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Manganese was extracted from leaching low-grade manganese ore in sulfuric acid medium. The effects of granule diameter, leaching time, liquid/solid ratio (V/W) and the concentration of sulfuric acid were investigated through orthogonal and single factor experiments. The experimental results showed that the optimal leaching conditions are, size of 0.054mm, 120 minutes of reaction time, 3(V/W) of liquid/solid ratio and 30% of the concentration of sulfuric acid (g/g). Under those conditions, the leaching efficiency is 96.73%. The kinetics of the leaching process is in accordance with the characteristics of fractal reaction. 相似文献
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通过尾矿砂微波加热硫酸溶解新方法的实验,研究了硫酸浓度、液固比及反应时间等因素对尾矿砂酸蚀率的影响。结果显示,在无需搅拌的情况下,微波加热实验的最佳条件酸浓度5mol/L、液固比5.0mL/g都较传统水热法有优势,在反应时间上,微波加热15min就能达到传统水热法2～3h的效果。在酸蚀率相同的情况下,微波加热所用时间仅为传统水热法的1/6,而所用的酸浓度、液固比都较传统水热法小很多。先对尾矿砂直接微波辐照一段时间,然后再加入硫酸进行微波辐照加热溶解,能够促进尾矿砂的溶解,提高尾矿砂的酸蚀率。微波加热条件下,尾矿砂在硫酸浓度8mol/L、液固比5.0mL/g条件下无需搅拌,反应1h后,除透闪石没有完全溶解外,绝大部分的矿物被溶解。与传统加热方式相比,微波加热可显著提高尾矿砂酸溶解速率。 相似文献
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为了粉煤灰的高附加值综合利用,采用酸浸法对准格尔地区循环流化床粉煤灰进行了镓的浸出实验研究,考察了多种因素对镓的浸出率的影响,包括粉煤灰粒度,酸的种类与浓度、酸浸温度与时间、固液比等。结果表明,提高镓的浸出率的适宜条件为:粉煤灰粒度200目,盐酸的浓度6 mol/L,酸浸温度应大于160℃,酸浸时间6 h,液固比在5∶1~6∶1之间为宜。在优选的工艺条件下,镓的浸出率可达80%以上。从粉煤灰中提取镓,使其作为一种资源加以利用,是提高粉煤灰综合利用价值的有效途径。 相似文献
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《矿物学报》2013,(Z1)
Fly ash was used to prepare alumina and silica white, The 3 stages of the process are as follows: ammonium sulfate calcining, acid leaching and alkali dissolution. The optimum conditions for the experiments to determine are as follows: molar ratio of (NH4)2SO4/Al2O3 is 6, the calcining time is 2h, he H2SO4 concentration is 20%, the leaching temperature is 80℃ and dissolution duration is 2h, the ratio of solution and solid reaction material is 6 for ammonium sulfate calcining and acid leaching stage, reaction time 30min, ratio of liquid to ore 5∶1, alkali concentration 45% and reaction temperature 95 ℃for the alkali dissolution stage. Under these conditions, the total leaching efficiencies of Al2O3 and SiO2 are 78.86% and 95%, respectively. The quality of the main products alumina and silica white can meet the national standards of GB/T24487-2009 and GB10517-89, respectively. 相似文献
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石煤钠盐焙烧气氛对钒转化的研究 总被引：3，自引：0，他引：3
石煤钠盐焙烧是复杂的气、液、固多相化学反应,焙烧气氛对钒焙烧转化率有重要影响。本文在钒价态研究工作的基础上,研究了在氧化—氯化焙烧下,提高气相中HCl或Cl_2分压,使钒转化率有不同程度的提高,小试表明能提高5～20％,半工业生产试验提高5.4％。该技术简易可行,实用性强,具有较好的经济效益。 本文对氧化—氯化焙烧的技术条件、焙烧温度、时间、含碳量及钠盐配比等作了全面研究。通过气体成分分析,焙砂浸出液成分分析和物相分析鉴定,对氧化—氯化焙烧提高钒焙烧转化率的机理作了探讨。 相似文献
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研究了铝土矿与硫酸进行加压反应制造硫酸铝的生产工艺，确立了最优工艺操作条件。当采用适当的矿粉粒度、硫酸浓度、原料配比、反应压力进行生产时，反应平稳，铝士矿的铝溶出率高。适当的矿渣洗涤次数可以使铝土矿的铝利用率得到提高。将本研究成果应用于生产实际，充分利用了矿产资源，节约了能源，从而降低了生产成本，达到了节能降耗的目的。 相似文献
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用Lix984萃取分离水钴矿浸出液中的铜钴 总被引：1，自引：0，他引：1
采用Lix984作为萃取剂、硫酸作为反萃剂,从低品位水钴矿还原酸浸液中萃取分离铜、钴。研究了平衡pH值、萃取剂浓度、相比（有机相与水相的体积比）、混合时间、反萃剂浓度、反萃相比等因素对萃取分离的影响,确定了Lix984萃取分离铜、钴的优化条件。结果表明,萃取剂Lix984萃取铜的优化工艺条件：p（Lix984）为40％;平衡pH为1．83;相比为1：1;混合时间为4rain;反萃剂为4．0mol／L的H2SO4。;反萃相比为1：1。在优化条件下,料液经三级逆流萃取和二级反萃,萃取率和反萃率分别达99．O％和99．5％。 相似文献
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钾长石粉酸浸除铁的实验研究 总被引：8，自引：1，他引：8
在利用钾长石粉合成沸石分子筛和制取碳酸钾技术中，铁的存在会降低沸石的白度。对北京平谷、天津蓟县、内蒙白云鄂博三地钾长石粉进行硫酸酸浸除铁实验，获得最大铁浸出率分别为88．6％、93．2％和64．6％，且前两地钾长石粉中铁的浸出行为相似，酸浸除铁效果均优于白云鄂博钾长石粉。采用正交实验法研究硫酸浓度、酸浸温度和时间对除铁效果的影响，表明三者对不同地区钾长石粉酸浸除铁效果的影响程度各不相同。钾长石酸浸除铁反应开始时，铁的溶解极快，反应速率主要由化学反应控制；其后溶解相对缓慢，反应速率则由扩散作用控制。 相似文献
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钾长石粉酸浸除铁废液资源化的实验研究 总被引：3，自引：0，他引：3
钾长石粉经过硫酸或盐酸酸浸除铁后, 废液中含有大量的金属离子和游离酸.对废硫酸进行循环利用实验, 钾长石粉铁的浸出率达到88.3%.对酸浸废液进行蒸发结晶, 分别制备得到了纯度为94.1%的绿钾铁矾和74%的三氯化铁.用硫酸酸浸废液蒸发结晶后, 铁的回收率为70.0%, 钾的回收率为96.5%, 整个工艺无需高温, 无需加压, 操作简单, 具有高效、低能耗、低污染的特点, 表明用本项技术处理酸浸废液行之有效.分析认为, 废液呈强酸性、溶液中存在大量Fe2+、碱金属离子含量偏低是蒸发结晶过程中形成绿钾铁矾而不是黄钾铁矾的主要原因. 相似文献
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Yun Chen Qiming FengYanhai Shao Guofan ZhangLeming Ou Yiping Lu 《International Journal of Mineral Processing》2006
Extraction of molybdenum and vanadium from ammonia leaching residue (main chemical composition: 2.05% Mo, 0.42% V, 65.6% Al2O3 and 10.7% SiO2) of spent catalyst was investigated by roasting the residue with soda carbonate, followed by hydrometallurgical treatment of the roasted products. In the roasting process, over 91.3% of molybdenum and 90.1% of vanadium could be extracted when a charge containing a sodium carbonate to spent catalyst ratio of 0.15 was roasted at 750 °C for 45 min and the roasted mass was leached with water (liquid to solid ratio of 2) at 80–90 °C for 15 min. After the purification of leach liquor, an extraction solvent consisting of 20 vol.% trialkylamine (N235, commercialized in China) and 10 vol.% secondary octyl alcohol (phase modifier) dissolved in sulfonated kerosene was used to extract molybdenum and vanadium in leach liquor. 10 wt.% ammonia water was used as stripping agent. Adding 30 g/l NH4NO3 to the stripping solution and adjusting the pH to 7–8.5, over 99% of vanadium can be crystallized as ammonium metavanadate. Over 98% of molybdenum can be crystallized as ammonium polymolybdate when pH is between 1.5 and 2.5 (pH is adjusted by HNO3). Ammonium metavanadate and ammonium polymolybdate were calcinated at 500–550 °C, the purity of MoO3 and V2O5 was 99.08% and 98.06% respectively. In the whole process, 88.2% of molybdenum and 87.1% of vanadium could be achieved. The proposed roasting, leaching and separation steps give a feasible alternative for the processing of ammonia leaching residue of spent catalyst and can be applied in the comprehensive utilization of low grade molybdenum ores. 相似文献
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江西玉山石煤烧结包裹与钒转化的研究 总被引：7，自引：0，他引：7
江西玉山含钒石煤既是低热值能源,又是一种新类型钒矿资源。为充分利用这种资源,在600～1150℃下进行了氧化培烧和钠化焙烧,测定了煤灰的松装密度d和钒转化率η,研究厂烧结包裹作用。结果表明:温度升高,煤灰烧结严重,对钒包裹几率加大,钒转化率降低。石煤作为能源利用时,燃烧温度不宜超过950℃,石煤钠化焙烧提钒最佳温度775℃。在两种焙烧条件下的石煤烧结机理作了探讨,并进一步得出玉山石煤烧结与包裹的关系式。 相似文献
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铀矿堆浸浸出法进行铀矿采集生产时存在着矿料浸出率过低、溶解酸耗量过大、产生较大环境污染等问题。以江西某铀矿作为研究对象,通过堆浸试验进行分析研究,优化铀矿堆浸的工艺参数:铀矿堆浸高度宜为3.5~4.0m,矿石粒径级配为6~8mm,在喷淋浸出初期(0~30d)采用20g/L的浓硫酸进行喷淋,喷淋强度宜为15L/(h·m2),在喷淋浸出期(30d~60d)采用20g/L的浓硫酸进行喷淋,喷淋强度宜为20L/(h·m2),在喷淋浸出后期(60d~120d)采用10g/L的浓硫酸进行喷淋,喷淋强度宜为8L/(h·m2)。 相似文献
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《Chemie der Erde / Geochemistry》2021,81(4):125773
As the most abundant copper containing resource and zinc containing resource, chalcopyrite and sphalerite/marmatite commonly coexist as Cu-Zn mixed ores in deposits. However, it is difficult to completely separate sphalerite and chalcopyrite by flotation, thus resulting in the existence of zinc impurity in copper concentrate. Sphalerite/marmatite existed in copper sulfide concentrate as impurity may lead to severe damage of the smelting equipment, and cause the waste of copper and Zn resources, it will also decrease of the sale price of copper concentrates. Therefore, the deep separation of zinc from zinc bearing copper sulfide concentrate is of great significance. In this work, selective chemical leaching was developed to efficiently remove zinc from zinc containing copper sulfide concentrate. Under the optimal condition (i.e., sulfuric acid concentration exceed 100 g/L, temperature of 80 °C, pulp density of 10%, leaching time of 48 h), over 85% Zn was extracted into the leaching solution together with only about 10% Cu and Fe, according to the leaching experiment. Leaching slurry had good solid-liquid separation characteristics, and zinc can be further effectively recovered from the leaching solution. According to X-ray diffraction (XRD) and scanning electron microscope/energy dispersive spectrometer (SEM/EDS) analysis, chalcopyrite was the main mineralogical phase in the residues, which can be regarded as high quality copper concentrate for metallurgy. Accordingly, a new process for deep and efficient separation of Cu-Zn mixed ores has been proposed. 相似文献
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《China Geology》2023,6(1):137-153
Lithium production in China mainly depends on hard rock lithium ores, which has a defect in resources, environment, and economy compared with extracting lithium from brine. This paper focuses on the research progress of extracting lithium from spodumene, lepidolite, petalite, and zinnwaldite by acid, alkali, salt roasting, and chlorination methods, and analyzes the resource intensity, environmental impact, and production cost of industrial lithium extraction from spodumene and lepidolite. It is found that the sulfuric acid method has a high lithium recovery rate, but with a complicated process and high energy consumption; alkali and chlorination methods can directly react with lithium ores, reducing energy consumption, but need to optimize reaction conditions and safety of equipment and operation; the salt roasting method has large material flux and high energy consumption, so require adjustment of sulfate ratio to increase the lithium yield and reduce production cost. Compared with extracting lithium from brine, extracting lithium from ores, calcination, roasting, purity, and other processes consume more resources and energy; and its environmental impact mainly comes from the pollutants discharged by fossil energy, 9.3–60.4 times that of lithium extracted from brine. The processing cost of lithium extraction from lepidolite by sulfate roasting method is higher than that from spodumene by sulfuric acid due to the consumption of high-value sulfate. However, the production costs of both are mainly affected by the price of lithium ores, which is less competitive than that of extracting lithium from brine. Thus, the process of extracting lithium from ores should develop appropriate technology, shorten the process flow, save resources and energy, and increase the recovery rate of related elements to reduce environmental impact and improve the added value of by-products and the economy of the process.©2023 China Geology Editorial Office. 相似文献
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A new method of reduction of core samples of uranium ore to their natural redox state is described. The results of laboratory experiments on agitated and filtration leaching of uranium from ore reduced by sulfide solution are considered in comparison with leaching of oxidized stored ore. The oxidation of ore samples during sampling and storage leads to the overestimation of the rate of uranium leaching from ore by sulfuric acid solutions. The more oxidized a sample, the higher the overestimation of this parameter in comparison with traditional borehole in situ leaching (ISL) with sulfuric acid. As is proved by determination of U and Fe species in reduced samples, the described technique of ore sample preparation for laboratory experiments enables the natural redox state of ore to be approached. The results of laboratory experiments with samples prepared according to the proposed technique are adequate to the real process of ISL with sulfuric acid at the Dalmatovo uranium deposit. 相似文献