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.

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.

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.

Preparation of sodium manganate from low-grade pyrolusite by alkaline predesilication-fluidized roasting technique

Low concentration alkaline leaching was used for predesilication treatment of low-grade pyrolusite. The effects of initial NaOH concentration, liquid-to-solid ratio, leaching temperature, leaching time and stirring speed on silica leaching rate were investigated and the kinetics of alkaline leaching process was studied. The results show that silica leaching rate reached 91.2% under the conditions of initial NaOH concentration of 20%, liquid-to-solid ratio of 4：1, leaching temperature of 180 ℃, leaching time of 4 h and stirring speed of 300 r/min. Shrinking-core model showed that the leaching process was controlled by the chemical surface reaction with activation energy Ea of 53.31 k J/mol. The fluidized roasting conditions for preparation of sodium manganate were optimized by the orthogonal experiments using the desiliconized residue. The conversion rate of sodium manganate was obtained to be 89.7% under the conditions of silica leaching rate of 91.2%, NaOH/MnO2 mass ratio of 3：1, roasting temperature of 500 ℃ and roasting time of 4 h, and it increased with the increase of silicon leaching rate.

Separation of alumina and silica from metakaolinite by reduction roasting-alkaline leaching process: Effect of CaSO_(4) and CaO

Limestone(CaCO3),which could promote sulfur fixation,was added to coal gangue during roasting in a circulating fluidized bed(CFB)boiler.CaO and CaSO_(4) were the main Ca-bearing minerals while metakaolinite was the major Al-bearing mineral in CFB slag.The effect of CaSO_(4) and CaO on the separation of alumina and silica from metakaolinite by reduction roasting−alkaline leaching process was studied.Results showed that metakaolinite was completely converted into hercynite and silica solid solutions(i.e.,quartz and cristobalite solid solutions)by reduction roasting with hematite.More than 95%of silica in the reduced specimen was removed by alkaline leaching.The addition of CaSO_(4) and CaO remarkably decreased the separation efficiency of alumina and silica in metakaolinite,which could be attributed to the formation of Si-bearing minerals:(1)Fayalite and anorthite were formed during the reduction roasting process;(2)Fayalite was stable while anorthite was converted into sodalite and wollastonite during the alkaline leaching process.This study demonstrates that sulfur in coal gangue should be fixed by treating the exhaust gas instead of controlling the combustion process of CFB to achieve the comprehensive recovery of silica and alumina from the CFB slag.

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.

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.

CFD simulation for reduction of pyrolusite in fluidized beds

In this study,a CFD model coupled with heterogeneous flow structure,mass transfer equations,and chemical reaction kinetics is established to forecast the pyrolusite reduction reaction behavior.Compared with the previous studies which ignore the volume change of solids phase,the influence of volume shrinkage on reaction and flow behavior is explored in this research.Volume shrinkage of pyrolusite is proved to be non-negligible in predicting the conversion rate.The negligence of volume shrinkage leads to the overestimation of conversion rate for its inaccurate estimation of surface area for reaction.Besides,the influence of volume shrinkage on the reaction is found smaller in the scaled-up reactor.

红土镍矿酸浸渣综合利用进展

近年来随着不锈钢及新能源用镍需求的增长,红土镍矿湿法提镍工艺得到更多应用,相应的红土镍矿酸浸渣堆存量也迅速增加,对此类固废进行科学消纳和资源综合利用迫在眉睫。针对红土镍矿酸浸渣有害元素硫、有价金属铁含量高的特点进行综合利用新技术、新装备研究对红土镍矿湿法冶金工业的持续发展具有重要的理论和现实意义。主要介绍红土镍矿酸浸渣的湿法冶金形成过程及其成分组成和物理化学性质,较为系统地梳理了红土镍矿酸浸渣在尾矿坝堆积、地下压滤回填和深海填埋等直接处理方法,提取利用铁、铬、铝、稀土等有价金属资源,消纳制备硅藻土、陶粒、矿物纤维等建筑材料、制备锂电池电极、沸石等高附加值材料综合利用技术上的研究进展,展望了红土镍矿酸浸渣流态化提铁降硫技术及装备的发展前景。

难选高铁软锰矿流态化磁化焙烧-磁选浸出工艺

【目的】提高难选高铁软锰矿的资源利用效率,缓解我国锰矿资源依赖进口的现状。【方法】采用流化床反应器作为焙烧装置,模拟发生炉煤气气体组分作为还原气体,对云南省某地区的难选高铁软锰矿进行流态化磁化焙烧。对还原后的焙烧矿进行磨矿-弱磁磁选,得到铁精矿和锰精矿。最后,对锰精矿中的锰进行酸浸浸出。【结果】在500℃流态化还原焙烧10 min,该难选高铁软锰矿即可达到最佳的还原状态效果,当样品的磨矿粒径<30μm的颗粒的质量分数为65%时,即可达到最佳的解离效果。选别后铁精矿铁品位和回收率分别为57.24%和76.55%,锰精矿锰品位和回收率分别为24.41%和73.84%,锰精矿的铁浸出率仅有1.94%。【结论】考察难选高铁软锰矿流态化磁化焙烧-磁选浸出规律,可以降低反应温度,缩短反应时间,低碳节能;抑制过还原反应的发生,可以提高铁资源利用率,降低浸出酸耗和除铁剂消耗。低能耗、低消耗和高利用率可以极大地提高难选高铁软锰矿的经济性,解决这部分资源难于利用的困境。

亚硫酸铵还原焙烧软锰矿提取锰试验研究

研究了以亚硫酸铵为还原剂,通过高温焙烧直接把软锰矿中的二氧化锰还原成硫酸锰,从而实现软锰矿的高效提取,考察了亚硫酸铵与软锰矿质量比、焙烧温度及时间、浸出温度及时间等对锰提取率的影响,确定了最佳工艺条件。结果表明:最优条件为亚硫酸铵与软锰矿质量比0.75∶1,焙烧温度500℃,焙烧时间1.5h,浸出温度25℃,浸出时间30min;在该条件下,锰浸出率约为93%,铁浸出率约9.5%。锰提取效果较好。

流化床处理粉铁矿工艺研究

评述了流化床处理粉铁矿的工艺,根据流化床反应器在工业中的应用实际,分别评述了流化床磁化焙烧生产铁精矿,流化床预热粉铁矿并低预还原度还原后添加到铁浴熔池,流化床生产直接还原铁(DRI)和碳化铁(Fe3C)作电炉炼钢原料,多级流化床生产较高预还原度DRI供熔融气化炉终还原冶炼铁水等工艺中流化床反应器的应用。

生物质焦焙烧还原低品位软锰矿及其动力学

采用生物质焦和活性炭粉作还原剂,在管式炉中进行了低品位软锰矿焙烧还原对比试验.分别研究了焙烧温度、焙烧时间、生物质焦用量等条件对软锰矿还原率的影响,对焙烧产物进行了X射线衍射分析.结果表明,生物质焦在焙烧时间和还原效率上优于活性炭粉;软锰矿焙烧还原依次经历Mn O2→Mn2O3→Mn3O4→Mn O过程;在焙烧温度为800℃,焙烧时间为50 min,生物质焦用量为10%时,软锰矿还原率可达98%以上,在此基础上导出了还原动力学方程,并证实还原过程由界面化学反应控制,表观活化能为43.896 k J·mol-1.

某进口软锰矿还原焙烧、浸出实验研究

根据某进口软锰矿石中锰矿物的赋存形态,进行了还原焙烧、硫酸浸出工艺实验研究。结果表明,适宜的还原焙烧条件为:还原剂烟煤加入量(相对锰矿质量)12%、焙烧温度900℃、焙烧时间60 min,该条件下所得焙烧矿在液固比9∶1、硫酸浓度120 g/L、搅拌速度450 r/min条件下浸出时间60 min,锰浸出率达96.83%。该工艺可为进口软锰矿的利用提供技术依据。

广西某低品位软锰矿还原焙烧—酸浸工艺研究

针对广西某低品位软锰矿,探究以煤为还原剂进行还原焙烧,焙烧矿采用硫酸浸出的工艺条件,通过单因素实验,考察了还原剂用量、焙烧温度与时间的组合、浸出酸量、时间、液固比、搅拌强度及浸出温度对浸出结果的影响。实验表明:在取300g软锰矿进行实验时,配煤11%、在温度750℃下焙烧60min的焙烧矿,在理论酸量、固液比5∶1、搅拌强度300r/min、常温下浸出45min后,锰的浸出率可达到95.57%的良好指标。

二氧化锰流态化还原试验研究

针对现有氧化锰矿还原处理工艺中效率低、能耗高、操作条件差、环境污染较严重等问题,利用气固流态化高效传质传热原理,将流态化快速还原技术应用于氧化锰矿的还原焙烧。广西大新锰矿矿石中的锰矿物主要以软锰矿(β-MnO2)和复水锰矿(MnO(OH)2)的形式存在,利用流态化状态下,气固传热效率高、接触面积大的特点,强化氧化锰矿石的还原焙烧反应,在750～850℃的温度条件下,CO含量3%～6%,获得了转化率大于90%的指标,焙烧时间只需10～60 s,证实了氧化锰矿(MnO2·nH2O)在数十秒钟实现流态化快速还原焙烧的科学设想,将对工业化开发利用低品位锰矿资源有一定的推动作用。

石油焦还原焙烧进口软锰矿的工艺研究

对某含锰40.73%的进口软锰矿进行了还原焙烧⁃硫酸常温浸出试验研究。开展了还原焙烧温度、焙烧时间、还原剂(石油焦)用量条件试验,确定最佳还原焙烧工艺条件为:焙烧温度900℃、焙烧时间60 min、石油焦用量15%,此条件下焙烧获得的焙烧矿经硫酸常温浸出,可获得锰浸出率95.61%的良好指标,为进口软锰矿的有效湿法利用提供了技术支持。

软锰矿流态化低温还原实验研究

采用流化床反应器作为焙烧装置,模拟发生炉煤气气体组分作为还原气体,考察了云南某地区软锰矿还原焙烧的影响因素。结果表明:煤气中氧化性气体组分的存在降低了该锰矿的还原效率。在700℃以下还原得到的产品,在隔绝氧气的条件下静置1 h或者直接投入水中,均表现出良好的抗氧化特性。通过对工艺参数进行优化,发现在模拟发生炉煤气气氛下,500℃流态化还原焙烧8 min获得的产品经过防氧化处理后,锰的浸出率可以达到95%以上。

某高镁低品位软锰矿锰镁分离试验

高镁低品位软锰矿石传统的锰镁分离工艺均存在污染严重、锰镁离子分离效率低等问题。为了实现高镁低品位软锰矿的高效、低污染开发利用,以广西某高镁低品位软锰矿为原料,对采用酸化还原焙烧+尾气(SO2)还原软锰矿矿浆—还原产物合并浸出—浸出液除杂工艺制得的高纯Mn SO4与Mg SO4混合溶液,进行了NH4HCO3沉Mn2+(锰镁高效分离)工艺条件研究。结果表明,在NH4HCO3与Mn SO4物质的量之比为2.25,反应时间为60 min,反应温度为30℃,搅拌速度为120 r/min情况下,锰镁分离率可达96.96%。该锰镁分离工艺既解决了酸化还原焙烧尾气(SO2)的高效回收利用问题,又高效地实现了浸出液中锰镁离子的分离。

回转炉还原焙烧低品位软锰矿及其尾气资源化利用的动力学研究

对某低品位软锰矿进行还原焙烧试验并考察低品位软锰矿浆对焙烧反应过程中产生的尾气(二氧化硫)的资源化利用及其动力学模型。实验结果表明:浓硫酸加入量为锰含量的1.5倍,并在400℃下焙烧120 min,软锰矿锰还原率在98%以上。当矿浆固液比为6:1,吸收温度为80℃时,低品位软锰矿浆中一级锰浸出率为98.60%,二级锰浸出率为98.61%;根据低品位软锰矿浆尾气资源化利用的反应动力学模型,反应的表观活化能为16.04k J/mol。本工艺利用回转炉高效还原低品位软锰矿并实现低品位软锰矿浆对尾气的充分资源化利用,对锰资源高效充分利用的同时,实现反应体系硫资源的循环利用,以达到对工艺流程的绿色创新,为低品位软锰矿资源的高效工业化利用提供了参考和依据。