红土镍矿真空碳热还原脱镁的热力学研究

高镁低品位红土镍矿开发利用的关键在于提取镍的同时,重视金属镁等金属的综合回收利用。本文从热力学角度针对真空碳热还原红土镍矿脱除金属镁的过程进行了分析及能耗估算,探索处理红土镍矿的新工艺的热力学可行性。热力学计算表明,真空碳热还原红土镍矿脱除金属镁是可行的。当体系压强100Pa左右的条件下,MgO被还原的温度大于1478K,Ni、Fe以及部分硅将优先被还原为金属;过程能耗理论估算表明:当红土镍矿∶煤炭(质量比)=100∶42,1600K~1800K条件下还原处理1吨红土镍矿,需消耗1.12吨~1.17吨标准煤。     

氢气作用下硅镁型红土镍矿的低温还原特性

以含镍0.82%、含铁9.67%的某硅镁型红土镍矿为原料开展氢气低温还原实验研究,考察还原温度、还原时间、氢气浓度及矿物粒度对镍、铁金属化率的影响。结果表明:在还原温度为600℃、还原时间90 min及氢气浓度为60%(体积分数)的条件下,红土镍矿中镍、铁金属化率分别达到95%和42%。当矿物粒度小于380μm时,矿物粒径对镍、铁金属化率的影响并不明显。随着还原温度的升高,镍铁合金([Fe,Ni])的衍射峰呈现先增强后减弱的趋势,在600℃时达到最大。且随着温度的进一步升高,无定型含镁硅酸盐重结晶生成镁橄榄石相,阻碍镍、铁的还原。通过氢气低温还原,矿物中的氧化镍几乎完全还原,部分铁被还原为金属铁与镍形成了镍铁合金,大部分的铁被还原为铁的低价氧化物。     

红土镍矿钠盐还原焙烧-磁选的机理

配加钠盐焙烧可改善红土镍矿的还原-磁选效果,显著提高磁性产品的镍、铁品位及回收率。通过热力学计算,并结合X射线衍射、光学显微镜以及环境扫描电镜分析,对硫酸钠和碳酸钠作用下红土镍矿的还原行为进行研究。结果表明:钠盐在红土镍矿还原焙烧过程中,可以破坏硅酸盐矿物的结构,有利于镍的还原富集。碳酸钠强化镍还原的能力强于硫酸钠的,硫酸钠则因还原过程中形成的硫具有降低镍铁金属颗粒表面张力的作用,因而其促进镍铁颗粒聚集长大的能力明显高于碳酸钠的,且硫酸钠作用下FeS的形成也有利于提高镍的品位。所以,硫酸钠和碳酸钠的共同作用下可获得高镍品位的磁性产品及较高的镍回收率。     

碳热还原法从红土镍矿中提取金属镁的研究

采用了HSC chemistry 5.0热力学分析软件、XRD、SEM及EDS等方法与手段,对碳热还原法从红土镍矿中提取金属镁过程进行了热力学分析及实验研究。研究结果显示,碳热还原提取金属镁过程主要由Mg2SiO4、Fe2O3、MgSiO3、MgFe2O4及少量NiO等参与反应。热力学研究表明,常压下MgFe2O4、Mg2SiO4与MgSiO3碳热还原生成金属镁蒸汽的初始温度在1373～2073K,Fe2O3、NiO碳热还原生成金属铁、镍的初始温度分别为923K、723K;在真空压力为10Pa时,MgFe2O4、Mg2SiO4与MgSiO3碳热还原生成金属镁蒸汽的初始温度均在923～1323K,Fe2O3、NiO碳热还原生成金属铁、镍的初始温度分别为673K、523K。试验结果表明,碳热还原法从红土镍矿提取金属镁过程是可行的,冷凝物含金属镁的平均含量达98.5%以上。     

某印尼低品位红土镍矿的微观结构及晶体化学(英文)

为深入研究红土镍矿的镍富集原理,利用电子显微镜、扫描电镜、X射线衍射分析以及电子探针微区分析对含镍0.97%的某印尼低品位红土镍矿的工艺矿物学进行研究,以了解镍钴有价金属的分布及赋存状态。实验表明:该矿样主要矿物为针铁矿(含量约为80%),镍含量约为0.87%;含镍、铁、镁的结晶水硅酸盐矿物((Mg,Fe,Ni)2SiO4)的含量约为15%,如利蛇纹石((Fe,Ni,Al)O(OH))和橄榄石((Mg,Fe,Ni)3Si2O5(OH))等,镍含量约在1.19%左右;其它含量较低的物相为赤铁矿、磁赤铁矿、铬铁矿和石英等,这些矿物的镍含量极低。钴土矿是含钴矿物,分析发现该矿物往往有较高的镍和钴含量。微观检测发现:红土镍矿微观结构复杂,不同矿物之间共生普遍,主要矿物的微观结构松散,因而传统选矿方式很难实现镍的富集。     

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。     

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.     

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。     

CO还原Fe_2O_3-NiO复合物形成Fe-Ni合金的非等温还原动力学(英文)

研究采用CO还原不同比例Fe2O3-NiO复合物的非等温还原动力学及机理。结果表明:随着NiO含量的增加,样品的还原程度不断提高,NiO的存在提高氧化铁还原率。在还原开始阶段,NiO优先被还原,Ni作为催化剂可以提高氧化铁的还原率。NiO含量的增加促进镍铁相(FeNi3)的增加,但导致铁纹石相(Fe,Ni)和镍纹石相(Fe,Ni)的减少。金属镍、金属铁及镍铁合金的形成导致微观颗粒具有均匀性。在还原初始阶段,气体产物中CO浓度大于CO2浓度,然后逐渐减小,当温度在400~500°C内,Fe2O3-NiO复合物的还原速率达到最大值,成核长大模型可以揭示还原机理。在温度低于1000°C的条件下,成核长大过程是还原反应速率的限制环节。     

CO还原Fe2O3-NiO复合物形成Fe-Ni合金的非等温还原动力学（英文）

研究采用CO还原不同比例Fe2O3-NiO复合物的非等温还原动力学及机理。结果表明:随着NiO含量的增加,样品的还原程度不断提高,NiO的存在提高氧化铁还原率。在还原开始阶段,NiO优先被还原,Ni作为催化剂可以提高氧化铁的还原率。NiO含量的增加促进镍铁相(FeNi3)的增加,但导致铁纹石相(Fe,Ni)和镍纹石相(Fe,Ni)的减少。金属镍、金属铁及镍铁合金的形成导致微观颗粒具有均匀性。在还原初始阶段,气体产物中CO浓度大于CO2浓度,然后逐渐减小,当温度在400~500°C内,Fe2O3-NiO复合物的还原速率达到最大值,成核长大模型可以揭示还原机理。在温度低于1000°C的条件下,成核长大过程是还原反应速率的限制环节。     

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.     

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.     

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%.     

Transformation and leaching kinetics of silicon from low-grade nickel laterite ore by pre-roasting and alkaline leaching process

The mineralogical phase transformation of a low-grade nickel laterite ore during pre-roasting process and the extraction of silicon during alkaline leaching process were investigated. The results indicate that the reaction activity of nickel ores is effectively improved by pre-roasting at 650 °C for 2 h, because of the transformation of lizardite into magnesium olivine and protoenstatite. When finely ground ore samples (44–61 μm) pre-roasted firstly react with sodium hydroxide solution (60 g/L) with a solid/liquid ratio of 1:5 at 140 °C for 120 min, the extraction of silicon can reach 89.89%, and the other valuable elements of magnesium, iron and nickel are accumulated in the solid residues. The leaching kinetics of nickel laterite ore can be described successfully by the diffusion through the product layer control model. The activation energy is calculated to be 11.63 kJ/mol and the kinetics equation can be expressed as 1–3(1–x)^2/3+2(1–x)=13.53×10^?2exp[–11.63/(RT)]t.     

选择性还原-磁选回收镍渣中的有价金属

采用选择性还原-磁选工艺富集某镍渣中的镍、铜,通过控制还原过程参数实现选择性还原。结果表明：添加熔剂并适当提高渣料的碱度（CaO与SiO2质量比）有助于镍、铜的富集;对碱度0.15、还原温度1200℃、还原时间20 min、内配煤量5%（质量分数）的优化条件下得到的还原样品,通过磨矿-磁选获得镍、铜、铁品位分别为3.25%、1.20%、75.26%的精矿,镍、铜、铁的回收率分别为82.20%、80.00%、42.17%,实现了镍、铜相对于铁的选择性富集;选择性还原-磁选没有显著降低S、P的含量,两者在工艺过程中的行为需要进一步研究。     

添加剂氧化钙对镍渣强化还原的影响

针对铁以铁橄榄石形式存在而难以直接还原磁选回收的问题,采用在镍渣中配加CaO提升镍渣碳热还原速率及铁金属化率的技术思路。将添加剂与镍渣按比例混合并配加适量还原剂后,在高温炉中进行还原,结合热力学计算及产物特征分析结果,研究添加剂强化还原的机理及还原效果。结果表明:CaO的加入有效促进了硅酸铁的解离,强化了还原过程,还原速率加快,金属相在渣相中的形态结构发生改变,金属铁的聚集长大明显,有利于后续分离回收。随着CaO添加量由0增加到10%(质量分数),还原产物的金属化率和Fe颗粒平均粒径呈先增大后减小的趋势,FeO含量和残碳量变化规律相反。在CaO添加量为8%时,还原产物中铁的金属化率和粒径达到峰值,分别为87.25%和41.3μm。     

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.