Low-temperature chlorination roasting technology for the simultaneous recovery of valuable metals from spent LiCoO_(2)cathode material

With the continuous increase in the disposal volume of spent lithium-ion batteries(LIBs),properly recycling spent LIBs has become essential for the advancement of the circular economy.This study presents a systematic analysis of the chlorination roasting kinetics and proposes a new two-step chlorination roasting process that integrates thermodynamics for the recycling of LIB cathode materials.The activation energy for the chloride reaction was 88.41 kJ/mol according to thermogravimetric analysis–derivative thermogravimetry data obtained by using model-free,model-fitting,and Z(α)function(αis conversion rate).Results indicated that the reaction was dominated by the first-order(F1)model when the conversion rate was less than or equal to 0.5 and shifted to the second-order(F2)model when the conversion rate exceeded 0.5.Optimal conditions were determined by thoroughly investigating the effects of roasting temperature,roasting time,and the mass ratio of NH_(4)Cl to LiCoO_(2).Under the optimal conditions,namely 400℃,20 min,and NH_(4)Cl/LiCoO_(2)mass ratio of 3:1,the leaching efficiency of Li and Co reached 99.43% and 99.05%,respectively.Analysis of the roasted products revealed that valuable metals in LiCoO_(2)transformed into CoCl_(2) and LiCl.Furthermore,the reaction mechanism was elucidated,providing insights for the establishment of a novel low-temperature chlorination roasting technology based on a crystal structure perspective.This technology can guide the development of LIB recycling processes with low energy consumption,low secondary pollution,high recovery efficiency,and high added value.

Precise regulation of the phase transformation for pyrolusite during the reduction roasting process

The mechanism involved in the phase transformation process of pyrolusite (MnO_(2)) during roasting in a reducing atmosphere was systematically elucidated in this study,with the aim of effectively using low-grade complex manganese ore resources.According to single-factor experiment results,the roasted product with a divalent manganese (Mn^(2+)) distribution rate of 95.30% was obtained at a roasting time of 25 min,a roasting temperature of 700℃,a CO concentration of 20at%,and a total gas volume of 500 mL·min^(-1),in which the manganese was mainly in the form of manganosite (MnO).Scanning electron microscopy and Brunauer–Emmett–Teller theory demonstrated the microstructural evolution of the roasted product and the gradual reduction in the pyrolusite ore from the surface to the core Thermodynamic calculations,X-ray photoelectron spectroscopy,and X-ray diffractometry analyses determined that the phase transformation of pyrolusite followed the order of MnO_(2)→Mn_(2)O_(3)→Mn_(3)O_(4)→MnO phase by phase,and the reduction of manganese oxides in each valence state proceeded simultaneously.

Effect of temperature on suspension magnetization roasting of hematite using biomass waste as reductant:A perspective of gas evolution

The magnetization reduction of hematite using biomass waste can effectively utilize waste and reduce CO_(2) emission to achieve the goals of carbon peaking and carbon neutrality.The effects of temperatures on suspension magnetization roasting of hematite using biomass waste for evolved gases have been investigated using TG-FTIR,Py-GC/MS and gas composition analyzer.The mixture reduction process is divided into four stages.In the temperature range of 200-450℃ for mixture,the release of CO_(2),acids,and ketones is dominated in gases products.The yield and concentration of small molecules reducing gases increase when the temperature increases from 450 to 900℃.At 700℃,the volume concentrations of CO,H_(2) and CH_(4) peak at 8.91%,8.90% and 4.91%,respectively.During the suspension magnetization roasting process,an optimal iron concentrate with an iron grade of 70.86%,a recovery of 98.66% and a magnetic conversion of 45.70% is obtained at 700℃.Therefore,the magnetization reduction could react greatly in the temperature range of 600 to 700℃ owing to the suitable reducing gases.This study shows a detail gaseous evolution of roasting temperature and provides a new insight for studying the reduction process of hematite using biomass waste.

Preparation of sodium molybdate from molybdenum concentrate by microwave roasting and alkali leaching

The preparation process of sodium molybdate has the disadvantages of high energy consumption,low thermal efficiency,and high raw material requirement of molybdenum trioxide,in order to realize the green and efficient development of molybdenum concentrate resources,this paper proposes a new process for efficient recovery of molybdenum from molybdenum concentrate and preparation of sodium molybdate by microwave-enhanced roasting and alkali leaching.Thermodynamic analysis indicated the feasibility of oxidation roasting of molybdenum concentrate.The effects of roasting temperature,holding time,and power-to-mass ratio on the oxidation product and leaching product sodium molybdate (Na_(2)MoO_(4)·2H_(2)O) were investigated.Under the optimal process conditions:roasting temperature of 700℃,holding time of 110 min,and power-to-mass ratio of 110 W/g,the molybdenum state of existence was converted from MoS_(2) to Mo O3.The process of preparing sodium molybdate by alkali leaching of molybdenum calcine was investigated,the optimal leaching conditions include a solution concentration of 2.5 mol/L,a liquid-to-solid ratio of 2 mL/g,a leaching temperature of 60℃,and leaching solution termination at pH 8.The optimum conditions result in a leaching rate of sodium molybdate of 96.24%.Meanwhile,the content of sodium molybdate reaches 94.08wt%after leaching and removing impurities.Iron and aluminum impurities can be effectively separated by adjusting the pH of the leaching solution with sodium carbonate solution.This research avoids the shortcomings of the traditional process and utilizes the advantages of microwave metallurgy to prepare high-quality sodium molybdate,which provides a new idea for the highvalue utilization of molybdenum concentrate.

Ash removal from inferior coal via ammonium fluoride roasting and simultaneous yield of white carbon black

The quality upgrading and deashing of inferior coal by chemical method still faces great challenges.The dangers of strong acid,strong alkali,waste water and exhaust gas as well as high cost limit its industrial production.This paper systematically investigates the ash reduction and desilicification of two typical inferior coal utilizing ammonium fluoride roasting method.Under the optimal conditions,for fat coal and gas coal,the deashing rates are 69.02%and 54.13%,and the desilicification rates are 92.64%and 90.27%,respectively.The molar dosage of ammonium fluoride remains consistent for both coals;however,the gas coal,characterized by a lower ash and silica content(less than half that of the fat coal),achieves optimum deashing effect at a reduced time and temperature.The majority of silicon in coal transforms into gaseous ammonium fluorosilicate,subsequently preparing nanoscale amorphous silica with a purity of 99.90%through ammonia precipitation.Most of the fluorine in deashed coal are assigned in inorganic minerals,suggesting the possibility of further fluorine and ash removal via flotation.This research provides a green and facile route to deash inferior coal and produce nano-scale white carbon black simultaneously.

Extraction of lithium from lepidolite using chlorination roasting-water leaching process

Chlorination roasting followed by water leaching process was used to extract lithium from lepidolite.The microstructure of the lepidolite and roasted materials were characterized by X-ray diffraction（XRD）.Various parameters including chlorination roasting temperature,time,type and amount of chlorinating agents were optimized.The conditional experiments indicate that the best mass ratio of lepidolite to NaCl to CaCl2 is 1：0.6：0.4 during the roasting process.The extraction of lithium reaches peak value of 92.86% at 880 °C,potassium,rubidium,and cesium 88.49%,93.60% and 93.01%,respectively.The XRD result indicates that the major phases of the product after roasting lepidolite with mixture of chlorinating agents（CaCl2 and NaCl） are SiO2,CaF2,KCl,CaSiO3,CaAl2Si2O8,NaCl and NaAlSi3O8.

Preparation of Ti-rich material from titanium slag by activation roasting followed by acid leaching

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.

Oxidation roasting of molybdenite concentrate

In order to investigate the oxidation roasting of molybdenite concentrate in pure oxygen atmosphere, experiments at 673, 723, 773, 873 and 973 K were carried out. The phase transitions and morphology evolutions of the samples obtained at different temperatures after reacting for different time were analyzed by X-ray diffraction（XRD） and scanning electron microscopy（SEM）. The results showed that molybdenite concentrate was oxidized directly to Mo O3 in pure oxygen atmosphere. There were remarkable changes of the morphologies of products with the increase of the roasting temperature. It was also found that sintering phenomenon occurred during the roasting process in pure oxygen when the temperature was above 873 K. The composition of sintered sample was mainly comprised of Mo O3 and some unreacted Mo S2.

Response surface optimization of process parameters for reduction roasting of low-grade pyrolusite by bagasse

The reduction roasting processes for low-grade pyrolusite using bagasse as the reducing agent was statistically analyzed. The central composite rotatable design （CCD） was used to optimize this reduction roasting processes. The three process parameters studied were the mass ratio of bagasse to ore, the roasting temperature and the roasting time. Analysis of variance （ANOVA） was used to analyze the experimental results. The interactions between the process parameters were done by using the linear and quadratic model. The results revealed that the linear and quadratic effects as well as the interaction are statistically significant for the mass ratio and roasting temperature but insignificant for the roasting time. The optimal conditions of 0.9：10 of mass ratio, the roasting temperature of 450 ~C, the roasting time of 30 min were obtained. Under these conditions, the predicted leaching recovery rate for manganese was 98.1%. And the satisfied experimental result of 98.2% confirmed the validity of the model.

Mechanism of phosphorus removal in beneficiation of high phosphorous oolitic hematite by direct reduction roasting with dephosphorization agent

High phosphorous oolitic hematite ore is one of typical intractable iron ores in China, and the conventional beneficiation methods are found to be impracticable to , remove phosphorus from the ore effectively. Better beneficiation index were gotten by direct reduction roasting with dephosphorization agent followed by two stages of grinding and magnetic separation. P content decreases from 0.82% in the raw ore to 0.06% in the magnetic concentrate, and the total iron grade increases from 43.65% to 90.23%, the recovery of iron can reach 87%. Mechanisms of phosphorus removal in the beneficiation of high phosphorous oolitic hematite ore by direct reduction roasting with dephosphorization agent were studied using XRD, SEM and EPMA. The results showed that about 20% of the apatite in the raw ore transferred into phosphorus and volatilized with the gas in the process of reduction roasting, while the rest 80% apatite was not involved in the reaction of generation of phosphorus, and remained as apatite in the roasted products, which was removed to tailings by grinding and magnetic separation. A small amount of phosphorus existed in the magnetic concentrate as apatite. The oolitic texture of raw ore was partly changed during roasting, resulting in the formation of nepheline in the reaction between the dephosphorization agent, SiO2 and Al2O3 in the raw ore, which greatly improved the liberation degree of minerals in the roasted products, and it was beneficial to the subsequent grinding and magnetic separation.

Reaction mechanisms of low-grade molybdenum concentrate during calcification roasting process

The effects of Ca-based additives on roasting properties of low-grade molybdenum concentrate were studied. The resultsshow that calcium-based additives can react with molybdenum concentrate to form CaSO4 and CaMoO4. The initial oxidationtemperature of MoS2 is 450℃, while the formation of CaMoO4 and CaSO4 occurs above 500℃. The whole calcification reactionsare nearly completed between 600 and 650℃. However, raising the temperature further helps for the formation of CaMoO4 but isdisadvantageous to sulfur fixing rate and molybdenum retention rate. Calcification efficiency of Ca-based additives follows theorder： Ca（OH）2〉CaO〉CaCO3. With increasing the dosage of Ca（OH）2, the molybdenum retention rate and sulfur-fixing rate rise, butexcessive dosages would consume more acid during leaching process. The appropriate mass ratio of Ca（OH）2 to molybdenumconcentrate is 1：1. When roasted at 650 ℃ for 90 min, the molybdenum retention rate and the sulfur-fixing rate of low-grademolybdenum concentrate reach 100% and 92.92%, respectively, and the dissolution rate of molybdenum achieves 99.12% withcalcines being leached by sulphuric acid.

Removal of magnesium and calcium from electric furnace titanium slag by H_3PO_4 oxidation roasting-leaching process

H3PO4 oxidation roasting followed by HCl acid leaching was proposed to remove magnesium and calcium from electric furnace titanium slag containing 3.12% MgO and 0.86% CaO. XRF, XRD and SEM techniques were used to characterize the composition, mineral phase component and microstructure of the titanium slag. The H3PO4 oxidation thermodynamic, mineral phase transformation, microstructure, element distribution in titanium slag during H3PO4 oxidation process and leaching process were investigated. The thermodynamic analysis indicated that H3PO4 could promote the decomposition of MgTi2O5 and CaSiO3. The results indicated that H3PO4 could effectively promote the transformation of titanium-bearing mineral to rutile and enrich the impurities in MxTi（3-x）O5 into phosphate which could be removed by acid leaching process. Under the studied conditions, the leaching rates of magnesium and calcium reached 94.68% and 87.19%, respectively. The acid leached slag containing 0.19% MgO and 0.13% CaO（mass fraction） was obtained.

Reaction kinetics of roasting high-titanium slag with concentrated sulfuric acid

A novel method of roasting high-titanium slag with concentrated sulfuric acid was proposed to prepare titanium dioxide, and the roasting kinetics of titania was studied On the basis of roasting process. The effects of roasting temperature, particle size, and acid-to-ore mass ratio on the rate of roasting reaction were investigated. The results showed that the roasting reaction is fitted to a shrinking core model. The results of the kinetic experiment and SEM and EDAX analyses proved that the reaction rate of roasting high-titanium slag with concentrated sulfuric acid is controlled by the internal diffusion on the solid product layer. According to the Arrhenius expression, the apparent activation energy of the roasting reaction is 18.94 kJ/mol.

Vanadium recovery from stone coal through roasting and flotation

A new process for vanadium recovery from stone coal by roasting-flotation was investigated based on the mineralogy. The process comprised four key steps： decarburization, preferential grinding, desliming and flotation. In the decarburization stage, roasting at 550 ℃ effectively avoided the negative effect of the carbonaceous materials in raw ore and generation of free CaO from calcite decomposition during roasting. Through preferential grinding, the high acid-consuming minerals were enriched in the middle fractions, while mica was enriched in the fine and coarse fractions. Through flotation, the final concentrate can be obtained with V2O5 grade of 1.07% and recovery of 83.30%. Moreover, the vanadium leaching rate of the final concentrate increased 13.53% compared to that of the feed. The results reveal that the decarburization by roasting at 550 ℃ is feasible and has little negative impact on mica flotation, and vanadium recovery from stone coal is conducive to reducing handling quantity, acid consumption and production cost.

Response surface optimization of process parameters for removal of F and Cl from zinc oxide fume by microwave roasting

Microwave was applied to roasting the zinc oxide fume obtained from fuming furnace for the removal of F and Cl. The effects of important parameters, such as roasting temperature, holding time and stirring speed, were investigated and the process conditions were optimized using response surface methodology (RSM). The results show that the effects of roasting temperature and holding time on the removal rate of F and Cl are the most significant, and the effect of stirring speed is the second. The defluorination rate reaches 92.6% while the dechlorination rate reaches 90.2%, under the process conditions of roasting temperature of 700 °C, holding time of 80 min and stirring speed of 120 r/min. The results indicate that the removal of F and Cl from fuming furnace production of zinc oxide fumes using microwave roasting process is feasible and reliable.

Preparation of potassium chromate by roasting of carbon ferrochrome

The oxidizing roasting process of carbon ferrochrome to prepare potassium chromate in the presence of potassium carbonate and air was investigated. The effects of reaction temperature, reaction time, mole ratio of potassium carbonate to carbon ferrochrome were studied, and thermodynamics and kinetics were also discussed. It was observed that the reaction temperature and reaction time had a significant influence on the roasting reaction of carbon ferrochrome. The reaction mechanism changed greatly as the temperature varied. A two-stage roasting process was favorable for the roasting reaction, and a chromium recovery rate of 97.06% was obtained through this two-stage roasting method. The chromium residue yielded from this method was only 1/3 of the product. Moreover, the component of Fe in the residue was as high as 55.04%. Therefore, it can be easily recovered to produce sponge iron, realizing complete detoxication and zero-emission of chromium residue.

Extraction of molybdenum and nickel from roasted Ni-Mo ore by hydrochloric acid leaching, sulphation roasting and water leaching

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.

Extraction of vanadium from high calcium vanadium slag using direct roasting and soda leaching

The extraction of vanadium from high calcium vanadium slag was attempted by direct roasting and soda leaching. The oxidation process of the vanadium slag at different temperatures was investigated by X-ray diffraction （XRD）, scanning electron microscopy （SEM）, and energy dispersive spectroscopy （EDS）. The effects of roasting temperature, roasting time, Na2CO3 concentration, leaching tem- perature, leaching time, and liquid to solid ratio on the extraction of vanadium were studied. The results showed that olivine phases and spinel phases in the vanadium slag were completely decomposed at 500 and 800℃, respectively. Vanadium-rich phases were formed at above 850℃. The leaching rate of vanadium reached above 90% under the optimum conditions： roasting temperature of 850℃, roasting time of 60 min, Na2CO3 concentration of 160 g/L, leaching temperature of 95℃, leaching time of 150 min, and liquid to solid ratio of 10：1 mL/g. The main impurities were Si and P in the leach liquor.

Roasting and leaching behaviors of vanadium and chromium in calcification roasting–acid leaching of high-chromium vanadium slag

Calcification roasting–acid leaching of high-chromium vanadium slag(HCVS)was conducted to elucidate the roasting and leaching behaviors of vanadium and chromium.The effects of the purity of Ca O,molar ratio between Ca O and V_2O_5(n(Ca O)/n(V_2O_5)),roasting temperature,holding time,and the heating rate used in the oxidation–calcification processes were investigated.The roasting process and mechanism were analyzed by X-ray diffraction(XRD),scanning electron microscopy(SEM),and thermogravimetry–differential scanning calorimetry(TG–DSC).The results show that most of vanadium reacted with Ca O to generate calcium vanadates and transferred into the leaching liquid,whereas almost all of the chromium remained in the leaching residue in the form of(Fe_(0.6)Cr_(0.4))_2O_3.Variation trends of the vanadium and chromium leaching ratios were always opposite because of the competitive reactions of oxidation and calcification between vanadium and chromium with Ca O.Moreover,Ca O was more likely to combine with vanadium,as further confirmed by thermodynamic analysis.When the HCVS with Ca O added in an n(Ca O)/n(V_2O_5)ratio of 0.5 was roasted in an air atmosphere at a heating rate of 10°C/min from room temperature to 950°C and maintained at this temperature for 60 min,the leaching ratios of vanadium and chromium reached91.14%and 0.49%,respectively;thus,efficient extraction of vanadium from HCVS was achieved and the leaching residue could be used as a new raw material for the extraction of chromium.Furthermore,the oxidation and calcification reactions of the spinel phases occurred at 592and 630°C for n(Ca O)/n(V_2O_5)ratios of 0.5 and 5,respectively.

Thermodynamics and kinetics of extracting zinc from zinc oxide ore by the ammonium sulfate roasting method

Thermodynamic analyses and kinetic studies were performed on zinc oxide ore treatment by （NH4）2SO4 roasting technology. The results show that it is theoretically feasible to realize a roasting reaction between the zinc oxide ore and （NH4）2SO4 in a temperature range of 573-723 K. The effects of reaction temperature and particle size on the extraction rate of zinc were also examined. It is found that a surface chemical reaction is the rate-controlling step in roasting kinetics. The calculated activation energy of this process is about 45.57 kJ/mol, and the kinetic model can be expressed as follows： 1 - （1 - α）1/3 = 30.85 exp（-45.57/RT）·t. An extraction ratio of zinc as high as 92% could be achieved under the optimum conditions.