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By K. Koike, N. Masuko, Y. Takasaki

At the Zinc & Lead'95 symposium, one of the authors presented several modifications of Pb-Ag-Ca alloy anodes for the zinc electrowinning process, and also, the behavior of some organic additives in the electrolyte. In this study, the mechanical properties of the electrode materials composed by Pb-0.5 wt% Ag-(0~0.8) wt% Ca were studied by the tensile and Vickers hardness methods. Furthermore, experimental research on the influence of rolling and annealing treatments on the electrolytic behavior and mechanical properties of these alloys was carried out. The tensile strengths of the annealed Pb-0.5 wt% Ag-(0.1-0.8) wt% Ca alloys were the same or higher than those of the as-cast Pb-0.9 wt% Ag alloy. The anode potential using the as-cast Pb-0.5 wt% Ag-0.6 wt% Ca alloy decreased compared to the as-cast Pb-0.9 wt% Ag alloy. Moreover, the annealed Pb-0.5 wt% Ag - 0.6 wt% Ca anode after rolling showed a lower anode potential than the as-cast Pb-0.5 wt% Ag - 0.6 wt% Ca anode.

Jan 1, 2000

By R. W. Lawrence

A method by which it is possible at an oxidation potential below 680 mV SHE to achieve quantitative conversion of sulphide sulphur to elemental sulphur instead of sulphate in a biological leach of chalcopyrite and bornite has been investigated. The process chemistry, biological phenomena and process flowsheet options are discussed. Several copper concentrates have been evaluated and have been found to be amenable to th process. Results to study the progression of chalcopyrite decomposition during the leach are presented and mineralogical factors affecting the requirement for silver addition are discussed. A chloride addition method for improving silver recovery is described. The results of a small scale continuous leach test have demonstrated that continuous operation can be carried out with high copper extraction and stoichiometric conversion of sulphide sulphur to elemental sulphur.

Jan 1, 1985

By Xiao Yang

Calcium oxide is one of the major fluxing agents used for steel refining. Because unreacted CaO remains in slag in most cases, discharged slag is difficult to be recycled smoothly. With the growth of environmental concerns, improving utilization efficiency of CaO is required to reduce the amount of discharged slag and energy consumption. The authors have reported a resource and energy-saving process by using multi-phase flux for dephosphorization. Reaction behavior of phosphorus in multi-phase CaO–FeOX–SiO2–P2O5 flux system was investigated. Formation mechanism of P2O5-rich phases was clarified at the interface between solid CaO or 2CaO, SiO2 and molten CaO–FeOx–SiO2–P2O5 slag. Phase relationship and thermodynamic properties of 2CaO, SiO2-3CaO, P2O5 solid solution were observed to confirm the condensation of P2O5 in solid phase. Based on these results, reduction of CaO consumption, discharged slag curtailment, and energy-saving effects in the novel refining process by using multi-phase flux were discussed.

By Bin Zhu, Mingmei Zhu, Hong Zhou, Xiaofei Dou, Bin Xie, Yu Wang, Tiancheng Lin

In this study, the formation of 2CaO·SiO2 and 3CaO·P2O5 solid solution (nC2S-C3P) in hot metal dephosphorization process is discussed. The variations of CaO and SiO2 mass, both in nC2S-C3P solid solution phase and liquid slag phase with increasing P2O5 mass in CaO-SiO2-FetO-P2O5 slags were calculated using Factsage software. CaO-SiO2-FeOt-P2O5 slag containing up to 18 % P2O5 was melted at 1823 K then cooled to 1673 K. The contents of CaO and SiO2 both in the solid solution phase and liquid slag phase of quenched samples were observed and analyzed by SEM/EDS. Both the calculated and experimental results show that the contents of CaO and SiO2 in the solid solution phase decreased, just opposite in the liquid slag phase, with increasing P2O5 content in the slags. Formation mechanism of nC2S-C3P solid solution in the slags was derived based on the ionic structure theory of molten slag. The SiO4 4- in the nC2S-C3P solid solution can be replaced by PO4 3- in the liquid phase resulting in the increasing of free Ca2+ and SiO4 4- in liquid phase.

Jan 1, 2015

By R. L. Miller

An industrial wastewater, including plating wastes, is treated with sodium sulfide and ferrous sulfate to form a sulfide-oxide precipitate containing chromium and other toxic metals. Hydrocarbons, in the water, coat the sulfide-oxide particles, impeding metal recovery. Froth flotation, without reagent addition, was found to recover 93.9% of the solids from the sludge with simultaneous rejection of 89% of the water. Methyl isobutyl carbinol (MIBC) improved recovery and potassium amyl xanthate improved both recovery and grade. The process design has wastewater feed (without MIBC) to the rougher circuit. The rougher concentrate is conditioned with MIBC and fed to a cleaner circuit to achieve a high grade concentrate. About 95% of the water is recirculated to the waste treatment plant.

Jan 1, 1992

By B. Yawney

A copper-nickel sulfide concentrate was leached in water at about 2000C under oxygen pressures of about 0.35 MPa. Oxidation of sulfide to sulfate was virtually complete in about 30 minutes, resulting in greater than 95% copper and 98% nickel extractions. Leaching temperature was shown to be the most important factor affecting metal extractions. Measured rates of oxygen consumption were used to develop batch kinetic data. Mini-plant continuous pressure leaching tests agreed with batch leach results. A proposed overall flowsheet is briefly examined.

Jan 1, 1976

By Masayasu Kawahara

"The extraction and separation of nickel, cobalt and copper were investigated with a solvent impregnated resin which was produced from UX84-I and XAD-4. The extraction behavior of metals with the solvent impregnated resin is similar to that with solvent extraction. The extraction ability of solvent impregnated resin improves when using finer particle resin and with a higher concentration of extraction solvent.IntroductionSolvent extraction is well known-as a method for separation of metallic elements in solution. However there are problems with intermediate phase and solvent loss. Ion exchange resin is also used in hydrometallurgy, but it is not fit for separating a large quantity of metals. Amberlite XAD-4 resin is known as a absorption agent of organic matter in water. It is expected that the resin which has the ability of extracting metals can be made by absorbing organic extractant on XAD-4 resin In this study, solvent impregnated resin (SIR) was produced from LIX84-I and XAD-4, and extraction and separation of nickel, cobalt and copper were conducted using SIR. The extraction properties of SIR were investigated, and a comparison between ordinary solvent extraction and SIR was made.Experimental ProcedureAmberlite XAD-4 was used as absorption resin for the organic extractant in this study. The physical properties of XAD-4 are shown in Table I. Amber lite XAD-4 has a large surface area and shows good ability to absorb organic matter in water, but the: resin does not have the ability of exchanging metallic ion."

Jan 1, 2000

By I. S. Blair

The Mufulira smelter (1, 2,' 3, 4,) is one of three smelters situated on the Zambian Copperbelt to treat concentrates and other intermediate products produced in the area. Two mining companies, Roan Consolidated Mines Limited (RCM) and Nchanga Consolidated Copper Mines Limited (NCCM) each have 51% Zambia Government participation. RCM operates two smelters, one at its Mufulira Division and the other at its Luanshya Dl vision. NCCM operates one smelter at its Rokana Division (5). The Mufulira smelter was commissioned in 1937 with two reverberatory furnaces and four converters. The final product, blister copper, was cast from a holding furnace. In 1952 two anode furnaces and a common casting wheel were installed to feed a new and adjacent electrolytic refinery. In 1956 the smelter was extended to accommodate increasing production from the new Chibuluma mine and expansion at Mufulira mine. A third reverberatory furnace, a fifth converter and duplicate anode processing and casting facility were added. The Copperbelt smelters suffered a loss of production potential when higher ash content coals were used in reverberatory smelting furnaces from 1966 onwards. This, and the planned production increases from Mufulira and other RCM mines, indicated a smelting shortfall of 55,000 tonnes of copper a year in the combined capacity of the RCM smelters.

Jan 1, 1976

The merging of flash smelting with spray converting in one single furnace is proposed as a continuous copper making process. The proposed process is expected to produce blister copper, clean slag and gas with high sulfur dioxide concentration. Some high grade matte is an intermediate product recycled within the proposed furnace, spray-atomized and converted. The thermodynamic equilibria of the simultaneous oxidations which take place during copper smelting and converting are discussed in order to establish the gradient of oxygen potential along the furnace required for smooth operation. The proposed process is a dynamic system designed to reach equilibrium only close to the slag skimming end of the furnace. The process feasibility is based on the predesigned significant difference between the rates of converting reactions in a finely dispersed system and the rates of deconverting reactions between immiscible liquid phases of restricted interface area. The converting of small stationary droplets of matte, in a silica saturated system, was studied in the laboratory for extremely short oxidation times (10 to 120 seconds). These experiments led to some interesting observations on reaction mechanisms and kinetic trends. The droplet size and the matte grade affected very strongly the oxidation rate and the proportion of the different products obtained.

Jan 1, 1976

By R. D. Osmanson

A series of batch roasting tests was performed on minus 1/2 inch gold bearing refractory samples from the Gold Acres deposit, and 67 select particles were polished and the sulfide alterations were examined microscopically. The suite of specimens represents two hour batch roast treatments at 600°C, 700°C, 725°C, 750°c, 800°C, and 900°C. Auriferous pyrites were found to alter sequencially to pyrrhotite, magnetite and finally to hematite. It was found that the decrease in cyanidation extraction with high temperature roasting is likely due to the increas ing stability of magnetite and more crystalline hematite, and the decrease in porosity with these phases which occlude the gold. Gold has been noted to occur very early in the mineralizing sequence, with gold being found in quartz and in clay/carbonate rock matrix. No gold was found in the two pyrite phases, but numerous gold grains were found in pyrrhotite after pyrite and in subsequent iron oxides. Gold mineralization had apparently ended by the time base metals were introduced into the deposit.

Jan 1, 1987

By Sergey Shabala

Green leaf tissues convert solar energy into the energy of chemical bonds of sugar molecules during the process of photosynthesis. The efficiency of this conversion can be much higher than efficiency of any of the currently known silicon-based solar panels. Importantly, vast amounts of CO2 are assimilated during this process. The efficiency of photosynthesis is critically dependent on the availability of a large number of nutrients, among which metal cations such as K, Ca, Mg, Cu, Zn, Fe, Mn, and Ni play a key role. In this paper I summarize basic requirements and major functions for each of these essential nutrients in plant photosynthesis, both at the whole-plant and molecular level. I will talk about how these requirements may be affected by the global climate trends and discuss the prospects of creating artificial photosynthetic “bioreactors” for efficient energy conversion and CO2 assimilation.

Jan 1, 2009

By D. A. Milligan

Ammonia leaching of copper sulfide concentrates, matte, scrap or other sources of copper yields copper amnine sulfate solutions which are converted to metal in a two-stage process involving precipitation of inter- mediate sulfites and their thermal decomposition to copper powder of high purity. Processing details and results are given. Considerable simplification in circuitry and cost savings are indicated. Energy requirements are reduced by substitution of copper powder precipitation and ammonium sulfate crystallization for solvent extraction, electrowinning and lime boil.

Jan 1, 1976

By E. O. Stensholt

The Hybinette electrorefining process for production of cathode nickel was in use at Falconbridge Nikkelverk A/S, Norway, until 1975 when conversion to the new chlorine leach process was started. In 1978 the conversion to electrowinning had been completed, but further major improvements to the process have been made up to 1987. After a major expansion program in 1985 - 87 the refinery has a nickel capacity of 54 000 mt Ni/year. The main feed materials to the process are converter matte from the Falconbridge smelter in Sudbury, Ontario, Canada, and from the BCL smelter in Selebi-Phikwe, Botswana. A typical composition of the mixed feed to the refinery is: 40 - 45% Ni, 25 - 30% Cu, 20 - 22% S, 2 - 3% Fe and 1.0 - 1.5% Co. The matte is chlorine-leached in the first process step, nickel being dissolved selectively in a train of leach-cementation tanks. Essentially all copper and sulphur that was present in matte appears in the leach residue as CuS. In 1986 an autoclave treatment stage was added to the process to improve nickel extraction from the leach residue. The resulting strong nickel chloride solution is purified by the process steps (1) precipitation of iron and arsenic, (2) solvent extraction of cobalt and other minor elements and (3) precipitation of lead and the remaining impurities. The nickel solution is then fed to the tankhouse, where various nickel products are produced by electrowinning. The chlorine gas that is produced during the electrolysis is returned directly to the chlorine leaching tanks to dissolve more matte. After stripping from the loaded organic phase and solution purification cobalt is also produced by electrowinning. The copper sulphide leach residue, which contains some nickel, is dead- roasted in fluidized-bed roasters. The sulphur is recovered and sold as sulphuric acid. The calcine is leached in spent copper electrolyte, from which cathode copper is produced by electrowinning. Precious metals and the remaining base metals are recovered from the copper leach residue. The Falconbridge chlorine process is very versatile. It can accept a wide range of impurities in the feed and still produce metals of exceptional purity.

Jan 1, 1988

By Peter Brown

The present work involves the study of Chemical Vapor Deposition (CVD) of silicon carbide from the methyltrichlorosilane (MTS) - hydrogen system. A computer program, SOLGASMIX has been employed to calculate the equilibrium compositions of the gaseous and condensed phases in the MTS - H2 system. The results indicate the range of parameters such as temperature, pressure and species concentrations, required to deposit the single phase Sic under thermodynamic equilibrium conditions. A mathematical model for describing the transport phenomena and the rate of deposition of Sic onto a thin rod in a vertically-oriented hot-walled CVD reactor has been developed. Exper iments were conducted using the same operating conditions as those used in the model in order to validate the mathematical predictions. The predicted rates of deposition were found to be in good agreement with those obtained experimentally. The model can be used to study the effects of various operating parameters on the rate of Sic deposition on a fiber.

Jan 1, 1992

By Brent W. Madsen

The acid-ferric sulfate leaching of low-grade sulfide copper mine waste provides more than 10 percent of the U.S. supply of new copper, but recoveries and leaching rates are very low. In Federal Bureau of Mines research to improve the leaching process, experiments on 5- to 10-ton lots of mine-run ore were more revealing than bench-scale tests on fine ore. This report discusses the results obtained by leaching run-of-mine and minus 6-inch plus 112-inch samples of low-grade chalcocite-bearing mine waste. The samples were leached for 461 days and 61.6 and 37.7 percent of the copper was extracted from the minus 6-inch plus -112-inch and run-of- mine samples, respectively. A mixed kinetics model was applied successfully to the leaching of the sample with the fines removed where bulk solution transport was not rate controlling. The model, when applied to the run-of- mine ore, gave somewhat higher predicted copper extractions than experimentally determined. Subsequently, oxygen was introduced into the run-of- mine ore, leaching continued for an additional 312 days, and an additional 53.0 percent of the copper was extracted. Before the leaching experiment was dismantled and the residue sampled, the ore was treated with sulfur dioxide and its permeability was improved markedly. The results of each of these procedures are interpreted in relation to the attendant condition of the test.

Jan 1, 1976

By R. A. Alock

The character, occurrence and genesis of the three types of nickel deposits, sulfide, laterite and deep-sea nodule are described, highlighting the differences in mineralogy and chemical composition that effect the extractive metallurgy of nickel and associated byproducts. World nickel production and current resources by deposit type and geographic location are also reviewed.

Jan 1, 1988

By Sue Ritz

Gold extraction has relied on coconut carbon in adsorption circuits for several decades. IXOS®, invented by 6th Wave Innovations of Salt Lake City, Utah, is a new type of adsorbent that exceeds the selectivity of carbon and improves on the mechanical properties of other available resins. IXOS selectivity against copper cyanide complexes allows it to be used in circuits that currently struggle with copper contamination. IXOS potential to eliminate or reduce the application of SART technology will expand the opportunities to exploit more complex precious metals deposits. This paper will describe the technology behind IXOS molecular imprint, its comparison to carbon adsorption in the field, and the effect of efficiency on long-term recovery from heap leach METSIM simulation software.

By H. -Peter Sattler

The laser sorting technique for replacing the handpicking of the non-ferrous metal scrap from auto shredders was first introduced 1990, The process is based on the use of atomic emission spectroscopy induced by lasers to identify the scrap particles and sort them with adequate systems into main fractions of Al-, Zn-and Cu-alloys, stainless steel and lead, In the meantime the handling technique for the scrap, i.e. the singling and sorting technique as well as the analytical detector were further developed for more efficiency and higher reliability, i,e. less maintenance, minor sorting errors etc. These developments were mainly based, on and initiated by operational experiences from the startup phase of a first production plant. Part of the on going research work was the development of an ,aluminium detector which makes it the first time possible to sort mixed aluminium scrap automatically into specific alloy fractions.

Jan 1, 1995

By L. J. Lutz

The U.S. Bureau of Mines has developed an electrorefining process to recover Ni, Co, Cr, and other metals from mixed and contaminated superalloy scrap. The superalloy scrap is cleaned then directly electrodeposited as a Ni -Co deposit. The process recovers approximately 90 pct of the Ni-Co available from the scrap as a 99t pct pure Ni-Co alloy. During electrorefining the electrolyte undergoes a purification process which results in Cr and other refractory metals reporting to the purification residue. Operating data is given for a 3-L compartmented electrolytic cell. Techniques for the removal of Fe, Cr, and trace impurities from the recycled electrolyte are outlined. The economics of a facility producing 15 mt/d of a Ni-Co alloy are also discussed.

Jan 1, 1993

By Chen Ya

Layered LiMn0.5Ni0.3Co0.2O2 was prepared from Mn(NO3)2, Ni(NO3)2, Co(NO3)2 and LiOH by co-precipitation and solid state reaction. XRD, C-V and galvanostatic charge-discharge methods were used to examine the structural and electrochemical characteristics. The LiMn0.5Ni0.3Co0.2O2 delivered a high discharge capacity of 167 mAh/g between 2.5 and 4.3 V with excellent cycling stability.

Jan 1, 2006