Aiming at the problems of low molybdenum grade, complex ore properties and poor enrichment effect by conventional mineral processing methods for a certain molybdenum deposit, on the basis of ore property research, acid leaching, alkaline leaching and enhanced leaching tests were carried out respectively. The high-acid enhanced leaching process was optimally selected. The effects of sulfuric acid dosage, ore particle size, leaching temperature, oxidant dosage, liquid volume to solid mass ratio and leaching time on molybdenum leaching efficiency were investigated, and the optimal leaching process parameters were determined. The results show that under the optimal conditions of ore particle size of -100 mesh, sulfuric acid dosage of 20%, pyrolusite dosage of 3%, leaching temperature of 80 ℃, liquid volume to solid mass ratio of 1.5∶1 and leaching time of 3 h, the molybdenum leaching rate calculated by leaching residue reaches 71.18% via the high-acid enhanced leaching process. To solve the problem of excessive residual acid in high-acid enhanced leaching solution, a two-stage countercurrent leaching process is adopted. The average molybdenum leaching rate calculated by leaching residue increases to 72.16%, which is about 1 percentage point higher than that of single-stage leaching. Meanwhile, the mass concentration of residual acid decreases by more than 40%, and the total consumption of sulfuric acid reduces by 25%. The ton-scale pilot test results of the two-stage countercurrent leaching process indicate that under suitable conditions, the total sulfuric acid dosage is 15%, the residual acid mass concentration of leaching solution is 48.8 g/L, and the molybdenum leaching rate calculated by leaching residue remains above 71%. This process proves stable and reliable with good reproducibility of leaching indexes, which serves as the technical basis for the industrial design of this molybdenum deposit.

Aiming at the problems of poor heap leaching permeability, low molybdenum leaching rate and difficult solid-liquid separation in stirring leaching of low-grade uranium-molybdenum ore in Guyuan, a process path of "mixing acid curing-low temperature roasting" synergistic enhanced acid heap leaching was proposed, and the key parameters such as acid concentration, curing temperature and time, roasting temperature and time, and spray intensity were optimized under column leaching conditions. Considering the energy consumption and process effect, the recommended conditions are as follows：ore particle size is -1 mm, acid mixing mass fraction is 12%, curing at 90 ℃ for 3 h, roasting at 300 ℃ for 30 min, acidizing water mass concentration is 10 g/L, spray intensity is 76 L/(m²·h). Under these conditions, the leaching rate of molybdenum slag can reach 82.89%, the molybdenum grade of tailings is reduced to 0.038%, and the uranium grade is reduced to 100.33×10^-6. The results of multi-batch ore verification test show that the process flow has adaptability to uranium-molybdenum ore with fluctuating grade and property. The molybdenum leaching rate of different low-grade uranium-molybdenum ore is stable at 72.51%–77.37%, which can achieve the process goal. The research results can provide technical support for the efficient extraction of low-grade complex uranium-molybdenum ore resources.

Aiming at the characteristics of complex intergrowth and intimate association of various mineral phases in titanium-iron tailings, which lead to difficult mineral separation, this study adopts benzohydroxamic acid as collector, sodium hexametaphosphate as depressant and 2^# oil as frother to realize flotation separation of valuable components from titanium-iron tailings. The influences of pulp pH, benzohydroxamic acid dosage, sodium hexametaphosphate dosage, 2^# oil dosage and flotation stages on the mass proportion of flotation products, TiO₂ recovery rate, TiO₂ grade and total iron mass fraction were systematically investigated, and the corresponding reaction mechanism was further discussed. The results show that under the optimal conditions of sodium hexametaphosphate dosage of 500 g/t, benzohydroxamic acid dosage of 35 kg/t, 2^# oil dosage of 1 mL and pulp pH=9, the mass proportion of floated product reaches 25.85%, the TiO₂ recovery rate is up to 42.23%, the TiO₂ grade is 14.93%, and the total iron mass fraction is 15.22%. After six stages of flotation, qualified flotation concentrate is prepared, with the floated product mass proportion of 71.82%, TiO₂ recovery rate of 75.81%, TiO₂ grade of 27.84% and total iron mass fraction of 24.49%. XRD and FT-IR analysis results demonstrate that the depressant sodium hexametaphosphate can exert an obvious inhibitory effect on pyroxene phase and silica. Benzohydroxamic acid is capable of generating adsorption behavior on the surface of titanium-iron tailings particles, which remarkably improves the overall flotation separation effect.

An antimony diaphragm electrorefining technology suitable for tartrate system was developed using refined antimony as raw material. The effects of different diaphragm types and electrolysis process parameters including current density, electrolysis temperature, electrolyte circulation speed and electrode spacing on cathode current efficiency, cell voltage, direct current consumption, cathode antimony purity and typical impurity removal rate are systematically investigated. The optimum process conditions are determined as follows:tartrate ion mass concentration is 250 g/L, antimony ion mass concentration is 45 g/L, current density is 150 A/m², electrolysis temperature is 35 ℃, electrolyte circulation speed is 300 mL/min, electrode spacing is 70 mm. The research results show that according to the single point impurity content requirements of 5N grade high purity antimony in the national standard' high purity antimony' (GB/T 10117—2021), the impurities that are easy to exceed the standard in the antimony tartrate electrolysis system are mainly arsenic, bismuth, lead and copper. Only regulating the electrolysis process conditions can not achieve the deep removal of the above impurities. Further electrolyte purification measures are needed to ensure the batch stability of high purity antimony products, while the remaining 16 main control impurities can meet the requirements of the national standard Sb-05 by electrolytic refining. Under the optimal electrolysis conditions, the electrolyte is deeply purified with antimony-based adsorbent (polyhydroxyantimonic acid with the main phase of H₁₄Sb₁₄O₂₁(OH)₄₂) to obtain cathode antimony with a purity of 99.995% and a dense and bright surface. The average cell voltage is 0.808 V, the current efficiency is 96.27%, and the direct current consumption is 485.85 kWh/t.

Extraction of aluminum from fly ash by ammonium sulfate roasting-leaching process was studied. The effects of roasting temperature, roasting time and the molar ratio of ammonium sulfate to Al₂O₃ in fly ash on the extraction of aluminum from fly ash were investigated. The interaction of various factors on the extraction rate of aluminum from fly ash was discussed by response surface method, and the optimum roasting parameters were determined. The phase and microstructure of fly ash raw materials and calcined products were characterized by X-ray diffractometer (XRD) and scanning electron microscope (SEM). The results show that the main product of fly ash activated by ammonium sulfate is NH₄Al(SO₄)₂, and the surface of the roasted product is corroded. The optimum process parameters of activation roasting of fly ash are as follows：roasting temperature is 438 ℃, roasting time is 63.8 min, molar ratio of ammonium sulfate to Al₂O₃ in fly ash is 7.31. Under the optimum process conditions, the leaching rate of Al is 95.30%, and the leaching effect is better.

In view of the recovery and utilization of associated thorium resources and the control of radioactive pollution in the process of bastnaesite smelting, the migration and distribution characteristics of thorium elements in each link of smelting separation were systematically studied. After the pretreatment of bastnaesite by alkali decomposition, the leaching acidity was adjusted to 3-5 mol/L, and the thorium leaching rate was increased to 98%. Fe³⁺ in the feed solution was selectively removed by N235 extractant in advance, and then thorium was directional extracted by P507 extractant. The results showed that the saturated loading of thorium oxide in the diluted P507-sulfonated kerosene extraction system could reach 25 g/L. After oxalic acid back extraction, washing and burning at 1 000 ℃, the purity of thorium oxide in the final product reached 98.88%. The total yield of thorium is more than 96% for the raw material of bastnaesite containing 0.2% -0.3% thorium. The method is efficient and environmentally friendly, and has certain application potential in the release treatment of thorium-containing solid waste.

The iron oxide loaded dolomite composite (DFC) was prepared by ball milling-calcination method and used to degrade malachite green (MG) in water. The degradation performance and mechanism of DFC were investigated. The results indicate that DFC has good degradation ability for MG, and the degradation rate is affected by the adding amount of DFC, solution initial pH, temperature, and coexisting anions. Under the optimal conditions of DFC adding amount of 30 mg, initial pH=7, and temperature of 40 ℃, the degradation rate of MG can reach 96% within 10 min. Free radical trapping experiments show that superoxide radical () is the dominant active species, and hydroxyl radical (·OH) plays a synergistic role. Intermediate analysis reveals that MG mainly achieves deep degradation through demethylation, chromophore destruction, and aromatic ring breakage. This study can provide a technical reference for the application of natural mineral-based iron oxide composites in dye wastewater treatment.

Addressing the technical challenges of poor settling performance and low solid-liquid separation efficiency of the collapse deposits in the MSESA orebody of the Democratic Republic of the Congo, the process mineralogy and settling characteristics of these deposits was systematically analyzed. Furthermore, it investigates the mineralogical features and settling behaviors of both the collapse deposits and their mixed systems with the MSESA run-of-mine ore. The results indicate that the collapse deposits are low-grade copper-cobalt oxide materials, with copper mainly existing as free copper oxide and cobalt primarily hosted in gangue minerals. Settling tests determined the optimal parameters for different systems:the optimal flocculant dosage for both raw collapse deposits and the mixed ore is 20 g/t, with final settling times of 40 and 25 min, respectively. For the leaching slurries of the collapse deposits and the mixed ore, the optimal flocculant dosage is 120 g/t, with final settling times of 20 and 40 min, respectively. These findings provide a reliable basis for the resource utilization and process design of collapse deposits in the MSESA orebody.

Lead-zinc ore is the main raw material for zinc hydrometallurgy, and cadmium is one of the major associated metal elements in lead-zinc ore. During the neutral leaching process, cadmium enters the leachate, producing a high-cadmium zinc sulfate solution. Aiming at the common issue in the zinc hydrometallurgy industry of the lengthy process for cadmium separation and recovery via zinc powder cementation, the two-stage countercurrent purification of cadmium from high-cadmium zinc sulfate solution using zinc powder was investigated. The appropriate zinc powder addition rate and residue return enrichment system were explored to increase the cadmium content in the cementation residue. The results show that the suitable zinc powder addition for the first-stage purification is 0.9 times the theoretical amount, and for the second-stage purification, it is 1.2 times the theoretical amount. Under optimal conditions, the cadmium concentration in the solution after the first-stage purification is approximately 70 mg/L, and after the second-stage purification, it is below 5 mg/L. The main phase of the first-stage purification residue is cadmium, with a cadmium mass fraction of 85.11% at the purification endpoint and a zinc mass fraction below 5%. The main phases of the second-stage purification residue are metallic cadmium and zinc. SEM-EDS analysis shows that the particles of the first-stage cementation residue are spherical with an encapsulation phenomenon. The encapsulation layer consists of basic zinc sulfate, while the inner layer is mainly unreacted zinc powder. The findings of this study can provide a technical reference for the separation and recovery of cadmium from high-cadmium leach solutions.

In order to accurately control the reduction, nucleation and growth process of platinum atoms in the preparation of high purity and high dispersion platinum powder, high purity and high dispersion platinum powder was prepared by chemical liquid phase reduction method with Na₂PtCl₆ aqueous solution as precursor, N₂H₄·H₂O as reducing agent and PVP as dispersant. The effects of temperature, reaction system and feeding method on the morphology, particle size distribution and dispersion of platinum powder were investigated. The results show that the nucleation and growth process of platinum atoms can be effectively separated by using PVP as the dispersion system and Na₂PtCl₆ and N₂H₄·H₂O as the synchronous feeding method at 80 ℃. The high purity and high dispersion micron platinum powder with good dispersion and narrow particle size distribution (D₉₀-D₁₀=1.18 μm) is prepared. This method provides a reliable theoretical basis and technical path for the stable and mass production of high purity and high dispersion micron platinum powder.