Hydrometallurgy最新文献

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.

Chalcopyrite dissolution in ferric sulfate media at ambient conditions is slow. The addition of iodide has been found to significantly improve chalcopyrite leaching in the potential range where triiodide (I₃⁻) or diiodine (I₂) is the predominant species. In the iodine-assisted chalcopyrite leaching process, elemental sulfur was proposed to be the form of sulfur product while iron precipitation was also observed. Given the potential impact of elemental sulfur and iron precipitates on chalcopyrite leaching, this study analyzed the solid leach residues collected from the iodine-assisted chalcopyrite leaching using different solid characterization techniques. The X-ray diffraction analysis of the bulk residues shows that elemental sulfur was the product of chalcopyrite leaching and that pyrite was unreactive. The cross-sectional analysis by a mineral liberation analyzer (MLA) shows that thick layers of jarosite and elemental sulfur coated the surfaces of the solid particles. Further analysis by X-Ray photoelectron spectroscopy (XPS) with a depth resolution of <10 nm confirmed that elemental sulfur was formed and that the formation of jarosite was favored at 40 and 45 °C. Despite the formation of elemental sulfur and iron precipitates, a near complete dissolution of chalcopyrite could be achieved. Further research is required to understand how the presence of iodine changes chalcopyrite leaching in such a way that surface passivation does not occur.

Jezechake Salt Lake brine in China is rich in lithium ions coexisting with relatively high concentrations of carbonate and borate anions. Both carbonate and borate anions have a pH-buffering ability to accept the exchanged H⁺ ions of lithium ion sieves, enhancing the adsorption process of Li⁺ ions. When Jezechake Salt Lake brine passes through a titanium-type ion-sieve-packed column, good lithium breakthrough occurs owing to the pH-buffering action in the brine, which is of great significance for effective lithium recovery from brines in industrial production. In addition, the boron content of Jezechake Salt Lake brine is high, and its recovery is valuable for industrial applications. Here, an N-methylglucamine resin-packed column was connected in series behind a titanium-type ion-sieve-packed column to obtain boron simultaneously with the aim of lowering the energy consumption and cost of lithium recovery. During the adsorption process, Jezechake Salt Lake brine was successively fed into the lithium and boron adsorption columns. The two packed columns were then desorbed independently through acidic solution washing to obtain the Li and B eluents. The breakthrough behavior and elution performance of Li and B in the packed columns were investigated experimentally, and the practicability and efficacy of the proposed process were assessed.

The changes in the micro-scale pore structure during the acid leaching of sandstone uranium ore significantly affect its reactive transport parameters, which, in turn, directly influence the leaching efficiency of uranium. Consequently, understanding and managing the dynamic evolution of the pore structure of sandstone at micro-scale during acid leaching is beneficial for enhancing the leaching efficiency of uranium. This paper employs a dynamic continuous water-rock reaction experiment on sandstone uranium ore to simulate the in-situ acid leaching process. Additionally, X-ray micro-computed tomography (μCT) scanning was utilized to produce three-dimensional (3D) images at various leaching stages. Subsequently, image processing techniques were employed to characterize and parameterize the pore structure within these images. The results revealed that, during the leaching process, mineral dissolution led to an increase in the interconnected pores, while there was a decrease in the isolated pores. Nonetheless, at the leading edge of the sandstone uranium ore, detachment of mineral grains occurred during the leaching process along the direction of fluid flow. The migration of mineral grains resulted in a reduction of interconnected pores and an increase in the isolated pores. The analysis of 3D images indicated that porosity, permeability, the percentage of connected pore area and the reactive surface area exhibited similar variation trends throughout the leaching process. Furthermore, the permeability, the percentage of connected pore area and the reactive surface area displayed a positive correlation with porosity. The study holds valuable insights to develop a deeper understanding of the evolutionary patterns regarding the pore structure during the in-situ acid leaching of sandstone uranium ore.

Owing to the characteristics of mineral paragenesis (equilibrium sequence of mineral phases), the hydrochloric acid leach liquors of Sn ores and Sn-containing waste materials typically contain large amounts of Pb. In this study, barium sulfate coprecipitation was employed to efficiently remove Pb from a tin chloride solution. The lead removal efficiency, reaction mechanism, and limitations were investigated by varying the Cl⁻ and SO₄²⁻ concentrations. The results showed that the structural formula of the precipitates was (Ba_xPb_y)(SO₄)₂ (x + y = 2), and its solubility product was in the range of 1 × 10⁻¹²–1 × 10⁻¹³, indicating that Pb²⁺ and Ba²⁺ were more easily precipitated in the form of complex salts. In addition, thermodynamic analysis revealed that when the Cl⁻ concentration was <1 mol/L in the Pb²⁺–Sn²⁺–Cl⁻–H₂O system, [Pb²⁺] and [PbCl⁺] were the primary species. These positively charged species were readily adsorbed onto the surface of BaSO₄ via electrostatic forces, resulting in a significant increase in the Pb content of (Ba_xPb_y)(SO₄)₂. Therefore, the Pb removal efficiency significantly improved as Cl⁻ concentration decreased. The lead removal mechanism involves a combination of adsorption and replacement processes. Based on the aforementioned fundamental study, a method involving dechlorination and coprecipitation was proposed to remove lead from an acid leaching solution of Kaldo slag. With the dechlorination via distillation, the Pb removal efficiency could reach 99.9% at 1 mol/L Cl⁻ and 0.5 mol/L SO₄²⁻.

This study reports the effect of sodium fluoride on the dissolution kinetics of pure metallic titanium in a citric acid solution. The effects of sodium fluoride concentration, citric acid concentration, temperature, stirring speed, and disc surface area were examined. The dissolution rate of titanium increases strongly with increasing citric acid and sodium fluoride concentrations. Fluoride ions react with hydrogen ions to form hydrofluoric acid, which removes the passive layer of titanium dioxide. An increase in the concentration of NaF and acid causes the formation of a brown layer on the surface of titanium. X-ray diffraction and SEM-EDX analyses showed that the layer composition is mostly of titanium fluoride (TiF₃) and titanium fluoride oxide (TiOF₂). A mixed kinetic model with an activation energy of 26.4 kJ/mol can be used to explain the reaction kinetics.