Enhanced properties of stone coal-based composite phase change materials for thermal energy storage

Phase change materials (PCMs) can be incorporated with low-cost minerals to synthesize composites for thermal energy storage in building applications.Stone coal (SC) after vanadium extraction treatment shows potential for secondary utilization in composite preparation.We prepared SC-based composite PCMs with SC as a matrix,stearic acid (SA) as a PCM,and expanded graphite (EG) as an additive.The combined roasting and acid leaching treatment of raw SC was conducted to understand the effect of vanadium extraction on promoting loading capacity.Results showed that the combined treatment of roasting at 900℃ and leaching increased the SC loading of the composite by 6.2%by improving the specific surface area.The loading capacity and thermal conductivity of the composite obviously increased by 127%and 48.19%,respectively,due to the contribution of 3wt% EG.These data were supported by the high load of 66.69%and thermal conductivity of 0.59 W·m^(-1)·K-1of the designed composite.The obtained composite exhibited a phase change temperature of 52.17℃,melting latent heat of 121.5 J·g^(-1),and good chemical compatibility.The SC-based composite has prospects in building applications exploiting the secondary utilization of minerals.

Effect of vanadium occurrence state on the choice of extracting vanadium technology from stone coal

The extraction technology of vanadium from stone coal by means of no-salt-roasting and dilute acid solution leaching treatment has the advantages of low pollution, low investment, as well as high vanadium leaching efficiency, which makes it the best technology for extending application. In the present study, the effects of vanadium occurrence state in stone coal, roasting temperature and additive agent on the leaching efficiency of vanadium were studied. The results indicate that the effect of vanadium occurrence state on the extracting vanadium technology is obvious. If the vanadium component in stone coal existed in amorphous phase form, this type of stone coal can be treated by no-slat-roasting and dilute acid solution leaching technology; while the vanadium-bearing crystalline phase existed in stone coal, the roasting additive must be added to destroy this crystalline form so as to acquire high vanadium leaching rate.

A new process of extracting vanadium from stone coal

A new process of extracting vanadium from the stone coal vanadium ore in Fangshankou, Dtmhuang area of Gansu Province, China was introduced. Various leaching experiments were carried out, and the results show that the vanadium ore in Fangshankou is difficult to process due to its high consumption of acid and the high leaching rate of impurities. However, the leaching rate can be up to 80% and the content of V2O5 in the residue can be between 0.22%-0.25% in the process of ore fme grinding→oxidation roasting→mixing and ripening→aqueous leaching→P2O4 solvent extraction→sulfuric acid stripping→oxidation and precipitation→decomposition by heat. Also, the quality of flaky V205 produced by this process can meet the requirements of GB3283--87. The total leaching rate of vanadium is 70%. Also, three types of wastes are easy to treat. The vanadium extraction process is better in relation to the aspect of environmental protection than the sodium method.

Bioleaching of vanadium from barren stone coal and its effect on the transition of vanadium speciation and mineral phase

This study determined the optimal conditions required to obtain maximum vanadium extraction and examined the transition of mineral phases and vanadium speciation during the bioleaching process. Parameters including the initial p H value, initial Fe^(2+) concentration, solid load, and inoculum quantity were examined. The results revealed that 48.92 wt% of the vanadium was extracted through bioleaching under optimal conditions. Comparatively, the chemical leaching yield(H_2SO_4, pH 2.0) showed a slower and milder increase in vanadium yield. The vanadium bioleaching yield was 35.11 wt% greater than the chemical leaching yield. The Community Bureau of Reference(BCR) sequential extraction results revealed that 88.62 wt% of vanadium existed in the residual fraction. The bacteria substantially changed the distribution of the vanadium speciation during the leaching process, and the residual fraction decreased to 48.44 wt%. The X-ray diffraction(XRD) and Fourier transform infrared(FTIR) results provided evidence that the crystal lattice structure of muscovite was destroyed by the bacteria.

Comparison of the mechanisms of microwave roasting and conventional roasting and of their effects on vanadium extraction from stone coal

Experiments comparing microwave blank roasting and conventional blank roasting for typical vanadium-bearing stone coal from Hubei Province in central China, in which vanadium is present in muscovite, were conducted to investigate the effects of roasting tempera- ture, roasting time, H2SO4 concentration, and leaching time on vanadium extraction. The results show that the vanadium leaching efficiency is 84% when the sample is roasted at 800℃ for 30 min by microwave irradiation and the H2SO4 concentration, liquid/solid ratio, leaching temperature, and leaching time are set as 20vo1%, 1.5：1 mL.g-1, 95℃, and 8 h, respectively. However, the vanadium leaching efficiency achieved for the sample subjected to conventional roasting at 900℃ for 60 min is just 71% under the same leaching conditions. Scanning electron microscopy （SEM） analysis shows that the microwave roasted samples contain more cracks and that the particles are more porous compared to the conventionally roasted samples. According to the results of X-ray diffraction （XRD） and Fourier-transform infrared （FTIR） analyses, neither of these roasting methods could completely destroy the mica lattice structure under the experimental conditions; however, both methods deformed the muscovite structure and facilitated the leaching process. Comparing with conventional roasting, microwave roasting causes a greater deformation of the mineral structure at a lower temperature for a shorter roasting time.

Calcified roasting-acid leaching process of vanadium from low-grade vanadium-containing stone coal

The low-grade vanadium-containing stone coal used in this experiment was collected from Wuxi Country, Chongqing City, China. The experiment focused on the vanadium recovery from roasted residue through optimizing the process conditions of an effective and environmentally-friendly technology, named calcified roasting-sulfuric acid leaching technology. By single-factor experiments and orthogonal experiments, the effects of roasting temperature, roasting time, sulfuric acid concentration and leaching time on the leaching ratio of vanadium were analyzed. The results showed that the leaching ratio of vanadium reached 85.5% under the proper technological conditions of roasting temperature=950℃, roasting time=4 h, 40% concentration of sulfuric acid and leaching time=6 h.

Extracting vanadium from stone-coal by oxygen pressure acid leaching and solvent extraction

Vanadium extraction from stone-coal was investigated by oxygen pressure acid leaching and solvent extraction.The mineralogy of the stone-coal from Tongren City of Guizhou Province,China,was investigated by various determination methods. The effects of leaching time,leaching temperature,leaching agent concentration,leaching L/S ratio,granularity of material,additive consumption were investigated based on the mineralogy.The results show that under the conditions of leaching time of 3-4 h, temperature of 150℃,sulfuric acid consumption of 25%?30%,ratio of liquid to solid of 1.2:1,the granularity less than 0.074 mm, additive consumption of 3%-5%,and oxygen pressure of 1.2 MPa,and the vanadium leaching rate can be more than 92%by the method of two-step pressurized acid leaching.The powdery V2O5 product with 99.52%in V2O5 content is obtained by the flowsheet of acid recovery,removing iron by reduction process,solvent extraction,precipitating vanadium with ammonium water,and pyrolysis from the stone-coal oxygen pressure acid-leaching solution.The total recovery efficiency of vanadium is above 85%,which is more than 20%higher than that obtained in the conventional process.Furthermore,the new process does not cause air pollution since no HCl or Cl2 is released by calcination of the raw material.

N235从高碳石煤脱碳渣酸浸液中萃取钒的工艺研究

以某地高碳石煤脱碳渣经空白焙烧-酸浸后的浸出液为原料,采用萃取剂N235和反萃剂Na_(2)CO_(3)对钒萃取工艺和反萃取工艺进行了研究。结果表明,最佳萃取工艺参数为:萃原液pH值2.3、有机相组成为15%N235+5%仲辛醇+80%磺化煤油、相比(O/A)1∶4、萃取时间4 min、氧化剂氯酸钠用量为萃原液V_(2)O_(5)质量的0.05倍、萃取级数五级,钒萃取率达99%;最佳反萃工艺参数为:反萃剂Na_(2)CO_(3)溶液浓度8%、反萃相比(O/A)4∶1、反萃时间10 min、反萃级数三级,钒反萃取率可达99.50%。在最佳萃取和反萃取工艺条件下,钒浓度由4.25 g/L富集到68.50 g/L,且pH值为8.5,无需中和处理。对富钒液进一步沉钒、煅烧得到的最终产品V_(2)O_(5)品位为98.95%,且杂质含量低,满足行业标准YB/T 5304—2017中98级V_(2)O_(5)的要求。

A review on vanadium extraction techniques from major vanadium-containing resources

The critical role of vanadium in metallurgy and the increasing commercialization of vanadium redox flow batteries have contributed to a rise in market demand for vanadium,emphasizing the need to ensure the sustainability of vanadium production.Converter vanadium slag and stone coal,generated during the smelting process of vanadium–titanium magnetite,serve as primary raw materials for vanadium extraction.This paper reviews the process mineralogy of converter vanadium slag and stone coal,encompassing the chemical and physical phase compositions of the raw materials and the state of vanadium occurrence.The aim is to establish a theoretical foundation for the mechanistic study of the vanadium extraction techniques and to provide a reference for the optimization of the process afterward.This paper systematically summarizes the current vanadium extraction technology from converter vanadium slag,covering roasting,direct leaching,sub-molten salt method,and the technology and research status of vanadium extraction from stone coal.Most importantly,it identifies key issues in the current vanadium extraction techniques and challenges faced in actual production,based on the analysis of the current vanadium extraction process and mechanism research.The forward feasibility and prospects for the development of vanadium extraction technology are outlined.

陕西某地石煤钒矿制备五氧化二钒产品工艺研究

采用硫酸熟化-水浸工艺从石煤钒矿中提钒,考察了熟化条件及浸出条件对钒浸出率的影响。结果表明,适宜的熟化工艺条件为:熟化温度130℃、熟化时间8 h、H_(2)SO_(4)用量25%。对熟化料进行水浸,在浸出温度90℃、浸出时间120 min、液固比2∶1条件下,钒浸出率为90.79%。对含钒浸出液用NaClO_(3)进行氧化,采用D202吸附氧化后液中的钒,再对树脂中钒进行解吸,解吸液经沉钒煅烧后五氧化二钒产品纯度达99.21%,钒总回收率为85.99%。

石煤钒矿低温硫酸熟化-水浸提钒工艺

这是一篇冶金工程领域的论文。以陕西某地含钒矿物为绢云母、伊利石的石煤钒矿为研究对象,采用低温硫酸熟化-水浸工艺提钒,考查了熟化温度、熟化时间、硫酸用量、氯化钠用量,浸出温度、浸出时间及浸出液固比对钒浸出率的影响。结果表明:向石煤钒矿中加入25%的硫酸、0.8%的氯化钠,在130℃下熟化8 h,熟化样品在浸出温度55℃、液固比1.5∶1的条件下浸出时间2 h,钒浸出率可达89%。这说明采用低温(130℃)硫酸熟化-水浸工艺提取石煤型钒矿是可行的。

石煤提钒尾渣基钙长石烧结砖的烧结及重金属固化机理

这是一篇陶瓷及复合材料领域的论文。石煤提钒尾渣是一种危险的工业废弃物,综合利用石煤提钒尾渣具有重要的环保与经济意义。本研究以石煤提钒尾渣为主要原料制备了建筑用烧结砖,研究了烧成温度对样品结构和性能的影响规律,探究了Cu、V和Zn重金属浸出活性的钝化机理。结果表明:烧成温度显著影响烧结砖的相组成、显微结构和性能。适当提高烧成温度不仅有助于石英与原料中的碱性物质反应,生成富硅液相,提高样品的致密度,而且促进了黄长石向钙长石转变及莫来石生成。这些晶相与玻璃相互胶结,赋予制品较高的致密度和抗压强度。此外,提高烧成温度增大了制品中Cu、V和Zn的残渣态含量,减小了制品的孔径与孔容积,从而实现了Cu、V和Zn的固化。采用烧成温度为1 200℃,保温时间为3 h的工艺制度,制备出的建筑用烧结砖的各项性能均满足《烧结普通砖》(GB/T 5101—2003)中Mu25等级的要求,且制品的Cu、V、Zn重金属浸出浓度均低于标准值。研究结果为石煤提钒尾渣的综合利用提供了新的思路。

添加石煤提钒尾渣对建筑用烧结砖性能的影响

本研究以石煤提钒尾渣为原料制备了建筑用烧结砖,研究了石煤提钒尾渣细度及添加量对制品抗压强度及吸水率的影响,采用XRD和SEM表征手段分析了石煤提钒尾渣添加量调控烧结砖相组成与显微结构的机理,检测了制品的抗冻融性与使用安全性。结果表明:提高尾渣细度改善了样品的致密度与抗压强度,引入尾渣使样品的抗压强度先增大后减小。B3样品(尾渣:黏土:粉煤灰=30:30:40)经1150 o C烧成后,其抗压强度≥35 MPa,吸水率小于≤13%,综合性能满足《烧结普通砖》(GB/T 5101-2003)中Mu30的要求。XRD与SEM分析表明,添加15%~30%尾渣有利于生成液相,促进钙长石与钙黄长石的生成,晶粒与液相相互胶结,使制品具有较高的抗压强度与致密度及较好的抗冻融性能。

钙盐添加剂焙烧对石煤提钒的影响研究

文章以湖南某地含钒石煤矿样作为研究对象,采用钙化焙烧工艺,经过水浸酸浸联合浸出,通过优化添加剂的组成比例、焙烧温度和时间等参数,结果表明:钙质添加剂为CaO和CaCl_(2),添加配比分别为6%和4%,在焙烧温度850℃的条件下焙烧90 min,浸出时间2.5 h的条件下,钒的浸出效果达到最佳,为91.1%。实现了钙盐添加剂对钒浸出率的显著提升。

从石煤酸浸液中萃取钒的工艺研究

研究了三正辛胺从石煤酸浸液中萃取钒的工艺过程 ,从萃取和反萃的 p H值、相比、有机相组成、澄清时间等方面进行了详细试验。研究表明 :用三正辛胺萃取钒时 ,其萃取率可达98%以上 ;而且易反萃 ,用 0 .5 M Na2 CO3反萃时 ,反萃率为 99.9%。经萃取后 ,浸出液中的钒可由每升几克富集到每升数十克以上 ,有利于后续的提钒工艺。

石煤氧压酸浸液萃钒除铁工艺研究

研究了从石煤氧压酸浸液中萃钒除铁的工艺过程,从萃取和反萃的相比、试剂组成、pH值、澄清时间等方面进行了详细试验。研究表明:浸出液经中和还原处理后,采用10%P204+5%TBP+85%煤油萃取钒时,经六级逆流萃取后萃取率为95%以上;负载有机相用15%H2SO4反萃时,经五级逆流反萃后反萃率可达99%以上。经萃取后,浸出液中的钒可富集到37.gL-1以上,铁可缩减至以0.6.gL-1以下,反萃水相中钒铁质量比大于60,钒铁分离效果较好。

我国石煤提钒研究现状及发展

综述了近年来我国在石煤选矿富钒、提钒主体工艺(包括焙烧、浸出、浸出液处理等)以及提钒尾渣和废水综合利用等方面开展的研究工作和取得的最新进展,同时展望了今后石煤提钒的发展。

采用HBL101萃取石煤高酸浸出液中的钒

针对现行石煤提钒萃取工艺及研究现状，提出采用新型萃取剂HBLl01从石煤高酸浸出料液中直接萃取钒的方案，考察料液的酸度、料液电位、萃取时间、相比以及温度对萃取率的影响，绘制HBLl010萃取等温曲线。结果表明：在料液酸度为1．458mol／L，萃取温度为35-45℃，萃取时间为10min，相比O／A=1／1（油相与水相体积比）的条件下，钒的单级萃取率达到95％以上。三级逆流萃取实验结果显示，钒的萃取率达到99．7％以上。采用NaOH对负载有机相进行反萃，反萃液经调节pH后直接加入NH4C1沉钒，得到的五氧化二钒纯度达到98．68％以上。

含钒石煤氧压酸浸中影响钒浸出率的主要因素研究

系统研究了石煤提钒中影响钒浸出率的主要因素。试验研究表明:浸出时间3~4 h、浸出温度150~180℃、硫酸用量25%~30%、液固比1.2∶1~1.5∶1、原料粒度150~250目是最为经济可行的工艺参数范围。钒浸出率最高可达91%。

石煤氧压酸浸提钒探索试验研究

探索了在常温常压下和在压力场下,贵州某地石煤中钒的浸出率分别在9.04%~17.50%和21.03%~38.43%之间,并在石煤氧压酸浸和石煤焙烧-酸浸两种条件下,设计了正交试验,且对浸出率最高的试验进行了验证。试验结果表明:氧压酸浸的浸出率最高可达77.30%;而焙烧-酸浸的浸出率最高为50.58%,说明贵州某地含钒石煤不适合焙烧-酸浸工艺流程。