印度尼西亚某红土镍矿还原焙烧-磁选试验

印度尼西亚某低品位红土镍矿含镍1.57%、含铁21.67%,其中镍主要以硅酸镍形式存在。为将该矿石的镍含量提高到6%以上以符合印度尼西亚政府对出口红土镍矿的规定,以硫酸钠和碳酸钠为助熔剂,进行了还原焙烧-弱磁选试验。试验结果表明,当煤用量为25%、硫酸钠+碳酸钠的配比和总用量分别为3∶1和20%、焙烧温度为1 200 ℃、焙烧时间为60 min、磨矿细度为-0.074 mm占85%、磁场强度为96 kA/m时,可获得产率为22.06%、镍品位为6.05%、镍回收率为85.03%、铁品位为65.74%、铁回收率为66.92%的镍铁精矿,其镍品位超过印度尼西亚出口红土镍矿的品位下限。     

Atmospheric leaching characteristics of nickel and iron in limonitic laterite with sulfuric acid in the presence of sodium sulfite

Atmospheric leaching of nickel from limonitic laterite ores is regarded as a promising approach for nickel production, despite its low nickel recovery and slower leaching rate than high pressure acid leaching. Sulfur dioxide can enhance the sulfuric acid leaching of laterite, but its behavior for enhancing atmospheric sulfuric acid leaching was uncertain due to SO₂ losses and emission. In this study, sodium sulfite was used as a substitute for SO₂ gas in the leaching and the sulfuric acid leaching characteristics of Ni and Fe from a limonitic laterite in the presence of sodium sulfite were investigated. A linear correlation exists between the extraction of Ni and Fe, indicating the difficulty in selective leaching of Ni over Fe. Most nickel is isomorphically substituted within the goethite and it is difficult to dissolve in a high oxidation–reduction potential solution environment, resulting in a low nickel recovery. SO₂(aq) generated from the reaction of sodium sulfite in sulfuric acid solution, lowers the potential for the reducing reaction of FeOOH to give Fe²⁺, accelerating the iron extraction and nickel liberation from goethite.     

基于BP神经网络技术的红土镍矿还原焙烧-磁选工艺条件的优化

还原焙烧—磁选工艺可有效提取红土镍矿中的镍和铁等有价金属,由于影响红土镍矿还原焙烧—磁选效果的因素较多,导致工业生产中的选矿指标不稳定。为进一步提高还原焙烧—磁选工艺处理红土镍矿的效果,本研究以青海某镍矿为原料,采用正交试验与BP神经网络相结合的方法,对还原焙烧—磁选工艺的还原剂用量、焙烧温度、料层厚度、焙烧时间及磁场强度等因素进行了优化。结果表明:通过BP神经网络模型优化后的试验条件为还原剂用量9.5%、焙烧温度1 070℃、料层厚度10.0 mm、焙烧时间65 min及磁场强度2.5 kA·m-1,在此条件下可获得产率为30.29%的镍粗精矿,比采用正交试验最优因素组合条件所得的镍粗精矿产率提高了2.83%。      

The reduction of nickel from low-grade nickel laterite ore using a solid-state deoxidisation method

The reduction of nickel from low-grade nickel laterite ore using a solid-state deoxidisation method was studied. The effects of temperature, time, reductant type and CaO content on the conversion percentage of the total nickel to metallic nickel (α_Ni) in the nickel laterite ore reduction process were investigated. The results showed that α_Ni was strongly influenced by the reaction temperature in both gas–solid and solid–solid reduction processes, and a higher temperature was more favourable for nickel reduction. Because the reduction mechanism of nickel laterite ore (NiO + CO → CO₂ + Ni) is indirect, a higher α_Ni (>80%) can be obtained by increasing the CO and anthracite content. In the gas–solid reduction process, a longer reaction time favoured nickel reduction, and the conversion percentage decreased when a gaseous reductant was used at 850 °C because of phase transformation. In the solid–solid reduction process, the conversion percentage of the total nickel to metallic nickel first increased and then decreased with increasing reduction time and CaO content. In both reduction processes, taenite was found by XRD in the reduced ore because of iron oxide reduction and metallic nickel formation. SEM revealed that the nickel laterite ore was transformed from large granular and sandwich structures to small granular and flocculent structures during the reduction process.     

An innovative process for extracting boron and simultaneous recovering metallic iron from ludwigite ore

Ludwigite ore has not yet been utilized on an industrial scale due to its complex mineralogy and fine mineral dissemination in China. Boron–iron separation and dissolution activity of boron-bearing minerals in alkaline liquor are the two key issues in the utilization of ludwigite ore, governing the boron recovery as well as operating cost. This paper proposes an innovative process for extraction of boron and iron from ludwigite ore based on coal-based direct reduction process with sodium carbonate (Na₂CO₃). The novel process involves reduction roasting, combined leaching and grinding of reduced ludwigite ore, followed by magnetic separation of leach residue, and experimental validation for each of the processing steps is demonstrated. Alkali-activation of boron and metallization of iron were synchronously achieved during carbothermic reduction of ludwigite ore in the presence of Na₂CO₃. Consequently, boron was readily extracted in the form of sodium metaborate (NaBO₂) with water at room temperature during ball mill grinding, and metallic iron powder was recovered from the leaching-filtering residue by magnetic separation. Boron extraction of 72.1% and iron recovery of 95.7% with corresponding iron grade of 95.7% in the magnetic concentrate were achieved when ludwigite ore was reduced with 20% sodium carbonate at 1050 °C for 60 min.     

Direct mechanochemical conversion of celestite to SrCO3

The celestite concentrate obtained from Barite Mining Company (Sivas-Turkey) was converted to strontium carbonate using sodium carbonate in a planetary ball mill operated both in dry and wet modes. After wet planetary milling of celestite concentrate and sodium carbonate mixture, solid strontium carbonate and soluble Na₂SO₄ were obtained. The conversion reaction were followed by dissolution and sulphate analyses after high-energy milling operations. The 10, 20 and 40 min of wet milling experiments were performed in a planetary mill and the results showed that over 90% of dissolution and conversion of initial celestite to strontium carbonate was achievable by high-energy milling.     

氯化钙作用下镁质贫镍红土矿选择性富集镍的研究

采用还原焙烧-磁选工艺, 对氯化钙作用下镁质贫镍红土矿选择性富集镍进行了研究, 考察了还原温度、还原时间、还原剂用量和氯化钙用量对富集镍的影响。结果表明, 在还原温度1 200 ℃、还原时间40 min、还原剂和氯化钙用量均为8%的条件下, 可获得镍品位8.67%、回收率92.01%的镍铁精矿; 相比于直接还原焙烧-磁选, 加入8%氯化钙后使镍的富集比由3增加到11, 显著提高了镍的富集效果。磁选产品物相分析显示, 镍主要以铁纹石形式存在于精矿中, 通过磁选实现了对镍铁精矿与脉石的有效分离。     

Mechanochemical processing of chrysocolla with sodium sulphide

Autogenous solid state processing of chrysocolla with sodium sulphide has been carried out at different CuSiO₃–Na₂S-ratios, particle sizes and temperatures. Water soluble sodium silicate (Na₂SiO₃) and covellite (CuS) were determined by means of chemical analysis and X-ray examination as reaction products. More than 84% of chrysocolla and 50% of associated malachite were converted to covellite at 80 °C, atmospheric pressure, 10 min contact time, stochiometric CuSiO₃–Na₂S-ratio and particle sizes of <100 μm. The mechanochemical treatment using a vibrating mill increases the conversion rate of chrysocolla and malachite up to 90% and 70%, respectively. After washing of the products, the sodium silicate solution was converted into sodium sulphate (Na₂SO₄) by adding sulphuric acid, with precipitation of silica as by-product. In addition, the carbothermic reduction of crystallised sodium sulphate led to sodium sulphide necessary for solid state processing.     

Mechanical activation of MoS2 + Na2O2 mixtures

A study of the mechanochemical activation of molybdenum ore concentrate (MoS₂) with sodium peroxide (Na₂O₂) shows that sodium molybdate dihydrate (Na₂MoO₄ · 2H₂O) is the final crystalline product. The mechanochemical formation of sodium molybdate dihydrate is evidenced by XRD, ²³Na MAS NMR and the increasing solubility of the molybdenum in water as the oxidative reaction proceeds.     

High pressure acid leaching of a refractory lateritic nickel ore

This paper describes the experimental findings of the extraction of nickel and cobalt by high pressure acid leaching (HPAL) of a refractory limonitic nickel laterite ore from the Gördes region of Manisa in Turkey. By optimizing the basic HPAL process parameters: leaching at 255 °C with 0.30 sulfuric acid to ore weight ratio with a particle size of 100% −850μ for 1 h of leaching, it was found that 87.3% of nickel and 88.8% of cobalt present in the ore could be extracted into the pregnant leach solution (PLS). However, these extraction results were found to be relatively low compared with other similar studies. In order to understand the possible reasons for this relatively lower extraction, further investigations have shown that together with a problem related to the kinetics of the dissolution reactions, a persistent acid resistant refractory mineral present in this sample also limited the leaching process. Attempts were made with different additives to solve this problem. The effects of chemical additives such as HCl, Na₂SO₄, FeSO₄, Cu⁺ and sulfur were tested and the effect of each addition on the degree of extraction of nickel and cobalt was determined.     

从红土镍矿中提取镍钴铁的新工艺研究

针对云南省元江红土镍矿的矿物组成特点,在比较国内外红土镍矿处理工艺的基础上,提出了还原—磨矿—选别—氧化浸出工艺处理该矿,并进行了全流程试验。首先进行了还原—磨矿—选别试验研究,主要考察了还原温度、还原时间、添加剂配比和还原剂配比对指标的影响;其次进行了综合试验。试验结果表明,还原—磨矿—选别可以抛弃红土镍矿中80%以上的脉石,同时实现镍钴铁富集,氧压浸出工艺可实现镍钴与铁的分离,并获得铁红产品。通过试验,获得的技术指标为:从原矿至氢氧化镍(钴)段,镍直收率大于75%、钴直收率大于70%和铁直收率大于80%;氢氧化镍产品镍的品位大于31%,氢氧化钴产品钴的品位大于0.70%,铁红产品铁含量大于62%,铁红达到铁精矿要求,可以作为铁精矿出售。该工艺实现了镍钴铁综合回收,资源利用率高,环境友好,为综合回收红土镍矿中镍钴铁提供一条新的工艺技术路线。     

红土镍矿冶金工艺现状及前景分析

本文简要介绍了镍资源的性质特点及现状,分析了红土镍矿回转窑-电炉还原熔炼、还原焙烧－磁选、还原硫化熔炼３种火法工艺以及常压酸浸、高压酸浸等湿法工艺的优势与短板,并指出生物浸出在红土镍矿处理中的应用,探讨了未来红土镍矿冶金工艺的改进发展,指出红土镍矿湿法冶金将在今后扮演重要地位。     

鄂西某高磷鲕状赤铁矿磁化焙烧及浸出除磷试验

针对鄂西某高磷鲕状赤铁矿(铁品位43.50%),在实验室条件下采用磁化焙烧—磁选工艺获取铁精矿,并对该铁精矿进行酸浸、生物浸出除磷试验。研究结果表明,在焙烧温度850℃,焙烧时间25min,还原剂用量为矿石质量的5%,磨矿时间4min,磁场强度120kA/m条件下,得到铁精矿铁品位为54.92%,铁回收率为86.78%,P含量为0.83%;酸浸试验中矿浆浓度2%,分别用0.1mol/L的H2SO4,HNO3,HCl,草酸(C2H2O4),柠檬酸(C6H8O7)除磷,其中H2SO4除磷提铁效果最佳,铁精矿品位为57.98%,回收率为96.47%,除磷率为95.30%;生物浸出试验中矿浆浓度2%,用嗜酸氧化亚铁硫杆菌(At.f菌)对铁精矿作用后,磷含量为0.23%,用黑曲霉菌滤液对铁精矿作用后,磷含量为0.20%。     

某硫铁烧渣浮选脱硫试验研究

重庆某硫酸厂硫铁烧渣硫含量高达5.79%,为了生产铁精矿,需要将硫含量降低至1.5%以下。试验在控制磨矿细度为-0.074mm占86%的条件下,以硫酸铜加硫酸为活化剂,丁基黄药为捕收剂,2#油作起泡剂,通过一粗二精的反浮选流程,使硫品位降低至1.32%。浮选精矿再经磁选,得到精矿S品位0.81%,Fe品位60.25%,回收率82.45%的良好指标。     

提高某超级铁精矿制备过程中的反浮选回收率

以安徽某粉末冶金公司超级铁精矿生产线所用普通铁精矿为原料,将其再磨、弱磁再选后进行生产超级铁精矿的反浮选试验,通过添加调整剂Na₂CO₃、水玻璃、淀粉来解决现场单纯使用十二胺引起的反浮选回收率偏低问题,并探讨通过提高矿浆浓度来提高反浮选的生产效率。试验结果表明：将2 000 g/t Na₂CO₃、200 g/t水玻璃、100 g/t淀粉与130 g/t十二胺配合使用,在24%矿浆浓度下,反浮选所获超级铁精矿的作业回收率可由单纯使用十二胺时的44.61%（对原料为43.44%）提高到89.58%（对原料为89.27%）;将矿浆浓度提高到28%,可使反浮选处理能力提高16.68%,而作业铁回收率仍达87.69%（对原料为87.38%）。     

Calcination of low-grade laterite for concentration of Ni by magnetic separation

With the continuous depletion of high-grade nickel ores such as millerite and niccolite, nickeliferous laterites have become the major source for the production of nickel metal. However, only 42% of the world’s production of nickel comes from laterites, since the concentration of Ni is relatively low (ca. 2 wt.%). In addition, other metals, such as magnesium, iron and silicon can be found in laterite, which make the concentration of nickel even more difficult.In this study, a low-grade nickeliferous laterite ore was first calcinated and then processed by using a wet magnetic separator in order to recover nickel. Since, the ore contains both Ni and Fe, the calcination of laterite is effective in altering the crystalline structure of Fe species and therefore its magnetic properties, which in turn enable the selective concentration of nickel by magnetic separation that is an easy and environmentally-friendly technique. The experimental results have indicated the importance of carefully controlling: (1) the calcination temperature; (2) the pulp density and (3) applied magnetic field strength. The main finding of this work was that magnetic separation is effective in recovering 48% of nickel from laterite, increasing the Ni grade in the recovered product from 1.5% to 2.9%, when prior to the separation the ore was calcinated at 500 °C for 1 h.     

Leaching of limonitic laterite ore by acidic thiosulfate solution

Cobalt is usually recovered as a by-product of copper and nickel processing, and only a small amount of cobalt is derived from laterites although a vast majority of cobalt resources in them. The exploitation of limonitic laterite containing high content of cobalt is becoming increasingly important. The mineralogy of a limonitic laterite ore was characterized by environmental scanning electron microscope (ESEM) and X-ray diffraction (XRD) in this paper. The results show that nickel occurs in goethite mainly, while cobalt is predominantly associated with manganiferous minerals. Thiosulfate is found to be able to selectively leach cobalt from limonitic laterite in the presence of sulfuric acid, and 91% Co, 22% Ni, 10% Fe are leached from an ore containing 0.13% Co, 1.03% Ni within the first 5 min at 90 °C under the conditions of 10 g/L sodium thiosulfate, 8% (w/w) sulfuric acid and 10:1 L/S ratio. The leaching kinetics of Mn and Co by acidic sodium thiosulfate solution can be characterized by the Avrami equation. In acidic solution, thiosulfate readily decomposes into sulfur and sulfur dioxide as intermediary reagents to reduce pyrolusite (MnO₂) and goethite (FeOOH); therefore, nickel and cobalt associated with goethite and pyrolusite respectively are extracted due to reduction dissolution. Furthermore, cobalt is selectively leached over iron and nickel because pyrolusite is preferentially reduced by acidic thiosulfate rather than goethite. The novel process may give an alternative method to selectively recover cobalt as the primary product from limonitic laterites at atmospheric pressure.     

新疆某磁铁矿选铁脱硫试验研究

介绍了新疆某磁铁矿的矿石性质和浮选脱硫的试验研究结果。试验结果表明,以丁基黄药为捕收剂,CuSO4+Na2S为组合活化剂,经先浮选后磁选工艺流程,可将铁精矿品位提高到65%以上,硫含量降至0.30%以下。试验实现了对铜、硫的综合回收。     

Microwave carbothermic reduction roasting of a low grade nickeliferous silicate laterite ore

The known resources of nickel sulphide ores are quickly diminishing and in order to satisfy future nickel demands, nickel laterite deposits are being investigated as an alternative. Currently, expensive leaching and smelting processes are used to process the nickel laterite ores. The objective of the present research was to produce a high grade nickel concentrate via microwave carbothermic reduction roasting followed by magnetic separation. A thermodynamic model was developed for the roasting process in order to determine the optimum experimental conditions. The experimental variables investigated were: microwave energy and argon shrouding for the reduction tests and the magnetic field strength for the concentration stage. The behaviours of nickel and cobalt were studied in the reduction and magnetic separation processes. By optimizing the reducing and magnetic separation conditions, a high grade concentrate containing 9.2% nickel with a nickel recovery of 88.8% was achieved.     

某新型抑制剂在铜锌分离中的试验研究

针对某复杂难选铜锌矿石,采用优先选铜原则流程进行铜锌分离,利用小分子有机抑制剂抑制锌矿物,分离效果良好。小型闭路试验可获得铜精矿含铜23.15%、铜回收率为77.61%、含锌5.61%,与常规锌组合抑制剂Zn SO4+Na2SO3相比,新药剂可使铜精矿中铜品位和回收率分别提高3.91和7.36个百分点、杂质锌的含量降低2.61个百分点,锌精矿的锌品位与回收率分别提高了0.66%、0.31%。