白钨矿焙烧转料的氢氧化钠浸出动力学研究

中国的钨矿种类以白钨矿为主,黑白混合钨矿与黑钨矿为辅,白钨矿虽储量丰富,但传统的白钨冶炼工艺冶炼难度大、成本高且环境污染严重。为此,提出了一种“氯化镁焙烧—氢氧化钠浸出”绿色环保新工艺。为进一步明确焙烧矿碱浸工序的反应机制,以白钨精矿与MgCl₂焙烧的转型物,即MgWO₄转料为研究对象,以氢氧化钠为浸出剂,详细考察了氢氧化钠浓度、反应温度、转料粒度、搅拌速度以及时间对MgWO₄转料碱浸过程的影响。在此基础上,开展动力学研究,结果表明：氢氧化钠浓度和反应温度为浸出反应过程中的主要影响因素,矿物粒度为次要影响因素。对MgWO₄转料的氢氧化钠浸出动力学实验数据进行拟合,可知该浸出过程符合整体反应模型(volume reaction model),计算可得反应活化能为89.9 kJ·mol^-1,氢氧化钠浓度反应级数为1.718,矿物粒度影响指数为-0.391。反应过程受化学反应控制,其反应动力学方程为：■     

焙烧钛渣的反应动力学研究

研究了钠化钛渣加碱焙烧过程的反应动力学、焙烧反应控制步骤,探讨了反应温度、碱渣质量比、钛渣粒度对TiO_2提取率的影响。试验结果表明:在适宜条件下,TiO_2提取率为92%,烧渣TiO_2品位达98.68%;钛渣焙烧优化工艺条件为焙烧温度900℃,碱渣质量比4∶6,钛渣粒度-200目,焙烧时间60min;扩散与化学反应的混合控制作用是影响钛渣焙烧效果的最主要因素,焙烧反应符合收缩未反应核模型,表观活化能为57.39kJ/mol,焙烧动力学方程可描述为1+2(1-x)-3(1-x)~(2/3)=3.532 3exp(-57 390/(RT))·t。     

硫酸铵焙烧钛渣的反应动力学研究

对硫酸铵焙烧钛渣提取TiO2的焙烧过程进行了动力学研究,考察了焙烧温度、钛渣粒度和硫酸铵与钛渣质量比对TiO_2提取率的影响。结果表明:硫酸铵焙烧钛渣过程受化学反应控制,焙烧反应的表观活化能为66.59kJ/mol,动力学方程为1-(1-x)~(1/3)=0.246 2×10~3exp[-66 587/(RT)]×t。     

高铁三水铝石矿溶出动力学

研究了高铁三水铝石矿常压条件下的溶出动力学。分别考察了矿石粒度、氢氧化钠浓度、溶出温度、搅拌速度对溶出性能的影响。研究发现,常压条件下矿石溶出率大于90%。提高溶出温度或氢氧化钠浓度、减小矿石粒度均可加快矿石的溶出速度。综合考虑矿石粒度、氢氧化钠浓度和搅拌速度的影响,推导出高铁三水铝石矿溶出动力学经验方程为:1-2/3α-(1-α)~(2/3)=1.28×10~(15)×(c_(NaOH))~(1.70)×d~(-0.36)×r~(0.74)×e~(-117402/RT)×t。该方程与内扩散控制的动力学模型吻合较好,反应速率受包覆在晶体表面的针铁矿和赤铁矿固体膜层所控制。     

焦炭还原焙烧锰矿制备硫酸锰及其机理

以非洲进口锰矿为原料,焦炭为还原剂,进行还原焙烧、硫酸浸出制备硫酸锰溶液。探究焦炭配比、焙烧温度、焙烧时间对锰还原效果的影响,确定最佳工艺参数,并对焙烧反应体系进行热力学分析及对浸出过程进行动力学分析。结果表明,最佳的焙烧条件为：焦炭配比15%、800℃焙烧60 min,锰的还原率达到97.84%。焙烧体系反应都是自发进行的,表观活化能为22.458 6 kJ/mol,且浸出过程受内扩散控制,动力学方程为：1-2x/3-(1-x)^2/3=8.48exp(-22458.6/RT)×t。     

硫酸铵焙烧红土镍矿工艺优化

本文采用硫酸铵焙烧-水浸工艺,将红土镍矿与硫酸铵混合焙烧,研究了焙烧温度、焙烧时间、物料配比和矿粉粒度对镍、镁、铝和铁提取率的影响.采用正交实验优化工艺条件,得到影响焙烧效果的各因素顺序为:物料配比、焙烧时间、焙烧温度、矿粉粒度.优化工艺条件为:焙烧温度450℃,焙烧时间120min,物料配比2∶1,矿粉粒度小于80μm,镍、镁和铝提取率大于98%,铁提取率大于94%.研究了焙烧过程镍、镁、铝和铁与硫酸铵反应过程的动力学.镍、镁、铝和铁的反应速率均符合未反应收缩核模型,表观活化能分别为19.93 k J·mol~(-1)、18.96 k J·mol~(-1)、17.86 k J·mol~(-1)和20.83 k J·mol~(-1).     

钠化焙烧钛渣的酸浸动力学研究

研究了钠化焙烧—酸浸工艺制备高钛渣过程中酸浸动力学、酸浸溶出规律和溶出反应的控制步骤,探讨了反应温度、盐酸浓度、搅拌速度对钛转化率的影响。结果表明,在下述最优工艺条件下,钛转化率达93%,TiO2品位达98.68%:搅拌速度400r/min、液固比10∶1、酸浓度20%、110℃酸浸90min。钛渣焙烧产物酸浸过程受固体产物层的内扩散控制,钛渣酸浸过程符合收缩未反应核模型,表观活化能为11.37kJ/mol,浸出钛渣动力学方程为1+2(1-x)-3(1-x)~(2/3)=0.244exp[-11370/(RT)]t。     

石煤提钒碱浸过程动力学研究

研究了石煤空白焙烧料NaOH溶液浸出过程中液固比、NaOH溶液初始浓度、反应搅拌速率、矿石粒径、温度对浸出速率的影响,确定了石煤提钒碱浸过程动力学方程式,计算了反应的表观活化能.结果表明:液固比、反应搅拌速率对浸出速率的影响不大,NaOH溶液初始浓度、矿石粒径、浸出温度对浸出速率的影响显著;V2O5的浸出动力学适用于收缩未反应芯模型,浸出反应服从化学反应控制过程;根据Arrhenius公式,使用曲线拟合法,近似求得该反应表观活化能约为58.5 kJ·mol-1,进一步表明该反应属于化学反应控制.     

白云鄂博稀土焙烧矿浸出动力学研究

以白云鄂博高品位稀土精矿为研究对象,对稀土精矿进行NaOH-Na_2CO_3焙烧,详细研究了焙烧矿中稀土在盐酸浸出过程中的化学与物理机制,分析动力学相关影响因素,确定了动力学浸出模型;结果表明:当HCl浓度4.0 mol·L~(-1),液固比(L/S) 4.5∶1.0,搅拌速度250 r·min~(-1),浸出温度90℃,酸浸时间为25 min时,稀土浸出率为93.2%,盐酸浓度和浸出温度为化学反应过程主要影响因素。根据扫描电镜(SEM)和X射线衍射(XRD)结果:稀土精矿与焙烧助剂NaOH-Na_2CO_3在660℃焙烧90 min,矿物中氟碳铈矿和独居石分解为稀土氧化物, Ce(III)被氧化为Ce(IV),焙烧矿盐酸浸出残渣中有少量未分解REPO_4和残留的CaF_2。计算出稀土元素表观活化能为37.52 kJ·mol~(-1), HCl浓度和液固比反应级数分别为2.446和2.226;根据焙烧矿中稀土浸出动力学相关试验数据推导出浸出模型为一种收缩核模型变体,化学反应过程受产物层界面传质和扩散共同作用。     

中浸渣铁的浸出过程动力学

研究中浸渣铁在热酸浸出中的浸出过程动力学,考察了搅拌转速、反应温度、初始H_2SO_4浓度、矿物粒度和Fe~(3+)离子浓度对铁浸出速率的影响。结果表明:该浸出反应的动力学方程符合收缩核模型。依据反应温度、矿物粒度等多个影响因素研究发现中浸渣热酸浸出过程受化学反应控制。浸出反应的表观活化能为45.657×10~3J/mol,浸出剂硫酸与Fe~(3+)离子浓度对反应的表观反应级数分别为1.09446和-0.07317。中浸渣热酸浸出的宏观动力学方程为:1-(1-α)~(1/3)=0.4442[H~+]~(1.09)[Fe~(3+)]~(-0.07)d_0~(-1)exp(-45657/RT)t。     

碱熔预处理对废旧稀土荧光粉中稀土提取的影响

针对废旧稀土荧光粉特性,采用高温碱熔与酸浸相结合的方法处理废旧稀土荧光粉,考察碱熔温度、氢氧化钠用量、碱熔时间对稀土浸出率和铝回收率的影响,并对碱熔过程中的物相及形貌变化进行分析.结果表明,在碱熔温度为1050℃、氢氧化钠与废粉质量比为2.5∶1、碱熔时间为2 h 条件下,其稀土浸出率可达98%以上,铝回收率可达98%.通过对碱熔产物物相和形貌分析表明,废粉晶体结构被有效破坏,其中稀土以稀土氧化物形式存在,碱熔产物变成无定型云絮状.     

从废荧光粉中回收稀土的工艺研究

研究了从废荧光粉(REO^12.00%)中回收稀土元素的工艺。采用碳酸钠焙烧-酸浸出工艺回收废荧光粉中的稀土,研究了碳酸钠加入量、焙烧温度、焙烧时间以及浸出条件对稀土回收率的影响。研究结果表明,碳酸钠焙烧试验的最佳条件为碳酸钠与荧光粉焙烧比例1∶2,焙烧温度700℃,焙烧时间1 h;焙烧产物用盐酸浸出,浸出试验最佳条件:盐酸浓度、液固比、浸出温度、浸出时间分别为3 mol/L、10∶1、70℃、2 h,在上述焙烧及浸出最优条件下,稀土总回收率(REO)达97%以上。     

Selective extraction and recovery of rare earth metals from waste fluorescent powder using alkaline roasting-leaching process

Waste management of rare earth metals(REMs) containing materials and recycling of rare earth metals(REM) from waste materials are becoming more and more important due to high demand and resource exhaustion. However, extraction of REM from waste fluorescent powder materials is difficult because of their special aluminate structure. A novel "alkaline roasting-acid leaching" process was developed in this study. The alkaline roasting process mechanism was examined using differential thermal analysis(DTA)-thermogravimetric(TG) measurements, and the roasting product was characterized by XRD analysis. In this process, Al_2O_3 was converted into water soluble NaAlO_2 via alkaline roasting, and NaAlO_ 2 in the roasting product could be easily dissolved in water, while the rare earth oxide(REOs) remained as solid. After filtration, REOs cake was leached using hydrochloric acid to achieve 99.8% of REM recovery. It was concluded that the alkaline roasting-acid leaching process could effectively separate Al_2O_3 and REOs with high REM recovery.     

Decomposition of mixed rare earth concentrate by NaOH roasting and kinetics of hydrochloric acid leaching process

A new clean extraction technology for the decomposition of Bayan Obo mixed rare earth concentrate by NaOH roasting is proposed.The process mainly includes NaOH roasting to decompose rare earth concentrate and HCl leaching roasted ore.The effects of roasting temperature,roasting time,NaOH addition amount on the extraction of rare earth and factors such as HCl concentration,liquid-solid ratio,leaching temperature and leaching time on the dissolution kinetics of roasted ore were studied.The experimental results show that when the roasting temperature is 550℃ and the roasting time is 60 min,the mass ratio of NaOH:rare earth concentrate is 0.60:1,the concentration of HCl is 6.0 mol/L,the ratio of liquid to solid(L/S) 6.0:1.0,and the leaching temperature 90℃,leaching time 45 min,stirring speed 200 r/min,and the extraction of rare earth can reach 92.5%.The relevant experimental data show that the process of HCl leaching roasted ore conforms to the shrinking core model,but the control mechanism of the che mical reaction process is different when the leaching temperature is different.When the leaching temperature is between 40 and 70℃,the chemical reaction process is controlled by the diffusion of the product through the residual layer of the inert material.The average surface activation energy of the rare earth element is E_a=9.96 kJ/mol.When the leaching temperature is 75-90℃,the chemical reaction process is controlled by the interface transfer across the product layer(product layer interface mass transfer) and diffusion.The average surface activation energy of rare earth elements is E_a=41.65 kJ/mol.The results of this study have certain significance for the green extraction of mixed rare earth ore.     

The soda-ash roasting of chromite ore processing residue for the reclamation of chromium

Sodium chromate is produced via the soda-ash roasting of chromite ore with sodium carbonate. After the reaction, nearly 15 pct of the chromium oxide remains unreacted and ends up in the waste stream, for landfills. In recent years, the concern over environmental pollution from hexavalent chromium (Cr⁶⁺) from the waste residue has become a major problem for the chromium chemical industry. The main purpose of this investigation is to recover chromium oxide present in the waste residue as sodium chromate. Cr₂O₃ in the residue is distributed between the two spinel solid solutions, Mg(Al,Cr)₂O₄ and γ-Fe₂O₃. The residue from the sodium chromate production process was analyzed both physically and chemically. The compositions of the mineral phases were determined by X-ray diffraction (XRD), scanning electron microscopy (SEM), and electron probe microanalysis (EPMA). The influence of alkali addition on the overall reaction rate is examined. The kinetics of the chromium extraction reaction resulting from the residue of the soda-ash roasting process under an oxidizing atmosphere is also investigated. It is shown that the experimental results for the roasting reaction can be best described by the Ginstling and Brounshtein (GB) equation for diffusion-controlled kinetics. The apparent activation energy for the roasting reaction was calculated to be between 85 and 90 kJ·mol⁻¹ in the temperature range 1223 to 1473 K. The kinetics of leaching of Cr³⁺ ions using the aqueous phase from the process residue is also studied by treating the waste into acid solutions with different concentrations.     

从荧光粉废料中提取稀土工艺研究

采用4种方案从荧光粉废料中提取稀土元素,并考查了盐酸法提取稀土时盐酸和双氧水用量对稀土浸出率的影响,随后采用碳酸钠焙烧法提取渣中较难浸出的铈、铽,最后采用中和法对酸浸出液进行除杂。结果表明,100g物料盐酸最佳用量为150mL,双氧水用量为20mL,钇、铕浸出率可达99%。经碳酸钠焙烧—盐酸浸出后铽浸出率达到55%,除杂后铁、硅、铝含量分别降至11.47mg/L、15.93mg/L和150mg/L。     

从钕铁硼废料中提取稀土工艺研究

采用氧化焙烧-盐酸分解法,研究从钕铁硼废料中提取稀土的工艺条件,探讨了焙烧温度和时间对铁的氧化率的影响,在浸出过程中考察了盐酸浓度、反应时间、反应温度以及液固比对稀土浸出率的影响,并分析了pH值和陈化时间对浸出液除杂效果的影响.结果表明:在700℃焙烧1.5 h,铁的氧化率最高,铁基本完全氧化成三价铁,在最佳浸出条件下稀土浸出率高达到99.33%,浸出液中和除杂时,调节pH值为3.5,陈化时间大于2 h,料液中非稀土杂质含量低,特别是铁仅为0.0014 g/L,浸出液完全达到稀土萃取的要求.      

Kinetics of nitric acid leaching of cerium from oxidation roasted Baotou mixed rare earth concentrate

The kinetics of nitric acid leaching of cerium was investigated for the oxidation roasted Baotou mixed rare earth concentrate. The effects of leaching temperature, HNO₃ concentration, liquid–solid ratio (L/S) and stirring rate on rare earth extraction were studied. The XRD and SEM mapping analysis of the samples before and after acid leaching shows that the roasted bastnaesite is completely leached. Besides, the decomposition process of oxidizing roasting was also obtained by TG–MS and XRD. Different kinetics models were applied in this leaching process. The results of dynamic fitting show that the leaching process can be described by a new variant of the shrinking-core model. And the leaching rate is controlled by both the interfacial transfer and diffusion through the product layer. The apparent activation energy is calculated as 76.78 kJ/mol and the reaction orders with respect to HNO₃ concentrations and liquid–solid ratio are determined to be 7.609 and 2.516, respectively. Besides, an empirical rate equation is obtained to describe the process.     

硝酸钠为氧化剂的碳酸稀土焙烧过程研究

对以硝酸钠为氧化剂的富铈混合碳酸稀土焙烧过程进行了实验研究。通过对氧化剂加入量、焙烧温度、焙烧时间的考察 ,得出了铈氧化率的变化规律。研究结果证明 ,硝酸钠对铈的氧化作用是显著的。     

Decomposition of Niobium Ore by Sodium Hydroxide Fusion Method

The decomposition kinetics of niobium ore in the NaOH system was studied experimentally. The results show that the reaction products are sodium metaniobate and sodium niobate formed by the reaction of pyrochlore with sodium hydroxide under roasting. The effects of temperature, particle size, and mass ratio of alkali-to-ore were studied. The conversion rate of niobium exceeded 99 pct after 20 minutes at 923 K (650 °C) with a mass ratio of alkali-to-ore 1.2:1 and with initial particle size 75 to 106 μm. The kinetic study indicates that the shrinking core model is applicable and the process is controlled by a chemical reaction. The activation energy was calculated to be 78.82 kJ mol^–1.