碳含量对煤矸石活化及酸浸提铝的影响

以煤矸石为原料提取氧化铝是煤矸石高值利用的主要途径之一,热活化可有效激发煤矸石中含铝矿物的反应活性,提高酸浸过程中铝的溶出率,是一种行之有效的活化方式。但该过程会受煅烧温度、氧气浓度、有机碳等因素影响。笔者考察了不同氧气浓度、煅烧温度对煤矸石热活化酸浸提铝效率的影响,并对比了煤矸石和无烟煤调配煤矸石中碳含量对煤矸石热活化酸浸提铝效率的影响。结果表明,增加氧气浓度(0~21%)可显著提高煤矸石热活化后铝的浸出率,氧气浓度提高到21%,即接近空气气氛中煅烧时,铝浸出率达到74.24%,氧气浓度进一步提高到50%时对铝浸出率的影响不大。煅烧温度在600~900℃,随煅烧温度升高,煅烧后煤矸石的氧化铝浸出率呈先升高后缓慢下降的趋势;煅烧温度为700℃时,氧化铝浸出率最高,为72.99%;煅烧温度提高到900℃时,铝浸出率大幅度降低,仅为31.77%。煤矸石的烧失量主要是由于碳质等有机质燃烧造成的,对活化后铝的浸出率有一定影响,烧失量随氧气浓度(0~21%)和煅烧温度的提高而增加,增大烧失量有利于提高铝浸出率,主要因为焙烧过程中产生的气体释放量增大,煤矸石变得疏松,使得活性组分与酸的接触更加充分,从而有利于铝的浸出。煤矸石和配碳煤矸石中碳含量对煤矸石热活化酸浸提铝效率有一定影响且规律相似,碳含量低于23%时,铝浸出率随碳含量的增加而降低;碳含量高于23%时,铝浸出率随碳含量的增加而提高。

高灰分碳粉的碱酸法提纯

对某电解铝企业废旧阴极材料进行浮选-酸洗-磨碎处理后获得含灰分高达27.3%的碳粉,然后采用加NaOH焙烧活化-水洗-盐酸酸浸-水洗工艺对该碳粉进行除灰提纯研究。考察了氢氧化钠加入量、焙烧温度、焙烧时间、盐酸酸浸温度对碳粉灰分脱除的影响。结果表明:当设定氢氧化钠/碳粉为0.4,在1000℃焙烧1h,然后采用浓度为1mol/L的盐酸在60℃酸浸1h,可使碳粉中的灰分由27.3%降低到4.04%,碳含量由68.4%提高至94.6%,处理后的碳粉可重返阴极配料利用。

复杂高含铟物料铟浸出工艺研究

某铅冶炼厂的复杂高含铟物料为冶炼的中间产物,此物料采用常规酸浸工艺铟的浸出率较低。针对此问题,本文采用盐酸-硫酸混酸浸出、加压酸浸法、硫酸化焙烧-低酸浸出等3种工艺对此复杂高含铟物料进行了试验,在各试验的较佳工艺条件下,试验结果为:采用盐酸-硫酸混酸浸出工艺,铟的浸出率可以达到75.87%,浸出渣铟含量可以降低至0.32%;采用加压酸性浸出工艺,铟的浸出率可以达到82.54%,浸出渣铟含量可降低至0.21%;采用硫酸化焙烧-低酸浸出工艺,铟的浸出率可以达到93.28%,浸出渣铟含量可降低至0.09%。综合分析对比3种铟的浸出工艺,"硫酸化焙烧-低酸浸出"工艺比较适合此复杂含铟物料提铟。

辽宁长石矿除铁工艺优化的研究

试验针对辽宁某地的长石的特点,首先用化学沉淀法确定长石中的铁含量,在参考已有的长石除铁方法的基础上,综合优化选择出合适的除铁工艺。本试验最终确定的除铁工艺为锤式破碎-球磨机细磨-酸浸-过滤-干燥-成品。采用正交试验的方法研究盐酸酸度、盐酸浸泡时间、盐酸浸泡温度对除铁效果的影响,最终确定20%的盐酸,在酸浸温度为94℃,酸浸时间为5h的条件下长石粉酸浸除铁的效果最好。对比长石除铁前后的X R D图可得,除铁之后长石中铁的峰值明显下降,二氧化硅的峰值由2900增加到17000,酸浸除铁效果显著。对比除铁前后的颜色可得,除铁之后长石的白度明显增加,长石质量明显提高。

攀枝花钛渣和钛精矿在氧化还原改性-酸浸工艺中对比研究

制备高品质沸腾氯化富钛料是钛白生产中的重要环节。采用X射线荧光光谱仪(XRF)、X射线衍射仪(XRD)和扫描电镜(SEM)等分析手段对攀枝花钛渣和钛精矿在氧化还原改性-盐酸法制备富钛料工艺进行了系统的对比研究。结果表明,钛铁矿和钛渣均保持致密结构,且都含有少量酸溶性差的硅酸盐相,但是两者的主要物相分别为酸溶性好的M2O3型固溶体(主要为Fe Ti O3)和酸溶性差的黑钛石M3O5型固溶体(包括Ti3O5,Mg2Ti O5和Fe Ti2O5等)。钛精矿经过改性后会变为多孔的Fe Ti O3,浸出反应活性非常高,常压浸出即可得到合格的沸腾氯化富钛料。而钛渣在现有的氧化还原-常压盐酸浸出工艺中的反应活性要比钛铁矿差很多,改性后钛渣仍有一定含量的难溶M3O5固溶体,且矿相致密结构未被完全破坏,常压浸出处理不能有效除杂,需要加压浸出才能得到高品位富钛料。另外,钛渣硅酸盐杂质相在整个改性和盐酸浸出过程中的变化不大。

Preparation of high pure tellurium from raw tellurium containing Cu and Se by chemical method

High pure tellurium was prepared from raw tellurium containing copper and selenium by chemical method containing oxidation with concentrated nitric acid, leaching with hydrochloric acid, reducing with sulfur dioxide and treating in hydrogen atmosphere at high temperature. Removal ratio of Cu in raw tellurium reaches 99% after raw tellurium is oxidized and leached with HNO3（69%） under the following conditions： 0.96 times stoichiometric quantity of concentrated nitric acid, 4：1 of ratio of liquid to solid, 20 °C of reaction temperature and 30 min of reaction time. Leaching ratio of Te reaches 99% after Te is leached with hydrochloric acid under the following conditions： 1.67 times stoichiometric quantity of hydrochloric acid, 4：1 of ratio of liquid to solid, 20 °C of reaction temperature and 30 min of reaction time. Tellurium powder（99.95%） is obtained when Te（IV） in leachate is reduced with sulfur dioxide. The purity of tellurium increases from 99.954% to 99.999 6% after tellurium（99.95%） is treated in hydrogen atmosphere at 723.15 K for 30 min.

湖北某白钨矿提高选矿精矿品位研究

由于原矿含CaCO3高,白钨粗精矿难以提高品位,试验研究先加药剂浮选出钨粗精矿,再加盐酸酸浸,可以产出回收率为71.18%、品位为69.55%的钨精矿。另外浮选尾矿回收率为27.5%、品位为14.2%,还可作为钨细泥,最大限度提高产品附加值。

高钛渣收尘灰中钛回收工艺研究

高钛渣收尘灰其钛含量是相当可观的,我国生产钛白粉厂比较多,生产过程中所用原料产生尘灰的数量也是值得关注的,如果不加以充分利用,造成的环境污染比较大.实验研究的目的是对二次资源的充分利用,减少环境污染.通过对比目前工业上应用较多的电炉熔炼法、硫酸浸出法、盐酸浸出法、还原锈蚀法的优缺点,再根据高钛渣收尘灰的的组成,实验选择盐酸浸出法,其盐酸在实验过程中可以实现循环利用,大大减少“三废”量.实验主要研究了酸浸浓度,酸浸温度,酸浸矿酸比,酸浸时间对钛的富集和回收的影响.结果表明：当盐酸浓度为6 mol/L,酸浸温度为90℃,矿酸比为1：1.3,酸浸时间为5h,除去铁锰等可溶性杂质的效果比较好,处理渣中的钛含量可以提高到43.17％.

Extraction of molybdenum and nickel from roasted Ni-Mo ore by hydrochloric acid leaching, sulphation roasting and water leaching

To extract molybdenum and nickel from the roasted Ni-Mo ore, a process of hydrochloric acid leaching, sulphation roasting and water leaching was investigated. The results showed that this process could get a high leaching rate of Mo and Ni. Under the optimum conditions of hydrochloric acid leaching （roasted Ni-Mo ore leached with 0.219 mL/g hydrochloric acid addition at 65 ℃ for 30 min with a L/S ratio of 3 mL/g）, sulphation roasting （51.9% sulfiaric acid addition, roasting temperature 240 ℃ for 1 h）, followed by leaching with the first stage hydrochloric acid leaching solution at 95 ℃ for 2 h, the leaching rates of Mo and Ni reached 95.8% and 91.3%, respectively.

Enhancement of CaMoO_(4) calcine decomposition and recovery of calcium resource by HCl cycle leaching

The thermodynamic equilibrium diagrams of CaMoO_(4)-CaSO_(4)-H_(2)SO_(4)-H_(2)O,CaMoO_(4)-HCl-H_(2)O and CaSO_(4)-CaCl_(2)-HCl-H_(2)O systems at 298 K were established.The calculation results demonstrated that HCl displays a much higher solubility of CaSO_(4)than H_(2)SO_(4).The leaching mechanism of Mo from CaMoO_(4)calcine was systematically investigated from the perspective of the micro particle properties variation.HCl exhibits an excellent leaching performance for Mo from CaMoO_(4)calcine due to the elimination of surface coating and the dissolution of a mass of Mo embedded in CaSO_(4)matrix.Excellent Mo leaching efficiency of 99.7%was achieved under the optimal conditions of decomposing CaMoO_(4)calcine by 2.4 mol/L HCl with a liquid/solid ratio of 10:1 at 50°C for 60 min.Based on the experimental results,a highly efficient and green cycle leaching process of molybdenum from molybdenite was proposed,which eliminated surface coating and physical entraining,and converted most of the calcium from CaMoO_(4)calcine into high purity gypsum by-product.

Leaching kinetics of calcium molybdate with hydrochloric acid in presence of phosphoric acid

Calcium molybdate(CaMoO4)is the main component of powellite and is a predominant intermediate in the pyrometallurgical and hydrometallurgical process of molybdenum.The extraction of Mo from CaMoO4 by a combination of phosphoric acid and hydrochloric acid was investigated.For further understanding of the leaching mechanism,the effects of five key factors were studied to describe the leaching kinetics.The results indicated that the dissolution rate of CaMoO4 was independent of the stirring speed.Mo extraction significantly increased with increasing HCl concentration and temperature,but decreased with increasing particle size.A shrinking core model with surface chemical reaction was found to withstand the dissolution of CaMoO4.The apparent activation energy was calculated to be 70.879 kJ/mol,and a semi-empirical equation was derived for the rate of reaction.

Ag recovery from copper anode slime by acid leaching at atmospheric pressure to synthesize silver nanoparticles

In this paper, recovery of silver from anode slime of Sarcheshmeh copper complex in lran and subsequent synthesis of silver nanoparticles from leaching solution is investigated. Sarcheshmeh anode slime is mainly consisted ofCu, Ag, Pb and Se. Amount of Ag in the considered anode slime was 5.4% （by weight）. The goal was to recover as much as possible Ag from anode slime at atmospheric pressure to synthesize Ag nanoparticles. Therefore, acid leaching was used for this purpose. The anode slime was leached with sulfuric and nitric acid from room to 90 ~C at different acid concentrations and the run which yielded the most recovery of Ag was selected for Ag nanoparticles synthesis. At this condition, Cu, Pb and Se are lea- ched as well as Ag. To separate Ag from leach solution HCI was added and silver was precipitated as AgCl which were then dissolved by ammonia solution. The Ag nanoparticles are synthesized from this solution by chemical reduction method by aid of sodium borohydride in the presence of PVP and PEG as stabilizers. The synthesized Ag nanoparticles showed a peak of 394 nm in UV-vis spectrum and TEM images showed a rather uniform Ag nanoparticles of 12 nm.

Electrochemical synthesis of Nb C-Sn composite powder in molten chloride

NbC-Sn composite powder was successfully prepared from SnO2,Nb2O5and carbon by electrochemical reduction andcarbonization in CaCl2-NaCl molten salt at900°C.The reaction pathway was investigated by terminating electrochemicalexperiments for various durations.The influence of carbon on the final products was considered.NbC particles were obtained byleaching the composite with acid.The results showed that the aggregated NbC-Sn composite powdev contained NbC particles about50-100nm and Sn particles about200nm.SnO2was reduced to Sn in the sintering process.Nb2O5was electrochemically reduced toNb in molten salt,experiencing some intermediate products of calcium niobates and niobium suboxides.Nb metal obtained wasconverted to NbC with assistance of carbon.The reduction of Nb oxides may be incomplete and Nb3Sn would be formed if carbon isinsufficient in the cathodic pellet.NbC with good dispersity is produced by leaching NbC-Sn with HCl.

Recovery of lead and silver from zinc acid-leaching residue via a sulfation roasting and oxygen-rich chlorination leaching method

A large amount of acid-leaching residue is produced during the conventional Zn hydrometallurgy process,and this residue has a large concentration of a variety of valuable metals.The purpose of this study was to evaluate the ability of a procedure that entails the use of sulfation roasting,water leaching,and chlorination leaching(blowing oxygen technique)to recover Pb and Ag,followed by cooling crystallization and the replacement of Ag with lead sheet,to realize the full recovery of all valuable metals from zinc acid-leaching residue;consequently,good results were achieved.The best results were obtained under the following conditions:a sulfuric acid at 70%of the raw material quality,roasting temperature of 300°C and roasting time of 2 h,followed by the process of leaching the roasted residue for 1 h by applying a water-to-solid ratio of 5꞉1 at room temperature.The recovery rates of Zn and Fe were 98.69%and 92.36%,respectively.The main parameters of the chlorine salt leaching system were as follows:Cl−concentration of 300 g/L,Fe3+concentration of 25 g/L,acid concentration of 2 mol/L,liquid-to-solid ratio of 9 mL:1 g,temperature of 90℃,and leaching time of 0.5 h;this leaching process was followed by filtration separation.These conditions resulted in high extents of leaching for Pb and Ag(i.e.,98.87%and 96.74%,respectively).Finally,the kinetics of the process of Ag leaching using Cl−ions in an oxygen-rich medium was investigated.It was found that the leaching process was controlled by the diffusion of the product layers,and the activation energy was 19.82 kJ/mol.

Cyclic metallurgical process for extracting Ⅴ and Cr from vanadium slag: Part Ⅰ. Separation and recovery of Ⅴ from chromium-containing vanadate solution

The separation and recovery of V from chromium-containing vanadate solution were investigated by a cyclic metallurgical process including selective precipitation of vanadium,vanadium leaching and preparation of vanadium pentoxide.By adding Ca(OH)_(2) and ball milling,not only the V in the solution can be selectively precipitated,but also the leaching kinetics of the precipitate is significantly improved.The precipitation efficiency of V is 99.59%by adding Ca(OH)_(2) according to Ca/V molar ratio of 1.75:1 into chromium-containing vanadate solution and ball milling for 60 min at room temperature,while the content of Cr in the precipitate is 0.04%.The leaching rate of V reaches 99.35%by adding NaHCO_(3) into water according to NaHCO_(3)/V molar ratio of 2.74:1 to leach V from the precipitate with L/S ratio of 4:1 mL/g and stirring for 60 min at room temperature.The crystals of NH_(4)VO_(3) are obtained by adjusting the leaching solution pH to be 8.0 with CO2 and then adding NH_(4)HCO_(3) according to NH_(4)HCO_(3)/NaVO_(3) molar ratio of 1:1 and stirring for 8 h at room temperature.After filtration,the crystallized solution containing ammonia is reused to leach the precipitate of calcium vanadates,and the leaching efficiency of V is>99%after stirring for 1 h at room temperature.Finally,the product of V_(2)O_(5) with purity of 99.6%is obtained by calcining the crystals at 560℃ for 2 h.

Pressure leaching of zinc silicate ore in sulfuric acid medium

Zinc silicate ore was characterized mineralogically and the results showed that zinc exists mainly as hemimorphite and smithsonite in the sample.Sulfuric acid pressure leaching of zinc silicate ore was carried out to assess the effect of particle size,sulfuric acid concentration,pressure,reaction time and temperature on the extraction of zinc and the dissolution of silica.Under the optimum conditions employed,up to 99.25% of zinc extraction and 0.20% silica dissolution are obtained.The main minerals in leaching residue are quartz and small amounts of undissolved oxide minerals of iron,lead and aluminum are associated with quartz.