Study on metal recovery process and kinetics of oxidative leaching from spent LiFePO_(4)Li-batteries

A green environmental protection and enhanced leaching process was proposed to recover all elements from spent lithium iron phosphate(LiFePO_(4)) lithium batteries.In order to reduce the influence of Al impurity in the recovery process,NaOH was used to remove impurity.After impurity removal,the spent LiFePO_(4) cathode material was used as raw material under the H_(2)SO_(4) system,and the pressure oxidation leaching process was adopted to achieve the preferential leaching of lithium.The E-pH diagram of the Fe-P-Al-H_(2)O system can determine the stable region of each element in the recovery process of spent LiFePO_(4)Li-batteries.Under the optimal conditions(500 r·min^(-1),15 h,363.15 K,0.4 MPa,the liquid-solid ratio was 4:1 ml·g^(-1)and the acid-material ratio was 0.29),the leaching rate of Li was 99.24%,Fe,Al,and Ti were 0.10%,2.07%,and 0.03%,respectively.The Fe and P were precipitated and recovered as FePO_(4)·2H_(2)O.The kinetic analysis shows that the process of high-pressure acid leaching of spent LiFePO_(4) materials depends on the surface chemical reaction.Through the life cycle assessment(LCA)of the spent LiFePO_(4) whole recovery process,eight midpoint impact categories were selected to assess the impact of recovery process.The results can provide basic environmental information on production process for recycling industry.

Oxidative leaching behavior of metalliferous black shale in acidic solution using persulfate as oxidant

The oxidative dissolution of metalliferous black shale in sulfuric acid solution using sodium persulfate as an oxidant was investigated. The effects of leaching factors including leaching temperature, leaching time, stirring speed, initial concentration of sodium persulfate and sulfuric acid and particle size on the leaching rate were studied as well. The leaching kinetics of molybdenum, nickel and iron from metalliferous black shale shows that the leaching rate is controlled by a chemical reaction through a layer on the unreacted shrinking core. The leaching process follows the kinetics model 1-（1-a）^1/3=kt with apparent activation energies of 34.50, 43.14 and 71.79 kJ/mol for Mo, Ni and Fe, respectively. The reaction orders in sodium persulfate are 0.80, 1.01 and 0.75 for molybdenum, nickel and iron, respectively, while in sulfuric acid, these orders are 0.45, 0.75 and 0.50 for molybdenum, nickel and iron, respectively. In addition, the reaction mechanism for the dissolution of the metalliferous black shale was discussed.

Indium recovery from zinc oxide flue dust by oxidative pressure leaching

Indium was recovered from zinc oxide flue dust(ZOFD)with sulfuric acid by oxidative pressure leaching in an autoclave, and the effects of different technological conditions on indium leaching were studied.Potassium permanganate and hydrogen peroxide were used as oxidants.The atmospheric pressure leaching experiments were also carried out.The experimental results show that the leaching rate of indium can be effectively improved by oxidative pressure leaching.The optimum conditions of pressure leaching are determined as sulfuric 5.10 mol/L acid,leaching time 150 min,temperature 90℃,and the H2O2 dosage of 0.5 mL/g or 2.5%KMnO4.The leaching rate of indium is more than 90%,which is increased by 13%compared with that of atmospheric pressure leaching process without oxidant under the optimum conditions.

Oxidation leaching of copper smelting dust by controlling potential

A study was conducted for metal extraction from copper smelting dust using the oxidation leaching and control of potential technology.The effects of H2O2 dosage,H2O2 feeding speed,initial HCl concentration,leaching temperature,liquid-to-solid ratio and leaching time on metals leaching efficiencies were investigated.The following optimized leaching conditions were obtained:H2O2 dosage of 0.8 mL/g(redox potential of 429 mV),H2O2 feeding speed of 1.0 mL/min,initial H2SO4 concentration of 1.0 mol/L,initial HCl concentration of 1.0 mol/L,leaching temperature of 80°C,initial liquid-to-solid ratio of 5:1 mL/g and leaching time of 1.5 h.Under the optimized conditions,copper and arsenic can be effectively leached from copper smelting dust,leaving residue as a suitable lead resource.The average leaching efficiencies of copper,arsenic and iron are 95.27%,96.82%and 46.65%,respectively.

Separation and recovery of copper in Cu-As-bearing copper electrorefining black slime by oxidation acid leaching and sulfide precipitation

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 ℃.

Selective recovery of lithium from spent lithium iron phosphate batteries using oxidation pressure sulfuric acid leaching system

Oxidation pressure leaching was proposed to selectively dissolve Li from spent LiFePO_(4) batteries in a stoichiometric sulfuric acid solution.Using O_(2) as an oxidant and stoichiometric sulfuric acid as leaching agent,above 97% of Li was leached into the solution,whereas more than 99% of Fe remained in the leaching residue,enabling a relatively low cost for one-step separation of Li and Fe.And then,by adjusting the pH of leachate,above 95% of Li was recovered in the form of the Li_(3)PO_(4) product through iron removal and chemical precipitation of phosphate.

Kinetics of reductive leaching of manganese oxide ore using cellulose as reductant

The kinetics of reductive leaching of manganese from a low-grade manganese oxide ore were studied using cellulose as reductant in dilute sulfuric acid medium.It was found that when the stirring speed was higher than 200 r/min,the effect of gas film diffusion on manganese extraction efficiency could be neglected,and the kinetic behavior was investigated under the condition of elimination of external diffusion influence on the leaching process.Effects of leaching temperature,mass ratio of cellulose and ore,and the sulfuric acid concentration on manganese extraction efficiency were discussed.The kinetic data were analyzed based on the shrinking core model,which indicated that the leaching process was dominated by both ash layer diffusion and chemical reaction at the initial stage,with the progress of leaching reaction,the rate-controlling step switched to the ash layer diffusion.It was also concluded that the sulfuric acid concentration had the most significant influence on the leaching rate,the reaction orders with respect to the sulfuric acid concentration were 2.102 in the first 60 min,and 3.642 in the later 90 min,while the reaction orders for mass ratio of cellulose and ore were 0.660 and 0.724,respectively.An Arrhenius relationship was used to relate the temperature to the rate of leaching,from which apparent activation energies were calculated to be 46.487 kJ/mol and 62.290 kJ/mol at the two stages,respectively.Finally,the overall leaching rate equations for the manganese dissolution reaction with cellulose in sulphuric acid solution were developed.The morphological changes and mineralogical forms of the ore before and after the chemical treatment were discussed with the support of SEM and XRD analyses.

Leaching kinetics of molybdenum from Ni-Mo ore in sulfuric acid solution with sodium peroxodisulfate as oxidant

The leaching kinetics of molybdenum from Ni-Mo ore in sulfuric acid solution with sodium peroxodisulfate was studied.The effects including leaching temperature, reaction time, particle size, stirring speed, and concentrations of sulfuric acid and sodium peroxodisulfate were investigated. The leaching process of molybdenum from Ni-Mo ore is controlled by the chemical reaction through the solid layer across the unreacted shrinking core. The apparent activation energy of the leaching of molybdenum is calculated to be 41.0 k J/mol and the leaching kinetics equation of molybdenum from Ni-Mo ore is expressed as1-(1-a)1/3=3405.7exp[-41030.0/(RT)]t.

Nitrogen mobility,ammonia volatilization,and estimated leaching loss from long-term manure incorporation in red soil

Nitrogen（N） loss from fertilization in agricultural fields has an unavoidable negative impact on the environment and a better understanding of the major pathways can assist in developing the best management practices. The aim of this study was to evaluate the fate of N fertilizers applied to acidic red soil（Ferralic Cambisol） after 19 years of mineral（synthetic） and manure fertilizer treatments under a cropping system with wheat-maize rotations. Five field treatments were examined： control（CK）, chemical nitrogen and potash fertilizer（NK）, chemical nitrogen and phosphorus fertilizer（NP）, chemical nitrogen, phosphorus and potash fertilizer（NPK） and the NPK with manure（NPKM, 70% N from manure）. Based on the soil total N storage change in 0–100 cm depth, ammonia（NH_3） volatilization, nitrous oxide（N_2O） emission, N plant uptake, and the potential N leaching loss were estimated using a mass balance approach. In contrast to the NPKM, all mineral fertilizer treatments（NK, NP and NPK） showed increased nitrate（NO_3~–） concentration with increasing soil depth, indicating higher leaching potential. However, total NH_3 volatilization loss was much higher in the NPKM（19.7%） than other mineral fertilizer treatments（≤4.2%）. The N_2O emissions were generally low（0.2–0.9%, the highest from the NPKM）. Total gaseous loss accounted for 1.7, 3.3, 5.1, and 21.9% for NK, NP, NPK, and NPKM treatments, respectively. Estimated N leaching loss from the NPKM was only about 5% of the losses from mineral fertilizer treatments. All data demonstrated that manure incorporation improved soil productivity, increased yield, and reduced potential leaching, but with significantly higher NH_3 volatilization, which could be reduced by improving the application method. This study confirms that manure incorporationis an essential strategy in N fertilization management in upland red soil cropping system.

A novel method for extracting vanadium by low temperature sodium roasting from converter vanadium slag

Long-term high temperature in conventional vanadium extraction process would cause particles to be sintered and wrapped, thus reducing extraction efficiency of vanadium. Based on the purpose of directional conversion and process intensification, this work proposed a combination of low temperature sodium roasting and high efficiency selective oxidation leaching in vanadium extraction. The investigation of the reaction mechanism suggested that the structure of vanadium slag was changed by roasting, which also caused the fracture of spinel.The addition of MnO2 promoted the directional oxidation of low-valent vanadium into high valence. It also found that Na2S2O8 could oxidize low-valent vanadium effectively in leaching. The leaching efficiency of vanadium reached 87.74% under the optimum conditions, including a roasting temperature of 650 ℃, a roasting time of 2.0 h, a molar ratio of sodium-to-vanadium of 0.6, a MnO2(roasting additive) dosage of 5 wt% and a Na2S2O8(leaching oxidant) dosage of 5 wt%. This percentage is 7.18% higher than that of direct roasting-andleaching under the same conditions.

A method for recovery of iron,titanium,and vanadium from vanadium-bearing titanomagnetite

An innovative method for recovering valuable elements from vanadium-bearing titanomagnetite is proposed. This method involves two procedures： low-temperature roasting of vanadium-bearing titanomagnetite and water leaching of roasting slag. During the roasting process, the reduction of iron oxides to metallic iron, the sodium oxidation of vanadium oxides to water-soluble sodium vanadate, and the smelting separation of metallic iron and slag were accomplished simultaneously. Optimal roasting conditions for iron/slag separation were achieved with a mixture thickness of 42.5 mm, a roasting temperature of 1200°C, a residence time of 2 h, a molar ratio of C/O of 1.7, and a sodium carbonate addition of 70 wt%, as well as with the use of anthracite as a reductant. Under the optimal conditions, 93.67% iron from the raw ore was recovered in the form of iron nugget with 95.44% iron grade. After a water leaching process, 85.61% of the vanadium from the roasting slag was leached, confirming the sodium oxidation of most of the vanadium oxides to water-soluble sodium vanadate during the roasting process. The total recoveries of iron, vanadium, and titanium were 93.67%, 72.68%, and 99.72%, respectively.

Extraction of copper,zinc and cadmium from copper–cadmium-bearing slag by oxidative acid leaching process

An intensified oxidative acid leaching of copper–cadmium-bearing slag featuring using high-efficient oxygen carrier, such as activated carbon, was investigated to achieve high leaching rate of valuable metals. The effects of leaching variables, including agitation rate, sulfuric acid concentration, temperature, slag particle size, activated carbon and cupric ion concentration, were examined. It is found that leaching rates of cadmium and zinc both exceed 99 % in a very short time, but for copper, leaching rate of 99 % is achieved under the optimized leaching parameters, which are agitation rate of 100 r·min^(-1), sulfuric acid concentration of 15 wt%, leaching temperature of 80 ℃, slag particle size of 48–75 lm, activated carbon concentration of 3 g·L^(-1),liquid-to-solid ratio of 4:1, oxygen flow rate of 0.16 L·min^(-1),and leaching time of 60 min. The macro-leaching kinetics of copper metal was analyzed, and it is concluded that the inner diffusion is the controlling step, with apparent activation energy of 18.6 kJ·mol^(-1). The leaching solution with pH value of 2–4 can be designed to selectively extract valuable metals without neutralization, and the leaching residue can be treated by prevailing Pb smelting process.

Influence of phase and microstructure on the rate of hydrochloric acid leaching in pretreated Panzhihua ilmenite

The present study investigated the influence of high temperature oxidation and reduction pretreatments on the leaching rate ofPanzhihua ilmenite. The as-pretreated ilmenite was leached with 20% HCI at 105 ℃, The leaching process was controlled by the phases and microstructures that evolved during the pretreatment processes. The leaching kinetics of pure hematite, ilmenite and pseudobrookite were characterized to clarify the phase effect on the iron-leaching rate; the rate of iron leaching occurs in the following order in the HCI solution： hematite （ferric iron） 〉 ilmenite （ferrous iron） 〉〉 pseudobrookite （ferric iron）. Therefore, the often-cited notion that ferrous iron dissolves faster in HCl solutions than ferric iron when explaining the pretreatment effects is inaccurate. Moreover, the oxidation pretreatment （at 600-1000 ℃ for 4 h） cannot destroy the dense structure of the Panzhihua ilmenite. Therefore, the influence exerted by the oxidation on the leaching process is primarily determined by the phase change; oxidation at 600 and 700℃ slightly increased the rate of iron leaching because the ilmenite was transformed into hematite, while the oxidation at 900-1000℃ significantly reduced the rate of iron leaching because a pseudobrookite phase formed. The reduction effect was subsequently investigated; the as-oxidized ilmenite was reduced under H2 at 750 ℃ for 30 min. The reduction significantly accelerated the rate of subsequent iron leaching such that nearly all of the iron had dissolved after leaching for 2 h in 20% HCl at 105 ℃. This enhanced iron-leaching rate is mainly attributed to the cracks and holes that formed during the reduction process.