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By Maija-Leena Metsiirinta

The behaviour of copper and lead in zinc roasting has been studied in laboratory scale roasting experiments as well as in large scale trials in an industrial fluidized bed, with the background of Outokumpu Kokkola Zinc smelter. Commercial zinc concentrates with a low and high ( up to 3 wt % ) copper content and with a low and high ( up to 3.5 wt % ) lead content were oxidized to various desulphurisation degrees in the fluidized bed to establish the reaction mechanisms and the paths for copper and lead oxidation in the roasting. Influences of the agglomerate size on the roasting kinetics and micro- scale phenomena were also determined. Optical microscopy was used on the calcine and the findings are presented as numerous photographs of the polished cross sections. SEM/EDS techniques were used for analyzing chemically the phases and particle morphologies present in the samples. Computational thermodynamics was applied to the complex phase equilibria occurring during roasting in the agglomerating calcine. A complete oxidation of zinc concentrate with a high copper and lead content requires a careful control of oxygen coefficient in the roasting in order to maintain a stable operation of the fluidized bed. Oxygen coefficient thus seems to have a significant impact on the hydrodynamic stability of the fluidized bed with fine raw materials.

Jan 1, 2003

By Vitalij K. Pecharsky

Recent interest in complex aluminohydrides as potential near ambient, ultra-high capacity, solid-state sources of hydrogen (e.g., for fuel cells) calls for an improved understanding of the processes occumng in these hydrogen-containing materials during both the release and the uptake of hydrogen. Lithium aluminohydride (LiAlH4) rapidly releases 5.2 wt. % of hydrogen (i.e., nearly four times more than LaNis-based alloys) during mechanochemical processing at ambient conditions in the presence of transition metal catalysts. Using a variety of solid-state characterization techniques, the mechanochemical transformations in a series of LiAlH4 - metal catalyst systems have been examined and preliminary insights on the mechanism have been obtained. Surprisingly, the well-known hydrogenating-dehydrogenating catalysts, such as Ni and Pt, are practically inactive in the mechanochemical dehydrogenation of LiAlH4.

Jan 1, 2002

By William Petruk

"Heap leaching is a developing technology for recovering native gold and electrum from low grade deposits and from tailings. Recovery depends on permeability of the rock and ore minerals, size of particles in the heap, and ore textures that influence exposure of gold grains to cyanide solutions. These ore textures include gold-bearing veinlets and microveinlets along fractures and microfractures, gold grains in porous minerals and rocks, and gold graim. occurring along grain boundaries of other minerals. On the other hand, small encapsulated gold grains arc not likely tote exposed even with fine grinding. Most gold grains that occtir along grain boundaries of other minerals can be exposed with a moderate to fine grind, but the. resulting material will be too fine-grained to be permeable in a heap. Such finely ground ore, as well as tailings from flotation operations, can however be prepared for heap leaching by agglomerating the material into porous briquettes. Oxidized and weathered zones above. orebodies are frequently amenable to heap leaching because the gold-bearing pyrite and arsenopyrite have become permeable by conversion to porous iron oxides and ferric oxyhydroxides. A review of ore textures in several types of gold deposits suggests that:(1) the gold in vein deposits would gtmerally be amenable to heap leaching of both oxidized ores and fresh non-refractory ores,(2) the gold in Carlin-type deposits and in volcanogenic massive sulfide deposits is not likely to be amenable to heap leaching fresh ore, but oxidized zones above the orebodies are generally amenable to cyanidation, and(3) the gold in porphyry copper-gold deposits and in Witwatersrand-type deposits would likely be amenable to heap leaching of both fresh and oxidized ores but with high cyanide consumption."

Jan 1, 1997

By Julian Johnson

The effect of changing redox potential (Eh) on the high temperature sulfuric acid leaching of Western Australian dry-land nickel laterite ores has been investigated. In particular, the effect of added reductant (sulfur, sodium sulfite) or oxidant (potassium chromate) on Eh, leaching kinetics, metal extractions, acid utilisation and leach residues has been determined. Three separate ores - nonhonitic ores from Murrin Mumn and Bulong, and limonitic ore from Cawse - were used in the study. Sulfur addition varied from 0 to 7.8 kg/t ore. The total acid loading was kept constant at 430 kg/t (Murrin Murrin ore) and 330 kglt (Cawse ore), whilst acid loading for Bulong ore was varied according to the presence of either reductant or oxidant. The extraction of both nickel and cobalt from limonite ore remained unaffected when small amounts of sulfkr were added, but addition of even small amounts of reductant to nontronitic ore decreased both nickel and cobalt extraction. Increased manganese and iron dissolution were observed as the levels of reductant increased, accompanied by a decrease in chromium dissolution. As the iron dissolution increased, free acidity decreased, with a subsequent decrease in nickel extraction. In contrast, addition of oxidant lead to reduced iron and manganese dissolution and slower extraction of nickel and cobalt. Keywords: Limonite, Nonhonite, Eh, Nickel Extraction

Jan 1, 2004

By Alan P. Prosser

The heap leaching of gold ores by cyanide solution was simulated by computer program. The program uses only basic m1neralogical and chemical data; no process testwork data, such as column leach results, are required. The simulation can be used to forecast the consequences of various changes before and during the testwork, design and operation stages of a particular ore-process system. Also, data about the state of different parts of the heap at various times help create an understanding of what is happening. The effects of mineralogical and process variables on the overall performance of the ?heap leaching process are illustrated. Comparisons are made to column leach tests and production data. The potential of the simulation strategy is discussed for making better use of operating experience and for suggesting specific research objectives. It has become clear that more attention should be paid to other metals that dissolve in the solution.

Jan 1, 1989

By F. T. Principe

In order for solvent extraction technology to be acceptable as an alternative to existing jarosite precipitation circuits, iron must be cleanly separated and concentrated into a feed stream suitable for downstream iron by-product recovery which meets market specifications and economics. Ten years ago, significant progress was made in a joint McCill University/CANMET project that involved iron extraction with mono(2-ethylhexyl) phosphoric acid and stripping with HCI. Our iron solvent extraction research was reactivated recently by looking into this and other members of the organophosphorus extractant family. Our research focuses on the mechanisms and control of anion (SO42/CI) transport across the various phases (Rafinate/Organic/Strip), iron build-up in the strip liquor, and extractant stability/selectivity. In this paper, these issues along with some other performance parameters are discussed using data derived from an extensive search of the two commercially available and best known members of the organophosphorus family, di-2-ethylhexyl (D2EHPA) and mono-2-ethylhexyl (M2EHPA) phosphoric acid. Further in this discussion, the challenges lying ahead in the development of sx systems that overcome the deficiencies of D2EHPA and M2EHPA are outlined. In particular, the necessity of considering secondary separation technologies that can further concentrate iron values in the chloride strip liquor to levels economically attractive for the production of saleable iron oxide by pyrohydrolysis is addressed.

Jan 1, 1998

By Paul C. Turner

In 1998, approximately 57,000 tons of titanium metal was consumed in the form of mill products (1). Only about 5% of the 4 million tons of titanium minerals consumed each year is used to produce titanium metal, with the remainder primarily used to produce titanium dioxide pigment. Titanium metal production is primarily based on the direct chlorination of rutile to produce titanium tetrachloride, which is then reduced to metal using the Kroll magnesium reduction process. The use of titanium is tied to its high strength-to-weight ratio and corrosion resistance. Aerospace is the largest application for titanium, and titanium cost has prevented its use in non-aerospace applications including the automotive and heavy vehicle industries.

Jan 1, 2001

By Corby G. Anderson

In the majority of industrial operations recovering gold and silver by hydrometallurgical means, cyanide is the predominant lixiviant. However, as an economic alternative, production of massive quantifies of silver chloride was done on an industrial scale in the nitrous-sulfuric acid pressure leaching of a refractory precious metals concentrate. This proven process required no cyanide. Thus, because of the large volumes encountered during day-to-day plant operation, the need for efficient and low cost methods of silver chloride precipitation, reduction and materials handling was mandatory. As such, this paper covers the evolution of plant scale precipitation, filtration, washing, reduction and control of silver chloride processing. As well, an in-depth description of the unique leaching system is included.

Jan 1, 1998

By John M. Gaines

Lead concentrates from the Viburnum Trend in Southeast Missouri are sintered, prior to smelting, to bind the particles together, homogenize the feed, and reduce the sulfur content. About 20% of the sinter at some plants is classified as bad sinter. A total of 130 samples of good and bad sinters were studied by reflected light microscopy, scanning electron microscopy-energy dispersive analysis, and electron microprobe analysis to characterize their differences in mineralogy and textures. Good sinters consist predominantly of hardysonite, franklinite, lead silicate matrix, and large crystals of calcium-magnesium-iron silicate, together with minor amounts of lead oxides, lead-zinc silicate, metallic lead, specular hematite, magnetite, lead-iron-bearing zincite, and zinc-lead-magnesium-bearing wustite in a tightly intergrown mesh of crystals and with an evenly distributed porosity of about 15-20%. Bad sinters contain larger amounts of lead silicate matrix, galena, metallic lead, and lead oxides, but smaller amounts of hardysonite, franklinite, and calcium-magnesium-iron silicate, lack intimate intergrowth of the phases, and have lower (0-5%) and unevenly distributed porosity.

Jan 1, 1992

By Katsutoshi Inoue

The sulfuric acid leaching of spent hydrodesulfurization catalysts yields an acidic solution rich in rare metals such as molybdenum, vanadium, cobalt, and nickel in addition to aluminum. For the purpose of separating and recovering the rare metals from the solution, basic solvent extraction characteristics of the metals involved with commercially available acidic organophosphorus reagents like TR-83, PC-88A, PIA-8 and CYANEX 272 as well as aliphatic alpha-hydroxyoxime LIX 63 from sulfuric acid media were investigated in the present work. It was demonstrated that CYANEX 272, PIA-8 and LIX 63 possessed excellent selectivity for molybdenum and vanadium over the other metals at comparatively low pH. Further, recovery of molybdenum and vanadium from the total leach liquor by the selected extractants was discussed. Additionally, in order to achieve preferential separation and recovery of cobalt and nickel over an appreciable amount of aluminum from the raffinate obtained after recovering molybdenum and vanadium, a systematic investigation was conducted on synergistic extraction behaviors of the mixed solvents of LIX 63 in combination with a series of acidic extractants such as CYANEX 272, PIA-8, PC-88A, D2EHPA, TR-83 and OPEHPA for cobalt, nickel and aluminum. It was elucidated that all of these mixtures could exhibit a good synergistic effect for extraction of cobalt and nickel, and that the lower the pKa of the acidic extractants, the greater the synergism. Contrarily, the extraction of aluminum, was considerably suppressed by LIX 63. Effective extraction of cobalt and nickel away from aluminum was accomplished from the sulfate solution at low pH (1.0 - 2.5) with the mixtures. In particular, the mixtures consisting of LIX 63 and CYANEX 272 or PIA-8 appeared to be the most feasible and economic from an industrial point of view, yielding acceptable separation efficiency for cobalt and nickel over aluminum and .a low acidity requirement for stripping cobalt and nickel as well as negligible degradation of LIX 63. Aluminum can be readily recovered as sulfate salt from the raffinate obtained after recovering molybdenum, vanadium, cobalt and nickel as outlined above by conventional crystallization method.

Jan 1, 1995

By Richard D. Hangi

Process mineralogy techniques are uniquely applicable to industrial problems related to the extractive metallurgy of nickel, cobalt, and associated metals. Reflected 1ight microscopy, together with selected determinations by scanning electron microscopy-energy dispersive spectrographic analysis and electron probe microanalysis, provides the means to identify the ore minerals cobalt and nickel ores, determine the abundance of each ore mineral, assess the extent of intergrowth between the cobalt and nickel minerals and other ore and gangue minerals in the ores and concentrates, and predict the potential success of cobalt and nickel extraction by grinding, flotation, pyrometallurgical, and hydrometallurgical treatments. In addition, the extent of locking of the cobalt and nickel minerals with each other and with the gangue minerals can be directly observed by reflected light microscopy. Where ores have been treated by hydrometallurgical procedures, reflected light microscopy can serve to determine the causes of excessive leach consumption. Finally, the partitioning of cobalt and nickel between high-temperature synthetic phases produced by pyrometallurgical treatment can be precisely determined by process mineralogy techniques. Case histories from cobalt and nickel deposits illustrate the value of applied mineralogical techniques in the study of cobalt and nickel ores, beneficiation, hydrometallurgial, and pyrometallurgical products.

Jan 1, 1993

By Bora Derin

The partial pressure ratios of sulfur and oxygen for several copper flash smelting plants were derived using the Reddy-Blander (RB) model and predicted dissolved sulfur contents available in the literature. The sulfide capacities (Cs) of six different flash smelting plants (FeO-CuO0.5- CaO-MgO-FeOwAIOl.s?SiO2) slags were calculated. The log (Po2/PS2) versus matte grades of various industrial processes was determined. The PoiPS2 ratio increased with an increase of copper grade. The calculated partial pressure ratios for the copper flash smelting are in good agreement with those of the experimental data.

Jan 1, 2003

By N. L. Piret

In copper smelting minimization of solid waste generation is receiving increased attention with more stringent environmental legislation and lower disposal space availability. Copper smelting slags generally being valorizable, solid waste generation is attributed to the presence in the feed of unvalorizable impurities which do not report to the slag and to the unavoidable formation of contaminated dilute acid solutions which need neutralization. As to unvalorizable impurities, solid waste minimization is achieved by rejecting them in the highest possible concentration. Dilute acid is mainly due to the formation of SO3 in the smelter gases and to the discard of decopperized bleed-off electrolyte. Its generation is substantially lowered by preventing SO3 formation, by converting the dilute acid to a valorizable product, e.g., by thermal decomposition for H2S04 recovery or by precipitation as chemical gypsum, and by suitable treatment of the bleed-off electrolyte for impurity removal. Industrial processes are evaluated and alternatives are discussed.

Jan 1, 1994

"The theoretical basis of ideal smelting reduction using coal as reductant in a process employing closed loop forced-circulation of hot metal is outlined. In the proposed process, an extensive area of relatively slag-free hot metal is available as a heat sink for absorbing the large amount of heat produced by complete combustion of CO within the furnace. Any slag that is formed within the oxidizing arm of the furnace loop is quickly swept away as a thin floating layer by the bulk movement of hot metal which is induced positively by forced melt circulation and sustained by causing the hot metal and the thin slag layer to overflow into a lower level hearth in which coal is carbonised and dissolved directly into the circulating melt.The proposed process is shown to offer very substantial cost savings over all other direct smelting technology. This is mainly because of its inherently low capital cost and its high energy efficiency. Furthermore, the favourable economics are not reliant on fuel gas credits as all the process gas is combusted within the reactor itself.IntroductionMetallurgical processes in the year 2000 will be compromise solutions to the metallurgical problems that already exist today. It is common knowledge that the mineralogical complexity of certain polymetallic sulphides, for example, precludes high degrees of separation and recovery by conventional mineral beneficiation. steps are therefore being taken towards the development of a truly polymetallic smelting process which will produce simultaneously zinc lead and copper as, metals from complex ores without prior beneficiation(l). In the ferrous extraction industry, the shortcomings of the highly efficient modern blast' furnace are well known and world-wide there is considerable effort being directed towards coal-based ironmaking without the requirement of metallurgical coke as a process intermediate. The current status 0f2Joal-based ironmaking has recently been reviewed by Smith and Corbett. (2)"

Jan 1, 1988

By Robert E. Moore

Synthetic spinel is proving to be a highly effective refractory for specific applications, but its cost limits applications. Formation of spinel from a raw material such as diaspore or bauxite could provide a simple and less costly starting material. In the study reported here, finely ground (-325 mesh) diaspore from Missouri and bauxite from Guyana and China are combined with hard burned MgO (-325 mesh) in appropriate proportions to yield spinel and spinel with secondary phases. A classical sintering and dilatometer study was completed, followed by a microstructure study to determine completeness of the reactions and morphologies of the reaction products.

Jan 1, 1993

By David M. Davidson

Seven intrusive phases have been mapped to date in the 1 b.y. Red Bluff Granite Complex (RBGC) near El Paso, Texas. This batholith is 20 km2 in area and intrudes older supracrustal units as well as a trachyte ignimbrite cap. Two early microgranite sills are spatially separated 6 km from a porphyritic, three-phase "ring complex" by a tin-bearing, biotite granite which cores the RBGC. These units are intruded, in turn, by a riebeckite granite which locally exhibits lineated and pegmatitic textures as well as exceptionally high light REE and Zr values (800 ppm La; 3,000 ppm Zr). The riebeckite granite is exposed in three separate occurrences within the Franklin Mountains. REE values in the RBGC are depleted in HREE and show distinct negative Eu anomalies. Typical La/Yb ratios are 5:1-2:1 with Eu values ranging from 10-80 ppm. The mineral source of the high REE values in the riebeckite granite has not yet been specifically identified, but is believed to be zircon.

Jan 1, 1988

By I. J. Moon

The reduction behavior of hematite by H2 and H2-CO gas mixtures was investigated to elucidate the reduction mechanism at 800-950 °C. The hematite compacts were made by Cold Isostatic Pressing to produce compacts of uniform shape and size. The compacts were sintered at 1000° C for 30min, and showed a contraction of 29%. Mercury pressure porosimeter was used for measuring the total porosity, pore size distribution and pore surface area of the compacts. Reduction was measured by means of weight-loss technique using Cahn Balance (Model : TG-171, capacity log, accuracy f 10pg). Microscopic examination and X-ray analysis were used to correlate the structure of reduced compacts with the mechanisms of reduction. Neither chemical reaction control nor diffusion control is adequate in describing the complete course of reduction of the hematite compact. The reduction of hematite compacts was initially controlled by the chemical reaction at oxidelmetal interface. The reaction goes through a transition phase .where it is undergoing a mixed control, followed by the diffusion control towards the end of the reduction. The values of apparent activation energies for chemical reaction control were 3.48-8.43 kcal/mol depending on the gas compositions.

Jan 1, 1997

By Nolberto Moreno Borjas

This paper outlines the detrimental impact that unblended ore can have on the Ni and Co recoveries and the ore slurry density with real data from the Caron low-pressure ammonia leaching process for the Punta Gorda ore. The author briefly discusses the reason for the decrease in slurry density that the process experiences. The lower metal recoveries in the process that are caused by the unblended ore chemical characteristics and variations are shown with a data base statistical analysis drawn from the results obtained from a pilot test, designed with different ore bypassed from the normal process flow and a similar ore from a similar metallurgical treatment. Keywords: Ni-Co Laterite, Caron Process, Ore Blending, Slurry Density, Metal Recovery

Jan 1, 2004

By Ahmad-Bitar¸ Riyad N., Shadia J. Ikhmayies

"CdSi-xTex thin films with O<x<0.3 were prepared on glass substrates by the annealing of CdS:In thin films produced by the spray pyrolysis technique. The annealing was performed in nitrogen atmosphere at 400 --,C in the presence of Te vapor. The compositions of the films were determined by energy dispersive X-ray detection (EDAX) measurements. Scanning electron microscopy (SEM) was used to investigate the morphology of the films. The transmittance of the films was measured at room temperature in the wavelength range 400-1100 nm and used to produce the absorbance of the films. The first derivative of the absorbance was used to estimate the optical bandgap energy Eg by using the minima in the first derivative. The relation between the bandgap energy and the fraction x was found to fit with a parabola and bandgaps in the range 2.196: =Eg: =2.480 e V were obtained. Urbach tailing in the bandgap was also investigated.IntroductionPolycrystalline thin film CdS/CdTe solar cells are promising candidates for low cost and high efficiency photovoltaic applications. The fabrication process of these cells involves the deposition of a p-type CdTe absorber layer on top of an n-type CdS one which acts as a window. The efficiency of the solar cell is affected by the method of preparation of both layers [1, 2].Interdiffusion processes taking place at the CdS/CdTe interface have largely been realized to significantly affect the device performance [3] by changing the bandgap of both CdTe and CdS which modifies the spectral response of the solar cell [ 4]. The formation of a mixed crystal or a graded layer causes a reduction in the lattice mismatch at the CdS/CdTe junction [5] which is about 9% [ 6] and then a reduction in the number of interfacial states and recombination centers, thus enhancing the solar cell performance [7, 8]. A complete understanding of the properties of the intermixed layers CdSxTe1-x that usually constitutes the interfacial region in the CdS/CdTe solar cell is essential. According to our knowledge, a lot of research was devoted to the investigation of the diffusion of S into CdTe [3-5, 9]; but little on the direct investigation of the diffusion of Te into CdS [6]."

Jan 1, 2011

By Masao Suzuki

Over 5 million cars are scrapped yearly in Japan. After dismantling. scrapped automobiles, they are put into a shredder for differential recovery of ferrous and nonferrous metals. The residue, which is called shredder dust, runs over 1.2 million tons per year. This paper reports a entire sequence of scrapping. cars. in Japan with the following sections: (1) production and scrapped car management, (2) material composition, (3) dismantling, (4) shredder plant, (5) differential recovery of metals including specific gravity separation and newly developed color separation.

Jan 1, 1995