The new process developed here consisting of sulfurization roasting transformation and water immersion can effectively realize the separation of rare earth elements (REEs) and impurities from spent Nd-Fe-B magnets. For the industrial application of the new process, it is critical to determine how to economically and efficiently prepare rare earth oxide (RE_xO_y) products with higher purity from the obtained water leaching solution. Therefore, according to rare earth sulfate (RE₂(SO₄)₃) solution characteristics, the oxalic acid precipitation–calcination method, sodium carbonate precipitation–calcination method, and double sulfates precipitation–alkali conversion–calcination method were optimized and compared. The results show that the recovery efficiency of REE recovery via the oxalic acid precipitation–calcination method is 99.44%, and the purity of RE_xO_y is 99.83% under optimal technological conditions. However, the cost of oxalic acid precipitation is higher. The process consisting of the double sulfates precipitation–alkali conversion–calcination method is relatively complicated, the recovery efficiency of REEs is 97.95%, and the purity of the RE_xO_y is 98.04%. The recovery efficiency of the REEs and the purity of the RE_xO_y obtained from the sodium carbonate precipitation–calcination method are 99.12% and 98.33%, respectively. Moreover, the recycling cost of sodium carbonate precipitation is the lowest among the three processes for preparing RE_xO_y, so it has industrial application potential. The obtained results for REE recovery from spent Nd-Fe-B magnets in this research can provide theoretical guidance for the innovation of the recycling process for REEs as secondary resources.

The environmental and economic benefits of recycling spent LiFePO₄ batteries are becoming increasingly important. Nevertheless, the reprocessing of this type of material by conventional processes remains a challenge due to the difficulties of Li and Fe separation and low product purity. Herein, a new approach for recovering Li to separate iron and phosphorus from spent LiFePO₄ cathode materials is developed. Selective separation of Li can be achieved by oxidation roasting followed by low-acid pressure leaching. During the oxidation-roasting stage, almost all the stable LiFePO₄ cathode materials were first transformed into Li₃Fe₂(PO₄)₃ and Fe₂O_3, with the most suitable oxidation-roasting temperature determined to be 550 °C. Then, >96% of Li could be extracted using 0.5 mol·L⁻¹ H₂SO₄ with an L/S ratio of 150 g·L⁻¹ at 110 °C for 1 h; in contrast, the leaching of Fe was 0.03%. The mineral-phase composition of the leaching residues mainly includes FePO₄·2H₂O, Fe₂O₃, and C, which can be used as a raw material for preparing battery-grade FePO₄. These findings demonstrate that the recycling process has the advantages of high selectivity for Li, excellent reaction kinetics, low acid consumption, and free oxidizing agent that may benefit the development of a circular economy.

To determine slag properties and the factors influencing these properties for optimization of operating conditions in the copper flash smelting process, the composition and microstructures of the quenched smelting and converting slags have been analyzed. Thermodynamic software FactSage 8.2 has been used to investigate the effects of matte grade, SO₂ partial pressure, and the Fe/SiO₂ ratio on the liquidus temperature and the copper content of the smelting slag. The possibility to recover valuable metals from the smelting and converting slags through pyrometallurgical reduction by carbon is also discussed. It was found that the flash smelting slag temperature is usually higher than its liquidus temperature and the copper (1.2% Cu) is mainly present in the slag as dissolved copper. In the copper flash smelting process, the copper content in the slag can be decreased by decreasing the Fe/SiO₂ ratio and temperature. In pyrometallurgical slag reduction, most Cu, Mo, and Ni can be recovered as an alloy. The conditions of recovery such as the ratio of smelting slag to converting slag, temperature, and reduction extent have been discussed.

The Kaldor furnace smelting process is currently the mainstream process for treating copper anode slime, but the existence of copper, tellurium and other impurities has adverse effects on the recovery of gold and silver during the Kaldor furnace smelting stage. Therefore, it is necessary to pretreat the copper anode slime to remove these impurities before Kaldor furnace reduction smelting. However, the current pretreatment process of copper anode slime generally has the problem of low removal efficiency of copper and tellurium, and little research on the occurrence state of main metals in copper anode slime. Therefore, this study quantitatively determined the phase composition of Cu, Te, Pb, Bi, As, Sb, Se, Ag and Au, and hydrogen peroxide was introduced to enhance the leaching of impurities. The leaching behavior of each metal in copper anode slime was investigated in detail. The results demonstrate that Cu and Te in the copper anode slime mainly exist in the form of CuO and CuSO₄ and Te and AuTe₂, respectively. More than 99% of the Cu and 97% of the Te were leached out using 250 g/L H₂SO₄ and 28.8 g/L H₂O₂ with a leaching pressure of 0.8 MPa at 150 °C for 2 h, while the leaching of Au and Ag was both < 0.03%. The removal of Cu and Te and the enrichment of precious metals were achieved. This study provides a rich theoretical reference for the optimization of the Kaldor furnace process.

The selective separation of lithium from spent ternary positive materials is achieved through hydrogen reduction followed by water leaching. Almost 98% of the Li is transformed into soluble LiOH⋅H₂O, while the Ni, Co and Mn species are all transformed into insoluble metals or their oxides, so the recovery of Ni, Co and Mn at this stage is challenging. The traditional acid leaching process has drawbacks such as high oxidant consumption, the low recovery of valuable metals and high production costs. Thus, sulfation roasting followed by water leaching was studied in this project. The leaching levels of Ni, Co, Mn and Al reached 87.13%, 99.87%, 96.21% and 94.95%, respectively, with 1.4 times the theoretical amount of sulfuric acid used at 180 °C for 120 min. To avoid the adverse effects of Mn and Al on the quality of the Ni and Co sulfate products, Mn²⁺ was first separated and precipitated via the KMnO₄ oxidation–precipitation method, and >98% of the Mn was removed and precipitated within 30 min with a K_p/K_t (ratio of actual usage to theoretical usage of KMnO₄) of 1.0 at pH = 2.0 and 25 °C. After removal of the Mn, the solvent extraction method was adopted by using P204 as an extractant to separate Al. More than 98% of the Al was extracted in 30 min with 20% (v/v) P204 + 10% (v/v) TBP with an A/O ratio of 1:1 at 30 °C. This optimized process for extracting lithium residues improved the hydrogen reduction process of waste lithium batteries and will enable industrialization of the developed processes.

Fluorinated carbon (CF_x) has been extensively served as promising positive electrode material for lithium primary batteries due to its high energy density. However, there are comparatively far less reports about the use of CF_x on other battery systems, let alone on the research of aqueous batteries. Herein in this study, we employed CF_x as the cathode active for aqueous zinc batteries for the first time and systematically investigated its electrochemical behavior under a series of aqueous zinc-ion electrolytes. As is discovered that the F/C ratio (the x value in CF_x) of CF_x have significant effects on the electrochemical performance of aqueous Zn/CF_x batteries. Specifically, CF_0.85 exhibits excellent electrochemical property with delivering a remarkable discharge capacity of 503 mA h g⁻¹ and energy density of 388 W h kg⁻¹ (at a current rate of 30 mA g⁻¹ under temperature of 25 °C), much better than several other CF_x electrode with F/C ratio of 0.70, 0.95, and 1.10, respectively. Besides, it also exhibits decent temperature performance with discharge capacities of 550 mA h g⁻¹ at 50 °C and 460 mA h g⁻¹ at 0 °C under current density of 30 mA g⁻¹. Furthermore, the electrochemical discharge mechanism based on conversion reaction was further uncovered by applying XPS, XRD, SEM and EDS elemental analysis characterization techniques. In conclusion, these results demonstrate the potential application value of CF_x in aqueous zinc primary batteries.