Extraction of lithium from alkali leaching solution of rare earth molten salt electrolytic residue by gas-liquid reaction crystallization method was studied. The effects of temperature, carbon dioxide flow rate, pH, aeration time and stirring speed on the carbonation rate were investigated. The results show that the carbonization rate is about 83% under the optimized conditions of temperature of 90 ℃, CO₂ inflow rate of 1 000 mL/min, ventilation time of 20 min and stirring speed of 450 r/min. The carbonization rate increases first and then decreases with the increase of CO₂ inflow rate and ventilation time. The product is lithium carbonate with no impurities and high crystallinity. This method overcomes the problems of local supersaturation and uneven crystal size caused by traditional precipitation method, and provides a new technical path for the resource treatment of complex alkaline lithium-containing wastewater. And it can promote the rare earth smelting industry to the direction of green, recycling, high value development.

Aiming at the issues of strong corrosiveness and environmental pollution associated with traditional inorganic acid leaching methods for recovering cathode powder from spent lithium iron phosphate batteries. The ultrasonic-enhanced leaching of lithium and iron from spent lithium iron phosphate battery cathode powder using a mixture of organic acids such as gluconic acid and tartaric acid as the leaching agent and 3 mol/L hydrogen peroxide as the oxidant was investigated. The effects of ultrasonic power, gluconic acid concentration, tartaric acid concentration, leaching time and leaching temperature on the leaching rates of lithium and iron were examined. The results show that the leaching rate of lithium and iron reach 99.1% and 2.5%, respectively, under the optimal conditions of gluconic acid concentration of 0.5 mol/L, tartaric acid concentration of 1 mol/L, ultrasound power of 600 W, liquid volume to solid mass ratio of 6∶1, leaching time of 90 min and leaching temperature of 60 ℃. Ultrasonic intensification has little effect on the chemical composition of the leaching residue. After heat treatment, the leaching residue forms iron phosphate with good crystallinity. The leaching process is mainly controlled by chemical reactions, and the leaching effect is good.

Aiming at the problems of serious pollution and low efficiency associated with the traditional roasting process for treating low-grade, high-sulfur and refractory gold concentrates, a combined process of oxygen pressure water leaching and cyanidation was proposed. The process parameters were systematically optimized, and the leaching mechanism was investigated. The effects of the process on gold recovery were evaluated through process mineralogical analysis and cyanidation experiments. The results show that under the optimal conditions of temperature of 180 ℃, oxygen partial pressure of 1.5 MPa, leaching time of 120 min, stirring speed of 700 r/min and liquid volume to solid mass ratio of 200∶1, oxygen pressure water leaching can effectively oxidize pyrite and other sulfide minerals into hematite, thereby destroying the encapsulated state of gold. After pretreatment, the gold grade in the leaching residue increases from 42.24 g/t to 63.78 g/t, with an enhancement of over 50%. Process mineralogical analysis indicates that gold is released from sulfides, occurring in exposed or semi-exposed states. The mineral structure becomes loose and porous, and the gold exposure significantly increases, providing a more favorable reaction interface for subsequent cyanidation. Cyanidation experiments further confirm the pretreatment effect, achieving a maximum gold leaching rate of over 99.50%, along with a marked reduction in cyanide consumption. These results demonstrate that oxygen pressure water leaching is an effective pretreatment technology for improving the gold recovery and reducing reagent consumption in low-grade, high-sulfur and refractory gold concentrates, providing technical guidance for the efficient and environmentally friendly utilization of such resources.

Blank roasting-carbon leaching process was used to extract gold from low-grade gold cyanide tailings in a certain place. The effects of roasting temperature, roasting time, dosage of 'Jinchan' mineral processing agent, leaching time and liquid volume to solid mass ratio on gold leaching rate were investigated. On the basis of small-scale experiments, an expanded experiment was carried out. The results show that the optimal roasting process parameters are roasting temperature of 550 ℃ and roasting time of 20 min. The optimum carbon leaching process parameters are as follows：dosage of 'Jinchan' mineral processing agent is 500 g/t, leaching time is 12 h, pulp pH is 12, bottom carbon density is 15 g/L, and liquid volume to solid mass ratio is 2∶1. Under the optimized roasting-carben leaching process parameters, Au leaching rate can reach more than 60%, which is significantly higher than the gold leaching rate obtained by direct leaching of cyanidation leaching residue without roasting. During the continuous operation of the rotary kiln roasting expansion test, the Au grade in each calcine process sample and the corresponding Au grade in the leaching residue fluctuate little, and the corresponding Au leaching rate is between 59.79% and 62.11%, and the continuous operation index is stable. The method can provide a technical reference for the recovery of gold from low-grade gold cyanide tailings.

The preparation of yttrium oxide (Y₂O₃) crystals suitable for industrial production of electronic ceramics by a two-step precipitation method using ammonia water and ammonium bicarbonate was studied. The effects of solution concentration, precipitant concentration, surfactant addition amount, reaction temperature, feeding rate, stirring speed, aging method and time, and calcination method and time of the precursor on the particle size of the prepared Y₂O₃ powder were investigated. The decomposition mechanism and kinetics of the precursor were also explored. The results show that the use of ammonium sulfide to purify 99.9% of yttrium oxide can effectively remove metal impurities. Under the optimal conditions of room temperature, Y(NO₃)₂ solution mass concentration of 100 g/L, NH₄HCO₃ solution mass concentration and addition rate of 200 g/L and 6 mL/min, respectively, stirring speed of 370 r/min, aging time of 1 h, surfactant addition amount of 2.5%, microwave drying for 10 min, calcination temperature of 800 ℃, and calcination time of 2 h, spherical cubic Y₂O₃ crystals with D₅₀ = 1.018 μm and BET = 31.32 m²/g can be prepared, which meet the requirements of Y₂ O₃ powder for electronic ceramics. The analysis of the decomposition process and mechanism of the precursor Y₂(CO₃)₃·2H₂O to Y₂O₃ indicates that Y₂(CO₃)₃·2H₂O is converted to Y₂ O₃ through the intermediate state of Y₂CO₅. The decomposition of the precursor to Y₂O₃ occurs in three stages, with activation energies of 61.68, 564.22, and 590.103 kJ/mol, and corresponding reaction orders of 1.836 9, 0.896 4, and 0.972 2, respectively. The method can achieve industrial production of Y₂O₃ for electronic ceramics. The entire process is simple to operate, low-cost, environmentally friendly, and pollution-free, with significant economic and social benefits and broad application prospects.

To address the issues of long leaching time and high consumption of hydrochloric acid in the conventional hydrochloric acid leaching method for neodymium-iron-boron (NdFeB) waste, the ultrasonic-enhanced hydrochloric acid solubilization method was studied for the leaching of rare earths from it. The effects of ultrasonic power, hydrochloric acid concentration, temperature, liquid volume to solid mass ratio, leaching time, particle size and other factors on the leaching rate of rare earths were investigated, and the optimal leaching parameters were determined. The results show that ultrasonic assistance significantly enhances the leaching process. Under the optimal conditions of ultrasonic power of 60 W, hydrochloric acid concentration of 2 mol/L, temperature of 85 ℃, liquid volume to solid mass ratio of 10 mL/1 g, leaching time of 105 min, and particle size of -0.074 mm, the rare earth leaching rate reaches 97.94%, which is 8 percentage points higher than that of the conventional hydrochloric acid leaching method under its best conditions. The method is an efficient and feasible green process for rare earth recovery and provides a new approach for the resource utilization of NdFeB waste.

The extraction of niobium and rare earths from niobium concentrate by concentrated sulfuric acid low-temperature roasting-water leaching process was studied. The effects of sulfuric acid dosage, calcination temperature, calcination time, liquid volume to solid mass ratio of water leaching, water leaching time and water leaching temperature on the leaching rates of niobium and rare earths were investigated. The results show that under the optimized conditions of 120% sulfuric acid dosage, calcination temperature of 180 ℃, calcination time of 4 h, liquid volume to solid mass ratio of water leaching of 4∶1, water leaching time of 4 h and water leaching temperature of 25 ℃, the niobium leaching rate is 91.4% and the rare earth leaching rate is 93.9%. The process can achieve efficient extraction of niobium and rare earths from niobium concentrate.

Based on the preparation of graphene oxide-ferroferric oxide composite (GO-Fe₃O₄) by chemical co-precipitation, hydrophobic modification of GO-Fe₃O₄ was carried out via physical adsorption using D2EHPA to obtain the D2EHPA@GO-Fe₃O₄ composite material (GOFP). Subsequently, a W/O-type GO-Fe₃O₄-based Pickering emulsion was successfully constructed using sulfonated kerosene as the oil phase, 1 mol/L hydrochloric acid as the internal aqueous phase, P507 (extractant) as the mobile carrier and GOFP as the solid emulsifier, which was used to extract heavy rare earth ion Ho³⁺. The effects of extraction time, mobile carrier dosage, emulsion-to-aqueous phase ratio, feed solution pH, and initial Ho³⁺ concentration on the extraction rate of Ho³⁺ were systematically investigated. The results indicate that under optimal conditions of extraction time of 3 min, feed solution pH ≥ 4.6, initial Ho³⁺ concentration of 0.002 mol/L, P507 (extractant) dosage of 0.5 mL, and emulsion-to-aqueous phase ratio of 0.3, the extraction rate of Ho³⁺ reaches 88%. The GO-Fe₃O₄-based Pickering emulsion liquid membrane method demonstrates promising application potential.

A gold-bearing arsenic concentrate from Inner Mongolia is a typical refractory gold ore, where gold exists in an encapsulated form within arsenopyrite (FeAsS) and pyrite (FeS₂). The encapsulation of arsenides not only severely reduces the efficiency of cyanide leaching of gold extraction but also affects environment. The method of removing arsenic and extracting gold by pressurized oxidation was studied. The influences of sulfuric acid concentration, oxygen partial pressure, reaction temperature, solid mass to liquid volume ratio, and reaction time on the leaching effect were investigated. The process conditions were optimized and the key constraints were revealed. The results show that pressure leaching can effectively destroy the coating structure of gold by sulfides and arsenides, and there is no emission of harmful gases such as As₂O₃ and SO₂ throughout the process. Its environmental friendliness is significantly better than that of oxidation roasting. The optimal process conditions are determined as follows:sulfuric acid concentration of 0.4 mol/L, oxygen partial pressure of 1.5 MPa, reaction temperature of 160 ℃, solid mass to liquid volume ratio of 1∶100, and leaching time of 120 min. Under the conditions, the arsenic leaching rate reaches 96%, and the mass fraction of gold in the leaching residue is 28 g/t, which is significant increased compared with 4.5 g/t of the raw ore. The persistent presence of ferric arsenate (FeAsO₄) precipitate in the leaching residue is the core factor restricting the complete removal of arsenic and the further improvement of gold grade. The research results can provide technical references for the clean and efficient utilization of similar refractory gold mines.

A process of solvent extraction-air roasting-sulfuric acid leaching-extraction with extractant-stripping was studied for the recovery of Mo and Fe from direct coal liquefaction residue (DCLR). The results show that after solvent extraction, the residual organic compounds in DCLR are efficiently removed, and the coal liquefaction residue extractant residue (ER) containing residual carbon and ash is obtained. ER is then air-roasted at 600-800 ℃ to completely remove the residual carbon, with Mo and Fe being enriched in the ash. The ash is leached with sulfuric acid to effectively extract Mo, Fe, Al, and other metals, with the highest leaching rates of Mo and Fe reaching 93.10% and 99.12%, respectively. The acid leaching solution is extracted with tri-n-octylamine (N235) extractant to simultaneously extract Mo and Fe, but NaCl is added to form complex anions such as and , achieving an ideal extraction effect. The optimized extraction process conditions are NaCl addition of 5%, N235∶sec-octyl alcohol∶kerosene ratio of 3∶2∶5, and hydrogen ion concentration greater than 12 mol/L, phase ratio V_O/V_A between 1/3 and 1/1, and extraction time greater than 2.5 min. Under the conditions, the single extraction rate of Mo is as high as 99.16%, and the extraction rate of Fe is close to 100%. Fe and Mo in the organic phase are sequentially stripped with sulfuric acid and ammonia, and the recovery rates of Fe and Mo in the entire extraction-stripping process are 97.90% and 94.01%, respectively. The method can provide a solution for the recovery of high-value metals in direct coal liquefaction catalysts.