石煤钒矿硫酸活化常压浸出提钒工艺

研究石煤钒矿的硫酸活化提钒方法。分别考察矿石粒度、硫酸浓度、活化剂用量、催化剂用量、反应温度、反应时间和浸出液固比等因素对钒浸出率的影响。结果表明:石煤提钒的优化条件为矿石粒度小于74μm的占80%、硫酸浓度150 g/L、活化剂CaF2用量(相对于矿石)60 kg/t、催化剂R用量20 g/L、反应温度90℃、反应时间6 h、液固比(体积/质量,mL/g)2:1,在此优化条件下,钒浸出率可达94%以上;在优化条件下,采用两段逆流浸出,可有效减少活化剂CaF2以及浸出剂硫酸的消耗量;经过两段逆流浸出萃取反萃氧化水解工艺,全流程钒资源总回收率可达86.9%;V2O5产品纯度高于99.5%。

石煤硫酸浸出过程中含钒矿物的浸出行为

以陕西省商南县石煤钒矿为研究对象,研究了直接酸浸、氧化酸浸及活化酸浸的提钒效果,分析了酸浸过程中含钒矿物钒的浸出规律和活化浸出机理。结果表明:在硫酸浓度250 g/L、液固比5 mL/g、温度90℃条件下,浸出12 h,钒浸出率仅为48.55%;在相同条件下,加入3%氯酸钠,氧化酸浸9 h,钒浸出率增加到61.42%;再添加9%氟化钙,仅活化浸出6 h,钒浸出率可达到92.90%。工艺矿物学研究表明:在活化酸浸过程中,氟化钙与硫酸反应生成的氢氟酸,促进了硫酸对含钒矿物结构的破坏,同时参与了含钒云母的反应,促进了含钒矿物中钒的浸出;矿物中含钒云母、钒钛矿、褐铁矿、钒硫化物中钒的浸出率分别为98.47%、42.22%、98.09%、89.70%,与氧化酸浸相比,分别增加了31.77、26.90、18.01、89.70个百分点。

活化-酸浸法提取粉煤灰中氧化铝的溶出规律研究

介绍了用活化-酸浸法提取粉煤灰中氧化铝的方法,考察了粉煤灰煅烧活化和硫酸浸出条件对粉煤灰中铝溶出率的影响。通过实验得出最佳工艺条件为:煅烧活化处理温度为750℃,碳酸钠和碳酸钾的混合物与粉煤灰的比例为2∶1,煅烧活化保温时间为45min,硫酸浓度为0.6mol/L,水浴溶出温度为80℃,氧化铝的最终溶出率为98.36%。

基于加压酸浸工艺的天然石墨低氟提纯方法

石墨经氢氟酸法提纯后环保成本高,且产品纯度不能满足现状。采用3种改进工艺对天然石墨提纯,研究改进工艺条件下不同石墨原料的浸出效果,得到最佳的石墨原料与工艺,并对相应样品进行X射线衍射和扫描电子显微镜表征。结果表明:在温度50℃、氩气压力3 MPa条件下,可将球形石墨尾料的纯度由97.69%提高至99.54%;改进的工艺中,以球形石墨尾料为原料的无氟酸浸-焙烧活化-加压酸浸法,不仅可得到99.979%的高纯石墨,而且可从源头减少环保成本。

蛇纹石短流程加工新工艺研究

作者对现有蛇纹石综合利用工艺流程所存在的问题进行了分析，并在此基础上，研究了蛇纹石短流程加工新工艺。结果表明．采用活化酸浸新技术和新的镁回收方法，可提高蛇纹石中各有用成份的利用率，缩短其加工工艺流程，降低原材料消耗。

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 HaPO4 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 ℃ 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 （futile） while impurities are transformed into acid-soluble phosphate and quartz.

Effects of mechanical activation and oxidation-reduction on hydrochloric acid leaching of Panxi ilmenite concentration

The effects of oxidation-reduction treatment and mechanical activation on the hydrochloric acid leaching performance of Panxi ilmenite concentration were investigated.The results show that both of oxidation-reduction treatment and mechanical activation significantly accelerate the extraction of Fe,Ca and Mg from Panxi ilmenite concentration;however,the CaO and MgO contents of the calcined residues obtained from oxidized-reduced ilmenite concentration are higher than the standard values required by chlorination process.The Ca and Mg in oxidized-reduced ilmenite concentration can be leached much faster after mechanical activation,yielding a synthetic rutile which meets the requirements of chlorination process containing 90.50% TiO2 and 1.37% total iron as well as combined CaO and MgO of 1.00%.The optimum oxidation and reduction conditions are as follows：oxidization at 900 ℃ in the presence of oxygen for 15 min and reduction at 750 ℃ by hydrogen for 30 min.

Effects of acid pretreatment on Fe-ZSM-5 and Fe-beta catalysts for N_2O decomposition

Two series of ZSM-5 and beta zeolites were pretreated in 1.0 mol/L HNO3 solution at room temper-ature for various time periods. The catalytic performances of their Fe-exchanged products in N2O decomposition were evaluated. The Fe-zeolite catalysts were characterized using N2 adsorp-tion-desorption, inductively coupled plasma optical emission spectroscopy, X-ray diffraction, ultra-violet-visible spectroscopy, temperature-programmed desorption of NH3, and scanning and trans-mission electron microscopies. For the ZSM-5 zeolite, acid leaching primarily takes place on the crystal surface and the particle size is reduced, therefore the pore channels are shortened. However, because of the good stability of MFI zeolites, the acid does not greatly penetrate the pore channels and new mesopores are not created. For the beta zeolite, because the amorphous material is in-clined to dissolve（deagglomerate）, some of the micropores are slightly dilated. The improved cata-lytic activities can be explained by the increased active Fe loading as a result of structural changes.

Effect of mechanical activation on thiosulfate leaching of gold from complex sulfide concentrate

The use of mechanical activation to enhance gold recovery from a CuPbZn complex sulfide concentrate was investigated. The effects of milling time, ball size, sample to ball ratio and milling speed on thiosulfate leaching were studied. Under optimum conditions of milling time 1 h, ball size 20 mm, sample to ball ratio 1/15 and mill speed 600 r/min, nearly 78% of sample is amorphized, particle size decreases from d100=30 μm to d100=8 μm, specific surface area increases from 1.3 m2/g to 4.6 m2/g and gold recovery enhances from 17.4 % in non-activated sample to 73.26 %.

Technological conditions and kinetics of leaching copper from complex copper oxide ore

The kinetic behavior of leaching copper from low grade copper oxide ore was investigated. The effects of leaching temperature, H2SO4 concentration, particle size of crude ore and agitation rate on the leaching efficiency of copper were also evaluated. And the kinetic equations of the leaching process were obtained. The results show that the leaching process can be described with a reaction model of shrinking core. The reaction can be divided into three stages. The first stage is the dissolution of free copper oxide and copper oxide wrapped by hematite-limonite ore. At this stage, the leaching efficiency is very fast （leaching efficiency is larger than 60%）. The second stage is the leaching of diffiuent copper oxides, whose apparent activation energy is 43.26 kJ/mol. During this process, the chemical reaction is the control step, and the reaction order of H2SO4 is 0.433 84. The third stage is the leaching of copper oxide wrapped by hematite-limonite and silicate ore with apparent activation energy of 16.08 kJ/mol, which belongs to the mixed control.

Conversion of coal gangue into alumina, tobermorite and TiO_2-rich material

A large amount of coal gangue from coal mining and processing is regarded as waste and usually stockpiled directly. In order to recycle the valuable elements from the coal gangue, an integrated process is proposed. The process consists of three steps: 1concentrating alumina from the coal gangue via activation roasting followed by alkali leaching of Si O2 which produces alumina concentrate for alumina extraction by the Bayer process; 2) synthesizing tobermorite whiskers from the filtrated alkali liquo containing silicate via a hydrothermal method and reusing excess caustic liquor; and 3) enriching titanium component from the Baye process residue by sulfuric acid leaching. Alumina concentrate with 69.5% Al_2O_3 and mass ratio of alumina to silica(A/S) of 5.9pure 1.1 nm tobermorite whisker and TiO_2-rich material containing 33% TiO_2 are produced, respectively, with the optimal parameters Besides, the actual alumina digestion ratio of alumina concentrate reaches 80.4% at 270 oC for 40 min in the Bayer process.