石煤提钒的浮选工艺及吸附机理

针对传统石煤提钒工艺中钒回收率低、环境污染严重等问题,在工艺矿物学研究的基础上,采用浮选的方法对含钒石煤矿进行条件实验和闭路实验,设计出提钒浮选流程,并利用分子动力学模拟研究捕收剂分子在矿物解理面的吸附过程。结果表明:石煤中主要含钒矿物为钒云母,主要的脉石矿物为石英;通过闭路浮选实验,得到了五氧化二钒品位为3.20%(质量分数)、回收率为74.50%的钒精矿;捕收剂分子容易在钒云母(001)面发生吸附,而在石英(001)面几乎不吸附,从而实现了石煤中钒云母和石英的浮选分离。     

焙烧-浮选富集石煤中的钒（英文）

在石煤工艺矿物学的基础上,采用焙烧-浮选新工艺富集石煤中的钒。新工艺包括焙烧脱碳、选择性磨矿、脱泥和浮选4个关键步骤。在焙烧脱碳过程中,550°C焙烧脱碳能有效避免原矿中的碳质物和焙烧中方解石分解生成游离CaO对浮选产生的不利影响。通过选择性磨矿,高耗酸性矿物在中间粒级富集,而云母在细粒级和粗粒级中富集。浮选得到的最终精矿V_2O_5的品位为1.07%,回收率为83.30%。此外,最终精矿的钒浸出率相比于脱碳样提高了13.53%。石煤原矿在550°C焙烧脱碳是可行的,对云母浮选几乎没有不利影响。通过对石煤中钒的浮选预富集,可以降低提钒的处理量、酸耗量和生产成本。     

低品位石煤矿中钒和碳的浮选回收（英文）

在工艺矿物学的基础上,进行高碳含钒石煤浮选工艺研究。浮选工艺包括选碳流程和选钒流程,可以分别对碳和钒矿物进行富集。在碳浮选流程中,通过再磨工艺可以将碳和钒矿物有效分离。结果发现,浮选精矿中,钒的品位和回收率分别为1.32%和88.38%,尾矿产率为38.36%。碳精矿品位为30.08%,回收率为75.10%,可以直接作为燃料使用。对浮选产品进行酸法浸出实验,浮选产品的浸出率都在85%以上,说明该浮选工艺对浸出影响比较小。该浮选工艺可以减小钒浸出的处理量,降低冶炼成本,对石煤浮选工艺有借鉴作用。     

石煤提钒工艺研究攀钢获技术突破

钒是冶金工业所需的重要资源，主要共生于钒钛磁铁矿和石煤矿中。石煤是一种含碳少、发热值低的劣质煤，又是一种低品位多金属共生矿，其中包括低品位的钒。含钒石煤遍布我国湘、鄂、川、黔、浙、桂、赣、皖、陕、晋、豫、甘等20余省。据中科院过程所的调查，我国石煤中的钒总储量占全国钒总储量的87％，比世界除中国外的钒总储量还多。仅浙江至广西一条长约1600余公里的石煤矿，就蕴含着1亿吨以上的五氧化二钒。但是绝大多数石煤含钒品位很低（V2O5含量〈0．8％），开发利用极其困难，即使含钒大于0．8％的石煤提钒也十分困难。     

磷酸盐对一水硬铝石和高岭石浮选的选择性作用

研究了磷酸盐对一水硬铝石和高岭石的浮选性能。六偏磷酸钠(SHMP)在低用量条件下，对高岭石的抑制作用大于对一水硬铝石的，可以实现铝土矿的正浮选脱硅；而在高用量的条件下，六偏磷酸盐对一水硬铝石的抑制作用很强，对高岭石的抑制作用与低用量条件相比没有明显变化，因而可以实现铝土矿的反浮选脱硅。通过磷酸盐的溶液化学、六偏磷酸钠对一水硬铝石和高岭石矿物的ξ-电位影响及其在矿物表面上的吸附量测定，同时通过显微镜对溶液中的矿物颗粒成像，探讨了(NaPO3)6在一水硬铝石和高岭石的胺阳离子反浮选体系中的作用机理。     

石煤提钒预富集分离过程中的松散-分层模型(英文)

在扫描电镜矿物定量评价(QEMSCAN)工艺矿物学研究的基础上,采用摇床分选工艺对石煤中的钒进行预富集,以实现提高钒品位的同时,降低耗酸矿物的含量;并利用Bagnold剪切理论和Kelly分层假说建立分离过程中的松散-分层模型,以阐述含钒基体矿物从分选体系中分离的机理,且对模型进行验证。模型结果表明,细粒级含钒基体矿物和粗粒级含钒基体矿物分别在粗选和扫选阶段实现分离;在模型计算过程中确定了影响粗选和扫选阶段分离的主导因素,分别为重力沉降和摇床床面的不对称往复运动产生的剪切分散压。     

粗粒辉钼矿工艺矿物学及流态化浮选抛废研究

粗粒预选抛废是实现选厂节能降耗的重要技术手段，借助流态化浮选技术对河南某选厂低品位辉钼矿(粒级-1 mm,Mo品位0.076%)进行选前抛废探索。通过工艺矿物学对辉钼矿嵌布和解离特征进行了全面分析，研究了操作因素对流态化浮选的影响规律，并在不同捕收剂用量下比较了其与机械式浮选的抛废效果差异。工艺矿物学结果表明，试验所用矿石中主要有价矿物辉钼矿为中粒-细粒不均匀嵌布，粗粒级矿石中辉钼矿解离度较低。流态化浮选结果表明，操作因素对于尾矿品位和抛尾率均可能产生影响，最优条件下流态化浮选尾矿品位为0.011%～0.015%，抛尾率为40%～55%，抛废效果远优于机械式浮选，其原因是流态化浮选构建了低紊流环境，提升了浮选粒度上限。     

某含银高硫铜矿的综合回收试验

某含银高硫铜矿含铜0.76%、硫24.35%及银34.92 g/t,有价矿物种类多、矿石性质复杂,采用抑硫优先浮选铜-活化浮选硫的原则工艺流程进行试验,配合石灰作为硫化铁矿物抑制剂以及筛选出丁基黄药+酯-105作为硫化铜矿物的组合捕收剂,强化了银在铜精矿中的富集。在选定工艺条件下,可获得铜品位21.60%、银品位602.84 g/t的铜精矿(铜和银回收率分别为89.30%和54.39%),硫品位45.60%、银21.55 g/t的硫精矿(硫和银回收率分别为89.79%和29.59%),实现了铜、硫和银的综合回收利用。     

微细粒嵌布独立银矿石的活性炭吸附浮选

以鲁甸微细粒独立银矿石为研究对象,采用岩矿鉴定、化学分析及X射线衍射分析(XRD)对原矿的矿物学性质进行研究。并采用活性炭吸附浮选法对银矿物进行回收。采用扫描电子显微镜(SEM)及能谱仪(EDS)对活性炭吸附浮选的机理进行分析。结果表明:开路试验中银精矿产率由3.19%降低到1.03%,银品位由1485.37 g/t升高到4736.52 g/t。通过闭路试验可以获得银品位为4116.73 g/t,银精矿回收率为54.83%。     

阴离子-非离子组合药剂对一水硬铝石浮选和泡沫稳定性的影响

采用单矿物浮选实验,接触角实验,两相、三相泡沫动态稳定性实验,研究组合药剂油酸钠(NaOL)-辛醇(OCT)对一水硬铝石浮选行为的影响规律及其作用机理。通过测定组合药剂与一水硬铝石作用前后溶液的TOC含量,分析组合药剂作用下一水硬铝石三相泡沫稳定性的原因。结果表明:在矿浆p H=7.0条件下,组合药剂与单一NaOL浮选一水硬铝石结果相比,NaOL物质的量在组合药剂总物质的量中所占的摩尔分数x(NaOL)≥40%时有利于提高一水硬铝石的浮选回收率。其中,x(NaOL)=40%和x(NaOL)=60%的组合药剂因通过形成更稳定的三相泡沫而提高一水硬铝石的浮选回收率;x(NaOL)=80%的组合药剂不仅能增强浮选泡沫稳定性,还能增大矿物表面疏水性从而提高一水硬铝石浮选回收率。     

Vanadium recovery from stone coal through roasting and flotation

A new process for vanadium recovery from stone coal by roasting–flotation was investigated based on the mineralogy. The process comprised four key steps: decarburization, preferential grinding, desliming and flotation. In the decarburization stage, roasting at 550 °C effectively avoided the negative effect of the carbonaceous materials in raw ore and generation of free CaO from calcite decomposition during roasting. Through preferential grinding, the high acid-consuming minerals were enriched in the middle fractions, while mica was enriched in the fine and coarse fractions. Through flotation, the final concentrate can be obtained with V₂O₅ grade of 1.07% and recovery of 83.30%. Moreover, the vanadium leaching rate of the final concentrate increased 13.53% compared to that of the feed. The results reveal that the decarburization by roasting at 550 °C is feasible and has little negative impact on mica flotation, and vanadium recovery from stone coal is conducive to reducing handling quantity, acid consumption and production cost.     

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%。     

Leaching of vanadium from stone coal with sulfuric acid

The effects of roasting, mass ratio of H2SO4 to stone coal, leaching temperature, liquid-to-solid ratio, grinding fineness of stone coal, and two-stage counter-current leaching on the vanadium leaching ratio were studied. The results show that the vanadium leaching ratio of roasted stone coal through two-stage counter-current leaching can reach 65.1% at the mass ratio of H2SO4 to stone coal of 20%, leaching temperature for the production of vanadium from stone coal.     

Selective Separation and Extraction of Vanadium (V) Over Manganese (II) from Co-Leaching Solution of Roasted Stone Coal and Pyrolusite Using Solvent Extraction

Based on the novel technology for selective separation and extraction of vanadium (V) over manganese (II) from co-leaching solution of roasted stone coal and pyrolusite using solvent extraction, the extraction effect of vanadium (V) and manganese (II) has been studied and many technical conditions have also been optimized. Meanwhile, countercurrent simulation experiments were conducted to verify the results of the experiments. The results indicated that with three countercurrent extraction stages, 99.21% vanadium (V) was extracted using 5% (v/v) N235 and 5% (v/v) secondary octyl alcohol at initial aqueous pH of 3.0 and O/A phase ratio of 1.0. Vanadium (V) could be completely stripped after three-stage countercurrent experiments with 20 wt.% NH₄Cl at O/A phase ratio of 1.0. The process flow sheet for the recovery of vanadium (V), as well as manganese (II), was proposed.     

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 extendingapplication.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.     

空心聚吡咯/V_2O_5复合材料的制备及其电化学性能

为改善晶态V2O5（c-V2O5）正极材料实际容量较低、循环性能较差等问题,制备了空心球聚吡咯/V2O5复合材料。利用导电吡咯单体（Py）在中空型V2O5层间发生原位氧化聚合反应制备聚吡咯（PPy）/中空型V2O5复合材料（HS-PPy/V2O5）。采用X射线衍射仪（XRD）、扫描电镜（SEM）、透射电镜（TEM）对样品进行表征,采用恒流充放电测试和电化学阻抗（EIS）测试样品的电化学性能。结果表明,Py单体已插入中空型V2O5层间,与纯中空型V2O5相比,制备的HS-PPY/V2O5复合材料比容量虽然有所减小,但是循环稳定性有较大的提高。     

循环流态化焙烧加压浸出从极难浸石煤中提取钒

以广西某极难浸石煤钒矿为研究对象,研究循环流态化焙烧试样在加压浸出条件下的钒浸出率。结果表明:在相同酸浸条件下,循环流态化空白焙烧试样的钒浸出率高于钠化焙烧的钒浸出率。系统的焙烧浸出工艺对比研究表明:该石煤钒矿只有在循环流态化焙烧并加压高浓度酸浸作用下才能获得最高的钒浸出率,应属于极难浸石煤钒矿。在V(H2SO4):V(HF)=1:1和MnO2添加量(质量分数)为3%的条件下,循环流态化空白焙烧矿的最佳酸浸条件为液固比1:1、浸出温度150℃、浸出时间6 h,钒浸出率可达98.11%。同时,研究循环流态化空白焙烧矿加压浸出的动力学模型、浸出控制步骤及表观活化能。循环流态化空白焙烧能避免钠化焙烧产生的Cl2及HCl等有害气体的排放问题。从焙烧反应设备的创新应用着手,探索试验工艺条件,为极难浸石煤钒矿的工业化利用提供参考和依据。     

Vanadium recovery from clay vanadium mineral using an acid leaching method

A technique including direct acid leaching,vanadium precipitation with alkaline,sodium hydroxide releaching,impurity removing by adjusting pH value,precipitation vanadium with ammonium chloride,and vanadium pentoxide by roasting steps was proposed according to the characteristic of Xichuan clay vanadium mineral.The factors influencing leaching vanadium such as temperature and the concentration of sulfuric acid were investigated and optimized.The experimental results indicate that the extract ratios of V2O5 can reach 94% and 92% at a sodium chlorate ratio of 3% and a manganese dioxide ratio of 3%,respectively.A completely chemical precipitation method was adopted to decontaminate and enrich the vanadium in the acid leaching solution.The X-ray diffraction (XRD) pattern and the purity analysis of vanadium pentoxide indicate that the purity of final vanadium pentoxide can reach 99% and meet the standard specifications.The total recovery can reach about 75%.The technique has the characteristics of simplicity,less investlnent,and more environment safety as compared with the traditional salt roasting method.     

微波加热制备氮化钒工艺

以五氧化二钒或偏钒酸铵为原料,炭黑为还原剂,采用微波法研究氮化钒的制备工艺.探讨在还原时间为60 min,还原最高温度为933 K时,混合物配碳比、氮化温度、氮化时间、氮气的流量、混合物成形压力等因素对产物氮含量的影响.由一步法结果表明:在混合物成形压力为20 MPa,配碳比为35%,氮化时间为120 min,氮化温度为1723 K,氮流量为2 L/min,产物氮化钒的氮含量为12.6%,钒含量79.2%,碳含量4.6%,密度为4.5 g/cm3.经 XRD检测产物为纯氮化钒.同时与传统的电阻炉加热方式相比,微波加热缩短了反应和冷却时间,节省能耗,简化工艺,降低成本.