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

  • Pre-print
    • Author can archive a pre-print version
  • Post-print
    • Author can archive a post-print version
  • Conditions
    • Pre-print allowed on any website or open access repository
    • Voluntary deposit by author of authors post-print allowed on authors' personal website, arXiv.org or institutions open scholarly website including Institutional Repository, without embargo, where there is not a policy or mandate
    • 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 .
    • Set statement to accompany deposit
    • 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 recent years, many researchers have studied processes for extracting gold and silver from ores using alternatives that are less harmful and aggressive to the environment than cyanides. Thiosulfate solutions represent one such alternative. Metallic silver can be dissolved in O2–thiosulfate in different experimental conditions. Silver dissolution was obtained without the formation of passivation layers on the surface of the silver plate used. The kinetic study conducted indicated that the process is affected by a stirring speed of between 8.3 and 15 s− 1 (the reaction rate increased threefold), and the reaction orders were 1 (oxygen partial pressure 0.2–1 atm), 0.41 (thiosulfate concentration 25–200 mol m− 3), 0 (thiosulfate concentration 200–600 mol m− 3), 0.35 (hydronium concentration 1 · 10− 8–1 · 10− 7 mol m− 3) and 0 (hydronium concentration 1 · 10− 6–1 · 10− 2 mol m− 3). The apparent activation energy of the process was 4.5 kJ mol− 1. Kinetics is controlled by a mass transfer of oxygen in the solid–liquid interface. In the presence of copper ions and oxygen, the reaction rate increased by 30% with respect to the process without copper. However, the thiosulfate partially decomposed and a layer of silver sulfide formed on the surface of the silver. The process was applied to a metallic silver powder, which showed a similar behavior to the silver plate.
    Hydrometallurgy 07/2015; 156. DOI:10.1016/j.hydromet.2015.05.009
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    ABSTRACT: A novel environment-friendly method based on stepwise roasting has been proposed to extract vanadium and chromium separately from vanadium slag with high chromium content (V–Cr slag), which consists of two stages of sodium roasting–water leaching. Fractional sodium roasting–water leaching is firstly conducted to extract vanadium from the V–Cr slag and 87.9% of vanadium is extracted in optimal conditions. Leaching residue of the fractional roasting stage is then roasted secondarily to extract chromium and 96.4% of chromium is extracted from the leaching residue in optimal conditions. Overall extractions of vanadium and chromium are 98.9% and 96.6%, respectively. Evolution mechanisms of V-bearing and Cr-bearing phases were elucidated by XRD, TG–DSC and SEM-EDS. It is elucidated that V(III) existing as V-spinels in the V–Cr slag is oxidized to water-soluble NaVO3 while Cr(III) remains in the form of Cr-spinels or R2O3 phase (R: Cr, Fe) during the fractional roasting; in secondary roasting, Cr(III) existing as Cr-spinels and R2O3 phase in the leaching residue of fractional roasting stage is oxidized to water-soluble Na2CrO4. Vanadium and chromium have thus been extracted separately and thoroughly at the extraction procedure by controlling the roasting extent of V–Cr slag via the proposed stepwise roasting method. The established method provides new insights for comprehensive utilization of complex minerals containing multiple valuable elements.
    Hydrometallurgy 07/2015; 156. DOI:10.1016/j.hydromet.2015.06.003
  • [Show abstract] [Hide abstract]
    ABSTRACT: The concentrated metal scraps, obtained by the mechanical process for waste printed circuit board (WPCB) recycling, were pressed as the anode, which was directly electrolyzed to produce copper powders. The effects of CuSO4·5H2O, NaCl and H2SO4 concentration, current density and electrolysis time on current efficiency and copper powder size were investigated in detail. The results indicated that current efficiency increased rapidly as the increase of CuSO4·5H2O, H2SO4 concentration and current density. The obtained copper powders became finer with the increase of current density and NaCl concentration; and copper powders became coarser with the increase of CuSO4·5H2O concentration. Under the optimum conditions (50 g/L CuSO4·5H2O, 40 g/L NaCl, 118 g/L H2SO4, 80 mA/cm2 for 3 h), current efficiency was the highest (98.12%) and the particle size was the finest (9.35 μm). The experimental studies also revealed that the purity of copper obtained at the highest current efficiency was higher to 98.06%. Produced copper powders were characterized by X-ray diffraction (XRD), scanning electron microscopy (SEM) and transmission electron microscopy (TEM). XRD analysis indicated that copper was the main phase; SEM analysis showed that dendritic structure was the main characteristic of the obtained copper powders; TEM results indicated that at the end of the dendritic structure, the copper surface was coated with a layer of dense Cu2O. It is believed that the process is effective for copper recovery from concentrated WPCB metal scraps.
    Hydrometallurgy 07/2015; 156. DOI:10.1016/j.hydromet.2015.06.006
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    ABSTRACT: Due to their large variety of applications, their low supply and high demand, the rare earth elements (REEs) are presently viewed as being among the most critical chemical elements. Because of this, their potential recovery from end-of-life waste products has been extensively discussed both in society and in the scientific literature. This concept of recovering elements contained in end-of-life products, known as urban mining, is regarded as an important step in achieving a sustainable, circular society. This review article discusses the perspectives of reclaiming the REEs from various waste streams using hydrometallurgical methods. Three main streams are discussed in detail (phosphor-containing products, NiMH batteries and permanent magnets), touching on the state-of-the-art of material pre-treatment, leaching and separation of REEs and refining. Comparisons with the extraction of REEs from ores are drawn, bringing forth both the advantages and some of the disadvantages of urban mining.
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.06.007
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    ABSTRACT: The present study investigated the removal of calcium from sulfate solutions containing magnesium and nickel using aqueous two phase systems (ATPS) at room temperature. In an attempt to evaluate the application range of this separation method, batch scale tests (bottom/top phase weight = 1) were carried out using synthetic test solutions at two concentration levels (low and high). The test solution of high concentration used as the bottom phase consisted of a solution containing [Ca] = 0.44 g·L− 1, [Mg] = 1.42 g·L− 1 and [Ni] = 80.0 g·L− 1, whose concentration level is similar to those typically found in high pressure acid leach liquors. The test solution of low concentration used as the bottom phase consisted of a solution of sodium tartrate in which a given volume of the solution of high concentration was added, resulting in a dilution of approximately 80 fold. The performance of some cationic extractants (Cyanex 272, 1N2N and PAN) at changing concentrations and the effect of the pH of the aqueous system (1, 6 and 11) were investigated in the dilute condition. This study found that, when ATPS operated in at least 3 contact stages, it proved to be a highly efficient and selective method to separate calcium from magnesium and nickel, using no extractant and at low pH values (around 1–2) for both dilute and concentrated solutions. The extraction of nickel depends on the pH and the type/concentration of extractant, while the extraction of magnesium was not affected by the studied operating variables.
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.06.010
  • [Show abstract] [Hide abstract]
    ABSTRACT: A three factor 2-level designed set of experiments was performed to determine the effects of inlet flow rate, temperature, and current density on impurity particle behavior in electrolyte and the associated distribution on the cathode in copper electrorefining. Laser-Induced Breakdown Spectroscopy (LIBS) was used to measure the concentration of impurities on the cathode. The results of the experiments were statistically analyzed using Minitab. The inlet flow rate was identified as the most significant factor. All three factors and their two-factor interactions have a significant effect on impurity concentration on the cathode. The impurity concentrations in corner positions of cathodes had higher impurity levels than those in the center position of cathodes. The current density exerts more effect on both impurity concentrations at corner positions than at the center position. A possible explanation for the phenomena observed is proposed.
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.06.005
  • [Show abstract] [Hide abstract]
    ABSTRACT: Rhenium (Re) was recovered as high purity ammonium perrhenate (APR) from a scrub liquor produced by a pilot plant treating 2,000 m3/h of a fume discharged from a molybdenite roaster. The hydrometallurgical process adopted incorporates first a neutralisation stage using lime to remove most of Mo, all Cu, As and other minor contaminants to produce a liquor containing 260–280 mg/L Re and 80–90 mg/L Mo. Solvent extraction was subsequently used to first load the Re and Mo into a 10 v/v.% tertiary amine (Alamine 304–1), 10 v/v.% Isodecanol in 80% Anysol-150. The organic phase which contained typically 6.2 g/L Re and 1.29 g/L Mo was stripped with 30% ammonium hydroxide to yield a liquor containing 29.9 g/L Re and 5.63 g/L Mo. High purity APR (> 99.8% purity) was recovered from this liquor by adding sulphuric acid to pH 6.8. The whole process would provide an efficient method to recover at least 85% of Re produced from the roaster, with some loss probably incurred due to the re-condensation and/or re-crystallisation of gaseous Re oxide in the cool part of the fume extraction circuit.
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.06.008
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    ABSTRACT: The feasibility and optimization of strategies of microbial re-inoculation were investigated to enhance chalcopyrite bioleaching. The application of microbial re-inoculation consisted of the re-addition of cultures of Acidithiobacillus caldus, Ferroplasma thermophilum or Leptospirillum ferriphilum (with different inoculum concentrations) into the two defined different bioleaching systems that separately represented the early and middle stages of chalcopyrite bioleaching. Changes in the bioleaching performance and microbial community structure after re-inoculation were compared to that in the control bioleaching experiment which without re-inoculation. Results of re-addition of pre-grown microbial cultures into the early stage indicated that the re-inoculated strain survived and/or grew in the leaching environments and meanwhile synergistically promoted the growth of other microorganisms, then accelerated the total iron/sulfur oxidation compared to the unamended control. Finally, all these factors resulted in a significant enhancement of copper extraction. Moreover, a higher re-inoculation concentration exerted a more significant improvement in copper leaching from chalcopyrite. In all experiments, re-inoculation with L. ferriphilum into the early stage showed the best enhancement in copper leaching, which significantly shortened the incubation time and improved the maximum copper extraction. Re-inoculation with iron or sulphur oxidisers into the middle stage exerted poor outcomes which could not or slightly improve the final leaching level of copper. These results demonstrate that re-inoculation can be a useful step to improve the bioleaching kinetics and level of chalcopyrite; however its efficacy is influenced by the functional strain selection and procedures of culture re-inoculation (including re-addition time and inoculum concentration).
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.06.009
  • [Show abstract] [Hide abstract]
    ABSTRACT: We investigated the separation of Mn with reduced extraction behavior of Co by PC88A/Versatic 10 acid from a leaching solution of spent lithium-ion batteries containing 11.4 g/L Co, 11.7 g/L Mn, 12.2 g/L Ni, and 5.3 g/L Li. The distribution coefficient and separation factor of Mn and Co were determined at different pH. The slope analysis method was used to determine the extraction mechanism in the PC88A/Versatic 10 acid system. A McCabe--Thiele diagram and countercurrent simulation batch extraction were also investigated to determine at what stage Mn is extracted. The results show a decrease in distribution coefficient of Co and Mn, however, the separation factor of Mn over Co increased. This is an example of the antagonistic effect of Versatic 10 acid addition to PC88A. Using this effect, Mn can be recovered with an organic/aqueous ratio of 1 in the fourth stage in a PC88A/Versatic 10 acid system.
    Hydrometallurgy 06/2015; 156. DOI:10.1016/j.hydromet.2015.06.002
  • [Show abstract] [Hide abstract]
    ABSTRACT: In order to recover the rare earths elements from the diluted source, a membrane dispersion micro-extractor was used to extract the Pr (III) in chloride acidic solution. The saponified 2-ethylhexyl phosphoric acid-2-ethylhexyl ester was used as extractant. Firstly, the dispersion of droplets was analyzed by using a photo imaging system. The results showed that Sauter diameter of the droplets increased with the growth of the aqueous acidity due to a high interfacial tension. However, a fixed phase ratio of 1:1 with a total flow rate increasing reduced the droplet diameter significantly. Secondly, the effects of the concentration of the extractant, aqueous pH, and total flow rate on the rare earths extraction were analyzed. At optimal conditions, the extraction efficiency can reach almost 100% in 2 s. Thirdly, by evaluating the mass transfer resistance, the overall mass transfer process was controlled by reaction process as the total flow rate over 80 mL/min and by mass transfer in organic phase as the total flow rate less than 80 mL/min. Finally, by using the membrane dispersion micro-extractor, a 90% stripping efficiency can be reached in less than 4 seconds as acidity of stripping phase is more than 0.5 M.
    Hydrometallurgy 06/2015; 156. DOI:10.1016/j.hydromet.2015.06.004
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    ABSTRACT: Manganese oxide ore is one of the most important resources. Owing to the depletion of high-grade ores, attention has turned to low-grade ones containing multiple elements such as iron, silicon, and aluminum. Conventional processes for extracting manganese are characterized by high production costs, intensive energy consumption, heavy environmental issues, or high co-leaching of impurities. In this study, selective sulfation roasting–water leaching is proposed for recovering manganese from iron rich low-grade manganese oxide ores using SO2 as reductant. Thermodynamic analysis indicated that manganese dioxide is readily transformed to sulfate. However, the sulfation of ferric oxide only occurs in the presence of both SO2 and O2. The thermodynamic stability region for MnSO4 and Fe2O3 demonstrated selective sulfation of manganese dioxide is feasible. The experimental validation for sulfation roasting and water leaching revealed that 90.6% of manganese and only 3.5% of iron were extracted when sulfation roasting was conducted at 500 °C for 60 min with SO2 partial pressure (SO2 / (SO2 + N2)) of 0.5%–1.0% (Vol.), and the leaching process was performed at 50 °C for 15 min with liquid-to-solid ratio of 5. This process is able to recover manganese from various low-grade manganese oxide ores.
    Hydrometallurgy 06/2015; DOI:10.1016/j.hydromet.2015.05.017
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    ABSTRACT: In the cyanidation of copper-gold ores, the presence of copper minerals can lead to soluble gold losses, the production of weak acid dissociable cyanide (WAD), as well as a number of operational challenges in CIP/CIL circuits with regards to competitive adsorption, and subsequent difficulties associated with elution, electrowinning and smelting. In addition, copper minerals are significant cyanide consumers, leading to higher cost in ore treatment. This paper presents a process to enhance the dissolution of gold using copper-cyanide solutions in the presence of glycine where the solution is cyanide starved. The effect of glycine addition on gold leaching kinetics in copper-cyanide solutions under different leaching conditions was studied. The results show that, in the presence of glycine, gold dissolution rate increases significantly in solutions containing copper-cyanide species at a very low, or zero, free cyanide concentration. In the presence and absence of glycine, gold dissolution rates in solutions containing 10 mM Cu(CN)32 − were 11.1 μmol/m2.s and 0.65 μmol/m2.s, respectively. It is shown that the average gold dissolution rate in a Cu-CN−-glycine system is about 6.5 times higher than the gold dissolution rate in the conventional cyanidation, the presence of cyanide being similar in each system. Kinetic and electrochemical studies were conducted to evaluate the effects of glycine concentration, pH, CN/Cu ratio, and initial copper concentrations on the dissolution of gold. It has also been shown that the gold dissolution rate increases by increasing the glycine concentration up to 1 g/L and any further glycine addition has a negative effect on the dissolution of gold. Increasing leaching pH up to 12 enhances gold dissolution in the copper-cyanide-glycine solutions. However, increasing leaching pH to 13 has an adverse effect on the dissolution of gold.
    Hydrometallurgy 06/2015; 156. DOI:10.1016/j.hydromet.2015.05.012
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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: 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 surface chemical information of massive chalcopyrite electrode during electrochemical oxidation was studied by SXPS, NEXAFS and Raman spectroscopy. The electrochemical studies show that there was an activated region for chalcopyrite anodic dissolution between 550 and 630 mV (vs. Ag/AgCl), accompanied by two passive regions nearby. The spectroscopic studies suggest a thin film of non-stoichiometric sulfur-rich layer formed in the first passive region, likely to be responsible for the passivation. In the active region, S22 − species and covellite were also found, which could be the cause of the potential surge. When the potential was increased to 650 mV, another passive region appeared. At the same time, S22 − and covellite started to dissolve, leaving a highly metal deficient polysulfide and elemental sulfur layer on the chalcopyrite surface.
    Hydrometallurgy 05/2015; 156. DOI:10.1016/j.hydromet.2015.05.011
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    ABSTRACT: The hydrothermal decomposition of K-feldspar in mixed alkaline solutions containing NaOH and KOH at a concentration of 8.5 M was studied. As the mole fraction of KOH (xKOH = KOH/ (KOH + NaOH) in mole) in the mixed alkaline solution decreased, metastable kalsilite, kalsilite and hydroxyl-cancrinite crystallized progressively at 240 °C. The metastable kalsilite disappeared when reaction temperature increased from 240 °C to 280 °C, while the crystallization zone of kalsilite was extended from xKOH = 0.8-0.3 to xKOH = 1.0-0.4, and the crystallization zone of hydroxyl-cancrinite was enlarged from xKOH = 0 to xKOH = 0.2-0. The K-feldspar dissolved under the action of OH− and released K+, Al(OH) 4− and H2SiO42 −. The kalsilite or hydroxyl-cancrinite was precipitated from solution.
    Hydrometallurgy 05/2015; 156. DOI:10.1016/j.hydromet.2015.05.014