Hydrometallurgy最新文献

Bauxite residue (BR) is a by-product of bauxite refining using Bayer process for alumina production, which contains numerous valuable components: iron (Fe), aluminum (Al), and rare earth elements such as scandium (Sc). The main issue of previous research for the Sc extraction from BR was the low extraction efficiency or high Fe co-extraction. This article shows that Fe co-extraction can be significantly reduced after reductive leaching of BR in the presence of Fe(II). This effect was achieved due to the formation of magnetite (Fe₃O₄) after reductive leaching of bauxite. Magnetite is practically insoluble in H₂SO₄ at pH > 2. The effect of temperature (T), time (t) and liquid to solid ratio (L:S ratio) on the leaching process was revealed. The results show that under optimum conditions (pH = 2, T = 90 °C, t = 2 h, L:S ratio = 10:1), Sc extraction can reach 57%, with a simultaneous co-extraction of Fe <1.3%. At the same parameters about 10% of Fe is extracted from the raw BR. The Fe content in the residue after leaching increased from 58.3 to 66.2 wt%, corresponding to the iron concentrate. Thus, the Sc extraction from a modified BR by dilute H₂SO₄ creates the conditions for the complete valorization of this technogenic waste, and improved extraction selectivity.

The effective and environmentally friendly recovery of gold from secondary resources has recently received increased attention. Hydrophobic deep eutectic solvent (HDES) composed of N,N-di-butylacetamide and 1-decanoic acid was used to construct the liquid-liquid extraction system to extract Au(III) from HCl medium. At first, DSC, TGA, FT-IR, and ¹H NMR were used to describe the evidence of internal hydrogen bond interaction. The above evidence was verified in the molecular dynamics study of HDES. Then, a series of optimization experiments were carried out to investigate the extraction behaviors of HDES. Based on the results of optimization experiments, response surface method was used to explore the best experimental conditions and the results showed that HDES extracted 99% Au(III) under 0.56 mol/L HCl, 2.19 mol/L NaCl, and 5.83 mmol/L Au(III), and the maximum loading reached 77.0 mg/g. Further studies on the kinetics and thermodynamics of the extraction were used to propose an ion association mechanism. In addition, using 0.1 mol/L thiourea as the stripping agent, HDES displayed outstanding Au(III) selectivity and maintained the extraction efficiency in tests which involved 5-cycles. Overall, this study provides a green and efficient Au(III) recovery strategy for secondary resources.

High efficiency extractants play a necessary role in the selective removal of uranium from rare earth minerals. In this paper, a novel cyclic bidentate diamine extractant, 1, 4-dioctyl-1, 4-dihydroquinoxaline-2, 3-dione (DODQ), was synthesized, and the complexation and extraction of UO₂²⁺ in nitric acid medium were systematically studied. The results indicate that DODQ exhibits a unique combination of properties in removing UO₂²⁺, with an extraction capacity 132 times higher than that of DODPhOA (acyclic structure). Extraction isotherm experiments and slope analyses illustrate the formation of a 1:1 DODQ/UO₂²⁺ regime during the extraction process. Electrospray mass spectrometry, fluorescence spectroscopy, and X-ray crystal diffraction analyses further corroborated its stoichiometric properties. In addition, DODQ can remove UO₂²⁺ from rare earth solutions containing UO₂²⁺ with high selectivity. This study provides an alternative method to separate uranium from the leaching solution of rare earth ore and has a new understanding of the coordination and complexing behavior of UO₂²⁺ with bidentate oxygen-containing ligands.

Mild leaching with HNO₃ acid can be suggested as a potential pre-treatment technique for removing most impurities found in ilmenite which has not been properly investigated yet. This study examines the mild HNO₃ leaching behavior of ball-milled ilmenite as a novel pre-treatment method to produce high-purity synthetic rutile. Ilmenite was mechanically activated for 1 h at 500 rpm which was then reacted with 1 M HNO₃ acid solution in the solid-to-liquid ratio (S/L) of 1:5 (w/v) at 90 °C, 300 rpm, for 2 h to obtain pretreated (ball-milled and ball-milled-mild-leached) ilmenite. The ball-milled ilmenite and ball-milled-mild-leached ilmenite were then reacted with varying concentrations of H₂SO₄, HCl, and KOH solutions at predetermined conditions. Titanium and iron extraction, TiO₂ purity, phase transformations, surface morphologies, and relative elemental abundances were determined by atomic absorption spectroscopy (AAS), X-ray fluorescence (XRF), X-ray diffraction (XRD), and scanning electron microscope equipped with an energy-dispersive X-ray spectroscopy (SEM/EDX), respectively. The initial mild leaching process removed V₂O₅, MgO, MnO, CaO, and Al₂O₃ impurities. The crystallinity indices of ball-milled ilmenite (39.7%) and ball-milled-mild-leached ilmenite (56.7%) indicate the effect of ball milling and mild leaching, respectively. The maximum Ti extraction of ball-milled-mild-leached ilmenite obtained from the sulfate method, chloride method, and alkaline method were 62.1%, 13.2%, and 41.8%, respectively. However, the Ti extraction of ball-milled-mild-leached ilmenite is relatively low compared to ball-milled ilmenite based on percentage extractions. Therefore, the HNO₃ mild leaching process is important to remove impurities, and mechanical activation is more suitable for efficient Ti extraction than a combined step with leaching.

There are significant technical challenges/shortcomings in recovering gold from thiosulphate leach solutions. This study presents a new approach for the recovery of gold from ammoniacal thiosulphate leach solutions using an environmentally friendly organo-sulphide reagent, trimercapto-s-triazine (TMT; C₃N₃S₃³⁻). Precipitation of gold by TMT was investigated using synthetic gold and multi-metal solutions (Au-Ag, Au-Cu, Au-Ag-Cu) to investigate the effect of the co-presence of metals commonly present in thiosulphate leach solution. The effect of TMT concentration (by up to 115 mM), temperature (25–60 °C) and ammonia (0.25–1 M NH₃) on the precipitation of gold was also investigated. The results showed that gold precipitation with TMT was relatively low (i.e., limited to only 37.7% over 15 min) from the solution containing only gold. However, the presence of silver and/or copper significantly improved the precipitation of gold. Silver concentrations at ≥20 mg/L led to a complete recovery of gold in 15 min, producing stable gold precipitates. The results have also indicated that silver readily precipitates when compared to gold. The effectiveness of the TMT precipitation process was also tested on a real leach solution from which complete recovery of gold and silver was readily achieved. The findings demonstrated that gold could be readily but unselectively recovered by precipitation with TMT from ammoniacal thiosulphate leach solutions containing silver, in particular, at the expense of co-precipitation of other metals such as copper and zinc.

The oxidation roasting pretreatment process, which is a mature and dominant industrial practice for treating complex polymetallic gold deposits, has a low recovery of associated silver and exhibits an unclear reaction sequence. This study takes a silver-containing gold ore as the research object and investigates the effect of (i) sodium sulfite as an additive, (ii) main gold-bearing mineral pyrite, and (iii) gangue mineral quartz, on silver during the roasting process in roasting and leaching tests. Characterization techniques included scanning electron microscopy–energy dispersive spectroscopy analysis, and thermogravimetry–differential scanning calorimetry analysis. The results based on thermodynamic calculations and characterization show that the main reason for the low extraction of silver is that the calcined product (i.e., iron silicate) formed by the reaction between pyrite and quartz during the roasting process is mutually absorbed and wrapped with silver oxide. In addition, silver minerals decompose some iron sulfides and iron oxides under high-temperature conditions, promoting the reaction of iron and quartz to form iron silicate while strengthening the secondary encapsulation of silver by products of roasting, thus deteriorating the extraction of silver. The addition of sodium sulfite reduces the formation of iron silicate and increases the extraction of silver by directly promoting the conversion of pyrite to iron oxide. This study provides a theoretical basis for further improving the roasting and leaching of silver.

In the present study, a method for cobalt recovery from Zinc Plant Residue (ZPR) has been developed. First, the residue was leached in a dilute sulfuric acid solution. >90% cobalt and 95% zinc were leached in two hours at room temperature. Then, leach liquor was purified before cobalt precipitation. First, copper and cadmium were removed using zinc dust. >99% copper and 80% cadmium were removed in 30 min at room temperature with slow agitation. Then, dissolved iron was removed using sodium persulfate as an oxidizing agent, and lime was used to maintain pH. Iron was precipitated (∼99%) in 1 h at 60 °C temperature by maintaining pH at 3.5 and Eh at 700 mV. After iron removal, cobalt was precipitated by using sodium persulfate as an oxidizing agent, and sodium hydroxide was used to maintain pH. The cobalt precipitate was calcined and then the final product was obtained with 58% Co. The filtrate, obtained after cobalt precipitation, had a zinc concentration of approximately 69 g/L. About 82% cobalt and 90% zinc were recovered in the entire process.

The use of hydroxylammonium chloride as a reducing agent is proposed to enhance the acid leaching of cathode material from spent lithium-ion batteries. The current study was conducted using real waste from scooter batteries. The main metals found in this cathode material are Mn (43.6%), Li (4.2%), Ni (8.3%) and Co (2.6%), expressed as % w/w. The effect of the initial concentrations of the extracting agent (HCl) and the reducing agent (NH₃OHCl) on the extraction yields of the target metals is investigated. The presence of NH₃OHCl in HCl solutions exerts a highly effective influence during the initial 15 min of the leaching process, with a complete solubilization of Mn within that timeframe, in contrast to the 20% achieved in its absence. During that period, over 70% of Li is solubilized, while Ni and Co reach maximum solubilities of 7% and 5%, respectively. Extending the contact time to 24 h between the extracting solution and LIB waste enables nearly complete extraction of Ni and exceeds 60% for Co. An analysis of variance was used to identify significant factors to be included in multivariable regressions to predict extraction yields. These regressions are used to carry out a preliminary economic analysis of the leaching process based on gross profit. The optimum outcome is achieved when the extraction is conducted through two consecutive leaching processes. In the first process, 100% Mn and 75% of Li are recovered, while the second process recovers the remaining Li, 96% of Ni, and 60% of cobalt. Additionally, the stoichiometry of the reduction of manganese(IV) by NH₃OHCl is studied through the correlation between the gas volume released during the leaching processes and the Mn solubilization reached. This reduction proceeds through two parallel reactions, resulting in the production of N₂O and N₂. The first of these reactions predominates, exhibiting an estimated selectivity of 87%.

The development of an efficient recycling route for spent polymer electrolyte membrane (PEM) electrolyzers is essential for the recovery of platinum-group metals (PGMs). Among the other refining processes, solvent extraction is a highly selective technique that can provide high-purity products. Moreover, the extraction behavior of Pt, Ru, and Ir from the PEM electrocatalysts is not studied extensively. In this work, the effect of extraction conditions for the efficient recovery of PGMs from model solutions was investigated. To determine the efficiency of extraction and stripping processes, metal contents in the feed solution, raffinate, and loaded strip solutions were quantified using inductively coupled plasma optical emission spectroscopy (ICP-OES). For achieving high separation efficiencies for Pt and Ru from Ir, the following conditions were found to be optimal: 100 mg/L of each PGM in 2 M HCl, 15 vol.-% Cyanex 923 in diesel, an organic to aqueous volume ratio (O:A) of 1:1, 30 min of stirring time, a temperature of 25 °C, and 400 rpm stirring rate. A short mixing time of 5 min resulted in a high separation factor of Pt/Ir. Stepwise recovery of PGMs from the organic phase was studied. The three-step stripping strategy was proposed to extract PGMs: 1) the separation of Pt and Ir in the aqueous phase using water, 2) the stripping of Ru using a mixture of ascorbic acid and HCl, and 3) the effective stripping of the remaining Pt using water. Furthermore, the separation of PGMs from the leach solution of the spent PEM electrocatalyst with other metal ions (Sb and Sn) was investigated.

There is a vast amount of globally underutilized low-grade mine tailings and leach residues, including those from primary processing of gold. In this research, the target is to recover the remaining gold (10.9 g/t) from weathered refractory iron-rich residue that had previously been subject to autoclave oxidation, subsequent cyanidation in a conventional carbon-in-leach (CIL) circuit as well as storage at tailings area. Chloride leaching has been considered as one of the most promising cyanide-free gold leaching methods and it has shown positive outcomes in treating primary gold ores, concentrates, and flotation tailings. Therefore, in the current study, the iron-rich residue investigated was subjected to chloride leaching combined with simultaneous carbon adsorption. The investigated parameters included leaching time (2–8 h), chloride concentration ([Cl⁻] = 0.2–5 M), type and concentration of oxidant ([Cu²⁺]/[Fe³⁺] = 0.1–1 M), as well as type and concentration of activated carbon (14–25 g/L), whereas S/L ratio (100 g/L), acidity (pH = 1), and temperature (90 °C) were kept constant. Leaching results indicate that up to 40% of the remaining gold could still be recovered from the investigated residue with optimized chloride leaching. According to the results, the most important parameter for gold recovery was the leaching time. Moreover, of the studied oxidants, cupric ions were shown to contribute more to gold recovery when compared to ferric ions (35% vs. 24% at [Cu²⁺]/[Fe³⁺] = 0.1 M). Nevertheless, an increase of cupric concentration from 0.1 M (low-concentrated) to 0.5 M, resulted in only a slight increase in gold recovery (from 36% to 40%), whereas no further improvement in gold recovery was achieved with a 1 M cupric concentration. Two studied activated carbon products showed equal effectiveness in gold adsorption. In-situ carbon adsorption was shown to occur effectively in chloride media, as all dissolved gold could be detected in the activated carbon, and the concentration of remaining gold in the pregnant leach solution was minimal (< 0.02 mg/L). These findings indicate that low-concentrated chloride leaching of leach residues from industrial gold processes can allow an enhanced recovery of gold from previously mined and treated raw materials.