微波辅助硫酸浸出红土镍矿的研究

以红土镍矿为原料,研究了微波辅助硫酸浸出镍钴的工艺条件。考察了硫酸浓度、微波功率、微波温度、辐射时间、液固体积质量比对镍钴浸出率的影响。结果表明,在硫酸浓度3.0mol/L、微波功率700 W、微波温度90℃、辐射时间2.5 h、液固体积质量比4:1的最佳工艺条件下,镍浸出率达91%,钴浸出率65%以上。     

Pressure acid leaching of arid-region nickel laterite ore: Part II. Effect of ore type

A comparison of the leach chemistry and residue mineralogy has been carried out on the pressure acid leaching of nontronite, limonite and saprolite ores, using hypersaline water. Results are also compared with a typical arid-region laterite feed from Bulong, which consists of a blend of these ore types. Particular emphasis is placed on the influence of ore type on liquor analysis of iron, aluminium and magnesium, residue mineralogy and nickel extraction. Microprobe evidence is presented that incomplete nickel extraction results from the presence of unreacted minor phases present in the original ore, or from the presence of nickel in the amorphous silica, in apparent association with magnesium.     

红土镍矿加压浸出工艺残积型红土镍矿中和试验研究

残积型红土镍矿是一种重要的红土镍矿,镁元素含量较高(10％～27％),在红土镍矿加压浸出项目中,通常用来中和加压浸出的矿浆,矿浆中硫酸浓度通常为30 ～ 50 g/L.但残积矿用量对镍、钴浸出率有较大影响,为了更好地利用残积型红土镍矿,本文进行了常规浸出试验和还原浸出试验.试验结果表明:随着残积矿用量的增加(液固比降低...     

Pressure acid leaching of arid-region nickel laterite ore: Part I: effect of water quality

The effect of water salinity on the reactions occurring during pressure acid leaching of an arid-region laterite ore, using hypersaline water, seawater, sub-potable water and tap water, is examined. Particular emphasis is placed on the mineralogy of the residue and its implications with regard to residue volume/mass, overall acid consumption and nickel extraction. Analysis of a pressure acid leach residue by electron microprobe indicates that the residual nickel is present in phases that contain silicon and varying concentrations of aluminium, but are deficient in sulphur. Incomplete extraction of nickel from the ore may not be attributed to any one mineral phase.     

Extraction of vanadium from black shale using pressure acid leaching

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO₄) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe³⁺ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Kinetics of vanadium dissolution from black shale in pressure acid leaching

The leaching kinetics of vanadium from black shale in the sulphuric acid-oxygen system is presented. The effects of agitation speed, leaching temperature in the range of 110-150 °C, sulphuric acid concentration, oxygen partial pressure and particle size on the rate of vanadium leaching were determined. The results indicate that the rate is nearly independent of agitation above 200 rpm and increases with increasing temperature, sulphuric acid concentration and oxygen partial pressure. As leaching occurs, there is a progressive dissolution of a vanadium-bearing alumino-silicate phase, while the inert quartz phase assembles onto the mineral surface and remains as an “ash” layer. The leaching kinetics was analyzed by using a new variant of the shrinking core model (SCM) in which both the interfacial transfer and diffusion across the product layer affect the leaching rate. The determined activation energy was found to be 40.14 kJ/mol and the reaction orders with respect to sulphuric acid concentration and oxygen partial pressure were 0.61 and 1.67, respectively. A semi-empirical rate equation was derived to describe the process.     

Comparative leaching of minerals by sulphuric acid in a Chinese ferruginous nickel laterite ore

The Yuanjiang nickel laterite ore containing mainly maghemite, goethite and lizardite was leached by sulphuric acid at atmospheric pressure and the residues were characterized using X-ray diffraction and scanning electron microscopy/X-ray energy dispersive spectroscopy. The relationship was discussed between the extraction of nickel, cobalt, iron, magnesium, aluminum, and the dissolution behaviour of the laterite minerals; as well as the extent of congruency of nickel, cobalt and iron extraction. The results show that the solubility of the laterite minerals in sulphuric acid decreases in the following order: lizardite > goethite > maghemite > magnetite ≈ hematite > chromite ≈ ringwoodite. Lizardite dissolved rapidly in 0.6 mol/L sulphuric acid at 60 °C whilst goethite dissolved completely in 2.5 mol/L sulphuric acid at 80 °C. The dissolution of the primary mineral maghemite was slow, but increased with increasing acid concentration and leaching temperature. Magnetite dissolved more slowly than maghemite; and hematite was only dissolved in > 6.2 mol/L sulphuric acid at 105 °C. Chromite and ringwoodite were not dissolved. The leaching behaviour of the laterite minerals may be explained by the bond strength differences of Me–O and the substitution of metal cations in the mineral structure.     

Recovery of valuable metals from jarosite by sulphuric acid roasting using microwave and water leaching

In this paper, jarosite residue (JR) blended with concentrated H₂SO₄ was subjected to a process comprising microwave roasting and water leaching. The effects of H₂SO₄/JR weight ratio, microwave roasting temperature and time, water leaching conditions on the recovery of Fe, Zn, In, Cu, Cd, Ag and Pb were investigated utilising a series of experiments.

Based on energy conservation and environmental protection, optimum conditions for metals recovery from JR were determined as: H₂SO₄/JR weight ratio?=?0.36, microwave roasting temperature, 250°C; roasting time, 30?min; leaching temperature, 50°C; leaching time, 1?h; and liquid–solid ratio, 4:1 (mL/g), thus, the extraction of Fe, Zn, In, Cu, Ag and Cd were 89.4, 80.7, 85.1, 90.7, 61.3 and 48.8% respectively, while the Pb was concentrated in the final residue. Scanning electron microscope-energy dispersive spectrometer (SEM-EDS) patterns were used to characterise and analyse the transformation of valuable metals in the residue after roasting and leaching.     

低品位红土镍矿焙烧-碱浸过程中硅的转化

针对目前红土镍矿碱法处理过程中存在的问题提出工艺改进,研究低品位红土镍矿焙烧活化-碱浸过程中含硅矿物的转化。考察了焙烧温度对红土镍矿活性的影响,探索了红土镍矿经焙烧后碱浸过程中温度、时间、搅拌强度、液固比以及碱初始质量浓度对硅转化的影响。结果表明,红土镍矿经650 °C焙烧2 h后,活性得到明显提高,红土镍矿经焙烧后采用初始质量浓度为60 g/L的碱溶液,在搅拌速度为400 r/min、浸出温度为140 °C、液固比为5∶1的条件下浸出120 min,硅的转化率可达89.42%。     

红土镍矿高压酸浸工程设计综述

红土镍矿高压酸浸工艺因为同时具有高温、高酸、高压力、高磨蚀等严酷工况条件,使得该工艺在工程设计以及工程应用上均具有极大的挑战.近十几年来,随着红土镍矿高压酸浸工程化研究的深入,新型设备的设计及制造,几个大型红土镍矿高压酸浸项目逐渐建成并投入运行.本文针对红土镍矿高压酸浸工艺的系统选择、关键设备选型、配置方案以及控制模式等方面进行了工程设计方面的综合论述.     

西澳大利亚三个镍红土矿项目的工程化比较

西澳大利亚的三个镍项目——布隆、考斯和穆林穆林,都是用加压酸浸法处理红土矿生产金属镍,每个主要工艺步骤都采用了不同的工程路径,本文比较了这些路径。     

红土镍矿绿色化综合利用

根据红土镍矿的化学组成和矿物结构,设计了绿色化综合利用的工艺流程.将红土镍矿和硫酸氢铵混合焙烧,红土镍矿中的镁、铁、镍、铝等生成可溶于水的盐,硅以二氧化硅的形式存在,且不溶于水.将焙烧产物用水溶出,过滤,使镁、铁、镍、铝的盐与二氧化硅分离.根据镁、铁、镍、铝等生成沉淀的pH值不同,向溶液中加入NH_3等碱性物质,调pH值,把镁、铁、镍、铝等分离提取.硫酸根转化为硫酸氢铵,返回焙烧,实现循环利用.对工艺流程的各工序进行了实验研究,给出了优化的工艺条件.     

锌精矿常压浸出与加压浸出工艺比较

比较了常压浸出与加压浸出两种工艺的机理、流程、技术经济指标、投资以及存在的问题。试验和生产数据表明,加压浸出在技术上和工艺上都更具有吸引力。     

锌精矿常压浸出与加压浸出工艺比较

比较了常压浸出与加压浸出两种工艺的机理、流程、技术经济指标、投资以及存在的问题。试验和生产数据表明,加压浸出在技术上和工艺上都更具有吸引力。     

红土镍矿湿法冶金工艺现状及前景分析

随着硫化镍矿资源的不断消耗以及镍需求量的持续增长,红土镍矿将是未来镍的主要来源。红土镍矿具有储量丰富、易开采、便于运输等特点,成为研究开发的热点。对还原焙烧 氨浸工艺、加压酸浸工艺、常压酸浸工艺等湿法冶金处理红土镍矿的工艺特点及现状进行了阐述,并分析了各工艺的优势与不足,介绍了红土镍矿湿法冶金工艺的研究进展。最后探讨了未来红土镍矿湿法冶金工艺的发展前景,指出加压酸浸工艺将在今后红土镍矿湿法冶金中扮演重要角色。     

Control of autoclave scaling during acid pressure leaching of nickeliferous laterite ore

An operating problem encountered at the Moa Bay operation in Cuba, where nickeliferous laterite ore is processed by sulfuric acid pressure leaching, is the formation of alunite and hematite deposits on the autoclave walls. The AMAX Extractive Research & Development, Inc., metallurgical laboratory (Golden, Colorado) has made substantial improvements in the Moa Bay process in the area of metal recovery, energy consumption, and feed versatility. One of the advantages of AMAX's process is its ability to treat substantial portions of nickel-and magnesium-rich serpentine while maintaining acid utilization efficiency. Scale formation is minimized by combining staged acid addition with vigorous agitation and 270 °C operation. This paper describes how advantage can be taken of MgSO₄· XH₂O precipitation both to inhibit alunite scaling and to disperse hematite scale within the MgSO₄ · XH₂O matrix. Cooling the autoclave from its 270 ·C operating temperature down to 180 ·C takes advantage of the reverse solubility of magnesium sulfate. The magnesium dissolves, liberating entrained hematite, thus providing a means for control of autoclave scale with minimum process disruption. P. B. QUENEAU, Formerly with AMAX Extractive Research and Development, Inc., Golden, CO, P. REY, Formerly with COFREMMI, Paris, France,     

蛇纹石硫酸法浸出MgO的热力学研究

本文通过对蛇纹石硫酸浸出过程中的热力学计算,分析了浸出MgO过程中各种矿物与酸的反应活性,及温度对各种矿物浸出的影响.结果表明,蛇纹石中主要矿物Mg6[Si_4O_(10)](OH)_8极易与硫酸反应;铁矿物中的FeO、FeO·SiO_2和FeCO_3易与硫酸反应,Fe~(2+)的析出将不利于Mg~(2+)的分离;磁黄铁矿FeS与硫酸有一定的反应;铝矿物Al_2O_3高温时不易与稀硫酸反应,低温时易与稀硫酸发生反应;而Fe_2O_3、黄铁矿FeS_2和黄铜矿CuFeS_2则不会与稀酸反应,它们的存在不影响Mg~(2+)的分离.     

湿法炼锌酸浸渣常规干燥实验研究

本文以湿法炼锌酸性浸出渣为研究对象,实验研究干燥温度、物料量、干燥时间对酸浸渣相对脱水率的影响,综合考虑干燥工艺的时效性、经济效益以及工艺效率,控制物料量为50g、干燥温度为90℃、干燥时间为80min时,酸浸渣相对脱水率可达到98.32%。实验研究对促进酸浸渣中有价金属的资源综合利用、改善环境等具有重要的指导意义。     

Upgrading of iron ores using microwave assisted magnetic separation followed by dephosphorization leaching

This paper describes the development of a flowsheet using a combination of sample preparation, magnetic separation (in a range 0.4–1?T), microwave treatment (in a range 0.54–0.9?kW), and leaching operations (HCl, in a range 0.25–1.25?M) for the beneficiation of iron ores (Total Fe [T_FE]: 55.48%). The work was aimed at recovering T_FE from the non-magnetic product by increasing its magnetic susceptibility through microwave treatment. It was found that goethite mineral in the non-magnetic product at a temperature of around 200^oC was converted to paramagnetic (hematite) or ferromagnetic minerals (maghemite and magnetite) by microwave treatment and overall beneficiation recovery was improved. The phosphorus (P) content in the final product was then removed by leaching in 0.25?M HCl. The iron loss from the final concentrate during leaching was found to be 0.78%, which was negligible. Overall, a final concentrate assaying 61.78% T_FE with a recovery of 94.97%, and containing 0.04% P with 94.73% removal was obtained providing satisfactory results for use in the industry. This study gives an alternative way for possible future studies to produce an iron concentrate with a high recovery from problematic iron ores, which can be categorised due to its P content.