氟碳铈矿精矿焙烧稀土浸出新工艺研究

以四川冕宁氟碳铈矿精矿为研究对象,提出了一种减少氟碳铈矿中Ce(Ⅲ)的氧化来提高稀土浸出率的新工艺。通过控制焙烧温度和焙烧助剂NaHCO_3用量,利用焙烧助剂分解产物Na_2CO_3熔融包覆矿物以减少Ce(Ⅲ)的氧化,经水洗、酸浸过程,促使氟碳铈矿中铈进入酸浸液而减少留在酸浸渣,从而实现铈的浸出回收,提高总的稀土浸出率。得到的优化工艺条件为:NaHCO_3与氟碳铈矿质量比为40%,于900℃焙烧2 h,水洗温度50℃,水洗液固比10∶1,盐酸浓度为2 mol·L~(-1),酸浸液固比为15∶1,温度75℃酸浸2 h。在此条件下,总稀土浸出率达93.23%,铈的浸出率为87.43%。     

HCl-H_2O_2溶液浸出氟碳铈矿的研究

以混合型稀土精矿为研究对象,采用HCl-H2O2溶液浸出氟碳铈矿,达到独居石与氟碳铈矿分离的目的。考察了焙烧温度、盐酸浓度、液固比、双氧水用量、浸出温度、浸出时间等对稀土精矿浸出率的影响规律,并得出最佳浸出工艺条件为:焙烧温度:600℃,焙烧时间:2 h,盐酸浓度:6 mol·L-1,液固比:30∶1,双氧水用量:10 m L,浸出温度:90℃,浸出时间:90 min。通过此实验条件,稀土精矿中氟碳铈矿的浸出率达到了98.69%,使稀土精矿中的氟碳铈矿大部分进入溶液中,达到与独居石分离的目的,为浸出氟碳铈矿的工艺提供新思路。     

硅酸钠焙烧提取钙热还原稀土冶炼渣中稀土的研究

采用钙热还原法生产稀土金属时会产生大量的稀土冶炼渣,渣中稀土的存在形式包括稀土金属单质、稀土氧化物和稀土氟化物,其中的稀土金属和稀土氧化物易于被无机酸直接浸出,而氟化稀土则难被酸浸出,成为从钙热还原稀土冶炼渣中高效回收稀土的瓶颈。在工业生产氟化钠晶体方法的启发下,开发了九水硅酸钠焙烧-HCl浸出提取钙热还原稀土冶炼渣中稀土的方法,通过九水硅酸钠焙烧钙热还原稀土冶炼渣,将渣中稀土氟化物转型为易溶于酸的稀土氧化物,而氟元素则转型为Na F被水洗除去,实现了稀土和氟的高效分离。考察了焙烧温度、时间和HCl浓度等因素对稀土提取率的影响,结果表明:在焙烧温度为850℃,焙烧时间为2 h,九水硅酸钠与钙热还原渣质量比为1∶1,酸浸温度为60℃,酸浸时间为1. 5 h,HCl浓度为4 mol·L~(-1),液固比为11∶1的条件下,稀土提取率高达99. 06%。     

氟碳铈矿氧化焙烧-盐酸催化浸出新工艺研究

针对氟碳铈矿盐酸处理过程中稀土浸出率低,高价值非铈稀土进入富铈渣造成高价元素低值利用,以及伴生资源氟综合利用等问题,研究开发了低温焙烧-催化浸出技术,考察了焙烧温度、浸出温度、盐酸用量、液固比及添加浸出助剂等对稀土浸出率的影响。在优化工艺条件:焙烧温度500℃,浸出温度50℃,酸矿质量比1.75∶1,液固质量比2∶1时,总稀土浸出率达到65.1%,非铈稀土浸出率为93.3%,浸出渣中CeO2/TREO为94.1%。本工艺简化了工艺流程,节省了大量化工原料消耗,降低了成本,整个过程实现了无氟排放,具有低消耗、高效能等特点,取得了良好的环保、经济和社会效益。     

HCl-AlCl_3溶液络合浸出稀土氧化物和氟化钙的研究

CaO-NaCl-CaCl2焙烧包头混合型稀土精矿的产物经脱磷处理后,矿物中的物相主要为稀土氧化物和氟化钙。以较低浓度AlCl3作为助浸剂用HCl溶液浸出矿物,利用低浓度的Al3+与氟形成氟铝配合物,有效地帮助难溶物相氟化钙浸出,同时避免浸出过程中氟离子与稀土形成氟化稀土沉淀而造成损失。结果表明:盐酸浓度为3 mol.L-1,Al3+浓度为0.25 mol.L-1,温度为60℃,液固比为10∶1,浸出时间为60min,搅拌速度为200 r.min-1时,稀土氧化物的浸出率为70.8%,氟化钙的浸出率为55.8%。五级逆流浸出后,RE2O3的浸出率为99.6%,CaF2的浸出率达到98.5%。     

硅酸盐稀土渣硫酸铵焙烧浸出研究

采用高温直接还原-磁选-(NH4)2SO4焙烧-水浴浸出的工艺流程,从低品位含稀土尾矿中提取稀土。主要考察了(NH4)2SO4配量,焙烧温度和焙烧时间等实验参数对La,Ce,Nd浸出率的影响。通过SEM-EDS,XRD和TG-DSC表征了尾矿经高温直接还原后稀土的赋存状态、不同温度(NH4)2SO4焙烧富稀土渣过程中的物相变化和浸出渣及浸出液烘干后析出晶的物相成分。结果表明:稀土尾矿经高温处理后分别得到了高品位铁精粉和富稀土渣。稀土元素主要以Ca2RE8(Si O4)6O2,CaRE2(Si O4)2硅酸盐形式存在。优化实验条件下,(NH4)2SO4与富稀土按8∶1混匀,400℃焙烧45 min,80℃热水浴浸出时间1 h,浸出液液固比10∶1,La,Ce,Nd浸出率分别为96.13%,98.88%,97.10%。经(NH4)2SO4焙烧处理后,稀土元素的最终产物变为可溶性的(NH4)RE(SO4)2和RE2(SO4)3,萤石部分转变为Ca SO4。     

包头磁尾矿的铝盐混合焙烧-酸浸萃取稀土的研究

以包头白云鄂博磁尾矿为原料,利用氟-铝极强的配位能力,进行铝盐混合焙烧,使氟-铝在焙烧过程中发生固相配位作用,达到活化稀土目的。然后利用酸溶液浸出稀土和氟的实验,考察稀土与氟的浸出率,最后P204萃取稀土。结果表明:采用Al2(SO4)3为磁尾矿抑氟剂,n(F)/n(Al)=3.0时,稀土矿的活化效果最佳。包头磁尾矿与铝盐混合物于700℃焙烧1.0 h,经2.0 mol·L-1H2SO4在80℃条件下浸出3.0h,氟和稀土的浸出效果较好。其中稀土浸出率在86.57%左右,氟可全部浸出。以该条件下产生的含氟稀土溶液为原料进行萃取实验,稀土全部进入有机相,而氟存留在浸出液中,基本实现了稀土和氟的分离。     

Na_2CO_3-NaOH焙烧高品位混合型稀土精矿的研究

以包头混合型稀土精矿为原料,运用差热分析、X射线衍射分析技术及化学分析等手段,研究了Na_2CO_3-NaOH体系下焙烧温度、焙烧时间、Na_2CO_3加入量、NaOH加入量对稀土浸出率的影响。结果表明:在焙烧温度为500℃,焙烧时间为90 min,Na_2CO_3加入量为20%(质量分数),NaOH加入量为16%(质量分数)的条件下,稀土浸出率达到了99%以上。     

HCl-柠檬酸配合浸出稀土氧化物的研究

以白云鄂博稀土精矿-钙化焙烧-HCl浸磷后的矿物为原料,用HCl-柠檬酸混合溶液浸出稀土并对浸渣进行了场发射扫描电镜和能谱分析。考察了HCl-柠檬酸混合溶液浸出过程中HCl用量、柠檬酸用量、反应温度、反应时间、搅拌速度及液固比对稀土、氟浸出率的影响,并采用五因素正交回归试验对工艺参数进行了优化。结果表明:在HCl浓度3.0 mol·L~(-1),柠檬酸浓度0.3mol·L~(-1),反应温度40℃,液固比9∶1,反应时间40 min,搅拌速度300 r·min~(-1)时,稀土的浸出率大于92.0%,氟的浸出率小于5.0%,实现了稀土与氟的分离。     

直接还原熔分-酸浸出处理稀土复合铁矿实验研究

采用高温直接还原熔分工艺处理白云鄂博稀土复合铁矿,在1400℃条件下还原12和15 min,可得到含Fe 94.35%~95.20%的珠铁和含稀土9.71%~10.55%的富稀土渣。直接还原熔分过程未添加CaCO_3获得的稀土渣其稀土相以Ce_(4.667)(SiO_4)_3O析出,当配加5%的CaCO_3时,稀土富集相转变为铈磷灰石(Ca_3Ce_2[(Si,P)O_4]_3F),但两种稀土富集相均属于硅酸盐系。将获得富稀土渣磨碎至74μm并置于HCl中浸出,实验结果表明:在HCl浓度1 mol·L~(-1),浸出时间2 h,固液比为1∶14和环境温度50℃时,稀土浸出率均能达到95%以上,还原熔分过程未添加CaCO_3设定还原时间12和15 min或还原熔分过程加入5%CaCO_3还原时间12 min,富稀土渣中Th的浸出率从21.92%分别增加到31.43%和51.58%。     

Leaching of rubidium from biotite ore by chlorination roasting and ultrasonic enhancement

Aiming at the problems of the low grade of rubidium (Rb) in biotite, and long leaching time and low leaching efficiency of Rb in mica ore, the chlorination roasting-assisted ultrasonic enhanced water leaching method was employed to extract Rb from biotite ore in this study. During the chlorination roasting process, the optimal conditions were obtained, namely roasting temperature 900 °C, roasting time 40 min, and the mass ratio of ore to calcium chloride 1:1, the optimum leaching rate was 96.75 %. Compared with conventional leaching, the ultrasonic field could greatly shorten the leaching time and realize fast and efficient leaching of Rb. The optimal conditions for ultrasonic enhanced leaching were: ultrasonic power 100 W, leaching temperature 60 °C, leaching time 20 min, liquid–solid ratio 4:1, the Rb leaching rate was 98.73 %, which was 40 min shorter than conventional leaching. The particle size and SEM results indicated that the samples by ultrasonic leaching were smaller, no agglomeration phenomenon in a large area, and the surface of the samples was relatively smooth.     

包头磁尾矿稀土浸出和制备冰晶石的研究

包头白云鄂博矿床产生大量磁尾矿,会产生很大的环境问题。采用纯碱焙烧法对其进行了稀土浸出实验和氟的资源化利用研究。实验结果表明,包头磁尾矿与纯碱混合物于700℃焙烧1.0 h,经稀盐酸洗涤后,以3.0 mol.L-1稀硫酸于65℃下浸出4.0 h,氟的洗脱和稀土浸出效果较好,其中稀土浸出率为97.97%,氟浸出率为99.54%,氟洗脱率为77.32%。以该条件下产生的含氟废水为原料,于90℃下制备了冰晶石,溶液中氟回收率为85.63%,并推算出磁尾矿中氟的总利用率为66.21%。该法使磁尾矿由废弃物变为二次资源成为可能。     

汽油活性炭基脱硫吸附剂的制备与评价

以250℃温度下浓硫酸改性后的活性炭为载体,采用浸渍法制备了以MnO₂为活性组分的活性炭基的汽油脱硫吸附剂MnO₂/AC,考察了吸附剂的制备条件及脱硫条件对脱硫效果的影响。研究结果表明,适宜的吸附剂制备条件为,以Mn(NO₃)₂为活性组分前驱物,Mn(NO)₂浸渍液浓度0.15mol/L、常温下浸渍24h、焙烧温度350℃、焙烧时间2h。该吸附剂在静态吸附温度120℃、吸附时间2h、剂油质量比0.10的条件下可使原料油硫的质量分数从628.6×10^-6降至221.5×10^-6,脱硫率达到64.8%;在动态吸附温度60℃、空速1.76h^-1的条件下,初始流出汽油硫的质量分数降至21.8×10^-6,初始脱硫率达到96.5%。     

磷酸硼固体酸催化合成邻苯二甲酸二乙酯的性能研究

以邻苯二甲酸酐和无水乙醇为原料,固体磷酸硼为催化剂,在醇酐比为4：1,催化剂用量为苯酐用量的20．3％,反应时间8h,反应温度80℃条件下合成邻苯二甲酸二乙酯（DEP）,探讨了影响催化剂催化效果的各个因素。实验表明,催化剂最佳的制备条件为：磷酸、硼酸和硫酸的摩尔比为1：1：1,焙烧温度为300℃,焙烧时间为3．5h。DEP最高产率可达75％。     

Leaching of zinc and germanium from zinc oxide dust in sulfuric acid-ozone media

A new process of leaching zinc oxide dust by ozone oxidation in a sulfuric acid system was studied. The main factors affecting the leaching rate, such as ozone time, leaching temperature, initial acidity, leaching time, and liquid/solid mass ratio, were comprehensively investigated. The results show that leaching efficiency depends on all the above factors. The optimum conditions for leaching Zn and Ge from zinc oxide dust are as follows: ozone time 10 min, leaching temperature 90 ℃, initial acidity 160 g/L, leaching time 60 min, and liquid/solid mass ratio 7:1. Under the optimum conditions, the leaching rates of Zn and Ge are 95.79% and 93.65%, respectively. The leaching rates of zinc and germanium in the ozone leaching are 4.05% and 10.49% higher than those of the atmospheric leaching, respectively. Therefore, it is determined that ozone in solution plays a key role in rapidly oxidizing sulfide and releasing encapsulated germanium. Sulfuric acid-ozone media can efficiently extract Zn and Ge from zinc oxide dust.     

Leaching behavior of uranium and vanadium using strong sulfuric acid from Korean black shale ore

The present scientific study focused on leaching behavior of uranium and vanadium from Korean domestic ore. The leaching process experimental conditions optimized for uranium and vanadium metals recovery from Korean domestic ore and developed the basic experimental procedures such as time, particle size, acid influence, temperature effect and pulp density (PD) behavior. Acid influence on leaching process was tested and noted that 2.0 M sulfuric acid concentration is the optimized conditions for present study. The time influence on leaching process was observed and its optimized 2 h for complete leaching process. The temperature influence tested and optimized the 80 °C for complete leaching process and PD is 50 % (wt%). The bench scale experiments developed in a laboratory and tested in pilot level each batch 100 kg of ore sample.     

金银花中绿原酸的酶法提取工艺优化

采用酶法优化提取金银花中的绿原酸,考察纤维素酶酶的用量、酶解时间、酶解温度及回流提取温度对绿原酸含量的影响;用高效液相色谱法(HPLC)测定绿原酸含量。用纤维素酶法提取金银花可提高绿原酸得率。酶法提取最佳条件为:加入纤维素酶3.0%,在46℃下酶解4 h,再在56℃下浸提1 h;其绿原酸含量为3.57%。     

氨浸取粗氧化锌粉中锌的工艺研究

研究以粗氧化锌为原料制取活性氧化锌的浸取工艺。采用氨配合法将粗氧化锌中的锌充分浸取,以低级氧化锌为原料,对固液比、反应时间、碳酸氢铵用量、反应温度等因素先进行单因素实验,考察对氨配合法浸取率的影响,并在此基础上对主要的影响因素进行正交实验,从而优化出最佳浸取工艺条件为:浸取温度40℃,浸取时间3 h,固液比70 g/mL,碳酸氢铵用量2.5 g,浸取率可达92%。     

Leaching of El-Missikat low-grade fluoritized uranium ore by sulfuric acid: mechanism and kinetic

A well-characterized low-grade fluoritized uranium samples from new occurrence in Gabal El-Missikat prospect, Eastern Desert, Egypt was subjected to sulfuric acid leaching. The effects of leaching parameters on uranium dissolution mechanism were investigated. The shrinking core model was used to model leaching reactions. The kinetics equations indicates that the reactions appear to be controlled by layer diffusion process. The activation energy for uranium dissolution was evaluated. Low activation energy value (2.54 kJ mol⁻¹) confirm the diffusion layer mechanism. The presence of fluoride ions in the solution increases the dissolution of uranium. The optimum process operating parameters were: sulfuric acid concentration: 1.5 M, solid–liquid ratio: 1:3, contact time 8 h; agitation speed rate 200 rpm; and ore particle size − 75 µm at temperature 60 °C, in the absence of an external oxidant. Under these experimental conditions, the extraction efficiency of uranium was about 91%.

     

Kinetics of aluminum extraction from aluminum ash by leaching with sulfuric acid

Abstract

To explore the high value of aluminum ash development and utilization, factors (granularity of aluminum ash, leaching temperature, leaching time, sulfuric acid concentration, and liquid solid ratio) affecting the leaching rate of aluminum were evaluated. The kinetics of the leaching process was also investigated. The results indicate that the leaching rate of aluminum in aluminum ash can reach 88.9% under the following conditions: leaching temperature, 100?°C; leaching time, 120?min; sulfuric acid concentration, 2.5?mol/L; and liquid–solid ratio (mass ratio), 1.3:1. Studies on the leaching dynamics of aluminum in aluminum ash showed that leaching was divided into two stages. At the initial stage (i.e. at the leaching rate x?≤?0.3), the external diffusion resistance and the diffusion resistance of the solid product layer are negligible, and leaching is controlled by chemical reaction at the interface. The apparent activation energy was 5733.25?J/mol. At the leaching rate x?=?0.3–0.9, the leaching rate was controlled by the non-steady diffusion of the liquid film in the porous solid by the fluid reactant H₂SO₄. The apparent activation energy was 25390.87?J/mol.