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By NEHA PANDEY, Sunil Kumar Tripathy, ARIJIT BISWAS

Due to the growing market of Electric Vehicles (EVs), there is a rise in battery demand. Nickel can be used as a potential element in cathode for lithium-ion batteries. Nickel can provide thermal stability, high energy density, and conductivity in cathodes of lithium-ion batteries. Low-grade chromite overburden from Sukinda mines contains less than 1 wt% of nickel, which can be utilized to extract nickel. Recovery of nickel from such low-grade overburden can help in value addition. The present study provides deeper insight into the aspects of extraction of nickel product derivatives (NPDs), mainly mixed hydroxide product (MHP), through the hydrometallurgical route. The current study focuses on extracting nickel and cobalt from a mine overburden with 0.41% nickel with major silicate gangue. MHP can be processed further for nickel sulfate production, a potential raw material for cathodes in EV batteries. The study discusses a process flowsheet to recover nickel and cobalt from overburden through atmospheric leaching with pre-calcination. It is observed that the step of pre-calcination improves the effective recovery of nickel from overburden. The leaching reactions are effective at a temperature of 90 °C and 1 atmosphere pressure. The orthophosphoric acid was used as a lixiviant for the selective recovery of nickel and cobalt from overburden. Recovery of nickel and cobalt obtained at optimum conditions is 84.0% and 39.9%, respectively. The iron impurity in the leachate solution is 1.27%. It is concluded that the process with pre-calcination along with acid leaching is effective for extracting nickel and cobalt at atmospheric pressure.

Dec 28, 2023

By A. Pawelczyk, D. Hreniak, E. Zych, W. Strek, K. Grabas, A. Ostrowski

"Mongolia is a country rich in rare earth resources. In the experiments carried out samples of the Mongolian ore were processed to obtain a lanthanide concentrate. The concentrate underwent hydrometallurgical processing. At the same time radioactive elements were removed. The lanthanides were obtained in the form of hydroxides that were further processed in order to get concentrated products. As a result cerium, lanthanum and neodymium-praseodymium concentrates were separated in the process. Also a concentrate of remaining elements: samarium, europium, gadolinium, terbium, dysprosium, holmium and erbium was obtained. INTRODUCTIONAccording to the assessment of the US Geological Survey, Mongolia possesses 16.77 % of the world total rare earth (RE) resources (Daly, 2011); that is the second largest RE reserve in the world, estimated to be 31 million tonnes in terms of rare earth oxides (REO). In the face of Chinese RE trade restrictions, Mongolia emerged as an important economical partner in mining and exploitation of these critical raw materials. This country is recognized as having more than five large RE deposits in the southern provinces. The most important are Lugeengol (Lugin Gol), Mushgia Khudag, Khotgor and Khalzan Buregtei (Eurasia, 2010). The Mongolian government has taken measures to facilitate western investments, nevertheless the country is economically dependent on China, the basic trading partner, particularly where exports are concerned.In the late 1970s Hydromet Kowary Enterprise, Poland, took up, along with the Wroclaw University of Technology, research and technological study on lanthanides recovery from raw materials of Vietnamese and Mongolian origin and from Indian monazite sands (Kowalczyk, 1990; Kowalczyk & Mazanek, 1987; 1990). For that purpose, a rare earth processing plant had been constructed with the capacity of about 10 t/year of hydroxide concentrate. The facility included several technological operations like grinding, flotation, hydroxide concentrate production in the hydrometallurgical processing, particular RE separation, and production of commercial products in form of CeO2, La2O3, Nd2O3, Pr2O3, and misch metal. Additionally work had been carried out in Kowary on the manufacture of yttrium and europium oxides of high purity, which were intended for color TV picture tube production in Poland."

Jan 1, 2012

By V. S. Melamud, A. G. Bulaev

"Sulfide ore processing is characterized by significant losses of metals, for example, with flotation tailings. Therefore, flotation tailings can present an important source of metals. The treatment of flotation tailings is a technological issue due to low content of valuable metals (e.g. copper and zinc) and high iron content. The goal of the present work was to investigate extraction of base metals and gold from old flotation tailings containing 17% iron, 0.26% copper, 0.22% zinc, and 0.7 g/t gold. Chemical leaching in columns with 0.5 to 10% sulfuric acid solutions and bioleaching in air-lift percolators were used for the extraction of base metals. 1% H2SO4 provided high extraction of base metals and best selectivity. Recovery of copper and zinc reached 35% and 40%. Increasing H2SO4 concentration did not lead to significant increase in base metals extraction, but led to increased iron concentration in the pregnant solutions. The 1% H2SO4 leaching solid residues were used for further experiments. Acid leaching and bioleaching did not result in significant increase in base metals recovery. Leaching of the residue with 10% H2SO4 provided extraction of 4% copper and 20% zinc, and bioleaching for 100 days provided extraction of 16% copper and 12% zinc. The recovery of gold by cyanidation from the flotation tailings, the 1% H2SO4 leaching solid residue and residue of two-step acid leaching were 50, 49, and 65%, respectively. The results suggest that acid leaching can be effective approach for treatment of old flotation tailings.INTRODUCTIONAbout 80% of base metals (copper, nickel, zinc) are produced through pyrometallurgical treatment of sulfide ores. Mining and metallurgical treatments of sulfide ores are presently characterized by significant losses of non-ferrous and precious metals into the various wastes. For example, flotation of sulfidic ores yields formation sulfide flotation tailings. These wastes pose a problem not just because of their sheer volume, but some of them may affect local ecosystems. Nowadays, mining and metallurgical wastes can be considered as technogenical source of base metals."

Jan 1, 2016

By G. Michael Lloyd

The term phosphate beneficiation often conjures up visions of processing technologies that physically separate the phosphate rock from some impurity and do not result in a change in the chemical nature of the phosphate rock. However, when we consider hydrometallurgical processing as a phos¬phate beneficiation technique we have to include all chemical treatments that convert phosphate rock into a more useful product. With most of the phosphate rock mined being used for fertilizer pro¬duction, the best known and most widely employed hydrometallurgical process is wet process phospho¬ric acid manufacture. In general, all phosphate rocks are so insoluble that there is no physical treat¬ment that will allow the phosphate content to be used by vegetation in a timely manner. However, after chemical treatment, 100% of the phosphate is readily available for vegetation nutrition. If there is a chemical beneficiation process that could be considered the standard method of treat¬ing phosphate rock it would have to be the wet pro¬cess for phosphoric acid production. In this process finely ground phosphate is reacted with sulfuric acid to give phosphoric acid and phosphogypsum. The phosphogypsum crystals are removed from the phosphoric acid solution and discarded; the phos¬phoric acid is further processed to give the finished fertilizer product. The primary reason that this pro¬cess was adopted so widely is because the phosphogypsum solid that is formed can be easily separated from the phosphoric acid, thereby removing the calcium portion of the phosphate rock and the sul¬fate portion of the sulfuric acid. Other acids have been used to treat phosphate rock but nitric acid is the only other acid used commercially. Nitric acid gives a mixture of phos¬phoric acid and nitrates that can be further pro¬cessed into N-P fertilizers, but the range of N/P grades that can be produced is limited. Hydrochlo¬ric acid has been used to react with phosphate rock, and it can be used to generate a pure phos¬phoric acid but only after solvent extraction to separate the calcium chloride from the phosphoric acid. And then there is the problem of what to do with the calcium chloride. The acid that has received the most attention as a substitute for sulfuric acid is phosphoric acid. In this process, the phosphoric acid reacts with the phosphate rock to give a mono-calcium phosphate solution. This process does not dissolve all the impurities in the phosphate rock, and these solids can be removed from the solution before sulfuric acid is added to convert the soluble calcium to insoluble phosphogypsum. Despite all the claimed advantages, this process has not found favor. A suc¬cessful application of this process to treat some of the lower grade phosphate rocks being used today would seem to have some distinct advantages. Other beneficiation schemes have been pro¬posed or practiced on specific phosphate rocks on a limited scale. These include calcination followed by water leaching to remove lime, weak acid

Jan 1, 1999

The NorthMet deposit of PolyMet Mining is located in northern Minnesota, adjacent to the Iron Range. The deposit is polymetallic with values of Cu, Ni, Co, Zn, Pt, Pd and Au. The processing route selected for extraction of these values utilizes bulk sulfide flotation followed by hydrometallurgical treatment. The hydrometallurgical process has been thoroughly tested on the bench and pilot plant scale at SGS Lakefield Research. The major steps in the process are; ? Chloride-assisted total pressure oxidation for extraction of base and precious metals ?Thickening/filtration/washing of the barren residue from total oxidation ? Reduction and precipitation of Au, Pt, Pd from the oxidation solution ? Partial neutralization of autoclave acid using limestone ? Copper SXlEW for production of copper cathode ? Treatment of a portion of the raffinate from Cu SX to recover NilCo/Zn as either a mixed Ni-Co-Zn hydroxide or as separate Ni hydroxide, Co hydroxide and a Zn sulfate solution. ? Recycling of the remaining raffinate to the autoclave as coolant The process flowsheet, chemistry and bench/pilot plant test results will be reported in this paper.

Jan 1, 2007

By A. N. Zagorodnyaya

The results of research on recovery of ammonium perrhenate from by-products of lead production are presented in the article. For the first time in the world, rhenium marketable products were produced on lead plant of Kazakhstan. Distribution of rhenium in lead production has been studied. During oxidative agglomeration of lead charge, 67.43% rhenium is sublimated and at that, 67.27% rhenium is trapped by the dusts of a system of gas purification, out of them 6.21% is transferred into different solutions, 1.89% into slime and only 0.09% into exhaust. During agglomerate smelting, up to 72% of its rhenium is sublimated and condensed in the dusts also. At leaching stage of sulphate product under dusts processing by the method of sulphatization, 15% of its content in the dusts is transferred into solution. The sorption method on the ionite ?N-21 has been proposed for recovery of rhenium from the solutions of sulphatization. At the same time, rhenium recovery from the dusts into commodity was 13%. The method of direct hydrometallurgy - leaching the dusts by solutions of sulphuric acid under presence of manganous concentrates, has been proposed as alternative method of recovery. This method has allowed increasing recovery of rhenium and other metals from the dusts into solution up to 90%~95%. Rhenium solvent extraction by ??? has been proposed for recovery of rhenium from the leaching solutions. At that, extraction of rhenium into marketable product from the dusts is above 75%.

Jan 1, 2014

By J. Brent Hiskey, Srini Raghavan

The physical chemistry of dissolution of silicate and complex oxide ores in aqueous lixiviants has been reviewed. Particular attention has been paid to the analysis of dissolution mechanisms and kinetic models and to the elucidation of the role of the porous nature of the materials on dissolution behavior. Importance of the physico-chemical characteristics of the leach residues on leaching efficiency has been discussed.

Jan 1, 1989

By Wasyl Kunda

High grade jig and low grade flotation concentrates used in the following experiments, are produced from silver ore at the mine. The study was carried out to recover the silver and other metals from these concentrates by hydrometallurgical methods. The concentrates were subjected to two stage oxida¬tion leaches in a HNO3-H2SO4 medium. The dissolved metals were recovered by selective precipitation of: (a) silver chloride, (b) iron-arsenic-bismuth residue, (c) copper sulphide, (d) cobalt-nickel sulphide. The effluent solution contained only ammonium sulphate and ammonium nitrate. The silver chloride was further processed to the high purity silver powder by the AgCl ? Ag2S ? Ag2SO4 ? Ag° route. The recovery of silver from the concentrates was over 98% and other metals from 95 to 99%.

Jan 1, 1984

By D. Lunt

Telfer bulk concentrates typically contain 8% chalcopyrite, 8% chalcocite, 45% pyrite and 80-120 g/t gold. Although very high copper and gold recoveries can be obtained by subjecting the concentrate to a "conventional" high temperature pressure oxidation step followed by copper recovery by solvent extraction and electrowinning, and cyanidation of the leach residue for gold recovery, the cost of lime required for neutralisation and power for oxygen generation was such that the process was uneconomic in terms of operating costs. The challenge therefore was to develop an alternative flowsheet that maximised copper and gold recovery in which the reagent costs were commercially attractive. An additional objective was to reduce the capital cost so that the overall NPV and IRR values would enhance the economic viability of the total project. Following a series of batch and locked cycle testwork programs, supported by Metsim modelling, it was established that the same high levels of overall copper and gold recoveries can be obtained by use of a novel two-stage split leaching circuit in which the chalcocite component is leached in a low temperature (90-110°C) circuit with the chalcopyrite component treated in a high temperature (220-250°C) circuit. This split circuit has a number of advantages, including minimising the required capacity of the high temperature autoclave, as well as reducing the overall lime consumption and simplifying the total process water balance. The paper outlines the basis for the development of the flowsheet and discusses the integration of the leaching stages into the overall circuit.

Jan 1, 2004

By A. V. Tarasov

Based on the data available in the literature and the practical experience of new metallurgical plants, a method has been proposed and relevant studies conducted for metallurgical processing of zinc sulfide concentrates. Process conditions and parameters have been identified, which ensure high levels of recovery of zinc into solution. This is achieved by leaching concentrates under the normal pressure at a temperature of >90°C in the presence of ferric iron ions. Sulfide sulfur is converted to elemental sulfur in this process. To produce elemental sulfur it is necessary to maintain oxidizing conditions in the system, characterized by a positive Eh value. In order to eliminate the interfering effect of elemental sulfur formed in the process, it is possible to add surfactants into the concentrate pulp. The proposed technology has certain advantages as compared with the high-temperature pressure leaching and is considered as a basic technology to be applied to increase zinc production at one of the Russian zinc plants.

Jan 1, 2006

By I. H. Warren, Forward F. A

The reduction of aqueous slurries of uranyl carbonate at elevated temperature with compressed hydrogen m the presence of metal catalysts and certain organic promoters has been shown. to yield uranium dioxide. The reduction takes place in a sequence of steps. The purity and sintering properties of the dioxide make it of interest as a starting material for fuel element production. Simple control can be exercised over the surface area and oxygen/uranium ratio of the product during reduction. URANIUM dioxide for fueling atomic power reactors 1s conventionally produced as a powder by simultaneously decomposing and reducing ammonium diuranate with hydrogen in a furnace at high temperature. Fuel elements are then generally made from this powder by compacting and sintering ( 1). In order that elements approaching theoretical density may be obtained,

Jan 1, 1961

By T. R. Heflin

The paper provides a comprehensive discussion of pump selection for the hydrometallurgical industry. Material suitability, in addition to proper size and design, will provide extended life for pumps in erosion /corrosion service.

Jan 1, 1979

By Viviane Tavares de Moraes, Denise Crocce Romano Espinosa, Jorge Alberto Soares Tenorio, Adriana Johanny Murcia Santanilla

"Increasing volumes of solid waste are generated worldwide, due to the continuous technological advances, the electronic scrap is part of these waste, requiring the study of recycling processes to stimulate this activity or the reuse of its components. The separation of nickel from sulphate leach solution, which also containing iron, zinc and aluminium was carried out using bis-(2,4,4-trimethylpentyl) phosphinic acid (Cyanex 272) as extractant, in order to determinate the effects of pH and temperature. The number of stages required for extraction was also assessed. The results showed that the metal extraction increased with increasing pH. On the other hand, increasing temperature was not favorable for all metals. However, under conditions used in this paper the purification of solution containing nickel was possible.IntroductionIncrease in the demand of metals in the world for use in industry causes a great exploration of natural resources. One outcome of this is increased investigations for the recovery of metals from industrial wastes. Nickel is found in wastes, such as batteries and electric and electronic devices, this kind of waste becomes a significant source for metals recovery. Since nickel is also present in printed circuit boards, which are associated with electric and electronic equipment, it can be recovered from this kind of waste through different techniques. It is important to highlight that iron is generally associated with this scrap [1] and for its removal chemical precipitation is often used. On the other hand, the solvent extraction process is widely used for treatment of leach liquors, because it can be produce high purity solutions [2, 3]. This technique is based on the purification of solution using contact between a liquor leaching and the immiscible organic solution, which contains an extractant that promotes the desired metal transfer [4]."

Jan 1, 2012

By Jean Goldschmidt

Recycling of fine oxidized particles of non-ferrous metals usually cannot be carried out by pyrometallurgical methods. 'Hydrometallurgy provides an alternative technology to reintroduce? the resulting products as secondary raw materials into the industrial cycle, thereby saving resources and energy while keeping the environment a little cleaner. The paper summarizes the methods employed by Hydrometal to achieve these objectives and? stigmatizes the latest counterproductive international regulations which create ever more obstacles to environmentally sound recycling.

Jan 1, 1995

By M. Wenzel, Dang Thanh Tung, Nguyen Anh Duc, Nguyen Huu Luong, Dang Van Sy, N. Kelly, K. Gloe, K. Kretschmer, Phuc Nguyen Le, K. Schnaars, S. Robles M, L. Götzke, J. J. Weigand

The recovery of rare earth metals from secondary sources has attracted much attention due to their ever expanding demand in the high-tech industry. The studies reported here focus on the hydrometallurgical recovery of lanthanum and cerium from spent fluid catalytic cracking (FCC) catalysts in a two-step process: leaching with nitric acid and solvent extraction by tri-n-butyl phosphate (TBP) and di(2-ethylhexyl)phosphoric acid (D2EHPA). The experiments show a high dissolution yield of about 93% lanthanum and 42% cerium in a single leaching step with 2 M (126 g/L) HNO3 at 80 °C; only 11% aluminum has been dissolved simultaneously. In the subsequent solvent extraction step the best results for this leach liquor could be achieved using a 1:1 mixture of 25% (v/v) TBP (0.92 M) and 25% (v/v) D2EHPA (0.76 M) in n-decane without the need for any pH adjustment. In that case La(III) and Ce(III) can be extracted with 60% and 74% yield respectively in one stage from the majority of accompanying matrix elements. In particular no extraction of Al(III) could be observed under these conditions.

Jan 1, 2016

By Caelen D. Anderson

An investigation of a hydrometallurgical method for the recovery of tellurium dioxide from cadmium telluride (CdTe) thin-film manufacturing scrap was conducted. Sulfuric acid/ hydrogen peroxide lixiviation was employed to selectively leach cadmium and oxidize tellurium to tellurium dioxide (TeO2). Experimental variables included solution pH, pulp density, oxidant (H2O2) concentration, temperature, agitation rate and residence time. After one hour of lixiviation, no noticeable increase in grade or recovery occurred; thus, all experiments were performed for one hour. Leaching results showed it was possible to recover at least 74% of the tellurium, while producing a 98% pure TeO2 product. Additionally, an apparent activation energy value of 12.1 kJ/mole (2.9 kcal/mole) was determined. This value, along with results from agitation rate experiments, indicated the reaction was primarily controlled by diffusion through the product layer. During lixiviation, partial dissolution of tellurium proved inevitable. Thus, a second step: ammoniacal precipitation was introduced to recover the aqueous tellurium as a precipitate. A statistical experimental design (DOE) was used to identify the significant effects from the residual effects. Experimental variables included precipitation pH, temperature, initial tellurium tenor and agitation rate. Precipitation results showed that it was possible to produce a low-grade (46% Te) precipitate, while recovering 98% of the aqueous tellurium species. The precipitation pH and temperature (°C) were determined as the significant effects for both the grade and recovery of tellurium. Agitation rate and initial tellurium tenor were determined as the residual effects and did not influence the grade or recovery of tellurium. An illustration of tellurium recovery, as a function of the significant effects, is illustrated in Fig. 1.

May 1, 2012

By J. -F. Blais, M. -O. Simonnot, J. Mocellin, G. Mercier, J. -L. Morel

Large quantities of ferromanganese sludge are generated as waste material by blast furnace during the manufacturing of ferromanganese. These slags are very rich in manganese, zinc and lead (5 to 40 %). These residues have been deposited in a pond in periphery of the steel industry. Considering the market value of the metals, these brownfield sites can now be considered as secondary resources. We have developed a hydrometallurgical process for selectively leaching Zn, Mn and Pb in order to produce compounds of Mn and Zn pure enough to be economically recoverable and a residue rich in Pb. Batch laboratory experiments were carried out to determine appropriate leaching conditions to maximize zinc extraction (100 %) and manganese extraction (90.0 %) with sulfuric acid and to generate a residue containing 15 to 30% of lead. This Pb residue is recyclable by another process.

Jan 1, 2015

By Cleusa Cristina Bueno Martha de Souza

Nowadays in Brazil the final disposal for spent batteries include sanitary and industrial landfill for hazardous waste from domestic waste and industrial waste respectively. The problems of environmental impact caused by spent batteries had been discussed of several years, considering the fact of metals contents, like Cd, Hg, Pb and Zn. Some techniques are been proposed for recycling this type of waste in order to protect the environment while bringing also economic advantages of recovering metals. The reclamation of metals from secondary raw materials had been employed and further applications of this flexible technology are planned to the future. The processing of metals solutions involving hydrometallurgical techniques is an efficient method for recovering metals, since it provides completely recovery of metals with low energy requirements. This paper discusses the spent alkalline batteries characterization and leaching stage experiments results with sulphuric acid as leachant. After batteries dismantling by mineral processing techniques, the black powder sample produced was submitted to diffraction X-ray characterization and atomic absorption spectrophotometry in order to verify all the substances presented. Some experiments were carried on to develop a hydrometallurgical process for initial metal recovery aiming to get Zn into a soluble form on aqueous solution in order to propose the next stages of purification and recovering Zn, as electroplated deposition. The sulphuric acid was chosen not only for being cheaper than other leachants but also for improving the extraction of soluble Zn as ZnS04 form at diluted acid concentration. Batch laboratory experiments leaching procedure were conducted to determine appropriate leaching conditions from the viewpoint of maximize zinc extraction. In these tests an amount of dry powder batteries was added to different acid concentrations and solid/liquid ratio and after leaching and filtration the aqueous solutions were submitted to atomic absorption spectrophotometry analysis to verify the Zn content.

Jan 1, 2000

By Bernard H. Coyle, Robert N. Anderson, Judith A. Koperski

Between 6 and 8 million cars are shredded annually in the U.S. This shredded material is magnetically separated leaving a mixture of nonmetallics and nonferrous metals which can be further separated into its metallic and nonmetallic constituents. The metallic portion is a mixture of metalr; and alloys containing predominately zinc, and some aluminum, copper, and residual iron. Hydrometallurgical strategies are proposed for recovering the aluminum, copper, and zinc. The thermodynamic and kinetic behavior of the most promising alternatives to separation are discussed.

Jan 1, 1978

By C. Sist, J. Shannon, J. Fossenier, J. F. Adams

With the cleanest and easiest to treat copper concentrates dwindling and becoming harder to find, electrorefineries must deal with the reality of higher impurity anodes. There are well-known strategies for managing nickel, bismuth, antimony, etc. using bleeds (Ni, Bi, Sb, As) or arsenic control (Bi, Sb) or hydrometallurgical add-on circuits (IX, MRT, SX for As, Bi, Sb). However, for integrated copper producers and those producing other base metals it is often worth considering a more holistic, global approach. This paper will discuss both the well-known strategies as well as some global strategies and the role that hydrometallurgy can play in them. Examples will include atmospheric and pressure leaching, selective precipitation and impurity recycling or fixation. INTRODUCTION In the electrorefining of copper, impure copper anodes are electrically corroded, and pure copper is deposited on cathodes by passing DC current through an acidic copper sulphate solution. The anode copper impurities either settle to the cell bottom as insoluble components (anode slimes) or remain in solution with minimal deposition at the cathodes when correct operating parameters are maintained. Consideration of impurities is important in all copper electrorefineries. A typical impure anode contains 98.5–99.5% copper (Schlesinger et al., 2011). Table 1 shows the range of impurities reported for all copper anodes in the 2007 survey of operating refineries (Moats et al., 2007). Also shown is an approximate “limit” of anode compositions within which the majority of refineries operate without additional impurity removal capacity (beyond conventional liberator/bleed circuits). The column is somewhat subjective and meant to be an approximation only since there are many factors that influence impurity deportments and refinery impurity tolerances. More recent world survey data from 2013 show the averages for all impurities trending upward from 2003 to 2013 (Moats et al., 2013). With a competitive market for good concentrate and more copper recycling from electronics (which contain high levels of impurities such as nickel, lead and antimony), this trend is expected to continue through the coming years. The traditional primary method of impurity control is refinery bleed through liberator cells which are employed in the vast majority of refineries (Moats et al., 2013). These are used to control copper (primary liberators), nickel, arsenic, antimony and bismuth (secondary and tertiary liberators). Other methods of impurity control include arsenic addition (upstream in the smelter deporting to anode or directly to electrolyte) to control antimony and bismuth (Kamath et al., 2003), ion exchange or molecular recognition technology (Hur et al., 2005; Navarro et al., 2012) for antimony and bismuth. Many refineries also operate nickel evaporators to remove nickel sulphate before recycling back acid. Some use, or are contemplating using, acid purification units to recycle acid and arsenic without nickel and making nickel carbonate (Kamath et al., 2003; Sheedy et al., 2006).

Jan 1, 2019