硫酸氢铵焙烧高钛渣反应动力学

研究了硫酸氢铵焙烧高钛渣的反应动力学过程,考察钛渣粒度、硫酸氢铵与钛渣质量比和焙烧温度对反应速率的影响,并对反应机理进行分析,建立了反应动力学模型。结果表明:硫酸氢铵焙烧高钛渣提取二氧化钛的动力学过程符合收缩未反应核模型,反应速率受化学反应控制,并由阿罗尼乌斯方程得到焙烧反应的表观活化能为87.95 kJ/mol。     

酸热法处理低品位铝土矿制取氟化铝的研究

采用酸热法处理低品位铝土矿,经过盐酸浸取除铁、加氟化铵焙烧脱硅、氢氟酸超声强化脱硅和浓硫酸精化除钛等步骤制取高纯度氟化铝,优化工艺条件∶控制加入浓盐酸液固比为1∶1,酸浸温度为60℃,酸浸时间为30min;控制氟化铵加入量为反应比的1.4倍,焙烧温度为500℃,焙烧时间为2h;控制加入氢氟酸液固比为3∶1,强化温度为60℃,强化时间为1h;控制加入浓硫酸液固比为3∶1,精化温度为90℃,精化时间为40min。在此条件下,铝的利用率达到90%以上,制得的氟化铝产品的纯度达到97.2%,满足相关化工产品的标准。     

高温焙烧高钛渣工艺的试验研究

以云南地区钛铁矿富集得到的高钛渣为原料,置于加热装置内焙烧,在1000℃下加热处理20min,利用XRD、SEM研究了高温焙烧对高钛渣相组成和微观结构的影响.结果表明:在高温条件下高钛渣中的锐铁型TiO2转化为金红石型TiO2.提出了一种高温焙烧高钛渣的新工艺,从生产源头消除了对环境的污染,实现钛资源的可持续发展.     

H_2SO_4分解富钛精矿的反应动力学

攀钢含钛高炉渣是一种尚待开发利用的钛资源。选用攀钢含钛高炉渣经选择性分离后的富钛精矿为原料 ,以液固异相反应的收缩未反应芯模型研究了硫酸法制取钛白的反应动力学过程。根据搅拌强度和温度对反应的影响判别反应过程为膜扩散控制过程 ;分析了反应速度的温度因素和粒度因素 ,导出了硫酸分解富钛精矿的动力学方程式 ,并按该方程计算得到反应温度在 6 0～ 80℃的活化能为 34.5 9kJ/mol。     

H2SO4分解富钛精矿的反应动力学

攀钢含钛高炉渣是一种尚待开发利用的钛资源,选用攀钢含钛高炉渣选择性分离后的富钛精矿为原料,以液－固异相反应的收缩未反应芯模型研究了硫酸法制取钛白的反应动力学过程,根据搅拌强度和温度对反应的影响判别反应为膜扩散控制过程,分析了反应速度的温度因素和粒度因素,导出了硫酸分离富钛精矿的动力学方程式,并按该方程计算得到反应温度在60－80℃的活化能力为34．59kJ/mol。     

硫酸铵焙烧粉煤灰反应动力学研究

探索了硫酸铵焙烧粉煤灰制备氧化铝的工艺条件,为强化焙烧过程,对硫酸铵焙烧粉煤灰的反应动力学规律进行了系统的研究,确定了焙烧温度、物料配比对铝提取率的影响,结果指出:硫酸铵焙烧粉煤灰的反应受固膜扩散控制,反应活化能为10.83 kJ·mol~(-1),并给出了该焙烧反应的动力学方程。     

富钛料加碳氯化制取TiCl_4的热力学和动力学研究进展

富钛料加碳氯化工艺技术是热力学和动力学理论研究的产物,高钛型高炉渣高温碳化低温氯化技术是氯化冶金的典型代表。通过TiCl_4加碳氯化过程Gibbs自由能分析和反应速率与机理探讨,综述了熔盐氯化过程的多种热力学研究和富钛料加碳氯化过程制备TiCl_4的动力学模型,指出了加碳氯化反应的热力学趋势大小、反应产物CO_2/CO气体的临界温度点以及氯化反应符合未反应核缩减模型的共性,提出了深入研究高温碳化低温氯化过程动力学和热力学的建议。     

高碳石煤流态化氧化焙烧提高钒的浸出率

采用流态化氧化焙烧方式预处理广西某难浸高碳石煤以提高钒的浸出率,对氧化焙烧过程的热力学和钒的氧化动力学进行了分析,并考察了流态化焙烧对钒浸出率的影响。结果表明:石煤在氧化焙烧过程中,碳、黄铁矿的氧化反应在热力学上比V(Ⅲ)氧化反应更易进行,它们的存在对钒氧化具有抑制作用。钒的氧化反应受扩散动力学控制,其表观活化能为347.00 kJ/mol。钒浸出率随焙烧温度的增加先增加后减小,当焙烧温度为700℃和750℃时,钒浸出率随焙烧时间的延长而增加;当焙烧温度为800℃,焙烧时间0.5 h时,钒浸出率最高,达97.51%,延长焙烧时间反而不利于钒浸出。与传统的钠化氧化焙烧法相比,浸出率高,环境污染少。     

硫酸氢铵焙烧高钛渣提取TiO_2

通过X射线衍射和扫描电镜,分析高钛渣的物相结构。通过单因素试验,研究原料粒度、硫酸氢铵与钛渣质量比、焙烧温度和焙烧时间对TiO2提取率的影响。正交实验结果表明:各因素对二氧化钛提取率的影响由大到小的顺序为焙烧温度、焙烧时间、硫酸氢铵与钛渣质量比。最佳的试验条件如下:硫酸氢铵与钛渣质量比4:1、焙烧温度480℃、焙烧时间60 min、原料粒度45~53μm。在最佳实验条件下,TiO2的提取率可达到85%。     

高硅硫化锌精矿氧化焙烧中硅酸锌生成反应的动力学

研究了高硅硫化锌精矿氧化焙烧过程中硅酸锌生成反应的动力学。确定了温度、粒度对硅酸锌反应速率的影响,结果指出：硅酸锌生成反应的动力学符合收缩核西式,其过程为固膜扩散控制。测定了各反应条件下的反应速率常数并测得其活化能为406KJ/mol。提出了硅酸锌生成反应的总动力学方程。限制硅酸锌反应速率的有效方法是适当提高精矿粒度和降低焙烧温度至860℃左右。     

Decomposition reaction kinetics of Baotou RE concentrate with concentrated sulfuric acid at low temperature

Baotou RE concentrate was decomposed with concentrated sulfuric acid by controlling the roasting temperature below 500°C.Thermogravimetry-differential thermal analysis(TG-DTA) and chemical analytical methods were used to study the thermal decomposition process and the thermal decomposition effect.The Freeman-Carroll method was applied to analyze the TG-DTA curves.The activation energy, reaction order, and reaction frequency factor at different stages were calculated.The Satava method was used to deduce the reaction mechanism and the relative reaction rate during the thermal decomposition process.     

Preparation of Ti-rich material from titanium slag by activation roasting followed by acid leaching

A method of activation roasting followed by acid leaching using titanium slag was introduced to prepare Ti-rich material. The effects of H₃PO₄ dosage, roasting temperature, and roasting time on TiO₂ grade were investigated. A Ti-rich material containing 88.54% TiO₂, 0.42% (CaO+MgO) was obtained when finely ground titanium slag was roasted with 7.5% H₃PO₄ at 1000 °C for 2 h, followed by a two-stage leaching in boiling dilute sulfuric acid for 2 h. The XRD patterns show that the product is titanium dioxide with a rutile structure. Mechanism studies show that structures of anosovite solid solution and silicate minerals are destroyed in the roasting process. As a result, titanium components in titanium slag are transformed into TiO₂ (rutile) while impurities are transformed into acid-soluble phosphate and quartz.     

Molybdenite–limestone oxidizing roasting followed by calcine leaching with ammonium carbonate solution

Oxidizing roasting of molybdenite with lime can significantly reduce SO₂ pollution compared with the traditional roasting without lime. However, the calcine is subsequently leached by sulfuric acid, resulting in serious equipment corrosion and abundant non-recyclable CaSO₄ slag. In this work, a novel process, in which the molybdenite was roasted with CaCO₃ followed by (NH₄)₂CO₃ solution leaching, was proposed to improve the art of lime roasting–sulfuric acid leaching. Oxidizing roasting of molybdenite with CaCO₃ was investigated through thermodynamic calculation, thermogravimetric analysis and roasting experiments. The results show that the products of the oxidizing reaction of MoS₂ in the presence of CaCO₃ and O₂ are CaSO₄, CaMoO₄ and CO₂ at 573–1000 K. The MoS₂ conversion rate achieves approximately 99% and the sulfur-retained rate attains approximately 95% with a CaCO₃-to-MoS₂ molar ratio of 3.6 at 500 °C for 1 h by adding 5% mineralizer A (mass fraction). The leaching results show that the leaching rate of Mo reaches 98.2% at 85 °C for 7 h with a (NH₄)₂CO₃ concentration of 600 g/L and a liquid–solid ratio of 10 mL/g. The results presented are potential to develop a novel cleaner technique for ammonium molybdate production.     

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.     

Recovering indium with sulfating roasting from copper-smelting ash

A technology for recovering indium from Jinchuan copper-smelting ash was developed. Indium in the ash was first enriched to the leaching-slag in leaching process,and then recovered by sulfating roasting. The method included mixing the leaching-slag with sulfuric acid,making them into particles,roasting the mixture,and then leaching the calcine with hot water. Above 90% of indium in calcine could be dissolved into the leaching solution. The optimized conditions were determined as follows： the mass ratio of sulfuric acid to leaching slag was 0.1,the roasting time was about 1 to 1.5 h in the temperature range of 200-250℃,and the calcine was leached for 1 h with 5：1 of liquid/solid ratio at 60℃. Over 99% of indium in leaching solution was finally enriched by Zn substitution or sulfide precipitation.     

焙烧氰化尾渣提金工艺研究现状

焙烧氧化过程中铁物相出现熔融或再结晶,对金造成二次包裹,使焙砂中部分金仍难以浸出,导致焙烧氰化尾渣金品位较高。破坏尾渣中铁氧化物对金的包裹可提高金的浸出率。综述了焙烧氰化尾渣主要提金工艺,包括直接酸溶法、还原焙烧法、氯化法、炼铁-电解法、硫酸熟化法和硫脲法等。直接酸溶工艺简单,金浸出效果较差;还原焙烧法金浸出率高,但工艺复杂、能耗大;氯化焙烧法对矿石适应性强,可综合回收有价金属,但基建及维护费用高;炼铁-电解法在富集金的同时可获得纯铁产品,对矿石有较高的要求;硫酸熟化法显著提高金银浸出率,与直接酸溶法相比,所需更高的温度与酸度;硫脲法反应速率快、选择性好,但生产成本较高。 关健词:有色金属冶金;氰化尾渣;铁氧化物;包裹金;提金     

Lithium and manganese extraction from manganese-rich slag originated from pyrometallurgy of spent lithium-ion battery

Mn and Li were selectively extracted from the manganese-rich slag by sulfation roasting?water leaching. The extraction mechanisms of Mn and Li were investigated by means of XRD, TG?DSC, and SEM?EDS. 73.71% Mn and 73.28% Li were leached under optimal experimental conditions: acid concentration of 82 wt.%, acid-to-slag mass ratio of 1.5:1, roasting temperature of 800 °C, and roasting time of 2 h. During the roasting process, the manganese-rich slag first reacted with concentrated sulfuric acid, producing MnSO₄, MnSO₄·H₂O, Li₂Mg(SO₄)₂, Al₂(SO₄)₃, and H₄SiO₄. With the roasting temperature increasing, H₄SiO₄ and Al₂(SO₄)₃ decomposed successively, resulting in generation of mullite and spinel. The mullite formation aided in decreasing the leaching efficiencies of Al and Si, while increasing the Li leaching efficiency. The formation of spinel, however, decreased the leaching efficiencies of Mn and Li.     

低钒转炉钢渣提钒湿法工艺的动力学研究

为了提高湿法浸出低钒钢渣中钒的浸出效率,并对湿法浸出低钒钢渣中钒提供理论依据,从动力学角度分析整个浸出过程。考察温度、液固比、硫酸质量分数和搅拌速率对浸出过程的影响。研究结果表明:在90℃、液固比为10?1以及硫酸浓度6.0mol/L时,浸取9h,低钒钢渣中钒的浸出率可达到95.3%。通过正交实验和动力学推导,得到描述浸出过程的经验方程,低钒钢渣湿法浸出钒的动力学模型为收缩核动力学模型,浸出过程的表观活化能为12.794kJ/mol,该模型表明浸出过程中的控制步骤取决于固膜扩散速率。提高温度、液固比和硫酸质量分数,均可加速钒的浸出速度,提高钒的浸出率。     

硫化锌银精矿富集银的工艺研究

基于国内外硫化锌矿处理的火、湿法研究进展,对含锌银精矿采用硫酸化焙烧、稀硫酸浸出工艺脱除锌、富集银,考察了焙烧和浸出过程中的主要影响因素。结果表明,硫酸配比为150%,在300℃焙烧90 min,以5%稀硫酸为浸出剂,液固比8:1,搅拌转速200~300 r/min,85℃浸出120 min,最终锌的浸出率可达到98%以上,浸出渣中银含量为7.24%,银被富集7倍。     

钢渣反应对高钛钢保护渣物化特性的影响

针对高钛钢连铸过程中钢渣反应及反应前后保护渣性能变化的问题,以高钛钢连铸保护渣为研究对象,通过实验室钢渣反应研究了钢渣的反应性,并对比了2种不同TiO₂含量的连铸保护渣与高钛钢反应前后组分和性能的变化。试验结果表明,在钢中钛质量分数为0.68%时,保护渣中的TiO₂质量分数由10%增加到15%,钢渣间均存在着钢中钛与保护渣中氧化物组分的反应,但TiO₂含量的增加明显抑制了钢渣反应的进行;钢渣反应前后保护渣熔点、黏度和转折温度略有增加;研究的高钛钢保护渣结晶矿相主要为LiTiO₂和BaTiO₃结晶相,钢渣反应对保护渣的主要结晶矿相种类影响较小。