By using high-alumina fly ash as raw material,a process was proposed for activating the fly ash with Na_(2)CO_(3)calcination and extracting aluminum from activated clinker with sulfuric acid leaching.The feasibility o...By using high-alumina fly ash as raw material,a process was proposed for activating the fly ash with Na_(2)CO_(3)calcination and extracting aluminum from activated clinker with sulfuric acid leaching.The feasibility of roasting process of activated fly ash by Na_(2)CO_(3)was discussed based on thermodynamic analysis.The experimental results showed that Na_(2)CO_(3)gradually reactes with mullite over 700 K to produce NaAlSiO_(4).The optimal process conditions for the activation stage are:a material ratio of 1:1 between sodium carbonate and fly ash,a calcination temperature of 900℃,and a calcination time of 2.5 hours.Under these conditions,the leaching rate of aluminum is 90.3%.By comparing the SEM and XRD analysis of raw and clinker materials,it could be concluded that the mullite phase of fly ash is almost completely destroyed and transformed into sodium aluminosilicate with good acid solubility.展开更多
To realize the resource utilization of the valuable metals in the titanium-containing blast furnace slag,the process route of “hydrochloric acid leaching-electrolysis-carbonization and carbon dioxide capture-preparat...To realize the resource utilization of the valuable metals in the titanium-containing blast furnace slag,the process route of “hydrochloric acid leaching-electrolysis-carbonization and carbon dioxide capture-preparation of calcium carbonate” was proposed.In this study,the influences of process conditions on the leaching rates of calcium,magnesium,aluminum,and iron and the phases of the leaching residue were investigated for the leaching process.The experimental results show that the HCl solution could selectively leach the elements from the titanium-containing blast furnace slag.The better leaching conditions are the HCl solution concentration of 4 mol/L,the leaching time of 30 min,the ratio of liquid volume to solid gas of 10 mL/g,and the stirring paddle speed of 300 r/min.Under the conditions,the leaching rates of calcium,magnesium,aluminum,and iron can reach 85.87%,73.41%,81.35%,and 59.08%,and the leaching rate of titanium is 10.71%.The iron and the aluminum are removed from the leachate to obtain iron-aluminum water purification agents,and the magnesium is removed from the leachate to obtain magnesium hydroxide.The leaching residue phase is dominated by perovskite,followed by magnesium silicate and tricalcium aluminate,and the titaniumrich material could be obtained from the leaching residue by desiliconization.展开更多
The leaching behavior of metals from a nickeliferous limonitic laterite ore was investigated by high pressure acid leaching process for the extraction of nickel and cobalt.The effects of sulfuric acid added,leaching t...The leaching behavior of metals from a nickeliferous limonitic laterite ore was investigated by high pressure acid leaching process for the extraction of nickel and cobalt.The effects of sulfuric acid added,leaching temperature,leaching time and liquid/solid（L/S） ratio on metals extraction were examined.More than 97% Ni,96% Co,93% Mn,95% Mg and less than 1% Fe are extracted under optimum conditions.Analysis of the high pressure acid leaching residue by chemical and XRD analysis indicates that the residual iron and sulfur are mainly present in phases of hematite and alunite,respectively.The high pressure leaching process provides a simple and efficient way for the high recovery of nickel and cobalt from laterite ore,leaving residue as a suitable iron resource.展开更多
A method of activation roasting followed by acid leaching using titanium slag was introduced to prepare Ti-rich material. The effects of HaPO4 dosage, roasting temperature, and roasting time on TiO2 grade were investi...A method of activation roasting followed by acid leaching using titanium slag was introduced to prepare Ti-rich material. The effects of HaPO4 dosage, roasting temperature, and roasting time on TiO2 grade were investigated. A Ti-rich material containing 88.54% TiO2, 0.42% （CaO＋MgO） was obtained when finely ground titanium slag was roasted with 7.5% H3PO4 at 1000 ℃ for 2 h, followed by a two-stage leaching in boiling dilute sulfuric acid for 2 h. The XRD patterns show that the product is titanium dioxide with a rutile structure. Mechanism studies show that structures of anosovite solid solution and silicate minerals are destroyed in the roasting process. As a result, titanium components in titanium slag are transformed into TiO2 （futile） while impurities are transformed into acid-soluble phosphate and quartz.展开更多
To analyze the thermodynamic characteristics of leaching process of converter slag, φ-pH diagram of V-Ti-H2O system at oxygen partial pressure of 0.5 MPa, ionic mass concentration of 0.1 mol/kg and temperatures rangi...To analyze the thermodynamic characteristics of leaching process of converter slag, φ-pH diagram of V-Ti-H2O system at oxygen partial pressure of 0.5 MPa, ionic mass concentration of 0.1 mol/kg and temperatures ranging from 60 to 200 ℃ was obtained by recently published critically assessed standard Gibbs energies and activity coefficients of various species. When pH2, stable regions of V3＋, VO2＋ and VO2＋ exist in the stable region of TiO2. The pH values of stable regions of vanadium and titanium decrease and redox potentials become more positive with the temperature increasing. Vanadium and titanium could be separated by one-step leaching based on thermodynamics. The experiment results of pressure acid leaching of converter slag show that leaching rates of vanadium and titanium are 96.87% and 8.76% respectively, at 140 ℃ of temperature, 0.5 MPa of oxygen partial pressure, 0.055-0.075mm of particle size, 15：1 of liquid to solid ratio, 120 min of leaching time, 500 r/min of stirring speed and 200 g/L of initial acid concentration. Vanadium and titanium could be selectively separated in the pressure acid leaching process, and the experiment result is in agreement with thermodynamic calculation result.展开更多
The effects of oxidation-reduction treatment and mechanical activation on the hydrochloric acid leaching performance of Panxi ilmenite concentration were investigated.The results show that both of oxidation-reduction ...The effects of oxidation-reduction treatment and mechanical activation on the hydrochloric acid leaching performance of Panxi ilmenite concentration were investigated.The results show that both of oxidation-reduction treatment and mechanical activation significantly accelerate the extraction of Fe,Ca and Mg from Panxi ilmenite concentration;however,the CaO and MgO contents of the calcined residues obtained from oxidized-reduced ilmenite concentration are higher than the standard values required by chlorination process.The Ca and Mg in oxidized-reduced ilmenite concentration can be leached much faster after mechanical activation,yielding a synthetic rutile which meets the requirements of chlorination process containing 90.50% TiO2 and 1.37% total iron as well as combined CaO and MgO of 1.00%.The optimum oxidation and reduction conditions are as follows：oxidization at 900 ℃ in the presence of oxygen for 15 min and reduction at 750 ℃ by hydrogen for 30 min.展开更多
The leaching processes of zinc plant purification residue in sulfuric acid solution were investigated with respect to the effects of sulfuric acid concentration, reaction temperature, and particle size. A particle siz...The leaching processes of zinc plant purification residue in sulfuric acid solution were investigated with respect to the effects of sulfuric acid concentration, reaction temperature, and particle size. A particle size of 75？80 μm was required to leach 99.8%cobalt and 91.97%zinc at 70 °C for 20 min when the sulfuric acid concentration was 100 g/L and the ratio of liquid to solid was 50？1 （mL/g）. The leaching kinetics of zinc plant purification residue in sulfuric acid solution system conformed well to the shrinking core model, and the dissolution rates of cobalt and zinc were found to be controlled by diffusion through a porous product layer. The apparent activation energy values of cobalt and zinc reaction were calculated to be 11.6931 kJ/mol and 6.6894 kJ/mol, respectively, according to the Arrhenius formula equation. The results show that diffusion through the inert particle pores is the leaching kinetics rate-controlling step.展开更多
Zinc neutral leaching residue（ZNLR） from hydrometallurgical zinc smelting processing can be determined as hazardous intermediate containing considerable amounts of Cd and Zn which have great threats to the environme...Zinc neutral leaching residue（ZNLR） from hydrometallurgical zinc smelting processing can be determined as hazardous intermediate containing considerable amounts of Cd and Zn which have great threats to the environment. The ZNLR contained approximately 35.99% Zn, 15.93% Fe and 0.26% Cd, and Cd mainly existed as ferrites in the ZNLR in this research. Reductive acid leaching of ZNLR was investigated. The effects of hydrazine sulfate concentration, initial sulfuric acid concentration, temperature, duration and liquid-to-solid ratio on the extraction of Cd, Zn and Fe were examined. The extraction efficiencies of Cd, Zn and Fe reached 90.81%, 95.83% and 94.19%, respectively when the leaching parameters were fixed as follows： hydrazine sulfate concentration, 33.3 g/L; sulfuric acid concentration, 80 g/L; temperature, 95 °C; duration of leaching, 120 min; liquid-to-solid ratio, 10 m L/g and agitation, 400 r/min. XRD and SEM-EDS analyses of the leaching residue confirmed that lead sulfate（Pb SO4） and hydrazinium zinc sulfate（（N2H5）2Zn（SO4）2） were the main phases remaining in the reductive leaching residue.展开更多
[Objective] This study was conducted to find a method for rapid determi-nation of Cd in squil a. [Method] Cd in squil a was determined by mixed acid leach-ing-flame atomic absorption method by digesting samples with H...[Objective] This study was conducted to find a method for rapid determi-nation of Cd in squil a. [Method] Cd in squil a was determined by mixed acid leach-ing-flame atomic absorption method by digesting samples with HNO3-HSO4 （5：1） and determining Cd in the samples by a flame atomic absorption spectrophotometer. [Results] The Cd contents in squil a were 1.0-2.0 mg/kg, and the recoveries of Cd were 86%-109%, and the relative standard deviations were 0-6%. [Conclusion] The method is simple and rapid with strong operability, good reproducibility and high ac-curacy.展开更多
To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leac...To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching （roasted Ni-Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 ℃ for 30 min with a L/S ratio of 3 mL/g）, sulphation roasting （51.9% sulfiaric acid addition, roasting temperature 240 ℃ for 1 h）, followed by leaching with the first stage hydrochloric acid leaching solution at 95 ℃ for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.展开更多
A technique including direct acid leaching, vanadium precipitation with alkaline, sodium hydroxide releaching, impurity removing by adjusting pH value, precipitation vanadium with ammonium chloride, and vanadium pento...A technique including direct acid leaching, vanadium precipitation with alkaline, sodium hydroxide releaching, impurity removing by adjusting pH value, precipitation vanadium with ammonium chloride, and vanadium pentoxide by roasting steps was proposed according to the characteristic of Xichuan clay vanadium mineral. The factors influencing leaching vanadium such as temperature and the concentration of sulfuric acid were investigated and optimized. The experimental results indicate that the extract ratios of V205 can reach 94% and 92% at a sodium chlorate ratio of 3% and a manganese dioxide ratio of 3%, respectivdy. A completely chemical precipitation method was adopted to decontaminate and enrich the vanadium in the acid leaching solution. The X-ray diffraction （XRD） pattern and the purity analysis of vanadium pentoxide indicate that the purity of final vanadium pentoxide can reach 99% and meet the standard specifications. The total recovery can reach about 75%. The technique has the characteristics of simplicity, less investment, and more environment safety as compared with the traditional salt roasting method.展开更多
Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various det...Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various determination methods. The effects of leaching time,leaching temperature,leaching agent concentration,leaching L/S ratio,granularity of material,additive consumption were investigated based on the mineralogy.The results show that under the conditions of leaching time of 3-4 h, temperature of 150℃,sulfuric acid consumption of 25%?30%,ratio of liquid to solid of 1.2:1,the granularity less than 0.074 mm, additive consumption of 3%-5%,and oxygen pressure of 1.2 MPa,and the vanadium leaching rate can be more than 92%by the method of two-step pressurized acid leaching.The powdery V2O5 product with 99.52%in V2O5 content is obtained by the flowsheet of acid recovery,removing iron by reduction process,solvent extraction,precipitating vanadium with ammonium water,and pyrolysis from the stone-coal oxygen pressure acid-leaching solution.The total recovery efficiency of vanadium is above 85%,which is more than 20%higher than that obtained in the conventional process.Furthermore,the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.展开更多
Hydrazine sulfate was used as a reducing agent for the leaching of Li,Ni,Co and Mn from spent lithium-ion batteries.The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and ...Hydrazine sulfate was used as a reducing agent for the leaching of Li,Ni,Co and Mn from spent lithium-ion batteries.The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined.97%of the available Li,96%of the available Ni,95%of the available Co,and 86%of the available Mn are extracted under the following optimized conditions:sulfuric acid concentration of 2.0 mol/L,hydrazine sulfate dosage of 30 g/L,solid-to-liquid ratio of 50 g/L,temperature of 80℃,and leaching time of 60 min.The activation energies of the leaching are determined to be 44.32,59.37 and 55.62 k J/mol for Li,Ni and Co,respectively.By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy,it is confirmed that the main phase in the leaching residue is MnO2.The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.展开更多
The effects of the acid leaching and alkali fusion on the leaching efficiency of Y,Eu,Ce,and Tb from the waste rare earth fluorescent powders were investigated in this paper.The results show that hydrochloric acid is ...The effects of the acid leaching and alkali fusion on the leaching efficiency of Y,Eu,Ce,and Tb from the waste rare earth fluorescent powders were investigated in this paper.The results show that hydrochloric acid is better than sulfuric acid in the first acid leaching,and NaOH is better than Na2CO3in the alkali fusion.In the first acid leaching,the Wloss is 20.94%when the waste rare earth fluorescent powders are acid leached in H?concentration 3 mol L-1and S/L ratio 1:3 for 4 h due to red powders dissolved.The better results of the alkali fusion can be got at 800℃ for 2 h when the NaOH is used.The blue powders and the green powders can be dissolved into NaAlO2and oxides such as rare earth oxide(REO).The REO can be dissolved in H?concentration 5 mol L-1,S/L1:10 for 3 h in the second acid leaching.The leaching rates of the Y,Eu,Ce,and Tb are 99.06%,97.38%,98.22%,and 98.15%,respectively.The leaching rate of the total rare earth is 98.60%.展开更多
A new hydrometallurgical route for separation and recovery of Cu from Cu-As-bearing copper electrorefining black slime was developed. The proposed process comprised oxidation acid leaching of Cu-As-bearing slime and s...A new hydrometallurgical route for separation and recovery of Cu from Cu-As-bearing copper electrorefining black slime was developed. The proposed process comprised oxidation acid leaching of Cu-As-bearing slime and selective sulfide precipitation of Cu from the leachate. The effects of various process parameters on the leaching and precipitation of Cu and As were investigated. At the first stage, Cu extraction of 95.2% and As extraction of 97.6% were obtained at 80 ℃ after 4 h with initial H2 SO4 concentration of 1.0 mol/L and liquid-to-solid ratio of 10 mL/g. In addition, the leaching kinetics of Cu and As was successfully reproduced by the Avrami model, and the apparent activation energies were found to be 33.6 and 35.1 kJ/mol for the Cu and As leaching reaction, respectively, suggesting a combination of chemical reaction and diffusion control. During the selective sulfide precipitation, about 99.4% Cu was recovered as CuS, while only 0.1% As was precipitated under the optimal conditions using sulfide-to-copper ratio of 2.4:1, time of 1.5 h and temperature of 25 ℃.展开更多
Through leaching from residue directly and leaching after a roasting treatment,respectively,the experimental research on sulfuric leaching of vanadium from residue of stone coal that came from power generation was con...Through leaching from residue directly and leaching after a roasting treatment,respectively,the experimental research on sulfuric leaching of vanadium from residue of stone coal that came from power generation was conducted.Factors which influence the leaching of vanadium such as concentration of sulfuric acid,leaching temperature,leaching time and liquid-to-solid ratio were investigated in both processes.In the process of direct leaching,to achieve a leaching rate of 74.49%,H2SO4 concentration of up to 5.4%,leaching temperature of 90℃and leaching time of 8 h were necessary reaction factors.The results show that after a roasting treatment at the optimum condition of 950℃at 1 h,76.88%vanadium can be leached under the experimental condition of 0.45% of H2SO4,30℃for 1 h with a liquid-to-solid ratio of 2 mL/g.Leaching after an oxidation roasting treatment is an efficient way to leach vanadium from the residue of stone coal,which has some advantages,such as high recovery,low economic cost and less impurities in leaching solution.展开更多
The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treat...The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in themineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.% and 94.9 wt.%,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.展开更多
A plumbomicrolite concentrate(PMC)was leached with the mixture of HF and H2SO4,HF and HNO3 acids,respectively.Optimal conditions ensuring high recovery of tantalum and niobium(up to 99%)into solution,and radionuclides...A plumbomicrolite concentrate(PMC)was leached with the mixture of HF and H2SO4,HF and HNO3 acids,respectively.Optimal conditions ensuring high recovery of tantalum and niobium(up to 99%)into solution,and radionuclides into insoluble residue were determined.Fluoride-sulfuric acid and fluoride-nitric acid schemes were proposed for PMC leaching by an extractive separation of tantalum form niobium,lead and impurities,and production of high-purity tantalum compounds.Octanol-1 was used as an extractant.Optimal conditions for production of high-purity tantalum strip solutions were defined for all stages(extraction-scrubbing-stripping).Produced tantalum compounds,such as tantalum pentoxide and potassium heptafluotanthalate,comply with the norms for high-purity substances in terms of impurities content.Final choice of the PMC processing scheme is determined by its profitability.展开更多
The chemical and mineral compositions of bauxite recovered from the Severoonezhsk Bauxite Mine(Arkhangelsk region,Russia)were studied by XRD,ICP-OES,TG/DSC,SEM,TEM,and Mössbauer spectroscopy.The iron-containing m...The chemical and mineral compositions of bauxite recovered from the Severoonezhsk Bauxite Mine(Arkhangelsk region,Russia)were studied by XRD,ICP-OES,TG/DSC,SEM,TEM,and Mössbauer spectroscopy.The iron-containing minerals of the bauxites were found to comprise alumogoethite(α-Fe_(1−x)Al_(x)OOH),alumohematite(α-(Fe_(1−x)Al_(x))_(2)O_(3)),alumoakaganeite(β-Fe_(1−x)Al_(x)O(OH,Cl)),and chromite(FeCr_(2)O_(4)).The efficiency of Fe extraction from the bauxite by HCl leaching was 82.5%at 100℃,HCl concentration of 10%,solid/liquid ratio of 1:10,and the process duration of 60 min,with aluminum loss from the bauxites below 4.5%of the total Al contents in the bauxite.Analysis of the kinetics of the iron leaching process proved diffusion to be the limiting stage of the process at 90−100℃.Bauxite residue after leaching presented traces of α-Fe_(1−x)Al_(x)OOH and β-Fe_(1−x)Al_(x)O(OH,Cl),and most of the iron content was in the FeCr_(2)O_(4).In bauxite residue after HCl leaching,in addition to iron oxide,the contents of chromium and calcium oxides significantly decreased.The iron chloride liquor after leaching contained the rare earth elements(REE)of 6.8 mg/L Sc,4.1 mg/L Ce and 2.3 mg/L Ga.展开更多
Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indica...Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.展开更多
基金Funded by the National Natural Science Foundation of China(No.U1710257)the Scientific and Technological Innovation Programs of Higher Education Institutions in Shanxi(No.2019L0656)+2 种基金the Doctoral Research Foundation of Taiyuan University of Science and Technology,China(No.20142001)the Open Foundation Program of Key Laboratory for Ecological Metallurgy of Multimetallic Mineral,Ministry of Education,China(No.2020003)the Supported by Fundamental Research Program of Shanxi Province,China(No.202103021224281)。
文摘By using high-alumina fly ash as raw material,a process was proposed for activating the fly ash with Na_(2)CO_(3)calcination and extracting aluminum from activated clinker with sulfuric acid leaching.The feasibility of roasting process of activated fly ash by Na_(2)CO_(3)was discussed based on thermodynamic analysis.The experimental results showed that Na_(2)CO_(3)gradually reactes with mullite over 700 K to produce NaAlSiO_(4).The optimal process conditions for the activation stage are:a material ratio of 1:1 between sodium carbonate and fly ash,a calcination temperature of 900℃,and a calcination time of 2.5 hours.Under these conditions,the leaching rate of aluminum is 90.3%.By comparing the SEM and XRD analysis of raw and clinker materials,it could be concluded that the mullite phase of fly ash is almost completely destroyed and transformed into sodium aluminosilicate with good acid solubility.
基金Funded by the National Natural Science Foundation of China Youth Fund(No.52204419)the Liaoning Provincial Natural Science Foundation(No.2022-BS-076)the Guangxi Science and Technology Major Project(No.2021AA12013)。
文摘To realize the resource utilization of the valuable metals in the titanium-containing blast furnace slag,the process route of “hydrochloric acid leaching-electrolysis-carbonization and carbon dioxide capture-preparation of calcium carbonate” was proposed.In this study,the influences of process conditions on the leaching rates of calcium,magnesium,aluminum,and iron and the phases of the leaching residue were investigated for the leaching process.The experimental results show that the HCl solution could selectively leach the elements from the titanium-containing blast furnace slag.The better leaching conditions are the HCl solution concentration of 4 mol/L,the leaching time of 30 min,the ratio of liquid volume to solid gas of 10 mL/g,and the stirring paddle speed of 300 r/min.Under the conditions,the leaching rates of calcium,magnesium,aluminum,and iron can reach 85.87%,73.41%,81.35%,and 59.08%,and the leaching rate of titanium is 10.71%.The iron and the aluminum are removed from the leachate to obtain iron-aluminum water purification agents,and the magnesium is removed from the leachate to obtain magnesium hydroxide.The leaching residue phase is dominated by perovskite,followed by magnesium silicate and tricalcium aluminate,and the titaniumrich material could be obtained from the leaching residue by desiliconization.
文摘The leaching behavior of metals from a nickeliferous limonitic laterite ore was investigated by high pressure acid leaching process for the extraction of nickel and cobalt.The effects of sulfuric acid added,leaching temperature,leaching time and liquid/solid（L/S） ratio on metals extraction were examined.More than 97% Ni,96% Co,93% Mn,95% Mg and less than 1% Fe are extracted under optimum conditions.Analysis of the high pressure acid leaching residue by chemical and XRD analysis indicates that the residual iron and sulfur are mainly present in phases of hematite and alunite,respectively.The high pressure leaching process provides a simple and efficient way for the high recovery of nickel and cobalt from laterite ore,leaving residue as a suitable iron resource.
基金Project(NCET-10-0834) supported by the Program for New Century Excellent Talents in University,China
文摘A method of activation roasting followed by acid leaching using titanium slag was introduced to prepare Ti-rich material. The effects of HaPO4 dosage, roasting temperature, and roasting time on TiO2 grade were investigated. A Ti-rich material containing 88.54% TiO2, 0.42% （CaO＋MgO） was obtained when finely ground titanium slag was roasted with 7.5% H3PO4 at 1000 ℃ for 2 h, followed by a two-stage leaching in boiling dilute sulfuric acid for 2 h. The XRD patterns show that the product is titanium dioxide with a rutile structure. Mechanism studies show that structures of anosovite solid solution and silicate minerals are destroyed in the roasting process. As a result, titanium components in titanium slag are transformed into TiO2 （futile） while impurities are transformed into acid-soluble phosphate and quartz.
基金Project(2007CB613504)supported by the National Key Basic Research Program of ChinaProjects(51004033,50974035,51074047)supported by the National Natural Science Foundation of ChinaProject(2008BAB34B01)supported by National Science and Technology Support Plan of China during the 11th Five-Year Plan
文摘To analyze the thermodynamic characteristics of leaching process of converter slag, φ-pH diagram of V-Ti-H2O system at oxygen partial pressure of 0.5 MPa, ionic mass concentration of 0.1 mol/kg and temperatures ranging from 60 to 200 ℃ was obtained by recently published critically assessed standard Gibbs energies and activity coefficients of various species. When pH2, stable regions of V3＋, VO2＋ and VO2＋ exist in the stable region of TiO2. The pH values of stable regions of vanadium and titanium decrease and redox potentials become more positive with the temperature increasing. Vanadium and titanium could be separated by one-step leaching based on thermodynamics. The experiment results of pressure acid leaching of converter slag show that leaching rates of vanadium and titanium are 96.87% and 8.76% respectively, at 140 ℃ of temperature, 0.5 MPa of oxygen partial pressure, 0.055-0.075mm of particle size, 15：1 of liquid to solid ratio, 120 min of leaching time, 500 r/min of stirring speed and 200 g/L of initial acid concentration. Vanadium and titanium could be selectively separated in the pressure acid leaching process, and the experiment result is in agreement with thermodynamic calculation result.
基金Project(2009FJ3082)supported by Research Project of Science and Technology in Hunan Province,ChinaProject(2007CB613606)supported by the National Basic Research Program of China
文摘The effects of oxidation-reduction treatment and mechanical activation on the hydrochloric acid leaching performance of Panxi ilmenite concentration were investigated.The results show that both of oxidation-reduction treatment and mechanical activation significantly accelerate the extraction of Fe,Ca and Mg from Panxi ilmenite concentration;however,the CaO and MgO contents of the calcined residues obtained from oxidized-reduced ilmenite concentration are higher than the standard values required by chlorination process.The Ca and Mg in oxidized-reduced ilmenite concentration can be leached much faster after mechanical activation,yielding a synthetic rutile which meets the requirements of chlorination process containing 90.50% TiO2 and 1.37% total iron as well as combined CaO and MgO of 1.00%.The optimum oxidation and reduction conditions are as follows：oxidization at 900 ℃ in the presence of oxygen for 15 min and reduction at 750 ℃ by hydrogen for 30 min.
基金Project(51072233)supported by the National Natural Science Foundation of China
文摘The leaching processes of zinc plant purification residue in sulfuric acid solution were investigated with respect to the effects of sulfuric acid concentration, reaction temperature, and particle size. A particle size of 75？80 μm was required to leach 99.8%cobalt and 91.97%zinc at 70 °C for 20 min when the sulfuric acid concentration was 100 g/L and the ratio of liquid to solid was 50？1 （mL/g）. The leaching kinetics of zinc plant purification residue in sulfuric acid solution system conformed well to the shrinking core model, and the dissolution rates of cobalt and zinc were found to be controlled by diffusion through a porous product layer. The apparent activation energy values of cobalt and zinc reaction were calculated to be 11.6931 kJ/mol and 6.6894 kJ/mol, respectively, according to the Arrhenius formula equation. The results show that diffusion through the inert particle pores is the leaching kinetics rate-controlling step.
基金Project(2012FJ1010)supported by the Key Project of Science and Technology of Hunan ProvinceChina+2 种基金Project(51474247)supported by the National Natural Science Foundation of ChinaProject(2012GS430201)supported by the Science and Technology Program for Public WellbeingChina
文摘Zinc neutral leaching residue（ZNLR） from hydrometallurgical zinc smelting processing can be determined as hazardous intermediate containing considerable amounts of Cd and Zn which have great threats to the environment. The ZNLR contained approximately 35.99% Zn, 15.93% Fe and 0.26% Cd, and Cd mainly existed as ferrites in the ZNLR in this research. Reductive acid leaching of ZNLR was investigated. The effects of hydrazine sulfate concentration, initial sulfuric acid concentration, temperature, duration and liquid-to-solid ratio on the extraction of Cd, Zn and Fe were examined. The extraction efficiencies of Cd, Zn and Fe reached 90.81%, 95.83% and 94.19%, respectively when the leaching parameters were fixed as follows： hydrazine sulfate concentration, 33.3 g/L; sulfuric acid concentration, 80 g/L; temperature, 95 °C; duration of leaching, 120 min; liquid-to-solid ratio, 10 m L/g and agitation, 400 r/min. XRD and SEM-EDS analyses of the leaching residue confirmed that lead sulfate（Pb SO4） and hydrazinium zinc sulfate（（N2H5）2Zn（SO4）2） were the main phases remaining in the reductive leaching residue.
基金Supported by Science and Technology Planning Project of Zhejiang Province(2015C32001)Special Fund for Research Institutes from Science and Technology Department of Zhejiang Province(2016F30021)Natural Science Foundation of China(21407127)~~
文摘[Objective] This study was conducted to find a method for rapid determi-nation of Cd in squil a. [Method] Cd in squil a was determined by mixed acid leach-ing-flame atomic absorption method by digesting samples with HNO3-HSO4 （5：1） and determining Cd in the samples by a flame atomic absorption spectrophotometer. [Results] The Cd contents in squil a were 1.0-2.0 mg/kg, and the recoveries of Cd were 86%-109%, and the relative standard deviations were 0-6%. [Conclusion] The method is simple and rapid with strong operability, good reproducibility and high ac-curacy.
基金Project(51104186)supported by the National Natural Science Foundation of ChinaProjects(2016zzts282,2016zzts283)supported by the Fundamental Research Funds for the Central Universities of Central South University,China
文摘To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching （roasted Ni-Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 ℃ for 30 min with a L/S ratio of 3 mL/g）, sulphation roasting （51.9% sulfiaric acid addition, roasting temperature 240 ℃ for 1 h）, followed by leaching with the first stage hydrochloric acid leaching solution at 95 ℃ for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.
基金This research was financially supported by the National Natural Science Foundation of China (No.20576137)the Major State Basic Research Development Program of China (No.2003CB716001).
文摘A technique including direct acid leaching, vanadium precipitation with alkaline, sodium hydroxide releaching, impurity removing by adjusting pH value, precipitation vanadium with ammonium chloride, and vanadium pentoxide by roasting steps was proposed according to the characteristic of Xichuan clay vanadium mineral. The factors influencing leaching vanadium such as temperature and the concentration of sulfuric acid were investigated and optimized. The experimental results indicate that the extract ratios of V205 can reach 94% and 92% at a sodium chlorate ratio of 3% and a manganese dioxide ratio of 3%, respectivdy. A completely chemical precipitation method was adopted to decontaminate and enrich the vanadium in the acid leaching solution. The X-ray diffraction （XRD） pattern and the purity analysis of vanadium pentoxide indicate that the purity of final vanadium pentoxide can reach 99% and meet the standard specifications. The total recovery can reach about 75%. The technique has the characteristics of simplicity, less investment, and more environment safety as compared with the traditional salt roasting method.
基金Project(2006AA06Z130)supported by the High-tech Research and Development Program of ChinaProject(50874053)supported by the National Natural Science Foundation of ChinaProject(2007GA010)supported by Science and Technology Bureau of Yunnan Province,China
文摘Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various determination methods. The effects of leaching time,leaching temperature,leaching agent concentration,leaching L/S ratio,granularity of material,additive consumption were investigated based on the mineralogy.The results show that under the conditions of leaching time of 3-4 h, temperature of 150℃,sulfuric acid consumption of 25%?30%,ratio of liquid to solid of 1.2:1,the granularity less than 0.074 mm, additive consumption of 3%-5%,and oxygen pressure of 1.2 MPa,and the vanadium leaching rate can be more than 92%by the method of two-step pressurized acid leaching.The powdery V2O5 product with 99.52%in V2O5 content is obtained by the flowsheet of acid recovery,removing iron by reduction process,solvent extraction,precipitating vanadium with ammonium water,and pyrolysis from the stone-coal oxygen pressure acid-leaching solution.The total recovery efficiency of vanadium is above 85%,which is more than 20%higher than that obtained in the conventional process.Furthermore,the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.
基金Project(51674298)supported by the National Natural Science Foundation of ChinaProject supported by Anhui Province Research and Development Innovation Program,China。
文摘Hydrazine sulfate was used as a reducing agent for the leaching of Li,Ni,Co and Mn from spent lithium-ion batteries.The effects of the reaction conditions on the leaching mechanism and kinetics were characterized and examined.97%of the available Li,96%of the available Ni,95%of the available Co,and 86%of the available Mn are extracted under the following optimized conditions:sulfuric acid concentration of 2.0 mol/L,hydrazine sulfate dosage of 30 g/L,solid-to-liquid ratio of 50 g/L,temperature of 80℃,and leaching time of 60 min.The activation energies of the leaching are determined to be 44.32,59.37 and 55.62 k J/mol for Li,Ni and Co,respectively.By performing X-ray diffraction and scanning electron microscopy in conjunction with energy dispersive X-ray spectroscopy,it is confirmed that the main phase in the leaching residue is MnO2.The results show that hydrazine sulfate is an effective reducing agent in the acid leaching process for spent lithium-ion batteries.
基金supported by the National Hi-Tech R&D Program of China (No. 2012AA063202)National Key Project of Scientific and Technical Support Program of China (Nos. 2011BAE13B07, 2012BAC02B01, and 2011BAC10B02)National Natural Science Foundation of China (Nos. 51174247 and 50972013)
文摘The effects of the acid leaching and alkali fusion on the leaching efficiency of Y,Eu,Ce,and Tb from the waste rare earth fluorescent powders were investigated in this paper.The results show that hydrochloric acid is better than sulfuric acid in the first acid leaching,and NaOH is better than Na2CO3in the alkali fusion.In the first acid leaching,the Wloss is 20.94%when the waste rare earth fluorescent powders are acid leached in H?concentration 3 mol L-1and S/L ratio 1:3 for 4 h due to red powders dissolved.The better results of the alkali fusion can be got at 800℃ for 2 h when the NaOH is used.The blue powders and the green powders can be dissolved into NaAlO2and oxides such as rare earth oxide(REO).The REO can be dissolved in H?concentration 5 mol L-1,S/L1:10 for 3 h in the second acid leaching.The leaching rates of the Y,Eu,Ce,and Tb are 99.06%,97.38%,98.22%,and 98.15%,respectively.The leaching rate of the total rare earth is 98.60%.
基金financial supports from the National Natural Science Foundation of China (51634010,51904354)the National Science Fund for Distinguished Young Scholars of China (51825403)+1 种基金the National Key R&D Program of China (2018YFC1900306,2019YFC1907405)Key Research and Development Program of Hunan Province,China (2019SK2291)。
文摘A new hydrometallurgical route for separation and recovery of Cu from Cu-As-bearing copper electrorefining black slime was developed. The proposed process comprised oxidation acid leaching of Cu-As-bearing slime and selective sulfide precipitation of Cu from the leachate. The effects of various process parameters on the leaching and precipitation of Cu and As were investigated. At the first stage, Cu extraction of 95.2% and As extraction of 97.6% were obtained at 80 ℃ after 4 h with initial H2 SO4 concentration of 1.0 mol/L and liquid-to-solid ratio of 10 mL/g. In addition, the leaching kinetics of Cu and As was successfully reproduced by the Avrami model, and the apparent activation energies were found to be 33.6 and 35.1 kJ/mol for the Cu and As leaching reaction, respectively, suggesting a combination of chemical reaction and diffusion control. During the selective sulfide precipitation, about 99.4% Cu was recovered as CuS, while only 0.1% As was precipitated under the optimal conditions using sulfide-to-copper ratio of 2.4:1, time of 1.5 h and temperature of 25 ℃.
基金Project(50974133)supported by the National Natural Science of Chinaproject(08zxgk06)supported by the Opening Foundation of Key Laboratory of Solid Waste Treatment and Resource Recycle(SWUST),Ministry of Education,China
文摘Through leaching from residue directly and leaching after a roasting treatment,respectively,the experimental research on sulfuric leaching of vanadium from residue of stone coal that came from power generation was conducted.Factors which influence the leaching of vanadium such as concentration of sulfuric acid,leaching temperature,leaching time and liquid-to-solid ratio were investigated in both processes.In the process of direct leaching,to achieve a leaching rate of 74.49%,H2SO4 concentration of up to 5.4%,leaching temperature of 90℃and leaching time of 8 h were necessary reaction factors.The results show that after a roasting treatment at the optimum condition of 950℃at 1 h,76.88%vanadium can be leached under the experimental condition of 0.45% of H2SO4,30℃for 1 h with a liquid-to-solid ratio of 2 mL/g.Leaching after an oxidation roasting treatment is an efficient way to leach vanadium from the residue of stone coal,which has some advantages,such as high recovery,low economic cost and less impurities in leaching solution.
基金the 11th Five-Year Plan of National Scientific and Technological Program of China (No.2007 BAB22B01)
文摘The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The pri-mary copper and leadminerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in themineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.% and 94.9 wt.%,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.
基金Project supported by the Federal Research Centre of Kola Science Centre of the Russian Academy of Sciences,Russian。
文摘A plumbomicrolite concentrate(PMC)was leached with the mixture of HF and H2SO4,HF and HNO3 acids,respectively.Optimal conditions ensuring high recovery of tantalum and niobium(up to 99%)into solution,and radionuclides into insoluble residue were determined.Fluoride-sulfuric acid and fluoride-nitric acid schemes were proposed for PMC leaching by an extractive separation of tantalum form niobium,lead and impurities,and production of high-purity tantalum compounds.Octanol-1 was used as an extractant.Optimal conditions for production of high-purity tantalum strip solutions were defined for all stages(extraction-scrubbing-stripping).Produced tantalum compounds,such as tantalum pentoxide and potassium heptafluotanthalate,comply with the norms for high-purity substances in terms of impurities content.Final choice of the PMC processing scheme is determined by its profitability.
基金Ministry of Science and Higher Education of the Russian Federation(scientific topic No.0137-2019-0023).
文摘The chemical and mineral compositions of bauxite recovered from the Severoonezhsk Bauxite Mine(Arkhangelsk region,Russia)were studied by XRD,ICP-OES,TG/DSC,SEM,TEM,and Mössbauer spectroscopy.The iron-containing minerals of the bauxites were found to comprise alumogoethite(α-Fe_(1−x)Al_(x)OOH),alumohematite(α-(Fe_(1−x)Al_(x))_(2)O_(3)),alumoakaganeite(β-Fe_(1−x)Al_(x)O(OH,Cl)),and chromite(FeCr_(2)O_(4)).The efficiency of Fe extraction from the bauxite by HCl leaching was 82.5%at 100℃,HCl concentration of 10%,solid/liquid ratio of 1:10,and the process duration of 60 min,with aluminum loss from the bauxites below 4.5%of the total Al contents in the bauxite.Analysis of the kinetics of the iron leaching process proved diffusion to be the limiting stage of the process at 90−100℃.Bauxite residue after leaching presented traces of α-Fe_(1−x)Al_(x)OOH and β-Fe_(1−x)Al_(x)O(OH,Cl),and most of the iron content was in the FeCr_(2)O_(4).In bauxite residue after HCl leaching,in addition to iron oxide,the contents of chromium and calcium oxides significantly decreased.The iron chloride liquor after leaching contained the rare earth elements(REE)of 6.8 mg/L Sc,4.1 mg/L Ce and 2.3 mg/L Ga.
基金Project(20050145029)supported by the Research Fund for the Doctoral Program of Higher Education of ChinaProject(2005221012)supported by Science and Technology Talents Fund for Excellent Youth of Liaoning Province,China
文摘Effects of particle size of the zinc sulfide concentrate,leaching temperature,solid-to-liquid ratio and additive amount on pressure acid leaching process of the zinc sulfide concentrate were studied.The results indicate that the additive can improve the reaction kinetics and the conversion rate.And sulfur can be successfully separated from the zinc sulfide concentrate as elemental sulfur.The reasonable experiment parameters are obtained as follows:the leaching temperature 150℃,oxygen partial pressure 1 MPa,additive amount 1%,solid-to-liquid ratio 1:4,leaching time 2 h,initial sulfuric acid concentration 15%,and particle size less than 44μm.Under the optimum conditions,the leaching rate of the zinc can reach 95%and the reduction rate of the sulfur can reach 90%.
