Search Documents

By Jay Black

In the thiosulfate gold leaching process gold is oxidized to Au+ by a copper(II) ammine catalyst and complexed by thiosulfate. Copper(I) complexes, formed during gold oxidation, are reoxidised to Cu(II) complexes by oxygen. In this study, we present the results of a spectrophotometric study of the Cu(I)-NH3-Cl system, which follows on from our recent investigations of speciation in the Cu(I)-Cl and Cu(I)-S2O3-Cl systems and is part of a detailed study of copper(I) speciation associated with the gold leaching process. Formation constants for Cu(I)-NH3-Cl complexes have been determined by modeling the systematic changes in the UV spectra that occur as the concentration of the components is varied. The analysis of the data indicates that three Cu(I)-ammonia complexes are present, i.e., [Cu(NH3)2]+, [Cu(NH3)Cl2]- and [Cu(NH3)2Cl] and their respective log Kf values are 11.13, 8.33 and 10.79 at 25°C. Investigations of the oxidation of Cu(I)-NH3-Cl solutions by oxygen in the atmosphere are also presented. These experiments indicate that the reaction of copper(I) complexes progresses through various intermediates. The presence of many copper(I) and copper(II) complexes in ammonia thiosulfate solutions complicates the task of determining optimal conditions for gold leaching. Further studies are needed to identify the complexes and chemical processes that are important in the Cu(I)-NH3-S2O3-Cl system.

Jan 1, 2003

By F. A. Myzenkov

The technology developed by the authors uses new methods for raw material smelting in electric furnaces, a new autogenous method and apparatus for subliming smelting for antimony trioxide production and separation of antimony and gold. Concentrate produced from the ore deposits in Yakutia has the following average contents of main constituents, % by weight: 29.4 Sb, 1.2 Fe, 13.1 S, 40.0 SiO2, 0.2 As, 2.6 CaO and 45 g/t of Au. Unlike the technology used abroad a new autogenous method makes it possible to increase considerably specific capacity and reduce gold losses with antimony trioxide by several times. At the same time mineralogical composition of trioxide has improved considerably ? it has mainly a cubic base with prevalence of senarmontite. The content of a main component (Sb2O3) in the final product has increased from 99.0% up to 99.6% which means better quality and makes it suitable for use in special technologies. Despite a lower antimony content of the concentrate (29.4% as compared with over 33% as common in the world practice), an antimony recovery of 85.3% into a commercial-grade product has been achieved, as well as a gold recovery of 96.65% into a gold-antimony alloy. The technology for processing gold-antimony materials developed and introduced at the Ryaztsvetmet plant is unique in the world with respect to gold recoveries.

Jan 1, 2006

By S. G. Kumar

The performance of nickel based high temperature materials is directly related to their microstructure, which is controlled by the various processing steps. Nickel base alloys are melted and refined by a variety of processes that include Vacuum Induction Melting (VIM), Vacuum Arc Remelting (VAR) and Electroslag Remelting (ESR). The limitation of VIM is that the metallrefractory reactions lead to the disintegration of the crucible and hence contamination of the product due to macrosegregation. This macrosegregation is alleviated to some extent by controlling the solidification in VAR. The main attraction of ESR is its ability to produce a clean product by controlling the deleterious inclusions. In recent years Electron Beam Melting (EBM) is emerging as an alternative to the VIM, VAR and ESR. This paper reviews the current status of these processes and outlines the scope for further improvements in the processing technology.

Jan 1, 1993

By Jinxing Ji

Several fundamental aspects related to the nickel electrowinning from chloride electrolytes have been addressed in this paper. The thermodynamic study includes the activity coefficient of the hydrogen ion and the speciation diagram of NiCI2 solutions. The activity coefficient of the hydrogen ion (YH+) was calculated and estimated experimentally using a combination glass pH electrode. It was found that YH+ was greater than one in concentrated NiCl2 solutions and increased significantly with increasing NiCl2 concentration. The speciation diagram for NiCl2 solutions was plotted with YH+ and the effect of the ionic strength on the equilibrium constants being taken into account. It was discovered that the predominant nickel species in the acidic region were Ni2+ and NiCI+ in 0.937 M NiCl2 solution at 25°C. Electrodeposition tests demonstrated that higher Nit' concentration in the electrolyte improved the current efficiency of nickel electrodeposition (CENi). The additions of H3BO3, and NH4Cl increased very little the current efficiency of nickel electrodeposition at pH 1.5 and temperature 60°C. However, the addition of sulfate decreased the current efficiency of nickel electrodeposition. Nickel electrodeposition was also found to be a steady-state process since the amount of acid added to the electrolyte at a constant pH increased linearly with time during galvanostatic electrodeposition. The cathode surface pH during nickel electrodeposition was measured using a flat-bottom combination glass pH electrode and a 500-mesh gold gauze as cathode. The cathode surface pH was found always to be higher than the pH in the bulk electrolyte, and if the current density was sufficiently large, it would eventually reach a level causing precipitation of insoluble Ni(OH)2(s) on the cathode surface. Rotating disc electrode studies revealed that the rate of nickel electrodeposition was first order with respect to the activity of nickel ion and zero order with respect to the activities of chloride and hydrogen ions. The rate of hydrogen evolution was found to be first order with respect to the activity of hydrogen ion and to be zero order with respect to the activities of nickel and chloride ions. These findings indicate that nickel deposition and hydrogen evolution proceed independently.

Jan 1, 1995

By J. Murdoch Mackenzie

Two commercial nickel solvent extraction plants treating laterite leach solutions and using LIX®84-I have been operated in Australia. These circuits employ similar extraction chemistry but differ in their leaching and stripping operations. An alternative process involving elements of both can be conceived. The preparation of the feed solution to solvent extraction plays an important role in nickel extraction using LIX®84-I and the special features of manganese removal and cobalt oxidation in the preparation of the PLS are discussed. This paper compares these three circuit configurations with special reference to the extraction and reductive stripping of cobalt, the transfer of zinc and copper to the strip aqueous, ammonia transfer and sulfur transfer. The potential degradation of the organic and remediation using re-oximation is discussed. Keywords: Nickel, Solvent Extraction, Oxime

Jan 1, 2004

By Tom W. Gonzales

Magma Metals Company commissioned a 3000 tpd Outokumpu flash smelting furnace in July 1988. Through September 1992 the PSP has smelted a first campaign record 4.1 million tons of concentrate. This paper discusses modifications made to the original process equipment to overcome design flaws as well as future design changes currently under consideration.

Jan 1, 1993

By A. V. Tarasov

In order to improve the efficiency of hydrometallurgical processing of pyrrhotite concentrates, an integrated set of research has been carried out to investigate the use of sulfur dioxide in individual unit operations and as a leaching agent in the autoclave process as a substitute for oxygen. Treating iron-containing raw materials for production of nonferrous metals are based on the selective reaction of SO2 with iron sulfides and oxides. Iron is dissolved as sulfate, which can be treated in a separate circuit to produce iron-containing pigments or it can be disposed of after neutralization. Nonferrous and precious metals contained in pyrrhotite concentrate do not react with sulfur dioxide and remain virtually completely in the solid phase, i.e. high-grade sulfide concentrate to be treated by hydro- or pyrometallurgical methods. Elemental sulfur formed in the process of leaching is separated from the pulp in granulated form as a result of the use of special surfactants. The process of sulfide concentrate leaching with the use of sulfur dioxide is conducted at temperatures above the sulfur melting point (119°C) with distributed feeding of surfactant and under a partial pressure of sulfur dioxide of up to 0.3 to 0.5 MPa.

Jan 1, 2006

By Jyothi Rajesh Kumar, Jin-Young Lee

21st century is electronic revolution in human lives as well as energy is one of the most significant tasks. Rare earth elements (REEs) occupy a key role in daily life as well as high-tech industrial applications. REEs are one of the energy critical elements; for sustainable growth, it depends on multiple extraction and separation technologies including solvent extraction and ionic liquid-based extraction methods.

Mar 1, 2017

By G. Fazel Jahromi

Enargite, known as one of the major arsenic containing copper minerals, with approximately 19% arsenic, introduces challenges to typical processing options. In most hydrometallurgical processing methods, arsenic remediation occurs after the copper recovery in an autoclave or in a stepwise neutralization process. In this study, the in situ precipitation of scorodite was investigated during an atmospheric leaching process in the presence of carbon-based catalysts and chloride media. The effective parameters of initial ferric addition, temperature, catalyst addition, and oxygen sparging rate were studied and the ferric behaviour was monitored during the scorodite precipitation. The most influential parameters on efficient scorodite precipitation are found to be: initial ferric addition, catalyst to concentrate ratio, and temperature. A 99% scorodite precipitation yield was achieved in this process.

By Lloyd R. Nelson

The feasibility of producing low carbon (LC) ferromanganese (FeMn) by lime-'ore' melt refining of high-grade ferromanganese silicon (FeMnSi) was investigated. Thermodynamic simulations demonstrated significant benefits in moving from a single-stage ladle operation to a two-stage counter-current reladling operation (to emulate the commercial Penin Process). The theoretical alloy silicon content was predicted to be refined from 1.57 to 0.005 per cent silicon, at a 6.2 per cent higher manganese yield (yield predicted to increase from 82.3 to 88.5 per cent). Extension of this counter-current refining principle, to a multistage Counter-Current Reaction Launder (CCRL) was investigated using the Slag Metal And Gas Klnetics (SMAK) software to model the mass transfer kinetics of the multicomponent coupled reaction system. Simulation results suggested that commercial production of 50 000 tonnes per annum (tpa) of refined LC FeMn, containing less than 0.5 per cent silicon, at manganese yields of up to 90 per cent, should be feasible in a compact CCRL (7 m long by 0.7 m wide).

Jan 1, 1994

By Lloyd R. Nelson

Experimental results of three-phase mixing in a physical model of a CCRL for metals refining are presented. The system was operated above a critical lower liquid height-to- width ratio (H,/W) of unity, with adequate bottom gas injection through a line of central, single tuyeres (Q,/A = 7.5 cm/min, or c = 17 W/m3) to sustain levels of interphase transport considered appropriate to an industrial pyrometallurgical refining operation. The application of a thermal tracer technique to elucidate both longitudinal mixing and interphase heat transfer is discussed. The results indicate that even under conditions of significant bottom gas injection, it should be possible to maintain acceptably low longitudinal mixing 1 (D, < 20 cm /s and D,/uL < 0.1) in the CCRL. Interphase mass transfer coefficients of the order of 0.004 cm/s are predicted from the measured interphase heat transfer coefficients (h = 1.3 to 2.0 kW/m2K), using the heat and mass transfer analogy. Adequate rates of rehning are predicted for an industrial CCRL operated under such conditions.

Jan 1, 1995

By Youn-Bae Kang

Evaporation mechanism of S from liquid Fe–C–S alloys at 1873 K was proposed by analyzing available experimental data. It has been known that increasing C content in liquid alloy increases activity coefficient of S ðfSÞ, and it could raise driving force for the evaporation reaction S = S(g). However, experimental data of the evaporation of S in the Fe–C–S alloys could not be accounted for only by considering the increases of fS. In the present study, formation of carbosulfides, CS(g) and CS2(g), was additionally taken into account in order to explain role of C for the accelerated S evaporation. Surface adsorption of S was also taken into account, which retards the evaporation rate of S. An evaporation model equation was formulated. It can be applied to calculate the evaporation rate of S over wider C content (from zero to its saturation to liquid alloy).

By Toshiyuki Hiranami, Keizou Uematsu

"AbstractionSpray freeze drying method was proposed to fabricate ·ceramic: powder granules for high performance ceramics. Obtained granules and their green compact were characterized by SEM and immersion liquid technique. Granules have a spherical shape of about l 00 µ m in diameter, ·and no binder segregation was found with the immersion liquid technique. The granule had sufficient strength for usual handling, while the strength was much lower than that of granules prepared by conventional spray-drying. Unlike the conventional green compacts, granule interfaces were not observed in the green body made with the spray-freeze dried granules.IntroductionIt was well kn9wn that spray-dried granules have binder segregation layer on their surfaces. The trace of this segregation layer survives in the later processes i.e. compaction and firing, and behaves as fracture origins in the sintered bodies. The origin of this problem is solvent drying process from slurry droplet surface in spray drying system. As alternative method we propose a freeze drying process for granule preparation, in which solvent is remove by sublimation though the process, binder segregation can be prevent.In this study, we applied spray freeze drying method to fabricate alumina powder granule, and characterized the obtained granules, and their green compacts and sintered bodies.Experimental procedure.Alumina powders and additives aqueous solution were mixed for 24 hours using a ball mill to· prepare the slurry for spray freeze drying. .The solid loading of the slurry was 30vo1%. Deflocculant (PAA-NH4 type) was added 0.5mass%, and the binder (PVA). 1.0, 2.0 or 3.0mass% of alumina. Obtained slurries were sprayed into liquid nitrogen to make frozen granules. The granules were freeze dried.For the characterization of the granules, immersion, liquid technique observation and single granule strength test were carried out. Strength of individual granule wa5 measured using micro compression testing machine (MCTM, SHIMADZU). Granule strength was calculated from following equation."

Jan 1, 2000

By Alejandro R. Alonso, Lizet Alvarado, Gretchen T. Lapidus

"Cobalt and lithium recovery from spent Li-ion secondary batteries has been increasingly studied in recent years due to the high value of both metals and the low proven reserves for lithium. The standard method employed until now has been the deactivation and crushing of the batteries, followed by leaching of the L1C0O2 spinel from the battery cathode using acid solutions, such as 1.5 to 2.0 M HC1 or H2SO4, in a temperature range from 120°C to 250°C. Once leached, the cobalt is separated by electrodeposition and the lithium by precipitation, using large amounts of carbonate to bring the solution to alkaline pH values. In the present work, citric-thiosulfate solutions are used for the spinel leach at room temperature, favoring a reductive environment at room temperature, which stabilizes the cobalt(II) as cobalt citrate and lithium(I) at pH 5, and recovers approximately 99% of both metals.IntroductionThe Li-ion batteries industry has become an important economic sector in recent years. This type of battery is the most promising system for energy storage, because of its high efficiency in charge-discharge cycles, and its adaptability to many configurations. The demand will probably increase because of the need for mobile energy sources in many types of devices ranging from cell phones, tablets, and laptops to medical, navigation and sophisticated communication equipment and even in automobiles.One major problem related to the use of these batteries is the method employed for recycling the components, mostly cobalt and lithium. Currently, there are two processes employed on an industrial scale: the Sony/Sumimoto process, used since 1990 [1] and the Toxco Process [2]. In the first case, incineration is used to open the batteries and at same time volatilize the organic solvents. After that, the metals (copper, iron and aluminum) are physically separated. The cobalt and lithium are recovered as oxides and then thermally treated in order to recycle them for new batteries manufacture."

Jan 1, 2012

By Somayeh Dashti, Ehsan Vahidi, Fereshteh Rashchi

"The main goals in zinc electrowinning process are decreasing of power consumption and increasing of current efficiency. In this study, the electrowinning of zinc in the presence of different additives in electrolyte were established and Arabic Gum, Gelatine, sodium lauryl sulphate and l-hexyl-3-methylimidazolium hydrogen sulfate in the five levels using Taguchi method was examined. In our previous researches, using orthogonal array, effect of zinc electrowinning parameters such as anode type, current density, sulphuric acid concentration, temperature and zinc ion concentration was studied and the optimized conditions for zinc electrowinning was obtained. Anode composition: Pb-Ag 0.5%-Sn2%-Sb1%, current density (A/m2):500, acid concentration (g/1):165, temperature: 60°C, zinc ion concentration (g/1):70. The main goals in zinc electrowinning process are decreasing of power consumption and increasing of current efficiency. ln this research, the effect of four additives: Gum Arabic, Gelatine, sodium lauryl sulphate and an organic material in the four levels with Taguchi method was examined. ln fact the effect of electrolyte environment and determining the best electrolyte under the optimized conditions is the main aim of this study.IntroductionDue to its low cost and abundance, zinc is widely used in the production of brass, galvanized steel, sacrificed anodes and primary and secondary batteries [I]. Some 80% of world primary zinc production is derived via roast-leach-electrowinning processes [2-4]. In the electrowinning process to both improve final product quality and current efficiency the leachate should be purified [5].To produce zinc a current method is to extractit from acidic zinc sulphate solutions by electrolysis with aluminum cathodes and insoluble lead-based anodes [6]. Though the electrolytic processing of zinc has been performed for many years, there are still current efficiency problems in the cathodic process. These low efficiencies (88- 93%) have a serious impact on the costs associated with the energy consumption. On the other hand, these low efficiencies increase energy consumption and decrease the quality of the surface morphology [7- 8]."

Jan 1, 2012

By Kiyoshi Shibata

A new assessment model has been proposed for metal production and its recycling system. The model consists of three processes, smelting of ore, cascading and upgrading of metal scrap. The present approach is to assess the total energy consumption of the system, by introducing two equations for describing the relationships between critical impurity content and the mass of production or scrap recycling. The total energy consumption in the system can be described as a function of recycling rate and performance of the upgrading process. The present model suggests that the impurity removal efficiency and yield in the upgrading process are essential to reduce the energy consumption in the systems. Especially, such points should be explicitly controlled in recycling of aluminum. The potential capability of this model was also discussed by obtaining several quantitative information required for the process design and the operational control of beneficial recycling of metals.

Jan 1, 1998

By Jingwei Lou, Bing Li, Suzhen Li, Miao Shen, Jianguo Yu

"As a main impurity in primary lithium metal, magnesium is difficult to remove because of the very close relative volatility of the two metals. Electrochemical refining provides a possibility to completely remove Mg in the LiCl-KCl melt prior to Li reduction. MgCl2 reduction processes in LiCl-KCl- MgCl2 melt were investigated by cyclic voltammetry (CV), square wave voltammetry (SWV), chronoamperometry and chronopotentiometry. Then potentiostatic electrolysis was carried out and Mg (II) ion was reduced onto a liquid lead cathode from the LiCl-KCl-MgCl2 melt at the Mg deposition potential. The results show that Mg ( II ) is reduced in one step with two-electron transfer. The reduction potential of Mg ( II ) was well defined. The reduction of Mg( II) results in Li ( I ) underpotential deposition at 0.2V more negative potential than that of Mg( II) reduction. Therefore, in order to prevent Li ( I ) from codeposition with Mg, accurately controlling Mg ( II) reduction potential during the potentiostatic electrolysis was necessary. Ninety three to 99 % removal rate of MgCl2 was achieved after 8-12 h electrolysis when liquid lead was used as cathode.IntroductionMetallic Li has been widely used in batteries, light weight alloys and in fusion reactors because of its excellent properties. Currently, the only way in industry to produce metallic Li is by electrowinning from molten LiCl-KCl at about 450 °C [l-5]. LiCl-KCl is the only electrolyte used, because they presents a lower eutectic temperature and KCl shows a slightly higher theoretical decomposition voltage than LiCl, as shown in Table l .The produced metallic Li generally contains 1-2 wt% impurities including Na, K, Al, Ca, Mg, Si, etc. [6l, which are derived from the corresponding metal chloride compounds or metal oxide compounds existing in the LiCl and/or KCl raw materials. These impurities have to be removed by subsequent vacuum distillation at high temperatures to meet the requirements for metal Li purity in the final products. Separation of Mg from Li by vacuum distillation is difficult because their relative volatilities are very close, as shown in Table 2."

Jan 1, 2011

By Kuniaki Murase, Tetsuji Hirato, Koji Nitta, Yasuhiro Awakura

"The redox behavior of copper species in an ""ammonium imide"" room temperature molten salt was examined in order to clarify the applicability of the salt to the electroplating media of various metals. Trimethyl-n-hexylammonium bis(trifluoromethanesulfonyl)- amide (TMHA-Tf2N), having an electrochemical window of 5.6 V (50 °C), and corresponding copper(II) salt were synthesized. In TMHA-Tf2N media monovalent copper ion, Cu(I), was stable and Cu metal was oxidized, or corroded, in the presence of Cu(II) species. Cathodic electrodeposition of copper metal and anodic dissolution of it were, thus, one-electron reaction and their current efficiencies were almost 100%.IntroductionElectrodeposition of metals from aqueous solutions is an important soft process for manufacturing thin-film materials, surface finishing, and extractive metallurgy. However, the electrodeposition of base metals, such as zinc, chromium, and nickel, from aqueous media inevitably accompanies hydrogen evolution, which lowers the cathodic current efficiency of metal deposition. Since the electrodeposition of aluminium, titanium, and rare earths from aqueous media is actually impossible due to their largely negative redox potentials, nonaqueous media such as aprotic organic solvent and high-temperature molten salts have been employed for the electrodeposition of these metals. Room temperature molten salts (RTMS),1, 2) or room temperature ionic liquids, have also been developed as a promising electrolytes for the electrodeposition of base metals. In this field, chloroaluminates of alkylpyridinium or alkylimidazolium cation have been well investigated for aluminium3) or alluminium-alloy4-5) electroplating media. On the other hand, in order to avoid high hygroscopicity arising from the use of aluminium chloride, attempts have been made to combine the alkylimidazolium cations with fluorine-containing hydrophobic anions, e.g. BF4-,6) CF3SO3- (TfO-),7) (CF3SO2)2N- (Tf2N-),8) and H(HF)n-ions,9) and investigated for electrodeposition media.'°) Recently, a new series of RTMS consisting of tetraalkylammonium cations and Tf2N- anion, sometimes abbreviated as TFSI- anion, was found independently by MacFarlane et al.11, 12) and Matsumoto et al. 13) This type of RTMS, which is commonly named ""ammoniumimide"" salt, has a wide electrochemical window of ca. 5 V and, thus, the RTMS seems to be suited for practical application as a solvent for the electrodeposition process. In this paper, we report the electrochemical behavior of copper species in an ammoniumimide RTMS, trimethyl-n-hexylammonium bis(trifluoromethanesulfonyl)amide (TMHA-Tf2N)."

Jan 1, 2000

By Guo-hua Bai, Hong-jing Yuan, Qian Li, Yong-bin Yang, Tao Jiang, Yan Zhang

"The effect of different additives on the bio-oxidation of arsenic-bearing gold ores has been carried out to solve the problems such as long period of leaching, low oxidation rate. The oxidants strengthen the cathodic reaction and thus improve the leaching rate by the increase of electron transfer. Meanwhile, silver catalysis is also effective for the decomposition of arsenicbearing gold ores due to the strengthening of anode reaction. Co-intensification by oxidants and Ag (I) is more efficient to the bio-leaching and the conversion of As (III) because the anode and cathodic reaction of biological oxidation of sulphide ores were both strengthened. The leaching rate reached to 63.94% after bio-leaching 13days by co-intensification of Ag (I) and Fe(III).IntroductionWith the worsening global ecological problems and the gradual exhaustion of available gold resources, metallurgical technology of refractory gold ore has received more and more attention and become possible due to its good advantage of the economic effects and negative effects to the environment. Refractory gold ore, the direct cyanide leaching rate of that is defined as less than 70%, has become an important raw material in gold extraction industry, in which sulfide bearing gold ore represents a large proportion, pyrite and arsenopyrite (FeAsS) is a common gold-bearing mineral. Since the gold is often encapsulated in the crystal lattice of these minerals, destruction or alteration of the ore to liberate the submicron and/or structurally bond gold is necessary[l]. Even if the ultrafine grinding approach also is generally not sufficient for these refractory ores and additional techniques like roasting, pressure oxidation and bacterial oxidation are used to liberate the gold [2]. In the three pretreatment processes, bacterial oxidation is characterized by its new ideas, low cost, no pollution, simple equipment and operation, high leaching index and become the most promising process which has received much attention over time [3-5]. But its further application is limited by some of its own shortcomings, such as long pretreatment time and the toxic effects of arsenic to bacteria on bio-leaching limit."

Jan 1, 2013

By B. Manana

In heap leaching, the oxidative dissolution of value minerals encapsulated deep within large particles cannot be fully understood on the basis of the bulk solution conditions relative to variable solution conditions near the mineral surface. In the present study, the diffusion of the oxidising species through inner particle pores is simulated in a model apparatus that separates an inert platinum electrode under controlled reduction potential from the bulk solution through narrow pores of varying length. The Cu-NH3 system was selected as a model system and Cu(II)/Cu (I) as the redox couple of interest. Platinum probes inserted in the pores allowed the measurement of varying redox potentials along the pore length. Results showed that in a 10 mm pore, there exists a narrow (0.5 mm) zone near the reduction surface that is depleted of the oxidant and in which its supply is strongly diffusion limited, even if it is in abundance in the bulk solution. The consistent presence of this narrow diffusion region across all experimental results indicated that the transportation of the reduced species away from the reaction surface had the most significant influence on the potential variations within the pore. This was further supported by the minimal effect of varying concentrations of both Cu(II) and dissolved oxygen on the solution potential within this region.