Hydrometallurgy最新文献

A complex approach to the extraction of gallium from a carbon concentrate (CC), a waste product of the aluminum industry, was considered. Ashing of CC made it possible to remove the main component - carbon and obtain ash, which is an oxide mineral-like compound, from which gallium was leached with solutions of inorganic acids and their mixtures. The maximum degree of leaching (98%) of gallium was achieved with 6 M HCl for ash after ashing the CC at 600 °C. It was shown that Purolite anion exchangers with highly basic tertiary and quaternary ammonium base groups recovered gallium(III) from 2 to 10 M HCl, where the sorption equilibrium was attained after 60 min. The maximum extraction was observed from 6 M HCl. Under optimal conditions, the maximum sorption capacities for gallium(III) were achieved for Purolite A300 (2.7 mmol g⁻¹) and Purolite A500 (2.2 mmol g⁻¹) sorbents. Gallium(III) was quantitatively (99%) eluted from the sorbents with distilled water. Sorption of gallium(III) and its subsequent desorption with water made it possible to separate it from the predominant amounts of transition metal ions: Fe(III), Ni(II), Co(II), Zn(II), V(V) and other cations: Al³⁺, Ca²⁺, Mg²⁺, Na⁺ and K⁺. The addition of sodium hydroxide to the concentration of 200 g L⁻¹ resulted in the cementation of gallium on aluminum gallama (liquid gallium‑aluminum alloy).

The current methods for separation of nickel from cobalt sulphate solutions have drawbacks of long flowsheets, release of large volumes of wastewater and incomplete separation of two metals. A synergistic solvent extraction (SSX) system consisting of dinonylnaphthalene sulfonic acid (DNNSA) and decyl 4-picolinate (4PC) is proposed in this work to extract nickel from cobalt sulphate solutions selectively. The extraction of nickel reached 99.7% after five-stage counter current extraction with an A/O volume ratio of 1/6 at 30 °C, resulting in a large increase of Co/Ni mass ratio to 25,700 in raffinate from 56 in feed. The loaded organic phase can be easily stripped using a dilute H₂SO₄ solution. Application of this process in industry is expected to make an important impact on the production of high-purity cobalt sulphate solutions and to deliver economic benefits.

Hydrometallurgical unit operations are typically used to recover palladium (Pd) from its ores and secondary resources with high selectivity owing to their low energy consumption, cost effectiveness, and volume flexibility. Herein, diluted HNO₃ solutions with added nitrate salts were used to dissolve Pd powders. Moreover, solutions of nitrates with same valency cations (such as HNO₃, LiNO₃, NaNO₃, KNO₃, CsNO₃, and NH₄NO₃) and same period cations (such as NaNO₃, Mg(NO₃)₂, and Al(NO₃)₃) were used to reveal the involved beneficial effects of the nitrates on the Pd dissolution in an environment friendly way with low acidity of the solutions. Among all added alkali metal nitrates, LiNO₃ resulted in the highest Pd dissolution efficiency, which was attributed to the higher dissociation constant of LiNO₃, resulting in a higher concentration of free nitrate and hence a higher oxidation potential of the overall system. The dissolution process was systematically investigated to determine the optimal temperature (353 K), LiNO₃ and HNO₃ concentrations (6 and 1 mol L⁻¹, respectively), stirring speed (500 rpm), and reaction time (5 h). These optimal conditions yielded a dissolution efficiency of 99.6%. Notably, as the reaction proceeded, the Pd powder surfaces corroded to form numerous holes, indicating that internal diffusion control also affected Pd dissolution.

This study investigated two bioleaching strategies for removing heavy metals from three mine tailings fractions generated by flotation processes. On the one hand, bioleaching with microbial consortia of acidophilic mesophiles and moderate thermophiles efficiently extracted Co, Cu, Zn, and As, while the leaching of Pb was facilitated through the use of organic acids produced by a heterotrophic bacterium and a fungus. Approximately 100% Co, 68% Zn, 63% As, and 31% Cu were bioleached with acidophilic mesophiles from the barite tailings (BT) sample after 14 days, whereas for the barite concentrate (BC) sample the results showed about 100% Co, 70% Zn and As, and 45% Cu removal at the same period. The sulfide concentrate (SC) sample underwent bioleaching with both consortia, acidophilic mesophiles and moderate thermophiles over 28 days. Approximately, 67% of Co, 28% of Zn, 56% of As, 28% of Cu, and 6% of Mn were extracted from the sample using mesophiles, whereas the leaching efficiency with the moderate thermophiles was about 72% of Co, 50% of Zn, 28% of As, 36% of Cu, and 5% of Mn in 20 L bioreactors. On the other hand, bioleaching of Pb was explored using the bacterium Gluconobacter oxydans and the fungus Penicillium simplicissimum for the production of gluconic acid and citric acid, respectively. Additionally, besides glucose-based media, glycerol and crystal sugar were tested as alternative and cheaper carbon sources. The metabolic activity of P. simplicissimum allowed a maximum Pb leaching of 39–43% from the BT sample in 28 days in glycerol-based medium, while for the BC sample, the maximum Pb extraction was around 60% in glucose-based medium. A lower extraction of Pb was achieved with G. oxydans for both samples. The maximum extraction of 34% and 39% of Pb was reached within 7 days when glucose was used as the carbon source. Further optimization should address both the enhancement of metals removal and – especially for the organic acid bioleaching – the reduction of costs related to media formulation and fungal biomass production on a larger scale.

Addition curing systems involve two-part silicones which require the mixture of a silicone polymer with a catalyst to initiate the cure. Platinum is the most commonly used metal catalyst for addition curing of silicones by hydrosilylation which involves the crosslinking by the addition reaction of silicon hydride species to unsaturated bonds, mainly CC, but also CO or CN double bonds. After crosslinking of the polymers, the platinum catalyst cannot be recovered but remains in the silicone materials throughout the entire product life. In the end, platinum is disposed of together with the silicones and is thus lost to the value chain. The overall objective of this work was to develop a recycling process for the recovery of platinum from addition-cured silicone elastomers. In the first step, this was achieved by efficient digestion methods and by optimizing the leaching processes for exemplary commercial silicone elastomer products. Two different silicone materials were investigated, both of which were crosslinked with a platinum catalyst. The initial Pt content in the tested samples was 12.6 ± 0.2 mg/kg for a commercial silicone impression material and 6.3 ± 0.5 mg/kg for a silicone baking mold, measured by graphite furnace atomic absorption spectrometry (GF-AAS). Samples were first frozen with liquid nitrogen to improve brittleness and then crushed with a simple food processor to obtain a silicone granule. Various acid mixtures, mainly based on sulfuric acid, were investigated as digestion methods in order to extract platinum from the silicone network. These had different effects on the dissolution behavior of silicone samples and the amount of platinum extracted in each case. The amount of platinum leached from the filtrate of the digested samples in each case was measured by ICP-OES to evaluate the efficiency of different leaching mixtures. In addition, the dissolved platinum species present in the solutions was identified by UV/VIS as tetrachloridoplatinate(II) complex. The best platinum leaching results so far were obtained with two methods, both of which used a leaching mixture based on sulfuric acid and hexamethyldisiloxane (M₂). In the presence of hydrochloric acid, 9.6 ± 1.6 mg platinum/kg was leached from the silicone impression material and 4.2 ± 0.8 mg platinum/kg from the silicone baking mold. With the additional use of aqua regia instead of hydrochloric acid, 10.4 ± 2.8 mg platinum/kg was extracted from the silicone impression material and 4.8 ± 1.0 mg platinum/kg was extracted from the silicone baking mold. These methods were replicated with n = 3. Using statistical evaluation methods (F-test, t-test, and confidence interval), no significant difference was found between these two best methods. Recovery of platinum(0) from leach mixtures has not yet been achieved due to high dilution and very low platinum concentration in samples and will be part of another study.

This study reports the kinetics of copper metal dissolution in citric acid-hydrogen peroxide solution using the rotating disc method. The effect of the concentration of citric acid as well as hydrogen peroxide, stirring speed, disc surface area, and temperature was investigated. The concentration of hydrogen peroxide (0.25–1.5 mol/L) has a strong effect on the dissolution rate (mg/min) of copper with a reaction order of 0.92, compared to a reaction order of 0.42 with respect to the concentration of citric acid (0.125–1.5 mol/L). The reaction rate is directly proportional to the disc area (A) and square root of the disc rotation speed (ω^1/2), as expected from the Levich equation. The activation energy of 33 kJ/mol, in the temperature range 30-60 °C is also comparable with the values reported in previous studies.

Due to its physical properties, metallic silver is present in numerous electronic wastes. Its recycling requires selective extraction, which involves leaching of silver as the first step. This work focusses on leaching silver with Cu(II)/NH₃/S₂O₃²⁻ which has been widely used for gold leaching. As recently shown for the leaching of copper, using foams whose aqueous phase consists of leaching chemicals is a promising way to reduce the environmental footprint, by improving the metal oxidation caused by the fast transfer of O₂ from bubble to bubble. In this work, metallic silver samples are dissolved by foams made of Cu(II)/NH₃/S₂O₃²⁻ solution with bubbles composed of O₂-N₂ mixtures. The main problem of the thiosulfate route is its degradation during metal oxidation hence it requires using large quantities of this reactant. The results obtained for our leaching foams show that the quantity of silver leached per quantity of thiosulfate used is about three times greater in comparison with a solution, which would bring a new approach to this problem. Moreover, we investigate the role of the bubble size and the gas composition (dioxygen partial pressure). Besides we find that the dissolved silver is inhomogeneously distributed between the foam column and the bottom solution, with an accumulation of 90 % of silver inside the foam, hence opening an interesting perspective for an easy separation of silver upon leaching. Comparing several surfactants, we show that only non-ionic surfactant polyoxyethylene oleyl ether, Brij O10, shows satisfying results, while dodecyl trimethyl ammonium chloride most likely binds with silver complexes and triggers a quick collapse of the foams.

The extraction properties of mono-hydroxy alcohols (2-butyl-1-octanol, 2-ethyl-1-hexanol, 1-octanol) and di-hydroxy alcohols (2-ethyl-1,3-hexanediol, 2-butyl-2-ethyl-1,3-propanediol, 2-methyl-2,4-pentanediol) towards boron, potassium, lithium, sodium, calcium and magnesium ions were investigated as a function of pH and phase volume ratio (O/A) by considering an aqueous phase with a composition mimicking the brine from the Salar de Hombre Muerto in Argentina, in the Lithium Triangle. The mono-hydroxy alcohols are soluble in kerosene and sulfonated kerosene whereas the di-hydroxy alcohols are only soluble in mixtures of kerosene and toluene or m-xylene. No significant effect of the diluent on the extraction properties is observed. All alcohols exhibit high selectivity for boron over potassium, lithium, sodium, calcium and magnesium at acidic pH (pH = 1–5.5). A significant decrease of the extraction efficiency of boron by the mono-hydroxy alcohols is observed at pH >5.5 while no significant decrease of the extraction efficiency occurs with di-hydroxy alcohols. Finally, the best conditions to selectively extract boron from the synthetic brine with 1 mol L⁻¹ mono-hydroxy alcohols or di-hydroxy alcohols at 35 °C are pH = 5.5 and O/A = 4 when the diluent is kerosene and O/A = 2 when the diluent is a mixture of kerosene and toluene.

This work proposes the roasting of amblygonite with calcium sulfate followed by water leaching to extract lithium and phosphorus. Various roasting parameters including roasting temperature, the mass ratio CaSO₄/ore, roasting time as well as leaching parameters, such as the liquid/solid ratio, temperature and time, have been carefully investigated. The lithium extraction efficiency can be as high as 99.8% with a mass ratio of CaSO₄/ore of 0.85 after roasting at 775 °C for 1 h and water leaching at a liquid/solid ratio of 3 mL/g at 100 °C for 2 h. The XRD analysis demonstrates that the fluoride in the amblygonite can be converted into Ca₅(PO₄)₃F and CaF₂. Lithium in the leach liquor can be recovered as Li₂CO₃. Meanwhile, the phosphorus in the water leach residue can be recovered as FePO₄ with a high recovery of 96.6% after acid dissolution and precipitation, which can be further treated to produce LiFePO₄.