Roasting bastnaesite concentrates is a crucial process in extracting rare earths.This study explored an efficient suspension roasting technology and investigated the bastnaesite pyrolysis and cerium(Ce)oxidation.Relev...Roasting bastnaesite concentrates is a crucial process in extracting rare earths.This study explored an efficient suspension roasting technology and investigated the bastnaesite pyrolysis and cerium(Ce)oxidation.Relevant analytical tests were applied to evaluate the phase and surface property variations of bastnaesite,and isothermal kinetic analysis of bastnaesite pyrolysis and Ce oxidation was performed.The results revealed that bastnaesite decomposed rapidly and accompanied by Ce oxidation,and the gas-solid products were identified as CO_(2),Ce_(7)O_(12),La_(2)O_(3),CeF_(3) and LaF_(3),with Ce oxidation restricted by bastnaesite pyrolysis.As roasting time prolonged,cracks and pores appeared on bastnaesite surface;the BET specific surface and pore diameter increased.In later roasting period,the pore diameter continued to increase but the specific surface decreased,assigned to particle fusion agglomeration and pore consolidation.Additionally,the surface C content reduced and Ce(Ⅳ)content increased gradually as roasting progressed.The reaction kinetics all followed Avrami-Erofeev equations,the reaction orders of bastnaesite pyrolysis and Ce oxidation decreased with decreasing reaction temperature.The calculated activation energies at lower temperatures were higher than those calculated at higher temperatures.This study analyzed the bastnaesite reaction mechanism to supply a reference for the application of suspension roasting technology in bastnaesite smelting.展开更多
In order to enhance the lead and zinc recovery from the refractory Pb-Zn oxide ore, a new technology was developed based on sulfidation roasting with sulphur by temperature gradient method. The solid-liquid reaction s...In order to enhance the lead and zinc recovery from the refractory Pb-Zn oxide ore, a new technology was developed based on sulfidation roasting with sulphur by temperature gradient method. The solid-liquid reaction system was established and the sulfidation thermodynamics of lead and zinc carbonate was calculated with the software HSC 5.0. The effects of roasting temperature,molar ratio of sulphur to lead and zinc carbonate and reaction time in the first step roasting, and holding temperature and time in the second roasting on the sulfidation extent were studied at a laboratory-scale. The experimental results show that the sulfidation extents of lead and zinc are 96.50% and 97.29% under the optimal conditions, respectively, and the artificial galena, sphalerite and wurtzite were formed. By the novel sulfidizing process, it is expected that the sulphides can be recovered by conventional flotation technology.展开更多
A novel process was proposed for the utilization of potash feldspar by roasting in the presence of sodium carbonate. The effects of roasting temperature, granularity, molar ratio of sodium carbonate to potash feldspar...A novel process was proposed for the utilization of potash feldspar by roasting in the presence of sodium carbonate. The effects of roasting temperature, granularity, molar ratio of sodium carbonate to potash feldspar and roasting time on the silica extraction rate were investigated. Under the optimal roasting conditions, the silica extraction rate was 98%. The optimal conditions, determined using an orthogonal experiment, were found to be roasting temperature of 875 A degrees C, potash feldspar granularity of 74-89 mu m, molar ratio of sodium carbonate to potash feldspar of 1.2:1, and roasting time of 80 min. The kinetics of potash feldspar roasting in the presence of sodium carbonate was described by the shrinking core model and the reaction rate was found to be controlled by the chemical reaction at the particle surface. According to the Arrhenius expression, the activation energy was 164.99 kJ/mol, and the process could be expressed as [1-(1-alpha)(1/3)]=2.66x10(5) exp[-164990/(RT)] t.展开更多
Suspension roasting followed by magnetic separation is a promising method to upgrade oolitic hematite ore.An oolitic hematite ore was roasted using suspension roasting technology at different temperatures.The phase tr...Suspension roasting followed by magnetic separation is a promising method to upgrade oolitic hematite ore.An oolitic hematite ore was roasted using suspension roasting technology at different temperatures.The phase transformation for iron minerals was investigated by XRD and Mossbauer spectrum,and the characteristics of roasted product were analyzed by VSM and SEM-EDS.Results indicate that the magnetic concentrate is of 58.73% Fe with iron recovery of 83.96% at 650 °C.The hematite is rapidly transformed into magnetite during the roasting with transformation ratio of 92.75% at 650 °C.Roasting temperature has a significant influence on the phase transformation of hematite to magnetite.The transformation ratio increases with increased temperature.After roasting,the magnetic susceptibility is significantly improved,while iron ore microstructure is not altered significantly.展开更多
The orthogonal test was used to optimize the reaction conditions of roasting zinc oxide ore with NaOH aiming to comprehensively utilize zinc oxide ore.The optimized reaction conditions were molar ratio of NaOH to zinc...The orthogonal test was used to optimize the reaction conditions of roasting zinc oxide ore with NaOH aiming to comprehensively utilize zinc oxide ore.The optimized reaction conditions were molar ratio of NaOH to zinc oxide ore 6:1,roasting temperature 450°C,holding time 150 min.The molar ratio of NaOH to zinc oxide ore was the most predominant factor affecting the extraction ratios of zinc oxide and silica.The mineral phase transformations were investigated by testing the phases of specimens obtained at different temperatures.The process was that silica reacted with molten NaOH to form Na_2SiO_3 at first,then transformed into Na_4SiO_4 with temperature rising.ZnCO_3 and its decomposing product ZnO reacted with NaOH to form Na_2ZnO_2.Na_2ZnSiO_4was also obtained.The reaction rate was investigated using unreacted shrinking core model.Two models used were chemical reaction at the particle surface and diffusion through the product layer.The results indicated that the reaction rate was combine-controlled by two models.The activation energy and frequency factor were obtained as 24.12 k J/mol and 0.0682,respectively.展开更多
Low recovery rate,high energy consumption and serious pollution are existed in traditional processes of vanadium extraction.To seek a new process with high recovery rate,low energy consumption and less pollution disch...Low recovery rate,high energy consumption and serious pollution are existed in traditional processes of vanadium extraction.To seek a new process with high recovery rate,low energy consumption and less pollution discharge is of great significance.Microwave-aided roasting for vanadium extraction is an environmentally-friendly technology.Non-pollution roasting processes with additives following sodium carbonate and potassium carbonate,under microwave irradiation were investigated with stone coal.The valence variation of vanadium in oxidation roasting of the vanadium bearing stone coal was studied.The results showed that the oxidation process of vanadium was a stepwise process following V(Ⅲ)→V(Ⅳ)→V(Ⅴ).The roasted stone coal was characterized by XRD,SEM.Factors associated with extraction rate were investigated following roasting temperature,roasting time,microwave power,and amount of roasting additive.Comparing with conventional roasting process,the recovery rate was higher without air pollution.展开更多
Zinc leaching residue(ZLR) contains high content of valuable metals such as zinc and iron. However, zinc and iron mainly exist in the form of zinc ferrite, which are difficult to separate and recover. This study propo...Zinc leaching residue(ZLR) contains high content of valuable metals such as zinc and iron. However, zinc and iron mainly exist in the form of zinc ferrite, which are difficult to separate and recover. This study proposed a new process involving sulfidation roasting, magnetic separation and flotation to recover zinc and iron in ZLR. Through sulfidation roasting of ZLR with pyrite, zinc and iron were converted into ZnS and Fe3 O4. The effects of pyrite dosage, roasting temperature and roasting time on the sulfidation of zinc in ZLR were investigated. The results showed that the sulfidation percentage of zinc reached 91.8% under the optimum condition. Besides, it was found that ball-milling was favorable for the separation and recovery of zinc and iron in sulfidation products. After ball-milling pretreatment, iron and zinc were enriched from sulfidation products by magnetic separation and flotation. The grade of iron in magnetic concentrates was 52.3% and the grade of zinc in flotation concentrates was 31.7%, which realized the recovery of resources.展开更多
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...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.展开更多
To explicate the thermodynamics of the chromite ore lime-free roasting process, the thermodynamics of reactions involved in this process was calculated and the phrases of sinter with different roasting times were stud...To explicate the thermodynamics of the chromite ore lime-free roasting process, the thermodynamics of reactions involved in this process was calculated and the phrases of sinter with different roasting times were studied. The thermodynamics calculation shows that all the standard Gibbs free energy changes of the reactions to form Na2CrO4, Na2O-Fe2O3, Na2O·Al2O3 and Na2O3 SiO2 via chromite ore and Na2CO3 are negative, and the standard Gibbs free energy changes of the reactions between MgO, Fe2O3 and SiO2 released from chromite spinel to form MgO-Fe2O3 and MgO·SiO2 are also negative at the oxidative roasting temperatures （1 173 1 473 K）. The phrase analysis of the sinter in lime-free roasting process shows that Na2O·Fe2O3, Na2O·Al2O3 and Na2O·SiO2 can be formed in the first 20 min, but they decrease in contents and finally disappear with the increase of roasting time. The final phase compositions of the sinter are Na2CrO4, MgO·Fe2O3, MgO·SiO2 and MgO. The results indicate that Na2CrO4 can be formed easily via the reaction ofNa2CO3 with chromite ore. Na2O·Fe2O3, Na2O-Al2O3 and Na2O·SiO2 can be formed as intermediate compounds in the roasting process and they can further react with chromite ore to form Na2CrO4. MgO released from chromite ore may react with iron oxides and silicon oxide to form stable compounds of MgO·Fe2O3 and MgO·SiO2, respectively.展开更多
The features of the techniques of fast reducing roasting （FRR） and conventional magnetic roasting, as well as tremendous demands of iron ores in iron and steel industry of China, were briefly described. The test equ...The features of the techniques of fast reducing roasting （FRR） and conventional magnetic roasting, as well as tremendous demands of iron ores in iron and steel industry of China, were briefly described. The test equipment suitable for FRR of fine-grained materials was introduced. Weakly magnetic materials with grain size of 〈0.30 mm were converted into strongly magnetic materials by FRR for several to dozens of seconds. In a weakly reducing atmosphere and at 740-800 ~C, refractory powder iron material （〈0.30 mm） which is rich in specularite, limonite and Mg-Mn siderite was subjected to FRR for a few seconds to 60 s. Concentrate with iron grade of 55.67%-55.21%, high contents of Mg and Mn in the ore is obtained and the yield of magnetic separation reaches 81.66%-86.57%. The results of X-ray diffraction （XRD） analysis and magnetism detection of the material before and after FRR indicate that weakly magnetic material is mainly converted into strongly magnetic material Fe304 with specific saturation magnetic moment. The efficiency of FRR is consistent with TFe recovery of magnetic separation; meantime, the specific sa^u＇ation magnetic moment increases from 33 to 42 times after FRR. Calculations show that speeds of flash magnetic roasting are obtained from several dozen to two or three hundred times, compared with rotary kiln or shaft furnace. This indicates that it is practicable to use the fast reducing roasting technique to improve the comprehensive utilization of iron ore resources.展开更多
Compared with natural magnetite concentrate, artificial magnetite with more lattice defects and higher activity tends to be oxidized. And the artificial magnetite pellet at the temperature of 400℃ has the oxidation d...Compared with natural magnetite concentrate, artificial magnetite with more lattice defects and higher activity tends to be oxidized. And the artificial magnetite pellet at the temperature of 400℃ has the oxidation degree approaching to natural magnetite concentrate pellet fired at 1000℃. Besides, two kinds of pellets displayed quite different roasting characteristics. When preheated at the same temperature for the same period of time, natural magnetite concentrate pellet and artificial magnetite concentrate pellet need to be roasted at the temperature of 1100℃ and 1250℃, respectively, for 25 min to reach the compressive strength of 3000 N per pellet. When roasted at the same temperature of 1200℃, natural magnetite pellet and artificial magnetite pellet need to be roasted for 15 min and 30 min, respectively, to reach the compressive strength over 3000 N per pellet. It can be seen from the test that artificial magnetite pellet has a faster oxidation, resulting in the high porosity in the produced pellet, and it requires a roasting process at higher temperature for a longer time to reach the desired compressive strength for industrial production.展开更多
Oolitic hematite is an iron ore resource with rich reserves,complex composition,low grade,fine disseminated particle sizes,and a unique oolitic structure.In this study,a microwave-assisted suspension magnetization roa...Oolitic hematite is an iron ore resource with rich reserves,complex composition,low grade,fine disseminated particle sizes,and a unique oolitic structure.In this study,a microwave-assisted suspension magnetization roasting technology was proposed to recover and utilize the ore.The results showed that under the conditions of microwave pretreatment temperature of 1050℃ for 2 min,a magnetic concentrate with an iron grade of 58.72%at a recovery of 89.32%was obtained by microwave suspension magnetization roasting and magnetic separation.Moreover,compared with the no microwave pretreatment case,the iron grade and recovery increased by 3.17%and 1.58%,respectively.Microwave pretreatment increased the saturation magnetization of the roasted products from 24.974 to 39.236(A∙m^(2))/kg and the saturation susceptibility from 0.179×10^(−3) m^(3)/kg to 0.283×10^(−3) m^(3)/kg.Microcracks were formed between the iron and gangue minerals,and they gradually extended to the core of oolite with the increase in the pretreatment time.The reducing gas diffused from outside to inside along the microcracks,which promoted the selective transformation of the weak magnetic hematite into the strong magnetic magnetite.展开更多
In this study,a roasting enhanced flotation process was proposed to recover LiMn_(2)O_(4) and grapite from waste lithium-ion batteries(LIBs).The effects of roasting temperature and time on the surface modification was...In this study,a roasting enhanced flotation process was proposed to recover LiMn_(2)O_(4) and grapite from waste lithium-ion batteries(LIBs).The effects of roasting temperature and time on the surface modification was investigated,and a series of analytical technologies were used to reveal process mechanism.The results indicate that LiMn_(2)O_(4) can be effectively separated from graphite via flotation after the roasting.The flotation grade of LiMn_(2)O_(4) was significantly increased from 63.10%to 91.36%after roasting at 550℃for 2 h.The TG-DTG analysis demonstrates that the difficulty in flotation separation of LiMn_(2)O_(4) from graphite is caused by the organic binder and electrolytes coating on their surfaces.The XRD,SEM,XPS,and contact angle analyses confirm that the organic films on the surfaces of those materials can be effectively removed by roasting,after which the wettability of LiMn_(2)O_(4) is regained and thus the surface wettability difference between the cathode and anode materials is increased significantly.The closed-circuit flotation test indicates that a LiMn_(2)O_(4) sample with high grade of 99.81%is obtained,while the recovery of LiMn_(2)O_(4) is as high as 99.40%.This study provides an economical and eco-friendly way to recycling waste LIBs.展开更多
A large amount of acid-leaching residue is produced during the conventional Zn hydrometallurgy process,and this residue has a large concentration of a variety of valuable metals.The purpose of this study was to evalua...A large amount of acid-leaching residue is produced during the conventional Zn hydrometallurgy process,and this residue has a large concentration of a variety of valuable metals.The purpose of this study was to evaluate the ability of a procedure that entails the use of sulfation roasting,water leaching,and chlorination leaching(blowing oxygen technique)to recover Pb and Ag,followed by cooling crystallization and the replacement of Ag with lead sheet,to realize the full recovery of all valuable metals from zinc acid-leaching residue;consequently,good results were achieved.The best results were obtained under the following conditions:a sulfuric acid at 70%of the raw material quality,roasting temperature of 300°C and roasting time of 2 h,followed by the process of leaching the roasted residue for 1 h by applying a water-to-solid ratio of 5꞉1 at room temperature.The recovery rates of Zn and Fe were 98.69%and 92.36%,respectively.The main parameters of the chlorine salt leaching system were as follows:Cl−concentration of 300 g/L,Fe3+concentration of 25 g/L,acid concentration of 2 mol/L,liquid-to-solid ratio of 9 mL:1 g,temperature of 90℃,and leaching time of 0.5 h;this leaching process was followed by filtration separation.These conditions resulted in high extents of leaching for Pb and Ag(i.e.,98.87%and 96.74%,respectively).Finally,the kinetics of the process of Ag leaching using Cl−ions in an oxygen-rich medium was investigated.It was found that the leaching process was controlled by the diffusion of the product layers,and the activation energy was 19.82 kJ/mol.展开更多
To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leach...To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leaching was proposed. Firstly, several factors such as pyrite dosage, roasting temperature, carbon powder dosage, holding time and particle size affecting on the flotation performance of Zn(Pb) and magnetic separation performance of Fe were simultaneously examined and the optimum process parameters were determined. A flotation concentrate, containing 17.46% Zn and 3.93% Pb, was obtained, and the Zn and Pb recoveries were 86.04% and 69.08%, respectively. The obtained flotation tailing was concentrated by a low-intensity magnetic separator. The grade of iron increased from 5.45% to 43.45% and the recovery of iron reached 64.87%. Hydrochloric acid leaching was then carried out for the magnetic separation tailing and a raw quartz concentrate containing 81.05% SiO2 was obtained. To further interpret the sulfidation mechanism of smithsonite, surface morphology and component of the sample before and after reactions were characterized by XRD and EPMA-EDS. The aim was to achieve the comprehensive utilization of the low-grade mining ore.展开更多
The separation of Al from the silicon-rich diasporic bauxite is of great significance in alumina production.Herein,we proposed a low-temperature ammonium sulfate roasting-water leaching process to extract aluminum fro...The separation of Al from the silicon-rich diasporic bauxite is of great significance in alumina production.Herein,we proposed a low-temperature ammonium sulfate roasting-water leaching process to extract aluminum from silicon-rich diasporic bauxite.Parameters including roasting temperature,dosage of ammonium sulfate,roasting time,and particle size of ore were investigated.Under the condition of roasting temperature of 400℃,roasting time of 5 h,ammonium sulfate dosage of 2.5 times of the theoretical value and ore particle size of 80−96μm,more than 98%leaching rate of aluminum was obtained.The phase transformation and mechanism during the roasting process were revealed by using X-ray diffraction,thermogravimetric analysis,differential thermal analysis,and scanning electron microscope methods.The diaspore and kaolinite phases in the silicon-rich diasporic bauxite could react with ammonium sulfate to form corresponding sulfates(NH_(4))_(3)Al(SO_(4))_(3) Al(SO_(4))_(2) and Al_(2)(SO_(4))_(3).The proposed technology could provide an effective method for the direct and separation of aluminum from silicon-rich diasporic bauxite.展开更多
This paper mainly investigated the antimony recovery from antimony-bearing dusts through reduction roasting process after the dust firstly oxidation roasted.CO–CO2 mixture gas was used as reducing agent,and the antim...This paper mainly investigated the antimony recovery from antimony-bearing dusts through reduction roasting process after the dust firstly oxidation roasted.CO–CO2 mixture gas was used as reducing agent,and the antimony-containing phase was reduced into Sb4O6,volatilized into smoke,and finally recovered through the cooling cylinder.The antimony recovery rate increased from 66.00 wt%to 73.81 wt%in temperature range of 650 to 800°C,and decreased with temperature increased further to 900°C due to the reduction of Sb4O6 to the nonvolatile Sb.Similarly,the CO partial pressure also played a double role in this test.Under optimized conditions of roasting temperature of 800°C,CO partial pressure of 7.5 vol%and roasting time of 120 min,98.40 wt%of arsenic removal rate and 80.40 wt%antimony recovery rate could be obtained.In addition,the“As2O3”product could be used for preparing ferric arsenate which realized the harmless treatment of it.展开更多
The present study aims at the recovery of potassium from muscovite mica(which contains K_(2)O;~10 wt%)using NaCl-roasting coupled with H_(2)SO_(4)-leaching process.The preliminary acid leaching studies applying differ...The present study aims at the recovery of potassium from muscovite mica(which contains K_(2)O;~10 wt%)using NaCl-roasting coupled with H_(2)SO_(4)-leaching process.The preliminary acid leaching studies applying different mineral acids resulted in a potassium recovery of 8%−18%.The optimum leaching conditions for the maximum recovery were 4 mol/L H_(2)SO_(4),60 min leaching time and liquid-solid ratio 4 mL/g at 90℃.However,the roasting of muscovite with additive NaCl(muscovite:NaCl mass ratio of 1:1.00,900℃,45 min)followed by H_(2)SO_(4)-leaching(95℃,60 min)extracted potassium to the tune of 98%.Under similar roasting conditions,the H_(2)O-leaching process extracted only 60%of potassium.The effects of various roasting and leaching parameters such as temperature,time,NaCl concentration,acid concentration,liquid-solid ratio on potassium extraction were evaluated.The appearance of the sylvite(KCl)mineral phase in the NaCl-roasted muscovite and its disappearance in the acid/water leached residue confirmed the physical and chemical distortions of the muscovite crystal structure.The possible mechanism of potassium release from the complex muscovite structure was elucidated based on available literature substantiated by characterizations using X-ray diffraction(XRD)and scanning electron microscopy with energy dispersive X-rays spectroscopy(SEM-EDX).展开更多
Siderite,as an abundant iron ore,has not been effectively utilized,with a low utilization rate.In this study,the in-situ kinetics and mechanism of siderite during suspension magnetization roasting(SMR)were investigate...Siderite,as an abundant iron ore,has not been effectively utilized,with a low utilization rate.In this study,the in-situ kinetics and mechanism of siderite during suspension magnetization roasting(SMR)were investigated to improve the selective conversion of siderite to magnetite and CO,enriching the theoretical system of green SMR using siderite as a reductant.According to the gas products analyses,the peak value of the reaction rate increased with increasing temperature,and its curves presented the feature of an early peak and long tail.The mechanism function of the siderite pyrolysis was the contraction sphere model(R_(3)):f(α)=3(1−α)2/3;E_(α)was 46.4653 kJ/mol;A was 0.5938 s^(−1);the kinetics equation was k=0.5938exp[−46.4653/(RT)].The in-situ HT-XRD results indicated that siderite was converted into magnetite and wüstite that exhibited a good crystallinity in SMR under a N_(2) atmosphere.At 620℃,the saturation magnetization(M_(s)),remanence magnetization(Mr),and coercivity(Hc)of the product peaked at 53.63×10^(-3)A·m^(2)/g,10.23×10^(-3)A·m^(2)/g,and 12.40×10^(3)A/m,respectively.Meanwhile,the initial particles with a smooth surface were transformed into particles with a porous and loose structure in the roasting process,which would contribute to reducing the grinding cost.展开更多
The growing characteristics of metallic iron particles during reductive roasting of boron-bearing magnetite concentrate under different conditions were investigated.The size of the metallic iron particles was quantita...The growing characteristics of metallic iron particles during reductive roasting of boron-bearing magnetite concentrate under different conditions were investigated.The size of the metallic iron particles was quantitatively measured via optical image analysis with consideration of size calibration and weighted ratio of image numbers in the core,middle and periphery zones of cross-section of pellets.In order to guarantee the measurement accuracy,54 images were captured in total for each specimen,with a weighted ratio of 1:7:19 with respect to the core,middle and periphery section of the cross-section of pellets.Increasing reduction temperature and time is favorable to the growth of metallic iron particles.Based on the modification of particle size measurement,in terms of time(t)and temperature(T)a predicting model of metallic iron particle size(D),was established as:D=125−0.112t−0.2352T−5.355×10^−4t^2+2.032×10^−4t∙T+1.134×10^−4T^2.展开更多
基金Projects(2022YFC2905800,2021YFC2901000)supported by the National Key R&D Program of ChinaProject(52174242)supported by the National Science and Technology of ChinaProject(52130406)supported by the National Science and Technology Major Project of China。
文摘Roasting bastnaesite concentrates is a crucial process in extracting rare earths.This study explored an efficient suspension roasting technology and investigated the bastnaesite pyrolysis and cerium(Ce)oxidation.Relevant analytical tests were applied to evaluate the phase and surface property variations of bastnaesite,and isothermal kinetic analysis of bastnaesite pyrolysis and Ce oxidation was performed.The results revealed that bastnaesite decomposed rapidly and accompanied by Ce oxidation,and the gas-solid products were identified as CO_(2),Ce_(7)O_(12),La_(2)O_(3),CeF_(3) and LaF_(3),with Ce oxidation restricted by bastnaesite pyrolysis.As roasting time prolonged,cracks and pores appeared on bastnaesite surface;the BET specific surface and pore diameter increased.In later roasting period,the pore diameter continued to increase but the specific surface decreased,assigned to particle fusion agglomeration and pore consolidation.Additionally,the surface C content reduced and Ce(Ⅳ)content increased gradually as roasting progressed.The reaction kinetics all followed Avrami-Erofeev equations,the reaction orders of bastnaesite pyrolysis and Ce oxidation decreased with decreasing reaction temperature.The calculated activation energies at lower temperatures were higher than those calculated at higher temperatures.This study analyzed the bastnaesite reaction mechanism to supply a reference for the application of suspension roasting technology in bastnaesite smelting.
基金Project(51204210)supported by the National Natural Science Foundation of ChinaProject(2011AA061001)supported by the High-Tech Research and Development Program of ChinaProject(2012BAC12B04)supported by the National Science&Technology During the12th Five-Year Plan Period,China
文摘In order to enhance the lead and zinc recovery from the refractory Pb-Zn oxide ore, a new technology was developed based on sulfidation roasting with sulphur by temperature gradient method. The solid-liquid reaction system was established and the sulfidation thermodynamics of lead and zinc carbonate was calculated with the software HSC 5.0. The effects of roasting temperature,molar ratio of sulphur to lead and zinc carbonate and reaction time in the first step roasting, and holding temperature and time in the second roasting on the sulfidation extent were studied at a laboratory-scale. The experimental results show that the sulfidation extents of lead and zinc are 96.50% and 97.29% under the optimal conditions, respectively, and the artificial galena, sphalerite and wurtzite were formed. By the novel sulfidizing process, it is expected that the sulphides can be recovered by conventional flotation technology.
基金Project(51204054)supported by the National Natural Science Foundation of ChinaProject(2007CB613603)supported by the National Basic Research Program of ChinaProject(N140204011)supported by the Ministry of Education Basic Scientific Research Business Expenses,China
文摘A novel process was proposed for the utilization of potash feldspar by roasting in the presence of sodium carbonate. The effects of roasting temperature, granularity, molar ratio of sodium carbonate to potash feldspar and roasting time on the silica extraction rate were investigated. Under the optimal roasting conditions, the silica extraction rate was 98%. The optimal conditions, determined using an orthogonal experiment, were found to be roasting temperature of 875 A degrees C, potash feldspar granularity of 74-89 mu m, molar ratio of sodium carbonate to potash feldspar of 1.2:1, and roasting time of 80 min. The kinetics of potash feldspar roasting in the presence of sodium carbonate was described by the shrinking core model and the reaction rate was found to be controlled by the chemical reaction at the particle surface. According to the Arrhenius expression, the activation energy was 164.99 kJ/mol, and the process could be expressed as [1-(1-alpha)(1/3)]=2.66x10(5) exp[-164990/(RT)] t.
基金Project([2011]01-69-07)supported by the China Geological Survey Project
文摘Suspension roasting followed by magnetic separation is a promising method to upgrade oolitic hematite ore.An oolitic hematite ore was roasted using suspension roasting technology at different temperatures.The phase transformation for iron minerals was investigated by XRD and Mossbauer spectrum,and the characteristics of roasted product were analyzed by VSM and SEM-EDS.Results indicate that the magnetic concentrate is of 58.73% Fe with iron recovery of 83.96% at 650 °C.The hematite is rapidly transformed into magnetite during the roasting with transformation ratio of 92.75% at 650 °C.Roasting temperature has a significant influence on the phase transformation of hematite to magnetite.The transformation ratio increases with increased temperature.After roasting,the magnetic susceptibility is significantly improved,while iron ore microstructure is not altered significantly.
基金Projects(51774070,51204054)supported by the National Natural Science Foundation of ChinaProject(N150204009)supported by the Ministry of Education Basic Scientific Research Business Expenses,ChinaProject(2007CB613603)supported by the National Basic Research Program of China
文摘The orthogonal test was used to optimize the reaction conditions of roasting zinc oxide ore with NaOH aiming to comprehensively utilize zinc oxide ore.The optimized reaction conditions were molar ratio of NaOH to zinc oxide ore 6:1,roasting temperature 450°C,holding time 150 min.The molar ratio of NaOH to zinc oxide ore was the most predominant factor affecting the extraction ratios of zinc oxide and silica.The mineral phase transformations were investigated by testing the phases of specimens obtained at different temperatures.The process was that silica reacted with molten NaOH to form Na_2SiO_3 at first,then transformed into Na_4SiO_4 with temperature rising.ZnCO_3 and its decomposing product ZnO reacted with NaOH to form Na_2ZnO_2.Na_2ZnSiO_4was also obtained.The reaction rate was investigated using unreacted shrinking core model.Two models used were chemical reaction at the particle surface and diffusion through the product layer.The results indicated that the reaction rate was combine-controlled by two models.The activation energy and frequency factor were obtained as 24.12 k J/mol and 0.0682,respectively.
基金supported by the National High Technology Research and Development Program of China(2008AA031202)Major Science&Technology of Chongqing(CSTC2008AA4026)foundation of Chongqing University for Young Talents to Promote the Innovative Force(qnjj2008-7)
文摘Low recovery rate,high energy consumption and serious pollution are existed in traditional processes of vanadium extraction.To seek a new process with high recovery rate,low energy consumption and less pollution discharge is of great significance.Microwave-aided roasting for vanadium extraction is an environmentally-friendly technology.Non-pollution roasting processes with additives following sodium carbonate and potassium carbonate,under microwave irradiation were investigated with stone coal.The valence variation of vanadium in oxidation roasting of the vanadium bearing stone coal was studied.The results showed that the oxidation process of vanadium was a stepwise process following V(Ⅲ)→V(Ⅳ)→V(Ⅴ).The roasted stone coal was characterized by XRD,SEM.Factors associated with extraction rate were investigated following roasting temperature,roasting time,microwave power,and amount of roasting additive.Comparing with conventional roasting process,the recovery rate was higher without air pollution.
基金Project(2018YFC1900305)supported by the National Key R&D Program of ChinaProject(51825403)supported by the National Science Foundation for Distinguished Young Scholars,China+1 种基金Projects(51634010,51474247,51904354)supported by the National Natural Science Foundation of ChinaProject(2019SK2291)supported by the Key Research and Development Program of Hunan Province,China。
文摘Zinc leaching residue(ZLR) contains high content of valuable metals such as zinc and iron. However, zinc and iron mainly exist in the form of zinc ferrite, which are difficult to separate and recover. This study proposed a new process involving sulfidation roasting, magnetic separation and flotation to recover zinc and iron in ZLR. Through sulfidation roasting of ZLR with pyrite, zinc and iron were converted into ZnS and Fe3 O4. The effects of pyrite dosage, roasting temperature and roasting time on the sulfidation of zinc in ZLR were investigated. The results showed that the sulfidation percentage of zinc reached 91.8% under the optimum condition. Besides, it was found that ball-milling was favorable for the separation and recovery of zinc and iron in sulfidation products. After ball-milling pretreatment, iron and zinc were enriched from sulfidation products by magnetic separation and flotation. The grade of iron in magnetic concentrates was 52.3% and the grade of zinc in flotation concentrates was 31.7%, which realized the recovery of resources.
基金Project(50674014) supported by the National Natural Science Foundation of China
文摘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.
基金Project(2009FJ1009) supported by the Major Science and Technology Program of Hunan Province,China
文摘To explicate the thermodynamics of the chromite ore lime-free roasting process, the thermodynamics of reactions involved in this process was calculated and the phrases of sinter with different roasting times were studied. The thermodynamics calculation shows that all the standard Gibbs free energy changes of the reactions to form Na2CrO4, Na2O-Fe2O3, Na2O·Al2O3 and Na2O3 SiO2 via chromite ore and Na2CO3 are negative, and the standard Gibbs free energy changes of the reactions between MgO, Fe2O3 and SiO2 released from chromite spinel to form MgO-Fe2O3 and MgO·SiO2 are also negative at the oxidative roasting temperatures （1 173 1 473 K）. The phrase analysis of the sinter in lime-free roasting process shows that Na2O·Fe2O3, Na2O·Al2O3 and Na2O·SiO2 can be formed in the first 20 min, but they decrease in contents and finally disappear with the increase of roasting time. The final phase compositions of the sinter are Na2CrO4, MgO·Fe2O3, MgO·SiO2 and MgO. The results indicate that Na2CrO4 can be formed easily via the reaction ofNa2CO3 with chromite ore. Na2O·Fe2O3, Na2O-Al2O3 and Na2O·SiO2 can be formed as intermediate compounds in the roasting process and they can further react with chromite ore to form Na2CrO4. MgO released from chromite ore may react with iron oxides and silicon oxide to form stable compounds of MgO·Fe2O3 and MgO·SiO2, respectively.
基金Project(20070497048) supported by China Scholarship Council,Ministry of Education of China
文摘The features of the techniques of fast reducing roasting （FRR） and conventional magnetic roasting, as well as tremendous demands of iron ores in iron and steel industry of China, were briefly described. The test equipment suitable for FRR of fine-grained materials was introduced. Weakly magnetic materials with grain size of 〈0.30 mm were converted into strongly magnetic materials by FRR for several to dozens of seconds. In a weakly reducing atmosphere and at 740-800 ~C, refractory powder iron material （〈0.30 mm） which is rich in specularite, limonite and Mg-Mn siderite was subjected to FRR for a few seconds to 60 s. Concentrate with iron grade of 55.67%-55.21%, high contents of Mg and Mn in the ore is obtained and the yield of magnetic separation reaches 81.66%-86.57%. The results of X-ray diffraction （XRD） analysis and magnetism detection of the material before and after FRR indicate that weakly magnetic material is mainly converted into strongly magnetic material Fe304 with specific saturation magnetic moment. The efficiency of FRR is consistent with TFe recovery of magnetic separation; meantime, the specific sa^u＇ation magnetic moment increases from 33 to 42 times after FRR. Calculations show that speeds of flash magnetic roasting are obtained from several dozen to two or three hundred times, compared with rotary kiln or shaft furnace. This indicates that it is practicable to use the fast reducing roasting technique to improve the comprehensive utilization of iron ore resources.
基金Project(51474161)supported by the National Natural Science Foundation of China
文摘Compared with natural magnetite concentrate, artificial magnetite with more lattice defects and higher activity tends to be oxidized. And the artificial magnetite pellet at the temperature of 400℃ has the oxidation degree approaching to natural magnetite concentrate pellet fired at 1000℃. Besides, two kinds of pellets displayed quite different roasting characteristics. When preheated at the same temperature for the same period of time, natural magnetite concentrate pellet and artificial magnetite concentrate pellet need to be roasted at the temperature of 1100℃ and 1250℃, respectively, for 25 min to reach the compressive strength of 3000 N per pellet. When roasted at the same temperature of 1200℃, natural magnetite pellet and artificial magnetite pellet need to be roasted for 15 min and 30 min, respectively, to reach the compressive strength over 3000 N per pellet. It can be seen from the test that artificial magnetite pellet has a faster oxidation, resulting in the high porosity in the produced pellet, and it requires a roasting process at higher temperature for a longer time to reach the desired compressive strength for industrial production.
基金Projects(51874071,51734005,52104257)supported by the National Natural Science Foundation of ChinaProject(161045)supported by the Fok Ying Tung Education Foundation for Yong Teachers in the Higher Education Institutions of China。
文摘Oolitic hematite is an iron ore resource with rich reserves,complex composition,low grade,fine disseminated particle sizes,and a unique oolitic structure.In this study,a microwave-assisted suspension magnetization roasting technology was proposed to recover and utilize the ore.The results showed that under the conditions of microwave pretreatment temperature of 1050℃ for 2 min,a magnetic concentrate with an iron grade of 58.72%at a recovery of 89.32%was obtained by microwave suspension magnetization roasting and magnetic separation.Moreover,compared with the no microwave pretreatment case,the iron grade and recovery increased by 3.17%and 1.58%,respectively.Microwave pretreatment increased the saturation magnetization of the roasted products from 24.974 to 39.236(A∙m^(2))/kg and the saturation susceptibility from 0.179×10^(−3) m^(3)/kg to 0.283×10^(−3) m^(3)/kg.Microcracks were formed between the iron and gangue minerals,and they gradually extended to the core of oolite with the increase in the pretreatment time.The reducing gas diffused from outside to inside along the microcracks,which promoted the selective transformation of the weak magnetic hematite into the strong magnetic magnetite.
基金Project(2021JJ20062) supported by the Natural Science Foundation of Hunan Province,ChinaProject(2019XK2304) supported by Landmark Innovation Demonstration Project of Hunan Province,China+3 种基金Project(2022GK4058) supported by High-tech Industry Science and Technology Innovation Leading Project of Hunan Province,ChinaProject(2020CX038) supported by the Innovation Driven Project of Central South University,ChinaProject(2019YFC1907301) supported by the National Key R&D Program of ChinaProject(202006375018) supported by the China Scholarship Council。
文摘In this study,a roasting enhanced flotation process was proposed to recover LiMn_(2)O_(4) and grapite from waste lithium-ion batteries(LIBs).The effects of roasting temperature and time on the surface modification was investigated,and a series of analytical technologies were used to reveal process mechanism.The results indicate that LiMn_(2)O_(4) can be effectively separated from graphite via flotation after the roasting.The flotation grade of LiMn_(2)O_(4) was significantly increased from 63.10%to 91.36%after roasting at 550℃for 2 h.The TG-DTG analysis demonstrates that the difficulty in flotation separation of LiMn_(2)O_(4) from graphite is caused by the organic binder and electrolytes coating on their surfaces.The XRD,SEM,XPS,and contact angle analyses confirm that the organic films on the surfaces of those materials can be effectively removed by roasting,after which the wettability of LiMn_(2)O_(4) is regained and thus the surface wettability difference between the cathode and anode materials is increased significantly.The closed-circuit flotation test indicates that a LiMn_(2)O_(4) sample with high grade of 99.81%is obtained,while the recovery of LiMn_(2)O_(4) is as high as 99.40%.This study provides an economical and eco-friendly way to recycling waste LIBs.
基金Projects(51804136,52064021,52074136,51564021,52064022)supported by the National Natural Science Foundation of ChinaProjects(2019T120625,2019M652276)supported by the China Postdoctoral Science Foundation+2 种基金Project(20202ACB213002)supported by the Jiangxi Province Science Fund for Distinguished Young Scholars,ChinaProject(2019KY09)supported by the Program for Excellent Young Talents,JXUST Young Jinggang Scholars of Jiangxi Province,Merit-based Postdoctoral Research in Jiangxi Province,ChinaProjects supported by the Distinguished Professor Program of Jinggang Scholars,Chinain Institutions of Higher Learning,Jiangxi Province,China。
文摘A large amount of acid-leaching residue is produced during the conventional Zn hydrometallurgy process,and this residue has a large concentration of a variety of valuable metals.The purpose of this study was to evaluate the ability of a procedure that entails the use of sulfation roasting,water leaching,and chlorination leaching(blowing oxygen technique)to recover Pb and Ag,followed by cooling crystallization and the replacement of Ag with lead sheet,to realize the full recovery of all valuable metals from zinc acid-leaching residue;consequently,good results were achieved.The best results were obtained under the following conditions:a sulfuric acid at 70%of the raw material quality,roasting temperature of 300°C and roasting time of 2 h,followed by the process of leaching the roasted residue for 1 h by applying a water-to-solid ratio of 5꞉1 at room temperature.The recovery rates of Zn and Fe were 98.69%and 92.36%,respectively.The main parameters of the chlorine salt leaching system were as follows:Cl−concentration of 300 g/L,Fe3+concentration of 25 g/L,acid concentration of 2 mol/L,liquid-to-solid ratio of 9 mL:1 g,temperature of 90℃,and leaching time of 0.5 h;this leaching process was followed by filtration separation.These conditions resulted in high extents of leaching for Pb and Ag(i.e.,98.87%and 96.74%,respectively).Finally,the kinetics of the process of Ag leaching using Cl−ions in an oxygen-rich medium was investigated.It was found that the leaching process was controlled by the diffusion of the product layers,and the activation energy was 19.82 kJ/mol.
基金Project(51604131)supported by the National Natural Science Foundation of ChinaProject(2017FB084)supported by the Yunnan Province Applied Basic Research Project,ChinaProject(2018T20150055)supported by the Testing and Analyzing Funds of Kunming University of Science and Technology,China
文摘To recover Zn, Pb, Fe and Si from a low-grade mining ore in the Lanping basin, Yunnan Province, China, a novel technology using the roasting with pyrite and carbon followed by beneficiation and hydrochloric acid leaching was proposed. Firstly, several factors such as pyrite dosage, roasting temperature, carbon powder dosage, holding time and particle size affecting on the flotation performance of Zn(Pb) and magnetic separation performance of Fe were simultaneously examined and the optimum process parameters were determined. A flotation concentrate, containing 17.46% Zn and 3.93% Pb, was obtained, and the Zn and Pb recoveries were 86.04% and 69.08%, respectively. The obtained flotation tailing was concentrated by a low-intensity magnetic separator. The grade of iron increased from 5.45% to 43.45% and the recovery of iron reached 64.87%. Hydrochloric acid leaching was then carried out for the magnetic separation tailing and a raw quartz concentrate containing 81.05% SiO2 was obtained. To further interpret the sulfidation mechanism of smithsonite, surface morphology and component of the sample before and after reactions were characterized by XRD and EPMA-EDS. The aim was to achieve the comprehensive utilization of the low-grade mining ore.
基金Projects(N 182304020,N 172304045)supported by the Fundamental Research Funds for the Central Universities,ChinaProject(E 2017501073)supported by the Hebei Province Natural Science Fund,ChinaProject(2019 CDXYCL 0031)supported by the New Materials and New Metallurgical Technology Frontier Scientific and Technological Innovation,China。
文摘The separation of Al from the silicon-rich diasporic bauxite is of great significance in alumina production.Herein,we proposed a low-temperature ammonium sulfate roasting-water leaching process to extract aluminum from silicon-rich diasporic bauxite.Parameters including roasting temperature,dosage of ammonium sulfate,roasting time,and particle size of ore were investigated.Under the condition of roasting temperature of 400℃,roasting time of 5 h,ammonium sulfate dosage of 2.5 times of the theoretical value and ore particle size of 80−96μm,more than 98%leaching rate of aluminum was obtained.The phase transformation and mechanism during the roasting process were revealed by using X-ray diffraction,thermogravimetric analysis,differential thermal analysis,and scanning electron microscope methods.The diaspore and kaolinite phases in the silicon-rich diasporic bauxite could react with ammonium sulfate to form corresponding sulfates(NH_(4))_(3)Al(SO_(4))_(3) Al(SO_(4))_(2) and Al_(2)(SO_(4))_(3).The proposed technology could provide an effective method for the direct and separation of aluminum from silicon-rich diasporic bauxite.
基金Project(51564034)supported by the National Science Fund for Distinguished Regional Scholars,China
文摘This paper mainly investigated the antimony recovery from antimony-bearing dusts through reduction roasting process after the dust firstly oxidation roasted.CO–CO2 mixture gas was used as reducing agent,and the antimony-containing phase was reduced into Sb4O6,volatilized into smoke,and finally recovered through the cooling cylinder.The antimony recovery rate increased from 66.00 wt%to 73.81 wt%in temperature range of 650 to 800°C,and decreased with temperature increased further to 900°C due to the reduction of Sb4O6 to the nonvolatile Sb.Similarly,the CO partial pressure also played a double role in this test.Under optimized conditions of roasting temperature of 800°C,CO partial pressure of 7.5 vol%and roasting time of 120 min,98.40 wt%of arsenic removal rate and 80.40 wt%antimony recovery rate could be obtained.In addition,the“As2O3”product could be used for preparing ferric arsenate which realized the harmless treatment of it.
文摘The present study aims at the recovery of potassium from muscovite mica(which contains K_(2)O;~10 wt%)using NaCl-roasting coupled with H_(2)SO_(4)-leaching process.The preliminary acid leaching studies applying different mineral acids resulted in a potassium recovery of 8%−18%.The optimum leaching conditions for the maximum recovery were 4 mol/L H_(2)SO_(4),60 min leaching time and liquid-solid ratio 4 mL/g at 90℃.However,the roasting of muscovite with additive NaCl(muscovite:NaCl mass ratio of 1:1.00,900℃,45 min)followed by H_(2)SO_(4)-leaching(95℃,60 min)extracted potassium to the tune of 98%.Under similar roasting conditions,the H_(2)O-leaching process extracted only 60%of potassium.The effects of various roasting and leaching parameters such as temperature,time,NaCl concentration,acid concentration,liquid-solid ratio on potassium extraction were evaluated.The appearance of the sylvite(KCl)mineral phase in the NaCl-roasted muscovite and its disappearance in the acid/water leached residue confirmed the physical and chemical distortions of the muscovite crystal structure.The possible mechanism of potassium release from the complex muscovite structure was elucidated based on available literature substantiated by characterizations using X-ray diffraction(XRD)and scanning electron microscopy with energy dispersive X-rays spectroscopy(SEM-EDX).
基金Projects(51874071,52022019,51734005)supported by the National Natural Science Foundation of ChinaProject(161045)supported by the Fok Ying Tung Education Foundation for Yong Teachers in the Higher Education Institutions of China。
文摘Siderite,as an abundant iron ore,has not been effectively utilized,with a low utilization rate.In this study,the in-situ kinetics and mechanism of siderite during suspension magnetization roasting(SMR)were investigated to improve the selective conversion of siderite to magnetite and CO,enriching the theoretical system of green SMR using siderite as a reductant.According to the gas products analyses,the peak value of the reaction rate increased with increasing temperature,and its curves presented the feature of an early peak and long tail.The mechanism function of the siderite pyrolysis was the contraction sphere model(R_(3)):f(α)=3(1−α)2/3;E_(α)was 46.4653 kJ/mol;A was 0.5938 s^(−1);the kinetics equation was k=0.5938exp[−46.4653/(RT)].The in-situ HT-XRD results indicated that siderite was converted into magnetite and wüstite that exhibited a good crystallinity in SMR under a N_(2) atmosphere.At 620℃,the saturation magnetization(M_(s)),remanence magnetization(Mr),and coercivity(Hc)of the product peaked at 53.63×10^(-3)A·m^(2)/g,10.23×10^(-3)A·m^(2)/g,and 12.40×10^(3)A/m,respectively.Meanwhile,the initial particles with a smooth surface were transformed into particles with a porous and loose structure in the roasting process,which would contribute to reducing the grinding cost.
基金Project(51804346)supported by the National Natural Science Foundation of ChinaProject(2019JJ50767)supported by the Natural Science Foundation of Hunan Province,ChinaProject(KY[2017]125)supported by Youth Foundation of Guizhou Education Department,China。
文摘The growing characteristics of metallic iron particles during reductive roasting of boron-bearing magnetite concentrate under different conditions were investigated.The size of the metallic iron particles was quantitatively measured via optical image analysis with consideration of size calibration and weighted ratio of image numbers in the core,middle and periphery zones of cross-section of pellets.In order to guarantee the measurement accuracy,54 images were captured in total for each specimen,with a weighted ratio of 1:7:19 with respect to the core,middle and periphery section of the cross-section of pellets.Increasing reduction temperature and time is favorable to the growth of metallic iron particles.Based on the modification of particle size measurement,in terms of time(t)and temperature(T)a predicting model of metallic iron particle size(D),was established as:D=125−0.112t−0.2352T−5.355×10^−4t^2+2.032×10^−4t∙T+1.134×10^−4T^2.
