Application of P507 and isooctanol extraction system in recovery of scandium from simulated red mud leach solution

The current recovery technique of Sc was complicated and the chemical consumption was high. This was due to the low content of Sc in resources and the difficulty of stripping. In this research, the isooctanol was added into the 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) extraction system to reduce the extraction and improve the stripping of Sc. The maximum stripping ratio of Sc from loaded organic phase by sulfuric acid can increase from 10% (without isooctanol) to 99% (with 15 vol% isooctanol). In the extraction test of the simulated red mud leaching liquor, the separation factors between Sc and Zr, Sc and Ti are 36 and 350, separately. At the same time, other metals are almost not extracted. The high selectivity and stripping of Sc suggest that the P507 with isooctanol extraction system can be applied in the practical Sc recovery process.     

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H~+concentration 2.0 mol/L, leaching time 60 min, liquid-solid ratio 4:1 and room temperature 25 ℃(only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 ℃). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204,15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y~(3+) and 42.60% Fe~(3+) are extracted from sulfuric acid leaching liquor(adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al~(3+), Ca~(2+), Mg~(2+)and Mn~(2+) are almost not extracted. The Y~(3+) in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe~(3+) is stripped. The Y~(3+) of loaded strip liquor can be precipitated by oxalic acid to further separate Y~(3+) and Fe~(3+). The precipitation efficiency of Y~(3+) in loaded strip liquor can be 98.56% while Fe~(3+) is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 ℃ for 3 h to obtain the oxide of Y in which the purity of Y_2 O_3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%.Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting,the yttrium yield is 82.04%.     

山西拜耳法赤泥脱铝提取氧化钪的研究

研究了拜耳法赤泥脱铝及提取氧化钪的工艺。赤泥配料∶碱比=1.65,钙比=2.4,烧结温度为1080±20℃,烧结时间为40 min。熟料经稀碱液溶出、洗涤后,赤泥中的氧化铝含量降低了78.73%。在60℃,L/S=5条件下用6.5 mol/L盐酸浸出1.5 h,用2%P507+磺化煤油萃取钪,钪萃取率达90.60%,分相快。经2 mol/L NaOH溶液反萃钪,草酸沉积,800℃焙烧1h后,得到纯度为95.20%的氧化钪。     

含钪料液中杂质的去除

含钪赤泥盐酸浸出液中杂质含量很高,在提取钪之前必须去除。研究了含钪赤泥盐酸浸出液中杂质Fe、Al、Ti、Zr的去除：首先在溶液中加入3%H_2O_2—3．0mol/L H_2SO_4溶液,将钛转化为不被P507萃取的过氧钛酸,然后用P507萃取钪,用6mol/L HCl和蒸馏水洗涤有机相,除去被同时萃取的Fe和Al,再用1%HF溶液洗涤除去Zr,最终使钪与杂质元素分离。最后得到的产品中Sc_2O_3质量分数为96．25%。     

钛白废酸与赤泥联合提取氧化钪的工艺研究

以钛白废酸、赤泥为原料,经过浸出、一次萃取反萃、酸溶水解、二次萃取、酸洗、二次反萃、草酸沉淀、精制等提纯工序,可得到纯度99.99%的Sc2O3。赤泥和钛白废酸中钪的回收率分别达到57.8%和93.3%。     

盐酸浸出赤泥回收铝的研究

研究了盐酸浸出拜耳法赤泥中铝的过程。考察浸出温度、浸出时间、液固比和酸度对铝浸出率的影响,并进行了赤泥回收铝的工业化试验。结果表明,影响铝浸出率的主次因素依次是浸出温度、液固比、盐酸浓度和浸出时间。赤泥在80℃下进行两段浸出,再经蒸发、除钛、除钙、中和及氢氧化钠溶液溶出,铝浸出率为88.13%,回收率为80.26%。     

Separation and recovery of scandium and titanium from red mud leaching liquor through a neutralization precipitation-acid leaching approach

Selective recovery of scandium and titanium from red mud leaching solution was achieved through neutralization precipitation followed by acid leaching approach.In the neutralization precipitation section,the effects of pH,temperature and reaction time on metal precipitation efficiency were investigated.Under the optimal co ndition,the precipitation efficiencies of scandium and titanium were 93.74% and 99.47%,respectively.In the acid leaching section,the effects of acidity,temperature,reaction time,and raffinate to acid solution ratio on leaching efficiency were investigated.Under the optimal condition,the leaching efficiencies of scandium and titanium were 99.97% and 5.44%,respectively.The loss of scandium and titanium were 6.3% and 5.9%,respectively.Compared with the traditional extraction procedure of scandium in red mud,this method could separate titanium from scandium effectively,which is beneficial for the purification of the products and improvement of value of the metal.     

盐酸浸出氧化铝赤泥回收镓

研究了拜耳法赤泥盐酸浸出镓的过程。采用正交试验考察浸出温度、时间、液固比和酸度对镓浸出率的影响。结果表明,在最佳浸出条件下:8mol/L盐酸、液固比4.0、109℃浸出5h,镓浸出率达到95.4%。用50%TBP+50%煤油一次萃取,镓萃取率达到98%。用0.5%食盐水反萃,镓反萃率为96.8%。反萃液用0.5mol/L NaOH溶液中和、过滤、烘干后,固体中镓的质量百分数为4.32%,从赤泥中富集了136倍。镓的总回收率达到85%以上。     

从氧化钴矿石中提取钴的试验研究

研究了从氧化钴矿石中回收钴.通过两段浸出,浸出渣中钴质量分数小于0.5%,钴浸出率达99%.通过黄钾铁钒法除铁,氟化钠法除钙、镁,亚硫酸钠法除铜,P204串级萃取法进一步去除杂质Fe、Ca、Mg、Cu、Zn、Mn、Pb、As等,P507萃取分离钴镍,最后通过沉淀得草酸钴产品,产品纯度符合要求.     

Leaching behavior of metals from a limonitic nickel laterite using a sulfation–roasting–leaching process

The leaching behavior of metals from a limonitic laterite was investigated using a sulfation–roasting–leaching process for the recovery of nickel and cobalt. The ore was mixed with water and concentrated sulfuric acid followed by roasting and finally leaching with water. Various parameters were studied including the amount of acid added, roasting temperature and time, sample particle size, addition of Na₂SO₄ and solid/liquid ratio in leaching process. More than 88% Ni, 93% Co and < 4% Fe are extracted under the determined conditions. Simultaneously, about 90% Mn and Cu, 70% Mg, 45% Al, 25% Zn, 4% Cr and Ca are extracted respectively. The pH of the leach solution is about 2. The leaching efficiency is independent of sample particle size due to decomposition of ferric sulfate formed during roasting. The roasted mass was characterized by various physico-chemical techniques such as DSC/TGA, XRD and SEM. This process provides a simple and effective way for the extraction of nickel and cobalt from laterite ore.     

Extraction of vanadium from black shale using pressure acid leaching

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO₄) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe³⁺ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Solvation Extraction by Amines and Synergistic Solvation Extraction with Neutral or Acidic Extractants

Abstract

Primary amine can be used to separate vanadium(V) from chromium(VI) effectively in weakly alkaline solution by solvation mechanisms, as shown by earlier work. Separation of rhenium(VII) from molybdenum, tungstenfVI) from molybdenun, etc. can be very effectively carried out by using primary amine mixed with neutral donor reagents by synergistic solvation extraction. Experimental results indicate the presence of solvation extraction in addition to the ordinary anion exchange extraction with amines as solvent. Iron present in the sulfuric acid leaching solution as impurity gets extracted into the organic phase and is difficult to strip. Several mixed solvent systems with an amine and a neutral donor cxtractant have been developed for iron removal. These have common feature that the iron in the organic phase can be stripped with dilute sulfuric acid.     

Purification of scandium from concentrate generated from titanium pigments production waste

Nowadays 80%of scandium in China is obtained from titanium pigments production waste through a complex purification process.The study mainly focused on the purification of Sc from its concentrate generated from titanium pigments production waste by solvent extraction.Several extractants have been tried and 10%D₂EHPA-5%TBP-85%sulfonated kerosene exhibited the best extraction performance towards Sc in 7 mol/L H₂SO₄solution,so it was selected as the oil phase.0.5%of H₂O₂was added into the concentrated solution which can effectively inhibit the extraction of Ti.Both the extraction and back extraction parameters are optimized.The preferred extraction conditions were obtained,i.e.,acidity:7 mol/L H₂SO₄,the phase ratio A/O:10,room temperature,mixed contact time:30 min,Sc concentrate:10 g/L,that the extraction rate of Sc in the above conditions was nearly 100%.NaOH was used for back extraction with the stripping rate 99%on the following conditions:5 mol/L NaOH stripping for 30 min at a phase ratio A/O:1 at 90℃.Finally,H₂C₂O₄was used to further purify the back extraction product and Sc₂(C₂O₄)₃precipitant fo rmed.The final product Sc₂O₃with a purity over 99.5%was obtained by calcining Sc₂(C₂O₄)₃at 1000℃for 2 h.A conceptual process for Sc purification was put forward and proved.The total recovery yield of Sc in the whole process is 95%.     

溶剂萃取法从含铍溶液中分离铍

采用溶剂萃取法从某复杂低品位铍矿的硫酸浸出液中进行铍的分离,研究了不同因素对铍的单级萃取效果的影响。结果表明,最佳条件为:水相pH=2～2.5、浸出液初始铍浓度1.5～2.5g/L、P204体积分数30%、萃取时间20min、相比为1。在此条件下四级逆流萃取后铍萃取率可达到98.50%。     

萃取法分离钪与铁的研究

赤泥是铝工业生产过程中的废弃物,其中含有有价值的金属钪,通过盐酸浸出、有机溶剂萃取可实现钪的富集分离。当选用w(P204)为1%和w(TBP)为1%的萃取剂、1∶1的V(O)∶V(A)(O为有机相,A为水相)时,在3 mol/L的酸度下进行10 min振荡萃取后,钪的提取率可达到90%以上,铁的提取率只有3.6%,钪与铁的分离系数为312,为后续钪的进一步提取提供了有力的条件。     

用LIX84从富钴结壳硫酸浸出液中选择性萃取铜

采用LIX84作萃取剂、硫酸作反萃剂 ,从大洋富钴结壳常温常压活化硫酸浸出除铁后液中萃取铜。试验考察了相比、平衡水相pH值、时间等因素对LIX84萃铜的影响。结果表明 ,相比、平衡水相 pH值、混合时间都对铜的萃取率有一定影响。最后优化出的萃取工艺条件为 (体积百分数 )有机相 12 %LIX84+ 88%煤油 ,室温 ,相比 (O/A)=1/ 2 0 ,出口水相pH2 60± 0 0 5 ,萃取级数为 2级 ,每级混合时间 5min。经过 2级萃取、1级洗涤、3级反萃后 ,可以得到完全符合电解沉积要求的硫酸铜溶液 ,从而使浸出液中的铜与其它金属彻底分离     

酸浸-萃取-氨沉淀法从石煤钒矿中提取钒

采用氧化焙烧脱炭-硫酸氧化浸出-P204+TBP溶剂萃取-氨水沉钒的工艺方法,从江西某石煤钒矿中提取V2O5,考察了硫酸用量、萃取及反萃次数、反萃液pH值对工艺过程的影响。试验结果表明:钒矿破碎后在硫酸溶液中用氯酸钠进行氧化浸出,钒的浸出率可达到96%以上;用P204+TBP溶剂萃取和稀硫酸溶液反萃,再用氨水沉淀钒,最终得到纯度98.0%以上的V2O5产品;从石煤钒矿到V2O5的总收率可达86.14%～93.09%。该工艺对钒的回收效果明显,操作简单,生产成本低,对环境污染较小。     

Recovery of rare earths from spent FCC catalysts by solvent extraction using saponified 2-ethylhexyl phosphoric acid-2-ethylhexyl ester(EHEHPA)

A process to recover rare earth(RE) metals from spent fluid catalytic cracking(FCC) catalysts by solvent extraction was studied, using saponified 2-ethylhexyl phosphoric acid-2-ethylhexyl ester(EHEHPA or P507). The recovery process involved three steps:(1) leaching REs(mainly lanthanum and cerium);(2) solvent extraction by applying saponified P507-kerosene system;(3) stripping. Experiments to assure optimal operating conditions were conducted. Results indicated that RE metals could be recovered effectively from spent catalyst with saponified P507-kerosene-HCl system. At room temperature of 25 oC, 10 g spent catalyst with 110 m L of HCl(1 mol/L) could achieve a leaching efficiency of 85%. For extraction, initial pH value of 3.17, organic/aqueous ratio(O/A ratio) of 2:1 with an extractants' saponification rate of 20% could obtain 100% efficiency. In the stripping process, 1 mol/L HCl with O/A ratio of 1:1 led to a stripping efficiency of 96%. In the present study, RE metals from spent FCC catalysts were effectively recovered, which avoided wasting a large amount of RE resources. It provides a theoretical support for commercial recycling of RE resources.     

P204+P507对钴浸出液协同萃取除钙锰铜锌

针对分步萃取法萃取钴工艺流程繁杂、萃取级数较多的问题,采用P204+P507为复配萃取剂从工业硫酸钴浸出液中一次分离出Zn2+、Ca2+、Mn2+、Cu2+等。探究了平衡pH、复配萃取剂配比、萃取相比O/A、有机相皂化率等对元素萃取率的影响。结果表明：以28%P204+7%P507为复配萃取剂,65%溶剂油为稀释剂,在有机相皂化率为50%、萃取平衡pH=3.57、相比O/A=2的条件下,Zn2+、Ca2+、Mn2+、Cu2+的单级萃取率分别达到99.97%、94.65%、88.42%、87.18%,Co2+萃取率仅有17.42%。后续使用1.5 mol/L硫酸在反萃相比O/A=20、两次洗涤条件下可以将99%的钴洗涤下来。     

赤泥-钛白废酸综合回收中钪的萃取工艺

针对赤泥-钛白废酸浸出液中钪及主要杂质的特点, 采用先除杂后萃取的工艺对溶液中钪进行萃取分离.首先, 将一定量的活性炭加入赤泥-钛白废酸浸出液中, 吸附去除浸出液中的硅, 硅的去除率可达96.70%, 而钪的去除率仅为1.25%, 这表明活性炭吸附除杂可在保证浸出液中钪含量基本不损失的情况下除去绝大部分的硅.除硅有效控制了浸出液的胶凝现象, 有利于下一步的钪萃取工艺.在萃取工艺过程中, 具体研究了除杂后液的酸度、萃取剂体积分数、相比、萃取时间对钪萃取率的影响.结果表明, 除杂后液酸度以1.81 mol/L为最佳, 既避免了有机相乳化, 又保证了钪的高萃取率; 相比在1/10~1/30之间时, 钪萃取率达到平衡, 但当相比为1/30时, 发生乳化, 难于分离, 因此, 相比1/25为最佳; 萃取时间为15 min时, 钪的萃取率达到平衡; 萃取剂体积分数为15% P204+ 6% TBP时, 钪的萃取率达到最大值.在最佳的萃取工艺条件下, 钪的萃取率达98.80%.