热酸浸出黄钠铁矾渣工艺

以黄钠铁矾渣为原料,研究硫酸浸出过程的工艺条件,分析浸出过程的热力学和动力学机理.实验结果表明,在液固比为5∶1,搅拌速率为350 r/min条件下,浸出黄钠铁矾渣的最佳工艺条件为:硫酸质量浓度为225 g/L,反应温度为95 ℃,反应时间为2.5 h,该条件下多组实验的渣中Fe、Zn浸出率均大于96%.黄钠铁矾渣硫酸浸出过程在动力学上属于收缩核模型,受化学反应控制.     

外加剂对水泥固化铁矾渣性能的影响

在硅酸盐水泥熟料中加入铁矾渣,制备成胶凝材料.分别以粉煤灰沸石、硫化钠和粉煤灰为外加剂,研究其对水泥固化体强度和浸出毒性的影响.在胶凝材料中铁矾渣加入量为60%时,加入沸石、硫化钠为稳定剂,均可提高重金属离子的稳定性,不同固化体的浸出毒性值均低于国家标准.在胶凝材料中加入粉煤灰,粉煤灰掺量增加,固化体强度下降,不同固化体的浸出毒性值也均低于国家标准.     

煤掺量对焙烧铁矾渣中硫物相和含量的影响

在含石膏铁矾渣中掺入不同质量分数的煤粉,用球磨机将两者混合均匀,在不同温度下焙烧,以探究煤掺量对焙烧含石膏铁矾渣中硫的物相和含量的影响。研究表明:在不同焙烧温度下,煤掺量不超过20%时,S含量随温度升高呈下降趋势;煤掺量为15%时的含石膏铁矾渣,在1 000℃焙烧2 h后Fe含量最高(为26.51%),S含量最低(为4.59%),S的物相主要为CaSO_4和CaS;将其烧渣研磨后过100目筛,水洗后测得该渣中S含量减少至2.13%。通过探究煤掺量对铁矾渣中硫物相和含量的影响,为炼锌厂资源化利用铁矾渣提供了新思路。     

黄钾铁矾渣浸出液还原和深度净化

以热酸直接浸出黄钾铁矾渣的浸出液为原料,采用单因素法对浸出液还原、硫化沉淀除隔和氟化沉淀除钙除镁的工艺进行研究.浸出液还原实验结果表明:在反应温度80℃,时间2 h,搅拌速度150~200 r/min,铁粉加入量为理论量1.15倍的条件下,浸出液中Fe3+全部被还原,铁锌比接近低功耗锰锌铁氧体的理论配方,杂质Cu2+的质量浓度降到1 mg/L,去除率在99%以上,Cd2+的去除率为17%.硫化沉淀除隔实验结果表明:浸出液还原反应完成后,直接在其中加入理论量1.4倍的(NH4)2S,反应时间为30 min,溶液中杂质Cd2+的质量浓度降到1 mg/L以下,去除率在98%以上.氟化沉淀除钙除镁综合实验研究中,溶液中Mg2+、Ca2+的除去率分别为96.73%和76.67%.     

用黄钠铁矾渣制备复合镍锌铁氧体

对黄钠铁矾渣制备复合镍锌铁氧体进行研究。研究结果表明：当黄钠铁矾渣与无烟煤按质量比为5：1均匀混合，在800℃还原焙烧0．5h，焙烧渣用0．5mol／L硫酸溶液按液固比7：1在70℃浸出40min时，渣中93％的铁和镍进入浸出液中。浸出液经过净化除杂后得到含镍的硫酸亚铁溶液，加入计量的硫酸镍和硫酸锌，采用共沉淀法，以NH4HCO3为沉淀剂，通过改变沉淀剂的用量控制pH值制备镍锌铁碳酸盐。将镍锌铁碳酸盐在800℃煅烧2h得到具有尖晶石结构的Ni0.5Zn0.5Fe2O4粉体，粒径约为59nm，粒度均匀，呈球形；煅烧4h时粉体粒径变大，粒度不均匀，呈棒状。     

铁矾法在Zn—MnO_2冶炼新工艺中的应用

本文结合成矾理论,论述了铁矾法在Zn—MnO_2冶炼新工艺中的应用。认为采用铁矾法除铁工艺简单,除铁效果好,且矾渣易于沉降、洗涤及过滤。     

杂铜阳极泥硫酸化焙烧-浸出工艺提取Cu和Ni的试验研究

杂铜阳极泥中含有大量的贵金属和稀有金属元素,是提取贵金属和稀有金属元素的重要原料,杂铜阳极泥处理的第一步即是提取Cu、Ni等贱金属,以富集贵金属和稀有金属元素.采用硫酸化焙烧-浸出工艺,从杂铜电解产生的阳极泥中提取Cu和Ni.考察焙烧温度、焙烧时间、浸出液固比、浸出硫酸浓度以及浸出时间等因素对Cu、Ni和Sn浸出率的影响.结果表明：当焙烧温度为400℃、焙烧时间3 h,浸出时液固比为4：1,100 g·L^-1硫酸、温度为80℃的条件下,Cu、Ni的浸出率>96.6%,可以有效地实现杂铜阳极泥中Cu和Ni的提取,而Sn的浸出率为13.0%,浸出渣可以作为提取Sn和贵金属的原料.     

锌浸出渣浮选试验研究

针对锌冶金工业产生的锌焙砂,基于富集浸出渣中的铁酸锌且不破坏铁酸锌的晶体结构的思想,采用浮选方法对锌浸出渣进行研究,考察了捕收剂阳离子十八胺、阴离子油酸钠、阴离子丁基黄药的单一浮选,以及全流程浮选对铁酸锌富集提纯的影响。结果表明:采用单一捕收剂浮选,可以除去部分杂质矿物,实现一定程度上的铁酸锌富集。利用全流程浮选,将铁酸锌含量由浮选前的83%提高到92%,铁酸锌富集程度明显提高。研究证明了在不破坏铁酸锌晶体的前提下,将铁酸锌从锌焙砂中富集出来作为独立产品是可能的。     

氯化物中黄钠铁矾除铁的热力学与动力学

对氯化物体系中锰结核浸出溶液采用黄钠铁矾净化除铁的热力学与动力学进行了研究．热力学分析表明，氯化物溶液中氯离子浓度升高能增大黄钠铁矾的溶解度，但在一般情况下（如ｃＣｌ－＜１０ｍｏｌ／Ｌ），完全可以将溶液中的铁以铁矾形式降到理想的程度；黄钠铁矾生成的最佳ｐＨ值范围为１．０～３．０；溶液中钠离子浓度升高有利于铁矾沉淀．动力学研究表明，氯化物中黄钠铁矾形成的表观活化能为９４．６６ｋＪ／ｍｏｌ，沉矾过程受化学反应控制．这些研究结果对锰结核浸出液中铁的分离以及氯化物体系中铁矾的沉淀具有指导作用．     

用还原焙烧法从硫铁矿烧渣中提取铁的研究

研究了高温还原焙烧、硫酸浸取提取硫铁矿烧渣中的铁。用碳作还原剂,800℃以上温度、反应时间约20min时铁的提取率达90%以上,得到的还原渣可以在温和条件下浸出。得到的FeSO4提取液不需再经还原,便于净化,适于作透明氧化铁颜料、磁性铁氧体等高档用品的原料。     

Roasting-induced phase change and its influence on phosphorus removal through acid leaching for high-phosphorus iron ore

In the present study, roasting-induced phase change and its influence on phosphorus removal via leaching has been investigated for high-phosphorus iron ore. The findings indicate that phosphorus in the ore is associated with goethite and exists mainly in amorphous Fe₃PO₇ phase. The phosphorus remains in the amorphous phase after being roasted below 300℃. Grattarolaite (Fe₃PO₇) is found in samples roasted at 600-700℃, revealing that phosphorus phase is transformed from the amorphous form to crystalline grattarolaite during roasting. Leaching tests on synthesized pure grattarolaite reveal a low rate of phosphorus removal by sulfuric acid leaching. When the roasting temperature is higher than 800℃, grattarolaite is found to react with alumina to form aluminum phosphate, and the reactivity of grattarolaite with alumina increases with increasing roasting temperature. Consequently, the rate of phosphorus removal also increases with increasing roasting temperature due to the formation of acid-soluble aluminum phosphate.     

Influence of acid leaching and calcination on iron removal of coal kaolin

Calcination and acid leaching of coal kaolin were studied to determine an effective and economical preparation method of calcined kaolin. Thermogravimetric-differential thermal analysis (TG-DTA) and X-ray diffraction (XRD) demonstrated that 900℃ was the suitable temperature for the calcination. Leaching tests showed that hydrochloric acid was more effective for iron dissolution from raw coal kaolin (RCK), whereas oxalic acid was more effective on iron dissolution from calcined coal kaolin (CCK). The iron dissolution from CCK was 28.78wt%, which is far less effective than the 54.86wt% of RCK under their respective optimal conditions. Through analysis by using M?ssbauer spectroscopy, it is detected that nearly all of the structural ferrous ions in RCK were removed by hydrochloric acid. However, iron sites in CCK changed slightly by oxalic acid leaching because nearly all ferrous ions were transformed into ferric species after firing at 900℃. It can be concluded that it is difficult to remove the structural ferric ions and ferric oxides evolved from the structural ferrous ions. Thus, iron removal by acids should be conducted prior to calcination.     

镍精矿氯气浸出液净化除铁工艺

硫化镍精矿氯气浸出一净化一电积工艺作为镍提取冶金的新方向,其净化过程大多采用溶剂萃取方法．该方法对于含铁较高的浸出液在运行一段时间后会出现萃取有机相性质变坏的问题,影响萃取剂的正常使用．因此,净化过程中必须先进行单独除铁．本试验研究了针铁矿法从镍精矿氯气浸出液中除铁的工艺．实验考察了在氯盐体系中氧化剂用量、反应温度、反应pH值、反应时间等因素对除铁效果的影响．确定了最佳工艺条件氧化剂用量5g／L,反应温度85℃,反应pH值2．5～3．0,反应时间2h．在该工艺条件下,除铁率高达99．8％以上,铁渣中Ni,Co损失小,除铁后液含铁小于0．01g／L,达到了净化要求．     

Comparison of the mechanisms of microwave roasting and conventional roasting and of their effects on vanadium extraction from stone coal

Experiments comparing microwave blank roasting and conventional blank roasting for typical vanadium-bearing stone coal from Hubei Province in central China, in which vanadium is present in muscovite, were conducted to investigate the effects of roasting temperature, roasting time, H₂SO₄ concentration, and leaching time on vanadium extraction. The results show that the vanadium leaching efficiency is 84% when the sample is roasted at 800℃ for 30 min by microwave irradiation and the H₂SO₄ concentration, liquid/solid ratio, leaching temperature, and leaching time are set as 20vol%, 1.5:1 mL·g-1, 95℃, and 8 h, respectively. However, the vanadium leaching efficiency achieved for the sample subjected to conventional roasting at 900℃ for 60 min is just 71% under the same leaching conditions. Scanning electron microscopy (SEM) analysis shows that the microwave roasted samples contain more cracks and that the particles are more porous compared to the conventionally roasted samples. According to the results of X-ray diffraction (XRD) and Fourier-transform infrared (FTIR) analyses, neither of these roasting methods could completely destroy the mica lattice structure under the experimental conditions; however, both methods deformed the muscovite structure and facilitated the leaching process. Comparing with conventional roasting, microwave roasting causes a greater deformation of the mineral structure at a lower temperature for a shorter roasting time.     

锌浸出渣的还原焙烧和人工硫化矿物浮选过程中硫酸铅和硫酸锌的反应行为

To evaluate the feasibility of recovering Pb and Zn sulfides and Ag-containing minerals from Zn leaching residue by the process of reduction roasting followed by flotation, the reaction behaviors of Pb and Zn sulfates during this process were investigated. Chemical analysis showed that the transformation ratios of PbSO₄ and ZnSO₄ could reach 65.51% and 52.12%, respectively, after reduction roasting, and the introduction of a sulfidation agent could improve the transformation ratios of these sulfates. scanning electron microscopy–energy dispersive spectroscopy (SEM–EDS) revealed that temperature obviously affects the particle size, crystal growth, and morphology of the artificial Pb and Zn sulfide minerals. Particle size analysis demonstrated that the particle size of the materials increases after roasting. Flotation tests revealed that a flotation concentrate composed of 12.01wt% Pb, 27.78wt% Zn, and 6.975 × 10?2wt% Ag with recoveries of 60.54%, 29.24%, and 57.64%, respectively, could be obtained after roasting.     

石煤钠化焙烧料酸浸动力学

研究了石煤钠化焙烧料硫酸浸出过程中,浸出剂初始浓度、搅拌速度和浸出温度对浸出率的影响,并对浸出过程动力学进行了分析. 结果表明:浸出剂初始浓度和浸出温度对钒浸出率有显著影响,搅拌速度对钒浸出率影响不大;该浸出过程符合核收缩模型,与化学反应控制动力学方程式相吻合,浸出反应的表观活化能为50.88kJ·mol-1,浸出过程控制步骤为化学反应控制.     

Deep cleaning of a metallurgical zinc leaching residue and recovery of valuable metals

Huge quantities of zinc leaching residues (ZLRs) generated from zinc production are dumped continuously around the world and pose a potential environmental threat because of their considerable amounts of entrained heavy metals (mainly lead). Most ZLRs have not been properly treated and the valuable metals in them have not yet been effectively recovered. Herein, the deep cleaning of a ZLR and recovery of valuable metals via a hydrometallurgical route were investigated. The cleaning process consists of two essential stages:acid leaching followed by calcium chloride leaching. The optimum conditions for extracting zinc, copper, and indium by acid leaching were a sulfuric acid concentration of 200 g·L-1, a liquid/solid ratio of 4:1 (mL/g), a leaching time of 2 h, and a temperature of 90℃. For lead and silver extractions, the optimum conditions were a calcium chloride concentration of 400 g·L-1, a pH value of 1.0, a leaching time of 1 h, and a temperature of 30℃. After calcium chloride leaching, silver and lead were extracted out and the lead was finally recovered as electrolytic lead by electrowinning. The anglesite phase, which poses the greatest potential environmental hazard, was removed from the ZLR after deep cleaning, thus reducing the cost of environmental management of ZLRs. The treatment of chlorine and spent electrolyte generated in the process was discussed.     

Thermodynamics and kinetics of extracting zinc from zinc oxide ore by the ammonium sulfate roasting method

Thermodynamic analyses and kinetic studies were performed on zinc oxide ore treatment by (NH₄)₂SO₄ roasting technology. The results show that it is theoretically feasible to realize a roasting reaction between the zinc oxide ore and (NH₄)₂SO₄ in a temperature range of 573-723 K. The effects of reaction temperature and particle size on the extraction rate of zinc were also examined. It is found that a surface chemical reaction is the rate-controlling step in roasting kinetics. The calculated activation energy of this process is about 45.57 kJ/mol, and the kinetic model can be expressed as follows:1-(1-α)1/3=30.85 exp(-45.57/RT)·t. An extraction ratio of zinc as high as 92% could be achieved under the optimum conditions.