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.

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.

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.

Pyrometallurgical recycling of spent lithium-ion batteries from conventional roasting to synergistic pyrolysis with organic wastes

The synergistic pyrolysis has been increasingly used for recycling spent lithium-ion batteries(LIBs)and organic wastes(hydrogen and carbon sources),which are in-situ transformed into various reducing agents such as H_(2),CO,and char via carbothermal and/or gas thermal reduction.Compared with the conventional roasting methods,this“killing two birds with one stone”strategy can not only reduce the cost and energy consumption,but also realize the valorization of organic wastes.This paper concluded the research progress in synergistic pyrolysis recycling of spent LIBs and organic wastes.On the one hand,valued metals such as Li,Co,Ni,and Mn can be recovered through the pyrolysis of the cathode materials with inherent organic materials(e.g.,separator,electrolyte)or graphite anode.During the pyrolysis process,the organic materials are decomposed into char and gases(e.g.,CO,H_(2),and CH_(4))as reducing agents,while the cathode material is decomposed and then converted into Li_(2)CO_(3) and low-valent transition metals or their oxides via in-situ thermal reduction.The formed Li_(2)CO_(3) can be easily recovered by the water leaching process,while the formed transition metals or their oxides(e.g.,Co,CoO,Ni,MnO,etc.)can be recovered by the reductant-free acid leaching or magnetic separation process.On the other hand,organic wastes(e.g.,biomass,plastics,etc.)as abundant hydrogen and carbon sources can be converted into gas(e.g.,H_(2),CO,etc.)and char via pyrolysis.The cathode materials are decomposed and subsequently reduced by the pyrolysis gas and char.In addition,the pyrolysis oil and gas can be upgraded by catalytic reforming with the active metals derived from cathode material.Finally,great challenges are proposed to promote this promising technology in the industrial applications.

Fluidized magnetization roasting of refractory siderite-containing iron ore via preoxidation-low-temperature reduction

Magnetization roasting is one of the most effective way of utilizing low-grade refractory iron ore.However,the reduction roasting of siderite(FeCO3)generates weakly magnetic wüstite,thus reducing iron recovery via weak magnetic separation.We systematically studied and proposed the fluidized preoxidation-low-temperature reduction magnetization roasting process for siderite.We found that the maghemite generated during the air oxidation roasting of siderite would be further reduced into wüstite at 500 and 550℃due to the unstable intermediate product magnetite(Fe_(3)O_(4)).Stable magnetite can be obtained through maghemite reduction only at low temperature.The optimal fluidized magnetization roasting parameters included preoxidation at 610℃for 2.5 min,followed by reduction at 450℃for 5 min.For roasted ore,weak magnetic separation yielded an iron ore concentrate grade of 62.0wt%and an iron recovery rate of 88.36%.Compared with that of conventional direct reduction magnetization roasting,the iron recovery rate of weak magnetic separation had greatly improved by 34.33%.The proposed fluidized preoxidation-low-temperature reduction magnetization roasting process can realize the efficient magnetization roasting utilization of low-grade refractory siderite-containing iron ore without wüstite generation and is unlimited by the proportion of siderite and hematite in iron ore.

Reaction behavior and non-isothermal kinetics of suspension magnetization roasting of limonite and siderite

In order to develop limonite and decrease CO_(2) emissions,siderite is proposed as a clean reductant for suspension magnetization roasting(SMR) of limonite.An iron concentrate(iron grade:65.92wt%,iron recovery:98.54wt%) was obtained by magnetic separation under the optimum SMR conditions:siderite dosage 40wt%,roasting temperature 700℃,roasting time 10 min.According to the magnetic analysis,SMR achieved the conversion of weak magnetic minerals to strong magnetic minerals,thus enabling the recovery of iron via magnetic separation.Based on the phase transformation analysis,during the SMR process,limonite was first dehydrated and converted to hematite,and then siderite decomposed to generate magnetite and CO,where CO reduced the freshly formed hematite to magnetite.The microstructure evolution analysis indicated that the magnetite particles were loose and porous with a destroyed structure,making them easier to be ground.The non-isothermal kinetic results show that the main reaction between limonite and siderite conformed to the two-dimension diffusion mechanism,suggesting that the diffusion of CO controlled the reaction.These results encourage the application of siderite as a reductant in SMR.

Interactive effect of hot air roasting processes on the sensory property,allergenicity,and oil extraction of sesame(Sesamum indicum L.)seeds

Sesame seeds are a healthy food ingredient and an oil crop for sesame oil production;however,it has recently been recognized as an essential allergenic food by FAO/WHO.This research investigated the relationship between the hot air roasting process(at 120,150,and 180℃ for 10,20,and 30 min)and several quality attributes of sesame seeds since roasting is the key process for preparing sesame seeds for both consumption and oil production.The hot air process followed the central composite design.The changes of sesame in terms of color,sensory properties(odor,texture,color,and taste),allergenicity caused by oleosins(ses i 4 and ses i 5),as well as oil extraction and quality were monitored using a colorimeter,sensory evaluation panelists,ELISA,as well as oil yield and acid value,respectively.Roasting temperature influenced the product quality more than roasting time,although the two processing parameters significantly interacted with each other(P<0.001).Sensory evaluation indicated medium roasting generated attractive flavor,order,appearance,and crispy texture.Allergenicity was high in sesame seeds after high-temperature roasting,according to IgE binding capacity test.Sesame oil extraction was favored by high-temperature roasting,which,however,adversely affected the oil quality.The optimal roasting conditions were 150.5℃ for 15 min for optimized sesame seeds quality in terms of sensory properties and allergenicity,while roasting at 158℃ for 10 min was optimal for sesame oil production.The finding will benefit the sesame seed industry.

Characterization and Pyrometallurgical Removal of Arsenic from Copper Concentrate Roasting Dust

This paper describes the experimental results of removing arsenic from the dust collected in electrostatic precipitators of a fluidized bed roasting furnace (RP dust). The fluidized bed roasting process generates 600 kilotons of copper concentrate per year with 3 - 6 wt% of concentration of arsenic, producing a roasted product with a low content of arsenic below 0.3 wt%. The process generates 27 kilotons of RP dust per year with a concentration of arsenic of the order of 5 wt% and copper concentration of around 20 wt%. Subsequently, the dust collected in the electrostatic precipitators is treated by hydrometallurgical methods allowing the recovery of copper, and the disposition of arsenic as scorodite. This work proposes to use a pyrometallurgy process to the volatilization of arsenic from RP dust. The obtained material can be recirculated in copper smelting furnaces allowing the recovery of valuable metals. The set of experiments carried out in the roasting of the mixture of copper concentrate/RP dust and sulfur/RP dust used different ratios of mixtures, temperatures and roasting times. By different techniques, the characterization of the RP dust determined its size distribution, morphology, and chemical and mineralogical composition. RP dust is a composite material of small particles (<5 μm) in 50 μm agglomerates, mostly amorphous, with a complex chemical composition of sulfoxides. The results of the roasting experiments indicated that for a 75/25 weight ratio of the mixture of the copper concentrate/PR dust under 700℃, 15 minutes of roasting time with injection of air, the volatilization of arsenic reached 96% by weight. The arsenic concentration after the roasting process is less than 0.3% by weight. For a 5/95 mixture of sulfur/RP dust, at 650℃, the volatilization of arsenic reached a promissory result of 67%. Even that this study was carried out for a particular operation, the results have the potential to be extended to dust produced in the roasting of concentrates of nickel, lead-zinc, and gold.

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.

Study on Roasting Decomposition of Mixed Rare Earth Concentrate in CaO-NaCl-CaCl_2

The decomposed process of bastnaesite, monazite and mixed rare earth concentrate in CaO-CaCl-CaCl2 was studied by means of TG-DTA method. The relationship among decomposition ratio, roasting temperature, and CaO-NaCl addition was studied by the quadratic regression orthogonal analysis, and then the regression equation was obtained. Through analysis, the optimum process conditions of mixed rare earth concentrate decomposed by CaO-CaCl-CaCl2 were obtained as follows: roasting temperature: 700 ℃, CaO addition: 15%, NaCl-CaCl2 addition: 10%, roasting time: 60 min, the decomposition ratio: 91.3%.

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.

Recovery of valuable metals from a low-grade nickel ore using an ammonium sulfate roasting-leaching process

Metal leaching from a low-grade nickel ore was investigated using an ammonium sulfate roasting-water leaching process. The nickel ore was mixed with ammonium sulfate, followed by roasting and finally leaching with water. During the process the effects of the amotmt of ammonium sulfate, roasting temperature, and roasting time on the leaching recovery of metal elements were analyzed. The optimum technological parameters were determined as follows： ammonium sulfate/ore ratio, 0.8 g/g; roasting temperature, 400℃; and roasting time, 2 h. Under the optimum condition the leaching recoveries ofNi, Cu, Fe, and Mg were 83.48%, 76.24%, 56.43%, and 62.15%, respectively. Furthermore, the dissolution kinetics of Ni and Mg from the nickel ore was studied. The apparent activation energies for the leaching reaction of Ni and Mg were 18.782 and 10.038 kJ.mo1-1, which were consistent with the values of diffusion control reactions. Therefore, the results demonstrated that the leaching recoveries of Ni and Mg were controlled by diffusion.

Thermodynamics and kinetics of alumina extraction from fly ash using an ammonium hydrogen sulfate roasting method

A novel method was developed for extracting alumina （Al2O3） from fly ash using an ammonium hydrogen sulfate （NH4HSO4） roasting process, and the thermodynamics and kinetics of this method were investigated. The thermodynamic results were verified experi-mentally. Thermodynamic calculations show that mullite present in the fly ash can react with NH4HSO4 in the 298-723 K range. Process op-timization reveals that the extraction rate can reach up to 90.95% when the fly ash reacts with NH4HSO4 at a 1：8 mole ratio of Al2O3/NH4HSO4 at 673 K for 60 min. Kinetic analysis indicates that the NH4HSO4 roasting process follows the shrinking unreacted core model, and inner diffusion through the product layer is the rate-controlling step. The activation energy is calculated to be 16.627 kJ/mol;and the kinetic equation can be expressed as 1-（2/3）α-（1-α）2/3=0.0374t exp[-16627/（RT）], whereαis the extraction rate and t is the roasting temperature.

Magnetizing Roasting Mechanism and Effective Ore Dressing Process for Oolitic Hematite Ore

Magnetizing roasting of oolitic hematite ore from western Hubei Province was investigated.The mechanism for reduction roasting of oolitic hematite ore was discussed and analyzed.It is found that flash magnetizing roasting-magnetic separation process is a promising approach for the processing of oolitic hematite ore from western Hubei Province.

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.

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

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

Upgrading and dephosphorization of Western Australian iron ore using reduction roasting by adding sodium carbonate

The technology of direct reduction by adding sodium carbonate （Na2CO3） and magnetic separation was developed to treat Western Australian high phosphorus iron ore. The iron ore and reduced product were investigated by optical microscopy and scanning electron microscopy. It is found that phosphorus exists within limonite in the form of solid solution, which cannot be removed through traditional ways. During reduction roasting, Na2CO3 reacts with gangue minerals （SiO2 and A1203）, forming aluminum silicate-containing phosphorus and damaging the ore structure, which promotes the separation between iron and phosphorus during magnetic separation. Meanwhile, Na2CO3 also improves the growth of iron grains, increasing the iron grade and iron recovery. The iron concentrate, assaying 94.12wt% Fe and 0.07wt% P at the iron recovery of 96.83% and the dephosphorization rate of 74.08%, is obtained under the optimum conditions. The final product （metal iron powder） after briquetting can be used as the burden for steelmaking by an alactrie a.re furnace to rer）la,ce scrar） steel.

Extraction of metals from complex sulfide nickel concentrates by low-temperature chlorination roasting and water leaching

The recovery of valuable metals from complex sulfide concentrates was investigated via chlorination roasting followed by water leaching. A reaction process is proposed on the basis of previous studies and the results of our preliminary experiments. During the process,various process parameters were studied,including the roasting temperature,the addition of NH_4 Cl,the roasting time,the leaching time,and the liquid-to-solid ratio. The roasted products and leach residues were characterized by X-ray diffraction and vibrational spectroscopy. Under the optimum condition,95% of Ni,98% of Cu,and 88% of Co were recovered. In addition,the removal of iron was studied in the water leaching stage. The results demonstrate that this process provides an effective approach for extracting multiple metals from complex concentrates or ores.

Fluidized roasting reduction kinetics of low-grade pyrolusite coupling with pretreatment of stone coal

Based on the fluidized roasting reduction technology of low-grade pyrolusite coupling with pretreatment of stone coal, the manganese reduction efficiency was investigated and technical conditions were optimized. It is found that the optimum manganese reduction efficiency can be up to 98.97% under the conditions that the mass ratio of stone coal to pyrolusite is 3：1, the roasting temperature of stone coal is 1000℃, the roasting temperature of pyrolusite is 800℃, and the roasting time is 2 h. Other low-grade pyrolusite ores in China from Guangxi, Hunan, and Guizhou Provinces were tested and all these minerals responded well, giving -99% manganese reduction efficiency. Meanwhile, the reduction kinetic model has been established. It is confirmed that the reduction process is controlled by the interface chemical reaction. The apparent activation energy is 36.397 kJ/mol.