Selective leaching of lithium from spent lithium-ion batteries using sulfuric acid and oxalic acid

Traditional hydrometallurgical methods for recovering spent lithium-ion batteries(LIBs)involve acid leaching to simultaneously extract all valuable metals into the leachate.These methods usually are followed by a series of separation steps such as precipitation,extraction,and stripping to separate the individual valuable metals.In this study,we present a process for selectively leaching lithium through the synergistic effect of sulfuric and oxalic acids.Under optimal leaching conditions(leaching time of 1.5 h,leaching temperature of 70°C,liquid-solid ratio of 4 mL/g,oxalic acid ratio of 1.3,and sulfuric acid ratio of 1.3),the lithium leaching efficiency reached89.6%,and the leaching efficiencies of Ni,Co,and Mn were 12.8%,6.5%,and 21.7%.X-ray diffraction(XRD)and inductively coupled plasma optical emission spectrometer(ICP-OES)analyses showed that most of the Ni,Co,and Mn in the raw material remained as solid residue oxides and oxalates.This study offers a new approach to enriching the relevant theory for selectively recovering lithium from spent LIBs.

Recovery of Li, Ni, Co and Mn from spent lithium-ion batteries assisted by organic acids: Process optimization and leaching mechanism

The proper recycling of spent lithium-ion batteries(LIBs)can promote the recovery and utilization of valuable resources,while also negative environmental effects resulting from the presence of toxic and hazardous substances.In this study,a new environmentally friendly hydro-metallurgical process was proposed for leaching lithium(Li),nickel(Ni),cobalt(Co),and manganese(Mn)from spent LIBs using sulfuric acid with citric acid as a reductant.The effects of the concentration of sulfuric acid,the leaching temperature,the leaching time,the solid-liquid ratio,and the reducing agent dosage on the leaching behavior of the above elements were investigated.Key parameters were optimized using response surface methodology(RSM)to maximize the recovery of metals from spent LIBs.The maxim-um recovery efficiencies of Li,Ni,Co,and Mn can reach 99.08%,98.76%,98.33%,and 97.63%.under the optimized conditions(the sulfuric acid concentration was 1.16 mol/L,the citric acid dosage was 15wt%,the solid-liquid ratio was 40 g/L,and the temperature was 83℃ for 120 min),respectively.It was found that in the collaborative leaching process of sulfuric acid and citric acid,the citric acid initially provided strong reducing CO_(2)^(-),and the transition metal ions in the high state underwent a reduction reaction to produce transition metal ions in the low state.Additionally,citric acid can also act as a proton donor and chelate with lower-priced transition metal ions,thus speeding up the dissolution process.

Extraction of Valuable Metals from Titanium-bearing Blast Furnace Slag by Acid Leaching

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.

Effect of Saline Water on Soil Acidity, Alkalinity and Nutrients Leaching in Sandy Loamy Soil in Rwamagana Bella Flower Farm, Rwanda

The necessity to saline and sodic waters is sometimes used for irrigating agricultural activities under certain circumstances, but it is important to note that the use of these waters comes with specific considerations and limitations. One way to decrease undesirable effects of sodic waters on the physical and chemical properties of soils is to apply organic and chemical amendments within the soil. This study aimed to assess the effectiveness of saline water on soil acidity, alkalinity and nutrients leaching in sandy loamy soil at Bella flower farm, in Rwamagana District, Rwanda. The water used was from the Muhazi Lake which is classified as Class I (Saline water quality). Column leaching experiments using treated soils were then conducted under saturated conditions. The soil under experimental was first analyzed for its textural classification, soil properties and is classified as sandy loamy soil. The t-test was taken at 1%, 5% and 10% levels of statistical significance compared to control soil. The results indicated that the application of saline water to soils caused an increase in some soil nutrients like increase of Phosphorus (P), Potassium (K+), Magnesium (Mg2+), Sulphur (S), CN ratio and Sodium (Na+) and decreased soil texture, physical and chemical properties and remained soil nutrients. Consequently, the intensive addition of saline water leachates to soil in PVC pipes led to decreased of soil EC through leaching and a raiser Soluble Sodium Percentage (SSP). The rate of saline water application affected the increase accumulation of SAR and Na% in the top soil layers. The study indicated that saline water is an inefficient amendment for sandy soil with saline water irrigation. The study recommends further studies with similar topic with saline water irrigation, as it accentuated the alkalinity levels.

Leaching of vanadium from stone coal with sulfuric acid

The effects of roasting, mass ratio of H2SO4 to stone coal, leaching temperature, liquid-to-solid ratio, grinding fineness of stone coal, and two-stage counter-current leaching on the vanadium leaching ratio were studied. The results show that the vanadium leaching ratio of roasted stone coal through two-stage counter-current leaching can reach 65.1% at the mass ratio of H2SO4 to stone coal of 20%, leaching temperature of 95℃, leaching time of 30 h, liquid-to-solid ratio of 1.1 mL·g^-1, and grinding fineness of 0.1 mm, which can serve as an experimental basis for the production of vanadium from stone coal.

Vanadium recovery from clay vanadium mineral using an acid leaching method

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.

Extracting vanadium from stone-coal by oxygen pressure acid leaching and solvent extraction

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.

Extraction of vanadium from vanadium slag by high pressure oxidative acid leaching

To extract vanadium in an environment friendly manner, this study focuses on the process of leaching vanadium from vanadium slag by high pressure oxidative acid leaching. Characterizations of the raw slag, mineralogy transformation, and the form of leach residues were made by X-ray diffraction, scanning electron microscopy, and energy dispersive X-ray spectroscopy. The result shows that the vana-dium slag is composed of major phases of fayalite, titanomagnetite, and spinel. During the high pressure oxidative acid leaching process, the fayalite and spinel phases are gradually decomposed by sulfuric acid, causing the release of vanadium and iron in the solution. Meanwhile, unreacted silicon and titanium are enriched in the leach residues. With the initial concentration of sulfuric acid at 250 g·L^-1, a leaching tem-perature of 140℃, a leaching time of 50 min, a liquid-solid ratio of 10：1 mL·g^-1, and oxygen pressure at 0.2 MPa, the leaching rate of vana-dium reaches 97.69%.

Process mineralogy characteristics of acid leaching residue produced in low-temperature roasting-acid leaching pretreatment process of refractory gold concentrates

To provide a theoretical basis for a suitable process to extract gold from refractory gold concentrates, process mineralogy on the acid leaching residue of gold calcine was studied by chemical composition, X-ray diffraction, scanning electron microscopy-energy spectrum, and mineral dissociation analysis. The results showed that the acid leaching residue contained Au 68.22 g/t, Ag 92.71 g/t, Fe 0.44%, As 0.10%, and S 0.55%. Gold and silver minerals existed as native gold, argentite, and proustite. Quartz, the main gangue mineral, accounted for 78.33 wt/%. The dissociation degree analysis showed that the proportions of monomer and exposed gold in acid leaching residue were 96.66 wt%. The cyanidation results showed that the cyanide gold leaching rate of acid leaching residues was close to 100 wt%. However, the maximum cyanide gold leaching rate of gold calcine was only 85.31 wt%. This suggests that acid leaching can increase the gold dissolution rate in the cyanide process.

Leaching Mechanism of the Spodumene Sulphuric Acid Process

Leaching mechanism of acid roasted ore in the spodumene sulphuric acid process was investigated. Experimental results of leaching rates along with variations of leaching temperature in the acidized and neutral leaching processes were reported. Applying ion exchange mechanism in acidized roasting and absorption entrainment mechanism at high temperature, leaching mechanism was discussed. A better explanation of experimental results was given.

Recovery of waste rare earth fluorescent powders by two steps acid leaching

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%.

Acid leaching of vanadium from roasted residue of stone coal

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.

Investigation on leaching of malachite by citric acid

Leaching of an oxidized copper ore containing malachite, as a new approach, was investigated by an organic reagent, citric acid. Sulfuric acid is the most common reagent in the leaching of oxide copper ores, but it has several side effects such as severe adverse impact on the environment. In this investigation, the effects of particle size, acid concentration, leaching time, solid/liquid ratio, temperature, and stirring speed were optimized. According to the experimental results, malachite leaching by citric acid was technically feasible. Optimum leaching conditions were found as follows： the range of particle size, 105-150 μm; acid concentration, 0.2 M; leaching time, 30 min; solid/liquid ratio, 1：20 g/mL; temperature, 40℃; and stirring speed, 200 r/min. Under the optimum conditions, 91.61% of copper was extracted.

Study on mechanisms of different sulfuric acid leaching technologies of chromite

The extraction of chromate from chromite via the sulfuric acid leaching process has strong potential for practical use because it is a simple and environmentally friendly process. This paper aims to study the sulfuric acid leaching process using chromite as a raw material via either microwave irradiation or in the presence of an oxidizing agent. The results show that the main phases in Pakistan chromite are ferrichromspinel, chrompicotite, hortonolite, and silicate embedded around the spinel phases. Compared with the process with an oxidizing agent, the process involving microwaves has a higher leaching efficiency. When the mass fraction of sulfuric acid was 80% and the leaching time was 20 min, the efficiency could exceed 85%. In addition, the mechanisms of these two technologies fundamentally differ. When the leaching was processed in the presence of an oxidizing agent, the silicate was leached first and then expanded. By contrast, in the case of leaching under microwave irradiation, the chromite was dissolved layer by layer and numerous cracks appeared at the particle surface because of thermal shock. In addition, the silicate phase shrunk instead of expanding.

Pressure acid leaching of zinc sulfide concentrate

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%.

Influence of acid leaching and calcination on iron removal of coal kaolin

Calcination and acid leaching of coal kaolin were studied to determine an effective and economical preparation method of calcined kaolin. Thermogravimetric-differential thermal analysis （TG-DTA） and X-ray diffraction （XRD） demonstrated that 900＆#176;C was the suitable temperature for the calcination. Leaching tests showed that hydrochloric acid was more effective for iron dissolution from raw coal kaolin （RCK）, whereas oxalic acid was more effective on iron dissolution from calcined coal kaolin （CCK）. The iron dissolution from CCK was 28.78wt%, which is far less effective than the 54.86wt% of RCK under their respective optimal conditions. Through analysis by using M？ssbauer spectroscopy, it is detected that nearly all of the structural ferrous ions in RCK were removed by hydrochloric acid. However, iron sites in CCK changed slightly by oxalic acid leaching because nearly all ferrous ions were transformed into ferric species after firing at 900＆#176;C. It can be concluded that it is difficult to remove the structural ferric ions and ferric oxides evolved from the structural ferrous ions. Thus, iron removal by acids should be conducted prior to calcination.

Primary beneficiation of tantalite using magnetic separation and acid leaching

Primary beneficiation was successfully performed prior to dissolution of manganotantalite （sample A） and ferrotantalite （sample C） samples obtained from two different mines in the Naquissupa area, Mozambique. Magnetic separation removed the majority of iron and titanium, whereas H2SO4 leaching removed a large portion of thorium and uranium in these samples. Analytical results indicated that 64.14wt% and 72.04wt% of the total Fe and Ti, respectively, and -2wt% each of Nb205 and Ta205 were removed from sample C （ferrotantalite） using the magnetic separation method, whereas only 9.64wt% and 8.66wt% of total Fe203 and TiO2, respectively, and -2wt% each of NbEOs and Ta2O5 were removed from sample A （manganotantalite）. A temperature of 50℃ and a leaching time of 3 h in the presence of concentrated HESOa were observed to be the most appropriate leaching conditions for removal of radioactive elements from the tantalite ores. The results obtained for sample A under these conditions indicated that 64.14wt% U3O8 and 60.77wt% ThO2 were leached into the acidic solution, along with 4.45wt% and 0.99wt% of Nb2O5 and Ta2O5, respectively.

Reductive Acid Leaching of Low Grade Manganese Ores

Manganese recoveries from low-grade ores using organic acids as reducing agents were investigated in the present work. The acid leaching potential of both oxalic acid and citric acid were estimated. Manganese leaching amount were measured by using standard manganese curve and estimated by titration method. Effects of various acid concentrations on leaching efficiency were studied. The observed result suggested prominent manganese recovery of 66% by oxalic acid at 2 M concentration whereas citric acid had less effect on leaching showing leaching percentage upto 40% in 6 days. Acid leaching of manganese ore with both the acids gave a comparative data stating that oxalic acid leached better than citric acid.