The recovery of valuable components manganese and lithium from retired lithium manganese oxide batteries by carbon-thermal reduction—acid leaching combined recovery process was studied.The mixed powder of lithium manganese oxide and graphite was roasted by carbon thermal reduction,and the roasted products were characterized by XRD,XRF,SEM,TG-DTA and other technologies.The results show that the best effect is achieved by roasting at 650 ℃ for 180 min,and the lithium manganese oxide in the roasted product is completely converted into manganese monoxide and lithium carbonate.The lithium carbonate in the sample powder can be extracted by water leaching,and the leaching rate of lithium carbonate is 86.15%.Manganese ions are extracted by acid leaching of sulfuric acid.Under the acid leaching conditions of acid leaching concentration of 3.5 mol/L,acid leaching temperature of 60 ℃,acid leaching time of 3 h,and liquid volume to solid mass ratio of 8/1,the highest leaching rate of manganese ions is 88%.The method can achieve the purpose of synchronous and efficient recovery of manganese and lithium from cathode materials,and has certain application value.

Preparation of alumina from fly ash by hydrochloric acid method is one of the processes with significant industrial application potential.Therefore,studying the ionic structure in acid leaching solution of fly ash is of great significance for iron removal process in the method.The ionic structure of the AlCl₃-FeCl₃-FeCl₂-HCl-H₂O system in hydrochloric acid leaching solution of fly ash at pH values of 1.0,1.5,and 2.0 was investigated by combining thermodynamic calculation,quantum chemical calculation,and Raman spectroscopy. The results show that at pH = 1.0,the main forms of Fe and Al complex ions in the system are [FeCl]⁺,[FeCl₂]⁺ and [AlCl]²⁺;when the pH rises to 1.5 and 2.0,[FeCl]⁺,[FeCl₂]⁺ and [AlCl]²⁺ transform into hydrolysis products such as [FeOH]⁺,[FeOH]²⁺ and [AlOH]²⁺ with hydroxyl ligands.The wave function analysis results indicate that at low pH,due to the high concentration of chloride ions,the chloride complexes are more stable.As the pH increases,the concentration of hydroxide ions increases,and the formation of high-coordination hydroxyl complexes becomes easier due to their higher metal-oxygen bond order and lower Gibbs free energy.

Address to the issue of high fluorine content in the flotation concentrate of a super-large Beryllianite-type uranium-beryllium co-associated ore in Xinjiang,the mineral and elemental composition,as well as the mineral dissemination characteristics were studied.Flotation process of "floating fluorite first and then beryllium" was adopted for flotation.Grinding fineness and flotation reagent system were optimized through systematic flotation condition tests. The results show that for the raw ore with a beryllium grade of 0.435% and a grinding finness of -325 mesh accounting for 93%,under the conditions of 1.5 kg/t for sodium silicate in roughing,2 kg/t for NaOH,700 g/t for oxidized paraffin soap in roughing,and 500 g/t for swept oxidized paraffin soap in scavenging,The technical indicators of flotation with beryllium grade of 3.36% and recovery rate of 81.99% are obtained. The process can achieve effective enrichment of beryllium minerals.

Iron and lithium were recovered from the cathode material of spent lithium iron phosphate battery using choline chloride,ascorbic acid,and ethylene glycol as a ternary deep eutectic solvent.The effects of the molar ratio of choline chloride/ascorbic acid/ethylene glycol,liquid volume to solid mass ratio,reaction temperature and time on the leaching rate of iron and lithium were investigated.The leaching mechanism was discussed through kinetic analysis and SEM characterization.The results show that under the optimal leaching conditions of choline chloride/ascorbic acid/ethylene glycol molar ratio of 1∶1∶3,liquid volume to solid mass ratio of 0.1 mL/1 mg,reaction temperature of 80 ℃ and reaction time of 1 h,the leaching rates of lithium and iron can reach 96% and 98%,respectively.The leaching process is mainly controlled by chemical reactions.The method is efficient and environmentally friendly,and can recover iron and lithium from spent lithium iron phosphate batteries.

In order to solve the problem of real-time and accurate parameter optimization in hydrometallurgical equipment operation,an optimization setting compensation method based on real-time data acquisition was proposed by combining the improved POPOA method and the improved JITL online learning method.The results show that compared with the traditional method,the retraining time of the modified JITL method is significantly reduced,the optimization rate is significantly increased,and the energy consumption is significantly reduced.The improved POPOA method significantly improves the performance of real-time data processing,and the processing time is about 40% shorter than that of the traditional method.The improved POPOA method reduces the load rate of the system significantly compared with the traditional method when the multi-task is running concurrently.This method can effectively improve the accuracy of operation performance evaluation,the real-time response ability of the system,and the generalization ability of the model,and reduce energy consumption and operation cost,so it has a certain application prospect.

Extraction and separation of valuable elements from the positive leaching solution of spent lithium-ion batteries by β-diketone/phosphate extraction system was studied.The optimal conditions for the extraction and separation of cobalt,nickel,manganese,and lithium were determined through equilibrium extraction.The results show that the β-diketone/phosphate extraction system can effectively separate cobalt,nickel,manganese,and lithium from the positive leaching solution of spent lithium-ion batteries by controlling the kinetics.Under optimized conditions,the extraction rate of cobalt,nickel,and manganese can reach 99%,and the yield of lithium can reach more than 95%.The method realizes the separation and recovery of cobalt,nickel,manganese and lithium by a single extraction system,which can provide a new process route for the recovery of waste ternary lithium batteries.

Aiming at the problems of high sulfur content and difficult utilization of high sulfur bauxite,the alkaline leaching desulfurization process of high sulfur bauxite was studied.The effects of alkali mass concentration,liquid volume to solid mass ratio,leaching temperature and time on the desulfurization effect were investigated,and the reaction kinetics were analyzed.The results show that under the conditions of base concentration of 180 g/L,liquid volume to solid mass ratio of 8 L/1 kg,leaching temperature of 160 ℃ and leaching time of 5 h,the bauxite sulfur mass fraction after desulphurization is 0.42%.The desulfurization process is controlled by reaction-internal diffusion,and the apparent activation energy is 18.23 kJ/mol.The method can effectively reduce the sulfur content in high sulfur bauxite and is beneficial to the wide utilization of high sulfur bauxite.

Complex component sandstone-uranium deposits are mainly composed of conglomerate,sandstone and Slate.The uranium minerals are mainly uranite and titanium-uranium ores,including a small amount of pitchblende.The process mineralogy of complex component sandstone uranium ore was studied,and the leaching process was optimized.The results show that most of the uranium in the ore exists in the tetravalent form,and it contains a relatively large amount of calcium,magnesium,aluminium,iron and carbonate.The test results of acid leaching,enhanced leaching and column leaching show that the acid leaching process has a better effect. When 40~50 g/L H₂SO₄ is used as the leaching agent,the slag leaching rate of the acid column leaching is all greater than 90%. Comprehensively considered,it is recommended that the heap leaching process of -10 mm particle size ore be adopted in industrial production,and the mass concentration of the leaching agent H₂SO₄ is preferably 40 g/L.

The leaching of silver from failed silver-containing spent catalyst using HNO₃+H₂O₂ system was investigated.The kinetics of the silver leaching process was analyzed using the shrinkage kinematics model of liquid-solid phase reaction.The effects of leaching temperature and nitric acid concentration on the leaching rate of silver were examined.The results show that under the optimal leaching conditions of nitric acid concentration of 1.1 mol/L,leaching temperature of 50 ℃,stirring speed of 300 r/min,n(H₂O₂)∶n(Ag)=1.5∶1,leaching time of 50 min,and liquid volume to solid mass ratio of 4 mL/1 g,the leaching rate of silver can reach 94.18%.The leaching is controlled by the diffusion of the solid film,and the apparent activation energy of the leaching reaction is 15.45 kJ/mol,and the reaction order of hydrogen ion is 1.13.The method can provide reference for the research of efficient resourcing utilization of silver-containing spent catalyst.

In view of the problems of high energy consumption,high equipment requirements and low flexibility in recovering rhenium from processing waste by traditional pyrometallurgical processes,the electrochemical enhanced leaching—precipitation crystallization method was studied to recover high-purity KReO₄ crystals from rhenium secondary resources.The results show that when 22%~24% HNO₃ solution is used as the electrolyte,there is no obvious passivation during the electrolysis process,and the energy consumption is stable at about 3.0 kWh/kg.During the electrolysis,Re atoms at the hexagonal lattice sites on the anode surface lose electrons,combine with hydroxyl groups and transform through low-valent oxidation states of Re to bridge oxygen connected Re(Ⅱ),and finally enter the electrolyte in the form of \mathrm{R} \mathrm{e} {\mathrm{O}}_4^{-} after reacting with the acid.When potassium salt is used as the precipitant to recover Re elements in the electrolyte,under the conditions of crystallization temperature of 25 ℃,precipitant flow rate of 6 mL/min,stirring rate of 500 r/min and crystallization time of 30 min,the precipitated KReO₄ crystals are in the shape of polyhedral spindle,with good uniformity in particle size,high recovery rate and purity of 99.95%,which can meet the requirements for hydrogen reduction to prepare metallic rhenium.The method can effectively recover rhenium processing waste and has certain promotion value.