Coordination-reduction leaching process of ion-adsorption type rare earth ore with ascorbic acid

The magnesium sulfate(MgSO₄)-ascorbic acid(Vc) compound leaching technique can extract rare earth elements(REEs) existing in ion-exchangeable phase and colloidal phase from ion-adso rption type rare earth ore through the synergy effect of coordination and reduction,but its reaction process and mechanism remain unclear.In this paper,the coordination-reduction leaching mechanism was analyzed from the perspectives of leaching thermodynamics and kinetics,which provide theoretical guidance...     

Leaching of vanadium from carbonaceous shale

The leaching of vanadium from carbonaceous shale using dilute H₂SO₄ was investigated, and the mechanism of leaching determined. The results showed that higher leaching efficiency of vanadium was obtained by increasing initial concentration of H₂SO₄, raising leaching temperature and prolonging leaching time. Addition of ammonium fluoride also enhanced recovery. A recovery of 92% was obtained using a liquid to solid ratio 4:1, initial H₂SO₄ concentration 18%, NH₄F addition 4.8 wt.% of carbonaceous shale, leaching temperature 95 °C and contact time 8 h; the recovery was only 56% without NH₄F. The presence of NH₄F enhanced the leaching of vanadiferous mica.     

Leaching of rare Earth elements from an Illinois basin coal source

The existence of rare earth elements(REEs) in coal sources at elevated concentrations has been the focus of several studies over the past decade. However, limited research has been conducted on methods to recover and refine the REEs. This paper reports the results of a detailed study into the potential of selectively recovering REEs in an Illinois basin coal source by leaching. Leaching characteristics are obtained for several segments obtained from a core sample as well as three different reject materials collected at a coal processing plant. Using a 1.2 mol/L sulfuric acid solution at 75 ℃,over 60% REE recovery is achieved from the direct floor and an inner parting material as well as the coal-rich core segments that are pretreated by low-temperature plasma oxidation to obtain access to the microdispersed mineral matter. In the leachable parting material, fluorapatite is detected by XRD analysis,which is one of the more soluble phosphate minerals with a documented association with REEs. For the three plant reject samples, the leaching recovery values obtained for the heavy REEs are higher than those obtained for the light REEs under the standard leaching conditions and when 0.1 mol/L(NH_4)_2SO_4 was used to extract REEs by an ion exchange mechanism. Thermal activation by roasting or chemical activation by pretreatment using 8 mol/L NaOH solution increases the total REE recovery with significantly higher gains obtained for the light REEs. Leaching kinetics are relatively fast within the first 2 h and then slow to provide relatively low overall recovery values under the standard test conditions for the coarse and fine reject samples. However, significantly higher recovery values are realized when treating mixed-phase(middling) particles existing within the coarse reject material.     

Decomposition of monazite concentrate in sulphuric acid

A sulphuric acid bake-leach process was investigated for the extraction of rare earth elements (REEs), uranium, and thorium from a monazite concentrate. X-ray diffraction analysis showed the formation of water-soluble sulphates that readily dissolved during the water leach step. Nearly complete extraction of REEs, uranium, and thorium was achieved when the concentrate was baked at 250°C for 4?h with a sulphuric acid to concentrate (S/C) weight ratio of 4:1. At the 1:1 sulphuric acid to concentrate ratio ～65% of REEs were leached while only ～2–5% of uranium and thorium were extracted. Temperature (180–250°C) had little effect on the extraction of REEs, but greatly affected the extraction of thorium and uranium. In general, the extraction of thorium and uranium decreased with an increase in temperature. The effect of particle size (48–137?µm) was also tested and found to have little effect on the extraction of REEs. Acid consumption was calculated and found to increase with longer baking times (0.3–0.8?g acid/g concentrate). The major mineralogical phases formed because of the sulphation reaction included potassium cerium sulphate hydrate (2K₂SO₄·Ce(SO₄)₂·2H₂O), lanthanum sulphate hydrate (La₂(SO₄)₃·2H₂O), sodium praseodymium sulphite hydrate (NaPr(SO₃)₂·2H₂O), and neodymium sulphate hydrate (Nd₂(SO₄)₃·5H₂O), all of which were water-soluble.     

Recovery of valuable metals from jarosite by sulphuric acid roasting using microwave and water leaching

In this paper, jarosite residue (JR) blended with concentrated H₂SO₄ was subjected to a process comprising microwave roasting and water leaching. The effects of H₂SO₄/JR weight ratio, microwave roasting temperature and time, water leaching conditions on the recovery of Fe, Zn, In, Cu, Cd, Ag and Pb were investigated utilising a series of experiments.

Based on energy conservation and environmental protection, optimum conditions for metals recovery from JR were determined as: H₂SO₄/JR weight ratio?=?0.36, microwave roasting temperature, 250°C; roasting time, 30?min; leaching temperature, 50°C; leaching time, 1?h; and liquid–solid ratio, 4:1 (mL/g), thus, the extraction of Fe, Zn, In, Cu, Ag and Cd were 89.4, 80.7, 85.1, 90.7, 61.3 and 48.8% respectively, while the Pb was concentrated in the final residue. Scanning electron microscope-energy dispersive spectrometer (SEM-EDS) patterns were used to characterise and analyse the transformation of valuable metals in the residue after roasting and leaching.     

Extraction of scandium from red mud by acid leaching with CaF_2 and solvent extraction with P507

The extraction of Sc by acid leaching with CaF_2 and solvent extraction with P507 from red mud was proposed.The influence of acid leaching and solvent extraction on recovery of Sc was investigated.The CaF_2 can obviously improve the leaching efficiency of Sc and reduce the acid consumption.The leaching efficiency of Sc increases from 74% to 92% and the dosage of acid reduces under suitable conditions by adding 5% CaF_2.The minerals in red mud can easily be decomposed and leached into the acid solution with CaF_2 through analysis of XRD pattern.The particles of red mud become smaller and multihole.The Sc can be selectively extracted with 10% P507 at the pH value of 0.1 from the acid leaching solution.More than 98% of Sc and less than 10% of Al and Fe are extracted.The SC_2O_3 with purity of 99% is obtained after the process of reverse extraction with NaOH,H_2SO_4 dissolution,precipitation by oxalic acid and roasting at 750℃.     

Extraction of cadmium from phosphoric acid solutions with amines Part I. Extractant selection,stripping, scrubbing and effects of other components

Cadmium can be effectively removed from wet-process phosphoric acid (WPA) by solvent extraction using long-chain amines, preferably tertiary amines, as extractants. Stripping can be performed either with water or with acids such as phosphoric or sulphuric acid depending on the final conversion of the cadmium waste. Scrubbing could be necessary in some cases and can be performed with alkali.A factorial design experiment was performed to determine the effects of four different components (H₂SO₄, HCl, HF, H₂SiF₆) in WPA on cadmium extraction with Alamine 336. The largest influence on cadmium extraction is seen for the hydrochloric acid, drastically increasing the distribution coefficient. Variations of the concentrations of H₃PO₄, HCl, HF, Fe and Al were also studied. The extraction of cadmium was seen to be strongly dependent on phosphoric acid concentration.A simple extraction model is proposed indicating that CdCl₄^2? ions are extracted.     

Effect of roasting activation of rare earth molten salt slag on extraction of rare earth,lithium and fluorine

A new process was proposed to extract rare earth elements(REEs),Li and F from electrolytic slag of rare earth molten salt by synergistic roasting and acid leaching.Firstly,the thermodynamic analysis of roasting reaction was carried out,then the effects of roasting factors on leaching REEs,Li and F in slag were investigated.In additions,the mineral phase and morphology of molten salt slag,roasting slag and acid leaching slag were characterized,and the migration mechanism of REES,Li and F minerals...     

Metal Recovery from Bottom Ash of an Incineration Plant: Laboratory Reactor Tests

The bottom ashes, produced by the industrial incinerators, are an essential secondary raw material resource that has been drawing attention to recover economically important metals. In the present study, a hydrometallurgical process (chemical leaching) has been discussed to recover some economically important metals using lab scale reactors (600 mL and 2L capacity). Prior to the leaching tests, the material was characterized for chemical composition and mineralogical phase analysis through XRD. In addition, the conventional process of magnetic separation was applied before leaching tests to remove some easily separable parts. Two acidic reagents (HCl and H₂SO₄) and an alkali reagent (NaOH) have been used to compare the recovery of metal values from the pre-treated bottom ash samples at varying concentrations. Process parameters such as acid/alkali concentration, working volume and temperature have been optimized and the recoveries of metal values under optimum conditions were recorded. The studies showed that 3M H₂SO_4, 1.5 L working volume and 80°C were sufficient to leach 88% Mo, 82% V, 37% Ni, 37% Fe, and 28% Cu. Acid leaching tests using H₂SO₄ were found to be an economical and appropriate solvent for metal recovery from bottom ash.     

Removal of Iron From Quartz Ore Using Different Acids: A Laboratory-Scale Reactor Study

The purification of quartz using chemical processes is extremely important for many industries, including the glass, electronic, detergent, ceramics, paint, refractory, and metallurgy industries, as well as for advanced technology products. The purpose of this work was to investigate the removal of iron as an impurity from quartz ores using a chemical leaching method with different reagents.

The iron content of the quartz ore sample was 310 ppm, and the iron was in the form of Fe₂O_3. In the first step, pre-enrichment studies were conducted based on the particle size, and the Fe₂O₃ content of the quartz ore was decreased to 88 ppm. A statistical design of the experiments and an ANOVA (analysis of variance) were performed in the second step to determine the main effects and interactions of the researched factors, which were the concentration of the leaching reagent (H₂SO₄, HCl, H₃PO₄, HClO₄, and NTA [Nitrilotriacetic acid]), solid/liquid ratio, leaching temperature, and leaching time. The highest Fe₂O₃ removal was 86.6%, and a 11.8 ppm Fe₂O₃ content quartz product with a whiteness index (WI) value of 90.6 was obtained after 120 min of treatment at 90°C with a 10% S/L ratio and 1 M H₂SO₄.     

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).     

Selective Leaching Process for Neodymium Recovery from Scrap Nd-Fe-B Magnet

Neodymium-iron-boron (Nd-Fe-B) magnets were most widely applied to permanent magnetic products in the world due to their high magnetic force. The increasing growth of scrap Nd-Fe-B magnets resulted in disposal problems and the reduction of neodymium (Nd) valuable resources. In this study, we developed a simple hydrometallurgical precipitation process with pH adjustment to separate and recover Nd 100 pct recovery from scrap Nd-Fe-B magnets. Several physical and chemical methods such as demagnetization, grinding, screening, and leaching processes were also adopted to investigate the recovery of Nd and other metals from scrap Nd-Fe-B magnets. The leaching process was carried out with four leaching reagents such as NaOH, HCl, HNO₃, and H₂SO₄. Batch studies were also conducted to optimize the leaching operating conditions with respect to leaching time, concentration of leaching reagent, temperature, and solid/liquid ratio for both HCl and H₂SO₄ leaching reagents. Nd was successfully separated and recovered with 75.41 wt pct from optimized H₂SO₄ leaching solution through precipitation. Further, the purity and weight percentage of the obtained Nd product was analyzed using scanning electron microscopy–energy-dispersive spectroscopy (SEM-EDS) analysis. An X-ray diffraction (XRD) study confirmed the obtained product of Nd was in the form of NdOOH and Nd(OH)₃.     

返酸浸出磷矿中的稀土

采用湿法磷酸工艺的返回酸浸出磷矿中伴生稀土与磷,稀土与磷以磷酸二氢盐的形式进入溶液。考察温度、浓度、时间、液固比等因素对稀土与磷浸出的影响。结果表明,适宜的浸出条件为:返回酸浓度25%、温度65℃、液固比10∶1、浸出时间8h,在此条件下,稀土浸出率可达90.4%,磷浸出率达98.7%。     

Dissolution studies of mechanically activated Manavalakurichi ilmenite with HCl and H2SO4

The effect of the change in phase constitution, particle size distribution, surface area, crystallite size, strain and lattice parameters introduced by mechanical activation of the altered beach sand ilmenite from Manavalakurichi region, India on the dissolution kinetics of HCl and H₂SO₄ was investigated. The altered ilmenite showed different physico-chemical characteristics and was found to be more resistant to acid leaching than the less altered ilmenite from the Chatrapur beach sands, India investigated earlier. The dissolution behavior was also different in H₂SO₄ and HCl. For sulfuric acid leaching, the dissolution of Fe and Ti increased monotonically with time of milling and showed a continuous increase with time of leaching, whereas hydrolysis of titanium occurs in HCl medium, especially for the activated samples at lower acid concentration, lower solid to liquid ratio and higher temperature leading to lower solution recoveries. The dissolution kinetics in both H₂SO₄ and HCl prior to hydrolysis conforms initially to the reaction rate control model and for higher leaching times to the shrinking core model where diffusion through the product layer is rate controlling. It is postulated that the anatase formed by hydrolysis in milled samples impedes the further progress of leaching. The activation energies for the dissolution of Fe and Ti decreased with time of milling and were marginally lower in HCl than in H₂SO₄. An attempt has also been made to correlate the decrease in activation energy to the increase in the energy input to the material through high-energy milling. The relative contribution of the increase in surface area and structural disorder on the enhancement of the dissolution rates has been evaluated.     

Uranium and Rare Earth Recovery from Phosphate Fertilizer Industry by Solvent Extraction

Abstract

The paper presents a comprehensive review of the fundamental and applied aspects of the recovery of uranium and rare earths from phosphoric acid used for fertiliser production, by solvent extraction, with emphasis on the data from the pilot plant operated in Romania. The occurrence of uranium and rare earths in various phosphatic deposits is described and variations discussed. The radioactivity of the phosphatic rock is shown to be a cause for concern for control over pollution. History of the commercial recovery operations is decribed and the role of economic factor highlighted. The acidulation of phosphate rock, behavior of uranium and rare earths, influence of iron and effect of humic matter in the sulphuric acid process and the nitric acid process is discussed- Uranium extraction from the phosphoric acid by acidic organophosphorus esters is discussed in detail. The industrial flowsheets are described. Details are presented for the single cycle extraction process developed in Romania. Cost estimates are given which show the superiority of the single cycle process and its capability to produce uranium at a competitive price, while also producing rare earths as a by-product. Process flowsheet with stream data is also given for treatment of the solution obtained by nitric acid attack of the phosphate rock. Analysis of the uranium green cake (UF₄) and the rare earth concentrate is given.     

Extraction of rare earth ions from phosphate leach solution using emulsion liquid membrane in concentrated nitric acid medium

It is very significant to recover rare earths (REs) from wet-process phosphoric acid, in terms of extraction rate and selectivity, the current carrier di(2-ethlhexly) phosphate (D2EHPA) out there is still inferior. Based on this question, our team modified D2EHPA to synthesize new extractants. This paper presents a comprehensive study on the extraction of rare earth ions (RE³⁺) from phosphate leach solution using emulsion liquid membrane (ELM) in concentrated nitric acid medium. The ELM system is made up of (RO)₂P(O)OPh-COOH as carrier, polyisocrotyl succinimide (T154) as surfactant, sulfonated kerosene as diluent, phosphoric acid (H₃PO₄) as stripping solution. Different chemical parameters such as type and concentration of carrier, surfactant, stripping solution, volume ratio of oil phase to internal phase, and volume ratio of emulsion ratio to external phase were analyzed. The extraction of RE³⁺ was evaluated by the yield of extraction. In addition, the demulsification process was also investigated. The proposed method of ELM using (RO)₂P(O)OPh-COOH as carrier can be expected to provide an efficient, simplify operation, and facilitated method for extracting RE³⁺.     

磷酸法中稀土溶出的动力学模型研究

利用磷酸浸出伴生稀土磷矿,研究了温度、时间、酸度、粒度与稀土浸出动力学的关系。结果表明,升高温度、延长反应时间、增大酸度、减小初始粒度均有利于稀土的浸出。界面化学反应控制为稀土浸出的控制步骤,表观反应级数0.993,表观活化能44.6kJ/mol,稀土浸出过程近似为基元反应。按反应核收缩核模型通过统计分析建立了关联温度、浓度、粒度三因素的稀土浸出数学模型。     

Reductive leaching of manganese from low-grade manganese ore in H2SO4 using cane molasses as reductant

Manganese extraction from a low-grade ore was investigated using cane molasses as a reducing agent in dilute sulfuric acid medium. The effects of concentrations of cane molasses and sulfuric acid, leaching temperature as well as reaction time were discussed. The results showed that high manganese recovery with low Fe and Al extraction yield could be obtained by analyzing the leaching efficiencies of Mn, Fe and Al during the leaching process. The optimal leaching condition was determined as 1.9 mol/L H₂SO₄ and 60.0 g/L cane molasses for 120 min at 90 °C while using particles smaller than 0.147 mm. The leaching efficiencies were 97.0% for Mn, whereas 21.5% for Al and 32.4% for Fe, respectively.     

Leaching of copper(I) sulphide by sulphuric acid solution with addition of sodium nitrate

Finely grained samples of copper(I) sulphide were leached by H₂SO₄ solution with added NaNO₃. The occurrence probability of chemical reactions was analysed based on literature data and products which were formed during the process and the overall leaching reaction was defined. The effect of temperature, concentration of NaNO₃ and H₂SO₄, stirring speed, phase ratio and time, on the leaching degree of copper was studied. The quantity of copper dissolved increases with growth of the values of all the parameters. Kinetic analysis shows that the leaching mechanism is very complex. By using appropriate mathematical kinetic models, it is found that the leaching rate is chemically controlled. It was concluded that the leaching reaction is first order with respect to the concentration of NaNO₃ and second order with respect to the concentration of H₂SO₄.     

Effects of iron and temperature on solubility of light rare earth sulfates in multicomponent system of Fe2(SO4)3-H3PO4-H2SO4 synthetic solution

Numerous light rare earth elements (LREE) minerals containing Fe and P were processed by sulfuric acid roasting method, and the leaching solution mainly comprises LREE sulfate, Fe₂(SO₄)₃, H₃PO₄, and H₂SO₄, however, the solubility data of LREE sulfates in this system is few. This work studies the solubility of LREE sulfates in independent LREE sulfate system RE₂(SO₄)₃-Fe₂(SO₄)₃-H₃PO₄-H₂SO₄ (RE = La, Ce, Pr or Nd) and mixed LREE sulfates system (La,Ce,Pr,Nd)₂(SO₄)₃-Fe₂(SO₄)₃-H₃PO₄-H₂SO₄ at different temperature (25–65 °C) and concentrations of Fe₂(SO₄)₃ (Fe₂O₃, 0–50.13 g/L), H₂SO₄ (0.5 mol/L), and H₃PO₄ (P₂O₅, 20.34 g/L) based on the industrial operating condition at low liquid and solid ratio 2:1. The solubility of each LREE sulfate in the independent system (La₂O₃, 12.25–20.88 g/L; CeO₂, 41.93–62.35 g/L; Pr₆O₁₁, 37.34–56.69 g/L; Nd₂O₃, 26.60–37.63 g/L) is much higher than that of the mixed system (La₂O₃, 6.95–11.03 g/L; CeO₂, 10.63–21.51 g/L; Pr₆O₁₁, 11.56–20.36 g/L; Nd₂O₃, 12.36–19.79 g/L) under the same other conditions. The results also indicate that, in the two systems, both Fe and the temperature have negative effects on the solubility of LREE sulfates. That may occur due to the complication reactions between the complexes of RESO₄⁺ and Fe(SO₄)₂^–. However, the influence degree of temperature and iron concentration on the LREE sulfates solubility varies in the two systems and among different LREE species. This research is of theoretical significance for optimizing the conditions of the sulfuric acid process for recovering the LREE from the mixed LREE bearing minerals as well as the single LREE containing secondary rare earth scraps.