Effect of bastnaesite as reductant on hematite reduction during in-situ suspension magnetization roasting of refractory iron ore under neutral atmosphere

The iron tailings of Bayan Obo mines are solid waste,which occupies land area and also causes environmental pollution;however,this waste can be recycled.In this study,based on the characteristics of iron minerals and fluorocarbonate contained in Bayan Obo iron tailings,clean magnetization roasting of iron minerals by bastnaesite from iron tailings during in-situ suspension magnetization roasting in a neutral atmosphere was explored.The results show that for iron tailings with a mass of 12 g,a N_(2) gas flow rate of 600 mL/min,and roasting for 5 min at 800℃,iron concentrate with a 60.44%iron grade at an iron recovery of 76.04%could be obtained.X-ray diffraction analysis showed that the weak magnetic hematite was reduced to strong magnetic magnetite in the neutral atmosphere,without additional reductant.The kinetics of the magnetization roasting of mineral mixtures(bastnaesite and hematite)in a neutral atmosphere showed that the optimal reaction mechanism function was the three-dimensional diffusion model with activation energy of 161.8838 kJ·mol^(-1);this indicates that the reaction was a heterogeneous,diffusion-controlled solid-state reaction.

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.

Thermal decomposition mechanism and kinetics of bastnaesite in suspension roasting process:A comparative study in N_(2) and air atmospheres

To investigate the thermal decomposition behavior and reaction kinetics of bastnaesite in suspension roasting,the gas and solid products of bastnaesite roasted in N2 and air atmospheres were examined using a gas analyzer,X-ray diffraction(XRD),scanning electron microscopy(SEM),and energy dispersive spectrometry(EDS).Subsequently,the kinetic parameters of bastnaesite in the suspension roasting process were derived and calculated using the isothermal method.The results show that the decomposition product of bastnaesite in N_(2) is CeOF.However,once the roasting temperature exceeds 600℃,CO is generated in addition to CO_(2),and all the XRD diffraction peaks of CeOF are shifted to the right,indicating that CO_(2) can oxidize CeOF and lead to the transformation of Ce(Ⅲ) into Ce(Ⅳ).When roasted in air,the decomposition product CeOF can be completely converted to CeF3 and Ce_(7)O_(12) as it easily oxidizes.Additionally,the reaction rate of bastnaesite in air is higher than that of N_(2),and the starting reaction temperature is lower than that of N_(2).A large number of irregular cracks and holes appear on the surface of solid-phase products following suspension roasting,which are due to the thermal decomposition of bastnaesite that produces CO_(2) as well as the reconstruction of the lattice of the solid-phase products.The reaction kinetic model of bastnaesite roasted in N_(2)(temperature range 600-750℃) and air(temperatu re range 500-575℃) confo rms to the A_(3/2) model with the mechanism function G(α)=-ln(1-α)^(2/3),and the reaction activation energy is 59.78 kj/mol and lnA is 1.65 s^(-1) in N_(2) atmosphere.In air,the reaction activation energy is 100.30 kj/mol and lnA is 9.63 s^(-1).

Kinetics of rare earth leaching from roasted ore of bastnaesite with sulfuric acid

Sulfuric acid leaching process was applied to extracting rare earth（RE） from roasted ore of Dechang bastnaesite in Sichuan,China.The effect of particle size,stirring speed,sulfuric acid concentration and leaching temperature on RE extraction efficiency was investigated,and the leaching kinetics of RE was analyzed.Under selected leaching conditions,including particle size（0.074-0.100 mm）,sulfuric acid concentration 1.50 mol/L,mass ratio of liquid to solid 8 and stirring speed 500 r/min,the leaching kinetics analysis shows that the reaction rate of leaching process is controlled by diffusion through the product/ash layer which can be described by the shrinking-core model,and the calculated activation energy of 9.977 kJ/mol is characteristic for a diffusion-controlled process.

Exploring the mechanism of a novel cationic surfactant in bastnaesite flotation via the integration of DFT calculations,in-situ AFM and electrochemistry

Effectively separating bastnaesite from calcium-bearing gangue minerals(particularly calcite)presents a formidable challenge,making the development of efficient collectors crucial.To achieve this,we have designed and synthesized a novel,highly efficient,water-soluble cationic collector,N-dodecylisopropanolamine(NDIA),for use in the bastnaesite-calcite flotation process.Density functional theory(DFT)calculations identified the amine nitrogen atom in NDIA as the site most susceptible to electrophilic attack and electron loss.By introducing an OH group into the traditional collector dodecylamine(DDA)structure,NDIA provided additional adsorption sites,enabling synergistic adsorption on the surface of bastnaesite,thereby significantly enhancing both the floatability and selectivity of these minerals.The recovery of bastnaesite was 76.02%,while the calcite was 1.26%.The NDIA markedly affected the zeta potential of bastnaesite,while its impact on calcite was relatively minor.Detailed Fourier-transform infrared spectroscopy(FTIR)and X-ray photoelectron spectroscopy(XPS)results elucidated that the―NH―and―OH groups in NDIA anchored onto the bastnaesite surface through robust electrostatic and hydrogen bonding interactions,thereby enhancing bastnaesite's affinity for NDIA.Furthermore,in situ atomic force microscopy(AFM)provided conclusive evidence of NDIA aggregation on the bastnaesite surface,improving contact angle and hydrophobicity,and significantly boosting the flotation recovery of bastnaesite.

Crystal structure transformation and lattice impurities migration of quartz during chlorine roasting

To investigate the effect of chlorine roasting on the migration and removal of trace elements in quartz lattice,firstly,an efficient pretreatment process,grinding–HCl washing–flotation–HF and HCl leaching,was used to remove the gangue minerals in quartz ore to obtain purified quartz for the preparation of high-purity quartz and the investigation of lattice impurities migration.The results showed that the high-purity quartz with total impurities less than 50μg/g could be obtained from purified quartz after being treated with chlorine at 1200°C.The variation of crystal structure and the lattice impurities migration of quartz during chlorine roasting were studied through in-situ XRD,TGA,SEM-EDS,ICP-MS,FT-IR and XPS analysis.It revealed that the decomposable impurities H_(2)O,-OH,and residual collectors in the crystal of purified quartz could be effectively removed through chlorine roasting above 900°C,which also had an obvious effect on the removal of low-valence trace elements Li,Na and K in the crystal of quartz but didn't affect the multivalent trace elements Al and Ti.This study revealed the removal and migration mechanism of the trace elements in quartz crystal during chlorine roasting.

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.

Characterization of fluidized reduction roasting of nickel laterite ore under CO/CO_(2)atmosphere

Fluidized reduction roasting is an efficient metallurgical technique.However,its application to nickel laterite ore has rarely been reported.In this paper,the effects of reduction temperature,reduction time,CO concentration,and material particle size on the roasting characteristics of ferronickel fluidization reduction were investigated.Combined with X-ray diffraction,scanning electron microscopy-energy dispersive spectrometry(SEM-EDS)characterization,the mineral phases and microscopic morphology of nickel laterite ore and its roasted ores were analyzed in depth.The results indicated that under the condition of a CO/CO_(2)ratio of 1:1,a reduction temperature of 800℃,and a reduction roasting time of 60 min,a nickel-iron concentrate with a nickel grade of 2.10%and an iron content of 45.96%was produced from a raw material with a nickel grade of 1.45%,achieving a remarkable nickel recovery rate of 46.26%.XRD and SEM-EDS analysis indicated that nickel in the concentrate mainly exists in the form of[Fe,Ni],while the unrecovered nickel in the tailings is primarily present in the form of[Fe,Ni]and Ni_(2)SiO_(4)in forsterite.This study established a theoretical foundation for further exploration of fluidized reduction roasting technology.

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.

Leaching kinetics of bastnaesite concentrate in HCl solution

The leaching kinetics of bastnaesite concentrate in HCl solution was investigated with respect to the effects of HCl concentration by changing HCl concentration,leaching temperature,liquid to solid ratio,and particle size.A particle size of 25 μm was required to leach 89.6% RE2（CO3）3 and 1.5%REF3 at 90 ℃ for 90 min,when HCl concentration was 6 mol/L and liquid to solid ratio was 15：1.The leaching kinetics of bastnaesite concentrate is represented by shrinking core model with diffusion through a porous product layer.The activation energies for the dissolution reaction of RE2（CO3）3 and REF3 were calculated to be 59.39 kJ/mol and 66.13 kJ/mol respectively.

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.