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Leaching of lepidolite and recovery of lithium hydroxide from purified alkaline pressure leach liquor by pHosphate precipitation and lime addition
Abstract
This study investigated the pressure leaching of lithium and other valuable metals from lepidolite using NaOH and Ca(OH)2. The findings showed that NaOH concentration, temperature, and stirring rate are the most significant process parameters. The addition of Ca(OH)2 facilitates the leaching of lithium and minimizes the concentration of silicate and fluoride in the leach liquor by forming solid phases identified as Ca3Al2Si3O12 CaF2 and NaCaHSiO4, with a loss of about 7% NaOH. The leaching efficiencies after 2 h were 94% Li, 98% K, 96% Rb, and 90% Cs under the most suitable conditions (250 °C, 320 g/L NaOH, liquid-to-solid ratio of 10, 300 rpm stirring speed, lime addition of 0.3 g/g). Precipitation of lithium as Li3PO4 by adding phosphoric acid and subsequent conversion of Li3PO4 to LiOH using Ca(OH)2 recovered 83% of lithium as LiOH allowing recycling the rest along with 93% of NaOH for leaching.
... This approach aims to diminish leaching impurities and enhance the leaching rate of lithium ions. James Mulwanda et al. [15] conducted a study on the pressurized leaching of lithium and other valuable metals from lithium mica using NaOH and Ca(OH) 2 . They found that under optimum conditions (250 • C, 320 g·L −1 NaOH, 10 mL·g −1 liquid to solid ratio, 300 rpm/min stirring speed, and 0.3 g·g −1 lime addition) with a 2 h reaction time, the leaching rates were as follows: Li + 94%, K + 98%, Rb 2+ 96%, and Cs + 90%. ...
... This approach aims to diminish leaching impurities and enhance the leaching rate of lithium ions. James Mulwanda et al. [15] conducted a study on the pressurized leaching of lithium and other valuable metals from lithium mica using NaOH and Ca(OH)2. They found that under optimum conditions (250 °C, 320 g·L −1 NaOH, 10 mL·g −1 liquid to solid ratio, 300 rpm/min stirring speed, and 0.3 g·g −1 lime addition) with a 2 h reaction time, the leaching rates were as follows: Li + 94%, K + 98%, Rb 2+ 96%, and Cs + 90%. ...
... They then converted the Li3PO4 to LiOH with Ca (OH)2, recovering 83% of the lithium in the form of LiOH. Figure 4 below shows the flow chart for this study. [15]. Reproduced with permission from Elsevier, 2021. ...
Ion Imprinting Technology (IIT) is an innovative technique that produces Ion-Imprinted polymers (IIPs) capable of selectively extracting ions. IIPs exhibit strong specificity, excellent stability, and high practicality. Due to their superior characteristics, the application of IIPs for lithium resource extraction has garnered significant attention. This paper discusses the following aspects based on existing conventional processes for lithium extraction and the latest research progress in lithium IIPs: (1) a detailed exposition of existing lithium extraction processes, including comparisons and summaries; (2) classification, comparison, and summarization of the latest lithium IIPs based on different material types and methods; (3) summarization of the applications of various lithium IIPs, along with a brief description of future directions in the development of lithium IIP applications. Finally, the prospects for targeted recovery of lithium resources using lithium IIPs are presented.
... Lithium-bearing lepidolite (K(Li,Al) 3 (Al,Si) 4 O 10 (F,OH) 2 ) mineral resource is highly abundant and contains varying Li content (% w/w) between 2.79 and 3.73%, thus a promising resource for lithium battery chemicals (Chagnes and Swiatowska, 2015;Garrett, 2004). At present, most of the published processes involving lithium production from lepidolite comprise direct leaching with acids or alkaline media or roasting with additives followed by water leaching (Mulwanda et al., 2020;Chagnes and Swiatowska, 2015;Yan et al., 2012b;Luong et al., 2013;Wietelmann and Bauer, 2000;Rosales et al., 2014;Martin et al., 2017;Zhang et al., 2019). A summary of lepidolite processing methods is reported elsewhere (Mulwanda et al., 2020). ...
... At present, most of the published processes involving lithium production from lepidolite comprise direct leaching with acids or alkaline media or roasting with additives followed by water leaching (Mulwanda et al., 2020;Chagnes and Swiatowska, 2015;Yan et al., 2012b;Luong et al., 2013;Wietelmann and Bauer, 2000;Rosales et al., 2014;Martin et al., 2017;Zhang et al., 2019). A summary of lepidolite processing methods is reported elsewhere (Mulwanda et al., 2020). The leach solution can be purified by adjusting the pH using NaOH to remove impurities such as Fe, Al, Mg, and Ca. ...
... However, the reagent consumption in the proposed alkaline pressure leaching method remains high (Catovic, 2018). Pressure leaching of lepidolite with NaOH in the presence of CaO and subsequent precipitation of Li 3 PO 4 by adding H 3 PO 4 allows the recycling of NaOH to leaching (Mulwanda et al., 2020). Other proposed methods involve roasting of lepidolite (750 • C to 1000 • C) with chemical additives (Luong et al., 2014;Yan et al., 2012aYan et al., , 2012bYan et al., , 2012c, followed by water leaching under atmospheric conditions to selectively extract lithium. ...
... These findings have encouraged authors to explore approaches that utilize the entire inventory of lithium minerals such as lepidolite [20], spodumene [21], or petalite [22] by alkaline leaching in high pressure autoclaves. It is noteworthy that these processes are carried out without thermal phase transformation under strongly alkaline conditions, usually hydrothermally in an autoclave, to decompose the silicates and enrich lithium in the solution [22][23][24]. In addition, Lv et al. [20] and Xing et al. [21] employed the alkaline treatment for the parallel synthesis of zeolites, while Qui et al. [25] produced KAlSiO 4 as a zeolite precursor as value-added by-products through the use of Al and Si derived from lithium silicate minerals. ...
... Precipitation and conversion of lithium compounds were achieved following the procedure of Mulwanda et al. [23]. During precipitation, small amounts of phosphoric acid (H 3 PO 4 , 85%, Carl Roth, analytical grade) were added through a glass cannula to the preheated caustic solution in a round bottom flask to produce lithium phosphate (Li 3 PO 4 ). ...
... Mechanochemical treatment combining ball milling of lithium silicates with alkaline leaching were investigated since it is known from the literature that alkaline solutions are able to extract lithium from aluminosilicate minerals such as lepidolite [20,23], spodumene [21,24], and petalite [22] by decomposing their silicate structure. Moreover, they are widely used for alkaline activation in the synthesis of zeolites [32]. ...
Lithium is in high demand: this is driven by current trends in e-mobility and results in increased global production and record prices for lithium ores and compounds. Pegmatite ores, in addition to brines, remain of particular interest because of their higher lithium content and lower geopolitical risks. In this work, we investigated lithium extraction via the mechanochemical treatment of the three most common lithium minerals: lepidolite, spodumene, and petalite. Indeed, we determine that the petalite crystal structure was much more suitable due to its less dense packing and the formation of cleavage planes along lithium sites, resulting in substantial lithium extraction of 84.9% and almost complete conversion to hydrosodalite after 120 min of ball milling in alkaline media. Further processing of the leach liquor includes desilication, the precipitation of lithium phosphate, and the conversion and crystallization of pure LiOH·H 2 O. Special attention was paid to a holistic approach entailing the generation of by-products, each of which has a specific intended application. The leaching residues were investigated by powder X-ray diffraction, Fourier transform infrared spectroscopy, N 2 adsorption/desorption, and scanning electron microscopy. Moreover, hydrosodalite was found to have a high potential as an adsorbent for heavy metal ions which were studied separately using aqueous solutions containing Cu 2+ , Ni 2+ , Pb 2+ , and Zn 2+ .
... (1) (Song and Zhao, 2018;Yanagase et al., 1983;Xiao and Zeng, 2018;Mulwanda et al., 2021;Shin et al., 2022;Sun et al., 2019) Lithium recovery from brine: Recent developments and challenges, Desalination, 528, p.115611. ...
... 대표적인 DLE 기술로는 용매추출법Shi et al., 2020;Meshram et al., 2014), 침전법(Yanagase et al., 1983;Xiao and Zeng, 2018;Yang et al., 2018;Mulwanda et al., 2021),흡착법(Abe and Chitrakar, 1987;Miyai et al., 1994;Ryu et al., 2013; Hong et al., 2018;Park et al., 2014;Xiao et al., 2015) 등이 검토되고 있으며 이외 에도 막여과(Gong et al., 2018; Guo et al., 2016), 전기흡착(Nie et al., 2017; Jiang et al., 2014;Shi et al., 2019; Kanoh et al., 1991, Ryu et al., 2015a 등의 기술들을 활용한 리튬 회수기술 개발이 진행되고 있다. 용매추출법 용매추출법의 경우 추출제의 리튬이온에 대한 선택도가 낮기 때문에 염호에 함유된 +2/+3가 이온을 분리하기 위해 알칼리 용제를 사용해 불순물을 분리해야 하는 전처리 과 정이 요구된다. ...
... The author reports a conversion higher than 99%. Mulwanda et al. (2021) showed a process (Fig. 17), which produced LiOH from purified alkaline pressure leach liquor by phosphate precipitation and subsequent lime addition. The authors used an industrial lepidolite feed and added Ca(OH) 2 and NaOH at 250 • C for two hours at 36 bar. ...
... Process schemes(Mulwanda et al., 2021;Lee et al., 2019). ...
... The Li + could be recovered as Li 3 PO 4 or Li 2 CO 3 . In both cases, the pH is raised above 12 with NaOH, and then H 3 PO 4 (Mulwanda et al., 2021) is added for the first case or carbonated with CO 2(g) for the second (Resentera et al., 2023a(Resentera et al., , 2023bRosales et al., 2014). The products (Li 2 CO 3 or Li 3 PO 4 ) are Li salts used industrially for Li-ion battery manufacture. ...
Numerous Li extraction methods from minerals and e-waste have been reported in the literature. Among them, direct fluorination processes appear to be a viable alternative due to their high lithium extraction efficiencies (>90%) as LiF. However, a drawback is the low water solubility of LiF, which requires acids for its separation and to obtain other commercial lithium salts. An interesting alternative for dissolving salts with low solubility is through the formation of coordination complexes. In this case, aluminum forms highly stable soluble complexes with the F− anion, such as AlF2+, AlF2+, AlF3, AlF4−, AlF52−, AlF63−.
This study proposes an acid-free LiF dissolution methodology using aluminum sulfate as a leaching agent. The LiF dissolution was modeled and optimized using Response Surface Methodology (RSM). The investigated operating parameters for LiF dissolution were the solid/liquid ratio (A), reaction temperature (B), and leaching time (C). Thus, a predictive mathematical model was successfully optimized (R2 = 0.9445). The results indicated that the S/L ratio negatively influences the dissolution of LiF, while temperature and time have a positive effect. The LiF dissolutions of 90 ± 3% were achieved with a leaching time of 31 min, a S/L ratio of 20 g/mL, and a temperature of 45 °C.
... Lepidolite is widely recognized as one of the primary lithium (Li) sources for Li extraction (Zhang et al., 2019). The process of Li extraction in the form of lithium carbonate from lepidolite mainly involves direct leaching in acidic or alkaline medium, or water leaching after roasting with salt assistance (Mulwanda et al., 2023;Mulwanda et al., 2021). Fluoride-bearing wastewater was inevitably produced from hydrometallurgical process of Li extraction, leading to the surrounding environmental risk. ...
Fluoride contamination is a serious environmental problem in lepidolite hydrometallurgy wastewater. The treatment of the fluoride-bearing wastewater is facing a great challenge due to the presence of co-existed ions including lithium (Li ⁺ ), rubidium (Rb ⁺ ), silicate (SiO 3 ²⁻ ), sulfate radical (SO 4 ²⁻ ) and so on. However, aluminum-modified zeolite (Al@zeolite) of sufficient hydroxyl group and high adaptability have exhibited unique advantages for fluoride elimination from lepidolite hydrometallurgy wastewater. Al@zeolite was prepared on natural zeolite by an atmospheric process, and it was then used for the adsorption of fluorine from fluoride bearing wastewater produced by lepidolite hydrometallurgy process. The results of material characterization confirmed the successful immobilization of aluminum within the zeolite pores, and indicating the formation of zeolite-Al-OH. The zeolite host significantly enhanced the chemical stability of Al@zeolite against pH changes with a wide pH of 2.0–10.0. The adsorbent possessed a surface area of 33.46 m ² /g and demonstrated excellent capacity and selectivity for fluoride adsorption. It is pertinent to mention that maximum adsorption of 98.6% has been observed with a pH value of 6.0 with 20 min time duration with fluoride dosage of 20 mg/L, and the equilibrium concentration dwindled to 0.4 mg/L. The results of fluorine adsorption showed that fluoride uptake onto Al@zeolite agreed well with the pseudo-first-order kinetic model and the Langmuir isotherm model. The reusability of the substance was evaluated for up to eight cycles following consecutive regeneration with 0.2 mol/L AlCl 3 . The exhausted Al@zeolite was effectively regenerated through a simple alkaline treatment for recycling. The above results verified that Al@zeolite is a new kind of efficient defluoridation adsorbent for lepidolite hydrometallurgy wastewater with practical application prospect.
... (90% ) [22][23][24][25] . Fig. 2 (Rb 2 O ) 7,[17][18][19][20][21][22][23][24][25][26][27][28][29][30][31][32][33][34][35][36] . ...
... The primary extraction technologies for valuable metals in lepidolite include limestone roasting method [8,9], sulfuric acid method [10,11], autoclave method [12,13], sulfate roasting method, and chlorination roasting method [14][15][16][17]. Their characteristics are shown in Table 1. ...
... Mulwanda, Senanayake [107] recovered lithium from an alkaline leach solution using phosphoric acid (H3PO4). With the purpose of increasing the efficiency of lithium extraction as Li3PO4, the study alternatively used H3PO4 instead of Na3PO4 to minimize the presence of sodium ions (Na + ) in the alkaline liquor. ...
Lithium is a vital raw material used for a wide range of applications, such as the fabrication of glass, ceramics, pharmaceuticals, and batteries for electric cars. The accelerating electrification transition and the global commitment to decarbonization have caused an increasing demand for lithium. The current supply derived from brines and hard rock ores is not enough to meet the global demand unless alternate resources and efficient techniques to recover this valuable metal are implemented. In the past few decades, several approaches have been studied to extract lithium from aqueous resources. Among those studied, chemical precipitation is considered the most efficient technology for the extraction of metals from wastewater. This paper outlines the current technology, its challenges, and its environmental impacts. Moreover, it reviews alternative approaches to recover lithium via chemical precipitation, and systematically studies the effects of different operating conditions on the lithium precipitation rate. In addition, the biggest challenges of the most recent studies are discussed, along with implications for future innovation.
... The leach liquor contained 1744 mg/L of Li, and Li 2 CO 3 with 99.9% purity could be produced after purification. Meanwhile, Li can be extracted using NaOH and Ca(OH) 2 pressurized leaching of lepidolite (Mulwanda et al., 2021). Here, Ca(OH) 2 can promote the leaching of Li and minimize the concentration of silicate and fluoride in the leachates by forming a solid phase. ...
... The concentration of NaOH had a significant effect and the XRD pattern described that the leaching residue was sodalite with high purity under optimal conditions. Zeolite NaA was successfully prepared by hydrothermal synthesis, the zeolite had adsorption properties for Pb 2+ and Cd 2+ , and the maximum adsorption capacities were 487.8 and 193.8 mg/g, respectively [89]. Mulwanda et al. [90] combined the above two processes, using NaOH and Ca(OH) 2 as co-leaching agents, as shown in reaction (47). ...
Lithium is considered to be the most important energy metal of the 21st century. Because of the development trend of global electrification, the consumption of lithium has increased significantly over the last decade, and it is foreseeable that its demand will continue to increase for a long time. Limited by the total amount of lithium on the market, lithium extraction from natural resources is still the first choice for the rapid development of emerging industries. This paper reviews the recent technological developments in the extraction of lithium from natural resources. Existing methods are summarized by the main resources, such as spodumene, lepidolite, and brine. The advantages and disadvantages of each method are compared. Finally, reasonable suggestions are proposed for the development of lithium extraction from natural resources based on the understanding of existing methods. This review provides a reference for the research, development, optimization, and industrial application of future processes.
... Furthermore, some researchers have proposed to extract lithium directly by using a NaOH solution at 250°C, 18,19 but it is difficult to obtain Li 2 CO 3 with high purity and high yield from the leachate with a relatively high sodium content using a conventional carbonation process. 20 Therefore, it is of great significance to develop a new technology to extract lithium from α-spodumene. In previous work, we found that α-spodumene could be directly decomposed by a NaOH solution, and the Li element in αspodumene was converted into acid-soluble intermediate solidphase product Li 2 SiO 3 under certain conditions (500 rpm, 250°C ...
... Existen publicaciones y hallazgos relevantes sobre la disolución de β-espodumeno en una autoclave. A temperaturas alrededor de 523 K con sales como Na2CO3 y NaCl (Mulwanda et al. 2021). ...
El mineral de litio de Macusani Puno Perú, se ha estimado un recurso de 4.7 millones de toneladas como carbonato de litio con una ley de 3500 ppm de Li. Tiene como objetivo obtener carbonato de litio mediante proceso de lixiviación, concentración, evaporación y precipitación. El tratamiento del mineral litio, se inició con las operaciones de chancado y molienda, en esta etapa se utilizó una chancadora de laboratorio y un pulverizador de anillos, el mineral fue molido hasta 87 % malla -200. Se obtuvo lixiviado de Li2SO4 mediante el proceso por tostación de mineral de litio con ácido sulfúrico concentrado a temperatura de 250 °C, el rendimiento de extracción alcanzó al 93 % de Li, así como la disolución del mineral con fluoruro de sodio y ácido sulfúrico en autoclave a temperatura de 125 ºC y alta presión (0,2 MPa) por 3 horas, el porcentaje de extracción de litio alcanzó al 92 %. En la segunda etapa se purificó los lixiviados de litio con cal, las impurezas fueron removidas al 99,92 %, se incrementó la concentración de calcio debido al uso de cal, la presencia de este elemento fue eliminado por precipitación con oxalato de amonio. Luego de la purificación la solución se encuentra libre de impurezas, en seguida de concentró el litio por evaporación de 1,2 g/L hasta 18,7 g/L de Li. El concentrado de litio se precipitó con carbonato de sodio a temperatura de 95 °C, obteniéndose el carbonato de litio con una pureza de 98,80 %. Se utilizaron tres diseños experimentales. Se obtuvo carbonato de litio de grado técnico mediante procesos de tratamiento químico-metalúrgicos.
... The XRD analyses in Fig. 3 shows that the quartz, albite, and muscovite decomposed completely under any stirring speeds as expressed in Eq. (4)-(6), while the diffraction peaks of α-spodumene gradually weakened and disappeared by 900 rpm. At the same time, a new phase of lithium silicate (Li 2 SiO 3 , PDF#29-0828, orthorhombic, Cmc2 1 ) appeared (Mulwanda et al., 2021;Xing et al., 2019), and sodium aluminosilicate hydrates with different compositions and structures became the dominant phases in the residues after hydrothermal treatment. According to the analysis by MDI Jade 6, the forms of sodium aluminosilicate hydrates were sodalite (Na 8 Al 6 Si 6 O 24 (OH) 2 (H 2 O) 2 , PDF#76-1639), faujasite-Na (Na 14 Al 12 Si 13 O 51 ⋅6H 2 O, PDF#28-1036), cancrinite (3NaAlSiO 4 ⋅NaOH, PDF#15-0734), hydroxycancrinite (Na 8 Al 6 Si 6 O 24 (OH) 2 ⋅2H 2 O, PDF#46-1457) and sodium aluminum silicate hydrate ((Na 2 O) 1.31 Al 2 O 3 (SiO 2 ) 2.01 (H 2 O) 1.65 , PDF#75-2318), etc., which were similar to the desilication products (DSPs) precipitating during the Bayer digestion process (Pan et al., 2016;Zeng and Li, 2012). ...
The key to the lithium extraction from α-spodumene is destroying the crystal structure. The conventional commercial process requires high-temperature calcination to transform α-spodumene into β-spodumene. This study aimed at finding an alternative treatment for destroying the structure of α-spodumene and extracting the lithium. The results indicated that α-spodumene could be directly decomposed by NaOH solution and the Li element in α-spodumene was converted into intermediate product Li2SiO3 in the form of solid phase under certain conditions. According to the experiments, the possible reactions during the hydrothermal alkaline treatment were predicted, which were thermodynamically possible by the calculation of Gibbs free energy using HSC Chemistry 10. The total extraction efficiency of Li2O was 87.3%, including the extraction efficiency of Li2O leached into liquid phase, 7.6%, and converted into Li2SiO3, 79.7%, obtained under optimal conditions: stirring speed of 500 rpm, leaching temperature of 250 °C, mass ratio of NaOH/ore of 1.5, initial NaOH concentration of 25 wt%, and leaching time of 24 h. Then the Li element in the Li2SiO3 could be extracted by acid leaching and precipitated using Na2CO3. The mother liquor obtained after hydrothermal alkaline treatment was reused for the subsequent hydrothermal alkaline cyclic leaching under the above conditions. During three cycles of alkaline treatment, the α-spodumene decomposition was stable, and the total Li2O extraction efficiency was around 86% (∼84% was converted into Li2SiO3). This process avoids high-temperature calcination and concentrated acid roasting, where the acid leaching process is carried out at room temperature.
... The existing methods for treating lepidolite can be classified into acid digestion (Gao et al., 2020;Guo et al., 2019;Guo et al., 2021;Liu et al., 2019;Liu et al., 2020;Vieceli et al., 2017;Zhang et al., 2019), chlorination roasting (Yan et al., 2012a;Yan et al., 2012b;Zhang et al., 2020a), high pressure leaching (Mulwanda et al., 2021;Yan et al., 2012d) and sulfation roasting (Hien-Dinh et al., 2015;Luong et al., 2014;Luong et al., 2013;Su et al., 2020;Vieceli et al., 2016;Yan et al., 2012c). Table 1 summarizes the reaction conditions and extraction efficiencies of different methods. ...
Lepidolite is an important mineral resource containing lithium, rubidium, cesium and potassium. Most researches focused on the extraction of lithium from lepidolite, while the co-extraction of rubidium, cesium and potassium was ignored. In this study, the extraction of lepidolite at lower temperature was investigated by using iron(II) sulfates feedstock, aiming at enhancing the extraction of lepidolite, especially for rubidium and cesium. Thermodynamic predictions indicated that the co-extraction of lithium, rubidium, cesium and potassium from lepidolite by roasting with iron(II) sulfate was feasible, but high temperature worked against the extraction. The effects of process parameters on the extraction were investigated systematically. It was found that the optimal roasting conditions were as follows: 50% lepidolite particles <74 μm, temperature of 675 °C, iron(II) sulfate to lepidolite mass ratio of 2:1, and holding time of 90 min. The extraction efficiencies of Li, Rb, Cs and K were 92.7%, 87.1%, 82.6% and 86.2%, respectively. The reaction between lepidolite concentrate and iron(II) sulfate followed two pathways, i.e. gas-solid and liquid-solid reaction. The formation of pyrosulfates with low smelting point promoted the reaction by reacting with lepidolite through a circular transformation between sulfates and pyrosulfates. Compared with the conventional extraction of lepidolite, the proposed process operates at a lower temperature and co-extracts various metals, exhibiting great potential in industrial application.
... 6b , 7 Various researchers are exploring alternative processes to overcome the mentioned challenges in extraction of Li from minerals. 7,8 One of these is the use of potassium sulfate for reaction with α-spodumene at elevated temperature. 9,10 This process was first developed and patented by Wadman. ...
The conventional process of lithium extraction from α-spodumene (LiAlSi 2 O 6) is energy-intensive and associated with high byproduct management cost. Here, we investigate an alternative process route that uses potassium sulfate (K 2 SO 4) to extract lithium while producing leucite (KAlSi 2 O 6), a slow release fertilizer. Presenting the first-ever in situ record of the reaction of α-spodumene with potassium sulfate, we use synchrotron X-ray diffraction (XRD) and differential scanning calorimetry (DSC) to document the reaction sequence during prograde heating. From 780°C, we observe a broad endothermic DSC peak, abnormal expansion of the α-spodumene structure, and an increase in α-(Li, K)-spodumene peak intensity during heating with potassium sulfate, indicative of the exchange between lithium and potassium in the spodumene structure. When 11 ± 1% K occupancy in the M2 site of α-(Li, K)-spodumene is reached, the mechanism changes from ion exchange to a reconstructive transformation of α-(Li, K)-spodumene into leucite, evidenced by a decrease in α-spodumene and potassium sulfate abundance concurring with formation of leucite over a narrow temperature range between 850 and 890°C. The increasing background intensity in synchrotron XRD above 870°C suggests that a lithium sulfate-bearing melt starts to form once >90% of α-spodumene has been converted during the reaction. This fundamental understanding of the reaction between α-spodumene and potassium sulfate will enable future development of lithium extraction routes using additives to significantly decrease energy intensity and to produce marketable byproducts from α-spodumene.
Demand for lithium‐ion batteries for use in electric vehicles is driving lithium demand. A large increase in demand between now and 2040 is expected due to increasing electrification of the transport and energy sectors of the economy. This is in part due to an interest in improving air quality in cities and in the reduction of greenhouse gas emissions when electricity is generated without fossil fuel combustion. Currently, the two major natural sources of lithium for use in lithium‐ion batteries are brines and hard‐rock minerals. Lithium‐bearing clays have also been investigated as a potential source of lithium. Battery recycling is another potential source. Lithium in hard‐rock minerals is predominantly found in pegmatites. Lithium‐bearing minerals that may occur in pegmatites include petalite, lepidolite, and spodumene. The latter is currently the predominant hard‐rock lithium source. Following comminution, a number of treatments or combinations of treatments can be used to produce materials suitable for the final stage of purification of lithium products. Most involve high‐ or very‐high‐temperature treatments with various reagents followed by water leaching, although a few treat lithium‐bearing feed directly with acid leaching. Other elements are co‐leached and must be separated from lithium to allow the production of high‐purity lithium carbonate or lithium hydroxide, which are then used to manufacture lithium‐ion battery cathodes.
A novel alkaline hydrothermal approach for low‐temperature conversion of α‐spodumene into Li2SiO3 residue was proposed, providing a promising method for extracting lithium from α‐spodumene as a pretreatment process. This work proposed a systematical investigation for extracting lithium from the residue by acid leaching and preparing lithium carbonate. The reaction feasibility between Li2SiO3 and acids (HCl and H2SO4) was firstly evaluated through thermodynamic calculation. Compared with the leaching effect of hydrochloric acid and sulfuric acid, sulfuric acid is the preferred leaching agent due to its higher extraction efficiency of lithium and less acid consumption. Lithium extraction efficiency from the residue achieved up to 87.48% under the optimized conditions: 0.75 mol/L H2SO4, 0.4 times the theoretical amount of acid, 10 min, 30°C, 100 rpm. Based on the optimized conditions, the lithium‐containing solution was concentrated through 3 consecutive cycles of leaching, which obtained a concentration of 17.78 g/L for lithium. The leaching solution was purified by CaO‐Na2CO3, resulting in the removal rates of SiO32‐, Mg2+, and Ca2+ of 84.22%, 95.51%, and 90.55%, respectively. Finally, the solution was precipitated with sodium carbonate to prepare Li2CO3. This paper facilitates the development of an economical process for efficient lithium extraction from spodumene at low temperatures. This article is protected by copyright. All rights reserved.
To supplement the increasing demand for lithium (Li) resources dictated by the growing market for lithium-ion batteries (LIBs), this study aimed to secure lithium carbonate (Li2CO3) using a waste box sagger generated from the process of manufacturing cathode materials for LIBs. Lithium was recovered through sulfation and subsequent wet conversion into high-purity Li2CO3. Lithium extractability (93%) from the waste box sagger powder was determined by the sulfation reaction using 5 M sulfuric acid (H2SO4) and distilled water (H2O) leaching of the resultant. The precipitation efficiency (85%) of Li into lithium phosphate (Li3PO4) in extracted Li solution was established using a sodium hydroxide (NaOH) and phosphoric acid (H3PO4) mixture. The maximum conversion efficiency (99%) of Li3PO4 to lithium chloride (LiCl) solution was calculated using a reflux method with calcium chloride (CaCl2) solution at a solid (g)/liquid (L) ratio of 100. The highest Li⁺ concentration reached 50,800 mg/L at a solid/liquid ratio of 400 and Cl/Li molar ratio of 0.95 in the range of tested conditions. Finally, high-purity Li2CO3 (99.8% in metal basis) was produced from the obtained Li solution through purification using NaOH, followed by carbonation using sodium carbonate (Na2CO3). The proposed method demonstrates the feasibility of securing a Li resource by recycling waste box sagger instead of disposing them as a designated waste.
Lithium has become a strategic element, as it plays
a crucial role in battery technology. Therefore, many efforts are
being made to extract lithium from primary sources, while lithium
recycling is almost nonexistent. In this paper, we report on a novel
alkaline mechanochemical approach for the recycling of end-of-life
glass−ceramics, which involves lithium extraction, zeolite synthesis,
desilication, and lithium precipitation. Special attention has
been paid to a holistic approach, with each byproduct having a
designated application. Optimal parameters for lithium extraction
such as sodium hydroxide concentration, rotational speed, and
ball-to-powder ratio were achieved at 7 mol/L, 600 rpm, and 50:1
g/g, respectively, resulting in high yields of 83.8% (60 min) and
92.4% (120 min). Desilication of the solution resulted in a removal
efficiency of 96.3% with calcium silicate as a value-added byproduct. Lithium could be recovered as phosphate with a high purity by
adding phosphoric acid. Moreover, the zeolites synthesized during this investigation were analyzed by powder X-ray diffraction,
Fourier transform infrared spectroscopy, and N2 adsorption/desorption analysis. In addition, selected zeolite samples were
investigated as potential adsorbents for the removal of heavy metal ions (Cu2+, Ni2+, Pb2+, and Zn2+) from aqueous solutions.
This work studies a membrane electrolytic process, where a diluted lithium chloride brine solution is concentrated, while simultaneously producing lithium phosphate. This innovative process takes advantage of the alkaline pH generated in the cathodic section, allowing added phosphoric acid to be precipitated as lithium phosphate. The influence of the operation variables, such as temperature and current density, was evaluated to optimise the precipitation process. Some key findings were observed: a) the optimal conditions for the precipitation of lithium phosphate were 25 °C and 7.3 mA/cm², as the chemical equilibrium was reached in a shorter time period (15 min) under these operating conditions; b) the efficiency and stability of the cationic membrane was favoured at 25 °C; c) it was possible to concentrate the lithium chloride brine solutions to precipitate lithium phosphate, while neutralising the alkaline pH of the cathodic section; and d) the experimental results could be explained using fundamental thermodynamics.
The recovery of lithium compounds from various Li resources is attracting attention due to the increased demand in the Li-ion battery (LIB) industry. This study aimed to secure lithium carbonate (Li2CO3) using the waste Li solution generated from the cathode manufacturing process. The effects of initial Li⁺ concentration, solution pH, and applied phosphate compounds were examined on converting Li-ion into lithium phosphate (Li3PO4) by a precipitation method. Li2CO3 was then obtained from Li3PO4 by reflux and subsequent carbonation methods, in which a concentrated Li solution was produced from Li3PO4 by reaction with metal chloride solutions. Among the tested metal chlorides (AlCl3, MgCl2, and CaCl2), CaCl2 showed the best performance to produce a concentrated Li solution. The highest conversion efficiency (97%) from Li3PO4 to Li solution was observed at 80 °C, and the highest Li concentration (55,900 mg/L) was obtained at a solid(g)/liquid(L) ratio of 500 by reflux reaction. High purity Li2CO3 was produced from the obtained Li solution by purification using sodium hydroxide, and its carbonation using sodium carbonate in turn. The proposed method shows an alternative way to secure Li2CO3 from waste Li solution, in which no acid leaching of Li3PO4 is required to prepare concentrated Li solution to be used Li2CO3 production.
A novel orthophosphoric acid stripping (OpaS) system has been developed for the stripping and crystallization of lithium from the synergistically produced organolithium compound consisting of lithium enolate in kerosene. In order to determine regression models and optimum conditions, a D-optimal experimental plan was constituted. Reduced quadratic, cubic and quartic models were selected for lithium and non‑lithium metal stripping responses. Accordingly, the linear terms of A and B (A: Organic/aqueous phase ratio [O/A], B: [H3PO4]) were determined as a major factor for all lithium responses while the only two-factor interaction (2FI) of AB was determined for all responses. In optimization, the non‑lithium responses were minimized while the lithium responses were maximized. The O/A was determined as 8.2 when the upper limit of 1.0 M was targeted for [H3PO4]. The optimum validated lithium stripping yield and mass fraction of lithium crystallized were predicted as 94.1% and 0.77, respectively. Based on the lithium strip McCabe Thiele diagram, two theoretical stages are required with a 1.0 M H3PO4 strip solution at an O/A ratio of 8.2. At the validated optimum equilibrium pH (pH[eq]) of 7.44, the Li3PO4 crystals of 92.2% purity of were obtained. As a result, the OpaS system has been proven to be a novel method for stripping and crystallization of lithium as Li3PO4 from the lithium enolate organic system without any requirement of thermal and extra chemical processes.
Sulfation and decomposition were proposed to selectively recover lithium, rubidium, and cesium from lepidolite ore. The purpose was to solve the problems of high acid consumption and the difficulty of separating lithium and aluminum in the sulfuric acid method. First, the theoretical feasibility of the process was verified by thermodynamic calculations. The optimal parameters were determined according to the theoretical and experimental results. The extraction rates of lithium, rubidium, and cesium were 90.5%, 91.2%, and 89.4%, respectively, whereas those of aluminum and iron were only 0.08% and 0.02%, respectively. The selective extraction of lithium, rubidium, and cesium was realized, and 90.4% of sulfuric acid could be recycled during the process. Subsequently, the mechanism was discussed by XRD and SEM-EDS analysis. The first-step roasting was the sulfation process of lepidolite, and the second-step roasting was the decomposition process of partial sulfates. The separation of alkali elements and impurity elements could be realized by simple deionized water leaching. The production of Li2CO3 and single-alkali sulfates (K2SO4, Rb2SO4, and Cs2SO4) were obtained through the efficient separation methods of carbonization precipitation and solvent extraction. This process achieved the selective recovery and efficient separation of lithium, rubidium, and cesium from lepidolite ores. At the same time, the recycling of sulfuric acid was realized; it greatly reduced the amount of reagents, such as acid and alkali. It is an efficient, clean, and sustainable process for the utilization of lepidolite ores.
The development of green energy technology has necessitated research on methods to recover and reuse spent Li-ion batteries. To increase the recovery efficiency of precious metals from spent Li batteries, the effects of different leaching parameters, such as the temperature, leaching time, leach liquor concentrations, and solid-liquid ratio, on the leaching efficiency of the NH3 + (NH4)2SO4 + Na2SO3 system were systematically investigated in this study. The most suitable leaching conditions were determined as follows: (NH4)2SO4 concentration = 1.5 mol/L, Na2SO3 concentration = 0.5 mol/L, NH3 concentration = 4 mol/L, solid-liquid ratio = 10:1, leaching time = 180 min, and temperature = 90 °C. Under the most suitable conditions, the leaching efficiencies of Li, Co, and Ni were 96.2%, 89.9%, and 90.1%, respectively, and up to 90.8% Mn remained as a leach residue with good crystallinity. The leach residue was characterised using X-ray diffraction, scanning electron microscopy, electron probe microanalysis, and other detection techniques. The kinetic modelling showed that the leaching process of Li, Co, and Ni was controlled by the surface chemical reaction with apparent activation energies of 62.0, 81.5 and 76.3 kJ/mol, for Li, Co and Ni, respectively. Moreover, Mn precipitated as (NH4)2Mn(SO3)2·H2O in the residue where higher leaching temperatures produced residues of better crystallinity. The optimum temperature to obtain a well-crystallised residue was 90 °C.
Two biomass carbon adsorbents ([email protected] and [email protected]) modified with potassium titanosilicate (PTS) were prepared by a hydrothermal deposition method and characterized by means of SEM-EDS, BET, FTIR, XRD, and XPS. The results show that the adsorbents have mesoporous structures, and potassium titanosilicate ((HK3)O2∙3SiO2∙4TiO2∙4H2O) was immobilized in the biomass carbon structure. The adsorption properties of these two adsorbents for rubidium (Rb⁺) and cesium (Cs⁺) from an aqueous solution and the tail liquor after extracting lithium from lepidolite were systematically studied. The adsorption amounts increased with increasing pH and decreased with increasing KCl concentration, which were mainly due to the competitive adsorption between H⁺ or K⁺ and Rb⁺ or Cs⁺ in solution. The adsorption kinetics of the adsorbents were consistent with the pseudosecondary kinetic model, which indicated that chemical adsorption played a leading role in the adsorption process. The adsorption capacities of Rb⁺ and Cs⁺ reached 2.57 mmol g⁻¹ and 2.12 mmol g⁻¹, respectively. The adsorption mechanism of Rb⁺ and Cs⁺ on the adsorbents was proposed as the ion-exchange between K⁺ in the adsorbents and Rb⁺/Cs⁺ in the solution. In addition, the adsorbents showed high extraction efficiency for Rb⁺ and Cs⁺ from two different lithium extraction tail liquors of lepidolite. Therefore, the obtained adsorbents in this work have excellent adsorption performance and show great potential as adsorption materials for Rb⁺ and Cs⁺.
A new composite adsorbent ([email protected](Sm)) grafted ammonium molybdophosphate onto MOF-76(Sm) was synthesized via a solvothermal method and well-characterized by SEM-EDS, particle size analysis, N2 adsorption-desorption isotherms, TG, FT-IR, XRD and XPS analysis. The adsorption behaviors of Rb⁺ and Cs⁺ on synthesized [email protected](Sm) were investigated depending on the parameters of solution pH, ionic strength, contact time, ion concentration and temperature. The adsorbent can adsorb Rb⁺ and Cs⁺ from the solution when the solution pH ranged from 4.0 to 10.0 and the coexisted ion (Na⁺) has little effect on the adsorption amounts of both Rb⁺ and Cs⁺. The adsorption kinetic data fitted well with pseudo-second order kinetic models. The adsorption equilibrium of Rb⁺ and Cs⁺ on [email protected](Sm) is better explained by the Langmuir isotherm model and the maximum adsorption amounts for Rb⁺ and Cs⁺ are 0.490 mmol g⁻¹ and 0.434 mmol g⁻¹, respectively. Furthermore, the [email protected](Sm) could be easily reused at least four times using 3.0 mmol L⁻¹ ammonia nitrate as the eluent. The mechanism of the adsorption of Rb⁺ and Cs⁺ is proposed to be a combination of electrostatic attraction and ion exchange mechanism.
Low-concentration lithium solutions are often produced during the production and use of lithium, and precipitation via lithium phosphate is an effective recovery method of lithium. However, lithium phosphate has a limited application market, which should be further transformed. Currently, the existing for processing lithium phosphate requires a large amount of chemical reagents, and the recovery rate of lithium is not high. In this study, a bipolar membrane electrodialysis (BMED) process was developed to convert lithium phosphate into lithium hydroxide and phosphoric acid. To verify the feasibility of BMED in lithium phosphate conversion, effects of solubility, current density and concentration of phosphate on the separation performance were investigated. The results suggested that the recovery rate of lithium was as high as 99%. The high solubility of lithium phosphate in phosphoric acid resulted in the superior BMED performance. The current efficiency decreased with the increase in current density, while the energy consumption increased as current density elevated. Typically, the lowest energy consumption of 10.54 kWh·kg⁻¹ was obtained at the current density of 20 mA·cm⁻². Besides, a higher lithium phosphate concentration led to the superior BMED performance. The total process cost of BMED process was estimated to be $2.941·kg⁻¹ LiOH under the optimal experimental conditions. Considering the price of lithium hydroxide, BMED is considered as a high-efficient, economical, and environmentally friendly technology for lithium phosphate treatment.
Rubidium is a rare alkali metal with high economic value and application potential. It is mainly recovered from the intermediate products of lithium/cesium extraction from lepidolite/pollucite, and thus its extraction process has always been accompanied by resource problem. The application potential and extraction difficulties of rubidium have necessitated the development of new methods for the extraction of rubidium from resources. However, to the best of our knowledge, the extraction of rubidium has received less attention than that of the main group elements, lithium and potassium, and there is no review of the rubidium extraction process to date. Hence, this state-of the-art review addresses the development of rubidium extraction technology and points out the current barriers and potential strategies to achieve the effective rubidium extraction. The review includes the occurrence of rubidium in various resources, and the processes and mechanism of rubidium extraction from mineral (lepidolite, zinnwaldite, pollucite, kaolin, muscovite and biotite) and brine (salt lake and seawater) resources. It is expected that the review can provide insights for the efficient utilization of the rubidium resources.
This contribution provides a detailed in-situ account of transformation reactions during calcination of a typical high-grade α-spodumene (α-LiAlSi2O6) concentrate, a pre-treatment step required to refine spodumene into commercial lithium chemicals. We observe four reaction pathways during the transition of spodumene, employing in-situ high-temperature powder XRD measurements using both cathode-tube and synchrotron radiation. At a relatively slow heating rate of 8 °C min⁻¹, we observe a close relationship between the development of γ-spodumene, with an onset temperature of 842 °C, and reduction of the amorphous background, in the collected XRD spectra. This demonstrates that, initially γ-spodumene recrystallises from amorphous spodumene. At the fast initial heating rate of 100 °C min⁻¹, γ-spodumene first appears at a higher temperature of 1025 °C. This mineral subsequently transforms into β-spodumene at high temperatures along the reaction pathways denoted as pathway (1) amorphous spodumene → γ-spodumene → β-spodumene and (3) crystalline α-spodumene → γ-spodumene → β-spodumene. The stability of γ-spodumene strongly depends on the mechanical treatment of the sample, and the heating rate of the calcination process, suggesting high and low activation energies for pathways (3) and (1), respectively. In another experiment, we observe rising peaks of β-quartz, a minor gangue mineral in the spodumene concentrate, that reflect the substitution of Li⁺ and Al³⁺ for Si⁴⁺ above 875 °C. This phase ultimately transforms to β-spodumene at 975 °C. The same experiment demonstrates the spectrum of β-spodumene continuously increasing in magnitude, above 975 °C, with decreasing abundance of α-spodumene, indicating a direct conversion of α- to β-spodumene. Thus, the two other reaction corridors comprise: (2) crystalline α-spodumene → β-quartzss → β-spodumene; (4) crystalline α-spodumene → β-spodumene. Heating of a finely ground sample results in faster and more-complete conversion of α-spodumene compared to a coarser specimen. Our experiments establish the characteristic temperatures of phase transformations during spodumene calcination and reveal the influence of amorphous material and thermal history on reaction sequences. Approaches that integrate the optimisation of grinding and heating thus bear the potential to reduce the energy requirements of the calcination process, including the extraction of lithium from γ-spodumene formed at a lower temperature.
The paper is devoted to the results of investigating the spontaneous Raman scattering spectra in the lithium compounds crystals in a wide spectral range by the fibre-optic spectroscopy method. We also present the stimulated Raman scattering spectra in the lithium hydroxide and lithium deuteride crystals obtained with the use of powerful laser source. The symmetry properties of the lithium hydroxide, lithium hydroxide monohydrate and lithium deuteride crystals optical modes were analyzed by means of the irreducible representations of the point symmetry groups. We have established the selection rules in the Raman and infrared absorption spectra of LiOH, LiOHH2O and LiD crystals.
This work describes the development of a new process for the recovery of Li, Al and Si along with the proposal of a flow sheet for the precipitation of those metals. The developed process is comprised of lepidolite acid digestion with hydrofluoric acid, and the subsequent precipitation of the metals present in the leach liquor. The leaching operational parameters studied were: reaction time, temperature and HF concentration. The experimental results indicate that the optimal conditions to achieve a Li extraction higher than 90% were: solid-liquid ratio, 1.82% (w/v); temperature, 123 °C; HF concentration, 7% (v/v); stirring speed, 330 rpm; and reaction time, 120 min. Al and Si can be recovered as Na3AlF6 and K2SiF6. LiF was separated from the leach liquor during water evaporation, with recovery values of 92%.
Mixed grinding with Na2S followed by water leaching was performed to extract Li from lepidolite. The leachability of Li increases dramatically in the ground mixture, regardless of the mixing ratio over the range of 1:1 to 3:1, while only 4.53% of Li was extracted in lepidolite ground without Na2S. The leachability increased with an increase of the grinding time, and ultimately, 93% of the Li was leached by water from the ground mixture with a weight ratio of 3:1 (Na2S:Lepidolite). In the process of the mixed grinding, the Li-contained lepidolite was destructured crystallographically, and it might have changed to different compounds. This process enables us to extract Li from lepidolite via a water leaching treatment.
Lithium is becoming a strategic metal due to its important applications in secondary battery electrodes used in electronic appliances and also in electric traction vehicles. Lithium primary resources are brines and rock minerals, the former being nowadays almost exclusively used in the production of lithium commodities. With the expected increase in lithium demand, the development of competitive technologies for recovery lithium from ores like pegmatites is getting imperative. The high energy and reagents consumption in processing minerals is an issue that should be considered. This paper presents some results on the comparison of two acid treatment routes for lepidolite, the H 2 SO 4 digestion and HCl leaching. Before both chemical treatments, lepidolite was calcined at 800ºC and was transformed in a more reactive species, β-spodumene. The H 2 SO 4 digestion at 175ºC (followed by water leaching) allowed 88% Li recovery into the solution in 30 min. By the contrary, the HCl leaching process carried out at 90ºC also achieved similar yields but only after 4 h of reaction. In both cases, an acid excess was used, but clearly higher for the essays with HCl. The H 2 SO 4 digestion process was also advantageous in what concerns to selectivity over other contained metals. Al, Mn and Fe concentrations in solutions were substantially higher in the hydrochloric acid leaching. These results showed that the digestion with sulphuric acid can be a more efficient and competitive process.
The ion-exchange properties of a synthetic hydrosodalite (Na-hS) have been investigated by kinetic and thermodynamic analysis of exchange reactions of the original sodium form for lithium, potassium and calcium forms. Kinetic curves, modelled by a Langmuir-type equation, revealed that exchange rate for lithium and for potassium are of the same order, whereas they are two order faster than for calcium. Thermodynamic analysis of the cation exchange isotherms pointed out that sodalite is selective for sodium over the other three cationic forms examined, which is consistent with the preference exhibited by the sodalite type for sodium environments, either in natural or in laboratory crystallization. Na/Li and Na/Ca exchanges are incomplete, whereas unexpectedly Na/K exchange turns out to be complete, even though K+ dimension exceeds the width of the access window to sodalite cages. The obtained results have been discussed in terms of Eisenman–Sherry theory, pointing out agreements and discrepancies.
Electrochemical etching of NiTi alloy in a pH 6 fluoride solution is proposed as a convenient, less dangerous procedure than traditional chemical etching in oxidizing HF media. Anodic polarization produces protons at the electrode surface, and local acidification increases the concentration of HF that efficiently dissolves the surface oxides, promoting a sustained dissolution current that decreases with increasing mass-transfer rate. The voltammetric pattern includes a "pseudoplateau" at low potentials, in which only a metal dissolution takes place, and a steady current increase at higher potentials due to the onset of oxygen evolution. The treatment produces a semilustrous rough surface, showing a structure of shallow microscopic cavities in scanning electron microscopy images. Impedance spectra recorded in the pseudoplateau region are similar to those observed during anodic polarization of Ti in HF, and their analysis indicates a formation of a thin barrier film whose thickness increases linearly with potential.
The continuously increasing demand for lithium has made it one of the strategic metals, rendering its exploitation of critical importance. Natural α-spodumene is still the primary resource of lithium extraction. The traditional process for the treatment of α-spodumene generates immense quantities of waste residue and needs a high-temperature heat treatment, leading to high energy consumption. In addition to lithium, α-spodumene is rich in aluminum and silicon, and thus it is a potential raw material for zeolite synthesis. Herein, a novel process was developed for the clean and efficient extraction of lithium from α-spodumene coupled with the synthesis of hydroxysodalite zeolite. By hydrothermal alkaline treatment, α-spodumene was converted into hydroxysodalite; the lithium in α-spodumene was released into the solution, and subsequently recovered by precipitation with Na2CO3. A lithium extraction efficiency of 95.8% was obtained under the optimum conditions: temperature 250 °C, NaOH concentration 600 g/L, liquid/solid ratio 5:1, stirring speed 500 rpm, and reaction time 2 h. In addition, the influences of various factors on the composition and textural properties of the product were analyzed using XRD, SEM, TG, N2 adsorption/desorption, and FTIR.
The lepidolite located in Yichun, Jiangxi Province, China, was adopted to investigate the recovery of alkali metals and leaching kinetics of lithium with sulphuric acid solution under atmospheric pressure. The results show that the recoveries of alkali metals were achieved under the leaching conditions: mass ratio of lepidolite with particle size less than 180 μm to sulphuric acid 1.2, leaching temperature 411 K, liquid−solid ratio 2.5:1, and leaching time 10 h. Under the selected conditions for leaching experiment, the leaching rates of lithium, potassium, rubidium and caesium are 94.18%, 93.70%, 91.81% and 89.22%, respectively. The X-ray diffraction analysis for leaching residue indicates that no insoluble product forms during leaching. The chemical compositions of leaching residue reveal that trace iron, manganese and calcium disappear after acid leaching. The kinetics of leaching process for lithium follows shrinking core model of mixed control and the apparent activation energy is 17.21 kJ/mol. The reaction orders with respect to sulphuric acid concentration and liquid−solid ratio are determined to be 2.85 and 1.66, respectively. A semi-empirical rate equation was obtained to describe the leaching process. The kinetic analysis shows that the leaching process is controlled by diffusion through the insoluble layer of the associated minerals.
Concrete is a composite material that is widely used in the construction industry due to its excellent mechanical and physical properties. Despite these benefits, concrete possesses several disadvantages including negative environmental impacts and mechanical durability (e.g., shrinkage, frost attack, and corrosion). To date, upgrading of concrete's properties to overcome such drawbacks has been one of the most challenging, but attractive, research topics for many researchers in this research field. As one of the effective means to meet such demand, the use of natural zeolites in the production of concrete has been made preferably to acquire the excellent performance along with improvement in terms of structure, durability, and mechanical properties. This review was designed to provide a comprehensive insight into the construction-related applications of natural zeolite. To this end, we discussed the structural and fundamental properties of natural zeolites and their applications to concrete production as natural pozzolans, internal curing agents, and lightweight aggregates. Also, through critical analysis of various researches made previously, we aim to offer a better understanding on the effect of zeolites' addition upon the performance of concrete in terms of workability, strength, durability, and permeability. Additionally, we describe the present challenges and future perspectives in the use of natural zeolites for concrete production.
Three lepidolite-1M and two lepidolite-2M2 mica structures from three localities were refined using single-crystal XRD data. The localities are: Radkovice, Czechoslovakia; Tanakamiyama, Japan; and Elba, Italy. Structural details are discussed and illustrated. Apparently different lepidolites have slightly different mica structures. Chemical analyses are given for lepidolites from Radkovice and Tanakamiyama. -K.A.R.
Mechanical activation in order to improve lithium recovery from a lepidolite ore concentrate was investigated. The concentrate was mechanically activated using different milling times and the effect on the resulting products was evaluated by X-ray powder diffraction, FTIR spectroscopy, TG-DTG analysis, SEM and particle size analysis by laser diffraction. The mechanical activation leads to lepidolite amorphization and concomitant improvement of the reactivity by sulfuric acid digestion and water leaching. About 85% of Li recovery was achieved when the concentrate was activated for 15 min and afterwards digested at 165 °C, during 4 h, using an acid/concentrate ratio of 650 g/kg. Li recoveries above 70% were reached even using digestion temperatures lower than 100 °C. However, to increase recovery, temperatures over 140 °C are more suitable. Results obtained showed that is possible to apply a low temperature technology for recovering Li from hard rock minerals, using mechanical activation.
In this study, solubility constants of hydroxyl sodalite (ideal formula, Na8[Al6Si6O24][OH]2·3H2O) from 25 °C to 100 °C are obtained by applying a high temperature Al—Si Pitzer model to evaluate solubility data on hydroxyl sodalite in high ionic strength solutions at elevated temperatures. A validation test comparing model-independent experimental data to model predictions demonstrates that the solubility values produced by the model are in excellent agreement with the experimental data.
The equilibrium constants obtained in this study have a wide range of applications, including synthesis of hydroxyl sodalite, de-silication in the Bayer process for extraction of alumina, and the performance of proposed sodalite waste forms in geological repositories in various lithologies including salt formations. The thermodynamic calculations based on the equilibrium constants obtained in this work indicate that the solubility products in terms of mΣAl×mΣSi for hydroxyl sodalite are very low (e.g., ∼10–13 [mol·kg–1]² at 100 °C) in brines characteristic of salt formations, implying that sodalite waste forms would perform very well in repositories located in salt formations. The information regarding the solubility behavior of hydroxyl sodalite obtained in this study provides guidance to investigate the performance of other pure end-members of sodalite such as chloride- and iodide-sodalite, which may be of interest for geological repositories in various media.
The aim of this study was to improve understanding of scale formation mechanisms and to provide theoretical support for development of scale inhibition strategies. Sodium aluminosilicate scaling on plant surfaces is a critical problem in Bayer refineries, particularly within digesters and heat exchangers. Although numerous studies have been carried out to understand the mechanisms of scale formation and to develop scale inhibition strategies in double-stream circuit Bayer plants, there is still insufficient relevant research on single-stream Bayer plants. We have studied the desilication behavior of a liquor of similar composition to spent cleaning caustic liquor, for which the aluminate concentration and A/C are both less than heat exchanger liquor. These concentration conditions were chosen in order to investigate the early stage of scale formation in single-stream Bayer plants by increasing the component of aluminosilicate bulk precipitate upon seeding and decreasing the relative degree of scaling.
Solution desilication experiments were performed using spent cleaning caustic liquor at 140 °C either without seeding or with sodalite (seed loading 1.125 g L⁻¹) or cancrinite seeding (seed loadings 1.125 g L⁻¹ and 2.850 g L⁻¹) in the presence of steel substrates. The solution desilication rate showed significant increase in the presence of seeds and was found to be greater in the presence of cancrinite seeds than sodalite of the same surface area. The desilication kinetics of the synthetic spent cleaning caustic liquor suggests predominant heterogeneous nucleation for unseeded desilication, a mixed secondary-nucleation and growth process for desilication on addition of sodalite seeding, and cancrinite seeding with both smaller and greater loading.
A first layer, possibly amorphous, with high Al and low Si concentrations was observed by SEM and FIB/SEM on the coupon surface after desilication with or without seeds. We note that this layers forms despite the low A/C and high Si supersaturation suggesting that under single-stream heat exchanger conditions this may also be the first layer to form. XRD confirmed that sodalite formed on the coupons in unseeded and sodalite seeded desilication systems while cancrinite formed on the coupons upon cancrinite seeded desilication with both smaller and greater seed loading. The coupon from the cancrinite seeded liquor with high seed loading is least covered by scale amongst the coupons, suggesting that cancrinite seeding may be a potential strategy for reducing scale formation in single-stream Bayer circuits.
This paper collates existing information from the literature on the reactions of lime added directly to high temperature Bayer digestions. It examines the reaction of calcium in lime with liquor components to form reaction products. Reactions are considered individually and in competition with each other. Because of the scarcity of information about competing reactions involving calcium/lime, the paper contains some speculation, drawing on different pieces of evidence from the literature to suggest possible reaction pathways. Three lime types have been considered — slaked lime (Ca(OH)2), calcite (CaCO3) and tri-calcium aluminate (TCA, Ca3Al2(OH)12).
Individual reactions of calcium in lime have been considered with alumina and carbonate, fluorine, vanadium, phosphorus, silica, titanium and iron. Consideration has also been given to the “catalytic” actions of lime in transforming goethite to hematite and enhancing boehmite dissolution.
A very broad vibrational band ranging from 1000 up to 4000 cm−1 and two relatively sharp bands at 5000 and 5027 cm−1 are found in the Raman scattering spectrum of hydroxyapatite-containing films obtained by gas detonation spray method. We developed a theoretical model that interprets the broad band as a result of strong interaction between the high-frequency hydrogen bond vibrations and lattice phonons. Both sharp bands around 5000 cm−1 are assigned to the overtones of v-OH vibrations. Copyright
Lithium (Li), an exceptional cathode material in rechargeable batteries, is an essential element in modern energy production and storage devices. The continuously increasing demand for lithium in these devices, along with their steady production, has led to the high economic importance of lithium, making it one of the strategically influential elements. The uneven distribution of mineral resources in the earth’s crust and the unequal concentration in brine and sea water reserves also causes lithium exploitation to be of critical importance. This situation requires the efficient processing of lithium resources either by the processing of minerals/brine/sea water or by the recycling of spent lithium-ion batteries. To explore new routes for the sustainable exploitation of lithium, it is imperative to review the methodologies that have already been studied and are currently in industrial practice. In this study, we present an overview of the processes investigated for the extraction, separation and recovery of lithium from not only a technological perspective but also from a chemical perspective.
The solubility of sodium aluminosilicate (sodalite), a major desilication product (DSP) in the Bayer process, in NaOH and NaOH–NaAl(OH)4 solutions was determined and modeled at temperatures from 303.2 to 348.2 K. Sodium aluminosilicate was synthesized in a batch crystallizer and the effect of temperature, solution concentrations, and aging time was investigated. Solubility was found to increase with increasing NaOH concentration while the solubility sharply decreases with the addition of Al(OH)3, reaches a minimum at about 0.8 mol·L–1, and then increases. A mixed-solvent electrolyte (MSE) model for the solubility of sodalite was developed with the help of the OLI platform via regression to obtain the model parameters. In addition, the desilication kinetics of NaOH–NaAl(OH)4 solutions by using sodalite as seeds was studied experimentally and modeled with the aid of a second order kinetic model. The activation energy of desilication over the temperature range 323.2–363.2 K was found to be 92 ± 14 kJ·mol–1. Under the optimal operation conditions, 80% silica was removed after 2 h.
The phase transitions of the main substances in Bayer red mud in high caustic sodium aluminate solutions were studied. Without addition of lime, cancrinite was not found to transform to CaNaHSiO4 even with 16 wt.% CaO present in the original residue up to 270 °C. However, this transition was verified to be a fast reaction which was completed in just 10 min with the temperature higher than 240 °C after adding lime, thus enabling the extraction of Al2O3 from the red mud. Whether additional CaO was supplemented or not, the isomorphous substitution of Fe to Al atoms occurred in hydrated andradite as long as the temperature was more than 240 °C, providing another way to extract Al2O3 from Bayer red mud. Accordingly the A/S (weight ratio of solid Al2O3 to SiO2) of red mud would reduce to 0.134 after the complete transition from Ca2.93Al1.97Si0.64O2.56(OH)9.44 to Ca3(Fe0.87Al0.13)2(SiO4)1.65(OH)5.4 in our study, assuming no other Al or Si containing phases. Furthermore, the addition of CaNaHSiO4 seed was confirmed to accelerate both the transition of cancrinite to CaNaHSiO4 and the reaction of isomorphous substitution in the andradite.
Iron sulphate roasting and water leaching were investigated to extract lithium from lepidolite in this study. HSC modelling was used to simulate the process of roasting lepidolite with FeSO4·7H2O and CaO. Based on HSC, three-dimensional models were derived to predict the effect of temperature, SO4/Li and Ca/F molar ratios on the production of S- and F-containing gases (SO2, SO3, HF) and Li species (Li2SO4, LiKSO4) during roasting. It is believed that temperature and SO2/SO3 gases controlled the extraction of lithium from lepidolite during roasting while more soluble Li sulphate species determined the recovery of lithium during leaching. Using optimum parameters selected from HSC, roasting tests were conducted to produce calcines for leaching. Optimum roasting conditions were experimentally determined as 850 °C, 1.5 h, and SO4/Li and Ca/F molar ratios of 3:1 and 1:1, respectively. Roasting in a closed environment led to more Li extracted than with an open system. Leaching the calcines obtained from the open and closed systems with a water/calcine mass ratio of 1:1 at room temperature for 1 h yielded leach liquors containing ~ 7.9 g/L Li and ~ 8.7 g/L Li, corresponding to ~ 85% and ~ 93% extractions of Li from lepidolite, respectively.
Chlorination roasting followed by water leaching process was used to extract lithium from lepidolite. The microstructure of the lepidolite and roasted materials were characterized by X-ray diffraction (XRD). Various parameters including chlorination roasting temperature, time, type and amount of chlorinating agents were optimized. The conditional experiments indicate that the best mass ratio of lepidolite to NaCl to CaCl2 is 1:0.6:0.4 during the roasting process. The extraction of lithium reaches peak value of 92.86% at 880 °C, potassium, rubidium, and cesium 88.49%, 93.60% and 93.01%, respectively. The XRD result indicates that the major phases of the product after roasting lepidolite with mixture of chlorinating agents (CaCl2 and NaCl) are SiO2, CaF2, KCl, CaSiO3, CaAl2Si2O8, NaCl and NaAlSi3O8.
Salt roasting with Na2SO4 + CaCl2 followed by water leaching was used to extract alkali metals from lepidolite. The experiments indicated that the best mass ratio of lepidolite /Na2SO4/CaCl2 during roasting was 1:0.5:0.3. The extraction of Li, Rb and Cs were all > 90% after 0.5 h at 880 °C. The recovery of Li was essentially constant when roasted at 830–930 °C. The flexible roasting condition is easily controlled in industrial application. After leaching in water, the solution was cooled to − 5 °C for 2 h to crystallise 92.1% of the sulphate and 3.9% of the chloride as Na2SO4·10H2O and NaCl, respectively. Evaporation and precipitation using Na2CO3 produced Li2CO3 crystals with > 99.5% purity and a solution from which Rb and Cs could be recovered.
A novel technique was developed to extract lithium from lepidolite.The lepidolite was pre-roasted at a high temperature with water steam atmosphere for defluorination. Then the defluorinated lepidolite was leached in a lime–milk autoclave. Various parameters including the defluorination percentage of lepidolite, milling time, temperature, time, lime-to-defluorinated lepidolite ratio, and liquid-to-solid ratio in the leaching process were optimized. The lithium extraction efficiency can reach 98.9% under the optimal conditions. The purity of the lithium carbonate obtained can be up to 99.9%.
Key factors controlling the extraction of lithium from a lepidolite concentrate (2.55% Li) using a two-stage process based on roasting with Na2SO4 and water leaching were determined in this study. HSC software was used to predict lepidolite decomposition characteristics and products yielded from roasting. From this simulation roasting was conducted at 850–1000 °C with different sodium sulphate additions (set at sodium sulphate-lithium molar ratios of 1:1 to 3:1) in 0.5–2 h to yield several calcines for leaching. Stabcal modelling was applied to investigate the stability of various Li species formed during water leaching. From the simulation it was identified that LiKSO4, as one of the main Li-containing products formed in the calcine, has a low solubility and therefore controls the release of Li into water during leaching. The low solubility of LiKSO4 at 25 °C indicated by Stabcal was confirmed during leaching at different water–calcine mass ratios from 5:1 to 15:1. Leaching under these conditions only yields liquors containing < 1.0 g/L Li, corresponding to Li extractions of 45–48%. The Li extraction was significantly improved at 85 °C, yielding liquors containing 1–3 g/L Li, corresponding to extractions of 47–90%, with lower extractions of Li achieved at higher Li concentrations. At best, 90.4% of lithium was extracted (at water–calcine mass ratio 15:1, 85 °C and 3 h) using calcines roasted at 1000 °C for 0.5 h (using Na2SO4 — Li molar ratio of 2:1).
Sulfation roasting followed by water leaching process was used to extract lithium from lepidolite. Various parameters including roasting temperature, the amount of additions, and solid/liquid ratio in leaching process were optimized. The lithium extraction efficiency of 91.61% could be reached with a mass ratio of lepidolite/Na2SO4/K2SO4/CaO of 1:0.5:0.1:0.1, and roasting at 850 °C for 0.5 h. XRD analysis showed that sulfation roasting caused the decompositions of the original aluminosilicate to NaSi3AlO8, KAlSi2O6 and CaAl2Si2O8. The phases of CaF2 and Ca4Si2O7F2 are observed due to the addition of CaO.
Phase pure, homogeneous, and well-crystallized lithium iron phosphate LiFePO4 was synthesized by aqueous co-precipitation of an Fe(II) precursor material and succeeding heat treatment in nitrogen. The particle morphology of the precursor is preserved during the heat treatments. Excellent electrochemical properties in terms of capacity, reversibility, cycling stability and rate capability have been achieved. The thermal stability of charged electrodes is superior versus other positive electrode materials. If reductive synthesis conditions are used in the heat treatments, problems arise from the possible generation of iron phosphide.
Equilibrium structures of Li2CO3 and K2CO3 were calculated using ab initio molecular orbital calculations carried out at the Hartree-Fock (HF) level. Of the four structures considered for Li2CO3 and K2CO3, the most stable was a structure with all five atoms in a plane. The harmonic frequencies were also calculated and found to be in agreement with the present Raman measurements. Structure factors, calculated from the ab initio data for each of the four structures considered, are compared with existing X-ray results.
An interpretation of the archaeological record, in particular that of a prehistoric cave site, is complicated by the diversity of depositional and post-depositional processes that affect the material deposited. Here we propose to use the authigenic minerals that form in situ within the cave sediments to reconstruct the ancient chemical environments in the sediments. This can be done by experimentally determining the conditions under which each of the authigenic minerals are stable. Although this information is not available to date for minerals formed in a prehistoric cave, we present calculated stability field data for the relevant minerals. The results clearly demonstrate the feasibility of this approach. This information, particularly if based on measurements of real authigenic cave minerals, will facilitate an assessment of the completeness of the cave archaeological record. This is particularly important for determining whether or not the distributions of archaeologically important materials, such as bones, teeth, plant phytoliths, charcoal and ash, reflect their original burial distributions or were altered as a result of secondary diagenetic processes.
Lime is used extensively within the Bayer process, where it performs many useful functions. In particular, it can enhance the extraction of alumina (improving the dissolution of boehmite and diaspore or the conversion of aluminogoethite), control liquor impurities (desilication, causticisation or phosphorous control), assist with the removal of impurities from the pregnant liquor (liquor ‘polishing’) and minimise soda losses in the red mud (formation of alternate desilication products or calcium titanates). This review examines these uses of lime and, where possible, uses the underlying chemistry to explain the observed benefits resulting from lime addition. A particular emphasis is placed on those factors influencing the efficiency of lime use (eg: surface area, reaction temperature, addition point and the presence and concentration of ‘impurities’).
The solubility of aluminum and silicon in aqueous, alkaline solutions was investigated at 95-degrees-C for solutions containing 0.1-4 mol/L NaOH and NaCl to ionic strengths of 4. The apparent solubility product, [Al][Si], increased with increasing [OH-] and decreasing ionic strength. It remained constant over a wide range (0.076-270) of [Al]/[Si] ratios at equilibrium. In this range, the solid phase consisted of sodalite or, at higher [Al]/[Si] ratios, of sodalite and Al(OH)3. At lower equilibrium [Al]/[Si] ratios, both sodalite and an amorphous product which contained higher ratios of Si/Al were found. An analysis of the chemical characteristics of aluminum and silicon in alkaline solutions in the solution composition range of interest showed that Al(OH)4- and HSiO43- are the predominant aluminum and silicon species, respectively. In addition, precipitates chemically similar to sodalite (hydroxysodalite, cancrinite) were reported by others when the anion matrix differed from that in this study but other conditions were similar. The solubility of aluminosilicates does not appear to be affected greatly by the composition of the anion matrix. A chemical model, employing Pitzer's activity coefficient equations, was developed to describe the equilibrium of sodalite over the range of composition of interest. Values for the apparent solubility product at 95-degrees-C and interaction parameters for the aluminum and silicon species were obtained by fitting the model to experimental solubility data. The predictions from the model fell well within the scatter of the experimental data. A concise graphical method was developed for estimating the apparent solubility product from gross solution composition. It gives reasonable predictions for solutions which contain other anions as well.
Data are reported for rare earth elements (REE), Y, Th, Zr, Hf, Nb and Ta in four geological reference materials using sodium peroxide (Na 2O2) sintering and inductively coupled plasma-mass spectrometry. The described procedure was used by students during their thesis work. A compilation of their reference material data acquired over one year of laboratory work demonstrates the ease and reliability of the method and the high reproducibility of the analytical results. Relative standard deviations of up to thirty six measurements of one reference material were lower than 5% for Y and the REE. Reproduciblities of Zr, Hf, Nb, Ta and Th were higher at between 5% and 10%, and can be attributed to the inhomogeneous distribution of zircon and other trace mineral phases and uncorrected drift effects. The concentration data are compared to reference and literature values and demonstrate that the procedure is also accurate. New data on G-3 show some systematic deviations from G-2, which are statistically significant.
The article contains sections titled:
1.
Introduction
2.
Properties
2.1.
Physical Properties
2.2.
Chemical Properties
3.
Occurrence
3.1.
Important Lithium Minerals
3.1.1.
Lithium Aluminum Silicates
3.1.2.
Micas
3.1.3.
Lithium Phosphates
3.1.4.
Other Lithium Ores
3.2.
Reserves of Lithium Minerals
3.3.
Lithium in Natural Brines
4.
Production of Primary Lithium Compounds
4.1.
Mining of Ore and Production of Concentrate
4.2.
Ore Digestion and Production of Lithium Compounds
4.2.1.
Acid Digestion
4.2.2.
Alkali Digestion
4.2.3.
Ion‐Exchange Processes
4.3.
Production of Lithium Carbonate from Brines
5.
Lithium Metal and Lithium Alloys
5.1.
Production of Lithium Metal
5.2.
Uses of Lithium Metal
5.3.
Lithium Alloys
6.
Lithium Compounds
6.1.
Inorganic Lithium Compounds and Lithium Salts
6.2.
Organolithium Compounds, Lithium Alkylamides, and Lithium Alkoxides
7.
Quality Specifications and Analysis
8.
Toxicology and Occupational Health
9.
Economic Aspects
Apatitic tricalcium phosphate Ca9(HPO4)(PO4)5(OH) is a calcium orthophosphate that transforms into β-tricalcium phosphate Ca3(PO4)2 by heating above 750 °C. This work deals with powder synthesis using a wet precipitation method. An experimental design is applied to precise the influence of the synthesis parameters on the chemical composition (e.g. the Ca/P molar ratio). The Ca/P ratio of the precipitates varies greatly according to the pH value and the temperature of synthesis. A more or less important increase of the Ca/P ratio can occur with the ripening time in dependence on the value of the previous parameters. A reproducible synthesis of pure apatitic tricalcium phosphate (TCP) powders is attained by refinement of the parameters. The study is completed by physicochemical characterizations and the thermal behavior of the powders. X-ray diffractometry and differential thermal analysis are necessary to insure the purity of TCP powders. The decomposition of the apatitic TCP into β-TCP during heating influences the sintering behavior.
A simple method to prepare nanocrystalline hydroxyapatite (nHAP) is performed using a precipitation method assisted with microwave heating method. This method can be reported notably with high reproducibility and productivity. The received ceramic powder possesses characteristic of needle-shaped nanocrystals with dimension about 50 nm in diameter and 200 nm in length. The particle size distribution has been confirmed being in the range of 28–159 nm. Thermal analyses revealed that nHAP has at least three thermal events influenced by elevated temperatures. Phase stability and microstructure evolution of the nHAP calcined at temperatures range between 700 and 1200 °C are discussed in terms of the formation of secondary phases, the decomposition of HAP releasing carbonate and water. Various experimental techniques have been employed in this work, including powder X-ray diffraction, IR spectroscopy, DSC and TGA thermal analyses, dynamic light scattering and scanning electron microscopy.
We describe a method for determining fluoride with ion-selective electrodes (ISEs). Tartrate and Tris-based total ionic strength adjustment buffers (TISABs) were found to lower the interference from aluminum to a greater extent than conventional citrate-based TISABs. We adopted a solid TISAB addition method that is simple to perform, and can be carried out without lowering the level of fluoride. The apparent recovery of fluoride was 95% or higher, even at 500 mg L(-1) of Al3+ when a tartrate and Tris-based TISAB was used. Interferences from common ions were not observed at 100 mg L(-1) levels. We determined the fluoride content in solid silicate samples with ISEs without preliminary steam distillation after alkali fusion processing. Adding a solid TISAB mixture consisting of tartaric acid, sodium tartrate, and Tris, however, eliminated any interference from high levels of aluminum and sodium and potassium carbonates. The proposed analytical method was also applied to the determination of fluoride in geochemical reference samples.
Method of recovering Lithium from Lepidolite
- Botton
Interpretation of the composition of Lithium micas: shorter contribution to general geology
- Foster
Recovery of Lithium values from Lithium bearing ores
- Kepfer
The manufacture of Lithia from Lepidolite
- Schieffelin
Lithium Extraction Techniques - A Look at the Latest Technologies and the Companies Involved
- M Bohsen
Caustic digestion process
- E Catovic