To address the issues of prolonged flowsheet,high costs,and inter-containment of concentrates in traditional molybdenum-bismuth mixed concentrate beneficiation process,the HCl-NaCl system was employed to leach bismuth from molybdenum-bismuth mixed concentrate to realize the separation of molybdenum and bismuth,and a thermodynamic analysis was performed on the leaching process.The influences of hydrochloric acid concentration,leaching temperature,leaching time,liquid-solid ratio,and NaCl concentration on the leaching rate of bismuth were investigated.The reaction kinetics were analyzed using leaching experimental data to identify the reaction rate-controlling step.The results show that the forms of Bi and S are BiCl₄^-and H₂S,respectively after the reaction of bismuth sulfide with hydrochloric acid without oxidant addition.The factors affecting bismuth leaching rate follow this order of significance:leaching temperature>hydrochloric acid concentration>liquid-solid ratio>NaCl concentration>leaching time.Under the optimized conditions of HCl concentration of 4 mol/L,liquid-solid ratio of 2.0,reaction temperature of 90℃,leaching time of 8 h,and NaCl concentration of 4 mol/L,the average bismuth leaching rate reaches 97.90%.The apparent activation energy of the leaching reaction is 84.55 kJ/mol,indicating an interface chemical reaction control mechanism,and elevated temperature promotes bismuth leaching.

Aiming at the problems of low recovery rate of high-value rare earth elements praseodymium and neodymium,long process and high consumption of auxiliary materials in the current mainstream leaching process of bastnaesite,a new complex leaching technology of bastnaesite with sulfuric acid as the complexing agent was proposed,greatly simplifying the leaching process and realizing the synergistic efficient leaching of F^-and rare earth in hydrochloric acid system.The effects of complexing agent types and dosages,acid-ore volume-mass ratio,leaching temperature,leaching time and liquid-solid volume-mass ratio on the rare earth leaching rate were systematically investigated.The obtained optimum process conditions are as follows:complexing agent is sulfuric acid,sulfuric acid dosage is 1%of the mass of bastnaesite roasted ore,acid-ore volume-mass ratio is 1.1 mL/g,leaching temperature is 25℃,liquid-solid volume-mass ratio is 4 mL/g and leaching time is 3 h.Under these conditions,the leaching rate of praseodymium and neodymium increases from 94.51%of the current process to 96.23%.

High-purity niobium metal was prepared by aluminothermic reduction-double electron beam melting,and the production process conditions were optimized to increase niobium recovery rate,save production costs and improve production efficiency.The results show that the recovery rate of niobium is relatively high under the following conditions:excess aluminum ratio of 8%~11%,slagging agent(80%CaO+20%CaF₂)dosage of 4%relative to the added aluminum mass,melting power of 250 kW,melting speed of 90~100 kg/h and melting vacuum level of≤10^-2Pa.And the quality of melted niobium products meets the NT-3 grade requirements of YS/T 1547—2022 standard.

In this paper,stone coal vanadium ore was used as raw material,and the vanadium was extracted by barium salt roasting-low acid leaching process from the concentrate after tailing discarding and enrichment.The effects of concentrate roasting and leaching conditions on the leaching rate of vanadium were studied.The results show that the average vanadium leaching rate can reach 87.93%under the optimal conditions of stone coal vanadium concentrate fineness of 70%passing 74μm,roasting temperature of 925℃,roasting time of 6 h,barium sulfate dosage of 4%concentrate quality,liquid-solid ratio in leaching process of 2∶1,sulfuric acid dosage of 5%calcine quality,leaching temperature of 50℃and leaching time of 1 h.The process has the advantages of high vanadium leaching rate,low acid dosage,low extraction cost and simplified subsequent vanadium extraction process.

Lithium in medium to low-grade lepidolite(0.93%Li)was extracted by roasting-water leaching method with sodium sulfate-calcium sulfate composite sulfate as roasting agent.The effects of roasting process parameters on the leaching rate of lithium were discussed.The thermodynamics and kinetics of the roasting process,the microscopic morphology and phase change of the roasted clinker were studied.The results show that the optimum roasting conditions are as follows:the mass ratio of lepidolite,Na₂SO₄and CaSO₄is 1∶0.197∶0.703,roasting temperature is 950℃and roasting time is 2 h.Under these conditions,the leaching rate of lithium can reach 94.8%by leaching with pure water.Under the optimum roasting conditions,the stable structure of lepidolite is destroyed,and its reactions with calcium sulfate and sodium sulfate mainly produce anorthite,albite and soluble lithium salt.Using unreacted shrinking core model to describe the kinetics of lepidolite roasting in the presence of sodium sulfate and calcium sulfate,the apparent activation energy of the roasting process is calculated to be 52.1 kJ/mol at 825~900℃,indicating that the roasting process is controlled by chemical reaction.

In this paper,gallium nitrate was taken as the raw material,gallium oxyhydroxide was successfully prepared by gallium salt precipitation method,and the effects of process parameters such as precipitant type,reaction temperature,pH value and gallium ion concentration on the preparation of gallium oxyhydroxide were systematically investigated.The results show that the precipitation rate of gallium reaches 79.7%,and the precipitated product is white powder of gallium oxyhydroxide under the optimal conditions of ammonia as precipitant,reaction temperature of 70℃,pH value of 8,and gallium ion concentration of 0.16 mol/L.Theβ-Ga₂O₃powder with uniform morphology and conforming to the Ga₂O₃3N standard in YS/T 741—2010 can be prepared by roasting the precipitated product at 800℃for 3 h.In this study,gallium nitrate can be efficiently converted into gallium oxyhydroxide with homogeneous morphology and size by gallium salt precipitation method,with simple process flow,low production cost,and prospect for industrialization.

In response to the problem of low germanium volatilization rate during the matte smelting of Gecontaining copper concentrates,taking high-sulfur low-grade Ge-containing copper concentrate from Yunnan as the research object,the effects of reaction atmosphere,slag composition,and smelting temperature on copper matte grade,Ge volatilization rate,and Ge distribution rules in dust,copper matte,and slag were systematically investigated.The results show that under the optimized conditions of carbon addition of 15%,oxygen concentration of 21%in top-blown gas,Fe/SiO₂mass ratio of 1.1,smelting temperature of 1350℃,and CaO addition of 10%,the copper matte grade obtained by top-blown smelting is 17.21%,and the Ge volatilization rate can reach 94.88%.A strongly reducing atmosphere is conducive to Ge volatilization,but not conducive to desulfurization.Lead and zinc enter the copper matte primarily in the form of sulfides,resulting in a lower copper matte grade.Appropriate Fe/SiO₂mass ratio and CaO addition promote the formation of pyroxene-based slag,reduce slag viscosity,and improve melt fluidity and copper matte settling efficiency,which is beneficial for increasing Ge volatilization rate and copper matte grade.

Antimony is an important strategic metal.China is the country with the richest antimony reserves and the largest antimony product production in the world.However,the long-term over-exploitation has led to a significant decline in China's current antimony resource advantage.Focusing on the current situation of antimony ore resources and industrial development in China,the technical features of the current dominant antimony smelting processes are systematically summarized,and the research progress of oxygen-enriched molten pool smelting process is highlighted.

The industrial test of“oxidation smelting-hot oxidized slag reduction-hot reduced slag fuming”using the oxygen-enriched side-blown triple furnace was carried out to process jamesonite,and the rules of element migration during the smelting process were studied.The test results show that the lead-antimony alloy with Pb and Sb mass fractions of 70.11%and 20.52%,respectively,can be produced in the reduction furnace and the direct yields of Pb and Sb are 95.1%and 88.2%,respectively.The fuming slag becomes harmless glass slag after water quenching treatment,the comprehensive recovery rates of Pb,Sb,Zn and Ag are 99.8%,97.6%,88.0%and 99.5%,respectively,and the S utilization rate is approximately 99.0%.The coal consumption for processing 1 t jamesonite is 259.3 kg powdered smokeless coal.This process has broad application prospects,due to the advantages of good environmental protection effect,high recovery rate of valuable elements,low energy consumption and high degree of automation.

由长沙有色冶金设计研究院有限公司自主研发的“沉积型软破围岩铝土矿床顶板安全管控综合技术”,解决了沉积型软破围岩铝土矿床地下开采顶板管控难度大、顶板垮冒事故易发多发、矿石损失率高等世界性难题。该技术实现了铝土矿采场安全回采,矿石大块率控制在5%以下,爆破成本降低0.27元/t,采场生产能力提高10%~15%,采矿回收率提高20%以上,技术经济指标先进,整体技术达国际领先水平。