富硼渣冷却速率与活性的关系

以硼铁矿13 m3高炉分离生产的富硼渣为原料进行了缓冷试验.通过改变富硼渣熔体的冷却条件,在1500~1 200℃区间控制冷却速率从0.76℃/min到20℃/min,在1200~900℃区间,冷却速率应小于2℃/min.富硼渣活性可由40.05%增加到83.72%.冷却速率越高,富硼渣的活性越大.这一试验结果与实验室研究结果相符合.试验结果表明现场可以实现两段式冷却提高活性,并为今后冷却装置设计指明了方向.     

Carbothermic reduction characteristics of ludwigite and boron-iron magnetic separation

Ludwigite is a kind of complex iron ore containing boron, iron, and magnesium, and it is the most promising boron resource in China. Selective reduction of iron oxide is the key step for the comprehensive utilization of ludwigite. In the present work, the reduction mechanism of ludwigite was investigated. The thermogravimetry and differential scanning calorimetry analysis and isothermal reduction of ludwigite/coal composite pellet were performed. Ludwigite yielded a lower reduction starting temperature and a higher final reduction degree compared with the traditional iron concentrates. Higher specific surface area and more fine cracks might be the main reasons for the better reducibility of ludwigite. Reducing temperature highly affected the reaction fraction and microstructure of the reduced pellets, which are closely related to the separation degree of boron and iron. Increasing reducing temperature benefited the boron and iron magnetic separation. Optimum magnetic separation results could be obtained when the pellet was reduced at 1300℃. The separated boron-rich non-magnetic concentrate presented poor crystalline structure, and its extraction efficiency for boron reached 64.3%. The obtained experimental results can provide reference for the determination of the comprehensive utilization flow sheet of ludwigite.     

Reduction-melting behaviors of boron-bearing iron concentrate/carbon composite pellets with addition of CaO

Although the total amount of boron resources in China is high, the grades of these resources are low. The authors have already proposed a new comprehensive utilization process of boron-bearing iron concentrate based on the iron nugget process. The present work de-scribes a further optimization of the conditions used in the previous study. The effects of CaO on the reduction–melting behavior and proper-ties of the boron-rich slag are presented. CaO improved the reduction of boron-bearing iron concentrate/carbon composite pellets when its content was less than 1wt%. Melting separation of the composite pellets became difficult with the CaO content increased. The sulfur content of the iron nugget gradually decreased from 0.16wt%to 0.046wt%as the CaO content of the pellets increased from 1wt%to 5wt%. CaO negatively affected the iron yield and boron extraction efficiency of the boron-rich slag. The mineral phase evolution of the boron-rich slag during the reduction–melting separation of the composite pellets with added CaO was also deduced.     

Effect of carbon species on the reduction and melting behavior of boron-bearing iron concentrate/carbon composite pellets

Iron nugget and boron-rich slag can be obtained in a short time through high-temperature reduction of boronbearing iron concentrate by carbonaceous material, both of which are agglomerated together as a carbon composite pellet. This is a novel flow sheet for the comprehensive utilization of boron-bearing iron concentrate to produce a new kind of man-made boron ore. The effect of reducing agent species (i.e., carbon species) on the reduction and melting process of the composite pellet was investigated at a laboratory scale in the present work. The results show that, the reduction rate of the composite pellet increases from bituminite, anthracite, to coke at temperatures ranging from 950 to 1300℃. Reduction temperature has an important effect on the microstructure of reduced pellets. Carbon species also affects the behavior of reduced metallic iron particles. The anthracite-bearing composite pellet melts faster than the bituminitebearing composite pellet, and the coke-bearing composite pellet cannot melt due to the high fusion point of coke ash. With anthracite as the reducing agent, the recovery rates of iron and boron are 96.5% and 95.7%, respectively. This work can help us get a further understanding of the new process mechanism.     

Extraction of vanadium from high calcium vanadium slag using direct roasting and soda leaching

The extraction of vanadium from high calcium vanadium slag was attempted by direct roasting and soda leaching. The oxidation process of the vanadium slag at different temperatures was investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM), and energy dispersive spectroscopy (EDS). The effects of roasting temperature, roasting time, Na₂CO₃ concentration, leaching temperature, leaching time, and liquid to solid ratio on the extraction of vanadium were studied. The results showed that olivine phases and spinel phases in the vanadium slag were completely decomposed at 500 and 800℃, respectively. Vanadium-rich phases were formed at above 850℃. The leaching rate of vanadium reached above 90% under the optimum conditions:roasting temperature of 850℃, roasting time of 60 min, Na₂CO₃ concentration of 160 g/L, leaching temperature of 95℃, leaching time of 150 min, and liquid to solid ratio of 10:1 mL/g. The main impurities were Si and P in the leach liquor.     

Extraction of Aluminum and Iron from Boiler Slag by Sulfuric Acid

This paper presents a new method of recycling aluminum and iron in boiler slag derived from plants that use coal as fuel. The new method integrates efficient extraction and reuse of the leached pellets together. An elemental analysis of aqueous solutions leached by sulfuric acid was determined by EDTA-Naz-ZnCl2 titration method. The components and microstructures of the samples were examined by means of XRF, XRD and SEM. An aluminum extraction efficiency of 86.50% was achieved when the sintered pellets were leached using 4 mol · L^- 1 H2SO4 at solid/ liquid [m（g）/V（mL）] ratio of 1 ： 5 at 80 ℃ for 24 h. An iron extraction efficiency of 94.60% was achieved in the same conditions for the maximum extraction efficiency of Al. The extraction efficiencies of Al and Fe increased with an increase in temperature, leaching time and acidity. The concentration of alumina and iron hydroxide in the final product was determined to be 99.12% and 92.20% respectively. This product of alumina would be used directly for the production of metallic aluminum.     

Reductive leaching of manganese from low-grade pyrolusite ore in sulfuric acid using pyrolysis-pretreated sawdust as a reductant

Manganese (Mn) leaching and recovery from low-grade pyrolusite ore were studied using sulfuric acid (H₂SO₄) as a leachant and pyrolysis-pretreated sawdust as a reductant. The effects of the dosage of pyrolysis-pretreated sawdust to pyrolusite ore, the concentration of sulfuric acid, the liquid/solid ratio, the leaching temperature, and the leaching time on manganese and iron leaching efficiencies were investigated. Analysis of manganese and iron leaching efficiencies revealed that a high manganese leaching efficiency was achieved with low iron extraction. The optimal leaching efficiency was determined to be 20wt% pyrolysis-pretreated sawdust and 3.0 mol/L H₂SO₄ using a liquid/ solid ratio of 6.0 mL/g for 90 min at 90℃. Other low-grade pyrolusite ores were tested, and the results showed that they responded well with manganese leaching efficiencies greater than 98%.     

铜冶炼闪速炉烟尘氧化浸出与中和脱砷

介绍了废酸氧化浸出铜冶炼闪速炉烟尘和漫出液中和沉淀砷、铁过程。从化学热力学和实验2方面研究了浸出液中以砷酸铁形式中和沉淀脱砷过程，并对砷酸铁沉淀的稳定性进行了研究。研究结果表明：闪速炉烟尘中铜、砷和铁的浸出率分别可达到83％，92％和30％，浸出液中的铁和砷的量比n（Fe）／n（As）约为1．50；控制适当的pH值中和沉淀砷、铁，可使铜存留于溶液中，而砷以砷酸铁形式进入固相中，从而达到铜、砷分离的目的；不稳定的砷酸铁沉淀物进一步转型后，则可作为无毒稳定渣丢弃。     

钒渣无焙烧浸出液中钒铁萃取分离

利用本课题组提出的钛白废酸无焙烧加压浸出钒渣提钒的新技术,以P204为萃取剂从废酸浸出钒渣的浸出液中进行了提钒研究.实验结果表明:采用亚硫酸钠为浸出液预处理还原剂,将浸出液中三价铁还原成二价铁,从而防止三价铁的共萃;常温条件下,当浸出液初始p H=2.5、水相与有机相体积比为1∶3,震荡时间为4 min时,采用有机相组成为20%P204及10%TBP协同萃取体系,钒的萃取率可达98.61%以上,钒铁的分离系数可达135.3.     

Extraction of vanadium from vanadium slag by high pressure oxidative acid leaching

To extract vanadium in an environment friendly manner, this study focuses on the process of leaching vanadium from vanadium slag by high pressure oxidative acid leaching. Characterizations of the raw slag, mineralogy transformation, and the form of leach residues were made by X-ray diffraction, scanning electron microscopy, and energy dispersive X-ray spectroscopy. The result shows that the vanadium slag is composed of major phases of fayalite, titanomagnetite, and spinel. During the high pressure oxidative acid leaching process, the fayalite and spinel phases are gradually decomposed by sulfuric acid, causing the release of vanadium and iron in the solution. Meanwhile, unreacted silicon and titanium are enriched in the leach residues. With the initial concentration of sulfuric acid at 250 g·L-1, a leaching temperature of 140℃, a leaching time of 50 min, a liquid-solid ratio of 10:1 mL·g-1, and oxygen pressure at 0.2 MPa, the leaching rate of vanadium reaches 97.69%.     

不同钙化剂对高钒渣酸浸提钒的影响

针对传统钒渣钠化焙烧-水浸提钒工艺的不足,确定对钒渣钙化焙烧-酸浸提钒进行研究。在理论分析的基础上,本研究以高钒渣为原料,研究了钙化焙烧-酸浸提钒过程中3种钙化剂(CaSO4、CaCO3、CaO)的焙烧机理以及对提钒效果的影响。研究结果表明:钒浸出率随焙烧温度的升高先增大后减小,且在1 450K时达到最大值;钙化剂配比为100%CaSO4时提钒率最大;在目前实验室研究条件下,钒的浸出率最大可达93.53%。     

酸浸法提钒新工艺的研究

研究了用稀硫酸直接浸出—萃取—反萃—氨水沉钒—煅烧的提钒工艺。结果表明,采用稀硫酸直接浸出,原矿渣中总钒的一次浸取率可达95%以上;用萃取-反萃方式净化和浓缩浸出液,同时使用萃取促进剂处理酸浸液,使萃取效率比传统方法有明显提高,萃取级数大大减少;沉钒步骤摒弃了传统的铵盐沉钒工艺,使用氨水直接沉钒,提高了产品的纯度。钒的总回收率达86%以上,比传统提钒工艺效率提高了20%以上,同时由于避免了焙烧从而解决了传统提钒过程中因焙烧等产生的HCl、Cl₂等污染问题。     

A method for recovery of iron,titanium, and vanadium from vanadium-bearing titanomagnetite

An innovative method for recovering valuable elements from vanadium-bearing titanomagnetite is proposed. This method involves two procedures:low-temperature roasting of vanadium-bearing titanomagnetite and water leaching of roasting slag. During the roasting process, the reduction of iron oxides to metallic iron, the sodium oxidation of vanadium oxides to water-soluble sodium vanadate, and the smelting separation of metallic iron and slag were accomplished simultaneously. Optimal roasting conditions for iron/slag separation were achieved with a mixture thickness of 42.5 mm, a roasting temperature of 1200℃, a residence time of 2 h, a molar ratio of C/O of 1.7, and a sodium carbonate addition of 70wt%, as well as with the use of anthracite as a reductant. Under the optimal conditions, 93.67% iron from the raw ore was recovered in the form of iron nugget with 95.44% iron grade. After a water leaching process, 85.61% of the vanadium from the roasting slag was leached, confirming the sodium oxidation of most of the vanadium oxides to water-soluble sodium vanadate during the roasting process. The total recoveries of iron, vanadium, and titanium were 93.67%, 72.68%, and 99.72%, respectively.     

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.     

Upgradation of bauxite by molecular hydrogen and hydrogen plasma

An approach was developed to upgrade the bauxite ore by molecular hydrogen and hydrogen plasma. A gibbsite-type bauxite sample was obtained from National Aluminium Company (NALCO), Odisha, India. The obtained sample was crushed and sieved (to 100 μm) prior to the chemical analysis and grain-size distribution study. The bauxite sample was calcined in the temperature range from 500 to 700°C for different time intervals to optimize the conditions for maximum moisture removal. This process was followed by the reduction of the calcined ore by molecular hydrogen and hydrogen plasma. Extraction of alumina from the reduced ore was carried out via acid leaching in chloride media for 2 h at 60°C. X-ray diffraction, scanning electron microscopy, thermogravimetry in conjunction with differential scanning calorimetry, and Fourier transform infrared spectroscopy were used to determine the physicochemical characteristics of the material before and after extraction. Alumina extracted from the reduced ore at the optimum calcination temperature of 700°C and the optimum calcination time of 4 h is found to be 90% pure.     

Boron separation from Si-Sn alloy by slag treatment

This study investigates a purification process for metallurgical-grade silicon (MG-Si) in which Si is alloyed with tin (Sn) and CaO-SiO₂-CaCl₂ slag is used to remove boron (B) impurity. Acid leaching was performed to remove the Sn phase after slag refining to recover high-purity Si from the Si-Sn alloy. The effect of refining time was investigated, and acceptable refining results were realized within 15 min. The effects of slag composition and Sn content on the removal of B were also studied. The results indicate that increasing Sn content favors B removal. With the increase of Sn to 50% of the alloy, the final B content decreased to 1.1×10-4wt%, 93.9% removal efficiency.     

Dissolution kinetics of iron oxides in clay in oxalic acid solutions

Clay samples containing 8.15% iron oxides and 27.49% alumina were leached in oxalic acid. Leaching experiments were performed in aqueous solutions of oxalic acid of 0.2-2 mol/L at 40-80℃ for up to 90 min. The mixed kinetic mechanism, i.e., t/τ=[(1-2X/3)-(1-X)2/3]+b[1-(1-X)1/3], seemed to be the most appropriate one to fit the kinetic data of leaching iron oxides contained in clay in the aqueous oxalic acid solutions. The Arrhenius activation energy for leaching in the 1.8 mol/L oxalic acid was found to be 41.035 kJ/mol.     

三相流态化浸出锑精炼渣的研究

本文对锑精炼渣进行了三相流态化浸出的小型及扩大试验研究,探讨了锑精炼渣水浸分离砷锑的控制环节。考察了锑精炼渣浆的流变特性,讨论了温度、粒度、时间液固比等因素对浸出过程的影响。小型及扩大试验均表明,在常温和液固比为2.0的条件下浸出30min,砷浸出率可达97%。三相流态化浸出可在低液固比下进行,能强化浸出过程,缩短浸出时间。     

我国硼工业关键技术的碳足迹分析

为分析我国硼工业对全球气候变化的影响,寻求节能减排的途径,采用德国GaBi 4.4生命周期评价软件对硼工业的关键技术进行了碳足迹的评价与分析.研究结果表明:在碳碱法生产硼砂工艺中,以硼精粉为原料产生的总碳足迹最大,以富硼渣为原料产生的总碳足迹最小,利用富硼渣生产硼砂既可以节约硼资源,又降低了环境影响;在硼酸生产工艺中,硼铁矿盐析法产生的总碳足迹最大,硼镁矿浮选分离法产生的总碳足迹最小,两步法产生的直接碳足迹最小,但总碳足迹则不占优势;降低能源消耗量,尤其是煤的消耗量对降低整个研究系统的碳足迹至关重要.     

Extraction and separation of nickel and cobalt from saprolite laterite ore by microwave-assisted hydrothermal leaching and chemical deposition

Extraction and separation of nickel and cobalt from saprolite laterite ore were studied by using a method of microwave-assisted hydrothermal leaching and chemical deposition. The effects of leaching temperature and time on the extraction efficiencies of Ni2+ and Co2+ were investigated in detail under microwave conditions. It is shown that the extraction efficiencies of Ni2+ and Co2+ from the ore pre-roasted at 300℃ for 5 h were 89.19% and 61.89% when the leaching temperature and time were about 70℃ and 60 min, respectively. For the separation process of Ni and Co, the separation of main chemical components was performed by adjusting the pH values of sulfuric leaching solutions using a NaOH solution based on the different pH values of precipitation for metal hydroxides. The final separation efficiencies of Ni and Co were 77.29% and 65.87%, respectively. Furthermore, the separation efficiencies of Fe of 95.36% and Mg of 92.2% were also achieved at the same time.