In this study, the distribution and evaporation principles of impurities in distilled magnesium metal were investigated using a low vacuum (8 × 10⁴ Pa) distillation purification experiment and theoretical analysis. Provided all other factors remain unchanged (distillation time, pressure, and total area of evaporation), the optimum temperature for preparing high-purity Mg (99.99%) is 750 °C. A detailed analysis of purified Mg was obtained by inductively coupled plasma mass spectrometer (ICP-MS) for 10 major impurity elements. After distillation at 750 °C, low vapor pressure impurities in condensed magnesium, including Si, Mn, Al, Fe, Cu, Ni, and Sn were significantly reduced as other impurities were slightly reduced. Our analysis confirmed a decrease in the following impurities: Fe, Si, Mn, Cu, and Al were reduced from 21.8, 78.6, 68.4, 4.4, and 39.4 ppm to 1.2, 9.3, 6.0, 1.0, and 5.4 ppm, respectively; satisfying the 99.99% Mg standard. The evaporation rate and separation coefficient were calculated under experimental conditions. To better describe the distillation process of metallic magnesium under low vacuum conditions, the mean free path is also calculated in this study under actual conditions.

Nickel laterite ore is divided into three layers and the garnierite examined in this study belongs to the third layer. Garnierite is characterized by high magnesium and silicon contents. The main contents of garnierite are silicates, and nickel, iron, and magnesium exist in silicates in the form of lattice exchange. Silicate minerals are difficult to destroy so are suitable for smelting using high-temperature pyrometallurgy. To solve the problem of the large amounts of slag produced and the inability to recycle the magnesium in the traditional pyrometallurgical process, we propose a vacuum carbothermal reduction and magnetic separation process to recover nickel, iron, and magnesium from garnierite, and the behavior of the additive CaF₂ in the reduction process was investigated. Experiments were conducted under pressures ranging from 10 to 50 Pa with different proportions of CaF₂ at different temperatures. The experimental data were obtained by various methods, such as thermogravimetry, differential scanning calorimetry, scanning electron microscopy, energy dispersive spectrometry, X-ray diffraction, and inductively coupled plasma atomic emission spectroscopy. The analysis results indicate that CaF₂ directly reacted with Mg₂SiO₄, MgSiO₃, Ni₂SiO₄, and Fe₂SiO₄, which were isolated from the bearing minerals, to produce low-melting-point compounds (FeF₂, MgF₂, NiF₂, etc.) at 1315 and 1400 K. This promoted the conversion of the raw materials from a solid–solid reaction to a liquid–liquid reaction, accelerating the mass transfer and the heat transfer of Fe–Ni particles, and formed Si–Ni–Fe alloy particles with diameters of approximately of 20 mm. The smelting materials appeared stratified, hindering the reduction of magnesium. The results of the experiments indicate that at 1723 K, the molar ratio of ore/C was 1:1.2, the addition of CaF₂ was 3%, the recovery of Fe and Ni reached 82.97% and 98.21% in the vacuum carbothermal reduction–magnetic separation process, respectively, and the enrichment ratios of Fe and Ni were maximized, reaching 3.18 and 9.35, respectively.

In the process of the comprehensive recovery and utilization of discarded lithium-ion batteries via acid leaching, a large number of NiS and CoS mixed materials are produced. To improve the metal recovery rate, the kinetics and rate-determining step of the oxygen-rich pressurized acid leaching of Ni and Co were investigated. The results showed that the leaching rates of Ni and Co were greater than 99% under the following conditions: sulfuric acid (leaching agent) concentration of 95 g/L, liquid-to-solid ratio of 6:1, leaching time of 120 min, temperature of 110 °C, stirring speed of 400 r/min, 90% oxygen-rich atmosphere, and pressure of 1.2 MPa. The leaching of Ni and Co was described by the shrinking unreacted core model, and the leaching rates of Ni and Co conformed to the kinetic equation 1 − (1 − x)^1/3 = k·t. The apparent activation energies of the Ni and Co leaching reactions were 50.87 and 45.6 kJ/mol, respectively, and the leaching process was found to be controlled by the chemical reaction at the interface.

In this paper, the key factors and the mechanism of the extraction of iron and manganese from the sulfuric acid leaching solution of waste ternary lithium-ion batteries were studied by combining P204 and N235 extractors. The experimental results showed that the optimal organic phase composition was 25% P204 + 15% N235 + 60% sulfonated kerosene, and the optimal pH of the pre-extraction solution was in the range of 3.0–3.5. Under the conditions of an extraction temperature of 25–35 °C and an extraction ratio of O/A = 1/1 and after mixing for 5 min, the removal rate of Mn and Fe exceeded 99%. Under the same extraction conditions, the extraction effect of the P204-N235 composite extractant on Mn and Fe was better than that of P204 alone. logD-log[P204] showed that the formulas of the Mn and Fe extracts were MnR₂(HR) and FeR₂(HR), respectively. The logD-pH diagram showed that only one free H⁺ was released for each metal ion during extraction, and extraction occurred via a cation exchange reaction.

In this study, we investigated the preparation process of lithium metal using a novel carbothermic reduction process. Herein, thermodynamic calculations and experimental research methods were used to investigate the novel carbothermal reduction process of LiAlO₂ for preparing lithium metal in a vacuum. The results suggest that the temperature of the chemical reaction can be significantly reduced by reducing the pressure in the system. When the system pressure was 10 Pa, the reduction temperature, reduction time, molar ratio of C/LiAlO₂, and reduction rate were 1623 K, 120 min, 0.9, and 79.01%, respectively. Furthermore, the lithium content in the reduction residue was 2.1 wt%, suggesting the presence of the LiAl₅O₈ phase. Under these conditions, when 20% CaO was added, the reduction rate of lithium increased to 99.38%, and the lithium in the reduction residue takes the form of CaAl₂O₄. This shows that CaO, as an additive, can promote reduction, compromise the effect of LiAl₅O₈, and increase the reduction rate of lithium.