percent copper. By September, a production of 6,000 stod of ore is expected
to be reached. Noranda will pay $10 million for the plant written off
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by Hecla and El Paso to the tune of hundreds of millions of dollars. An ini-
tial $1.6-million lease and preroyalty payment is also involved.
(The Northern Miner, Engineering Mining Journal)
(13) The Nevada mines division of Kennecott Copper Corporation has
begun engineering work on the planned McGi11 tailings retreatment project.
Accumulating for over 70 years from the Ruth mine which closed in 1978, natu-
ral classification has resulted in a deposit of minable grade copper-bearing
material. Whether or not Kennecott will proceed with this will depend upon
forthcoming engineering data and economic analysis.
(Skillings1 Mining Review)
(14) Smelting operations at the Kennecott Copper Corporation's McGi11,
Nevada, plant will be suspended for at least 3 months, beginning in February,
to make major repairs on reverberatory furnaces and boilers.
(New Mexico Paydirt)
(15) Anaconda Company will temporarily curtail operations during April
at its copper smelter in Anaconda, Montana, due to lack of feed. Increase
in freight rates to ship concentrates from Canada is blamed. Carr Fork will
provide feed, commencing late in 1979.
(USBM Minerals and Materials)
(16) The Duval Corporation has permanently closed down the sulfide cop-
per ore operation at its Battle Mountain, Nevada, plant. The sulfide milling
operation was converted to process precious metal ores after the sulfide ore
production tapered off because of low ore grades.
(New Mexico Paydirt)
(17) Park City Ventures, a joint venture of Anaconda and Asarco has
abandoned, at this time, plans to reopen the Ontario No. 3 mine. It was
closed on February 15, 1978, because of high water and unstable rock
conditions.
(Engineering Mining Journal)
(18) Asarco's Amarillo, Texas, copper refinery has begun to refine
copper scrap as well as blister copper. The scrap feed facility was built
in late 1978. Blister copper is being sent to Amarillo from the Tacoma,
Washington smelter following the shutdown of the Tacoma refinery. The
Amarillo refinery produced 270,000 tons of refined copper last year and is
expected to be producing at a rate of 378,000 tpy by the end of 1979.
(Arizona Paydirt)
(19) Kennecott Corporation's Flambeau Mining Company has no plans to
mine copper in the near future at the site near Ladysmith, Wisconsin. The
Wisconsin Department of Natural Resources rejected a 1976 application for
a mining permit, and Kennecott will not appeal this.
(Engineering and Mining Journal)
(20) Hydrometallurgical and pyrometallurgical processes for copper
smelting and refining are being examined at the University of Utah in a
$233,000 renewal of a Department of Energy study.Energy usage is being
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assessed from the time the ore is mined until it becomes high-purity, refined
copper. The goal of the research is to determine the least energy-intensive
copper producing methods.
(Mining Engineering)
(21) A report has been released by the Inform organization, strongly
criticizing copper companies who operate 16 smelters in the U.S. for failing
to adequately protect their workers from exposure to cancer-causing agents.
Half of the 2,800 workers examined were exposed to hazardous levels of
arsenic and a fourth to hazardous levels of sulfur dioxide. Many others are
exposed to dangerous levels of copper dusts, fumes, cadmium, silica, and
noise. Reactions from industry: "replete with errors and misstatements";
none of the findings is correct"; "unqualified sources, hearsay, gossip, and
conjecture", etc.
(American Metal Market)
(22) Reversing its previous stand on the subject, the U.S. in late
February proposed an international agreement to stabilize the price of cop-
per by fixing a median price to be maintained by means of a buffer stock of
at least 1 million metric tons.
(Arizona Paydirt)
(23) The cost of implementing federal regulations in the copper indus-
try, over the next 10 years, could close three major copper smelters and
force those remaining to face a "no-growth situation" according to a review
by the Department of Commerce. Those which possibly will close are operated
by Kennecott at McGill, Nevada, by Phelps Dodge at Douglas, Arizona, and by
Asarco at Tacoma, Washington. Adherence to the regulations will require
investments of $1.8 billion and total expenditures of $3.5 billion, and will
increase the price for copper.
(American Metal Market)
(24) III 1974. Kennecott's Ray Mines Division in Arizona discontinued
the practice of charging copper-rich converter slag to the reverb to allay.
air pollution, and instead, now returns copper values to the head of the
line (crushing) operation. This has caused logistics problems. RMD is
adapting to slag inventory buildup by additional truck haulage and the
installation of a new grizzly to handle this excess inventory.
(New Mexico Paydirt)
Government Regulations
(25) In order to meet national air quality standards by the end of
1982, the New Mexico Environmental Improvement Board has adopted a plan
which focuses particular attention on 10 areas that fail to meet air quality
standards. One of these areas is near Silver City, location of the Hurley
smelter. To meet this plan, Kennecott will probably have to spend $140 mil-
lion to control sulfur dioxide and particulate matter at Hurley. (Added in
proof: Kennecott has big plans for Chino Mines Division, including Hurley
smelter--see next Bulletin.)
(New Mexico Paydirt)
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Lead
(26) Mr. D. B. Craig, president of St. Joe Minerals Corporation, has
suggested the following points on which both government and the lead indus-
try must agree to reach long- and short-term approaches to regulatory
enforcement. These are: environmental improvements made with existing tech-
nology without exorbitant costs; research and development for alternative
conversion methods; and financing both programs so that industry should not
have to bear the entire cost. He also called for new technology for develop-
ment of scrap and stressed the need to make workers in the user and produc-
tion industries aware of the problem.
(American Metal Market)
(27) The District Court of Appeals of Washington, D.C. has stayed the
most costly provisions of the OSHA standard for lead exposure. Requirements
for engineering controls, compliance plans, and construction of new hygiene
facilities were stayed, while the standard which eventually requires a limit
of 50 micrograms of Iead/m3 of air over an 8-hour period went into effect on
March 1, 1979.
(American Metal Market)
(28) Three years after OSHA proposed a standard limiting worker expo-
sure to lead, the controversy concerning it still rages. The prime target
is the 50 microgram/m^ of air standard for exposure. Industry claims the
standard should be based on actual lead in the blood, which should be kept
below 80 micrograms per gram. Legal challenges have been filed by several
groups, and the processes will drag on for some time. (A Supreme Court deci-
sion concerning benzene this year, in which it will decide whether OSHA must
come up with strict risk/benefit analyses before issuing a standard, could
destroy the lead standard rule.)
(American Metal Market, Lead and Zinc Supplement)
(29) The EPA ambient lead standard is the principal area of concern of
six primary lead smelters and 50 secondary ones. EPA states there has been
no positive response from the industry in the form of engineering data to be
used as guidelines. EPA and NIOSH have initiated joint studies of the occu-
pational & ambient lead problem. Contact Fred Craig, 513-684-4491 or
Jim Gideon at 513-684-4295. (American Metal Market Lead & Zinc Supplement)
(30) ILZRO (International Lead and Zinc Research Organization) last
year spent $480,000 in research and development in the environmental field.
ILZRO states three concepts which are particularly onerous to industry:
(1) the zero threshhold concept which holds that there are no safe concentra-
tions of toxic substances; (2) the proof of safety, in which the burden of
proof is shifted from those contending a substance is harmful to those con-
tending a particular substance is safe; and (3) the adversary law requiring
scientists to be identified with one side or the other in a particular issue.
At stake are 9,300 direct jobs and 13,950 indirect jobs; the closing of all
primary smelters; 40 of 50 secondary smelters and 6 of 200 battery manufac-
turing companies.
(American Metal Market Lead and Zinc Supplement)
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(31) The strike by the production and maintenance employees of
Kennecott's Ozark Lead Company, Sweetwater, Missouri, continues after rejec-
tion of an eleventh-hour proposal. Negotiations for contract renewal began
in January; however, the present contract expired March 1 with no settlement.
No date has been set for resumption of negotiations.
(American Metal Market)
Miscellaneous
(32) Laser Analytics, Inc., of Lexington, Massachusetts, has been
granted a $60,000 contract by EPA's Environmental Sciences Research
Laboratory for development of a flue gas sulfuric acid vapor monitor. A
monitoring instrument is needed to detect sulfuric acid in power plant and
industrial smokestack emissions; however, conventional infrared techniques
cannot separate gases whose wavelengths are too close in the spectrum. An
infrared system using a tunable diode laser as a light source is sought by
the EPA.
(Environment Reporter)
Molybdenum
(33) The development stage, estimated to cost over $200 million, ha_s
been initiated by Cyprus Mines Corporation in their Thompson Creek, Idaho,
molybdenum mine. Located near Clayton in Custer County, the deposit has
reserves estimated at 233 million mt assaying 0.18 percent Mo$2 or recover-
able reserves of 155 million mt averaging 0.201 percent Mo$2. Japanese
steel interests are studying the possibility of financial assistance to
Cyprus in developing this project.
(Engineering and Mining Journal, The Mining Record)
(34) The second largest molybdenum deposit in the world, containing
over 700 million st, with grades ranging from 0.15 to 0.22 percent Mo$2, .
lies in Quartz Hill, Alaska, which has recently been placed in the Misty
Fjords National Monument, an unminable preserve. With a potential value of
$7 billion, U.S. Borax and Chemical Corporation has already invested
$7 million in work at the site. The company feels that it has demonstrated
willingness to protect the area by rehabilitating the adjacent Keta River.
In March, the House Interior Committee approved an Alaska lands bill that
would move the boundary line for the wilderness area 15 miles north to
exclude the deposit. U.S. Borax is optimistic and will proceed with
$3 million of additional exploration work this year.
(Engineering and Mining Journal)
Nickel
(35) The Amax Nickel Refining Company, Inc., at Port Nickel, Louisiana,
received the first nickel matte shipment from the Agnew mine in Hestern
Australia on February 14. This initial 400 mt shipment is part of a 10-year
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contract with the Agnew mine for its entire production up to an annual maxi-
mum of 15,800 mt. The matte contains 72 percent nickel, 6 percent copper,
and 0.6 percent cobalt.
(Ski 11 ings' Mining Review)
Precious Metals
(36) The reconditioned Nancy Lee mill began processing stockpiled ore
on April 5. The 2,000 tons of ore were mined at the Nancy Lee mine near
Superior, Montana, and consists of copper, silver, and gold. Concentrates
will be shipped to East Helena for refining.
(The Mining Record)
(37) Treasure Hill Exploration, Inc., has paid the Silver King Mines
an advance royalty payment of $100,000 on its East Hamilton property near
Ely, Nevada. Treasure Hill plans to process 100,000 tons of ore per year
at its plant near the property. About 225,000 tons of ore are stockpiled,
with another 200,000 to 300,000 tons near the surface, averaging 6.3 ounces
of silver per ton.
(Skillings1 Mining Review)
(38) Gulf Oil is continuing exploration drilling on Silver King Mines'
Ward Properties (14,500 acres) near Ely, Nevada.Indicated reserves of 15
million tons of copper, silver, and zinc have been found in which Gulf can
earn a 51 percent interest by expending $10.5 million over a 7-year period.
(Skillings' Mining Review)
(39) Canadian Superior Mining is conducting further drilling and metal-
lurgical work for Ranchers Exploration and Development Corporation on
Rancher's leased property near Stibm'te, Idaho. Determination of the extent
of an oxidized gold deposit on the property is the objective of the work by
CSM. Production by open-pit mining, heap leaching, and precipitation of
gold concentrates could begin soon if no metallurgical problems develop. .
(Engineering and Mining Journal)
(40) The Mary Nevin Mine and mill made the first shipment of gold from
the Cripple Creek, Colorado District since 1962 on February 14, 1979. On
March 28, operations at the Mary Nevin Mill were expanded to a 24-hour per
day schedule. The mill capacity is presently 25 tons of ore per day but
will soon be expanded to TOO tpd. Concentrates are shipped to an Asarco
buying station in Denver.
(The Mining Record)
(41) Deposits of gold, silver, zinc, and copper have been discovered
in the abandoned Bluejacket mine near Grangeville, Idaho. Approximately
2^5-million tons of ore, with a value of $70 million, lies in this area;
however, the site is within Hells Canyon National Reservational area, so a
renewal of mining may not be permitted. Attempts are being made in Congress
to exclude the mine area from the wilderness classification, but it is being
opposed by the Sierra Club.
(Engineering and Mining Journal)
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(42) Plans to begin trial mining at the Escalante silver mine in south-
west Utah are under way by the Ranchers Exploration and Development
Corporation. The increase in silver prices has added impetus to this project
as the mine is estimated to contain 20-million ounces. Trial mining on its
Alaskan gold placers this summer is also planned by Ranchers.
(New Mexico Paydirt)
(43) The Mikado and perhaps the Little Squaw, two lode deposits belong-
ing to the Little Squaw Gold Mining Company will begin production this year,
financed and operated by Whelan Mining and Exploration, Inc. Areas along
Big Creek, Little Squaw Creek, Tobin Creed, and Big Squaw Creek will be
explored and mined where results are favorable. Each creek has small rich
sections remaining unmined after producing 45,000 to 50,000 ounces of gold
in the past.
(The Mining Record)
(44) Helena Silver Mines has acquired an electric flotation mill near
Missoula, Montana. It will be used to custom mill at its present location
if the Cape Nome-Hidden Treasure properties are reactivated, or to mill
stockpiled material from Helena Silver's Gregory Mine. Summit Silver, Inc.,
would like to rent it for processing silver ore from its Baltimore Mine near
Boulder, Montana.
(The Mining Record)
(45) The Sutro Tunnel at Virginia City, Nevarda, is being reopened in a
renewed search for gold and silver in the famous Comstock mining district.
Restored and retimbered by the Houston Oil and Minerals Corporation, the
gates were opened in January for crews working to restore the 4-mile-long
tunnel for exploration and perhaps production. Completed in 1878, it was
used until the mines closed down in 1942.
(The Mining Record)
(46) The eight largest U.S. mine producers of silver in 1977 and 1978
in millions of ounces were: Sunshine mine in Coeur d'Alene, Idaho, 3.79
and 4.95; Callahan's Galena mine in the Coeur d'Alene, operated by Asarco,
3.70 and 3.99; Hecla's Lucky Friday mine, in the Coeur d'Alene, 2.64 and
2.46; Coeur d'Alene Mines Corporation, Coeur mine, adjoining the Galena mine
and operated by Asarco, 2.38 and 2.40; Anaconda's Berkeley pit copper mine
at Butte, Montana, 3.19 and 2.35; Kennecott's open-pit copper mine at
Bingham Canyon, Utah, 1.96 and 2.22; and Homestake's Bulldog mine in
Colorado, 2.04 and 2.10.
(The Mining Record)
Titanium
(47) With the lead time at the end of last year of 40 to 50 weeks for
titanium alloy, sheet and plate, and up to 70 weeks for forgings, producer's
ability to deliver will depend on the availability of titanium sponge and
scrap. RMI. the only U.S. titanium sponge producer selling externally, has
"maximized its output", and will undertake a $3.5-mil1ion expansion of its
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sponge plant at Ashtabula, Ohio, to be completed by the spring of 1980.
will expand its current 15-million pounds annual capacity by about 25
percent.
(American Metal Market Aerospace Metals and Machines Supplement)
This
Vanadi urn
(48) Ranchers Exploration and Development Corporation has suspended-
plans to extract vanadium and uranium from old mine tailings at Durango,
Colorado, because of delays in licensing by the Colorado Department of
Health.
(USBM Minerals and Materials)
Zinc
(49) There is no expectation of a settlement in the near future of the
11-month-old strike at St. Joe Zinc Company's Balmat-Edwards mining operation
in upstate New York. The Edwards mine is the largest zinc mine complex in
the U.S., supplying 17 percent of the total U.S. mine production of zinc in
1977.
(American Metal Market)
(50) Exploration began about a year and a half ago on St. Joe Minerals
Corporation's Smith County zinc property in Tennessee to assess its profit-
ability. About $5 million has been spent on shaft sinking near Carthage to
determine the size and grade of the ore body. A decision is expected by
late spring or early summer on opening the mine. If favorable, the joint
venture of St. Joe Minerals and Freeport Zinc Company will spend $40 million
to develop the mine.
(USBM Minerals and Materials)
REFERENCES
(1) American Metal Market, 87(54):29, 34, March 19, 1979.
U.S. Bureau of Mines, Minerals and Materials, 47, February, 1979.
(2) Engineering and Mining Journal, 179(11):29, 32, November, 1978.
(3) Chemical Engineering, 86(7):103, March 26, 1979.
Chemical and Engineering News, 57(11):9, March 12, 1979.
American Metal Market, 87(47):!, 16, March 8, 1979.
(4) American Metal Market, 87(43):!, 7, March 2, 1979.
American Metal Market, 87(60):5, March 27, 1979.
American Metal Market, 87(64):28, April 2, 1979.
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(5) American Metal Market, 87(53):!, 10, March 16, 1979.
(6) U.S. Bureau of Mines, Minerals and Materials, 48, March, 1979.
(7) The Mining Record, 90(13):6, March 28, 1979.
(8) New Mexico Paydirt, (22):38, 40, March, 1979.
Sellings' Mining Review, 68(10)1:18, March 10, 1979.
Arizona Paydirt, (476):5, February, 1979.
American Metal Market, 87(37):2, February 22, 1979.
Mining Engineering, 31(4):349, April, 1979.
The Mining Record, 90(9):3, February 28, 1979.
U.S. Bureau of Mines, Minerals and Materials, 3, March, 1979.
(9) Arizona Paydirt, (477):5, March, 1979.
New Mexico Paydirt, (23):40, April, 1979.
American Metal Market, 87(45):!, 13, March 6, 1979.
(10) American Metal Market, 87(81):13, April 25, 1979.
(11) Engineering and Mining Journal, 180(3):45, 49, March, 1979.
Skillings' Mining Review, 68(11):6, March 17, 1979.
(12) The Northern Miner, 65(2):3, March 22, 1979.
Engineering and Mining Journal, 180(4):41, April, 1979.
American Metal Market, 87(52):10, March 15, 1979.
The Mining Record, 90(12):!, March 21, 1979.
Skillings1 Mining Review, 68(12):22, March 24, 1979.
Arizona Paydirt, (477):5, March, 1979.
(13) Skillings1 Mining Review, 68(13):10, March 31, 1979.
(14) New Mexico Paydirt, (22):40, March, 1979.
(15) U.S. Bureau of Mines, Minerals and Materials, 45, March, 1979.
(16) New Mexico Paydirt, (23):73, April, 1979.
(17) Engineering and Mining Journal, 180(4):167, April, 1979.
(18) Arizona Paydirt, (477):44, March, 1979.
(19) Engineering and Mining Journal, 180(3):240, March, 1979.
(20) Mining Engineering, 31(4):349, April, 1979.
Engineering and Mining Journal, 180(3):233, March, 1979.
(21) American Metal Market, 87(71):11, April 11, 1979.
(22) Arizona Paydirt, (477):20, March, 1979.
New Mexico Paydirt, (23):18, April, 1979.
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(23) American Metal Market, 87(81):!, 24, April 25, 1979.
(24) New Mexico Paydirt, (23):59, April, 1979.
(25) New Mexico Paydirt, (21):3, February, 1979.
(26) American Metal Market, 87(72):7, April 12, 1979.
(27) American Metal Market, 87(44):34, April 5, 1979.
Chemical and Engineering News, 57(11):17, March 12, 1979.
Arizona Paydirt, (477):24, March, 1979.
(28) Stern, L., American Metal Market, Lead and Zinc Supplement, 87(70):7A-
8A, April 10, 1979.
(29) Yafie, R. C., American Metal Market, Lead and Zinc Supplement, 87(70):
4A, 6A, April 10, 1979., A. B. Craig, Jr., EPA, Industrial Environmental
Research Laboratory, Cincinnati, Ohio, 513-684-4491.
(30) Radtke, S. F., American Metal Market, Lead and Zinc Supplement, 87(70):
9A-10A, 19A, April 10, 1979.
(31) American Metal Market, 87(44):29, April 5, 1979.
(32) Environment Reporter, 2048, March 2, 1979.
(33) Engineering and Mining Journal, 180(3):239, March, 1979.
The Mining Record, 90(15):!, April 11, 1979.
(34) Engineering and Mining Journal, 180(4):35, 37, April, 1979.
(35) Sellings' Mining Review, 68(8) :16, February 24, 1979.
(36) The Mining Record, 90(16):!, April 18, 1979.
The Mining Record, 90(15):8, April 11, 1979.
(37) Sellings' Mining Review, 68(10):12, March 10, 1979.
(38) Skillings' Mining Review, 68(10)12, March 10, 1979.
(39) Engineering and Mining Journal, 179(11):241, November, 1978.
(40) The Mining Record, 90(9):6, February 28, 1979.
The Mining Record, 90(14):5, April 4, 1979.
(41) Engineering and Mining Journal, 179(11):268, November, 1978.
(42) New Mexico Paydirt, (22):52, March, 1979.
Skillings1 Mining Review, 68(11):14, March 17, 1979.
(43) The Mining Record, 90(9):1, February 28, 1979.
The Mining Record, 90(12):!, March 21, 1979.
The Mining Record, 90(15):8, April 11, 1979.
-------
(44) The Mining Record, 90(12):3, March 21, 1979.
(45) The Mining Record, 90(9):1, February 28, 1979.
(46) The Mining Record, 90(15):7, April 11, 1979.
(47) Man, A., American Metal Market, Aerospace Metals and Machines
Supplement, 87(49):14A-16A, March 12, 1979.
(48) U.S. Bureau of Mines, Minerals and Materials, 9, February, 1979,
(49) American Metal Market, 87(75):6, April 17, 1979.
(50) U.S. Bureau of Mines, Minerals and Materials, 48, March, 1979.
RECENT PATENTS
4,141,159
METHOD AND APPARATUS FOR DEEP SEA MINING
Wilford V. Morris and George W. Sheary, III, assignors to Summa Corporation,
Las Vegas, Nevada.
(This describes a sea-bed mining system comprising a nodule collector and
crusher and a pneumatic lift system. Crushed nodules are pumped to a storage
line from which they are fed batchwise to a pneumatic lift feed chamber from
which they are pumped to the surface.)
4,141,721
METHOD AND APPARATUS FOR COMPLEX CONTINUOUS PROCESSING OF POLYMETALLIC RAW
MATERIALS
Jury F. Frolov and 15 other USSR inventors.
(A continuous smelting process for complex zinc sulfide/oxide ores is claimed.
This involves reduction of the smelt by a plasma at 4,000 to 5,000 C, with
the bath surface maintained at 1,500 to 1,600 C!)
4,141,804
PROCESS FOR ELECTROWINNING METAL FROM METAL BEARING SOLUTIONS
Michael M. Avedesian and Anthony P. Holko, assignors to Noranda Mines Limited,
Toronto, Canada.
(This describes the apparatus and electrolyte handling system for fluidized
bed electrowinning of metal values from leach solutions.)
4,142,952
COPPER EXTRACTION WITH SALICYLALDOXIME-P-NONYLPHENOL MIXTURES
Raymond F. Dalton, assignor to Imperial Chemical Industries Limited, London,
England.
-------
RECENT PATENTS (Continued)
(A process is claimed for extracting and electrowinm'ng copper using this
class of lixiviant.)
v
4,143,865
FLASH SMELTING FURNACE
Thomas N. Antonioni, assignor to The International Nickel Company, Inc., :
New York, New York.
(A method for cooling a critical area--matte-slag interface—of refractories
in the Inco flash smelting furnace is claimed.)
4,144,055
METHOD OF PRODUCING BLISTER COPPER
Stig A. Petersson and Bengt S. Eriksson, assignors to Boliden Aktiebolag,
Stockholm, Sweden.
(The Boliden TBRC process for blister copper production is described and
claimed. This includes operations of the TBRC, slag-treating, and converter
units.)
4,144,056
PROCESS FOR RECOVERING NICKEL, COBALT AND MANGANESE FROM THEIR OXIDE AND
SILICATE ORES
Paul R. Kruesi, assignor to Cato Research Corporation, Wheatridge, Colorado.
(Heating oxidic or silicate ores of manganese, nickel, and cobalt in the
absence of air,with ferric chloride and other chloride salts to suppress
volatilization,converts the desired metal values to soluble chlorides.)
4,144,310
HIGH SLURRY DENSITY SULFIDIC MINERAL LEACHING USING NITROGEN DIOXIDE
Theodore C. Frankiewicz and Robert E. Lueders, assignors to Kennecott
Copper Corporation, New York, New York.
(Complex sulfidic base metal concentrates in a highly loaded slurry form are
oxidized to return at least one Teachable metal value. Nitric oxide is the
oxidant, and is restored to its higher valence form by the presence of
oxygen.)
4,145,187
TREATMENT OF MATERIAL WITH HYDROGEN CHLORIDE
Raymond E. Oliver and George McGuire, assignors to Matthey Rustenburg
Refiners (Pty.) Ltd., Johannesburg, South Africa.
(A hydrogen-chlorine burning reactor is claimed for the conversion and sepa-
ration of silver from silver-containing base metal concentrators. The
reacted product is leached to remove soluble base metal chlorides.)
-------
RECENT PATENTS (Continued)
4 145 212
PROCESS FOR RECOVERING SILVER AND OPTIONALLY GOLD FROM A SOLID STARTING
MATERIAL CONTAINING SAID METALS
Fernand J. J. Bodson, assignor to Societe des Mines et Fonderies de Zinc
de la Vielle Montagne, Angleur, Belgium.
(Team silver and gold resources, at least partly sulfidic, when oxidation
leached in an aqueous thiouria solution, are converted to soluble complexes
which can be cemented as a 90+ percent precious metal product.)
4,145,398
BAUXITE DIGESTION BY CAUSTIC ALKALI WITH IMPROVED HEAT TRANSFER IN TUBULAR
REACTORS
Ludolf Plass, assignor to Vereinigte Aluminum-Werke A. G., Bonn, Federal
Republic of Germany.
(This claims and describes the VAW energy-conserving tubular reactor for
bauxite treatment—see Technology and Trends item 3.2.9/1.)
4,146,389
THERMAL REDUCTION PROCESS OF ALUMINUM
Bela Karlovitz, assignor to Bela Karlovitz and Bernard Lewis, both of
Pittsburgh, Pennsylvania.
(This appears to be a plasma reactor/condenser system for the suspended-
particle reduction of metal oxides.)
4,146,572
SIMULTANEOUS EXTRACTION OF METAL VALUES OTHER THAN COPPER FROM MANGANESE
NODULES
Paul H. Cardwell and William S. Kane, assignors to Deepsea Ventures, Inc.,.
Gloucester Point, Virginia.
(A variant of U.S.P. 4137291. A reduction roast converts the copper values
to the cuprous state whereby they are not removed by ammonium complexing,
and thus can be separated from other metal values in the first-stage leach.)
4,146,573
RED MUD TREATMENT
James Kane, assignor to Nalco Chemical Company, Oak Brook, Illinois.
(Less than 0.02 percent of an organic ammonium salt mixed with Bayer process
red mud waste will "solidify" the red mud.)
4,147,390
NODULE DREDGING APPARATUS AND PROCESS
Jacques Deliege, Michel Giot, Vsevolod Obolensky, and Marc Lejeune, assignors
to Union Miniere S.A., Brussels, Belgium.
-------
RECENT PATENTS (Continued)
(A sea nodule dredging apparatus is claimed.)
4,147,614
AQUEOUS MIXTURE OF DIESEL OIL, PINE OIL AND DIAMINE FOR CONDITIONING OF
CRUSHED MAGNESITE ORE IN MAGNETIC BENEFICIATION PROCESS
Theodor Gambopoulos and Antonios Frangiskos of Athens, Greece.
(Conditioning of crushed magnesite ore in an aqueous mixture of diesel oil,
pine oil, and a diamine allows the occlusion of particulate ferromagnetic
material on gangue material, but not on the magnesite mineral. Gangue may
be magnetically separated, leaving a high-quality magnesite concentrate.)
4,147,644
COLLECTOR COMBINATION FOR NON-SULFIDE ORES
Samuel S. Wang, Eugene L. Smith, Jr., and Ernie F. Huliganga, assignors to
American Cyanamid Company, Stamford, Connecticut.
(Nonsulfide ores may be treated with a fatty acid and an anionic perfluoro-
alkyl compound as claimed to permit concentration by flotation.)
4,148,630
DIRECT PRODUCTION OF COPPER METAL
Charles Arentzen, assignor to The Anaconda Company, Denver, Colorado.
(Claims a method for pneumatically injecting a slurry of sulfidic copper
into a molten bath in a process for continuous, autogenous smelting and
reduction to produce metallic copper.)
4,148,632
TREATMENT OF DISSOLVED BASIC NICKEL CARBONATE TO OBTAIN NICKEL
Willie Seibt and Donald R. Weir, assignors to Sherritt Gordon Mines Limited,
Toronto, Canada.
(This appears to be a variant of the Sherritt process for nickel in which
secondary nickel values are recovered from the unleached residue.)
4,148,813
NICKEL AND COBALT SOLVENT EXTRACTION WITH MERCAPTIDES
Alkis S. Rappas and J. Paul Pemsler, assignors to Kennecott Copper
Corporation, New York, New York.
(Claims mercaptides as the complexing agent in solvent extraction of nickel
and/or cobalt from hydrometallurgical leach liquors.)
4,148,815
AMINO-THIOL NICKEL AND COBALT SOLVENT EXTRACTION
Alkis S. Rappas and J. Paul Pemsler, assignors to Kennecott Copper
Corporation, New York, New York.
-------
RECENT PATENTS (Continued)
(A variant of 4,148,813.)
4,148,816
ALKALI METAL MERCAPTIDE SOLVENT EXTRACTION OF COBALT, NICKEL
Alkis S. Rappas and J. Paul Pemsler, assignors to Kennecott Copper
Corporation, New York, New York.
(A variant of 4,148,813.)
4,148,862
HYDROMETALLURGICAL TREATMENT OF SOLUBLE SILICATE-BEARING ZINC MATERIALS
Sigmund P. Fugleberg and Jaakko T. I. Poijarvi, assignors to Outokumpu Oy,
Outokumpu, Finland.
(Physical characteristics of the silicate burden in siliceous zinc ores are
controlled by the rate of addition of silicates during sulfuric acid leach-
ing to promote good settleability and filtration properties.)
4,149,880
RECOVERY OF COPPER FROM ARSENIC CONTAINING METALLURGICAL WASTE MATERIALS
John D. Prater and Barry A. Wells, assignors to Kennecott Copper Corporation,
New York, New York.
(This claims a method for processing toxic flue dusts and or refinery sludges
from copper smelters. Oxidative pressure leaching with sulfuric acid is
conducted under conditions to solubilize copper and some arsenic, with most
of the other toxicants and precious metals remaining as solids. Downstream
cementation with iron recovers cement copper without the liberation of
arsine. The iron-laden raffinate containing arsenic is recycled.)
4,149,947
PRODUCTION OF METALLIC LEAD
John C. Stauter and William K. Tolley, assignors to UOP, Inc., Des Plaines,
Illinois.
(In a variant of USP 4124461 for converting crystallized lead halide first
to carbonate, then to lead fluosilicate from which it is electrowon, the
carbonate conversion may be accomplished by treating the halide with C02-)
4,150,091
MANGANESE ORE LEACHING PROCESS
Henry J. Peterson, assignor to Sun Ocean Ventures, Inc., Radnor,
Pennsylvania.
(This claims an improved HC1 leaching process for treating sea nodules. By
regulating pH and chlorine pressure, the desired base metals may be con-
verted to soluble chlorides, leaving insoluble manganese.)
-------
FOREIGN TECHNOLOGY
(The articles listed are from foreign language journals with titles and
abstracts printed in English.)
Vogg, H., H. Braun, and A. Lubecki, Ore Prospecting, Exploring, and Processing
Techniques by Use of Radionuclides, Erzmetall, 32(j);8-12, January, 1979.
(German)The application of radionuclide excited X-ray fluorescence, and
neutron activation analysis using neutron isotope sources, with a view to in
situ borehole analysis and on-line analysis of slurry systems is reviewed.
Wobking. H., and H. Worz, Theoretical Bases and Practical Impacts of the PCR
Process in the Copper Refinery Electrolysis, Erzmetall, 32(2):53-57, February,
1979. (German) The operational results of the Brixxlegg-developed PCR pro-
cess in relation to anodic passivation is reported.
Zanardi, G. M., Exploitation and Extraction of Copper Oxide Ores by
Hydrometallurgical Methods in Cerro Verde, Peru, Erzmetall, 32(2):58-63.
February, 1979.(German)Mining and processing of the copper oxide ore cap-
ping the sulfide ore body includes crushing, heap leaching, ion exchange, and
electrolysis. The geology, mineralogy, and costs are also discussed.
Galitis, N., M. Clement, and R. Bertram, Investigations for the Leaching of
Sulfidic Copper Ores with Agitation Reactors, Erzmetall. 32(2):74-77,
February, 1979.(German!The influences of leachant/electrolyte concentra-
tion, the solid electrolyte rate, and temperature on graphite, titanium/
platinum, and lead anodes and copper cathodes during simultaneous pocked bed
reactor leaching and electrowinning of Mitterberg chalcopyrite and Rammelsberg
copper concentrate are discussed.
Trytko, 0.. V. Masek. Emissions and Immissions of SOg in Metallurgical Plant.
Hutnicke Listy, (2): 127-130, February, 1979. (Czechoslovakian) This presents
a study of workplace and ambient S02 levels associated with a "metallurgical"
plant. For each ton of material "rolled or forged", 22 Kg S02 was emitted
(apparently from the power plant. The abstract does not otherwise describe
the "metallurgical plant").
Solozhenkin, P. M., and R. D. Kupeyeva, Investigating the Sorption Layer on
Galena and Sphalerite and of Surface Conditions for Xanthogenate Desorption,
Izvestiya Vuz Tsvetnaya Metallurgiya, (6):9-14, 1978.(Russian, title only)
Alkatsev, M. I., Production of Standard Conditioning Metal Deposits by
Cementation, Izvestiya Vuz Tsvetnaya Metallurgiya, (6)-.21-27, 1978. (Russian,
title only)
Kutvitsky, V. A., V. V. Mechev. and V. Yu. Taskin, Copper Refining from Nickel
by Directed Crystallization in the Magnetic Field, Izvestiya Vuz Tsvetnaya
Metallurgiya, (6):27-31, 1978.(Russian, title only)
-------
Zapuskalova, N. A., and E. V. Margulis, The Influence of Separate Parameters
on Hydrolytic Precipitation of Iron, Copper, and Zinc from Sulfate Solutions,
Izvestiya Vuz Tsvetnaya Metallurgiya, (6) .-31-37, 1978. (Russian, title only)
Slutsky, I. Z.. V. P. Kiseiyov. E. S. Yakubovski. and M. G. Tsypkin, Pilot
Plant Test of Electrolytic Refining the Aluminum-Silicon Alloy Resulting in
Aluminum Production, Izvestiya Vuz Tsvetnaya Metal!urgiya, (6):59-64, 1978.
(Russian, title only)
Markov. Yu. F.. V. A. Ivanov, and A. F. Gavrilenko, On Controlling Nickel
Precipitation in the Flow Sheet of the Autoclave Treatment of the Nickel
Magnetic Pyrite Concentrates. Izvestiya Vuz Tsvetnaya Metal!urgiya, (6):115-
121, 1978. (Russian, title only)
Sergiyevskaya, Ye. M., and I. A. Vorobyova, Thermodynamics and Kinetics of
Nickel Oxide Dissolving in the Sulfuric Acid Solution, Izvestiya Vuz
Tsvetnaya Metallurgiya, (6):126-129, 1978.(Russian, title only)
Kravets. M. V., and E. V. Margulis, Effect of Lead Compounds on the Indium
Behavior in Zinc Sulfate Solutions. Izvestiya Vuz Tsvetnaya Metal!urgiya,
(6J.-129-131, 1978. (Russian, title only)
Novozhenov, V. M., T. I. Tataurova, and S. I. Kuznetsov, To the Question of
Reducing the Chemical Losses of Alkali and Aluminum Oxide at Alumina
Production by Bayer Method,, Izvestiya Vuz Tsvetnaya Metal!urgiya, (1):40-45,
1979.(Russian, title only)
Volkov, V. V., and N. I. Eremin, The Thickening Rate Calculation of Nephelite
Slime in Aluminate Solution. Izvestiya Vuz Tsvetnaya Metallurgiya, (l):45-48.
1979.(Russian, title only)
Pol.yakov, P. V.. V. M. Mozhayev. V. V. Burnakin. V. A. Kryukovsky. and
V. E. Nikolayenko. On Bubble Liberation at Cryolite-Alumina Bath
Electrolysis. Izvestiya Vuz Tsvetnaya Metallurqiya, (1):55-61. 1979.
(Russian, title only)
Leonova, M. L., B. L. Egorov, M. D. Ivanovsky, and M. A. Meretukov, Effect
of Heat Treatment on Gold Extraction from Grinding Powders, Izvestiya Vuz
Tsvetnaya Metallurgiya, (l):75-78, 1979.(Russian, title only)
Ivanov. V. A., L. B. Berliner, V. I. Kesisoglu, M. P. Shapirovsky, and
E. I. Sterenberg, The Mathematical Description of Basic Regularities'!^
Copper-Nickel Concentrate Sinter Roasting, Izvestiya Vuz Tsvetnaya
Metal!urgiya, (1):128-134, 1979.(Russian, title only)
Mukhina, T. M., and N. V. Krupkin, The Analysis of the Actual and Planned
Efficiency of Introducing New Equipment into Non-ferrous Metallurgy,
Izvestiya Vuz Tsvetnaya Metallurgiya, (1):134-139, 1979.(Russian, title
only)
-------
Molnar, I., Some Aspects of the Selection of the Technical Solution for a
New Aluminum Electrolysis Plant, Banyaszati Es Kohazati Lapok-Kohaszat,
112(l):33-37, January, 1979.(Hungarian) An analysis of the energy and
environmental protection aspects in the design of an aluminum electrolysis
plant is presented.
Zambo, J., Technical Development in the Hungarian Alumina Refineries,
Banyaszati Es Kohazati Lapok-Kohaszat, 112(1):43, January, 1979.(Thlngarian)
Energy economics and utilization of red mud will result in a major change.in
equipment for new Bayer plant and equipment design.
Paschen. P., Metal Refining by Selective Evaporation. Metal!, 33(2):137-140,
February, 1979.(German, title only)
Stepniak, B. P., Copper Extraction by 2-Hydroxy-5-Nonylobenzoic Aldoxime,
Rudy Metale, 24(1):2-5, January, 1979.(Polish)A higher degree of copper
extraction in a snorter period of time is obtained by the use of 2-hydroxy-
5-nonylobenzoic oxime rather than 2-hydroxy-5-nonylobenzophenomic oxime.
Kuczumow, A., M. Majdan, and S. Radzki, The Role of Rare Earth Elements in
Chemistry and Technology, Part 1. Resources,Winning and Distribution, Rudy
Metale, 24(l):24-28, January, 1979.(Polish)Methods of isolating rare-
earth elements by classical and fractional methods applied in industrial
processes are reviewed. Local resources of rare-earth elements are also
discussed.
Kubicki, J., and S. Kuczkowska, The Effect of Ammom'um Sulfate on
Intensificationof the Copper Electrorefining Process, Rudy Metale, 24(2):
56-60, Februaryj 1979. (Polish) Ammonium sulfate added to copper electro-
lytes containing nickel improves the electrochemical refining process.
Krupkowa, P., Cementation Purification of Industrial Zinc Sulfate Solutions,
Rudy Metale, 24(2):65-75, February, 1979.(Polish)This discusses cementa-
tion purification of zinc electrolytes using zinc dust. In the first stage,
cobalt, nickel, and other impurities are eliminated at 90 C with high lead-
antimony zinc dust, while small quantities of cadmium are eliminated at
60 C in the second stage with "columnar" zinc dust.
Wanielista, K.. Algorithm to Determine Criteria for Recoverability of Ore
Deposits. Rudy Metale. 24(2):76-82, February, 1979.(Polish)A procedure
is proposed to determine the minimum ore content of a deposit necessary for
industrial utilization. Profitability is emphasized.
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US EPA TECHNOLOGY AUV TREWDS ABSTRACT
for the NONFERROUS METALS INDUSTRY
Index
Page
1.2.9
3
Subject: Conditioning Effects of Flotation of Finely Divided
Non-Sulphide Copper Ore
The effect of several physical and chemical parameters on the grade and
percent of copper recovery from nonsulfide ores during the flotation process
are described in this paper.
The ore--a siliceous, malachite-prominent ore from Zaire—was ground to
60 percent-9y. The following parameters were varied during the experiments:
(1) stirring speed, (2) stirring time, (3) quantity of collector (K-amyl
xanthate), (4) quantity of sodium sulfide, (5) feed grade of copper, (6)
pulp density, and (7) the quantity of calcite, dolomite, or both, added.
The results are shown graphically.
The results showed that flotation was successful only when the fines
were conditioned at stirring speeds above 900 rev/min. A minimum satisfac-
tory stirring time of 40 minutes at 2060 rev/min was defined. A stirring
time of 60 minutes at 2060 rev/min was chosen to be used in most experiments.
The minimum quantity of K-amyl xanthate necessary to obtain efficient separa-
tion was found to be 3.75 kg/t. The addition of sodium sulfide was found to
be detrimental to the flotation process. As the feed grade of copper
decreased from 8.5 to 0.4 percent, only a 10 percent decrease in recovery and
a 12 percent reduction in grade was noted. However, there was a significant
decrease in both grade and recovery when pulp density was increased to about
10 percent in weight. No effect on grade or recovery was found when calcite
or dolomite made up 0-25 percent of the gangue.
Reference: Rubio, J., Institution of Mining and Metallurgy, Section C,
87:C284-287, December, 1978.
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USEPA TECHNOLOGY AMP TREWPS ABSTRACT Index 1 3.2
for the NONFERROUS METALS INDUSTRY Page 17
Subject: Metal and Sulfur Recovery from the Copper Sulphide
Precipitate in the Sherritt Gordon Process
The Sherritt Gordon pressurized ammonia leach process for the extrac-
tion and production of nickel and cobalt is employed at the Western Mining
Corporation's Kwinana plant in western Australia. A by-product of this
operation is about 2,600 annual tonnes of copper sulfide precipitate produced
in the "copper boil" stage of the S.G. process. This precipitate carries
with it some entrained nickel and cobalt values, and also significant quanti-
ties of precious metals (Pd, Ag, Au, Pt). In present practice, the CuxS
product is washed, filtered, dried, and toll-smelted and refined by the
Electrolytic Refining and Smelting Co. Pty Ltd, N.S.W., Australia. However,
the freshly precipitated and repulped CuxS product is a very reactive, finely
divided slurry that is ideally suited to further hydroprocessing. The eco-
nomics of further processing will depend on the scale of operations, and have
not yet been determined.
This technical article describes laboratory studies of the further
hydrometallurgy of the copper boil product to recover metallic copper, return
nickel and cobalt values to the main S.G. process circuit, and to concentrate
precious metal values for refining. The preferred hydroprocess involves a
low-temperature (<40 C), moderate-pressure (30 psig) oxidizing NH3/(NH4)2$04
leach, which releases sulfur in elemental form. Precious metals precipitate
as sulfides when the liquor is stripped of NH3- S02 treatment and heating
of the stripped liquor crystallizes copper as Chevreul's salt, and nickel and
cobalt values (still in solution) are returned to the S.G. circuit. Aceto-
nitrile treatment under nitrogen at 55 C yields cuprous sulfate, which is
steam disproportionated to yield copper powder of high purity. Various
alternatives or options were also investigated, and are discussed. Elements
of the process were individually examined, and a 500 gm demonstration was •
conducted. A flow sheet and material balance are presented in the article.
RpfFilmer, A. 0., A. J. Parker, M. Ruane, and L. G. B. Wadley,
e' Proceedings Australasian Institute of Mining and Metallurgy,
(268):39-46, December, 1978.
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USEPA TECHNOLOGY MK> TRENVS ABSTRACT Index 1.3.2
for the NONFERROUS METALS INDUSTRY Page 18
Subject: stnriipq nn Segregation of Roasted Hopper Concentrates:
Separate Na and C1 Additions and Effects of Sulfate
The segregation process for winning metallic copper from dead-roasted
concentrates involves halide-activated reduction by carbon, with CO? and '
copper as the ultimate reaction products. This paper reports expanded
laboratory studies of the role of halide source, reductant, and influence of
residual sulfur (as sulfate) on the rates and extent of reaction at about
700 C.
Sodium chloride is the usual halide source. When cuprous chloride is
used as the halide source, reaction rate is improved, but with petroleum coke
as the reductant, a continuous layer of segregated copper blocks further
reduction when only about half the copper value has been recovered. The
coaddition of CuCl and active carbon (e.g., coconut charcoal) results in
essentially complete reaction at suitably high rates.
Coadditions of CuCl and N32C03 with mol ratios of Na:Cl varying from
0.5 to 2 were studied. Excess sodium slowed the reaction, and excess
chloride blocked it. At a molar ratio of 1, reaction proceeded as if Nad
had been the additive—moderate rate; complete reaction.
Experiments with calcine containing a higher sulfate content (as CuO •
CuS04?) suggested that the activity of petroleum coke is enhanced by a
moderate amount of residual sulfate in the calcine. The presence of sulfate
increases the reaction rate. However, the presence of halide is still
necessary to allow reduction to go to completion.
(The dead-roasting/segregation process offers potential for abating the
economic impact of S02 removal from conventional copper smelting. These .
reported experimental results may help to define the potential commercial
operation of the segregation process.)
Reference: Marcuson, S. W., Inst. of Min. and Met. Trans. C., 87:(26)-C265,
December, 1978.
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USEPA TECHNOLOGY AND TRENDS ABSTRACT Index 1.3.2
for the NONFERROUS METALS INDUSTRY Page 19
Subject:utah Copper and the $280 Million Investment in Clean Air
Kennecott's essentially new, 270,000 st/yr, $280-million copper smelter
at Garfield, Utah, has been the subject of prior Awareness reports (see
T&T 1.3.2/1, 1.3.2/15, selected news articles). This editorial feature from
E&MJ presents an excellent overall summary of this new, two-stage, Noranda-
process-based facility.
Two, 1,650 stpd (dry concentrate) reactors and three converters (exclud-
ing a spare reactor and a spare converter) form the working heart of the sys-
tem. However, these are dwarfed by the dimensions of the gas handling and
processing system, which require in excess of 33,000 horsepower to move the
gas through various stations, finally exhausting at a rate of 1.3 million
scfm through the new, 1,200-foot stack.
A lucid flowsheet and schematic layout are included in the article.
The presentation centers about reasons for the original selection of the new
system, reactor and converter processes, slag reprocessing, gas handling
(including the four single-contact acid plants, scrubbers, mixers, cyclones,
etc.), process control and data management.
This large investment by Kennecott was directed toward achieving the
then-mandated air regs (which EPA subsequently intensified, much to
Kennecott's chagrin). The smelter now captures 86 percent of the input sul-
fur, compared with a 55 percent prior capture.
Production capacity is about the same as previously. Despite consider-
able savings in direct process energy consumption, net energy savings are
essentially offset by energy demands for pollution control and essential
ancillary process demands.
Reference: Dayton, S.} Engineering and Mining Journal, 180(4):72-83,
April, 1979.
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USEPA TECHNOLOGY AMP TREWS ABSTRACT Index 1.3.
for the NONFERROUS METALS INDUSTRY Page 20
Subject: Aqueous Oxidation of Chalcopyrite in Hydrochloric Acid
Process kinetics for the oxidation leaching of chalcopyrite by hydro-
chloric acid are explored in the laboratory, as described in this paper.
The effects of oxygen pressure (to >2000 k Pa), temperature (90-130 C), agi-
tation, and acid strength (CuFeS2'.HCl molar ratios from about 0.1-2) were
explored. Optimum conditions were defined to be 110 C, 2010 k Pa of 02»
highly agitated, and molar ratio of chalcopyrite/HCl of 0.8-0.9.
The reaction rates are faster than with the similar ^$04 process.
Under optimum conditions, copper recovery was found to be only about 70 per-
cent, but the pregnant liquor was very low in iron, which hydrolyzed and
precipitated as F6203 or FeOOH. Increased acid strengths resulted in the
generation of copper ferrite or contamination of the leachate with iron. The
sulfur reports as free sulfur, which at temperatures greater than 110 C was
held responsible for decreased reaction rates due to coating of the particles
with molten sulfur.
(This report of academic investigation was not concerned with economic
or corrosion problems.)
Reference: Habashi, F., and T. Toor, Metallurgical Transactions, 10B(1):49-
56, March, 1979.
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USEPA TECHNOLOGY ANP TRENDS ABSTRACT Index 1.3.2
for the NONFERROUS METALS INDUSTRY Page 21
Subject: Afton's Copper Smelter Proves Economic at 27.000 Tons
Yearly
The Afton copper smelter at Kami oops, British Columbia is the first com-
mercial use of the TOD Blown Rotary Converter (TBRC) Process for the com-
bined smeltina and converting of concentrates. (See the "In the News Section"
of the September, 1977 issue of the Technical Awareness Bulletin). In
1979, 27,000 tons of 99.6 percent copper product is expected to be shipped
from this $25 million smelter.
The ore, native copper, bornite (CusFeS/j), chalcocite (Cu2S), and chal-
copyrite (CuFeS2) is crushed and processed to produce two distinct types of
copper concentrate: a high grade metallic concentrate containing 80-90 per-
cent copper, and a conventional flotation concentrate of 55-60 percent cop-
ner with 35 percent sulfur. The conventional concentrate is mixed with iron
and limestone,then fed into the converter where it is melted by an oxy-fuel
flame. This is later replaced by oxygen blowing into the molten bath; how-
ever, the major quantity of oxygen is supplied above the slag, in a highly
exothermic reaction attaining 1350 C. Tilted at 15-20 degrees, the furnace
rotates up to 40 rpm to protect the refractories from the intense heat.
After pouring of the first slag, smelting resumes, then the slag is again
removed. Then the native copper concentrate is charged into the converter
and the procedure repeated,blowing the charge to blister copper (99.6 percent
Cu).
All exhaust gas is collected and processed in a four-stage dual-alkali
scrubbing system. The first stage reduces the temperature by waterspray to
340 C and removes particulates in an electrostatic precipitator; the second
stage contacts the qas with sodium sulfite solution to absorb the sulfur
dioxide; the third cools the gas to condense any mercury present; and the'
fourth stage includes the use of a proprietary filter. The sulfur control
system also includes sulfur dioxide storage so that a continuous flow is
available for a sulfuric acid plant.
(This news article may not be wholly consistent in a technical sense;
several questions arise. However, it presents a useful description of the
overall operation at Kami oops.)
Reference: Canadian Chemical Processing, 63(2):22-24s March 21, 1979.
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USEPA TECHNOLOGY AND TREWPS ABSTRACT Index 1.3.4
for the NONFERROUS METALS INDUSTRY Page 3
Subject: Some Aspects of Inn Exchange in Copper Hydrometallurgy
Ion exchange is not at present a leading copper hydrometallurgy process
but is an attractive alternate to the solvent extraction process, as solvent
extraction is usually limited to liquors containing more than 1 g/1 of
copper. The purpose of this reported work was to compare the copper hydro-
metallurgical properties of (1) chelating resins, 8-hydroxyquinoline, sali-
cyloldoxime, and XF-4196; (2) as-received impregnated foams, Kelex 100 (33
percent) and LIX 65N (23.5 percent); and (3) amberlite beads impregnated with
Kelex 100 (33 percent) and LIX 65N (33 percent).
The ion exchange systems were examined for their high capacity for the
metals of interest, high selectivity, and rapid equilibration. Total
capacity was reported as copper uptake at pH 2 from a dilute sulfuric acid
medium or at pH 4 from a sodium acetate/acetic acid buffered solution.
Selectivity ratios were reported as the respective weights of copper and iron
from solutions which were equimolar in copper (II) and iron (III). Kinetics
or rates of equilibrium were time values (t-|/2J taken for the metal to
occupy one-half of the available sites on the exchanges.
Results of the work show that only chelating resins are viable; impreg-
nates have too low a capacity and high bulk volumes; impregnated beads are
too high in cost. However, the impregnates were far superior to the chelat-
ing resins in selectivity. Solution to the problem appears to be an attempt
to improve the selectivity of the XF-4196 (See this issue "Technology and
Trends" 1.3.9, page 10) -
It seems unlikely that ion exchange will replace solvent extraction in
copper winning. However, ion exchange could complete the removal of copper
from the solvent extraction raffinate and ensure complete copper removal
from discharge liquors where excess iron must be bled from the system. A
flow sheet of a process incorporating this idea is included in the paper.
Reference: Vernon, F., Hydrometallurgy, 4(2):147-157, 1979.
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Subject: Copper-Selective Ion-Exchange Resin with Improved Iron
Rejection
A description of a new chelating ion-exchange resin with higher copper
leaching efficiency and lower affinity for iron is given in this paper.
The resin, XFS 43084, (N-(2-Hydroxypropyl)-pecolyamine) has a several
fold increase in copper/iron (III) selectivity and retains the copper absorp-
tion characteristics of the original XFS 4196 resin. Elution of copper can
be done with 2N sulfuric acid at 15 times the rate of absorption. Aqueous
ammonia can also be used to elute the copper. Equilibrium data for metal-
ion absorption were determined, as well as the kinetics of metal exchange.
These data are presented in eight graphs, together with a table comparing the
properties of XFS 43084 with those of XFS 4196, the original ion-exchange
resin.
Reference: Grinstead, R. R., Journal of Metals, 31(3):13-16, March, 1979.
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USEPA TECHNOLOGY AMP TRENDS ABSTRACT
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Index
Page
1.3.9
11
Subject: A New Hydrometallurgical Method for the Processing of
Copper Concentrates Using Ferric Sulfate _
This paper describes laboratory tests and a 3-week pilot-plant opera-
tion of a process for electrolysis of copper from leaching solutions in
diaphragm cells, with double-stage anodic recovery of the spent leaching
agent.
Leaching is carried out at 90-95 C with copper sulfides decomposed as
follows:
Chalcocite, Cu2S + Fe2(S04)3 -> CuS04 + 2FeS04 + CuS
Covellite, CuS + FeJSCLK + CuSO. + 2FeSO. + S
L. T1 0 T1 T"
Bornite, CugFeS4 + 6Fe2(S04)3 -> 5CuS04 + 13FeS04 + 4S
Chalcopyrite, CuFe$2 + 2Fe2(S04)3 -»• CuS04 + 5FeS04
2S
The extraction of copper reaches 96 percent in 100-120 minutes except
for chalcopyrite. After 4-6 hours, the yield of copper into solution may
reach 97-98 percent. About 70-74 percent of the product is treated to
crystallize FeS04 • 1^0 by heating the solution to 170-175 C. The pregnant
liquor is then delivered to the cathode compartment of the electrolyser where
copper is won in one or two stages. The spent catholyte is combined with a
portion of previously crystallized FeS04 • h^O and delivered to the anode
compartments where the initial anodic oxidation of FeS04 to Fe2(SQ4)3 takes
place. This ferric sulfate is utilized in the second leaching stage. The
spent leaching agent in the second leaching stage is reoxidized anodically
before the first leaching stage.
The essential feature of the process is the double anodic recovery of
the leaching agent, decreasing the quantity of ferric ions necessary to
leach a given quantity of copper to about 50 percent of the required stoi-
chiometric ferric ions to leach the same amount of copper.
Electrolysis of the copper solutions, containing a high concentration of
ferrous sulfate after leaching the concentrate, is made possible by the appli-
cation of diaphragms. These diaphragms (0.5 mm thick PVC) restrict mixing
between iron-depleted catholyte and recharged anolyte.
Silver and lead occurring with the copper are not leached by the ferric
sulfate solutions, but require leaching with concentrated brine solution in
an additional leaching process.
Reference: Letowski , F.5 B. Kolodziej, M. Czernecki , A. Jedrczak, and
Z. Adamski, Hydrometallurgy, 4(2) : 169-1 84, 1979.
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Subject: A Hydro-Saline (Chloride-Ion) Cycle for Copper-Bearing
Haste Leaching
This paper describes the laboratory experiments conducted to determine
the characteristics of KC1 and NaCI solutions in the leaching of low-grade,
cooper-bearing wastes and the subsequent recovery of the copper by cementa-
tion onto aluminum and aluminum-alloy scrap and beverage cans.
KC1 and NaCI were chosen for the tests because of their natural abun-
dance. Sea water, geothermal brines, brackish-water in underground saline
reservoirs, and in some ground-water in the U.S. southwest, all could be
utilized in leaching the low-grade copoer-bearing waste.
After leaching, the copper solution is recovered by cementation on a
more electropositive metal, which typically has been iron. Recently, alu-
minum has been suggested for this use as the copper recovery is more effi-
cient with aluminum (although a higher temperature is needed). At least
75 ppm Cl" in solution is necessary to break down the protective aluminum
surface oxide film. The efficiency of cementation of copper on aluminum
doubles between 30 and 40 C. The efficiency of copper cementation on iron is
less than half that of aluminum. Cementation on aluminum is ideally suited
for dilute solutions, the cement copper is purer than that cemented by iron,
and the cement is physically easier to remove from aluminum.
The proposed hydro-saline leach cycle and cementation on aluminum may
not be generally economical at the present time; however, in the next decade
with abundant aluminum scrap and the price difference between iron and alu-
minum decreasing, this process will become more attractive and profitable,
in the authors' opinion. (This is no;t this reviewer's opinion.)
Reference: Murr, L. E., V. Annamalai, and P-C. Hsu, Journal of Metals,
31(2):26-32, February, 1979.
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USEPA TECHNOLOGY AMP TRENDS ABSTRACT
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Index
Page
1.3.9
13
Subject:
Noranda Reprocesses Slags to Enhance Recovery
In 1973, Noranda Mines Limited, at its Quebec copper smelter, began
slag-milling.
The slag at Noranda is cooled slowly to promote agglomeration of metal
values to promote agglomeration of metal values to permit easier recovery.
The slag is then crushed to -325 mesh and floated with sodium amyl xanthate
and Dow 1263. A concentrate containing 35-40 percent copper is produced.
Only 0.28 percent copper reports to the tailings.
At Inco, a similar procedure of flotation of slags produces concentrates
containing 15-50 percent copper, with only about 0.25 percent copper lost in
the tails.
Reference: Canadian Chemical Processing, 63(2):24, March 21, 1979.
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US EPA TECHNOLOGY ANP TREWPS ABSTRACT Index
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1.4.9
2
Subject: Recent Advances in Copper Electrowinning
Interest in hydrometallurgical processes is increasing, with nearly.
15 percent of copper production at the present time by this method. This
may increase to about 40 percent by 1985. This paper reviews the advances in
copper electrowinning in detail. The first section is devoted to methods
used to permit copper to be electrowon at high current densities, with the
following section detailing the consequences of this operation on the various
components of the system. The design and development of electrochemical
cells is considered, with effects of impurities, the role of additives, and
the morphology and growth of copper deposits discussed. A bibliography of
125 articles is included.
The depletion of copper sulfide ores is necessitating the use of lower
grade or mixed oxide-sulfide ores, and material from old tailings dumps.
Solvent extractants, such as LIX and Kelex, which have been recently devel-
oped, are a significant aid in processing these ores by hydrometallurgy.
Impurity of product, buildup of iron in the dump, acid consumption, and
increased cost of scrap iron all have led to a preference for the solvent
extraction/electrowinning route for copper recovery. However, all processes
have problems associated with them and the solvent extraction/electrowinning
process is no exception. These problems are (1) organic burn of the
cathodes, (2) acid misting in the tankhouse, and (3) alternative anode mate-
rials which can resist high acid concentration and methods to overcome the
inclusion of lead in the cathode copper.
Further study is needed in the interaction of the solvent extraction
stage with the electrowinning operation, the effects of organics and flota-
tion agents in the electrowinning stage, and the feasibility of the use o'f
diaphragm cells, cyclic voltammetry, and scanning electron microscopy.
Reference: MacKinnon, D. J., and V. I. Lakshmanan, Canada Center for
Mineral and Energy Technology, Canmet Report 76-10, Jan. 1976.
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Subject: Hydromet May Free Small Lead Districts From The Smelter
The chloride-electrolysis process of the U.S. Bureau of Mines (see -
Technology & Trends Index 2.3.2, page 1) could permit small mining districts
to recover the lead from their own concentrates, saving millions of dollars
each year.
At the present time, small mining districts must ship their concentrates
to smelters, often at long distances, thus requiring shipping charges in
addition to the smelter's fees. In this article, the author analyzes the
economic feasibility of the installation of the new Bureau of Mines' hydromet
process in two lead districts of Mexico.
The U.S. Bureau of Mines' Chloride-Electrolysis Process leaches the sul-
fide concentrates with a solution of ferric chloride, sodium chloride, and
hydrochloric acid which, in 30 minutes, can dissolve 98 percent of the lead
and 96 percent of the silver. The residue (50 percent of which is sulfur) is
removed by filtration and pure lead chloride crystals are formed from the
filtrate. These crystals are fused and electrolyzed at 320 C to produce
lead. The leaching solution is regenerated with chlorine gas which is pro-
duced in the electrolytic cells. No sulfur dioxide air pollution results
from this process, nor are workers exposed to lead fumes.
The two Mexican lead concentrators considered produce about 12,000 and
18,500 stpy, respectively. These would require $3.4 and $4.3 million capi-
talization for the USBM process, of which an assumed 60 percent would be
financed (at 10 percent per year!). Over a 14-year assumed life, the esti-
mated operating costs would be $13.2 and $18.1 million, respectively. Includ-
ing ore depletion charges, gross profits accruing to the hydromet operatidn
installation would be $10.3 and $12.5 million for the 14-year period. It is
concluded that supplanting freight and custom smelting fees with the USBM
hydromet process would be a winning proposition for small lead concentrators.
Reference- Rovirosa, N. Y., adapted from "The Hydrothermal Processing of
Lead Concentrates in Small Scale Mining", Mining Engineering,
31(2):147-149, February, 1979.
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USEPA TECHNOLOGY AWP TREMVS ABSTRACT
for the NONFERROUS METALS INDUSTRY
Subject: Cominco Buildina World's First Zinc
Plant of 70,000 TPY Capacity
Index 2.3.2
Paqe 8
Pressure Leaching
!
As part of an 8-year $425+ million modernization and expansion program,
Cominco announced the beginning of construction of a $23-million, 70,000-ton
per year pressure leaching zinc plant at its Trail, B.C., operations. The
plant represents the first commercialization of the S-C (Sherritt-Cominco)
process for zinc.
In this process, zinc sulfide concentrates are directly leached (without
roasting) in oxygenated ^504 at high temperatures and pressures in a closed
system. This removes the sulfur as free sulfur which is extracted and easily
stored for later use or marketed. In Cominco's case, the added flexibility
in fertilizer production will be a major feature. The new operation will
greatly improve industrial hygiene and environmental issues. About one-fourth
of the Trail zinc concentrate will be processed in the new facility after the
second reactor comes on-stream in 1983. The balance of concentrates will
continue to be processed by conventional roast-leach processing. (The modern
fluidized bed roasters were installed in 1971. The leaching plant is being
modernized as part of the current program.)
The pressure-leached product will join the main flow in the conventional
leaching plant for purification and refining.
The Northern Miner, 65(6) :A14, April 19, 1979.
American Metal Market, 87(78):9, April 20, 1979.
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Subject: Sulfation Roast Boosts Recovery from Lean Concentrates
Ore reserves in the New Brunswick area of northeastern Canada total
400 million tons, grading 13 percent lead, zinc, and copper, yet only two
operations are producing zinc concentrates. The reason more mining develop-
ment has not taken place in this area is the startling losses which occur
during the processing of the ore; as much as 35-45 percent of the copper,
30 percent of the lead, 13-20 percent of the zinc, and 35-50 percent of the
silver.
Three years' work was recently completed to find and assess processes
to better treat these low-grade zinc concentrates. Two processes were
developed, one by Sherritt Gordon (a zinc pressure leaching technique) and
a sulfation roasting process by the Research and Productivity Council (RPC)
of New Brunswick.
The process sulfation roasts ore at a relatively low 685 C then leaches
the roasted ore in two stages with weak, then strong sulfuric acid. The
strong acid filtrate is recycled to the roaster to eliminate the iron sul-
fate which is the key to the improvement. The RPC process accepts bulk
concentrates, eliminates the problem of iron in solution common to zinc
smelters. It can be used in existing facilities without extensive modifica-
tions and is a low-energy process.
Pilot-plant operation at about a 10-ton per day rate is the next step
in testing the RPC process.
(This news article presented no bottom-line data; improved recoveries
from the complex, lean ores are merely inferred.)
Reference: Branch, S., Canadian Mining Journal, 100(3):51, 53, March, 1979.
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for the NONFERROUS METALS INDUSTRY Page 1
Subject: Jersey Miniere Zinc Opens A New Refinery Capable of
Producing 90.000 tpy of Slabs
The Clarksville, Tennessee, zinc refinery, a joint venture of New Jersey
Zinc Company and Union Miniere, was dedicated in mid-March. The $200-miTlion
facility (refinery and mines), largest private industrial investment ever
made in Tennessee, will produce 90,000 tpy of zinc slabs and can include
140,000 tpy of acid, 300 tpy cadmium, and 200 tpy copper. The zinc capacity
can be doubled to 180,000 tpy if needed. (See "In the News", Technical
Awareness Bulletins, 1978.)
The zinc ore (4-4.5 percent zinc) is mined and concentrated at the
Elmwood and Gordonsville, Tennessee, mines to 60 percent zinc and 30-35 per-
cent sulfur and shipped to Clarksville for refining. Here it is roasted in
a Lurgi fluid-bed roaster at 1650 F to remove the sulfur as sulfur dioxide.
This gas, after cleaning in an electrostatic precipitator, moves to the
double-contact acid plant where it is converted to 400 tpd of sulfuric acid.
The calcined ore, after grinding, is leached with sulfuric acid to form
a zinc sulfate solution. Zinc dust is added in first a cold, and then a hot
stage to precipitate contaminants (cadmium, copper, and cobalt). The zinc
sulfate solution is then electrolyzed, depositing metallic zinc on the alu-
minum cathodes. Sheets of zinc, 1/4 in. in thickness, are removed from the
cathodes and melted in induction furnaces into 2,400 Ib ingots or 50 Ib
slabs.
The air pollution control system cost $12.2 million at Clarksville.
Particulate matter is trapped in baghouses in the concentrate handling area,
the zinc dust production unit, the dross grinding area, and in the cadmium
smelting and casting areas. Total efficiency is 99.9 percent. In the acid
plant, tailgas sulfur dioxide is less than half the 650 ppm allowed by the
EPA. Water, just under 1 million gpd, is mostly cleaned and recycled, but
some water is treated and directed to three settling ponds before it is
returned to the Cumberland River.
Reference: 33 Metal Producing, 17(4):60-61, April, 1979.
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USEPA TECHNOLOGY AMP TRENPS ABSTRACT Index 5.2.4
for the NONFERROUS METALS INDUSTRY Page 2
Subject: Flotation Reagent Practice in Primary
and Byproduct Molybdenum Recovery
This article is a review of the reagents used as collectors, frothers,
and depressors in the flotation of porphyry copper for molybdenum recovery.
A large table details the primary flotation circuit, concentrate retreatment,
and metallurgical results of plants in the U.S. (16), Canada (5), Chile (4),
Peru (1), and the Communist World (4).
Examination of the table shows no consistent pattern in copper-
molybdenum recovery except that better molybdenum recovery generally comes
from ore with the lowest copper content. This is probably due to the greater
attention paid to optimize molybdenum recovery because of low copper values
in the ore.
This review revealed that about 60 percent (19 of the 30 plants) use
xanthate as a collector and about the same percent use MIBC as the primary
frother; however, the author's analysis of all the available published data
showed no statistically significant correlation between the collector or
frother used and molybdenite recovery. The reason is probably because copper
is the most valuable mineral in the porphyries and the collector and frother
are chosen primarily to increase recovery of the copper rather than
molybdenum.
Another factor in reagent selection is the depression of copper in the
molybdenum circuit. Preconditioning can be thermal or chemical. The thermal
treatment is used by a third of the mines and includes roasting or steaming
to decompose the collector on the surface of the sulfide mineral and vaporize
the frother. The majority use sodium sulfide or hydrosulfide depression of
copper sulfides. An oxidation process utilizes hypochlorites, peroxides,.
permanganates, or dichromates to oxidize copper and pyrite surfaces and to
possibly destroy adsorbed collectors. The oxidizers are always combined with
some form of cyanide (sodium or potassium ferro or ferri) to provide the
desired HCN level with f^Og in the conditioning step and to remove the col-
lector coating on the copper mineral.
(In primary molybdenite mines~e.g., Climax and Henderson—the reagent
selection is completely different. In early days, only pine oil was used
for collector and frother. Today, a neutral oil called "Vapor Oil" is used
for the collector and Artie Syntex L, a sulfated coconut oil, for some froth
control. Sodium cyanide and Nokes reagent (P2$5 plus caustic) is used to
depress gangue minerals.)
Reference: Crozier, R. D., Mining Magazine, 140(2) .-174-178, February, 1979.
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USEPA TECHNOLOGY ANP TRENDS ABSTRACT
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Subject: How to Select a Byproduct Moly
Index 5.2.9
Page 4
Recovery Process
(This article is the second in a series on byproduct molybdenum
recovery. The first was abstracted in The Technical Awareness Bulletin
Index 5.2.9, page 2.)
Recovery of molybdenite from copper operations, once considered uneco-
nomical, today, with the very strong molybdenum market, it is a different
story. With the many byproduct recovery processes available, it is difficult
to choose the proper procedure to use for a particular ore. The author
reviews the most attractive processes for treating different types of ores
during initial startup.
Only established processes should be considered for a new mine, and
operations should be kept as simple as possible. After the plant is in
operation, more effective, procedures can be developed. Because several of
the reagents used are poisonous and known pollutants, environmental aspects
must be considered. Despite the fact that potential pollutants from moly
circuits are nearly always contained, local policy and the fear of public
reaction are strong forces. These could be the most important factors in
deciding which process to use.
The metallurgical effectiveness of a process is directly influenced by
(1) copper mineralization, (2) reagents used in the copper recovery circuit,
(3) the mode of occurrence of molybdenite in the ore body, and (4) the pres-
ence of natural floaters in the ore. Natural floaters are talc and pyro-
phyllite, sulfur, graphite, coal and native asphalt, and covellite. These
cause unique design and process problems for their suppression.
The author's recommendations for flotation reagents and procedures can
be summed up as follows:
Chalcopyrite ores: 1. Sodium hydrosulfide; 2. Arsenic Nokes
Chalcocite ores: 1. Sodium ferrocyanide with (a) acid-peroxide
conditioning or (b) steaming or cooking; 2. Phosphate Nokes with
(a) acid-peroxide conditioning or (b) steaming or cooking.
Mixed ores: 1. Phosphate Nokes with steaming or cooking pretreat-
ment; 2. arsenic Nokes.
Reference: Shirley, J. F., World Mining, 32(4):44-47, April, 1979.
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for the NONFERROUS METALS INDUSTRY Page "f "
Subject: Fuming of Stannous Oxide From Slags
An accepted process for the separation of tin from slags is the fuming
of stannous sulfide. The recovery of tin is high, 90-95 percent, but the
process results in the production of sulfur dioxide. To avoid pollution of
the environment by sulfur dioxide, there is interest in fuming of SnO to
recover tin from slags. This paper describes the laboratory experimental
work on the fuming of tin from slags as stannous oxide (SnO).
Four types of experiments were run: (1) fuming of liquid slags under
an atmospheric pressure of CO-CO? or CO-Ar mixtures, (2) sparging with CO,
(3) sparging with hydrogen, and (4) under reduced pressure. The results
showed that, at atmospheric pressure, both with and without gas injection,
the fuming of SnO was greater under reducing conditions. The reason for
this is not clear unless a volatile hydroxide was being formed.
Whether the operation is commercially attractive depends on the control
of oxygen potential and the rate at which SnO is fumed in the absence of
reduction or oxidation of the slag. A range of oxygen potential exists
within which virtually all tin removal from slag is by the vapor phase;
therefore, volatilization of SnO is nearly as important a mechanism of tin
removal from slag as tin reduction. Fuming of SnO from slags by purging with
a nonreducing gas is not practical. Treatment of slags under a vacuum is
concluded to offer an alternative to sulfide fuming, however. Much more
study is needed to adequately assess the economic feasibility of such a pro-
cess if sulfide fuming becomes unacceptable because of S02 emissions.
R°y> T- D-> s- R- Chandrashekar, and D. G. C. Robertson,
Inst1tution of Mining and Metallurgy, Section C, 87:C225-C230,
December, 1978.
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Subject: ''Thar's Gold in Them Thar Wells!"
The greatest potential of the Sal ton Sea geothermal field in California's
Imperial Valley is not the 5000 MW electrical generation capacity, but the
2 ppm of gold the waters contain, according to this article. If 80 percent
of the gold were recovered from these waters, the Sal ton Sea field alone
could produce 2340 short tons of gold annually, double the present world's
output. Because geothermal resources are exploited by power companies, most
are not authorized to extract minerals; consequently, the mineral potential
is neglected.
Minerals may be recovered from the sinters and precipitates where they
have been concentrated by evaporation, and from the gases, as well as from
the waters. The waters offer the lowest cost recovery; however, the mineral
content is generally low. Geothermal waters carry arsenic, boron, sulfur,
gold, antimony, silver, and base metals, and often high quantities of salt,
potash, and manganese. Sinters and precipitates carry high concentrates of
gold (85 ppm), antimony, and silver, with some values of mercury, zinc, and
copper. Gases generally contain large quantities of carbon dioxide.
There are few data as to the life of the geothermal fields. However,
one in Italy is still operating after 70 years, and some in the U.K. were
once utilized by the Romans. The author claims, if the minerals are obtained
from hot water leaching from host rock, concentration will change little with
time. If, on the other hand, the minerals are derived from magma, mineral
content depends on length of time those sources are operative.
The potential is great for geothermal areas to become a source of
mineral supply, especially gold.
Reference: Barnea, J., Adapted from "Geothermal Minerals - The Neglected
Minerals", Mfning Engineering, 31(2):146-147, February, 1979.
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for the NONFERROUS METALS INDUSTRY Page " l
Subject: Cyanide and Arsenic Removal from Gold Mine Effluent
This brief article previews a forthcoming report relative to a joint
industry-government (Canadian) demonstration project to lower cyanide and
arsenic levels in gold milling effluents from the Giant Yellowknife Mines,
Ltd., operations.
Details of the project were not given; however, at a chlorine-to-cyanide
ratio of 5:1 and,with a (pond) retention period of 5 hours' total,cyanide was
less than 1 mg/ji in the barren solutions. (Cyanate, rather than C02 and N2,
is the oxidation product.) A C1:CN~ ratio of 9:1 was required for pond over-
flow. Iron precipitation reduced arsenic to the 0.2 to 0.5 mg/£ level when
iron: arsenic ratios of 7:1 to 4:1 were used.
"The 96-hour toxicity tests on treated effluents indicated that non-
acutely lethal (??) effluents were possible after dechlorination". The added
cost for treatment at Giant Yellowknife would range from $0.94 to $6.32/oz of
gold, depending on treatment.
(This brief is not particularly rational in itself, but the forthcoming
report should be of interest.)
Reference: Water and Pollution Control, 117(2):20, February, 1979.
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USEPA TECHNOLOGY AMP TRENDS ABSTRACT Index 9.3.4
for the NONFERROUS METALS INDUSTRY Page 2
Subject: Smoky Valley Operations at its Round Mountain Mine
in Nevada
This paper presents a history and description of gold mining at Smoky
Valley's Round Mountain Mine near Tonopah, Nevada.
Operating intermittently since 1906, a joint venture of Copper Range,
Case Pomeroy & Company,and Felmont Oil Corporation began full-time production
mining in January, 1976. In 1979, production is expected to reach 55,000 oz
of gold from 1,800,000 tons of ore. The cutoff for ore is 0.02 oz of gold
per ton with material below that considered waste. Reserves total 12,000,000
tons of ore averaging 0.062 oz of gold per ton with 0.053 oz recoverable and
0.07 oz of silver per ton. Six years of production is planned at this rate.
The ore, after crushing to minus 2 inches, is piled in five, 50,000-ton
heaps on an impervious membrane of ground rubber and asphalt to begin the
heap leaching process. A dilute solution of 1 pound of sodium cyanide and
0.9 pound of lime per ton of solution is sprayed at the rate of 400 gallons
per minute on each heap for 27 days. This is followed by a 2-day water wash
when the residual gold content falls below the economic level. The leaching
effluent flows through five activated-charcoal-loaded adsorption tanks at
1,600 gallons per minute where the cyanide complexes are adsorbed on the car-
bon at the rate of about 250 troy ounces of gold and silver per ton of
charcoal.
The gold-and silver-laden carbon is stripped by hot sodium hydroxide and
sodium cyanide solution at 190 F, then reactivated in a kiln and returned to
the system. The pregnant solution is plated out on steel wool bats used as
cathodes. The recovered gold and silver (in a 2 to 1 ratio) is then combined
with fluxes and heated in crucible reduction furnaces to produce about 2,200
ounces of dore buttons weekly. These buttons are remelted and cast into two,
1,100 ounces bars per week for shipment to a refinery.
Water source is from Jett Canyon (8.5 miles west and 760 feet above the
plant) within the Toiyabe Mountain range,and two wells. A deep well situated
down slope from the waste dump is tested periodically for possible seepage of
the cyanide leach solution.
Reference: Skillings, D. N., Jr., Skillings1 Mining Review, 68(9):8-9,
16-17, March 3, 1979.
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USEPA TECHNOLOGY AWD
for the NONFERROUS
Subject: Eberle
TREWPS
METALS
Mine:
ABSTRACT
INDUSTRY
A Heap Leachinq
Index
Page
Case History
9
.3
4
.8
This article examines in detail (including a flow sheet) the heap leach-
ing process operated at the Eberle mine by Challenge Mining Company in
western New Mexico's Cooney district. Here, because the low-grade ore could
not sustain a mining and milling operation, the district lay virtually dor-
mant for 30 years until the advent of the cyanide heap leaching process.
Total investment at the Eberle operation is $600,000, with about 34,000
short tons of ore capacity per year. A 60-day leaching period is expected
to recover 70 percent of the gold and silver. Washing requires another 20
days at the Eberle operation. Eighty-five percent of the silver is precipi-
tated with sodium sulfide before the pregnant leach solution is pumped
through the charcoal stripping columns, and the balance is precipitated after
stripping and before electrowinning to recover gold.
Preliminary analysis indicates operating costs (based on surface vein
mining and processing of old mine dumps) of $10.41 per ton. Projected
profits are such that the investment should be recovered in about 2 years.
Future operations will be supported by underground mining at the old Eberle
mi ne.
Eveleth, R. W., Adapted from "New Methods of Working an Old Mine-
Case History of the Eberle Group, Mogollon, N. M.," Mining
Engineering, 31(2):138-140, February, 1979.
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USEPA TECHNOLOGY AMV TRENDS ABSTRACT Index 938
for the NONFERROUS METALS INDUSTRY Page 5
Subject: Heap Leaching is Small Miner's Golden Opportunity
If the ore is relatively free of carbonaceous cyanide-consuming and
acid forming agents, and also free of excessive fines, heap leaching is
ideal for the small scale gold and silver miner. This article explains the
process, including costs, in some detail with considerable emphasis on stack-
ing of the ore. (An actual gold heap leaching operation is described in this
issue of the Technical Awareness Bulletin, "Smoky Valley Operations At Its
Round Mountain Mine in Nevada".)
The author estimates that an entire heap leaching/recovery operation
sized to leach 10,000-ton heaps in about 2 months can be put into production
for $200,000(exclusive of mining costs) with minimal labor requirements, make-
up sodium cyanide cost of about $0.15 per ton per month and a power require-
ment of 0.0003 kw per ton. The low capital cost and relative ease of opera-
tion is attributed to two recent developments: (1) plastic membranes for
the leach pads and (2) an activated carbon process for recovery of the gold
and silver.
The most important factor in the leaching operation is the permeability
of the rocks and of the heap as a unit. Permeability of the heap as a unit
cannot be adequately determined in the laboratory. Therefore, heap construc-
tion is a critical operation and should be piloted in the field. Compaction
of the heap should be avoided. The building of successive layers of cones
as high as 16 feet, by the use of a mobile stacker/conveyor or a clamshell
bucket on a small crane, is usually preferred.
Cyanide leaching of the heaps is best accomplished by a sprinkler sys-
tem which guarantees good distribution, although it is more complicated and
requires more maintenance than the ponding technique.
Single-use plastic leach pads,where the heap is leached as long as it
is returning values, allows better flexibility than periodic removing and
rebuilding of heaps on multiple use pads. Expended and washed heaps can be
sealed with clay or contoured and vegetated for environmental compliance.
Reference: Kappes, D. W., Adapted from "Leaching of Small Gold and Silver
Deposits", Mining Engineering, 31(2):136-138, February, 1979.
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USEPA TECHNOLOGY AND TRENPS ABSTRACT Index 10.2.8
for the NONFERROUS METALS INDUSTRY Page 3
Subject: New Swedish Scheelite Flotation Process Results
In a prior Issue of the Technical Awareness Bulletin. Index 10.2.8,
page 2, the article "How Swedish Metallurgists Developed Selective Scheelite
Flotation Process" was presented. After almost a year of operation, the
present article describes the process in more detail with flowsheets of the
grinding circuit, the copper flotation circuit, and the scheelite flotation
circuit.
After grinding in a rod mill and in a ball mill in circuit with a
cyclone, the 80 percent passing size is about 160 microns. Chalcopyrite is
recovered as a valuable byproduct by a long scavenger series and four cleaner
stages. Rejects from the copper flotation are conditioned with fatty acid
which flocculates the scheelite into very stable aggregates which require
special efforts to remove entrapped particles of calcite and fluorite.
Tailings are flocculated with caustic lime before being pumped to the tail-
ings pond. Effluent from a subsequent clarification pond has a turbidity of
<10 FTU and contains <6 mg per liter of suspended matter, with only about
50 ppm of copper. Fluorine tests show 4-5 mg per liter. Attempts to lower
this are in progress.
Selectivity in the 26 to 125 micron range is very good with a 97.6 per-
cent recovery in a concentrate containing 73.4 percent W03 or 91.2 percent
CaW04.
Future development includes: (1) improved stability of operations by
using on-line analyses and reporting, (2) additional future development in
the grinding stages where losses occur in the ultra fines; perhaps pebble
grinding will replace ball milling, and (3) recoveries of 70 to 80 percent
of the fluorite at concentrate grades of 98 and 95 percent, respectively.
The latter will be achievable only at considerable investment, and will
depend on recovery of the currently depressed fluorite market.
Reference: Grasberg, M., World Mining, 32(3):54-58, March, 1979.
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USEPA TECHNOLOGY AND TRENVS ABSTRACT Index
for the NONFERROUS METALS INDUSTRY Page
11.3.9
1
Subject: Carbothermic Reduction of Domestic Chromites
All of the chromite used to produce ferrochrome in the U.S. must be
imported. Because of the increasing demand for chromium and our dependence
on foreign sources, the USBM has undertaken laboratory research to maximize
iron and chromium recovery while decreasing energy requirements, from low-
grade ore found in the U.S. This paper describes the work to determine the
optimum types of carbonaceous reductants, temperatures, and duration of
reduction to prereduce the domestic chromites in preparation for smelting.
(Advantages of prereduction as opposed to direct smelting of concentrates
include lower electrical energy requirements, improved smelting control and
productivity, improved slag quality, and better overall recovery.)
Two chromite materials were used, representing the two major types of
deposits; Mouat concentrate from Montana, a strateform, peridotite-gabbro
complex containing 34 percent chromium oxide and 21 percent iron oxides, and
the High Plateau ore from northern California, a podiform deposit, containing
54 percent chromium oxide and 17 percent iron oxides. Five reductants were
investigated: coal char, coke breeze, metallurgical coke, petroleum coke,
and shell carbon.
The reduction mechanism of both chromites cannot be described by simple
kinetic equations; however, it is believed to be nucleation controlled.
Based on the degree of reduction and metallization, coal char is the pre-
ferred reductant for both chromites up to 1300 C. At 1400 and 1500 C metal-
lurgical coke is the choice for the Mouat chromite while coal char remains
the choice of the High Plateau chromite. Reduction for both of the chromites
is greatest during the first 15 minutes,and increasing temperatures tend to
increase the degree of reduction and metallization.
The maximum reduction and metallization for 1 hour of both chromites are
shown in the following tabulation:
Material
Mouat chromite
(High iron)
High Plateau chromite
(Low iron, metallurgical
Temp C Reduction % Metallization %
grade)
1200
1300
1200
1300
32
57
25
49
27
44
12
33
Reference:
Nafziger, R.
Transactions
H., J. E. Tress, and J. I. Paige, Metallurgical
(B), 10B(1):5-14, March, 1979.
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USEPA TECHNOLOGY ANV TREHVS ABSTRACT
for the NONFERROUS METALS INDUSTRY
Index _L3,2,9
Page 4
1
Subject: Cobalt and Nieke 1 Recovery fromjlissourj J.ead Be 11
._ Chajcopyrite Concentrates
The Missouri Lead Belt reserves are estimated at 325 million tons,
grading (in addition to lead and zinc) 0.3 percent copper, 0.02 percent
nickel, and 0.015 cobalt. Copper recovered accounts for about 1 percent of
the national primary copper production, and the copper concentrate fraction
carries up to 30 percent of the nickel and cobalt content of the ore. Prior
USBM studies on the recovery of nickel and cobalt were ineffective on the
copper concentrates, and this investigation concentrated on this subject.
The payoff would be the recovery of several tens of millions of pounds each
of nickel and cobalt.
The Bureau found that leaching of the chalcopyrite concentrate from J:his
source with ferric chloride concentrated nickel and cobalt in the residues
at concentrations from 10 to 16 percent Ni and 8 to 11 percent cobalt, with
recoveries estimated at about 80 percent. Froth flotation of reground and
pretreated copper concentrates resulted in 67 percent of the nickel and
cobalt reporting in concentrated forms to the tailings. Although the chlo-
ride leach appears to be more effective, it is not so adaptable to Missouri
Lead Belt mill practice.
A moderately scaled-up demonstration (60 kg/hr) conducted at a commer-
cial mill confirmed the initial laboratory finding relative to the applica-
bility of flotation.
Clifford, R. K., and L. W. Higley, Jr., USBM-RI 8321, U.S.
Department of the Interior, 1978.
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USEPA TECHNOLOGY AWP TREWPS ABSTRACT Index 13.3.2
for the NONFERROUS METALS INDUSTRY Page 2
Subject: Selective Recovery of Copper from Copper-Nickel Sulfide
Concentrates by Applying Segregation Technology
Technology for the "segregation" process (chloride-activated reduction
of copper on coke in a solid-state operation) is well-documented for treat-
ment of oxide ores and for dead-roasted chalcopyrite calcine. This report
by AMAX researchers gives the results of laboratory studies directed toward
selective "segregation" of copper from copper-nickel-iron sulfide concen-
trates. The studies were conducted on low-grade (4.6 percent Cu, 3.1 percent
Ni, 45 percent Fe, 28 percent S) concentrates from Botswana, and higher-
grade concentrates from Minnesota.
Concentrates were dead-roasted to 0.2 percent S max., and the calcine
was blended with 0.5 percent NaCl and about 5 percent coke. Also, the addi-
tion of up to 10 percent Si02 to the silica-lean Minnesota concentrate aided
segregation. Small-scale (50 gm of calcine) tests were conducted under a
flowing moist CO atmosphere for 4 hours at 675 C. In these tests, 85-92 per-
cent of the copper was reduced to metal, and greater than 99 percent of the
nickel remained with the unreacted calcine. Tests of 600 gram lots were
conducted, with products separated by flotation. (Copper was activated with
Cyanamid Aerofloat 208 in conjunction with Aerofroth 65.) Under optimum
conditions, about 80 percent of the copper and 0.1 percent of the nickel were
collected in the flotation concentrate. Middlings, with intermediate amounts
of copper and nickel were suitable for recycle through the segregation treat-
ment. Simple melting of concentrate samples yielded blister copper--98 per-
cent Cu, 0.75 percent Fe, 0.05 percent Ni. Thus, energy-efficient segrega-
tion treatment has been shown capable of separating metallic copper from
mixed copper-nickel sulfides, with nickel concentrated in the tailing. S02
emissions are restricted to the enriched roaster offgas, which is readily
processed in an acid plant.
Reference: Opie, W. R., L. D. Coffin, D. L. Armant and 0. F. Cimler,
Metallurgical Transactions, 10B(1):27-32, March, 1979.
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USEPA TECHNOLOGY AMP TREWPS ABSTRACT Index 13.3.9
for the NONFERROUS METALS INDUSTRY Page 5
Subject: Amax will Use Energy-Saving Nickel Extraction Process
in New Caledonia
A new nickel extraction process has been developed and piloted tested
by Amax which will reduce the energy required to extract nickel by 40 per-
cent when compared to a pyrometallurgical process. A unique feature of the
sulfuric acid leaching process is that it can economically treat both the
limonitic layer (low nickel and magnesium content) and the garnieritic
fraction (high nickel and magnesium content) of nickel-bearing laterites.^ A
combination of high-pressure and atmospheric-pressure leaching is used, with
the residual acid in the solution neutralized by the readily soluble mag-
nesium from the garnieritic fraction at atmospheric pressure. The key to the
process is the consideration of elemental sulfur as a source of primary
energy with credits for the sulfuric acid by-product.
Reference: Engineering and Mining Journal, 179(11):32, November, 1978.
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USEPA TECHNOLOGY AMP TREWPS ABSTRACT Index 13.3.9
for the NONFERROUS METALS INDUSTRY Page 5
Subject: The Kennecott Process for Nickel-Slag Cleaning
Since 1975, it has been confirmed that a stirred electric furnace for
cleaning copper and molybdenum from oxidized slags, produced by converters
or the Noranda Process, was feasible. Because nickel recovery from oxidized
slag is thermodynamically similar to the case for molybdenum, the application
to nickel was considered.
This paper describes the thermodynamics and laboratory experiments on
nickel-slag cleaning considering the Kennecott Process from two points of
view: (1) modifying existing smelters by treating nickel converter slag
directly in a stirred electric furnace and (2) developing a direct-smelting
process including a pyrometallurgical slag cleaning step. Experimental data
were obtained on the rates of magnetite reduction and metal recovery as a
function of slag temperature and composition, matte composition, and furnace
and stirrer design.
By adding the stirred electric furnace to existing equipment, an
increase of about 15 percent in plant capacity can be expected. A commercial
operation built to incorporate the stirred electric furnace has been designed
for a nickel smelter along the lines of Kennecott's Utah Copper Division fur-
nace that is in use.
The advantages of the Kennecott slag-cleaning process are (1) fast mag-
netite reduction, (2) fast digestion rate of concentrate or flux, and (3) a
higher recovery of metal values.
Reference: Arnmann, P. R., J. J. Kim, and T. A. Loose, Journal of Metals,
31(2):20-25, February, 1979.
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USEPA TECHNOLOGY AND TRENDS ABSTRACT Index 15.2.9
for the NONFERROUS METALS INDUSTRY Page 2
Subject: Galena-Sphaleri'te-Chalcopyrite Flotation
at St. Joe Minerals Corporation
This paper discusses the significant variables in the differential flo-
tation of galena, sphalerite, and chalcopyrite at St. Joe's Bushy Creek mill
in Missouri.
The ore is normally 3-10 percent lead, 0.4-2.0 percent zinc, and 0.05-
0.40 percent copper in a dolomitic limestone gangue. Isopropyl xanthate is
the primary collector for galena and chalcopyrite flotation. Zinc sulfate
depresses sphalerite. Copper recovery is increased by the addition of Z-200;
however, excess is avoided to prevent floating sphalerite.
High zinc recovery requires 5 minutes' conditioning of the tailings with
copper sulfate to activate the sphalerite. Insufficient copper sulfate pro-
duces a high zinc tailing assay. Residual xanthate and Z-200 from the lead
copper circuit is sufficient collector for good recovery and grade for zinc-
lean tailings, but added collectors are necessary at >0.4 percent zinc.
Sodium cyanide is added to the zinc recleaner to depress iron minerals.
Cyanide is necessary for zinc recovery, although the mechanism is not under-
stood. Water added at this point increases the zinc concentrate grade to
58-60 percent zinc.
Lead-copper separation is made by a 3-5 minute conditioning period with
causticized starch and SC>2 in which the chalcopyrite is floated away from
the depressed galena. Preconditioning with potassium dichromate improves the
grade by lowering the lead assay 1-2 percent. The S0£ is added to keep the
pH at 4.5-5.0, which also helps to depress the lead. Starch is a general
depressant and will depress the copper if insufficient S02 is used.
One of the most significant variables is the xanthate level; if the
optimum is used, good separation is not difficult. However, if excess is
used, lead is difficult to depress.
Reference:Clifford, K. L., E. J. Haug, and K. L. Purdy, Mining Engineering,
31.(2): 180-182, February, 1979.
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USEPA TECHNOLOGY AND TREWPS ABSTRACT Index
for the NONFERROUS METALS INDUSTRY Page
Subject: The Characterization of Mercury
in Some Sulfide Concentrates
The presence of mercury in many sulfide ores, although in small quanti-
ties, could pose an environmental problem in both smelting and hydrometallur-
gical processing because of the toxicity of its vapor and compounds. This
paper examines in detail mercury distribution in the copper-gold ore of the
Timmons, Ontario, Pamour Porcupine operations and in the pyritic zinc-lead-
copper-silver ore from the mill of the Brunswick Mining and Smelting Corpora-
tion, Bathurst, New Brunswick. Flowsheets of both operations and several
tables of analytical data are presented.
In both milling operations, most of the mercury reports to the various
concentrates, which may contain in the range of 5-15 ppm mercury. Mercury
concentrations in plant tailings typically range from about 0.1-0.7 ppm.
A number of analyses were made of the minerals present in the concen-
trates of the two operations to determine which of the minerals present were
responsible for the mercury content. The copper concentrate of the Pamour
Porcupine Mines contained 60 percent of the mercury in less than 2 percent
of the input weight, with the three mercury-bearing minerals identified as
native silver-electrum, tennantite-tetrahedrite, and sphalerite. In the con-
centrate from the Brunswick operation, the zinc concentrate contained 67 per-
cent of the mercury in 11 percent of the input weight. More than 90 percent
of the mercury was contained in the sphalerites with tetrahedrite and pyrite
as possible minor carriers.
During pyrometallurgical processing of either copper or zinc concen-
trates, most of the roaster-volatilized mercury is trapped by the weak acid
scrubber as a (Hg,Cu)x (Se,S)y solid, but a small amount passes into the
sulfuric acid. Only about 1 percent of the mercury input remains with the
calcine.
Reference: Dutrizac, J. E., and T. T. Chen, CIM Bulletin, 72(803):201-208,
March, 1979.
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US EPA TECHNOLOGV ANP TRENVS ABSTRACT
for the NONFERROUS METALS INDUSTRY
Subject:
A New Carbothermic Process
Index
Page
15.3.9
3
Carbothermic reduction of oxides of highly reactive metals (I), Ti02?
1102, A1203, Si02, ZK)2, and MgO, etc.) usually results in the formation of
carbide or mixtures of the metal and carbide, neither of which is particu-
larly useful or saleable.
The authors have recently developed a new method for the production of
commercially pure, highly reactive metals which eliminates the formation of
the carbide and the use of very high temperatures. The method uses a liquid
metallic solvent (tin) to lower the chemical activity of the reduced, dis-
solved reactive metal, eliminating formation of the carbide. The reaction is
as follows:
MC-2 + 2C -> 2CO + M (dissolved in Sn)
The choice of the liquid metal solvent is very important. Sn has been found
a suitable solvent for uranium and several other metals; however, additional
research is needed to find the best solvent for each case.
The paper describes the laboratory work done using uranium to establish
the overall reaction of the process and the effects of temperature and pres-
sure. The oxide and carbon were pelletized before insertion in a quartz tube
and heated by induction to between 1550 and 1630 C under a pressure of 1-10
torr of carbon monoxide. The U-Sn was then heated by induction in a yttria
crucible to 1300 C in a vacuum of 10~5 torr to distill the tin, thus allowing
recovery of the uranium.
Reference: Bakshani, N., N. A. D. Parlee, and R. N. Anderson, Industrial
Research/Development, 21(2) :122-126, February, 1979.
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USEPA TECHNOLOGY AND TREWPS ABSTRACT Index 16.1.8
for the NONFERROUS METALS INDUSTRY Page 1
Subject: Future Effects on the Mining Industry
of 1977 Clean Water Act
The author, in this article, examines the Clean Water Act of 1977 and
its predecessor, The Federal Water Pollution Control Act of 1972, to deter-
mine the implications they will have on the mining industry. The Clean Water
Act of 1977 amended parts of the 1972 Act. Important factors include exten-
sions of deadlines for compliance.
A proposal by the EPA in August of 1978 was made to revise the existing
National Pollutant Discharge Elimination System regulations. In these revi-
sions, the EPA has shown that its intention is to propose and work for the
most stringent controls possible on each and every pollutant from each indus-
trial plant. The industry will have to assume "worst case" interpretation by
the EPA of every proposed regulation. In short, mining companies must expect
to be required to achieve zero discharge of pollutants to surface waters and
to the ground.
The cost of developing a new mine will often not be worth the economic
risk, with the long and costly environmental studies that will be required
before construction can begin, according to this analysis. Existing mines
could easily be closed with job losses and adverse impact on the balance of
payment position of the U.S.
Reference: Pickering, I. G., Mining Congress Journal, 65(2):75-80,
February, 1979.
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USEPA TECHMOLOGV AMV TREWS ABSTRACT Index 16.3.9
for the NONFERROUS METALS INDUSTRY Page 2
Subject: Plasma Process is Ready for Metals Recovery
Plasma technology, which seems to have something going for it in numer-
ous potential pyrometallurqical applications, appears to be on the verge of
another breakthrough. In Farinqdon, Oxfordshire, England, a unit will make
trial runs this spring to recover platinum-group metals from chromite ore.
The technique, called the expanded precessive plasma process (EPP),
centers on a water-cooled, metal sheathed, plasma gun with a doped tungsten
electrode and a small flow of argon. Mounted at 15 degrees from the vertical,
and revolving at 1500 or more rpm, the gun produces a precessing cone of
plasma between the gun and a counter electrode or the hearth of the furnace.
Smelting, or any other chemical action, occurs during the 300 millisecond
time that the fallina curtain of fine material takes to pass through the
high-temperature zone.
The trial runs this spring will take place in a 1,400-kVA demonstration
unit (a 300-kVA unit has already completed successful testing). Tetronics, a
British firm, and the U.K. affiliate of Foster Wheeler, developed this tech-
nique and is believed to be ahead of competitors in application of EPP.
Success in this test, sponsored by Texas Gulf, could trigger a big break-
through for the EPP process.
A future use of EPP could be in ferrochrome or steel manufacturing
which would call for a 6,000-kVA unit costing $5 million. It is projected
that the cost of ferrochrome production would be lowered by $0.04/lb.
Three distinct advantages of EPP should be noted: (1) the use of fine
particulate ore (<0.5-mm diameter) which is the size many ores are reduced to
during beneficiation, (2) the high temperature which permits strongly endo-
thermic reactions in very short periods of time, and (3) precise control of
furnace atmosphere.
Reference: Chemical Engineering, 86(5):75-77, February 26, 1979.
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USEPA TECHNOLOGY AMP TREWPS ABSTRACT
for the NONFERROUS METALS INDUSTRY
Index 16.9.2
Page 1
Subject: Baghouses: Separating and Collecting Industrial Dusts
This two-part treatise reviews in comprehensive fashion baghouse tech-
nology. The subjects discussed in the first part include:
Fabric properties, material, weaves (including staple
construction), comparative performance, cleaning
methods, design requirements, housing construction,
construction, testing for fire hazard.
The second part tells how one goes about buying a baghouse installation-
the analysis of bids, guarantees, flow-rate estimating, performance assess-
ment, etc. Monitoring of installation, startup, and testing for emissions
and leaks are also covered.
(This review is not specific to nonferrous metals operations, many of
which require baghouses, but the ABC's of buying, owning, and operating these
facilities covered in the pair of articles is a worthwhile addition to the
personal library of staff with baghouse responsibilities.)
Reference: Kraus, M. N., Chemical Engineering, 86(8):94-106, April 9,
and 86(9):133-142, April 23, 1979.
-------
NONFERROUS METALS RESEARCH REPORTS AVAILABLE THROUGH
NATIONAL TECHNICAL INFORMATION SERVICE
Assessment of Technology for Possible Utilization of Bayer Process Muds-
REF. NO. EPA-600/2-76-301
Control of Sulfur Dioxide Emissions from Copper Smelters: Volume I-Steam
Oxidation of Pyritic Copper Concentrates-REF. NO. EPA-650/2-74-085a
Control of Sulfur Dioxide Emissions from Copper Smelters: Volume II-Hydrogen
Sulfide Production from Copper Concentrates-REF. NO. EPA-650/2-74-085b
Copper-REF. NO. EPA-600/1-77-003
Determination of Hazardous Elements in Smelter-Produced Sulfuric Acid-
REF. NO. EPA-650/2-74-131
Energy Consumption: The Primary Metals and Petroleum Industries-REF. NO.
EPA-650/2-75-032b
Environmental Considerations of Selected Energy Conserving Manufacturing
Process Options. Vol. VIII. Alumina/Aluminum Industry Report-
REF. NO. EPA-600/7-56-034h
Environmental Considerations of Selected Energy Conserving Manufacturing
Process Options. Vol. XIV. Primary Copper Industry Report-
REF. NO. EPA-600/7-76-034n
Industrial Process Profiles for Environmental Use: Chapter 25. Primary
Aluminum Industry-REF. NO. EPA-600/2-77-23y
Industrial Process Profiles for Environmental Use: Chapter 26. Titanium
Industry-REF. NO. EPA-600/2-77-023z
Measurement of Sulfur Dioxide Particulate, and Trace Elements in Copper
Smelter Converter and Roaster/Reverbatory Gas Streams-REF. NO.
EPA-650/2-74-111
Metallic Recovery from Waste Waters Utilizing Cementation-REF. NO.
EPA-670/2-74-008
Methodology for Assessing Environmental Implications and Technologies:
Nonferrous Metals Industries-REF. NO. EPA-600/2-76-303
Operation of a Sulfuric Acid Plant Using Blended Copper Smelter Gases-
REF. NO. EPA-600/2-76-199
Process Modifications for Control of Particulate Emissions from Stationary
Combustion, Incineration, and Metals-REF. NO. EPA-650/2-74-100
Reclamation of Sulfuric Acid from Waste Streams-REF. NO. EPA-670/2-75-016
Regeneration of Chromated Aluminum Deoxidizers-REF. NO. EPA-660/2-73-023
Systems Study of Conventional Combustion Sources in the Primary Aluminum
Industry-REF. NO. EPA-R2-73-191
SO Control Processes for Nonferrous Smelters-REF. NO. EPA-600/2-76-008
Trace Pollutant Emissions from the Processing of Metallic Ores-REF. NO.
EPA-650/2-74-115
Treatment and Recovery of Fluoride Industrial Waste-REF. NO. EPA-660/2-73-024
Water Pollution Control in the Primary Nonferrous Metals Industry, Vol. 1-
Copper, Zinc, and Lead Industries-REF. NO. EPA-R2-73-247a
Water Pollution Control in the Primary Nonferrous Metals Industry, Vol. II-
Aluminum, Mercury, Gold, Silver, Molybdenum, and Tungsten-REF. NO.
EPA-R2-73-242b
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USEPA
Industrial Environmental Research Laboratory
Metals and Inorganic Chemicals Branch
Cincinnati, Ohio 45268
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TECHNICAL AWARENESS BULLETIN - NONFERROUS METALS
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