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By N. S. Keith

SOME time ago, a firm engaged largely in the manufacture of copper sulphate, applied to me for information as to the practicability of obtaining the copper from their mother liquors by means of electricity ; having reference, more especially, to obtaining the electric current from some magneto-electric or dynamo-electric machine. The mother liquors were the result of several solutions of commercial scrap copper, containing impurities, the quantity of which

Jan 1, 1878

By L. R. Bottomer, G. M. Leary

"The Copper Canyon property is located in the Galore Creek area of northwestern British Columbia. Porphyry copper-gold mineralization on the property is notable for its high gold content and close spatial relationship with a distinctive suite of high level syenitic and monzonitic intrusions, and pseudoleucite-bearing volcanic country rocks. The style of alteration and mineralization is similar in many respects to the nearby Galore Creek deposits.The current geological resource of 32.4 million tonnes grading 0.75% Cu, 1.17 g/ t Au and 17.1 g/t Ag is insufficient to support a mining operation, but definition of additional reserves combined with improved transportation infrastructure in the region may change this in the future. The climate, topography and proximity to an active glacier indicate that the most likely production scenario is an underground bulk mining operation, with the ore being milled offsite.IntroductionThe property is located at 57°07' North latitude, 131 °21' West longitude, on NTS sheet 104G/ 3W, approximately 160 km northwest of Stewart and 90 km south of Telegraph Creek (Fig. 1). The Galore Creek camp and airstrip are located 6 km to the west. Current access to the property is via helicopter either from Galore Creek, or from the Bronson Creek airstrip 32 km to the south-southeast.The deposit occurs within the Boundary Ranges of the Coast Mountains, on the north side of the east fork of Galore Creek, which flows northerly into the Scud River and thence westerly to the Stikine River.Topography is rugged. In the deposit area, elevations range from 1130 m to 1750 m. Valley glaciers and permanent snow and ice fields are characteristic of the region at these elevations, and at its southern limit, mineralization is truncated by the East Fork Glacier.The property is entirely above the treeline, and except for areas of moraine cover and talus slopes, outcrop is abundant."

Jan 1, 1995

By V. Annamalai, J. B. Hiskey, L. E. Murr

The kinetics of copper cementation on rotating disks of body and tear-top beverage can aluminum alloys were studied as a function of temperature and copper ion concentration. Both alloys followed the two-step kinetic regions, an initially slow rate followed by an enhanced rate. Experimental activation energy determinations (9.04 to 19.62 kJ/mol) indicated that ionic diffusion controlled the reaction mechanism between 20° and 60°C. At low temperatures (< 20°C), the system appeared to be controlled by the presence of an aluminum oxide surface film. The results of this study showed that cementation kinetics for the beverage can aluminum alloys are similar to those for pure aluminum. It was also found that the chemical removal of the oxide layer is less important for the alloys than /or the pure aluminum system. The characteristics of the resultant copper deposits were studied under a scanning electron microscope, and the reaction kinetics seemed to depend strongly on the nature and morphology of the deposits. In addition, maximum rates occurred when the deposit structures were fine feathery dendrites.

Jan 1, 1980

By J. J. Bean

The development drilling program of the Copper Cities mine indicated that the molybdenum content of the ore would be about 0.011% MoS2&apos; The molybdenum content of the Miami mine at the same time was about 0.025% MOS2&apos; The Miami mill had a very successful molybdenum operation which used bleeder turbine steam from the Miami power plant for the steaming operation. The much lower grade of the Copper Cities ore, the absence of a source of low cost steam and the use of sodium aerofloat as the copper collector which Miami experience indicated did not respono readily to the steaming technique made very doubtful the success of a similar operation at Copper Cities. However, the relatively few pounds of copper that could be recovered from the low grade Copper Cities heads made very desirable the recovery of the molybdenwa,as a source of additional revenue. Accordingly, shortly after the Copper Cities mill began operation, a research program was instituted to develop a suitable method of molybdenum recovery. This program explored first, the successful steaming process then in use at Miami and as expected, found it unable to decompose the sodium aerofloat copper collector. The known chemical depressants were then tried without much encouragement. Finally the copper concentrate was dried, not roasted, to 2-3% mOisture and this along with sodium ferro-cyanide and cyanide resulted in sufficient copper depression that a final molybdenum concentrate containing 1.0% CU could be made. The bench scale testing was followed by pilot plant testing. Sixteen

Jan 1, 1972

By Arthur Notman

IN VIEW of the wide publicity given to the charges by the Couzens Committee of the United States Senate of discrimination by the Bureau of Internal Revenue in favor of the copper companies, it becomes a matter of interest to consider the basis for these charges. The point to be determined is whether they rest on a foundation of fact or are purely and simply matters of opinion. The whole question boils down to the purpose of the Tax Law as applied to mines, which is to afford the same protection to them as provided for all other taxpayers, namely: that they were to be allowed to pay back as depletion to their owners the values of their properties as they existed or were estimated to exist on May 1, 1913, free from tax. To determine these valuations it was? necessary to select from a wide range of possibilities a set of values for a number of essential factors.

Jan 1, 1925

By R. W. Bartlett

Hydrometallurgical processes for copper flotation concentrates avoid the SO2 emission problems associated with smelting, but they require oxidation of copper to obtain solubilization during leaching. The resulting copper ions must be reduced to metal, which normally requires large amounts of energy or energy equivalent reagents such as hydrogen or iron. The conventional reduction method is electrowinning which is labor intensive and involves 35 to 40% of both the entire operating and capital cost of the better hydrometallurgical processes. There is considerable incentive to reduce the cost and energy required for producing copper ions from solution. Furthermore, it is thermodynamically possible to do this using sulfur contained in the concentrate itself. In fact, the thermodynamic potential of oxidizing sulfur from S+4 to the S+6 state is greater than required to reduce cupric ion to metallic copper.1 Realization of this potential in a practical process would significantly reduce the cost of hydrometallurgical processing of sulfide concentrates because sulfur would become a cost-free reducing agent. An apparently practical method is to use SO: and HSO3 ions, obtained by ammonia scrubbing of SO2 from a concentrate roaster gas, to reduce leached copper to metal in an autoclave. This reduction step is the Jumau 1907 A series of steps, including sulfite reduction, constituting a crude flowsheet was suggested by Parker in 1976.3 The present paper describes a detailed complete. sulfite process flowsheet based on materials and energy balances, discusses the chemistry, and discusses further work that is required to develop this process. Process materials, utilities costs, and energy requirements are projected to be unusually low because of the utilization of sulfur from the concentrate. Process Description The sulfite reduction process requires a copper concentrate containing approximately as much sulfur on a weight basis as copper. For those few concentrates that are deficient in sulfur, additional sulfur can generally be obtained by allowing more pyrite to float in the concentration process. The process steps, shown in [Fig. 1], are (1) a fluid bed roast of the concentrate, (2) a leach of the roaster calcine using recycled barren liquor from the reduction autoclave, (3) scrubbing of the SO2 gas with ammonium sulfite liquor to produce ammonium bisulfite, (4) reduction and precipitation of copper in an autoclave by combining part of the SO2 scrubber effluent (SO5

Jan 1, 1980

By W. A. Yuill

A process has been developed for increasing the grade of chalcopyrite concentrates from about 25 percent copper to about 60 percent copper. The concentrates are treated with two steps of autoclaving and then by flotation. Because of the lower associated weight, copper in the concentrates could be shipped and smelted at less cost than untreated concentrates. Use of the process can also increase overall yield from an orebody if a lower grade concentrate is sent to enrichment than would be sent to a smelter. Potential applications of the process would be for treating concentrates that must be transported a long distance for smelting or for treating concentrates so as to increase the capacity of a smelter.

Jan 1, 1984

By John O. Marsden

In early 1998, Phelps Dodge embarked upon a program to investigate alternative technology for the extraction and recovery of copper, and other metal values, from copper concentrates. A review of all of the available process alternatives indicated that many were unsuitable for large-scale commercial application and others were clearly targeted at niche applications where contaminants (and consequently smelter penalties) were driving the decision-making. The result of this development program is the world?s first commercial application of high temperature pressure leaching of chalcopyrite concentrate at Bagdad, Arizona, which was commissioned in the second quarter of 2003. Other related processes are in various stages of development. In addition to the obvious considerations, the economic aspects of this technology have been considered in detail at each step in the development of the program. The technology developments in this area are reviewed with emphasis on the economic viability and commercial application.

Jan 1, 2003

By R. E. Brewer

Phelps Dodge Mining Company (Phelps Dodge) is operating a pressure leach vessel (PLV) for copper concentrates at the Bagdad Mine facility in Arizona. This single PLV unit was engineered and constructed based on continuous laboratory testing, reaction chemistry and kinetics, and available pressure vessel technology. The challenge of designing for optimum process retention time, agitation, oxygen mass transfer, temperature and pressure will be reviewed in this paper. The PLV interior lining, agitation equipment and compartment configuration chosen for the PLV will also be reviewed.

Jan 1, 2004

By G. Zárate, G. Tapia

Anglo American Chile has been doing copper concentrate leaching studies at the laboratory and small pilot plant scale since 1992. Among the processes evaluated are bacterial leaching and pressure leaching using Mantos Blancos, Salobo Project and Quellaveco Project concentrates. Late in the nineties, it was decided to focus the research on pressure leaching, either medium or total pressure oxidation. Preliminary scouting tests were run with Mantos Blancos concentrate samples during 2000, which were repeated during 2003 together with samples of Los Bronces and El Soldado concentrates, in order to define the most suitable concentrate for running a pilot plant campaign at AARL, which was carried out in two campaigns, a first one from November 2003 through January 2004 and a second one during April 2004. Mantos Blancos was the selected concentrate and the laboratory and pilot plant results are discussed in this paper together with some preliminary economic estimates.

Jan 1, 2004

By T. Yeomans

NVI Mining’s Myra Falls operation is a forty-year-old underground mining operation feeding a 4,000 metric tonnes per day concentrator. The current mill, built in 1985, was designed to process a massive sulphide ore-body. This orebody, known as the HW, contains about 50% pyrite. Minerals of value are chalcopyrite and sphalerite, with precious metals occurring in all mineral types. By 1997, mining of a new ore-body, known as Battle-Gap, commenced. Mineralogically complex, the new orebody contains about 15% pyrite copper minerals, including primary chalcopyrite and secondary bornite, chalcocite and tennantite, combined with increased galena and sphalerite. Flotation performance generally decreased with increasing amounts of Battle-Gap ore types, with both copper concentrate quality (increased lead and zinc content) and copper recoveries being affected. Following significant metallurgical test-work, a new, fully sequential copper-lead-zinc flotation flow sheet and reagent scheme was developed. This sequential flow sheet was progressively installed and commissioned during 2006. Copper concentrate lead and zinc smelter penalty contents have been reduced by some 40%, while increasing copper recovery by 10%.

Jan 1, 2009

By Trevor Yeomans

"NVI Mining’s Myra Falls Operation is a forty year old underground mining operation feeding a 4000 mtpd concentrator.The current Mill, built in 1985, was designed to process a large massive sulphide ore-body. This ore-body, known as the “HW”, contains about 50% pyrite. Value minerals are chalcopyrite, and sphalerite, with precious metals occurring in all mineral types.By 1997, mining of a new ore-body, known as “Battle-Gap” commenced. Mineralogically complex, the new orebody contains about 15% pyrite, copper minerals including primary chalcopyrite, secondary bornite, chalcocite and tennantite, combined with increased galena and sphalerite.Flotation performance generally decreased with increasing amounts of Battle-Gap ore types, with both copper concentrate quality (increased lead and zinc content) and copper recoveries being affected.Following significant metallurgical testwork, a new fully sequential copper-lead-zinc flotation flowsheet and reagent scheme was developed.This sequential flowsheet was progressively installed and commissioned during 2006.Copper concentrate lead and zinc smelter penalty contents have been reduced by some 40%, whilst increasing copper recovery by 10%.INTRODUCTIONNVI Mining operates the Myra Falls Operation near Campbell River on Vancouver Island, BC. The complex has been in operation since 1966 with mining being predominantly underground. One unique feature of Myra Falls is that the mine operates within a BC Provincial Park.In 1985, a new large massive sulphide ore-body (HW) had been defined, and metallurgically tested. A new concentrator was built based upon this massive pyrite, chalcopyrite, and sphalerite mineralization producing copper and zinc concentrates by conventional flotation methods."

Jan 1, 2008

By Mauricio Guimarães Bergerman

At the regrind or secondary grinding stages of base metal ores, the concern about energy efficiency is usually extremely important. In general, the specific energy consumption at this stage is higher than during the primary grinding stages. In addition, a tendency of finer grinding has been observed in the new projects of metallic minerals, due to the finer mineral liberation, which leads to increasing capital and operational costs. At Vale?s new copper projects in Brazil, regrind sizes with a P80 of approximately 40 to 20 µm have been necessary. Conventional ball mills, adequate for primary grinding, usually bring about low energy efficiency with products bellow 50 µm. Vale is using the vertical mill on its copper projects ? Sossego and Salobo, in order to optimize the regrind of rougher concentrates and reduce energy consumption. This paper presents the results of industrial surveys carried out on the Vertical mill circuits of Sossego plant during its initial years of operation, and compares these results with the design criteria and laboratory regrind tests used for the mill scale up. Keywords: grinding, regrinding, copper, vertical mill, stirred mill

Sep 1, 2012

By Toshiyuki Kawai

Copper concentrate smelting directly in P-S converters was started at Onahama Smelter in 1983. Thereafter, the tonnage of concentrate smelted through the P-S converters increased, such that by 2002, it equaled the tonnage smelted through the reverberatory furnaces. In 2003, 56% of the total copper concentrates treated at Onahama Smelter were processed directly through the P-S converters. Technical improvements are reviewed and advantages of direct smelting in P-S converters are reported.

Jan 1, 2005

By Hare S. M, Griffin P. T, O&apos

Chalcopyrite ore and converter slag treatment to produce copper concentrates is undertaken at the Copper Concentrator of Mount Isa Mines Limited. This concentrator was commissioned in May 1973, replacing the original No 1 Lead-Zinc Concentrator, for the processing of all chalcopyrite ore coming from Mount Isa Mine.Treatment of chalcopyrite ore at rates of up to 125 000 l/week is described here; the treatment of converter slag is described later.

Jan 1, 1993

By Hare S. M, Lumsdaine I, O&apos

Converter slag assaying 3.7 per cent Cu is produced by Mount Isa copper smelter at a rate of up to 180 000 tonnes per annum.Priority is given to the processing of ore, however concentrator capacity periodically exceeds that of the mine and the opportunity is taken to treat converter slag. It is processed in batch campaigns, typically through the Copper Concentrator, but the Hilton Concentrator has also been used. Slag and ore are kept as separate as possible, since each has an adverse effect on the flotation performance of the other.

Jan 1, 1993

By E. N. Mounsey

Ausmelt Technology was selected by Bindura Nickel Corporation (BNC) as the basis for a new copper circuit to process nickel plant copper leach residue to blister, anode and finally cathode copper and recover PGM values at the Bindura Smelter Refinery (BSR). The process selected, based on Ausmelt Technology to smelt and convert the residue to matte and blister copper and produce a discard slag, is novel incorporating smelting, converting and slag reduction process steps in a single Ausmelt vessel. The process demands a high quality blister copper product with low levels of arsenic and nickel. This paper reviews the project and the initial operational results to the end of 1997 of the Ausmelt Top Submerged Lancing (TSL) smelter and discusses the implications for the wider TSL converter market.

Jan 1, 1998

By Kim Fagerlund

Boliden Harjavalta Oy operates a copper smelter in Finland, which is based on flash smelting together with Peirce-Smith converting of copper matte. This paper reviews a case study when a high arsenic feed mixture was successfully eliminated in order to guarantee the anode quality. Based on theoretical and practical knowledge in slag chemistry at Outokumpu Research, the CaO-fluxing practice was introduced. Description of a thermodynamic simulation using MTDATA and implementation of the results into converting practice is given. Results showed that the impurity fluctuations in raw materials could be effectively controlled by efficient utilization of different slag chemistry.

Jan 1, 2005

By Lauri Holappa

Copper converting is mostly performed in Peirce-Smith converters. The process is still very similar to the original one developed in the early 20th century. It is a batch process in a horizontal cylindrical reactor with air blowing via tuyeres along the length of the reactor side. The Peirce-Smith converter has its roots in Bessemer&apos;s inventions in the 1850&apos;s as he tested both horizontal and vertical vessels for steel converting. However, for steel making a vertical pear-shaped converter with air bottom-blowing was established. In the course of time, there was a distinct necessity for oxygen usage instead of air due to harmful effects of nitrogen to steel properties and its cooling effect. In-bath blowing of oxygen did not succeed, however, and top-blown oxygen converter with oxygen lance (LD or BOF) was first developed in he 1950&apos;s and emerged rapidly to a dominating position. Somewhat later oxygen bottom-blowing via special shrouded tuyeres was developed (OBM or Q-BOP) but has attained only fractional share. Bottom-blowing technique had, however, impact on the current combined blown converters which adopted the benefits of both top and bottom blowing. In copper converting Peirce-Smith type converters have kept their position until today although new processes have been introduced and at least few of them seem to be emerging. In one respect copper converting has a clear leading position i.e. in the progress of continuous converting processes (Mitsubishi, Ausmelt). These processes utilise top-blowing lances instead of in-bath blowing tuyeres. A different principle for continuous converting is flash converting by Kennecott Outokumpu, which makes smelting and converting independent of each others. Solid ground matte is used as a feed for flash converting in a reactor shaft similar to Outokumpu matte flash smelting. A comparable process has been presented by Inco too. The paper discusses principal physico-chemical and process technical similarities and differences in copper and steel converting. Although these processes are metallurgically quite different it seems that during the modem history of ferrous and non-ferrous metallurgy, starting from the 1850&apos;s, the exchange of knowledge between these disciplines has been quite intensive. In long-term perspective, many similarities in the development can be found but also divergences due to different ambient circumstances.

Jan 1, 2003

By R. B. Nourse, R. Hundermark, B. Masters, H. Joubert

The new 30 MVA slag cleaning furnace at Waterval Smelter, Rustenburg Platinum, South Africa, has been in operation since 10 March 2003. The furnace design and operation face a number of challenges including a wide variety of feed materials and resulting bath conditions. Furnace feed varies in combinations of granulated Waterval Ausmelt Converter Slag (WACS), hot fed Peirce-Smith converter slag, concentrate and reverts. Coke is added as reductant. The resulting slag composition varies considerably and regularly. For the case feeding WACS an almost pure fayalitic slag (60%+ FeO, 30%+ SiO2) with a very low viscosity results. Due to the widely varying slag conditions, including a superheated fayalitic slag with high fluidity, a robust sidewall design is required adjacent to the slag bath. For this reason the furnace sidewall is equipped with Pyromet&apos;s MAXICOOL® high intensity copper cooling system. The design and installation of the copper coolers as well as tap holes are discussed. To cope with the variation in bath conditions and high slag superheat (350°C when tapped at 1650°C), the copper coolers are capable of removing peak sidewall heat fluxes above 500 kW/m2. Feedback is included on the performance of the Pyromet MAXICOO® copper cooling system for the first two years of operation. The paper further includes an overview on the design and performance of the tap holes, furnace shell, furnace roof and copper slag spouts.

Jan 1, 2005