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By A. P. Paiva

Studies focused on the recovery of precious metals by solvent extraction have recently increased. either to devise suitable solvents or to clarify the involved extraction mechanisms and kinetics. In this work, a comparison of the performance exhibited by triphenylphosphine (TPP), bis-(penty1thio)-ethane (BPTE) and triisobutylphosphine sulfide CI"I'BPS) in the extraction of silver from acidic chloride media is carried out. Extraction and stripping isotherms involving the three extractants are presented together with the results obtained on the selectivity of silver extraction against high contents of iron, copper and lead. Attempts to clarify the mechanism of silver extraction by TIBPS were performed by studying the distribution coefficient dependence on ligand and chloride concentrations. The effect of temperature on extraction and stripping is also considered. The interfacial tension for the system, measured with different concentrations of extractant in the organic phase, is analysed.

Jan 1, 1994

By Nagahiro Saito, Takahiro Ishizaki, Daisuke Yata, Akio Fuwa

"Ternary compound-semiconductor, CuInTe2, film was deposited cathodically under potentiostatic conditions on titanium substrate from aqueous solution containing CuCl2, InCl3, TeO and HCI in this study.The deposition parameters such as electrolytic solution composition, potential and temperature were optimized· for electrodeposition of CulnTe2. Structural characterization of the: deposited film was also carried out using XRD and SEM. Electrodeposited films Were prepared and analyzed for their chemical composition using ICP. The ICP analyses showed that the stoichiometry of the films could be controlled by (1) the deposition potential and/or by (2) the electrolytic bath composition.1. IntroductionThe ternary chalcopyrite type compound semiconductors exhibit such electrical, chemical and optical properties that make them promising for photovoltaic devices (Kazmerski and Wagner 1985)(1). Many researches have been devoted to the characterization of these materials for over one and a half decades. High efficiency solar cells based on p-CulnSe2/n-CdS have been reported (Mickelson and Chen 1983)(2).The electrodeposited fil ns of semiconductors have more advantage than other physical and chemical deposition techniques; it is relatively easy and economic technique and the semiconducting properties such as band-gap variation, concentration of dopant stoichiometric composition and so on can be controlled with some accuracy. The electrodepositions of the ternary CuInS2, CulnSe2 have been reported by many workers (Hodes and Cahen 1986, Bhattachaya 1983, Singh et al 1986, Lokhande and can be prepared as both p-type and n-type. Therefore, its potential for application in optical and photovoltaic devices is one of the great interests. Bhattacharaya and Rajeshwar (1986)(9) reported the electrodeposition of CuInTe2 from an acidic bath· on indium tin oxide (ITO) and Mo substrate. Triethanolamine (TEA) was used as a complexing agent for indium. The electrodeposited CuInTe2 films were non-stoichiometric in composition under all their experimental conditions by electron-probe micro analysis (EPMA) and Auger depth profile analysis. In this paper, we report our results on the behavior .of Cu-In-Te electrodeposition from ·acidic baths; the preparation of CulnTe2 and the effect of bath composition on electrodeposited films obtained."

Jan 1, 2000

By Corby G. Anderson

Antimony has been utilized by mankind for thousands of years. However, due to recent environmental concerns, antimony is often considered to be an impurity element, Since antimnoy is commonly found in precious metal refractory ore bodies, extraction is essential. The Sunshine Mine in Kellogg, Idaho has recovered antimony from its precious metal tetrahedrite deposits since the 1940's. A plant was initially constructed as a primary antimony source during World War II. This facility also decreased the downstream smelting penalty for the silver copper-lead concentrate and provided an industrial commodity from a potential environmental nuisance.' This paper describes the Sunshine antimony processing- facility. Specifically, the relevant process chemistries involved in concentrate leaching, metal electrowinning, and sodium antimonate production will be discussed. In addition, ongoing areas of research to improve unit operations and treatments implemented to meet environmental' discharge requirements will be covered.

Jan 1, 1992

By Charles H. Bucknam

A gold sampling optimization study was carried out in support of open pit gold mining operations for the Newmont Gold Company Bootstrap-Capstone deposits in Northeastern Nevada. Six bulk samples were tested using fire refining techniques at field particle size, to determine coarse particle heterogeneity of the ores, and by gravity .concentration methods at 10 mesh particle size, to observe any particulate gold liberation. An overall sampling model is presented for the deposit based on the pooled test information, to reconcile the results of coarse particle heterogeneity testing and the fine particle 'liberation testing. The sampling model is used to recommend procedures to control the variability of gold grade estimation due to the fundamental error (nugget effect) during sampling, sample preparation and assaying of the ores and other fundamental variances. The technique of simplified quantitative risk analysis (SQRA) is used to provide an economic justification for implementation of. recommended improvements to field sampling practices.

Jan 1, 1999

By A. Shemi

Tungsten carbide scrap (WC–Co) is material generated from the tungsten carbide manufacturing process and worn out or used carbide parts from industry. Tungsten (W) ores contain about 3 wt% WO3 and their W concentrates about 10– 75 wt% WO3. Tungsten carbide scrap typically contains 40–95 wt% W. The high intrinsic value of W and the fact that the least valuable scrap contains more W than the average tungsten ore makes tungsten carbide scrap a worthy material for re-use. This study has investigated the selective dissolution of the cobalt (Co) binder phase, and recovery of tungsten monocarbide (WC) and Co using an acid leach process. The Co component was selectively leached leaving the WC solid particles intact. Further investigation using a complexing agent-aided acid leach process that takes advantage of organic chelating agents that form a soluble tungsten complex in place of the passive insoluble tungsten oxide is envisaged. The optimal dissolution of the Co binder phase using an inorganic acid in conjunction with an organic complexing agent is a novel approach with the potential of reduced leaching times and possibility of regenerating the unreacted organic acid.

By Walter Curlook

An improved process of hydrometallurgical treatment of laterite ores of the limonitic type for the recovery of nickel and cobalt using sulphuric acid has been developed and patented [I]. The major improvement is brought about by recycling a significant portion of the "mother liquor" emanating from the pressure leaching reaction back to the feed preparation stage thereby substituting for all or at least a major proportion of the fresh water that must be added. The amount of process effluent waters is correspondingly reduced. Concomitantly with the major savings in fresh water requirements, a degree of reduction in new sulphuric acid requirements is effected, along with a savings in limestone and lime required for subsequent neutralizations. Keywords: Acid pressure leaching; Nickel laterites

Jan 1, 2004

By Gui-fang ZHANG, Zhe SHI, Lei GAO

Secondary V-recycling slag, an industrial waste containing high gallium is being dumped continuously, which causes the loss of gallium. Thus, the alkali fusion process was employed to recover gallium from this slag. The effects factors on extraction of gallium such as roasting temperature, roasting time, alkali fusion agent concentration and CaO concentration were investigated in the paper. The experimental results indicated that excessive roasting temperature and roasting time is unfavorable to the recovery rate of gallium. The appropriate roasting temperature and duration are 1000 and 2 hours, respectively; The appropriate proportioning of Na2CO3: NaOH is 2:1 when the concentration of alkali fusion agent weighs 0.4 times the mass of the slag; In order to remove SiO2 from the leaching liquor, CaO should be used as an additive in the roasting process. The appropriate concentration of CaO should weigh 0.2 times the mass of the slag. Employing these optimal alkali fusion conditions in the roasting process, gallium recovery is above 90%

Jan 1, 2015

By Yongbin Yang, Meixia Cai, Yufeng Guo, Guanzhou Qiu, Tao Jiang, Jianjun Fan

"A series of sintering and metallurgical experiments were carried out to optimize the blast furnace burden structure in this study. The results indicated that the sinter disintegrated seriously with basicity around 1.4, and after this “turning point”, the sintering properties were improved gradually with the increase of basicity of sinter. Based on the results, an idea was put forward which focused on producing two different kinds of sinter with different basicity. Different with the original burden structure of 85% sinter with basicity 1.4-1.5 + 15% pellets, an alternative new BF burden structure was proposed by the following structure?44% sinter with basicity 1.05 + 44% sinter with basicity 1.87 + 12% pellets. This new alternative BF burden structure has been adopted by some steel plants.IntroductionIn recent years, conventional blast furnace (BF) process was put into operation to produce hot metal in “A” steel plant. “A” steel plant owns two sets of sinter strands at present. Because the sinter production capacity exceeds the actual requirements of BF, a burden structure with 85-90% self-made sinter + 10-15% purchased pellets is adopted to take advantage of the low–cost sinter. But the serious problems encountered with the BF operators are that the sinter disintegrated easily during transportation and storage process. Therefore, the rate of return fines is very high after sinter being screened."

Jan 1, 2012

By Shilai Yuan, Huimin Lu, Liyuan Zhao, Zhijiang Gao

"A purification process was developed to removal impurity element boron from metallurgical grade silicon by electromagnetic induction slag melting. Vacuum melting furnace was used to purify boron in both Al2O3-MgO-CaO-SiO2 slag system and Al2O3-CaO-SiO2 slag system. The relationship between different slag chemistry and the removal of boron in silicon were studied using Back Propagation (BP) Neural Network model. The best slag chemistry for the removal of boron was predicted by Genetic Algorithm (GA) contributed by the use of Matlab. The results show that the mass fraction of Boron in silicon is reduced from 11.7496×10-6 to 2.3259×10-6 after slag melting in 28.96%Al2O3-3.43%MgO-36.24%CaO-31.37%SiO2 slag system. The relative error obtained with GA-BP Neural Network model was below 0.35%.IntroductionSolar energy, due to its low green house gas emissions, has been paid great attention. Accordingly, the demand for solar grade silicon (SoG-Si) has been growing rapidly. The quest for developing a low-cost method of producing SoG-Si has been moving forward in the last decade. Refining of metallurgical-grade silicon (MG-Si) attracts research interest due to its low material and energy costs, and more environmentally friendly technology in comparison to the traditional Simens process. As such, impurity elements in MG-Si must be removed to low levels for the photovoltaics (PV) cell to operate at optimum efficiency. Most impurities in metallurgical-grade silicon (MG-Si), especially metal elements, can be separated by directional solidification and the zone refining plasma-arc method [1, 2]. The main focus in refining MG-Si to SoG-Si is on the impurities which are the most difficult to remove. Boron has relatively high segregation coefficients (KB=0.8, concentration of Boron in solid silicon divided by its concentration in liquid silicon) [3] and higher vapor tension (the saturated vapor pressure of boron and silicon are 6.78×10-7 Pa and 0.4 Pa, respectively). The vacuum melting and directional solidification methods hence have no effects on boron removal.Slag refining involves melting MG-Si in the presence of a flux to produce a slag phase that can take up the impurity elements. The mechanism of slag refining is considered to include two main steps: the oxidation of boron and the absorption of boron oxide by slag materials. The oxidation of boron, described by the oxygen potential (PO2), resulting from the equilibrium Si and SiO2. The absorption of boron oxide, described as the slag basicity, is interpreted as expected to facilitate their extraction to alkali and alkali-earth oxides in the slag phase. Thus, effective removal may be obtained for a certain slag composition. Researchers [4-5] examined the effect of boron removal in CaO-SiO2 based ternary systems and showed that the addition of excess CaO can decrease the activity of SiO2. Researchers [6-8] studied the boron behaviors in CaO-SiO2-Al2O3, CaO-SiO2-Na2O, Al2O3-BaO-SiO2, CaO-SiO2-Li2O and Al2O3-CaO-MgO-SiO2 slag systems by adding boron in the silicon to a degree. Li2O, Na2O, MgO and BaO are thought to be associated with B2O3."

Jan 1, 2014

By Hideya Sato

Refractory wear caused by melts is one of the major determining factors for the interval of shut-down repair of the metallurgical furnaces in general. Calcium ferrite slag employed by M it subishi Continuous Copper Converter, shows hi& fluidity and corrosive characteristics against most of the refractories. In order to improve the corrosion resistance of the refractories, the mechanism of MgO-based refractory wear by the calcium ferrite slag was intensively studied. The dissolution behavior of the refractory into the molten slag and the penetration of the molten slag into the refractory were observed by crucible tests and newly developed oscillating furnace tests. It was found that the wear of MgO-based refractories by the calcium ferrite slag is mainly dominated by the internal structural failure caused by the penetration of CaO-Cu20 rich slag which is formed as a result of the preferential absorption of FeO, from original slag by the spinel of the refractory.

Jan 1, 1999

By James E. Hoffmann

Oxidative pressure leaching is gradually supplanting other techniques for the decopperizing of slimes. The reasons for this include: much more rapid reactions, more compact equipment, a reduced environmental burden, and perhaps most important, a much lower final copper concentration in the decopperized slimes. This paper will discuss the reactions occurring during decopperization, the acidity and temperatures encountered and the techniques employed for discharging non condensable gasses. A spreadsheet is provided which allows adjusting the operating conditions depending upon raw (undecopperized) slimes composition, solids concentration and lixiviant composition. A typical process equipment flowsheet is provided along with a process narrative.

Jan 1, 2000

By J. Phillip Mackey

The Peter Hayes Symposium on Pyrometallurgy, held as part of Extraction 2018, is an acknowledgement of his contributions to the basic science and principles underpinning modern pyrometallurgical processing. After completing his PhD at Strathclyde University, Scotland, in 1974, Peter took up a teaching position at The University of Queensland, Brisbane, Queensland, Australia in 1976. Peter Hayes’ research at the University since then has been focussed on: • The mechanisms and kinetics of heterogeneous solid/solid, gas/solid and liquid/ solid reactions, and • Experimental measurement of phase equilibria in complex high temperature metallurgical systems. This paper provides a brief biography of Dr. Peter Hayes and gives a full listing of Peter Hayes’ publications.

By Versiane A. Leão

The hydrochloric leaching of chrysotile tailings and soapstone powders) is discussed. This two materials are magnesium compounds associated with carbonates and iron oxides. In this work the chemical and physical characterization of formers is presented and the thermodynamic analyses of leaching is carried out. Magnesium extraction is almost complete for chrysotile tailings while, for soapstone powders, one can be reached 60%. The behavior of iron and calcium impurities is discussed.

Jan 1, 1997

By M. J. Soria-Aguilar, F. R. Carrillo-Pedroza, S. Preciado-Nuñez

"Dust originated from the Iron Blast Furnace (BF) and the Basic Oxygen Furnace (BOF) containing alkalis and zinc are not completely recycled because due affects the process efficiency. These types of dust represents an economical lost as a consequence of values contents of iron and carbon. Several processes in aqueous media have been developed to treat the steelmaking dust. This work presents a laboratory study for reduce zinc and alkali containing in BF and BOF dust, using different acid and alkaline solutions and oxidants as ozone. The results obtained show that the elimination of zinc and alkalis can vary from 5 % to 80 % depending on the treatment conditions (the solution, the oxidant and the temperature). The dust obtained from this treatment can to be recycling to the steelmaking processes to recovery iron and carbon values.IntroductionEach one of the steel making industry processes generate several sub-products and residues containing useful units of iron, carbon and other elements that are or could be recycled to the productive process. Some of those sub products are the dust from the gas cleaning system of the blast furnace and the dust from BOF reductive process; such dusts are sent to the sedimentary lagoons. The dust represent a lost of 2 % of the total iron and coal mineral’s content feed to the process.The typical iron, carbon and zinc content of the blast furnace dust is approximately 40, 24 and 0.3 % and the content of the BOF dust are 50, 15 and 0.9 % weight respectively, with physics - chemical characteristics that became them susceptible to be agglomerated and feed to the blast furnace. But the high content of zinc does not allow recycling it into the process. At the blast furnace the presence of zinc propitiate the formation of film in the furnace wall; these elements are vaporized in the low section of the cube of the furnace and they are dragged by the gases to the top and condensed again at the cube. Part of these condensed materials strongly adheres to the furnace cover harming it and reducing the empty space of that section, causing eventually a severe perturbation of the ascendant gas flux [1]."

Jan 1, 2008

By Stein Tore Johansen

The flow of a liquid through a large grain filter has been considered. An improved form of the fluid dynamic equations describing flows through porous media is proposed. The theory is applicable to all Reynolds numbers. Particulate inclusions are described by passive scalar fields where only very simple mechanisms for particle removal are included. However, the deposition of particles are allowed to change the "porosity" of the filter. Hence, the flow pattern through the filter will change as the filter ages. The theory is employed to the filtration of aluminum in an industrial grain filter. The model shows that the filter is not effectively designed and there is room for considerable improvement. The dynamic "aging" process of the filter is shown to change the filtration from bed filtration to cake filtration. Finally, the predicted inclusion collection efficiencies are compared to efficiencies measured in an industrial filter. The predicted initial efficiencies are in very good agreement with experimental results.

Jan 1, 1990

By J. Pérez

A lead smelter slag with high copper, lead, arsenic and silver contents was treated at laboratory scale for arsenic removal. The slag was first leached at atmospberic pressure with sulfuric acid and oxygen. According to arsenic contents leaching of most of the arsenic was achieved in either one or two stages, producing a lead-silver rich residue and a solution high in copper and arsenic concentrations. Alternative routes were considered to treat the resulting solution. Firstly, solvent extraction of arsenic with a mixture of TBP and 2-etbyl-bexanol was tested and extractions of over 95 percent of arsenic were achieved. Secondly, a copper arsenate precipitation treatment was done to eliminate arsenic and copper from solution.

Jan 1, 1996

By Michael L. Free

Bacterial oxidation of an arsenopyritic gold-ore .concentrate was carried out with continuous-flow .operation in 14-liter, mechanically agitated reactors. Samples from the slurry reactors were separated into solid and liquid fractions and analyzed separately to determine the electrochemical rates of sulfide conversion by the bacteria and the rates of direct bacterial oxidation of sulfide. The rates of the individual mechanisms are compared to the overall rates of biooxidation in the continuous-flow slurry reactors. The results show that the commonly hypothesize4 direct and indirect mechanisms of biooxidation both play a significant role in the bacterial oxidation of sulfide ores.

Jan 1, 1991

By Nural Kuyucak

Acid mine drainage (AMD) occurs at operating and abandoned mine sites as a result of oxidation of sulphide minerals, and is characterized by low pH and heavy metal contamination (e.g. Fe, Zn, Cu, Cd, Al, Mn). It usually requires treatment before release in order to meet regulatory requirements. Chemical methods, such as lime neutralization and sulphide precipitation, are commonly used to treat AMD. However, the conventional processes require high capital and operating costs, and constant attention for many years into the future. In addition, the amount of sludge produced by conventional lime treatment processes poses an increasingly significant disposal problems at decommissioned mining operations. There is, therefore, an urgent need to develop passive, cost-effective processes for mitigating AMD and seepages originating from tailings and waste rock areas. Noranda Technology Centre (NTC) has been evaluating various methods being developed by other institutions, and exploring new ideas, to achieve passive in situ processes for treating AMD. These methods include anoxic limestone (ALS), lime-organic mixture (LOM), biosorbents (BIOS) and biotrench (BT). Current investigations at NTC involve preliminary bench-scale testing. Based on the outcome of the bench tests, a field test will be performed at a selected site. Details of the bench tests are discussed in the paper. Key Words: Acid mine drainage, treatment, metals removal, passive treatment, anoxic lime, limestone, biological treatment, sulphate reducing bacteria, biosorption, biosorbent

Jan 1, 1993

By F. J. Arriagada

Where gold is present in a finely disseminated form within sulfide particles, the porosity and surface area of the calcined material strongly influence leaching since they control the penetration of solution inside the gold-bearing grains. On the basis of a stability diagram for the Fe-O-S in the temperature range of 400 to 750aC, the expected sequence chemical transformations associated with the oxidation of pyrite is discus-sed. By using analytical techniques which include scanning electron microscopy, X-ray diffraction and BET gas adsorption, the chemical and structural changes experienced by auriferous pyrite-arsenopyrite concentrates and pure pyrite samples are studied under an oxidizing roasting atmosphere of com-position 80% N + 20% O2 in the temperature range of 450 to 750°C. The results indicate that as the roasting temperature increases, the surface area of the calcines decreases sharply after a maximum value attained at about 450-500°C. SEM micrographs show that the needle-shaped tips which characterize the surface of the oxide product become wider and the size of the pores between them is decreased when the roasting temperature is in-creased from 550 to 750°C. This effect suggests that the formation of a more "closed" hematite product at higher temperatures is due to sintering associated with a faster growth and crystallization of the oxide. A strong correlation is shown between increasing roasting temperature and increase in gold lost in cyanidation tailings, an effect which may be partly attributed to the loss in surface area of the calcines.

Jan 1, 1982

By H. Soto, J. Jara, F. Nava

A method has been developed to recover cyanide and copper from solutions containing copper complexes and thiocyanate. Copper is precipitated as copper thiocyanate or as copper cyanide depending on pH and the concentrations of thiocyanate and cyanide in the effluent. The copper precipitate assays over 50% copper and presents good dewatering characteristics. After settling it can be separated from the effluent by filtration. The solution containing the bulk of the cyanide is then oxidized with ozone to transform the remaining thiocyanate into cyanide. This results in a regenerated solution rich in free cyanide that can be recycled to the cyanidation process. Careful control of the pH of the solution is required. First, pH of the effluent is lowered to optimize the precipitation and oxidation stages, then the pH is restored to an alkaline value to regenerate the cyanide solution. Recoveries of over 96% copper and cyanide were obtained in laboratory tests using industrial solutions. With this procedure, cyanocidic gold/copper ores could be treated economically by heap leaching, since the cyanide consumed by the copper impurities could be regenerated at very low cost (most of the cost of cyanide regeneration would be compensated by the copper credits). In order to achieve this, a bleed of the barren solution could be treated by the proposed method, and then recycled to the leaching step. The volume of solution to be treated should be large enough to keep a cyanide/copper molar ratio higher than 3:1. This would assure that copper does not cause undue interferences in the leaching or the carbon adsorption stages. If the thiocyanate content is low, the oxidation stage can be avoided, and the solution can be recycled to the leaching stage directly after the solid-liquid separation stage.

Jan 1, 1997