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选择性催化还原反应(SCR)的废催化剂中含有钒、钨和钛等有价值的金属,为最大限度回收其中的钒和钨资源,采用草酸分级浸取废SCR催化剂中的钒和钨,系统考察了草酸浓度、浸取温度、液固比和浸取时间对钒和钨浸出率的影响,并通过X射线衍射(XRD)和X射线光电子能谱(XPS)等表征手段分析浸出机理。废SCR催化剂直接用草酸浸取,钒浸出率为76.95%,钨和钛浸出率仅为5.31%和0.22%,钒被浸出。滤渣经焙烧后用草酸浸取,钨浸出率为56.70%,钒和钛浸出率仅为16.36%和0.12%,钨被浸出,最后滤渣主要成分为锐钛矿型TiO2。该流程实现了废SCR催化剂的分级浸取,钒和钨在不同步骤中浸出,避免了钒和钨分离困难的问题,简化了后续处理工作,降低了二次污染。 相似文献
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废加氢催化剂中含有大量的有机污染物和金属元素,如钼、钒、镍和铝,若处理不当,会造成严重的生态污染和资源浪费。本研究采用空气焙烧-碳酸钠浸出法处理废弃加氢催化剂以回收其中的钼和钒。通过热力学计算可知低温碳酸钠浸出可以实现废催化剂中钼、钒与铝、镍的分离。单因素实验考察了空气焙烧温度、碳酸钠浓度、反应时间、浸出温度、液固比等工艺条件对钼和钒浸出率的影响。实验结果表明,在焙烧温度500℃,碳酸钠浓度4 mol/L,浸出温度80℃,反应时间90 min,液固比为20:1的条件下,钼和钒的浸出率可分别达到98.02%和94.36%。为了最大限度地回收钼和钒,采用二段逆流浸出流程处理废加氢催化剂,可将钼和钒的浸出率维持在98%和97%。浸出渣中主要含有Al2O3, NiO和NiAl26O40,而绝大部分钼和钒被转移至浸出液中。 相似文献
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以Na_2CO_3为钠化剂,对高炉富硼渣采用低温钠化焙烧—水浸方法制取硼砂,考察了焙烧温度、焙烧时间、Na_2CO_3加入量、高炉富硼渣粒度、浸出温度、浸出时间、液固比等对硼浸出率的影响。高炉富硼渣中主要组分为镁橄榄石(Mg_2SiO_4),硼元素主要以玻璃态存在。试验结果表明,低温钠化焙烧过程和水浸过程对硼浸出率有显著影响,这是因为钠化焙烧使硼转化成了可溶性的硼酸钠盐,有利于硼的浸出。试验获得的最佳工艺参数如下:高炉富硼渣颗粒200目通过率为98.56%、Na_2CO_3加入量为理论量的4倍、焙烧温度为700℃、焙烧时间为4h、浸出温度为95℃、水浸时间为2h、液固比为10∶1;在此条件下,硼的一次常压水浸浸出率为71.81%,水浸滤液经除杂、蒸发浓缩后获得了结晶良好的硼砂产品,纯度为96.3%。 相似文献
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通过对铝系钒铁炉渣碳酸钠焙烧-水浸全过程的矿物分析、热力学计算及对比实验,研究了炉渣中钒、铝同步转化、溶出的机理与规律. 结果显示,焙烧进程中渣中镁铝尖晶石MgO×Al2O3相、CaO×2Al2O3相逐渐消失,MgO相生成,并生成碱熔相Na2O×Al2O3和钒酸盐. 随焙烧温度及时间增加,Na2O×Al2O3和钒酸盐相明显增多,钒、铝溶出率增加. 焙烧熟料经水浸后,液相呈碱性,钒、铝分别以可溶性钒酸钠和铝酸钠的形式进入水相,固相残留物为少量未反应的镁铝尖晶石及新生成的MgO和Ca(OH)2. 在磨矿粒度<75 mm、配碱系数1.0、焙烧温度1000℃及焙烧时间4 h的优化条件下,钒的溶出率可达90%,铝的溶出率可达75%. 相似文献
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转炉钒渣焙烧提钒技术效率低,过程涉及化学反应、传递及相变过程,蕴含物相分形生长的动力学行为。对钒渣分形变化规律的研究有助于促进钒的定向转化,进而对工业提钒具有指导意义。根据金相电镜图,使用“周长-面积法”对不同焙烧条件下钒渣粉体分形维数进行计算,得到分形维数变化与物相转化的规律。结果表明,焙烧前硅相、钒相紧密包裹,分形维数数值为1.60~2.00;加入碳酸钠焙烧后尖晶石破坏,钒相逐渐分离,使分形维数小于1.20;随着钠盐加入量的增加,物相分形维数逐渐下降;二次焙烧后,稳定的钒酸钠生成,体系趋于稳定,使得分形维数进一步下降为1.10~1.20。 相似文献
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钒渣钠化焙烧过程低价钒氧化不充分,不能被浸出,降低了钒渣的浸出率。实验采用蒽醌磺酸钠(ADA)和栲胶作载氧体实现氧的传递,强化低价钒的氧化行为。通过X射线衍射、扫描电镜、紫外光谱以及紫外可见漫反射光谱等检测方法,分析了转炉钒渣浸出反应前后物相变化行为,探索了反应过程机理,证实了其可行性。结果表明,采用ADA和栲胶作载氧体,能将转炉钒渣中的低价钒氧化成可溶的高价钒,实现空气催化氧化高效浸钒。此时,钒浸出率由89.47%分别提高到92.84%和93.64%,且催化剂对体系后续工艺没有不良影响,转炉钒渣中的尾渣含钒量由1.1%分别降至0.52%和0.47%。 相似文献
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采用高钙钒比钒渣[钙钒比ω(CaO)/ω(V2O5) 0.32]在适宜钠化焙烧条件下的熟料,通过单因素控制法,进行水浸出和碳酸铵浸出实验. 对比钒渣熟料两种浸出的适宜条件和浸出效果,分析其特点. 对浸出前后的钒渣进行物相分析,考察和对比两种浸出的浸出机理. 结果表明,钒渣熟料水浸适宜条件为,温度90℃,时间30 min,液固比8.0 mL/g. 此条件下的钒浸出率为89.4%;钒渣熟料碳酸铵浸出适宜条件为,温度60℃,时间20 min,碳酸铵含量12%. 此条件下钒的浸出率为90.2%;与熟料水浸相比,碳酸铵浸出钒的浸出率提高0.8%,浸出温度下降30℃,浸出时间缩短10 min;熟料水浸时只有水溶性钒酸盐被浸出,而碳酸铵浸出时水溶性钒酸盐和部分水不溶性钒酸盐都被浸出. 相似文献
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This study presents a synergistic extraction and recovery of vanadium and tungsten by mixed extractants (LIX984 and N263) from an acid leaching solution which was obtained from spent denitrification catalysts by chlorination heat treatment and acid leaching. Through thrice counter-flow extraction, 99.5% V and 99.7% W were extracted, while 3.0% Fe, 2.9% Ti, 2.1% P, and 3.3% Mg were co-extracted at a LIX984:N263 volume content of 22.0%, the phase ratio of 2.5 and the mother liquor's pH of 2.5. Then, at 0.80 mol/l NaOH and a phase ratio of 1, 99.9% tungsten and 99.1% vanadium were stripped from the organic phase to the aqueous phase. Subsequently, the aqueous phase's tungsten of 99.3% and vanadium of 98.1% were separated as calcium tungstate and ammonium metavanadate, respectively. In contrast, the residual solutions containing tungsten and vanadium can be returned to the following purification separation process to recover the valuable metals from the solution. Roasting converts the precipitated calcium tungstate and ammonium metavanadate to V2O5 and WO3 products. In addition, the thermodynamic analysis found the separation and recovery of tungsten and vanadium from the acid leach solution with LIX984:N263 to be an exothermic process. This method can be effectively extended for the separation of vanadium and tungsten from spent denitrification catalysts by the proposed process and validates the conclusion that metals with similar properties can be extracted using a mixture of extractants. 相似文献
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随着选择性催化剂还原法(SCR)脱硝技术在国内的普及以及SCR脱硝工程的大量建设,废弃SCR脱硝催化剂的高效处置已引起了广泛关注,针对高附加值成分的元素回收是较为合理的处置方案。本文总结了目前关于废弃SCR脱硝催化剂3种主要元素Ti、V、W(或Mo)回收的主要技术方案,其中Ti元素的回收主要是通过钠化焙烧法或浓碱浸出法首先分离Ti元素,而后通过酸洗法回收获得二氧化钛;V元素的回收方法主要包括铵盐沉淀法、萃取法和电解法,从而得到五氧化二钒或者偏钒酸铵;W元素的回收方法主要包括钙盐沉淀法、钠盐结晶法和酸沉法,从而得到三氧化钨。在此基础上,对各技术方案进行了比较,为开发高效合理的元素回收技术提供依据,并指出后续研究中还需要优化酸洗法回收Ti元素的酸洗条件以及V、W元素的纯化技术,从而进一步提高回收产品的纯度。 相似文献
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Compared with traditional sodium or calcification roasting process for vanadium extraction from raw vanadium slag (V-slag), ammonium sulfate (AS) roasting could reduce about 470℃ roasting temperature and avoid Cl2, HCl, sodium-containing waste-water and waste gypsum discharging. To reduce the amount of AS added in vanadium extraction process, an efficient AS two-stage cyclic roasting and acid leaching process was proposed. The result of TG analysis indicates V-slag could be decomposed in 275-380℃ using AS roasting process. Using 2.03:1 total mass ratio of AS to V-slag, 90.86% V and 80.54% Ti could be extracted after 380℃ roasting for 30 min and 8% initial concentration of H2SO4 leaching at 70℃ for 100 min. XRD analysis indicates V-containing spinel phase in the 1st stage leaching residue would be efficiently decomposed by the cyclic two-stage roasting and leaching process. Furthermore, the valence of V(III) in raw V-slag was not changed after the 1st AS roasting stage, but a part of V(III) in the 1st leaching residue was oxidized to V(V) after 2nd roasting process. 相似文献
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A novel method for extracting vanadium by low temperature sodium roasting from converter vanadium slag 下载免费PDF全文
Rongrui Deng Hao Xiao Zhaoming Xie Zuohua Liu Qiang Yu Geng Chen Changyuan Tao 《中国化学工程学报》1982,28(8):2208-2213
Long-term high temperature in conventional vanadium extraction process would cause particles to be sintered and wrapped, thus reducing extraction efficiency of vanadium. Based on the purpose of directional conversion and process intensification, this work proposed a combination of low temperature sodium roasting and high efficiency selective oxidation leaching in vanadium extraction. The investigation of the reaction mechanism suggested that the structure of vanadium slag was changed by roasting, which also caused the fracture of spinel. The addition of MnO2 promoted the directional oxidation of low-valent vanadium into high valence. It also found that Na2S2O8 could oxidize low-valent vanadium effectively in leaching. The leaching efficiency of vanadium reached 87.74% under the optimum conditions, including a roasting temperature of 650 ℃, a roasting time of 2.0 h, a molar ratio of sodium-to-vanadium of 0.6, a MnO2 (roasting additive) dosage of 5 wt% and a Na2S2O8 (leaching oxidant) dosage of 5 wt%. This percentage is 7.18% higher than that of direct roasting-andleaching under the same conditions. 相似文献