高磷鲕状赤铁矿煤基还原焙烧制备直接还原铁

针对鄂西高磷鲕状赤铁矿,采用煤基还原焙烧-磁选工艺制备直接还原铁,研究了还原剂用量、焙烧温度、焙烧时间、助熔剂等对还原焙烧效果的影响规律。研究结果表明:在焙烧温度为1 100℃,焙烧时间为50 min,还原剂用量为30%,助熔剂为碳酸钠和硫酸钠、用量分别为15%和30%时,磨矿磁选后获得直接还原铁的铁品位91.13%,铁的回收率78.87%,残留S含量0.03%,P含量0.09%,满足电炉炼钢原料要求。本文方法为同类型铁矿石的综合开发利用提供了充分的技术支持。     

焙烧温度对高铁提钒尾渣煤基直接还原效果的影响

利用X射线衍射(XRD)、扫描电镜(SEM)等手段考察了焙烧温度对提钒尾渣煤基直接还原效果的影响,并利用磨矿-磁选方法对焙烧产物进行了金属铁的分离实验.结果表明:最有利于金属铁析出、兼并、长大以及金属铁和渣相单体解离的焙烧温度是1200℃,在此温度下,提钒尾渣中的Fe2O3基本还原成了金属铁,Fe2TiO5基本转变成了金属铁和TiO2.铁质量分数36.54%、TiO2质量分数9.28%的提钒尾渣,经1200℃焙烧所得产物经过二段磨矿-二段磁选可获得铁质量分数90.90%、TiO2质量分数0.56%的金属铁粉.     

宣龙式铁矿焙烧还原-磁选工艺及其影响因素

基于煤基焙烧还原-磁选工艺,进行了宣龙式难选鲕状赤铁矿石提铁过程及其影响因素的实验研究.以铁精矿品位和铁回收率为评价指标,确定了适合于该类矿石的最佳工艺条件:焙烧还原温度为1 200℃,还原剂用量为30%,焙烧还原时间为60min,焙烧产物磁选前的磨矿细度为-45μm占96.19%,磁选的磁场强度为111kA.m-1.在该工艺条件下,可以使铁精矿品位达到92.53%,铁回收率达到90.78%.     

煤种对钛磁铁矿直接还原-磁选钛铁分离的影响

研究了烟煤和无烟煤对海滨钛磁铁矿直接还原-磁选钛铁分离的影响机理.结果表明,在试验用量范围内,两种煤对还原铁指标的影响规律相近,煤用量低时钛磁铁矿还原不充分.随煤用量增加,被还原的金属铁越来越多,但粒度较小,与其他颗粒嵌布紧密,因此还原铁Fe品位低,Ti O2品位高,铁回收率则先提高后基本不变.所有煤用量下所得金属铁颗粒均纯净.和无烟煤相比,烟煤固定碳较低,还原气氛较弱,但灰分较高,有利于金属铁颗粒的聚集长大;因此相同用量的烟煤为还原剂时,焙烧矿中金属铁颗粒较少,但粒度较大,还原铁中Fe品位较高,铁回收率较低,Ti O2品位较低.     

红土镍矿直接还原焙烧磁选回收铁镍

采用添加助熔剂直接还原焙烧-磁选方法,对镍主要以硅酸镍形式存在的低品位红土镍矿中镍和铁的富集进行了研究. 结果表明,同时添加助熔剂,可获得较好的技术指标. 最佳工艺条件为:煤作还原剂,质量分数为15%;KD-2为助熔剂,质量分数为20%;焙烧温度为1200℃;焙烧时间为40min. 在此条件下可以得到镍品位10.83%、铁品位52.87%、镍回收率82.15%和铁回收率54.59%的镍铁精矿. 用X射线衍射(XRD)和透射电镜(TEM)对还原过程中助熔剂和煤的作用机理进行了研究. 发现KD-2可以与原矿中含镍的石英和硅酸盐矿物反应,释放出其中的镍;煤用量太多时可生成部分不含镍的金属铁,会造成镍的回收率降低.     

深度还原-弱磁选回收稀土尾矿中铁的试验研究

对某全铁品位为1625%的稀土尾矿进行了深度还原-弱磁选回收铁试验研究,研究了还原剂种类及用量、焙烧温度及时间、磨矿细度及磁场强度对铁精矿品位和回收率的影响,并采用SEM,XRD等手段对稀土尾矿、焙烧产物、铁精矿进行了测试.结果表明,在烟煤质量分数30%,焙烧温度1300℃,焙烧时间60min,磨矿细度-0074mm占75%,磁场强度118kA/m的条件下,所得铁精矿TFe品位可达8076%,铁回收率可达9324%;稀土尾矿经深度还原后,其中的赤、褐铁矿、硅酸铁等含铁矿物转化为单质铁,铁精矿品位和回收率较常规选矿方法大幅度提高,同时脉石矿物组成简单,有利于萤石的富集回收.     

某高磷铁矿提铁降磷研究

以湖南某地高磷铁矿为原料,采用还原焙烧一磁选一硫酸浸出工艺进行提铁降磷试验研究.对还原焙烧一磁选粗精矿进行硫酸浸出工艺参数优化,对浸出时间、液固比、硫酸用量和搅拌速度等因素对提铁降磷效果的影响进行研究.研究结果表明:对原矿品位为47.28%Fe(质量分数)和磷含量为1.59%的高磷铁矿石经过还原焙烧一磁选得到的粗精矿,在浸出时间为2h、液固比为2.5、硫酸用量为50 kg/t和搅拌速度为500 r/min的条件下进行酸性浸出提铁降磷,最终得到铁精矿品位达62.35%Fe,磷含量为0.20%,铁总回收率为90.54%和脱磷率为87.42%.     

鄂西高磷鲕状赤铁矿直接还原焙烧同步脱磷机理

采用XRD和SEM分析方法研究了鄂西高磷鲕状赤铁矿添加脱磷剂后直接还原焙烧的产物及磁选后的最终产品. 结果表明,还原焙烧过程中添加的脱磷剂除具有脱磷效果外,对铁的还原也有促进作用. 脱磷剂可以使部分磷转化为易去除的可溶性磷酸盐,同时破坏鲕粒结构,使细磨-磁选后铁相易与脉石矿物分离从而达到脱磷效果,并且可以提高产品中铁的品位和回收率.     

高铁铝土矿直接还原—溶出工艺

提出了一种以Na2CO3为添加剂、以煤为还原剂的还原分离方法,将原矿中铁的氧化物还原为铁单质粉末通过磁选分离回收,将水铝石矿物转化为铝酸钠溶出分离回收.通过单因素实验考察了还原温度、还原时间、Na2CO3用量和还原剂用量对粉末铁品位、铁回收率和氧化铝溶出率的影响,并用X射线衍射分析、扫描电镜观察和能谱分析等方法研究了反应的过程和机理.通过正交试验优化了实验参数,获得的最优条件为还原温度1150℃,还原时间45 min,Na2CO3用量40.47%,还原剂用量11.9%;在最优条件下,粉末铁品位为95.88%,铁回收率为89.92%,氧化铝溶出率为75.92%.     

云南惠民铁矿微波磁化焙烧工艺

基于云南惠民铁矿主要为细粒浸染结构、氧化矿的主要铁矿物为褐铁矿,以氧化矿为研究对象,采用微波磁化焙烧-弱磁选工艺分选铁矿石,考查微波焙烧温度、还原剂用量、磁选磁场强度对分选指标的影响。研究结果表明:在焙烧温度为800℃,还原剂用量为12%,还原时间为12 min,磁选磁场强度为119.37 kA/cm的条件下,获得铁精矿品位为59.31%,回收率为81.92%,证实微波磁化焙烧铁矿石的方法可行,为难选铁矿石的分选提供了一种新的思路。     

Effects and mechanisms of fluorite on the co-reduction of blast furnace dust and seaside titanomagnetite

The co-reduction roasting and grinding magne -tic separation of seaside titanomagnetite and blast furnace dust was investigated with and without fluorite addition at a reduction roasting temperature of 1250℃ for 60 min, a grinding fineness of -43 μm accounting for 69.02wt% of the total, and a low-intensity magnetic field strength of 151 kA/m. The mineral composition, microstructure, and state of the roasted products were analyzed, and the concentrations of CO and CO₂ were analyzed in the co-reduction roasting. Better results were achieved with a small fluorite dosage (≤ 4wt%) in the process of co-reduction. In addition, F- was found to reduce the melting point and viscosity of the slag phase because of the high content of aluminate and silicate minerals in the blast furnace dust. The low moisture content of the blast furnace dust and calcic minerals inhibited the hydrolysis of CaF₂ and the loss of F-. Compared with the blast furnace dust from Chengdeng, the blast furnace dusts from Jiugang and Jinxin inhibited the diffusion of F- when used as reducing agents, leading to weaker effects of fluorite.     

深度还原分离回收含铌铁尾矿中铌和铁的试验研究

以某稀土综合尾矿经磨矿-磁选-浮选处理后的含铌铁尾矿为对象,采用深度还原焙烧的方法分离回收铌和铁,研究还原焙烧条件对铌、铁分离效果的影响。结果表明,还原剂种类对铁回收率的影响较为显著,对铌的分离回收影响相对较小,还原剂为褐煤时铁回收率最高;还原时间的延长、焙烧温度的升高以及助熔剂用量的增加均有利于铌、铁的分离回收;在还原剂褐煤用量为10%、助熔剂用量为15%、还原时间为60min、还原温度为1300℃的条件下可实现含铌铁尾矿中铌、铁的高效分离回收,得到w(TFe)为94.82%的铁精矿,铁回收率为99.53%,同时还得到w(Nb2O5)为0.3519%的铌粗精矿,铌回收率为99.62%。     

Purification of specularite by centrifugation instead of flotation to produce iron oxide red pigment

This study used specularite, a high-gradient magnetic separation concentrate, as a raw material in reverse flotation.An iron concentrate with a grade of 65.1 wt% and a recovery rate of 75.31% were obtained.A centrifugal concentrator served as the deep purification equipment for the preparation of iron oxide red pigments, and its optimal rotating drum speed, feed concentration, and other conditions were determined.Under optimal conditions, a high-purity iron oxide concentrate with a grade of 69.38 wt% and a recovery rate of 80.89% were obtained and used as a raw material for preparing iron oxide red pigment.Calcining with sulfuric acid produced iron red pigments with different hues.Simultaneously, middlings with a grade of 60.20 wt% and a recovery rate of 17.51% were obtained and could be used in blast furnace ironmaking.High-value utilization of specularite beneficiation products was thus achieved.     

红土镍矿深度还原-磁选富集镍铁实验研究

采用深度还原-磁选工艺,以煤粉为还原剂,添加氧化钙作助溶剂,在微熔化,不完全造渣的条件下,将矿石中镍和铁的氧化物还原成金属镍铁,然后经磁选方法使金属镍铁在磁性产品中得到富集.结果表明,深度还原最佳工艺条件为:还原温度1 300℃,还原时间60 min,配煤过剩倍数2.在此工艺条件下得到镍、铁质量分数分别为5.01%,22.46%的镍铁产品,镍、铁回收率分别为96.05%,79.69%.对深度还原过程研究表明,还原物料中镍和铁以金属合金颗粒形式存在,高温有利于镍铁金属相凝聚,适当延长还原反应时间有利于镍铁颗粒的还原和聚集长大,进而有利于磁选富集.     

用离心法代替浮选提纯镜铁矿制备氧化铁红色素

This study used specularite, a high-gradient magnetic separation concentrate, as a raw material in reverse flotation. An iron concentrate with a grade of 65.1wt% and a recovery rate of 75.31% were obtained. A centrifugal concentrator served as the deep purification equipment for the preparation of iron oxide red pigments, and its optimal rotating drum speed, feed concentration, and other conditions were determined. Under optimal conditions, a high-purity iron oxide concentrate with a grade of 69.38wt% and a recovery rate of 80.89% were obtained and used as a raw material for preparing iron oxide red pigment. Calcining with sulfuric acid produced iron red pigments with different hues. Simultaneously, middlings with a grade of 60.20wt% and a recovery rate of 17.51% were obtained and could be used in blast furnace ironmaking. High-value utilization of specularite beneficiation products was thus achieved.     

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 Al₂O₃), 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 electric arc furnace to replace scrap steel.     

Recovery of iron from copper tailings via low-temperature direct reduction and magnetic separation: process optimization and mineralogical study

Currently, the majority of copper tailings are not effectively developed. Worldwide, large amounts of copper tailings generated from copper production are continuously dumped, posing a potential environmental threat. Herein, the recovery of iron from copper tailings via low-temperature direct reduction and magnetic separation was conducted; process optimization was carried out, and the corresponding mineralogy was investigated. The reduction time, reduction temperature, reducing agent (coal), calcium chloride additive, grinding time, and magnetic field intensity were examined for process optimization. Mineralogical analyses of the sample, reduced pellets, and magnetic concentrate under various conditions were performed by X-ray diffraction, optical microscopy, and scanning electron microscopy-energy-dispersive X-ray spectrometry to elucidate the iron reduction and growth mechanisms. The results indicated that the optimum parameters of iron recovery include a reduction temperature of 1150℃, a reduction time of 120 min, a coal dosage of 25%, a calcium chloride dosage of 2.5%, a magnetic field intensity of 100 mT, and a grinding time of 1 min. Under these conditions, the iron grade in the magnetic concentrate was greater than 90%, with an iron recovery ratio greater than 95%.     

还原剂对高磷鲕状赤铁矿还原焙烧铁磷分离的影响

研究还原剂种类及用量对高磷鲕状赤铁矿还原焙烧铁磷分离的影响.添加脱磷剂Na2CO3,在提铁降磷的同时能降低还原铁的硫含量;还原剂用量的增加都能促进铁还原,但使用灰分和固定碳含量较高或挥发分含量较低的还原剂时,不利于降磷.焙烧产物的X射线衍射分析表明:添加脱磷剂Na2CO3时,随着还原剂用量的增加,焙烧产物中金属铁含量增加,浮氏体和石英含量降低;使用灰分含量较高的还原剂时,随其用量的增加,灰分会消耗Na2 CO3,从而减弱其对于铁还原的促进作用;还原剂用量相同时,石煤、烟煤、焦炭和褐煤所得焙烧产物中金属铁含量逐渐增加,浮氏体含量逐渐降低.总体来看,褐煤作为还原剂时铁磷分离效果最好,其次为烟煤,焦炭和石煤.