Behavior of phase transformation of Baotou mixed rare earth concentrate during oxidation roasting

Based on the new process named “Combination Method” for metallurgy and separation of Baotou mixed rare earth concentrate (BMREC), the aim of this paper is to clearly elucidate the phase change behavior of BMREC without additives during oxidative roasting at 450–800 °C. The results indicate that the bastnaesite in BMREC is decomposed at 450–550 °C, the weight loss is about 10.3 wt%, and the activation energy (E) is 144 kJ/mol. The bastnaesite in BMREC is decomposed into rare earth fluoride, rare earth oxides (La₂O₃, Ce₇O₁₂, Pr₆O₁₁ and Nd₂O₃), and CO₂, particularly, with the increase of roasting temperature, bastnaesite in BMREC is more completely decomposed into LaF₃, which causes a decrease in leaching rate of La during the HCl leaching process. Additionally, the maximum cerium oxidation efficiency reaches about 60 wt% when the roasting temperature is equal to or above 500 °C, and the oxidation reaction rate of cerium increases with the increasing roasting temperature.     

浸出条件对稀土和铁浸出率的影响试验

以稀土精矿浓硫酸焙烧工艺中焙烧矿水浸过程为对象,研究了焙烧矿浸出温度、浸出时间、焙烧矿粒度等条件对稀土、铁浸出率的影响,并对水浸渣中稀土赋存状态进行了研究。研究表明,浸出温度和焙烧矿粒度对稀土、铁的浸出速率有较大影响,但对其浸出率没有影响,延长浸出时间,焙烧矿中的可溶性稀土、铁均可被浸出。水浸渣中的稀土主要以磷酸盐和氟氧化稀土形式存在,铁主要以磷酸铁形式存在,并含有少量硫化铁。     

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.     

Rare earth recovery from fluoride molten-salt electrolytic slag by sodium carbonate roasting-hydrochloric acid leaching

Fluorinated rare earth molten-salt electrolytic slag contains a considerable amount of rare earth elements,as well as a variety of heavy metals and fluorides that cause environmental pollution.Therefore,it is of great importance to fully utilise this resource.In this study,the transformation mechanism of fluorinated rare earth molten-salt electrolytic slag roasted with sodium carbonate,and the regulation mechanism of rare earth leaching under different roasting conditions were investigated with ...     

Extraction and back-extraction behaviors of 14 rare earth elements from sulfuric acid medium by TODGA

The unique physical and chemical properties of rare earth elements lay the foundation for their extensive application. N,N,N',N' Tetra-octyl-3-oxopentanediamide(TODGA) is excellent in its ability of extracting rare earth elements and it is favored for green initiative. In this paper, the extraction and back-extraction of14 rare earth elements by TODGA were studied. Experiments show that in conditions of 6 mol/L sulfuric acid, the extraction temperature of 25 ℃,the phase ratio of 1:1 and 0.04 mol/LTODGA(aviation kerosene as the diluent), the extraction rates of 14 rare earth elements including lanthanum, cerium, praseodymium,neodymium, europium, gadolinium, terbium, dysprosium, holmium, erbium, thulium, ytterbium, lutetium, and yttrium were 99.00%-99.73%. Mixed with hydrochloric acid and nitric acid(HCl 3.5 mol/L, HNO_30.5 mol/L), the recoveries of the 14 rare earth elements are 91.52%-99.91% when the extraction temperature is 25 ℃ and the ratio is 1:1. The following application is based on the optimum conditions above with practical samples(from the roasting production line of China North Rare Earth High-tech Company Limited) for extraction and back-extraction experiments. Experiments show that TODGA has excellent enrichment effect on 14 rare earth elements, the extraction rates are 91.36%-99.80%, the back-extraction rates are 87.29%-99.64% and the total recoveries are 81.19%-99.44%.     

Beneficiation of High-Aluminium-Content Hematite Ore by Soda Ash Roasting

High-aluminium-content iron ore is one of typical intractable iron ores, and magnetic separation and floatation processes are found impracticable to remove alumina from the ore effectively. In this article, a new process, roasting with addition of soda followed by leaching, is developed to remove aluminium from the ore. Results show that Al₂O₃ content decreases from 8.16% in raw ore to 2.13% in iron concentrate, and total iron grade increases from 48.92 to 63.21% when the ore is roasted at 1000°C for 15 min with the addition of 14.0% (wt.) sodium carbonate. Mechanisms of aluminium–iron separation were studied by using XRD, SEM, and thermodynamic methods, and it is shown that aluminium is transformed into sodium aluminosilicate, sodium aluminate, and corundum during roasting; sodium aluminate is able to be leached by water, so is sodium aluminosilicate by dilute acid solution, while corundum remains in the iron concentrate.     

硫代硫酸铵法从焙烧后的某含铜硫金精矿中回收金

研究了硫代硫酸盐法从焙烧后的含铜硫金精矿中回收金的工艺过程,研究了表明,焙烧前金的回收率很高,焙烧后,常温下硫代硫酸盐浸出,既可是较好的金浸出指标,又可降低硫代硫酸耗量。     

Co-extraction of tungsten and molybdenum from refractory scheelite–powellite blend concentrates by roasting with Na2CO3 and SiO2 and leaching with water

Studies on the recovery of tungsten and molybdenum from refractory scheelite–powellite blend concentrates – mainly composed of scheelite, powellite, and fluorite – were performed using soda-silicon roasting and water leaching processes. However, significant amounts of scheelite and powellite ore are present in an intergrowth state and the application of mineral dressing is not suitable for the separation and extraction of tungsten and molybdenum from this refractory ore. The effects of roasting parameters including sodium carbonate addition, roasting temperature, roasting time, and mass ratio of SiO₂/concentrate (w_SiO2/w_conc.) and the effects of leaching parameters including leaching temperature, leaching time, and liquid-to-solid ratio on the leaching efficiency of tungsten and molybdenum were investigated. The results demonstrated the efficiency of this process for the extraction of tungsten and molybdenum from the ore. Under the optimum experimental conditions where soda-silicon roasting is performed for 2?h at 850°C with three times, the stoichiometric ratio of Na₂CO₃ (Na₂CO₃:WO₃ and Na₂CO₃:MoO₃) and w_SiO2/w_conc of 12%, and water leaching is subsequently performed for 1?h at 70°C with a liquid-to-solid ratio of 3:1, the leaching ratios of W and Mo are 98.89% and 99.41%, respectively.     

Non-ammonia enrichment of rare earth by magnesium oxide from rare earth leaching liquor in magnesium salt system

In order to solve the problem of ammonia-nitrogen pollution in the enrichment process of the ionadsorption type rare earth ore,the technology of non-ammonia precipitation with magnesium oxide precipitant was carried out.It is determined that the rare earth precipitation efficiency is 99.6% and the purity of rare earth concentrates is only 85.89 wt%under the optimum precipitation conditions.And the contents of MgO,SO_3 and Al_2O_3 in the rare earth concentrates are 5.12 wt%,6.77 wt%and 1.78 wt%,respectively.Furthermore,the thermo-decomposition process of precipitates was investigated by TGDSC,XRD and FI-IR.The thermal decomposition process consists of two stages:the dehydration of rare earth hydroxide and alkaline rare earth sulfate within 900 ℃ and the thermal decomposition of RE_2O_2SO_4 at 900-1300 ℃.Therefore,a high-temperature calcinations method for removing SO_3 from precipitates is proposed.When the precipitates are calcined at 1300 ℃ for 2 h,the rare earth concentrates with a purity of 92.03 wt%can be acquired.Moreover,the content of SO_3 in the concentrate is only 0.46 wt%.In the MgO precipitation and high-temperature calcinations process,the raw material cost is low and the quality of rare earth concentrates is acceptable.It could have great significance for nonammonia enrichment of rare earth from the rare earth leaching liquor,and finally solve the problem of ammonia nitrogen in the extraction process of the ion-adsorption type rare earth ore within magnesium salt system.     

Rare earth extraction from bastnaesite concentrate by stepwise carbochlorination-chemical vapor transport-oxidation

A stepwise carbochlorination-chemical vapor transport-oxidation process is developed for the green rare earth extraction from a bastnaesite concentrate using carbon as reductant, chlorine gas as chlorination agent, SiCl₄ gas as defluorination agent, AlCl₃ as vapor complex former, and (O₂+H₂O) mixed gas as oxidant. Between 500 °C and 800 °C, the apparent activation energy of the carbochlorination within 2 hours changed from 17 to 10 kJ/mole roughly for the initial 20 minutes and final 1.5 hours, respectively, in the absence of SiCl₄, but these values reduced to 15 and 5.9 kJ/mole under 10 kPa of SiCl₄ gas, while the rare earth chloride conversion for 2 hours was 43 to 81 mol pct in the absence of SiCl₄ and 55 to 99 mol pct under 10 kPa of SiCl₄ gas. After carbochlorination at 550 °C for 2 hours in the (Cl₂+SiCl₄) atmosphere for efficient rare earth extraction and thorium-free volatile by-product release, throium was removed by chemical vapor transport at 800 °C for 0.5 hours in the (Cl₂+SiCl₄+AlCl₃) atmosphere and alkaline earths were separated from rare earths by oxidation at 700 °C to 1000 °C in the (O₂+H₂O) atmosphere for 0.5 hours, followed by water leaching at room temperature. Their combination allows a clean and efficient rare earth extraction from the concentrate.     

Rare earth extraction and separation from mixed bastnaesite-monazite concentrate by stepwise carbochlorination-chemical vapor transport

A stepwise carbochlorination-chemical vapor transport (SC-CVT) process is proposed for the rare earth extraction and separation from a mixed bastnaesite-monazite concentrate based on thermodynamic and kinetic analysis using carbon as reductant, chlorine gas as chlorination agent, SiCl₄ as defluorination agent, and AlCl₃ as vapor complex former. Between 500 °C and 800 °C, apparent activation energy of the carbochlorination within 2 hours changed from 22 to 16 kJ/mol roughly for the initial half hour and final 1 hour, respectively, in the absence of SiCl₄; but these values reduced to 15 and 2.1 kJ/mole under 2 kPa of SiCl₄ gas. The rare earth chloride yield for 2 hours was 56 to 88 mol pct in the absence of SiCl₄ and 92 to 99 mol pct in the presence of SiCl₄; but carbochlorination at above 1000 °C yielded a large amount of acid-insoluble residue. This, together with the negligible equilibrium vapor pressure of ThCl₄ at below 600 °C, suggests that carbochlorination of the mixed concentrate at temperatures as low as 500 °C in the (Cl₂ + SiCl₄) atmosphere is suitable for rare earth extraction and thorium-free volatile by-product release, which is different from the conventional Goldschmidt process at 1000 °C to 1200 °C. The CVT reaction of the carbochlorination product was performed at 800 °C for 0.5 hours in the (Cl₂ + SiCl₄ + AlCl₃) atmosphere and then at 1000 °C for 6 hours in the (Cl₂ + AlCl₃) atmosphere along different temperature gradients, leading to complete thorium removal and efficient rare earth separation, respectively. Their combination allows an efficient and environmentally conscious extraction and separation of rare earth elements from the mixed concentrate.     

萃取废水再利用探讨性研究

将[H+]为0.2⁰.3 mol/L的萃取捞稀土废水用生石灰中和至pH为7.0后过滤,滤液用于稀土焙烧矿的浸出。通过实验来反映焙烧矿的浸出率、稀土REO收率以及水浸液产品质量等指标,并同自来水浸出焙烧矿的情况相对比,通过对比来反映二者之间的指标差异情况,从而为稀土分离企业萃取废水的处理提供技术依据。     

High-efficiency simultaneous extraction of rare earth elements and iron from NdFeB waste by oxalic acid leaching

Iron can not be recovered at high value because only rare earth elements are effectively recovered from NdFeB waste via oxidation roasting-hydrochloric acid leaching process.In this study,a new method for leaching NdFeB waste with oxalic acid was developed.The high-efficiency,simultaneous and high-value recovery of rare earth elements and iron was realized to simplify the process and improve the economic benefit.Results of the oxalic acid leaching experiments show that under the optimum leaching conditions at 90℃ for 6 h in the aqueous solution of oxalic acid(2 mol/L) with a liquid-solid ratio of60 mL/g,the iron leaching efficiency and precipitation rate of rare earth oxalate reach 93.89% and 93.17%,respectively.Rare earth oxalate and Fe(C₂O₄)3^3- were left in the residue and the leaching solution,respectively.The leaching mechanism was further analyzed by characterising the leach residues obtained through X-ray powder diffraction(XRD) and scanning electron microscopy-energy dispersive X-ray spectroscopy(SEM-EDS).Results of the leaching kinetics study indicate that the process of oxalic acid leaching follows the shrinking nucleus model,and the leaching kinetics model is controlled by the mixed factors of diffusion and chemical reaction.The leaching residue was calcined at 850℃ for 3 h and then decomposed into rare earth oxide,which can be directly used to prepare rare earth alloy via molten salt electrolysis.For the leaching solution,ferric oxalate solution was reduced using Fe powder to prepare the ferrous oxalate(FeC₂O₄-2H₂O).     

Precipitation transformation of rare earth sulfate into chloride with p-dodecylphenoxy carboxylic acids

Methods for transforming rare earth(RE) sulfate into chloride mainly include extraction process with organophosphonic mono-acids or aliphatic acids and precipitation process with ammonium bicarbonate(NH_4 HCO_3).In this paper,alkylphenoxy carboxylic acids(HAs) ofp-dodecylphenoxy acetic acid(HA-Ⅰ),pdodecylphenoxypropanoic acid(HA-Ⅱ) and p-dodecylphenoxybutyric acid(HA-Ⅲ),which were liquid at room temperature were synthesized and characterized.The precipitation mechanisms of RE elements with the HAs were investigated and the HA/RE molar ratios of the solid complexes were determined as3:1 by equi-molar series method which accord with the principle of charge balance.Applicability of HAs for the transformation of RE sulfate from concentrated sulfuric acid roasted RE concentrate into chloride via precipitation method was discussed.100% HA-Ⅱ was selected as the liquid organic precipitant without dilution of volatile solvent soracceleration of phase separation by phase-modifiers.The RE sulfate solution can be precipitated by HA-Ⅱ after neutralization with liquid NaOH and stripped with concentrated HCl at room temperature.High concentration of RE chloride of 218.1 g/L with low residue of sulfate radical of 0.536 g/L was obtained.The residual organic precipitant in the raffinate solution was tested to be lower than 8 mg/L at 25℃and the chemical oxygen demand(COD) in wastewater was less than 50 mg/L.     

Optimization of Lithium Extraction from Lepidolite by Roasting Using Sodium and Calcium Sulfates

The lithium extraction from a lepidolite concentrate using roasting, followed by water leaching, was studied. Several alternative additives were initially tested. The use of sodium and calcium sulfates as additives was evaluated in more detail. The influence of some process variables, namely the roasting time, roasting temperature and the additive/concentrate mass ratio, was studied applying a design of experiments. The lithium extraction was modelled and the fitted and validated model was used to optimize the process response. The increase in the additive/concentrate mass ratio, roasting time and temperature seems to result in solid state reactions and transformations that lead to phase, morphological and particle size distribution modifications, which were assessed by XRPD, SEM, and particle size analyses. In this process, lithium sodium sulfate formation constitutes a crucial step enabling the Li water leaching. High lithium extractions were estimated for several combinations of factors. At 850°C, lithium extractions over 90% are obtained when the roasting time is above 1.90 hour and the additive/concentrate mass ratios are over 0.77. An increase in the temperature to 875°C also leads to lithium extractions over 90% for a roasting time of 1 hour and an additive/concentrate mass ratio of 0.60.     

焙烧-酸浸-氰化法从复杂金精矿中回收金银铜

采用焙烧-酸浸-氰化工艺综合回收复杂金精矿中的金、银、铜.结果表明,焙烧温度、焙烧时间、焙烧添加剂种类和用量对金、银、铜浸出率影响显著.实验确定了较优工艺条件为:焙烧添加剂NaOH用量为6 %,温度630 ℃,焙烧时间3 h,硫酸浓度1 mol/L,酸浸液固体积质量比5:1,酸浸温度50 ℃,酸浸4 h,氰化纳浓度3 ‰,氰化浸出液固体积质量比5:1,常温氰化72 h.在上述条件下,金、银、铜浸出率分别达到93.53 %、75.37 %、94.23 %.      

Studies on Extraction of Potassium from Feldspar by Roast-leach Method Using Phosphogypsum and Sodium Chloride

The extractions of potassium value from feldspar via roasting and leaching route was studied with a focus on the effects of the roasting time, temperature, additives, and particle size. Sodium chloride and phosphogypsum (PG) were used as a source of chloride and calcium, respectively, and played the important role during the roasting of feldspar. When the feldspar sample was roasted at 900°C with sodium chloride alone, the extraction of potassium was limited to 61%. The extraction could go up to 92.5% by the addition of phosphogypsum along with sodium chloride. The optimum conditions of potassium extraction were found to be, particle size 100 µm, roasting temperature 900°C and roasting time of one hour. The X-ray diffraction study indicated the formation of sylvite (KCl) in the roasted product and its disappearance in roast-leach residue due to its high water solubility. The morphological changes during the roasting process were clearly observed by field emission scanning electron microscope (FESEM) images. The extraction of potassium from feldspar was best fitted by the Ginstling and Brounshtein kinetic model. The activation energy of 238.6 KJ/mole and 28.73 KJ/mole for low and high-temperature regions indicated that the overall extraction process follows two-steps reaction path.     

Metallization of siderite ore in reducing roasting

The behavior of the initial ore and the concentrate of magnetoroasting beneficiation during metallization under the conditions that are close to those for reducing roasting of iron ores in a rotary furnace is studied in terms of works on extending the field of application of Bakal siderites. A difference in the mechanisms of the metallization of crude ore and the roasted concentrate is observed. The metallization of roasted concentrate lumps is more efficient than that of crude siderite ore. In this case, the process ends earlier and can be carried out at higher temperatures (1250–1300°C) without danger of skull formation.     

Iron recovery and rare earths enrichment from Bayan Obo tailings using Coal-Ca(O H)2-NaO H roasting followed by magnetic separation

The recovery of iron and enrichment of rare earths from Bayan Obo tailings were investigated using CoalCa(OH)_2-NaOH roasting followed by magnetic separation.The influences of roasting temperature,roasting time,coal content,milling time,Ca(OH)_2 dosage and NaOH dosage on the iron and rare earths recovery were explored.The results showed that the magnetic concentrate containing 70.01 wt.% Fe with the iron recovery of 94.34% and the tailings of magnetic separation containing 11.46 wt.%rare earth oxides(REO)with the REO recovery of 98.19% were obtained under the optimum conditions(i.e.,roasting temperature of 650°C,roasting time of 60 min,coal content of 2.0%,milling time of 5 min,and NaOH dosage of 2.0%).The Ca(OH)_2 dosage had no effect on the separation of iron and rare earths.According to the mineralogical and morphologic analysis,the iron and rare earths of Bayan Obo tailings could be utilized in subsequent ironmaking process and hydrometallurgy process.     

The use of computational thermodynamic for yttrium recovery from rare earth elements-bearing residue

The rare earth elements are considered critical metals,due to the risk of supply interruption.The recycling of waste electrical and electronic equipment can be an alternative to supply the rare earth market.Several processes have been developed,and by aqueous media is the most prominent,which makes possible the extraction and separation of elements even in low concentration(traces).As an example of thermal processing,the use of thermodynamic simulations might benefit the metal extraction in hydrometallurgical processing.For this reason,the goal of this work is to evaluate the use of FactSage 7.2 software for the leaching of fluorescent lamp powders by sulfuric acid.The effect of concentration and temperature was evaluated.Results comprise that the thermodynamic software wellpredicted the solid phase formed in all residues of leaching experiments-gypsum was predicted by the software and identified in XRD analyses.It demonstrates that FactSage software can be explored for metals extraction in aqueous media,being important for trace-elements extraction.Yttrium extraction reaches up to 95%at 45℃using H₂SO₄2 mol/L.