Hydrometallurgy最新文献

There are two well-defined processes for the types of copper ores. One for the processing of copper sulfides and another for the processing of the copper oxides. There are similarities between the two, particularly in the preparation stage of the run of mine (ROM) such as the crushing. However, significant differences exist in how to concentrate the elements of interest. Flotation is the most used concentration process to separate copper sulfide minerals from other minerals using air bubbles. For oxides, after the dissolution of copper, the solvent extraction (SX) process is the preferred path using two immiscible liquids to separate the copper.

A line of investigation for solvent extraction is the use of a bubble coated with solvent to extract the metal of interest from the aqueous solution, some devices have been built and tested for this purpose using different principles to generate a coated bubble swarm. However, those equipments have been tested on laboratory and have not been scaled up to an industry level.

The Hollow Drop (HD) concept was born from the idea of building a device to generate coated bubbles in a continuous swarm that could be scalable to an industry level. In this paper two columns were built and operated: a proof-of-concept column and a scale-up attempt for the extraction of Cu(II) fom an aqueous solution of 2.5 g L⁻¹ using ACORGA® M5640 (25% v/v) in the Kerosene.

The results show that we could generate a bubble swarm and conduct the solvent extraction process at a 97% recovery using our proposed coated bubble generator. However, in our scaled prototype test only a 70% recovery was achieved, which shows that our column is working but the scaling-up needs more investigation regarding the dimensions and flows of the process.

Base metal hydrometallurgy in a chloride medium has considerable advantages, as metals can be recovered by solvent extraction through a neutral complex mechanism without extra reagent addition, leading to ready lixiviant recycling and significantly reduced wastewater discharge. However, the recovery of Ni(II) using this hydrometallurgical method is challenging because of the unavailability of appropriate extraction reagents. In this study, a new reagent, N-2-ethylhexylpyridine-3-formamide (MEH3), was used to efficiently extract Ni(II) from weakly acidic (pH > 2.5) or neutral chloride solutions in the form of neutral complexes without pH adjustment. Some other metal ions can also be extracted; the selectivity of the reagent for each metal decreased in the following order: Cu(II) > Zn(II) > Ni(II) > Fe(III) > Co(II) > Mn(II). The reagent had weak ability to extract other metal ions such as Ca, Al, Cr(III), Mg, and Li. The extraction of Ni(II) was positively correlated with the Cl⁻ concentration in the aqueous solution, indicating that Cl⁻ is a driving force for Ni(II) extraction. The Ni(II) loaded organic phase was stripped with water, and approximately 50 g/L of Ni(II) in the loaded strip liquor was obtained at an O/A ratio of 15:1. The Ni(II) extraction mechanism was further studied using crystal structure analysis, FT-IR spectroscopy, and the maximum loading capacity. The results showed that a Ni(MEH3)₂Cl₂ complex was formed via Ni(II) extraction. In conclusion, the MEH3 system can selectively extract Ni(II) from a chloride solution without consuming alkaline reagents, and no new substances are produced, which is beneficial for the realization of mother liquor recycling and has a positive impact on clean hydrometallurgy.

Numerous Li extraction methods from minerals and e-waste have been reported in the literature. Among them, direct fluorination processes appear to be a viable alternative due to their high lithium extraction efficiencies (>90%) as LiF. However, a drawback is the low water solubility of LiF, which requires acids for its separation and to obtain other commercial lithium salts. An interesting alternative for dissolving salts with low solubility is through the formation of coordination complexes. In this case, aluminum forms highly stable soluble complexes with the F⁻ anion, such as AlF²⁺, AlF₂⁺, AlF₃, AlF₄⁻, AlF₅²⁻, AlF₆³⁻.

This study proposes an acid-free LiF dissolution methodology using aluminum sulfate as a leaching agent. The LiF dissolution was modeled and optimized using Response Surface Methodology (RSM). The investigated operating parameters for LiF dissolution were the solid/liquid ratio (A), reaction temperature (B), and leaching time (C). Thus, a predictive mathematical model was successfully optimized (R² = 0.9445). The results indicated that the S/L ratio negatively influences the dissolution of LiF, while temperature and time have a positive effect. The LiF dissolutions of 90 ± 3% were achieved with a leaching time of 31 min, a S/L ratio of 20 g/mL, and a temperature of 45 °C.

Cyanidation, a conventional process to extract gold from gold ores, has been used for over 130 years in industrial mining because of the high efficiency and rate of formation of Au(CN)₂⁻ and the high recovery efficiency by adsorption of Au(CN)₂⁻ on activated carbon. However, carbonaceous refractory gold ores are not targeted because Au(CN)₂⁻ is easily adsorbed on carbonaceous matter in the ores, resulting in high recovery loss. In this study, the flotation concentrates of a carbonaceous refractory gold ores was subjected to biooxidation at 45 °C using a mixed culture containing iron-oxidizing and sulfur-oxidizing bacteria, followed by gold extraction using thiourea under strongly acidic conditions. The gold extraction efficiency reached ∼100% in 12 h without re-adsorption. Finally, the quantitative recovery of the Au(CS(NH₂)₂)₂⁺ complex was confirmed by adsorption on strongly cationic exchange resin. Biooxidation reduced the amount of Fe-containing metal sulfides, which minimized the decomposition of thiourea, and the Au(CS(NH₂)₂)₂⁺ complex had a low affinity toward carbonaceous matter, different to Au(CN)₂⁻. Since the process described in this study does not require roasting to remove carbonaceous materials in pretreatment and does not use cyanide for gold extraction, it is environmentally friendly and should be considered for practical applications in carbonaceous gold ore-producing mines.

The effective recovery and utilization of elemental sulfur in the direct leach residue (DLR) from zinc oxygen pressure leaching poses a significant challenge. This study analyzes the distribution characteristics of sulfur in DLR and determines its solubility in n-decane at various temperatures. Results indicate a gradual increase in sulfur solubility with temperature, reaching a maximum of 6.84 g/100 mL at 150 °C Utilizing the Apelblat model, a fitting equation of lnX = 88.3–7155.9/T − 12.0lnT is derived. Under conditions of 130 °C, a liquid–solid ratio of 8:1, a reaction time of 3 min, and a stirring speed of 300 rpm, 99.2% of sulfur in the residue can be dissolved in n-decane. Additionally, this separation process naturally enriches other valuable elements in the residue. The cooling rate significantly influences sulfur purity, with elemental sulfur forming high-quality crystals exhibiting a positive octahedral rhombic morphology at a cooling rate of 0.018 °C/s. Finally, the dissolution mechanism of sulfur in n-decane involves sulfur complexation, and the reliability of the sulfur solubility model is verified.

This study focuses on the ore characteristics and occurrence status of lithium in lithium-poor clay-type ores by employing activation pretreatment by calcination followed by leaching with tartaric acid. This study investigates the influence of factors such as calcination temperature, calcination time, and leaching temperature on the leaching yield of Li. The findings show the optimal leaching conditions for Li extraction as follows: calcination temperature, 600 °C; calcination time, 5 min; leaching temperature, 100 °C; ore-tartaric acid mass ratio, 5:7; leaching time, 5 h; and ore-water ratio, 1:3 (g/mL), resulting in a leaching yield of Li of 85.0%. According to the results of the three-cycle leaching experiments, the Li concentration in the leach liquor increased from 40.2 mg/L to 125 mg/L, indicating efficient utilization of tartaric acid and successful Li enrichment. Moreover, the XRD, SEM, TG-DSC, and FTIR analyses of the samples reveal that tartaric acid dissociates into C₄H₅O₆⁻ and C₄H₄O₆²⁻, which then form complexes with ions such as Li⁺, Al³⁺, Ca²⁺, and Fe³⁺ that are dissolved during the ore leaching process. With an increase in leaching time, complexes involving Al, Ca, Fe, and tartaric acid radicals result in precipitation, leading to a reduction in the content of these ions in the leach liquor. This increases the selectivity of Li extraction, which is beneficial for the subsequent separation and extraction of Li.

Recovery of cobalt and lithium from end-of-life Li-ion battery wastes have been evaluated in batch and semi-batch leaching systems. In this preliminary study, HCl and H₂O₂ were used as leaching and reducing agents, respectively. The comparison of batch and semi-batch processes was carried out, obtaining an improvement from 40% to 70% in the metal mass extracted (i.e. Co and Li) for semi-batch experiments under the same experimental conditions. Effects of the initial concentration of reducing and leaching agents were evaluated for a semi-batch system in which the acid was continuously fed to maintain a constant pH value. From experimental results, it was found that the concentration of H₂O₂ plays an important role in the leaching process in terms of selectivity. For the experiments carried out using 0.1 M of HCl and 1 M of H₂O₂, the percentage of Li and Co extracted was 90% for a leaching time of 30 min. The double-controlled addition of HCl and H₂O₂ to the semi-batch system allows the reduction of the H₂O₂ concentration to 0.5 M. The optimization of reactants entails not only the decrease of their consumption but also maximize the selectivity of the reactions desired, which represents promising results for the environmental sustainability of the process. Further work will examine the fate of chloride ions in the process.

The manganese production industry produces a large amount of sulfide purification sludge (SPS) every year, representing a hazardous solid waste but also a valuable secondary resource for Ni, Co, and Mn. In this paper, the recovery of Ni, Co, and Mn from the SPS was achieved by (i) leaching with a solution of hydrogen peroxide, (ii) selectively extracting nickel, cobalt and manganese, (iii) solvent extraction for zinc removal, and finally obtained the battery-grade Ni-Co-Mn sulfate solution. During the hydrogen peroxide solution leaching stage, the leaching efficiency of Co, Mn, and Ni reached 98.5%, 98.6%, and 95.6%, respectively. A synergistic extraction system (SES) consisting of decyl 4-picolinate and dinonylnaphthalene sulfonic acid was used to selectively extract Ni, Co, and part of Mn, and the extraction of Ni and Co was >99.8% and 95.5%, respectively. The loaded organic was subjected to four-stage countercurrent scrubbing using a 5 g/L H₂SO₄ solution, resulting in nearly 100% removal of Ca and Mg. After that, 150 g/L H₂SO₄ was used to strip Ni, Co, and Mn from the loaded organic to obtain a crude Ni-Co-Mn sulfate solution. Finally, D2EHPA was utilized for the extraction of impurity Zn from the stripping solution to obtain a battery-grade Ni-Co-Mn sulfate solution with <5 × 10⁻⁴ g/L of Zn. Compared to traditional technology, the novel process not only enables the recovery of valuable Ni, Co, and Mn in SPS but also facilitates their direct preparation into battery-grade nickel‑cobalt‑manganese sulfate solution, which has the advantages of a short process and high added value.

This paper presents a multiscale reactive flow model to simulate in-situ leaching of copper in heterogeneous porous microstructures. The introduced workflow combines fluid flow simulation with advection-diffusion-reaction simulation, both required to model reactive flow. The proposed workflow can include the fluid flow in resolved and unresolved pore structures and utilizes required parameters from molecular simulation (ionic diffusivity) and reaction databases (reaction rate parameters). The modeling approach is validated by comparing results to other open-source codes for a model calcite dissolution on acid injection. This model is applied to copper mining by leaching to analyze the reactive flow through a fractured digital rock model of a subsurface sample. Results are analyzed by tracking the concentration distribution along the pore space structure and calculating the outlet concentration of copper to conform the leaching path. Several sensitivity studies are performed to show the robustness of the modeling framework as well as to investigate the importance of each of these parameters on copper production. The complexity of the model is systematically increased from a single scale surface reaction model, to consider the influence of competitive bulk solution reactions, and finally to include flow through porous media to model multiscale reactive flow. This study shows that a multi-scale flow model with homogeneous bulk and heterogeneous surface reactions is required to accurately model copper leaching.