Hydrometallurgy Journal Impact Factor & Information

Publisher: Elsevier

Journal description

Hydrometallurgy aims to compile studies on novel processes, process design, chemistry, modelling, control, economics and interfaces between unit operations, and to provide a forum for discussions on case histories and operational difficulties. Topics covered include: leaching of metal values by chemical reagents or bacterial action at ambient or elevated pressures and temperatures; separation of solids from leach liquors; removal of impurities and recovery of metal values by precipitation, ion exchange, solvent extraction, gaseous reduction, cementation, electro-winning and electro-refining; pre-treatment of ores by roasting or chemical treatments such as halogenation or reduction; recycling of reagents and treatment of effluents.

Current impact factor: 2.22

Impact Factor Rankings

2015 Impact Factor Available summer 2015
2013 / 2014 Impact Factor 2.224
2012 Impact Factor 2.169
2011 Impact Factor 2.027
2010 Impact Factor 1.917
2009 Impact Factor 2.078
2008 Impact Factor 1.747
2007 Impact Factor 1.324
2006 Impact Factor 1.227
2005 Impact Factor 1.163
2004 Impact Factor 1.088
2003 Impact Factor 1.14
2002 Impact Factor 1.087
2001 Impact Factor 0.654
2000 Impact Factor 0.846
1999 Impact Factor 0.693
1998 Impact Factor 0.662
1997 Impact Factor 0.575
1996 Impact Factor 0.483
1995 Impact Factor 0.555
1994 Impact Factor 0.59
1993 Impact Factor 1.255
1992 Impact Factor 0.811

Impact factor over time

Impact factor
Year

Additional details

5-year impact 2.34
Cited half-life 8.00
Immediacy index 0.22
Eigenfactor 0.01
Article influence 0.54
Website Hydrometallurgy website
Other titles Hydrometallurgy (Online)
ISSN 0304-386X
OCLC 38901127
Material type Document, Periodical, Internet resource
Document type Internet Resource, Computer File, Journal / Magazine / Newspaper

Publisher details

Elsevier

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  • Post-print
    • Author can archive a post-print version
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    • Deposit due to Funding Body, Institutional and Governmental policy or mandate only allowed where separate agreement between repository and the publisher exists.
    • Permitted deposit due to Funding Body, Institutional and Governmental policy or mandate, may be required to comply with embargo periods of 12 months to 48 months .
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    • Published source must be acknowledged
    • Must link to journal home page or articles' DOI
    • Publisher's version/PDF cannot be used
    • Articles in some journals can be made Open Access on payment of additional charge
    • NIH Authors articles will be submitted to PubMed Central after 12 months
    • Publisher last contacted on 18/10/2013
  • Classification
    ​ green

Publications in this journal

  • [Show abstract] [Hide abstract]
    ABSTRACT: In this research, W and Mo were separated and recovered from high-molybdenum synthetic scheelite by leaching with HCl solutions containing H2O2 as the chelating agent. The effects of the leaching parameters, such as temperature, leaching time, HCl concentration and H2O2 concentration, on the recovery of W and Mo were investigated. The results show that W and Mo were effectively extracted at room temperature and low acidity. With a leaching temperature of 30 °C, an L:S ratio of 10:1, an HCl concentration of 1.5 mol/L, an H2O2 concentration of 2.5 mol/L and a leaching time of 40 min, the recovery of W and Mo was 97.7% and 99.4%, respectively. Thermal decomposition was then employed to treat the leaching solution. W was transformed into tungstic acid, whereas Mo remained in solution. At 60 °C, 97.47% of W was precipitated, and the separation factor reached 105.
    Hydrometallurgy 05/2015; 155. DOI:10.1016/j.hydromet.2015.03.020
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    ABSTRACT: The regeneration of strong (≥ 6 mol/L) hydrochloric acid (HCl) is an important but capital and energy intensive unit operation in chloride metallurgical facilities. This paper discusses a novel low temperature process capable of producing super-azeotropic hydrochloric acid via the reactive crystallization of calcium sulfate α-hemihydrate, a valuable by-product, in a continuous stirred-tank reactor (CSTR). The design of the process was enabled via model-based water activity calculations that allowed for selecting process conditions promoting the metastability of calcium sulfate α-hemihydrate in strong HCl solutions, while preventing the formation of the undesirable calcium sulfate anhydrite phase. The CSTR crystallizer was operated in the range from 13 °C to 60 °C by reacting concentrated CaCl2 feed solution (up to 5 mol/L) with concentrated H2SO4 (up to 18 mol/L). Super-azeotropic HCl (up to 9.5 mol/L, ~ 30 wt.%) and calcium sulfate α-hemihydrate (α-HH) was produced at 60 °C. Calcium sulfate dihydrate (DH) could also be produced but with lower strength HCl (sub-azeotropic ≤ 6 mol/L) and at ≤ 40 °C. By maintaining a low CaCl2 concentration (0.1–0.4 mol/L) the growth of metastable α-HH crystals (30–60 μm crystal size) was promoted due to a favorable supersaturation regime. Finally, it was determined that the super-azeotropic HCl producing crystallization process can be carried out in a thermally autogenous mode at 60 °C taking advantage of the exothermic character of the reaction.
    Hydrometallurgy 05/2015; 155. DOI:10.1016/j.hydromet.2015.03.019
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    ABSTRACT: Lead (Pb) is known for its high and long-term toxicity even at very low concentrations and its presence limits the reusability and recyclability of industrial wastewaters. In this study, Pb was recovered from synthetic wastewater as PbCO3 crystals via the fluidized-bed crystallization process. Pb removal exceeded 99% when the operating pH was kept within a range of 8–9 and the carbonate-to-lead molar ratio was set to 3:1. Higher Pb removals were obtained with the gradual reduction of recirculation flow rate. The present study demonstrates the potential of Pb recovery from contaminated wastewaters using fluidized-bed crystallization technology.
    Hydrometallurgy 05/2015; 155:6-12. DOI:10.1016/j.hydromet.2015.03.009
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    ABSTRACT: This paper reports on continuous testing of a process for the recovery of vanadium from a H2SO4-HF solution, generated from the leaching of stone coal, by solvent extraction using a D2EHPA/TBP mixture as the extractant. Three unit processes are included: fully continuous countercurrent solvent extraction and stripping; oxidation and precipitation; and calcination. The solvent-extraction circuit was operated for 96 h. With six stages of countercurrent extraction, 97.7% vanadium, 98.3% molybdenum, and 97.9% titanium were extracted, while the sodium, potassium, fluorine, and nickel impurities were barely extracted. Scrubbing removed co-extracted and aqueous-entrained zinc, magnesium, silicon, and arsenic impurities. Most co-extracted impurities were scrubbed in two stages. Using five stages of stripping, more than 99.8% of the vanadium was stripped, yielding a vanadium concentration in the loaded strip solution of 34.1 g/L. Co-extracted molybdenum(VI), titanium(IV), and iron(III) remaining in the stripped organic phase were removed by a saturated sodium carbonate solution and the regenerated organic was reused for vanadium extraction. Vanadium was selectively precipitated with 98.5% efficiency from the loaded strip solution by oxidation and precipitation with ammonium sulfate. A high-purity V2O5 (99.61%) product was successfully produced by this flowsheet.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2014.11.008
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    ABSTRACT: This work is focused on the behavior of zinc, iron and calcium during leaching EAF dust in sulfuric acid solutions. The influences of leaching time, temperature, sulfuric acid concentration and L:S (Liquid:Solid) ratio on the extraction of zinc, iron and calcium were studied. The leaching experiments were preceded by a thermodynamic study consisting of calculations of ΔG0 and E–pH diagrams of the Zn–Fe–Ca–S–H2O system. A thorough characterization of the input sample of EAF dust was performed. The highest zinc extraction, 87%, was achieved by using 1 M H2SO4 at the temperature of 80 °C and L:S ratio 50. From the perspective of a selective leaching of zinc, where no iron is passing into the solution, concentration 0.1 M H2SO4 at L:S ratio = 50 and 0.25 M H2SO4 at L:S ratios = 20 and 10 can be considered as optimal concentrations. The calculated values of the activation energy confirmed different mechanisms of leaching zinc, iron and calcium. The rate limiting step in the case of zinc and calcium is represented by a rate of diffusion, while in the case of iron it is a rate of chemical reaction.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.03.008
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    ABSTRACT: Sodium tungstate solution can be prepared by decomposing wolframite with sodium hydroxide. A new technique is developed for precipitating scheelite from the sodium tungstate solution by the addition of calcium hydroxide. In this study, the effects of several variables on the conversion of sodium tungstate are investigated. Results show that the effects of temperature and sodium tungstate concentration are significant. Specifically, 96.4% WO42 − was converted to scheelite in 2 h at 100 °C with a calcium hydroxide stoichiometric ratio of 1.4 and a stirring speed of 350 rpm. The parent solution containing sodium hydroxide can be recycled as a leaching agent return to the wolframite decomposition process.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.03.010
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    ABSTRACT: In this work the comparative studies between solvent extraction and transport across polymer inclusion membranes (PIMs) have been carried out for the removal of cadmium(II) from aqueous chloride solutions. Cyphos IL 104 (trihexyl(tetradecyl)phosphonium bis(2,4,4-trimethylpentyl)phosphinate) was used as extractant/ion carrier in both processes. Effect of different parameters such as hydrogen ion concentration, chloride ion concentration as well as hydrochloric acid concentration in aqueous phase on cadmium(II) extraction has been investigated. Cellulose triacetate membranes with Cyphos IL 104 as carrier have been prepared and applied for removal of Cd(II) from HCl and NaCl solutions. Cd(II) ions were effectively removed from the source phase by transport through PIM into 1M H2SO4 as the receiving phase. Also separation of Cd(II) was carried out from 1 M HCl solution containing Cu(II), Co(II) and Ni(II) and the separation coefficients were found in order of SCd/Cu < SCd/Co < SCd/Ni. The data generated from the transport of Cd(II) study were compared with that of conventional solvent extraction investigation.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.04.007
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    ABSTRACT: High-grade natural chalcocite (Cu2S) was leached in column under controlled redox potential (Eh). To satisfy the requirements of the shrinking core model, the conditions in the column were kept as uniform as possible by adopting a short height of the middle chalcocite layer, high acid Fe2(SO4)3 feed and high irrigation rate. The effects of temperature, Fe3 + concentration and Eh on the second stage of chalcocite leaching were investigated. The second stage is sensitive to temperature whereas insensitive to the variations in Eh and Fe3 + concentration above a certain level. The second stage was divided into two “sub-stages” based on the inflection point at approximately 70% copper dissolution. The excellent linear relationship plotted by the mixed-kinetics indicated that the second stage is controlled by the rates of both chemical reaction and diffusion, in which the limitation of the chemical reaction rate is dominant. The second “sub-stage” (> 70% dissolution) is responsible for the slow kinetics of the second stage. Mineralogical study of the residues confirmed that a S0 product layer covered around a shrinking CuS core during the second stage. In addition, high temperature resulted in more porous S0 layer which remarkably influences ion diffusion rate. A dissolution model based on the shrinking core model was proposed. Implications of the findings for heap bioleaching were discussed.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.022
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    ABSTRACT: Recycling of electroplating wastewater is very important because of the large amounts of wastewater generated, the enormous economic value and large environmental concerns when conventional neutralization-precipitation processes are used to dispose of the wastewater. This work describes the recovery of valuable metals from industrial wastewater collected at a galvanizing industrial site using a solvent extraction process. Several extractants such as DEHPA, TBP, LIX 984N-C, Cyanex 272 and Aliquat 336 were tried for the separation and recovery of valuable metals from the wastewater. Optimum extraction parameters for zinc, copper, iron, nickel and chromium were determined as 10% Aliquat 336, 5% LIX 984N-C, 10% DEHPA, 15% LIX 984N-C and 10% Cyanex 272 in kerosene at equilibrium pH 1.45, 1.20, 1.00, 5.25 and 6.00 with 99.6%, 100%, 100%, 99.9% and 100% extraction efficiencies at almost 100% selectivity, respectively. These extraction efficiencies were attained at two extraction stages for zinc, copper and chromium and at one extraction stage for iron and nickel. Copper-loaded LIX 984N-C, iron-loaded DEHPA, nickel-loaded LIX 984N-C and chromium-loaded Cyanex 272 were stripped with 450 g/L, 2 M, 150 g/L and 1 M H2SO4 respective strip solutions with relative 100% stripping efficiencies. A two-stage process stripped the zinc-loaded Aliquat 336 with a 2 M NaOH solution with 100% efficiency. Based on the data, a flow sheet for the separation of the five metal ions is provided.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.021
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    ABSTRACT: In (III) ion-imprinting polymer (In (III)-IIP) was prepared with the surface imprinting technique by using indium (III) ion as the template ion, vinylphosphonic acid (VPA) as the functional monomer and silica gel grafted as the support. The In (III)-IIP prepared was characterized by Fourier transform infrared spectroscopy (FTIR), scanning electron microscopy (SEM), Brunauer–Emmett–Teller (BET), thermogravimetric analysis and energy dispersive X-ray spectrometer (EDX). The adsorption behavior of the In (III) ion-imprinting polymer and non-imprinting polymer (In (III)-NIP) was investigated in detail by batch and column experiments. Optimum pH for maximum sorption was found to be 3.0. Kinetic studies revealed that both adsorption processes could be better described by pseudo-second-order kinetic model, while the sorption equilibrium data agreed well with the Langmuir isotherm model. The maximum adsorption capacities calculated from the Langmuir isotherm were 47.39 and 31.11 mg·g− 1 for In (III)-IIP and In (III)-NIP, respectively. The selectivity coefficients of the In (III)-IIP for indium (III) ion in the presence of Cu (II), Pb (II), Zn (II) and Fe (II) are found to be 189.17, 67.94, 886.63 and 2479.71, respectively. In addition, it was observed that the dynamic adsorption capacity of the In (III)-IIP was close to the static adsorption capacity due to the fast adsorption rate. Indium (III) ion can be separated selectively from the solution containing various foreign ions by column packed with the In (III)-IIP and can be pre-concentrated, and the enrichment factor reaches up to 80.0. The In (III)-IIP prepared was repeatedly used and regenerated six times without a significant decrease in adsorption capacity, demonstrating that the sorbent is of high stability and reusability. Moreover, the developed method was successfully applied to the selective extraction of indium (III) ion from the real water samples with satisfactory results.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.03.011
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    ABSTRACT: Enhanced separation of sodium and aluminum in Bayer spent liquor has been a challenge of the alumina industry for decades in its effort to obtain a stream of caustic solutions with high MR (molar ratio of Na2O to Al2O3) to treat red mud. The existing methods are too energy-intensive to use in practice. Solvent extraction technology has attracted considerable attention owing to its high selectivity and efficiency. However, the extraction capacities of extractants as reported in the literature are insufficiently high, requiring multiple-stage extraction and high organic-to-aqueous phase ratio. In this paper, the mechanisms for the extraction of sodium from Bayer liquor using alkyl phenol weak organic acids are investigated in terms of their structures and acidities, and an optimized extractant consisting of 2-tert-butyl-4-methylphenol dissolved in 1-octonal was selected due to its high sodium extraction capacities. The equilibrium relations between equilibrated organic sodium concentration and aqueous alkaline concentration for sodium extraction from sodium aluminate solutions under different concentrations of organic acid at various temperatures were systematically investigated. The equilibrium data obtained using 1 mol/L 2-tert-butyl-4-methylphenol diluted by 1-octonal at 40 °C for extraction and stripping were used to design the extraction parameters to achieve a sodium extraction efficiency of 50%. The results of the extraction of a two-stage countercurrent experiment in laboratory scale verified the design parameters. Sodium extraction efficiency of 40.4% was achieved. Concurrently near-pure alkaline solution with about 210 g/L NaOH was obtained from the stripping unit, which could be evaporated to treat red mud.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.03.013
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    ABSTRACT: The flotation separation is the general method to recover molybdenite from a concentration containing molybdenite and bismuthinite, although the whole circuit could be long and complicated. In this paper, a novel selective electro-oxidation of molybdenite process has been introduced and systematically investigated for bismuthinite purification. It was found that Mo could be selectively separated by the electro-oxidation method under the conditions of pH not less than 9.0 and the applied potential at between 3.0 V and 4.0 V, by this circumstance Bi was hard to extract. The effects of NaCl concentration, liquid/solid ratio and stirring speed on electro-oxidation leaching were investigated. Optimum leaching conditions were found as follows: operated at room temperature, leaching time = 2.5 h, NaCl concentration = 4 mol/L, pH = 9–10, applied potential = 3.4 V, liquid/solid ratio = 30 mL/g, stirring speed = 400 rpm. Under these conditions, Mo extraction yield was obtained at 98.4% and 99.3% of Bi remained in the residue. The chlorine evolution reaction at the anode which was effected by leaching pH was studied by linear scan voltammetry. Mechanism of electro-oxidation leaching of Mo was studied by cyclic voltammetry. Furthermore, the mass transfer of Mo from leachate to organic phase was introduced, and the two phase transfer could obtain as high as 99.6%.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.04.012
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    ABSTRACT: This study was conducted to establish the possible application of bioleaching to recover valuable metals from the sulfidic tailing of Golgohar Iron Mine (Kerman, Iran). Shake flask leaching experiments were carried out in the presence of a mixed culture of moderately thermopilic microorganisms at a stirring rate of 150 rpm and a temperature of 45 °C, with the addition of yeast extract (%0.02 (w/w). The influence of bacterial inoculation, pH, nutrient medium type and pulp density on the recovery of copper, nickel and cobalt from the tailing was investigated. The results showed that 55.0% of copper, 98.2% of nickel and 59.5% of cobalt could be extracted from the tailing through the bioleaching process after 30 days at 5% (w/v) pulp density. The recovery of valuable metals from the tailing in the presence of microorganisms was approximately three times higher than that in the un-inoculated leaching experiment under similar conditions. It was also found that the recovery of copper at the initial pH of 1.2 was 17% higher than that at the pH of 1.8, while nickel and cobalt recoveries were 37% and 23% more at the pH of 1.8, respectively. It was also revealed that the recoveries of valuable metals were approximately similar in both 9 K and Norris nutrient media. Moreover, the results showed that in both nutrient media, copper recovery at the higher pulp density was significantly higher than that at the lower pulp density; this was mainly attributed to the lower redox potential. On the other hand, nickel and cobalt recoveries were higher at the lower pulp density, probably due to the higher redox potential values. A two-stage bioleaching process in which the redox potential is controlled at a low level in the primary reactors, followed by the secondary reactors with a high redox potential is supposed to achieve the maximum recovery of valuable metals.
    Hydrometallurgy 04/2015; 154. DOI:10.1016/j.hydromet.2015.03.006
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    ABSTRACT: Column leaching test was carried out for uranium ore samples obtained from a uranium in-situ leaching (ISL) mining site using 2.45 bed volumes (BVs) of 0.255 M sulphuric acid to study the evolution of uranium concentration and its isotopic 234U/238U activity ratio in the pregnant leach solution. The activity ratio of the two uranium isotopes in the pregnant leach solution was measured by alpha-spectrometry following radiochemical separation. Three phases of uranium leaching were identified based on the observed uranium concentration and the 234U/238U activity ratio. The first phase, up to 0.8 BVs, was characterised by low uranium extraction at approximately 3% and a 234U/ 238U activity ratio in the range of 1.21 − 1.25, which can be attributed to hexavalent uranium leaching from mineral grain surfaces. The second phase, up to 1.8 BVs, with somewhat over 80% uranium extraction is related to the oxidation and dissolution of tetravalent uranium from uranium-bearing minerals, mostly uraninite, with a 234U/238U activity ratio of 0.92 − 1.00. Uranium leaching was very slow in the third phase (between 1.8 and 2.45 BVs), with only 0.5% leached and the 234U/238U activity ratio raised to 1.05 − 1.16. The total recovery of uranium at 2.45 BVs was 85%.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.017
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    ABSTRACT: Coal gangue is regarded as a potential source of aluminum and silica. In this work, acid leaching behaviors of aluminum, iron, and silica in coal gangue were investigated using hydrochloric acid as leaching agent. Chemical and mineralogical compositions of coal gangue and the leached residue were determined by X-ray diffraction (XRD), X-ray fluorescence, and Fourier transform infrared spectroscopy. XRD results indicated that kaolinite and boehmite were the predominant minerals in coal gangue. During elevated temperature acid leaching, mineral phases gradually decomposed, and aluminum and iron ions were dissolved in the acid leaching solution. However, the carbonaceous material and the product of silica were not dissolved in acid and were abundant in leached residue. The extraction efficiency of Al and Fe from coal gangue depended on coal gangue size, liquid–solid ratio, temperature and leaching time. Extraction efficiency of Al and Fe increased with increased liquid–solid ratio, temperature, and leaching time. Under the same conditions, SiO2 content in the ash of leached residue also increased. The optimum technological conditions for aluminum and iron extraction were obtained with a coal gangue size of 74 μm at a liquid–solid ratio of 3:1 to react at 180 °C for 4 h. After evaporation and crystallization, selective precipitation by adding sodium hydroxide solution, carbon decomposition, and calcination, alumina with a purity of 98.70% was prepared from acid leaching solution. Finally, a suitable amount of low ash anthracite was introduced to the leached residue, and the product of SiC was produced by carbonthermal reduction method. The product had a yield and purity of 72.72% and 76.01%, respectively.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.018
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    ABSTRACT: The recovery of rare earths from ores and scraps requires the selective separation of dilute rare earth ions from solutions containing high concentrations of base metal ions. Using column tests under conditions mimicking practical use, we investigated a novel adsorbent comprising silica gel particles modified with diglycol amic acid groups (EDASiDGA). We evaluated the influence of flow rate on dysprosium adsorption–desorption behavior and the selective adsorption of dysprosium ions from a simulated solution containing dilute rare earth ions and high concentrations of base metal ions at low pH (pH 1.0). The EDASiDGA-packed column had a high adsorption–desorption rate. Moreover, the EDASiDGA-packed column enriched middle and heavy rare earths more than 10- to 20-fold from even a dilute solution. In addition, the molar ratios of the rare earths to base metals in the solution after desorption were three to four orders of magnitude higher than the initial molar ratios.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.015
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    ABSTRACT: In 2008, the nitric acid pressure leaching (NAPL) technology was patented and developed to treat laterite ores in China. In the following year, a pilot plant with an annual processing capacity of 330,000 tons of dry ores was assembled and tested. The pilot-scale tests were documented to illustrate the innovative technology. NAPL consists of six process steps: raw ore preparation, selective pressure leaching, pregnant leach liquor purification, Ni/Co intermediate product preparation, Mg precipitation, and HNO3 regeneration/recycle. The results of the treatment of limonitic laterite ores with NAPL are as follows: (i) the recoveries of both Ni and Co were over 82%, (ii) Ni/Co hydroxide with 25.4% Ni and 2.6% Co was obtained, (iii) above 85% of HNO3 could be regenerated/recycled, and (iv) several valuable by-products could be produced. The leach iron residue without sulfur in particular is marketable because of its application in iron making. Meanwhile, fibrous calcium sulfate used in papermaking was produced in the regeneration of HNO3. In the processing of high magnesium-bearing laterite ores, nickel, cobalt, and iron extractions reached over 98%, approximately 99%, and less than 1.5%, respectively. Such advantages make the NAPL technology for laterite processing profitable, as proven by preliminary economic accounting.
    Hydrometallurgy 04/2015; DOI:10.1016/j.hydromet.2015.04.016