Hydrometallurgy最新文献

Silver-bearing manganese ore (Mn-Ag ore) is an important resource for the extraction of the precious metal Ag. However, the efficient and economical utilization of low-grade Mn-Ag ore poses challenges due to its polymetallic co-occurrence, complex associated mineral structures, and lower Ag grade compared to an industrial Ag grade of 80 g/t. In contrast, the Ag grade of commercially viable concentrate, achieved through physicochemical beneficiation, generally exceeds 1000 g/t. This work briefly discusses the metallogeny and resource characteristics and the processing technologies of Mn-Ag ore to produce high grade Ag concentrates. Based on the metallogeny characteristics of Mn-Ag deposits, this work delves into the challenges and difficulties in the physical separation of Ag and Mn. By regulating the differences in the chemical properties of Mn and Ag constituents, chemical beneficiation processes, including unit operations of pyrometallurgy and hydrometallurgy, result in favorable Ag and Mn enrichment and separation. The chemical principles, technical parameters, Mn-Ag separation efficiency, and advantages and disadvantages of chemical beneficiation (blast furnace smelting, chlorination roasting, one-step leaching, and two-step leaching) were systematically summarized and discussed. This work can provide theoretical and technical guidance for the effective treatment of low-grade Mn-Ag ores.

This paper reports the solid-liquid phase equilibria of the CoSO₄-H₂O and CoSO₄-H₂SO₄-H₂O systems at low temperatures. Binary and ternary phase diagrams, including the stable solid phases CoSO₄·6H₂O and CoSO₄·7H₂O were established using experimental data and thermodynamic modeling applying the mixed-solvent electrolyte (MSE) model. The results showed that the addition of H₂SO₄ shifts the eutectic temperature and concentration to lower values for cobalt sulfate and ice crystallization. The trends obtained from the experimental data and the modeling are consistent for the binary CoSO₄-H₂O system with good agreement, but the ternary CoSO₄-H₂SO₄-H₂O system shows some deviations. In general, the MSE model is shown to be reliable for inferring and establishing the phase diagram of the low-temperature system. The phase diagrams are helpful for designing the pathways of cooling crystallization and eutectic freeze crystallization and assessing the performance of the low-temperature crystallization process in the production of CoSO₄ hydrates. In addition, some practical examples of cooling crystallization and eutectic freeze crystallization of CoSO₄ solutions are provided.

A microbial method using commercially available baker's yeast was developed for efficiently and selectively collecting aqueous Au(III) ions in aqua regia leachates from waste printed circuit boards (PCBs) through biosorption under air at temperatures ranging from 10 °C to 34 °C. Even when the total concentration of base metals was much higher than the initial Au concentration in the PCB leachate with high acid concentrations of 4.7–5.6 mol/L, commercial dry baker's yeast exhibited an excellent ability to selectively collect aqueous Au ions from the leachate within 60 min. When the biosorption test was repeated as a three-stage batch operation, the percentage biosorption of Au from the PCB leachate increased from 61% to 99% with 36 g/L dry yeast cells at 34 °C. The experimental results for the three-stage batch biosorption test were consistent with theoretical predictions based on the material balance of Au in multistage equilibrium operations and the distribution coefficient of Au. Equilibrium data of the Au biosorption conformed to the linear isotherm, regardless of the yeast concentration and the initial Au concentration in the leachate. The distribution coefficient K_Au at 34 °C decreased by 30%, from 53.7 to 37.3 L/kg-dry cells, as the total base metal concentration in the leachate was increased tenfold, from 1.94 to 16.4–19.1 g/L. Moreover, the distribution coefficient of Au at 10 °C to 34 °C was analyzed to determine thermodynamic parameters according to the van't Hoff equation. The thermodynamic studies indicated that the biosorption of aqueous Au ions by baker's yeast was spontaneous and exothermic.

The residue known as jarosite-alunite (JAR) is produced when the leach solution of spent lithium-ion battery is neutralized. This residue typically consists of Fe, Al, Na₂SO₄, Ni, Co, and Mn. It is classified as both a hazardous solid waste and a secondary resource. A unique hydrometallurgical technique was implemented to recover Na₂SO₄ and use Al extracted from JAR in high value applications. This extraction process involves phase transformation and NaOH leaching, with the pH adjusted in the range from 10.9 to 14. Initially, the JAR compound underwent dissociation to isolate SO₄²⁻ as Na₂SO₄ by means of NaOH at a moderate pH, while the other metals were preserved as a hydroxide residue. Afterwards, aluminum in the hydroxide residue was selectively leached with NaOH leaving Ni, Co, and Mn in the remaining residue. The results indicated that over 93% of Na₂SO₄ and 86% of Al in JAR were effectively recovered as Na₂SO₄ and high value-added γ-AlOOH, respectively. Additionally, the enriched Ni, Co, and Mn in the alkaline leach residue were selectively recovered by H₂SO₄ leaching. The suggested procedure led to a significant decrease in waste by more than 67%, offering a fresh approach to effectively reduce waste and recover metals from JAR.

Direct solvent extraction (DSX) was applied to produce purified concentrated Ni and Co solutions from a synthetic sulfuric liquor containing Ca, Cu, Mg, Mn, and Zn as impurities, which simulates the solution obtained by the sulfation-roasting-leaching process after precipitation of Fe, Al, and Cr. The commercial extractants Versatic 10, Cyanex 272, D2EHPA, TBP were used in 3 solvent extraction circuits. In the first circuit, operated at 40 °C, Versatic 10 (0.5 M) fully extracted Cu, Zn, Ni, and Co in 3 stages at O/A = 2:1, and pH 6.5, leaving remaining Mn (54% was extracted) and most of the Ca and Mg in the raffinate. The co-extracted Ca and Mg were fully scrubbed off the loaded Versatic 10 in 2 stages at O/A = 5:1, and pH 6.5. The other metals were stripped out of the Versatic 10 extract using a synthetic Ni spent electrolyte (60 g/L Ni, 2 M H₂SO₄) in 2 stages at O/A = 9:1. This loaded strip liquor was subjected to a second circuit with Cyanex 272 (0.64 M) operated at 50 °C. Three stages were required to fully extract Co, Cu, Mn, and Zn (O/A = 2:1, pH 4), whereas the raffinate containing 82 g/L Ni was deemed suitable for electrowinning. The co-extracted Ni(II) was fully scrubbed off the loaded Cyanex 272 in only 1 stage at O/A = 5:1, and pH 4. All Co, Cu, Mn, and Zn were stripped out from the scrubbed loaded Cyanex 272 in 2 stages at O/A = 10:1 using a synthetic Co spent electrolyte (45 g/L Co, 1 M H₂SO₄). The raffinate containing 58.4 g/L Co was submitted to a third circuit using a D2EHPA (0.6 M) + TBP (0.73 M) synergistic system operated at 50 °C. Zinc(II) was fully extracted by the D2EHPA + TBP system in 2 stages at pH 2 and O/A = 1:3, while Mn(II) and Cu(II) were fully extracted from the Zn-depleted raffinate in 2 stages at pH 3.5 and O/A = 2:1. The raffinate containing 58.3 g/L Co was deemed suitable for electrowinning. Copper(II), Mn(II), and Zn(II) were stripped out from the loaded D2EHPA + TBP in 3 stages at O/A = 2.5:1 using 1 M H₂SO₄. The real number of moles of extractants involved in the extractions and apparent equilibrium constants were estimated for all circuits. A flowsheet of the purification conceptual route is presented.

Spent catalyst of molybdenum with silicon dioxide, commonly used as the carrier, is an important secondary resource for recovery of molybdenum. This work proposes a process to recycle molybdenum and remove silicon simultaneously by solvent extraction. The sulfuric acid leachate of the spent catalyst was contacted with the trialkylamine N235 (R₃N, R = C₈–C₁₀), to extract molybdenum and silicon. Several vital parameters were investigated to explore the influence on extraction and stripping. The extraction efficiencies of molybdenum and silicon were up to 99.6% and 77.1% after three-stage countercurrent extraction under optimized condition. The extraction reactions were determined by maximum loading capacity and FT-IR. The peak at 802.1 cm⁻¹ was caused by the stretching vibration of Si-O-Si, indicating the co-extraction of silicon. Molybdenum and silicon in the loaded organic phase can be stripped by the mixture of solution containing 7.00 mol/L NH₄OH and 0.80 mol/L (NH₄)₂CO₃, and the stripping efficiencies were >99.0%. Ammonium molybdate was prepared by removing silicon and evaporating, and the purity was 99.9%.

The behavior of calcium lignosulfonate (CLS) in the oxygen pressure acid leaching process of ZnS concentrate was investigated using total organic carbon assessment (TOC), UV–visible spectrophotometry, Fourier transform infrared spectroscopy (FTIR), and gas chromatography-coupled mass spectrometry (GC–MS) as characterization methods. The effect of the CLS degradation products on zinc electrowinning was also discussed. The results showed that the temperature was positively correlated with the degradation of CLS, while the initial acidity had only significant effects in the range of 0–50 g/L and oxygen partial pressure range of 0–0.1 MPa. At an oxygen partial pressure of 0.2 MPa, an acidity of 160 g/L, and a reaction temperature of 150 °C, about 82.4% of CLS was degraded. In the oxygen pressure acid leaching process, CLS underwent polymerization and decomposition reactions, and its aromatic rings and side chain groups were damaged to varying degrees. At 120 °C, CLS was partially converted into sulfonic acids, phenols, and esters of higher molecular weights. At 150 °C, CLS further degraded into lower-molecular-weight aromatic ethers and sulfonic acids with shorter carbon chains. These organics were relatively stable and were the main sources of organic compounds during the oxygen pressure leaching process of zinc concentrates. The addition of CLS had a significant negative impact on zinc electrowinning, which was related to the adsorption of CLS on the cathode surface, enhancing cathodic polarization and inhibiting zinc reduction kinetics.

Secondary aluminum dross (SAD) has been identified as a hazardous waste because it contains refractory AlN, fluoride, and other salts. Considering that Al(OH)₃ produced during the hydrolysis of AlN in SAD hinders spontaneous AlN hydrolysis, which is amplified by the partial embedment of AlN in other oxide phases present in SAD particles, the complete removal of AlN from SAD is difficult. Herein, we propose a new catalytic hydrolytic denitrification process of SAD, which requires a low additive dose. By changing the time at which the additive (NaOH) was added to the SAD slurry, the proportion of nitrogen removed from SAD was increased. The additive (NaOH) reacted with Al(OH)₃, thereby mitigating its hindering effect on spontaneous AlN hydrolysis and facilitating the complete hydrolysis of the semi-encapsulated AlN in SAD particles. Concurrently, the reaction between NaOH and Al₂O₃ and SiO₂ phases present in SAD was mitigated, resulting in the maximal retention of residual materials (phases other than Al resources) in the solid phase. A SAD denitrification efficiency of 99.0% was achieved under optimal processing conditions (T = 95 °C, t = 240 min, Liquid-to-solid ratio = 6:1, stirring speed = 400 rpm, and the addition of 7 wt% NaOH at 0.5 h), and a slightly lower efficiency of 98.2% was achieved when 50% lower amount of NaOH (3 wt%) was used. Thus, approximately 50% less additive is required for nitrogen removal from SAD, relative to that used in traditional catalytic water-washing processes.

Selective leaching behavior of the spent NdFeB magnets was investigated to find out how controlled phase system through the roasting conditions gave an impact on the performance of the selective leaching. The pre-formed Nd₂O₃ by the selective oxidation made it possible to prevent the formation of NdFeO₃ at high roasting temperature. This resulted in the ideal Nd₂O₃ and Fe₂O₃ phase system, which was obtained by roasting at 660 °C, suitable for the selective leaching of Nd. While the roasting at 600 °C was still not enough to get full oxidation, the roasting at 720 °C made the NdFeO₃ formed. These characteristics with the roasting conditions affected leaching behavior. The NdFeB powder sample of 32 μm particle diameter, roasted at 600 °C showed the highest Nd leaching efficiency as well as the highest Fe leaching efficiency, which seemed to be ascribed to the existence of the metallic Fe. The poor leaching efficiency for the large powder size after roasting was ascribed to the poor diffusion characteristic of the H₂SO₄ solution. With the controlled phase system, the 100% leaching efficiency of Nd could be obtained by reducing the particle size to 16 μm. The formation of NdFeO₃ as a co-product, which had resistance to the reaction with the leaching agent, could prevent the dissolution of Nd₂O₃ around the NdFeO₃ layer. Thus, the 720 °C roasting condition showed the lowest leaching efficiency. The co-product Fe₂O₃ also showed good resistance to leaching so its leaching efficiency was just 5–6% after the leaching time of 3–4 h.

Vanadium is a strategic metal with extensive applications in steel production and emerging energy technologies. In vanadium metallurgy, the pivotal steps encompass the roasting of vanadium slag, leaching, and precipitation of vanadium. The roasting process, which involves elements such as sodium, calcium, manganese, and magnesium, facilitates the phase transformation and extraction of vanadium. Considering the phase separation behavior of vanadium-enriched phases (MV₂O₆, MV₂O₇, or MV₂O₈) in various leaching media, including acid, alkali, and water, the wet decomposition of these phases can be classified into two categories: (i) those yielding insoluble M and soluble V and (ii) those resulting in both soluble M and V. Thermodynamically, the reaction equilibrium constants and temperature profiles of the vanadium-rich phases in various acid and alkaline decomposition processes were calculated and juxtaposed. This review also reports the limiting factors of leaching kinetics of vanadium-rich phases in acid and alkaline decomposition processes, particularly the separation and transformation of vanadium-rich phases in calcified vanadium slag. The vanadium precipitation process encompasses a detailed elaboration of the mechanisms behind the precipitation of hydrolyzed vanadium product and ammonium‑vanadium product. Finally, the vanadium slag roasting-leaching‑vanadium precipitation process was evaluated from four aspects: principle, laboratory and plant practice, resource and environment, and cost and benefit.