采用硫酸-还原剂浸出工艺从废旧锂离子电池中回收LiNi_(0.6)Mn_(0.2)Co_(0.2)O_2

采用H_2SO_4-还原剂浸出工艺处理废旧锂离子电池正极材料(LiNi_(0.6)Mn_(0.2)Co_(0.2)O_2),研究了H_2SO_4浓度以及浸出温度对有价金属元素浸出的影响,确定了浸出过程中适宜的H_2SO_4浓度为2 mol/L,浸出温度为40℃。在H_2SO_4浓度为2mol/L、原料与浸出剂比例为100g/L、浸出时间为2h、浸出温度为40℃、搅拌速度为500r/min的优化条件下,通过单因素实验考察了还原剂H_2O_2、C_6H_(12)O_6、Na_2SO_3对有价金属浸出的影响。结果显示,还原剂H_2O_2、C_6H_(12)O_6、Na_2SO_3的最佳添加量分别为4.5%、80g/L、60g/L。通过正交实验考察了混合还原剂的影响,结果表明当还原剂组成为120g/L C_6H_(12)O_6和100g/L Na_2SO_3时,Co、Li、Mn、Ni浸出率分别为93.51%、92.68%、95.61%、92.93%,Al浸出率仅达到18.57%。与单个还原剂相比,通过改变混合还原剂的组成,可在不明显降低有价金属(Li、Ni、Mn、Co)浸出率的情况下,控制杂质金属Al的浸出。另外,对于Al、Li的浸出,影响因素的主次关系分别为:Na_2SO_3>C_6H_(12)O_6>H_2O_2、H_2O_2>Na_2SO_3>C_6H_(12)O_6;对于Co、Mn、Ni的浸出,影响因素的主次关系为C_6H_(12)O_6>H_2O_2>Na_2SO_3。

软锰矿二次还原浸出制备软磁铁氧体用碳酸锰

采用"二次还原浸出-净化脱杂-碳酸铵沉锰"工艺,对云南某软锰矿制备软磁铁氧化体用碳酸锰工艺条件进行了研究。考察了还原剂用量、酸度对浸出率的影响以及浸出液的净化除杂条件、净化后溶液的沉锰条件等工艺参数对产品质量的影响,并获得了最佳工艺参数。结果表明,在m(MnO2):m(FeS2):m(H2SO4):m(Fe)=1:0.35:1.5:0.07时,锰的浸出率大于99%,碳酸锰总回收率为86%。产品质量符合HG/T 2836-1997《软磁铁氧体用碳酸锰》标准要求。

钛渣物相组成对其改性浸出的影响

氧化-还原-酸浸工艺是处理攀西钛渣最具工业应用潜力的工艺。以攀西钛铁矿为原料,采用还原熔炼工艺制备不同Fe/Mg(摩尔比)的钛渣,系统研究了Fe/Mg对钛渣氧化还原改性-盐酸浸出升级工艺的影响。结果表明,1 000℃高温氧化时,Fe/Mg对钛渣的氧化率基本无影响;800℃低温氧化时,降低Fe/Mg可使钛渣的氧化速率降低,且氧化不完全。在还原过程中,Fe/Mg对Fe^(3+)还原为Fe^(2+)的还原速率无明显影响,但对其物相转变影响显著,Fe/Mg<1.9的钛渣还原后仍存在大量(M_(3)O_(5))^(3)相,使其酸浸除杂后无法达到氯化钛白粉的质量要求。

Manganese extraction by reduction-acid leaching from low-grade manganese oxide ores using CaS as reductant

The extraction of manganese from low-grade manganese oxide ores using Ca S derived from Ca SO4 as reductant was investigated. The effects of mass ratio of Ca S to ore, reduction temperature, reduction time, liquid to solid ratio（L/S ratio）, stirring speed, leaching temperature, leaching time and H2SO4 concentration on the leaching rates of Mn and Fe during the reduction–acid leaching process were discussed. The leaching rates of 96.47% for Mn and 19.24% for Fe were obtained under the optimized conditions of mass ratio of Ca S to manganese oxide ore 1：6.7, L/S ratio 5：1, stirring speed 300 r/min, reduction temperature of 95 °C for 2.0 h in the reduction process and leaching stirring speed of 200 r/min, H2SO4 concentration of 1.5 mol/L, leaching temperature of 80 °C for 5 min in the leaching process. In addition, this process can be employed in the recovery of manganese from various manganese oxide ores, and Mn leaching rate above 95% is obtained.

Effects of mechanical activation and oxidation-reduction on hydrochloric acid leaching of Panxi ilmenite concentration

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.

Manganese extraction from high-iron-content manganese oxide ores by selective reduction roasting-acid leaching process using black charcoal as reductant

Reduction roasting-acid leaching process was utilized to process high-iron-content manganese oxide ore using black charcoal as reductant. The results indicate that, compared with the traditional reductant of anthracite, higher manganese extraction efficiency is achieved at lower roasting temperature and shorter residence time. The effects of roasting parameters on the leaching efficiency of Mn and Fe were studied, and the optimal parameters are determined as follows: roasting temperature is 650 °C, residence time is 40 min, and black charcoal dosage is 10%(mass fraction). Under these conditions, the leaching efficiency of Mn reaches 82.37% while that of Fe is controlled below 7%. XRD results show that a majority of MnO2 and Fe2O3 in the raw ore are reduced to MnO and Fe3O4, respectively.

Priority recovery of lithium and effective leaching of nickel and cobalt from spent lithium-ion battery

The cathode materials of spent lithium-ion batteries(LIBs)were recovered via reductive roasting,Na_(2)CO_(3) leaching,and ammonia leaching.The effects of roasting parameters,Na_(2)CO_(3) leaching parameters,and ammonia leaching parameters on the leaching efficiencies of metals were explored.The results show that the mineral phase of spent LIBs is reconstructed during reductive roasting,and more than 99%of Li can be preferentially leached via Na_(2)CO_(3) leaching.Ni(99.7%)and Co(99.9%)can be leached via one-step ammonia leaching,and Mn cannot be leached.Thus,good leaching selectivity is achieved.The kinetic study shows that the leaching of Ni and Co conforms to chemical reaction control.

Reductive acid leaching of valuable metals from spent lithium-ion batteries using hydrazine sulfate as reductant

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.

Kinetics of Reductive Leaching of Low-grade Pyrolusite with Molasses Alcohol Wastewater in H_2SO_4

The kinetics of reductive leaching of manganese from low grade pyrolusite in dilute sulfuric acid in the presence of molasses alcohol wastewater was investigated. The shrinking core model was applied to quantify the effects of reaction parameters on leaching rate. The leaching rate increases with reaction temperature, concentrations of H 2 SO 4 and organic matter in molasses alcohol wastewater increase and ore particle size decreases. The leaching process follows the kinetics of a shrinking core model and the apparent activation energy is 57.5 kJ·mol –1 . The experimental results indicate a reaction order of 0.52 for H2SO4 concentration and 0.90 for chemical oxygen demand (COD) of molasses alcohol wastewater. It is concluded that the reductive leaching of pyrolusite with molasses alcohol wastewater is controlled by the diffusion through the ash/inert layer composed of the associated minerals.

Kinetics and Modeling of Chemical Leaching of Sphalerite Concentrate Using Ferric Iron in a Redox-controlled Reactor

This work presents a study for chemical leaching of sphalerite concentrate under various constant Fe3+ concentrations and redox potential conditions. The effects of Fe3+ concentration and redox potential on chemical leaching of sphalerite were investigated. The shrinking core model was applied to analyze the experimental results. It was found that both the Fe3+ concentration and the redox potential controlled the chemical leaching rate of sphalerite. A new kinetic model was developed, in which the chemical leaching rate of sphalerite was proportional to Fe3+ concentration and Fe3+ /Fe2+ ratio. All the model parameters were evaluated from the experimental data. The model predictions fit well with the experimental observed values.

Extraction of valuable metals from low-grade nickeliferous laterite ore by reduction roasting-ammonia leaching method

Nickel and cobalt were extracted from low-grade nickeliferous laterite ore using a reduction roasting-ammonia leaching method.The reduction roasting-ammonia leaching experimental tests were chiefly introduced,by which fine coal was used as a reductant.The results show that the optimum process conditions are confirmed as follows:in reduction roasting process,the mass fraction of reductant in the ore is 10%,roasting time is 120 min,roasting temperature is 1 023-1 073 K;in ammonia leaching process,the liquid-to-solid ratio is 4:1(mL/g),leaching temperature is 313 K,leaching time is 120 min,and concentration ratio of NH3 to CO2 is 90 g/L:60 g/L.Under the optimum conditions,leaching efficiencies of nickel and cobalt are 86.25% and 60.84%,respectively.Therefore,nickel and cobalt can be effectively reclaimed,and the leaching agent can be also recycled at room temperature and normal pressure.

Stepwise recovery of magnesium from low-grade ludwigite ore based on innovative and clean technological route

A novel and clean technological route for the comprehensive utilization of low-grade ludwigite ore was proposed, inwhich magnesium was extracted by metallizing reduction？magnetic separation, sulfuric acid leaching and ethanol precipitationoperation. Meanwhile, iron-rich product, silicon-rich product and boron-rich product were obtained, respectively. In the process ofmetallizing reduction-magnetic separation, 94.6% of magnesium was enriched in the non-magnetic substance from the ore reducedat 1250 ℃ for 60 min with the ore size of 0.50-2.00 mm and coal size of 0.50-1.50 mm. When the non-magnetic substance wasleached at 90 ℃ for 15 min with the liquid-to-solid ratio of 7：1, 87.4% of magnesium was leached into the liquor separated fromsilicon gathering in leaching residue. The ethanol precipitation was conducted for 30 min with the ethanol-to-original liquid volumeratio of 1.5：1 at room temperature. 97.2% of magnesium was precipitated out with the initial concentration of 0.8 mol/L in the formof MgSO4·7H2O.

Reductive leaching of manganese oxide ores using waste tea as reductant in sulfuric acid solution

Manganese oxide ores from Gabon and Xiangxi were leached with waste tea as reductant in dilute sulfuric acid solution. The effects of waste tea dosage, concentration of sulfuric acid, liquid-to-solid ratio, leaching temperature and reaction time on leaching process were explored. The leaching efficiency of Gabonese manganese oxide ore reached almost 100% under the optimal condition which was determined as follows： manganese oxide ore to waste tea mass ratio of 10：1, sulfuric acid concentration of 2.5 molFL, liquid-to-solid ratio of 7.5：1, leaching temperature of 368 K, time of 8 h. The leaching efficiency of Xiangxi manganese oxide ore reached 99.8% under the optimal condition which was determined as follows： manganese oxide ore to waste tea mass ratio of 10：1, sulfuric acid concentration of 1.7 mol/L, liquid-to-solid ratio of 7.5：1, leaching temperature of 368 K, time of 8 h. The leaching process followed the internal diffusion controlled kinetic model, and the apparent activation energies of Gabonese manganese oxide ore and Xiangxi manganese oxide ore were calculated to be 38.2 kJ/mol and 20.4 kJ/mol, respectively. The morphological changes and mineralogical forms of the ore before and after the chemical treatment were discussed with the support of XRD analysis and SEM analysis.

Reductive leaching of zinc and indium from industrial zinc ferrite particulates in sulphuric acid media

Zinc ferrite is the principal constituent in zinc neutral-leach residue（NLR） which is commonly treated by hot-acid leaching in electrolytic zinc plants. Reductive leaching of zinc ferrite with sphalerite concentrate as a reducing agent was performed. It was found that leaching of zinc ferrite in the presence of sphalerit concentrate was a viable process that effectively extracted zinc and indium and converted Fe^3＋ into Fe^2＋ at the same time. Reflux leaching tests by two stages were performed to achieve extractions of 98.1% for zinc and 97.5% for indium, and a Fe^2＋/Fe^3＋ molar ratio of 9.6 in leach solution was also obtained. The leaching behaviors of other elements, such as iron, copper and tin were also studied. The results showed that iron and copper were completely leached, whereas tin presented lower extraction values.

TiO_2 nanoparticles prepared by hydrochloric acid leaching of mechanically activated and carbothermic reduced ilmenite

A new combination method consisting of ball milling, carbothermic reduction and hydrochloric acid leaching was proposed for the preparation of nanosized synthetic rutile from natural ilmenite. The ball milling was employed to grind ilmenite into small particles. The carbothermic reduction was carried out to yield a high titanium slag, which would be easily purified by subsequent leaching procedure. Factors affecting the hydrochloric acid process, namely the leaching time, temperature, and acid concentration, were studied. After leaching and calcining the milled and annealed mixture of FeTiO3/C under the optimal conditions, the TiO2 nanoparticles with size of 10-200 nm and purity〉98.0% were obtained.

Selective reduction process of zinc ferrite and its application in treatment of zinc leaching residues

The traditional zinc hydro-metallurgy generates a large amount of zinc ferrite residue rich in valuable metals. The separation of iron is crucial for resource recycling of valuable metals in zinc ferrite residue. A novel selective reduction roasting？leaching process was proposed to separate zinc and iron from zinc leaching residue which contains zinc ferrite. The thermodynamic analysis was employed to determine the predominant range of Fe3O4 and ZnO during reduction roasting process of zinc ferrite. Based on the result of thermodynamic calculation, we found thatV（CO）/V（CO＋CO2） ratio is a key factor determining the phase composition in the reduction roasting product of zinc ferrite. In the range ofV（CO）/V（CO＋CO2） ratio between 2.68% and 36.18%, zinc ferrite is preferentially decomposed into Fe3O4 and ZnO. Based on thermogravimetric （TG） analysis, the optimal conditions for reduction roasting of zinc ferrite are determined as follows： temperature 700？750 °C, volume fraction of CO 6% and V（CO）/V（CO＋CO2） ratio 30%. Based on the above results, zinc leaching residue rich in zinc ferrite was roasted and the roasted product was leached by acid solution. It is found that zinc extraction rate in zinc leaching residue reaches up to 70% and iron extraction rate is only 18.4%. The result indicates that zinc and iron can be effectively separated from zinc leaching residue.

Preparation of high pure tellurium from raw tellurium containing Cu and Se by chemical method

High pure tellurium was prepared from raw tellurium containing copper and selenium by chemical method containing oxidation with concentrated nitric acid, leaching with hydrochloric acid, reducing with sulfur dioxide and treating in hydrogen atmosphere at high temperature. Removal ratio of Cu in raw tellurium reaches 99% after raw tellurium is oxidized and leached with HNO3（69%） under the following conditions： 0.96 times stoichiometric quantity of concentrated nitric acid, 4：1 of ratio of liquid to solid, 20 °C of reaction temperature and 30 min of reaction time. Leaching ratio of Te reaches 99% after Te is leached with hydrochloric acid under the following conditions： 1.67 times stoichiometric quantity of hydrochloric acid, 4：1 of ratio of liquid to solid, 20 °C of reaction temperature and 30 min of reaction time. Tellurium powder（99.95%） is obtained when Te（IV） in leachate is reduced with sulfur dioxide. The purity of tellurium increases from 99.954% to 99.999 6% after tellurium（99.95%） is treated in hydrogen atmosphere at 723.15 K for 30 min.

Relationship and effect of redox potential,jarosites and extracellular polymeric substances in bioleaching chalcopyrite by acidithiobacillus ferrooxidans

The changes of pH,redox potential,concentrations of soluble iron ions and Cu^2＋ with the time of bioleaching chalcopyrite concentrates by acidithiobacillus ferrooxidans were investigated under the different conditions of initial total-iron amount as well as mole ratio of Fe（III） to Fe（II） in the solutions containing synthetic extracellular polymeric substances （EPS）.When the solution potential is lower than 650 mV （vs SHE）,the inhibition of jarosites to bioleaching chalcopyrite is not vital as EPS produced by bacteria can retard the contamination through flocculating jarosites even if concentration of Fe（III） ions is up to 20 g/L but increases with increasing the concentration of Fe（III） ions;jarosites formed by bio-oxidized Fe3＋ ions are more easy to adhere to outside surface of EPS space on chalcopyrite;the EPS layer with jarosites acts as a weak diffusion barrier to further rapidly create a high redox potential of more than 650 mV by bio-oxidizing Fe^2＋ ions inside and outside EPS space into Fe^3＋ ions,resulting in a rapid deterioration of ion diffusion performance of the EPS layer to inhibit bioleaching chalcopyrite severely and irreversibly.

Leaching of chalcopyrite: An emphasis on effect of copper and iron ions

Many researchers found that the Fe2+together with less amount of Cu2+can accelerate the leaching of chalcopyrite.In this work,the leaching of chalcopyrite with Cu2+was investigated.The leaching residuals were examined by Raman spectroscopy.Based on the leaching experiments,the chemical equilibrium in solution was calculated using Visual MINTEQ.The results showed that the Fe in chalcopyrite lattice was replaced by Cu2+;therefore,the chalcopyrite transformed into covellite.Furthermore,the formation of chalcocite occurred when Fe2+and Fe3+were added to the solution containing Cu2+.The copper extraction increased with a decrease of the initial redox potential(or the ratio of Fe3+/Fe2+).

Stepwise dissolution of chalcopyrite bioleaching by thermophile A.manzaensis and mesophile L. ferriphilum

Chalcopyrite dissolution was evaluated by bioleaching and electrochemical experiments with thermophile A. manzaensis（Acidianus manzaensis） and mesophile L. ferriphilum（Leptospirillum ferriphium） cultures at 65 ℃ and 40 ℃, respectively. It was investigated that the bioleaching of chalcopyrite was stepwise. It was reduced to Cu2 S at a lower redox potential locating in the whole bioleaching process by A. manzaensis at high temperature while only at initial days of bioleaching by L. ferriphilum at a relative low temperature. No reduced product was detected when the redox potential was beyond a high level（e.g., 550 m V（vs SCE）） bioleached by L. ferriphilum. Chalcopyrite bioleaching efficiency was substantially improved bioleached by A. manaensis compared to that by L. ferriphilum, which was mainly attributed to the reduction reaction occurring during bioleaching. The reductive intermediate Cu2 S was more amenable to oxidation than chalcopyrite, causing enhanced copper extraction.