红土镍矿还原焙烧-磁选制取镍铁合金原料的新工艺

采用钠盐添加剂强化红土镍矿的还原焙烧-磁选,确定了添加剂存在下适宜的焙烧和磁选技术参数,开发出红土镍矿还原焙烧-磁选制取镍铁合金原料的新工艺.结果表明:钠盐添加剂具有显著降低焙烧温度、大幅提高产品镍、铁品位和回收率的作用;对一种含镍1.58％、铁22.06％的红土镍矿配加添加剂后,在还原温度1 100℃、还原时间60 min、磁场强度0.1T的条件下,磁性产品的镍、铁品位可分别从无添加剂时的2.0％、57.2％提高到7.5％、80.5％,镍、铁回收率也相应从19.1％、33.6％增加到82.7％、62.8％.XRD结果表明:红土镍矿在无添加剂作用下经还原焙烧-磁选所得的磁性产物中仍有部分镁橄榄石及顽火辉石存在;而有添加剂存在时,还原生成的镍铁合金通过磁选可与非磁性脉石成分得到更为有效的分离,产品可作为不锈钢的生产原料.     

低品位红土镍矿直接还原焙烧-磁选试验研究

近年来,随着高品位硫化镍矿的枯竭及国内不锈钢产业的快速发展,低品位红土镍矿已经成为生产镍铁产品的主要原料.为了解决红土镍矿的合理利用问题,以红土镍矿为原料,煤粉为还原剂,采用直接还原法将矿石中的镍还原成了金属镍,经磁选分离使镍得到富集.考察了还原温度,还原时间,原料粒度,配煤量对镍回收率的影响.通过试验得出的最佳工艺条件为:原料粒度-0.074 mm、配煤量4％、还原剂粒度0.177～0.25 mm、还原温度1200℃、还原时间90min;得到的焙烧产物细磨至-0.048 mm,并在0.4T的磁场强度下扫选.在0.1T精选后,镍的品位为6.4％,镍的回收率为90％.     

Na_2SO_4在高铁低镍型红土镍矿选择性还原焙烧中的作用机理

在较大Na_2SO_4用量范围内,研究Na_2SO_4对高铁低镍型红土镍矿选择性还原焙烧的作用机理。结果表明:Na_2SO_4的作用随其用量变化有本质的区别;随其用量增加,磁选镍铁产品中镍的品位和回收率以及镍铁回收率差均先提高后降低,而铁产率和回收率则是先降低后升高;当Na_2SO_4用量为5%时,选择性还原效果最佳;Na_2SO_4会损耗煤中的固定碳,减弱还原气氛,使试样中的铁矿物还原为不具磁性的FeO;而当Na_2SO_4过量时,会导致部分铁矿物还原为含镁磁铁矿,造成铁回收率呈现先降低后提高的变化规律,同时过量的Na_2SO_4还会生成多余的Na2S,与焙烧体系中的NiO和FeO发生反应生成NiS和FeS,二者混熔生成(Ni,Fe)S,导致镍的品位和回收率都降低。     

铁质和镁质红土镍矿直接还原-磁选工艺对比

主要进行铁质红土镍矿和镁质红土镍矿的直接还原-磁选工艺的对比研究。结果表明:在不使用添加剂的情况下,铁质试样中的镍比较容易还原和回收,镍回收率能达到90%,但镍品位较低。镁质试样中的镍较难回收,镍的品位和回收率都不理想,镍回收率只有45%左右。进一步试验发现,硫酸钠可作为铁质试样的理想添加剂,但对镁质试样作用效果不佳。氟化钙对镁质试样作用效果显著,但对铁质试样作用效果不明显。分析发现,两种试样镍的赋存状态、铁元素含量以及还原所得金属颗粒大小的差异是造成上述差别的主要原因。     

钠盐对高铝褐铁矿还原焙烧铝铁分离的影响

研究钠盐对高铝褐铁矿还原焙烧过程中铝铁分离的影响。结果表明:高铝褐铁矿经还原后,铁的金属化率为87.13%,焙烧产物经磨矿磁选后,金属铁粉中铁品位和Al2O3含量分别为68.07%(质量分数)和7.94%,铁的回收率仅为19.77%;添加硫酸钠14%(质量分数)、辅助添加剂BS2.5%(质量分数)后还原高铝褐铁矿,铁的金属化率可达95.69%,焙烧产物经磨矿磁选后,金属铁粉中铁品位升高到91.3%,Al2O3含量降低为1.27%,铁的回收率达到93.64%。XRD、EDX及微观结构研究表明:未添加钠盐时,高铝褐铁矿中铁氧化物易被还原为无磁性的γ-Fe,且铁与铝、硅结合紧密,磁选分离难度大;添加的钠盐能与Al2O3和SiO2反应生成铝硅酸钠,破坏矿石结构,有利于改善高铝褐铁矿的还原效果,但在碳酸钠作用下铁晶粒较小且易与脉石矿物结合,而在硫酸钠作用下金属铁颗粒长大,与脉石矿物解离性能好,有利于铝铁分离。     

蛇纹石矿物的高温相变对红土镍矿直接还原的影响

红土镍矿所含的蛇纹石矿物在焙烧过程中会出现脱羟基和重结晶等相变。选取两种不同试样进行直接还原焙烧-磁选实验,研究蛇纹石的高温相变对直接还原焙烧红土镍矿的影响。对两种红土镍矿进行热重分析、XRD衍射分析和扫描电镜分析,研究两种红土镍矿中的蛇纹石矿物在焙烧过程中相变过程的异同及其对直接还原的影响。结果表明:两种试样所含主要矿物为蛇纹石和针铁矿,其热重分析曲线相似;在焙烧过程中,试样2在较低温度下出现橄榄石相。在最终的焙烧矿物相中,与试样1相比,试样2中出现石英相;与试样2相比,试样1的蛇纹石颗粒内部在焙烧后形成较多裂隙。因此,试样2的镍回收率较低。     

低品位红土镍矿深度还原机理

采用扫描电子显微镜和EDS能谱研究低品位红土镍矿深度还原过程中金属颗粒的生长行为,并在此基础上分析其还原机理。结果表明,金属铁和镍逐渐聚集生长为Fe—Ni颗粒,并且颗粒粒度随着还原温度的升高和还原时间的延长而明显增大。还原后,红土镍矿明显变为Fe—Ni金属颗粒和渣相基体两部分。铁镁橄榄石的还原与其晶体化学特性密切相关。铁和镍的氧化物被还原剂还原为金属铁和镍,同时,橄榄石的晶格结构被破坏。红土镍矿深度还原包含金属氧化物还原和金属相生长两个过程。     

硫酸钠和碳酸钠对高铝铁矿中铝铁分离作用的差异

研究硫酸钠和碳酸钠对高铝铁矿石还原焙烧铝铁分离作用机理的差异。结果表明:硫酸钠或碳酸钠均可显著改善高铝铁矿石的还原效果,添加硫酸钠可获得较好的铝铁分离效果,添加碳酸钠可获得较高的铁回收率。碳酸钠作用下,铁晶粒较小且与脉石矿物结合;而硫酸钠作用下金属铁颗粒长大,与脉石矿物界限分明,解离性能好,后者有利于铝铁分离。硫酸钠存在的还原体系形成新生相S、Na2S和FeS,在体系内以液相存在,为Fe2+离子的扩散提供液相环境,降低了Fe2+离子迁移的势垒,有利于Fe2+离子的扩散,从而为铁晶粒和铝硅酸钠的聚集提供有利途径;而碳酸钠存在的还原体系没有液相生成,Fe2+离子的迁移只能通过固相扩散进行,迁移阻力大,因此,铁晶粒与脉石矿物的界限不及添加硫酸钠时的分明。     

气体选择性还原红土镍矿的热力学（英文）

由于硫化镍矿生产镍铁在经济和环境上不断出现的问题,采用红土镍矿生产镍铁越来越受到重视。但是红土镍矿制备镍铁的火法工艺中,在提高铁镍产品中的镍含量方面的理论研究仍存在许多不足。出于这方面的考虑,假设Fe2O3、Fe O和Fe3O4的活度为1,计算了CO2/CO、H2O/H2和CO2/H2三种气氛下选择性还原红土镍矿时,不同铁活度下铁-铁氧化物的平衡条件。从已有的热力学数据出发,利用Miedema二元合金生成热模型,计算了Ni-Fe固态二元合金中铁的活度系数。并以活度系数为纽带,最终计算出这三种还原气氛下,镍铁合金产物中的铁含量与还原气体分压、还原温度的关系。并用CO2/H2还原红土镍矿,得到的实验数据与理论值进行了对比分析与讨论,热力学计算结果很好地解释了选择性还原红土镍矿时铁金属化无法避免的原因,并较好地预测了红土镍矿还原产物中铁含量随温度和气体组分的变化趋势。     

Na2CO3对红土镍矿的H2还原影响规律与作用机理

以红土镍矿为研究对象,重点考察添加Na₂CO₃对红土镍矿的H₂还原影响规律。对还原焙烧矿物采用X射线衍射(XRD)、扫描电子显微镜(SEM)和热重-质谱联用(TG-MS)等技术进行表征。结果表明：在还原温度为1000℃,还原时间为90 min,H₂浓度为45%(体积分数),Na₂CO₃的添加量为15%(质量分数)时,可得镍品位为3.02%、镍回收率96.75%的精矿。Na₂CO₃对红土镍矿的修饰作用机理的本质为,Na₂CO₃中的Na+通过与红土镍矿中的Mg-Si-O以及Ni-Mg-O体系发生反应,取代全部Ni2+以及部分Mg²⁺,从而破坏硅镍酸盐及硅镁酸盐的结构,进而使赋存于硅酸盐类中的镍元素被释放出来,有利于后续镍的富集提取。     

Thermal behaviors and growth of reduced ferronickel particles in carbon-laterite composites

The thermal behaviors of single laterite ore and graphite-laterite mixtures were investigated by thermogravimetry (TG), derivative thermo-gravimetry (DTG), and differential thermal analysis (DTA). Four mass loss steps maximized at about 78, 272, 583, and 826°C are observed for the laterite ore, representing the vaporization of free water, the dehydroxylation of goethite, the decomposition of serpentines, and the second dehydroxylation of serpentines, respectively. The reduction reactions of the graphite-laterite mixtures start at around 700°C and can be divided into three major temperature regions. Coal-laterite composites with an addition of 10 wt.% CaO were roasted at 1100-1350°C for 30 min, and the reduced samples were characterized by X-ray diffraction (XRD) and scanning electron microscopy (SEM). The results indi-cate that the reduction reactions proceed more completely at higher temperatures. The growth of the reduced ferronickel particles is greatly influenced by the roasting temperature. Obvious growth of the reduced ferronickel particles appears with the formation of worm-like crystals for the sample reduced at 1250°C, and spheric particles are observed for the sample reduced at 1300°C. When the reduction temperature in-creases to 1350°C, the reduced ferronickel particles agglomerate to ferronickel granules of 3-8 mm in diameter. The main elements in the granules include iron, nickel, chromium, carbon, and sulfur, with the content of nickel and that of iron of 9.08 wt.% and 85.21 wt.%, respec-tively.     

Carbothermic Reduction of Nickeliferous Laterite Ores for Nickel Pig Iron Production in China: A Review

Both the consumption and production of crude stainless steel in China rank first in the world. In 2011, the nickel production in China amounted to 446 kilotons, with the proportion of electrolytic nickel and nickel pig iron (NPI) registering 41.5% and 56.5%, respectively. NPI is a low-cost feedstock for stainless steel production when used as a substitute for electrolytic nickel. The existing commercial NPI production processes such as blast furnace smelting, rotary kiln-electric furnace smelting, and Krupp-Renn (Nipon Yakin Oheyama) processes are discussed. As low-temperature (below 1300°C) reduction of nickeliferous laterite ores followed by magnetic separation could provide an alternative avenue without smelting at high temperature (~1500°C) for producing ferronickel with low cost, the fundamentals and recent developments of the low-temperature reduction of nickeliferous laterite ores are reviewed.     

Efficient enrichment of nickel and iron in laterite nickel ore by deep reduction and magnetic separation

The process of deep reduction and magnetic separation was proposed to enrich nickel and iron from laterite nickel ores. Results show that nickel–iron concentrates with nickel grade of 6.96%, nickel recovery of 94.06%, iron grade of 34.74%, and iron recovery of 80.44% could be obtained after magnetic separation under the conditions of reduction temperature of 1275 °C, reduction time of 50 min, slag basicity of 1.0, carbon-containing coefficient of 2.5, and magnetic field strength of 72 kA/m. Reduction temperature and time affected the possibility of deep reduction and reaction progress. Slag basicity affected the composition of slag in burden and the spilling and enriching rate of nickel–iron from a matrix to form nickel–iron particles. Nickel–iron particles were generated, aggregated, and grew gradually in the reduction process. Nickel–iron particles can be effectively separated from gangue minerals by magnetic separation.     

Reduction roasting–magnetic separation of vanadium tailings in presence of sodium sulfate and its mechanisms

Reduction roasting with sodium sulfate followed by magnetic separation was investigated to utilize vanadium tailings with total iron grade of 54.90 wt% and TiO_2 content of 17.40 wt%. The results show that after reduction roasting–magnetic separation with sodium sulfate dosage of 2 wt% at roasting temperature of 1150 °C for roasting time of 120 min, metallic iron concentrate with total iron grade of 90.20 wt%, iron recovery rate of 97.56 % and TiO_2 content of 4.85 wt% is obtained and high-titanium slag with TiO_2 content of 57.31 wt% and TiO_2 recovery rate of 80.27 % is also obtained. The results show that sodium sulfate has a catalytic effect on the reduction of tailings in the novel process by thermodynamics, scanning electron microscopy(SEM) and X-ray diffraction(XRD) and reacts with silica and alumina in the tailings to form sodium silicate and sodium aluminosilicate. Migration of elements and chemical reactions destroy the crystal structures of minerals and promote the reduction of vanadium tailings, resulting in that iron grains grow to large size so that metallic iron concentrate with high total iron grade and low TiO_2 content is obtained.     

Phase transformation in reductive roasting of laterite ore with microwave heating

Selective reduction of laterite ores followed by acid leaching is a promising method to recover nickel and cobalt metal, leaving leaching residue as a suitable iron resource. The phase transformation in reduction process with microwave heating was investigated by XRD and the reduction degree of iron was analyzed by chemical method. The results show that the laterite samples mixed with active carbon couple well with microwave and the temperature can reach approximate 1000 ℃ in 6.5 min. The reduction degree of iron is controlled by both the reductive agent content and the microwave heating time, and the reduction follows Fe2O3→Fe3O4→FeO→Fe sequence. Sulphuric acid leaching test reveals that the recoveries of nickel and iron increase with the iron reduction degree. By properly controlling the reduction degree of iron at 60% around, the nickel recovery can reach about 90% and iron recovery is less than 30%.     

Co-reduction of Copper Smelting Slag and Nickel Laterite to Prepare Fe-Ni-Cu Alloy for Weathering Steel

In this study, a new technique was proposed for the economical and environmentally friendly recovery of valuable metals from copper smelting slag while simultaneously upgrading nickel laterite through a co-reduction followed by wet magnetic separation process. Copper slag with a high FeO content can decrease the liquidus temperature of the SiO₂-Al₂O₃-CaO-MgO system and facilitate formation of liquid phase in a co-reduction process with nickel laterite, which is beneficial for metallic particle growth. As a result, the recovery of Ni, Cu, and Fe was notably increased. A crude Fe-Ni-Cu alloy with 2.5% Ni, 1.1% Cu, and 87.9% Fe was produced, which can replace part of scrap steel, electrolytic copper, and nickel as the burden in the production of weathering steel by an electric arc furnace. The study further found that an appropriate proportion of copper slag and nickel laterite in the mixture is essential to enhance the reduction, acquire appropriate amounts of the liquid phase, and improve the growth of the metallic alloy grains. As a result, the liberation of alloy particles in the grinding process was effectively promoted and the metal recovery was increased significantly in the subsequent magnetic separation process.     

Efficient Utilization of Nickel Laterite to Produce Master Alloy

To lower the smelting temperature associated with the carbothermic reduction processing of laterite, the optimization of slag and alloy systems was investigated to enable the reduction of laterite ore in the molten state at 1723 K. The master Fe-Ni-Mo alloy was successfully produced at a lower temperature (1723 K). The liquidus of the slag decreased with the addition of oxide flux (Fe₂O₃ and CaO) and that of the ferronickel alloy decreased with the addition of Mo/MoO₃. More effective metal–slag separation was achieved at 1723 K, which reduces the smelting temperature by 100 K compared with the current electric furnace process. A small addition of Mo/MoO₃ not only decreased the melting point of ferronickel alloys but also served as a collector to aggregate the ferronickel sponges allowing them to grow larger. The FeO concentration in the slag and the nickel grade of the alloy decreased with increasing graphite reductant addition.     

Recovery of Co（Ⅱ） and Ni（ Ⅱ） from hydrochloric acid solution of alloy scrap

A hydrometallurgical process was developed for recovery of nickel and cobalt from the hydrochloric acid leaching solution of alloy scraps. The process consists of five maj or unit operations： 1） leaching with 6 mol/L hydrochloric acid under the L/S ratio of 10：1 at 95 ℃ for 3 h; 2） copper replacement by iron scraps under pH value of 2.0 at 80 ℃, and stirring for 1 h, 3） removal of iron and chromium by chemical precipitation： iron removal under pH value of 2.0 at 90 ℃ by dropwise addition of sodium chlorate and 18% sodium carbonate solution, then chromium removal under pH value of 4.0 at 70 ℃ by addition of nickel carbonate solution, stirred by air flow for 2 h; 4） selective separation of cobalt from nickel by extraction using 30% trialkyl amine＋50% kerosene （volume fraction） and tri-n-butylphosphate （TBP） as a phase modifier with the O/A ratio of 2：1, and stripping of cobalt with 0.01 mol/L HCl; 5） crystallization of nickel chloride and electrodeposition of cobalt. It is found that the nickel recovery of 95% and the cobalt recovery of approximately 60% with purity over 99.9% are obtained by this process.