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By C. A. Summers

The U.S. Bureau of Mines applied process mineralogy to investigate a lead and zinc deposit located near Story Creek, Colville Mining District, Alaska. Mineralization is hosted by the Kayak Shale in veins, veinlets, and breccia. Sphalerite and galena, with supergene anglesite, occur in symmetrical quartz veins and between breccia clasts. The study sample analyzed 2.11 percent Pb, 5.37 percent Zn, 0.014 percent Cu, and 1.2 oz/ton Ag. Analyses of crushed products showed that 33 percent of the sample weight, and more than 50 percent of the Pb and 65 percent of the Zn, are in the minus 850 um fraction. Liberated grains of sphalerite and galena were found in crushed products finer than 300 um. Beneficiation tests indicate the material is amenable to processing by traditional gravity and flotation techniques.

Jan 1, 1991

By Kenichi Moriyama

The Sumitomo Toyo Copper Smelter started operation in 1971, as an epoch-making smelter at which 9996 of the charged sulfur could be captured. In the past ten years, the amount of treated copper concentrate has-been increased from 1,300 l/ day to 1;900 t/ day by the extremely effective application of the oxygen enrichment for a flash smelting furnace and converter furnace. The sulfuric acid production capacity at the acid plant has also been increased up to 1,900 t/ day corresponding to the increase of copper concentrate smelting rate without installing an additional acid plant. Furthermore the sulfur fixed ratio is kept at a high level, more than 99.996. despite the SO. strength fluctuation of the wide range of 596 to 1396 which is attributed to the combination of flash smelting furnace and, P.S. converter. These -operational results at the? acid plant were achieved by several modifications. and two examples are briefly described as follows; (1) Converter modifications in order to cope with the high SO, strength -gas (2) Introduction of the anodically protection shell and tube type acid coolers In addition. the quality of sulfuric acid has been positively increased. For example, the concentration of iron and arsenic, which were the main Impurities in ?sulfuric acid, was successfully .reduced.

Jan 1, 1995

By G. A. Brooks, R. W. Matusewicz, Nazmul Huda

"A computational fluid dynamic (CFD) model of a conventional tuyere blown slag fuming furnace has been developed in Eulerian multiphase flow approach. The model was developed by coupling CFD with kinetics equations developed by Richards et al. for zinc fuming furnaces. The model integrates submerged coal combustion and chemical reactions with the heat, mass and momentum interfacial interaction between the phases in the system. The model predicts the velocity, temperature field of the molten slag bath, generated turbulence, vortex and coal utilization behavior from the slag bath. The highest coal penetration distance was found to be at l/L = 0.2, where / is the distance from the tuyere tip and L is the total length (2.44m) of the modeled furnace. The model also predicts that 10% of the injected coal bypasses the tuyere gas stream un-combusted and carried to the free surface, which contributes to ZnO reduction near the free surface.IntroductionSignificant metal values are often discarded with slag in the form of metal oxide produced by the smelting industries. These metal values can be recovered from the slag by proper treatment. Zinc slag fuming is such a process, where a reductant source (usually coal) is injected in the molten slag bath to reduce dissolved ZnO from the bath into metallic zinc vapour. The process can be carried out in either conventional tuyere blown furnace or top submerged lance (TSL) smelting furnace. Conventional slag fuming processes are carried out on batch basis in a rectangular tuyere blown furnace. Approximately 50 tonnes of charge (either molten slag or solid slag) is fed to the furnace at the beginning of each fuming cycle. Two opposing sets of submerged tuyeres are employed to inject air and pulverized coal into the molten slag bath to carry out combustion and reducing reactions."

Jan 1, 2012

By D. Zhao

High temperature deformation characteristics of austenitic 304L, and martensitic 410 and 416 stainless steels were investigated by conducting compression tests over a wide range of temperatures and strain rates (950-1250 C and 0.001-20 s-1). Flow behaviors of the three alloys were characterized. Optical metallography was performed to determine the as-deformed microstructures. The flow behaviors were related to microstructural phenomena during high temperature deformation. Dynamic recrystallization was responsible for the flow softening. Both dynamic and static recrystallization were operating, and the recrystallized grain size increased with increasing temperature and decreasing strain rate. Strain rate sensitivity was found to increased with increasing temperature and decreasing strain rate. No defects such as cracking and shear bands were observed at those test conditions. Preferred processing conditions can be selected by considering as-deformed microstructure and strain rate sensitivity.

Jan 1, 1994

Roasting treatments were investigated to remove arsenic from a nickel concentrate. The main arsenic minerals and their proportions in the concentrate were niccolite (NiAs)-50%, gersdorffite (NiAsS)-40% and arsenopyrite (FeAsS)-10%. Thermal analysis showed that gersdorffite and niccolite began decomposing when heated above 640 and 830°C respectively. The maximum roasting temperature was limited by agglomeration of the concentrate above 900°C. A small vibrated bed reactor was used to investigate the effects of temperature and gas flowrate, the latter to sweep away evolved arsenic. At 900°C, 24% of the original arsenic remained in the concentrate. SEM examination of the calcine showed that the residual arsenic was associated with the calcium in the gangue minerals. Additions of sulphur (either as elemental sulphur or from pyrite) facilitated the decomposition of the arsenic-containing minerals, decreased reaction with calcium in the gangue and produced calcine containing 4-6% of the original arsenic. Treatment with chlorine during roasting was effective in reducing the arsenic content to similar levels. Generation of chlorine in the bed by sulphation of CaC12 also produced encouraging results. Arsenic, Nickel Concentrate, Roasting, Chlorine, Sulphur, Gersdorffite, Niccolite

Jan 1, 2005

By Tsembel Darjaa, Toru H. Okabe, Yoshiaki Umetsu

"With the purpose of extracting molybdenum and rhenium from molybdenum sulfide concentrate; .electrolysis of slurry, solvent extraction and ·precipitation operations were carried out. The slurry of the concentrate in 10 mass% NaCl aqueous solution was continuously circulated into a vertical type 10 stack bipolar electrolytic cell consisted of carbon electrode disks, and, the concentrate was dissolved by electrochemically generated oxidizers at 313 K. The current efficiency for molybdenum dissolution was 73 % after 10 hours electrolysis, and the leach solution was found to contain 0.2 mol/l of molybdenum and 4x10-5 mol/l rhenium. This result shows the dissolution is more than 8 times faster than that obtained using a single cell. The leach solution from electro-oxidation was then subjected to a solvent extraction for rhenium which is then recovered as sulfides. After rhenium separation, molybdenum was precipitated by pH adjustment, and solid MoO3 (H2O)x crystals, free of sulfate, were obtained from the leach solution containing large amount of so42-.IntroductionIn the past, electro-oxidation using a bipolar cell was applied for leaching molybdenite concentrate and ores at U. S. Bureau of Mines [1-3] and also at Bhabha Atomic Research Center [4]. In· these previous studies, Moebious type bipolar cells, in which electrode plates were vertically inserted into the free surface of the electrolyte, were employed and its effectiveness was demonstrated.Recently, the present authors studied on leaching of Mongolian molybdenite concentrate, and the effectiveness of leaching of the concentrate by electro-oxidation in alkaline and acidic NaCl solutions was demonstrated [5]. The leaching reaction during the process can be explained by the 18 electron transfer reaction of molybdenite (MoS2) with the oxidizers produced in-situ by electrolysis of NaCl aqueous solution. Further investigation was made with respect to various process parameters for electrooxidation such as pH, pulp density and temperature of the slurry, and the dissolution behavior of minor elements such as iron, copper, and rhenium [6]. Solubilities of molybdenum in the leach solution at, various temperatures and pH .was · studied, and feasibility of molybdenum recovery by hydrolysis of molybdenum species was demonstrated. By adjusting the pH of the leach solution, molybdenum oxides or hydroxides free from sulfate were successfully obtained from the solution containing large amount of sulfate ion. Rhenium present in the concentrate at a level of several tens of mass ppm was found to dissolve into the leach solution, and co-precipitate with the molybdenum compounds. Rhenium in the leach solutions, therefore, has to be recovered before precipitation of molybdenum, otherwise it is lost in the precipitate."

Jan 1, 2000

By Yongfeng Jia

Lime neutralization and co-precipitation of As(V) with iron(III) is the most widely used and most economical method for the removal of arsenic from mineral processing solutions and effluents. However, the arsenate fixation mechanism in the coprecipitated solids is not fully understood. It was suggested by various researchers to be a ferrihydrite precipitation process during which arsenate is removed by adsorption on ferrihydrite. In this study, XRD, EXAFS and microprobe were employed for the speciation of arsenic and the effect of various factors on arsenic solubility was also investigated. The results indicate that coprecipitation is a more complicated process than simple adsorption resulting in lower arsenic solubilities. Poorly crystalline ferric arsenate was observed for the first time to be present in the coprecipitated solids. In case of lime as base, there exists some type of Ca-Fe-arsenate association. A general term ?coprecipitate? is suggested to describe the precipitated arsenic-bearing solids instead of those widely used specific terms ?arsenical ferrihydrite? or ?arsenic-bearing ferrihydrite?.

Jan 1, 2005

By Tam Tran

Most new gold mines in Australia are required to treat tailings to reduce cyanide levels before they are discharged to storage ponds to meet the stringent environmental management guidelines set by government statutory authorities. In New South Wales, the average level of weak acid dissociable cyanide allowed in open tailings dams is now limited to 20 ppm, As stringent environmental guidelines continue to be enforced in Australia as well as in other parts of the world, the management of cyanide wastes has emerged as a key issue for the gold mining industry. This paper provides a critical evaluation on technical aspects related to the use of ion exchange resins and affiliated processes in cyanide management and tailing treatment. Particular attention is given to the use of ion exchange resin technologies in the processing of mixed gold-copper ores, where cyanide usage is the highest. Information presented in this paper is heavily drawn from the experience on the application of ion exchange resin processes at May Day mine, in New South Wales, Australia and associated pilot plant test programs. Issues related to the management of thiocyanate, especially with respect to the recovery of cyanide from thiocyanate are also discussed.

Jan 1, 1998

By M. A. Verhaege, Stanislav S. Naboychenko, Vladimir I. Skorohodov, Olga Yu. Goriaeva

A new technology for the processing of secondary raw materials containing precious and Platinum Group Metals (PGMs) has been designed at the Laboratory of Heavy Metals at USTU. The technology is based on hydrometallurgical processing of the materials followed by sorption separation of the PGMs from the leach solution. Different versions of the technology can be realised depending on the type of processed material but the fundamentals are similar. Some parts of the technology can easily be integrated into existing flowsheets. The stage of sorption removal of platinum and palladium was tested on the site of the Joint Stock Company (JSC) "Uralelectromed" on pilot plant scale. A process was developed to obtain fine gold, silver, platinum and palladium. No waste is produced because all solutions are recycled. There are no gas emissions. Selective sorption of platinum and palladium from silver-containing solutions is possible even if the concentration ratio of silver to PGMs is one thousand to one. Sorption behaviour of the PGMs as a function of the bulk metal concentrations was studied. Recovery of 90-95% of platinum and 70-80% of palladium is possible.

Jan 1, 2000

By Henk Dijkman

Existing metallurgical operations can be optimized by implementing engineered bioreactor systems in the process configuration. For instance, metals can be recovered safely and economically from process and waste streams using biogenic sulfide. This paper focuses upon possible applications of high rate biotechnology in metallurgy and mining. Metal recovery from liquid streams (process bleed streams, leach water, waste streams), integrated removal of (fugitive) SO2, treatment of acid blowdown and other options will be described. Practical examples are given. A large-scale application of biotechnology at Pasminco's Budel Zink refinery is described. Here, acid plant blowdown is treated together with electrolyte magnesium bleed. Concentrated zinc sulfide (10 tpd) is produced biologically, eliminating the production of gypsum and closing the refinery sulfur cycle. Recovery of metals from solution at Kovehute Pribram in the Czech Republic is the second practical example. In this case, H2S is produced on-site for metal removal from an alkaline slag dump leachate. At the Caribou mine in New Brunswick, Canada, acid mine drainage is treated with biogenic sulfide. This plant is situated upstream of a lime neutralization plant. As a result, reusable ZnS is produced, the volume of the sludge (gypsum and hydroxides) from the lime plant is reduced by 50%, the sludge is free of heavy metals and the final effluent concentration of heavy metals is in the ppb range. In the second phase of this project, metals will be recovered from stored tailings, resulting in significant revenues.

Jan 1, 2002

By Hiroshi Maeta, Takao Kino, Hirota Osono, Kikuo Matsusaka, Azusa Shida

"Prior to refine an ultra-high purity material, it is necessary to estimate the surface cleanness of a base material. In this report, the cleaning method of a surface for refining of an ultra-high purity aluminum has been studied. Aluminum sample of 99.999% purity was p9lished mechanically and electrically and washed with various solvents. Surface composition and a depth profile near the surface were examined. Based on the experimental results, surface structure of an ultra-high purity aluminum was discussed.IntroductionRecently, studies on high purity metals have become of interest, and we have reported .various properties of an ultra- high purity aluminum1- 2)In order to purify a metal, it is necessary to clean a work area and to make vacuous as possible. So, we are studying the purification of aluminum by an ultra-high vacuum refining system and a zone refining system set up in a clean room with cleanliness class of 1000.Now, in order to avoid a mixing of impurities into the refining metal, a surface cleanness of a base metal is very important factor. The purity of the base aluminum used our work is 99.999%, but the surface state of the base aluminum is not clear. Aims in this work .are to investigate the surface state and to obtain the cleaning method of the base aluminum.Experimental procedureSamples used in this work were 99.9% (3N), 99.99%(4N) and 99.999% (SN) aluminum produced industrially and 99.9999% (6N) aluminum purified by us."

Jan 1, 2000

By Kongfang Feng, Fei Ruan, Zhiyu Liu, Jieyu Zhang, Xin Hong, Shaobo Zheng, Jixu Wang, Jinyin Xie

"Mixing phenomena in smelting reduction furnaces with asymmetric side blowing were studied in this paper. Mixing phenomena in a smelting reduction furnace was represented as the mixing time by electrical conductivity measurements. Experiments were performed in a water model under various operating conditions to determine the effects of horizontal angle, radial angle, inserting depth, flow rate of the upper and the lower side nozzle, the included angle between the upper side nozzle, and the lower side nozzle on mixing phenomena in smelting reduction furnace. Comparing mixing time and mixing phenomena, the optimal conditions were obtained for an asymmetric side blowing process.IntroductionIn recent years, many countries have put forward development strategies for saving resources and environment-friendly, the development of various industries had to face its influence for resources and the environment .As we all know, the steel industry is a pillar industry of the global economy, in which ironmaking is an important part of it. In the modern ironmaking process, the blast furnace is regarded as the most important equipment and tool. But with the increasing scarcity of metallurgical coke resources and considering the environmental impact of such processes, metallurgists proposed the concept of smelting reduction. Smelting reduction processes to produce hot metal from iron ore and coal were being developed as a possible alternative to the traditional blast furnace. At present, a number of smelting reduction processes had been developed all over the world, they were mainly included COREX, HISMELT, FINEX, DIOS, etc."

Jan 1, 2012

By Ana E. Bohé

Chlorination process is an attractive and efficient method for the metal recovery from complex materials. The process that is performed at high temperatures, allows converting all the metals contained in the material into their respective chlorides which are further separated in a selective way. Aluminum metal matrix composites with discontinuous ceramic reinforcement of short fibers, whiskers, and particles have been gaining wide acceptance as an important engineering material in automotive, aerospace, and electronic industries. Since the ceramic reinforcements are still quite expensive and contribute significantly to the final cost of composite materials, the feasibility of their recovery from the manufacturing scrap and further recycling is a factor of increasing economic interest. In the present work, a theoretical and experimental study is given for the recovery of Al2O3 and SiC particles and short fibers from three commercially available aluminum-matrix composites produced by molten metal mixing route. The process consists of transforming the constituents of metallic alloy matrices in gaseous chlorides that are dragged by the gaseous stream, while the ceramic reinforcement remains as a solid residue in the reaction bed.

Jan 1, 2006

By P. M. Tait, N. A. Warner

"The refining of thermally produced zinc is currently performed using a highly energy intensive and arguably outmoded process, distilling zinc under atmospheric pressure at temperatures of 900°C to 1000°C. In order to produce refined zinc more profitably and with less environmental impact a new process has been proposed. This process is generic to all thermal zinc smelters and aims to match the purity of zinc from current technology (99.99%-99.999%). The objective is to perform all separation processes using only the heat contained in the smelter product stream. Vacuum conditions are extensively used lowering operating temperatures and enhancing separations. As part of this process zinc is distilled under vacuum from high concentration obtaining a refined product with respect to lead content. A mathematical model has been developed from fundamental principles to describe this stage, the distillation of zinc across a temperature difference between two circulating zinc streams in open channels under vacuum. Heat for the distillation is supplied by the liquid phase, necessitating a large recycle to maintain high operating temperatures, and keep zinc concentrations high. In determining this model and the operating conditions required, relationships between temperature, pressure and composition are considered.IntroductionCurrent technology (the Refluxer) used to refine thermally produced zinc is both expensive and highly inefficient. At present 80% of the worlds zinc is produced via the electrolytic process reflecting the poorer performance of the thermal route, the zinc-lead blast furnace and Refluxer. A new process is proposed to overcome the drawbacks of current processes. These drawbacks are both economic and environmental. The Refluxer requires 6.6 GJ of energy to refine one tonne of SHG zinc 1, other disadvantages of the Refluxer stem from high operating temperature, around 900°C, and liquid zinc being highly corrosive towards ferrous metals. The Refluxer is constructed from silicon carbide sections which join together by means of flat machined faces making liquid metal leaks inevitable. The Refluxer columns are very sensitive to thermal cycling, thermal shock and mechanical vibration, which easily lead to structural failure. To avoid thermal stress on the columns the feed rate and feed temperature need to be kept constant which means the feed baths to the columns must be able to accept solid and liquid feed as the Refluxer has to operate when the smelter is shut down."

Jan 1, 1998

By V. Gornitz

The Orphan mine uranium deposit, Grand Canyon, Arizona, is a cylindrical collapse-breccia pipe, 70 to 150 m in diameter and 500 m deep. Pipe fill was derived from the Supai Group, Hermit Shale, and the Coconino Sandstone. Uraninite, with Cu, Fe, Pb, Zn and Ni-Co sulfides is concentrated along the pipe boundary ring fracture (annular ring) and within the central porous sandstone matrix. There is some suggestion of vertical and lateral zonation of U with respect to Cu. Reduction of ferric oxide cement from red beds and recrystallization of carbonate, with formation of kaolinite preceded major ore deposition. Early Ni sulfides, arsenides and disseminated pyrite were followed by massive pyrite, and botryoidal uraninite, which in turn was suc-ceeded by Cu, Zn and Pb sulfides. A second, later generation of uraninite was closely intergrown with hematite and base metal sulfides. Oxidation of primary minerals may have been associated with Tertiary regional uplift. The paragenetic sequence closely resembles that of other Grand Canyon breccia pipes, except for the extensive development of the distinctive uraninite-hematite association. Mineral assemblages at the Orphan suggest possible mixing of the metal-bearing solution(s) with a more oxidized fluid, containing soluble uranyl complexes. Homogenization temperatures of fluid inclusions in calcite (GO-lIO?C) at the Orphan fall within the lower end of the range observed at other breccia pipes.

Jan 1, 1987

By L. -C. Lai

A series of LaNi5-based alloys, with partial substitution of Ni by Al, Co, or V, were prepared The gas-phase hydrogen absorption kinetics and pressure-composition-temperature curves were measured The activation became easier and the plateau pressure was reduced as Ni was partially substituted by Al Substitution with Co reduced the hysteresis as well as the hydrogenation capacity, but enhanced the absorption kinetics Vanadium is often considered as a hydride former, but when it was added to LaNi5, it was found that V occupied the B site. A small amount of substitution of Ni with V led to a higher hydrogenation capacity and faster activation rate Based on these properties, an alloys LaNia 25CO0.5V0.1Al0. 15 was fabricated as an electrode. Compared with LaNi4.85Al0.15, it was found to exhibit faster activation, higher discharge capacity, and longer cycle life. The variation of the hydrogenation characteristics is discussed based on the physical and chemical properties of the alloying elements.

Jan 1, 1997

By M. Benavides

An experimental study is carried out With sulphide ore obtained from Bailadores-Venezuela as well as artificial concentrates. Concentrates of the minerals from froth flotation are fused to study equilibrium between the matte and slag phases. The first part of the work deals with obtaining a concentrate by flotation experiments with collective and selective reagents, with acceptable metallurgical recovery, studying the influence of different parameters in the process. The second part gives the results of the experimental investigation. on thermodynamic equilibrium in smelting which facilitates knowledge of the elemental processes, giving bases for the possibilities of reducing the loss of metal in slag.

Jan 1, 1983

By Miller, Geoffrey W.

The Ravensthorpe Nickel Project (RNP) flow sheet provides for the upgrading of Ni & Co in a size based beneficiation process prior to leaching. The upgrading effect on Ni and Co makes the beneticiation process a powerful addition to the flow sheet. Multiple linear regression (MLR) analysis based on 351 individual batch tests, 52 large scale composites and 21 pilot plant tests has been used to successfully model beneticiation product grade, recovery, upgrade and mass recovery. Statistically, 83% and 96% of the variability in nickel recovery and nickel product grade respectively was accounted for. Algorithms have been developed for Ni, Co, Mg, Al and Fe, which represents the revenue and operating cost factors for the process. The algorithms have subsequently been applied to the resource model, allowing prediction of elemental grade and recovery for the development of the mine schedules based upon these parameters and not head grades alone. This has allowed optimisation of the mine schedule for improvement to the project's NPV. Keywords: Ni Laterite, Prediction, Multiple Linear Regression, Beneficiation

Jan 1, 2004

By Matti Leiponen

During the last few years Outokumpu Oy has put forth efforts to develop the material handling in the Pori tankhouse. The final objective is full automation including cell loading. The fully automated material handling provides high quality anodes and cathodes as well as equipment and handling systems capable of very accurate cell loading. The new machines and equipment for accurate cell loading and the methods to produce high quality anodes and starting sheets are described in this paper.

Jan 1, 1987

By Gökhan Demirci, Ishak Karakaya

"Energy consumption in electrolytic magnesium cells is well above the theoretical values. However, the mechanism of the back reaction, the main cause of the current losses, has not been fully understood yet. Magnesium dissolution into the melt is generally regarded as the limiting step for the back reaction. Accordingly, present cell designs are made with priority given to fast magnesium removal from the cell. However, experimental data and modeling results in this study indicate that the limiting step for the back reaction between magnesium and chlorine is the chlorine dissolution into the electrolyte when physical contact between electrolysis products was eliminated. Effects of chlorine bubbles on the current efficiency and the cell potential were investigated. Results show that the extent of back reaction is proportional to the chlorine surface area in contact with the electrolyte per unit time, and the cell potential increases with the amount of chlorine bubbles inside the electrolyte.1. IntroductionMagnesium is less dense than the electrolyte in electrolytic magnesium production. Therefore, both electrode products, magnesium and the chlorine gas, move upwards inside the molten salt electrolyte during electrolysis. The upward motion of electrode products promotes the most important current loss mechanism, the ""back reaction"". Separation of the highly reactive species; liquid magnesium and the chlorine gas in the cell, to prevent their recombination is one of the main concerns in cell designs. The employment of large inter-electrode distance or a separation wall in early cell designs; decreased the extent of recombination, but increased the cell voltage and energy consumption. Recent cell designs, on the other hand, make use of lifting action of chlorine gas [1,2] by utilizing smaller inter electrode distances to separate the magnesium and chlorine gas faster. Although energy consumption was reduced, but more physical interactions between electrode products make electrolytic magnesium cells still highly inefficient when energy consumptions are compared with the theoretical values. Today, the most efficient industrial cells utilizing high purity feed operate at 10.5-11.5 kWh kg-1 Mg [3-5] and it is considerably higher than the theoretically attainable value of about 6.0 kWh kg-1 [6]. This large gap demonstrates the necessity of more research on physics and chemistry of magnesium cell electrolysis."

Jan 1, 2011