Hans Oskierski

This study reports on the development of three spodumene reference materials referred to as Curtin University (CU) Tri-1, Kun-A-1 and MG-1 for calibrating the instrumental mass fractionation for oxygen isotope measurement by secondary ionisation mass spectrometry (SIMS). Millimetre-sized fragments were prepared from gem-quality spodumene crystals with chemical compositions close to the ideal spodumene end member (LiAlSi2O6). From these fragments, six subsamples were extracted for laser fluorination isotope ratio mass spectrometry (LF-IRMS) to determine the bulk oxygen isotope ratio of the three materials and confirm their homogeneity on the mg scale. The bulk δ18O (VSMOW) values are 11.29 ± 0.04‰, 9.12 ± 0.10‰ and 15.13 ± 0.08‰ (all 2SE, standard error of the mean) for materials Tri-1, Kun-A-1 and MG-1, respectively. A random selection of twenty or more fragments from each material was subjected to SIMS analysis. Repeatability of SIMS δ18O measurements was at most 0.24‰ (2s), indicating oxygen isotopic homogeneity on the ng scale. Crystal orientation effects potentially affecting SIMS analysis were systematically evaluated and found to be absent at levels equivalent to the repeatability of the method. Applications of the new materials lie mainly in furthering research on pegmatite formation and tracing of raw materials by providing the means for accurate, in situ oxygen isotope measurement in spodumene.

The conventional process of extracting lithium by decrepitating the concentrates of α-spodumene and then baking β-spodumene with concentrated sulfuric acid is the most economic to operate but nonsustainable because of its feedstock, energy, and waste byproduct intensity. Directly extracting lithium from α-spodumene by roasting it with potassium sulfate (K2SO4) followed by leaching it with water offers a more sustainable alternative. We optimized the potassium sulfate process (PSP) at a ratio of K2SO4 to spodumene concentrate of 0.6:1 (w/w), 1050 °C, and 30 min roasting time, achieving a lithium extraction efficiency of 96.3 ± 1.4% (w/w), in comparison to 96.66 ± 0.37% (w/w) for the conventional process, for the same feedstock of spodumene concentrate. While the purification of the leach liquor from PSP is more complex, requiring the addition of aluminum sulfate to recover potassium as potash alum, its byproducts have high economic value. Both processes display a similar energy demand, based on 200 kt y–1 of the spodumene concentrate feed. The use of aluminum sulfate increases the overall cost of PSP by $12.8 million, but sales of potassium alum elevate the revenue by $45.8 million. We reveal that the key advantage of PSP lies in its capability of leveraging the byproducts (leucite and potash alum), while the sulfuric acid process (SAP) may incur disposal cost for its hydrogen aluminosilicate (HAS) byproduct. For PSP to breakeven with SAP, leucite must be converted to a fertilizer and sold at a price of $102.6 t–1 if HAS requires no disposal cost. With the process development focused on the byproduct value, the proposed PSP provides an efficient, more sustainable, and potential near zero-waste alternative to the conventional refining of the lithium chemicals from spodumene.

Kilns consume about half of the energy necessary to operate lithium refineries and their decarbonisation requires accurate modelling of the calcination of spodumene concentrates fed to the process. This contribution applies the isoconversional methodology to investigate the kinetic parameters of the transition of α-spodumene to its high-temperature polymorph of β-spodumene, using the heat flux measurements from the differential scanning calorimetry (DSC). The normalised energy demand (αH), presented as a function of temperature, characterises these measurements. The activation energy Eα and the product of the reaction model fα and the frequency factor Aα, (fA)α, depend on αH. As the process involves multi-step reactions, we deploy the Friedman differential method and the accurate flexible integral method of Vyazovkin to obtain the kinetic parameters. We also modify the method of Ortega to acquire additional estimates of Eα and (fA)α and apply the rigid integral method of Starink for comparison. The Friedman, Vyazovkin and modified methods deliver the same estimates of the kinetic parameters within their error bands. The Starink method works surprisingly well for predicting the conversion time despite the inaccuracies in the derived values of Eα and (fA)α. This comes to pass because of the compensation effect between these parameters. The activation energy declines rapidly from around 1000 kJ mol−1 at the commencement of the heat treatment to 668 kJ mol−1 at αH = 0.22, then, decreases gently to 577 kJ mol−1 at αH = 0.98, during successive recrystallisation events. Average uncertainties in these results amount to 13 kJ mol−1. The frequency factors fall between 58.5 (±1.0) min−1 and 51.0 (±3.2) min−1, as computed at αH = 0.23 and 0.98, respectively. The so-called false compensation analysis reveals that the first-order reaction model (in αH) governs the energy demand for calcination for αH ≥ 0.23, but, initially, the transformation proceeds through the dissociation-diffusion regime that is not part of the established reaction models. This regime must not be ignored in modelling the calcination of α-spodumene, as it consumes around 20 % of energy required for the transformation reactions. The results reveal significant differences in the predictions of the treatment time, by more than two orders of magnitude, from the existing kinetic models, and explain the differences by the experimental conditions to collect the data for the models. The dissociation of Si-O bonds and diffusion of Si4+ ions out of their tetrahedral cages govern the onset of the thermal treatment of α-spodumene and account for the elevated values of the activation energy in the dissociation-diffusion regime. The two recrystallisation events are limited by the multicomponent diffusion, especially Si4+, in partly crystallised structure. The recrystallisation of γ- to β-spodumene defines the required retention time of concentrate particles in the kiln, to maximise the effectiveness of the subsequent recovery of lithium from the treated material. Fitting Eα and ln(fA)α to polynomials allows a convenient integration of the model equation for making predictions under any heating program. The model forecasts well the transformation of particles of α-spodumene characterised by d80 = 315 µm, studied in additional experiments, using the X-ray powder diffraction to quantify the conversion of α-spodumene. The predictions from the isoconversional model also concur well with other conversion measurements available in the literature, within the expected variability of different spodumene concentrates.

This paper presents the extraction process of lithium from lepidolite concentrate by roasting with NaHSO4 followed by water leaching. Results from TG-DSC analysis of feed material and XRD scans of solids justified the thermodynamic modelling of chemical reactions predicted using HSC software package. The roasting of NaHSO4/lepidolite concentrate mixture of mass ratio 2.5/1.0 at 500 °C followed by water leaching at 95 °C yielded 96% of Li extraction. Under these conditions, about 79% Rb and 73% Cs were also leached from calcined lepidolite material. Purification of the lepidolite leach liquor by adjusting the solution pH to 9 using caustic soda resulted in the precipitation of cryolite (NaAlF6) as a primary phase. The purified liquor with pH adjusted to 12 was treated with phosphoric acid to precipitate Li3PO4 to achieve a recovery of about 93%.

This contribution develops a process for recovering lithium from β-spodumene by sequential pressure leach with solutions of carbonic acid, as a replacement for digestion of β-spodumene with concentrated sulfuric acid. Six sequential steps, each operated at 200 °C and 100 bar for 1 h, retrieve 75 % of lithium from the initial feed material. This compares with a 12.6 % Li recovery from a single-stage (batch) operation. We develop both the thermodynamic and mass-transfer models of the leach process and argue that the lithium extraction is mass-transfer controlled. The underlying mechanism involves: (i) near-instantaneous ion exchange between H+ for Li+ at the onset of the digestion; (ii) direct formation of amorphous Li-depleted rinds; (iii) precipitation of secondary minerals, boehmite (AlO(OH)) and skin-forming amorphous silica (SiO2), the latter only in the batch operation; (iv) counter-current diffusion of H+ and Li+ through the surface layer (surface layer = depleted sublayer + precipitation sublayer); (v) dissolution of the surface layer at the solid-solution interface, through the reactions with water. The evidence comes from the near-proportionality between the lithium recovery and mass of dissolved β-spodumene, no detection of keatite-HAlSi2O6, particle morphology, precipitation of amorphous skins of SiO2 decorated with boehmite specks observed by electron microscopy and predicted from thermodynamic modelling as well as from a mass-transfer analysis of the Li-recovery rates. Direct observations confirm the existence of a three-phase reacting system, as predicted from thermodynamic calculations, with the absence of a supercritical fluid phase. Lithium recoveries, of at least 28 % in a single step, are attainable, but only for long contact time of 48 h. Digestion lasting hundreds of hours is needed for the batch process to reach Li extraction corresponding to that of the sequential leach. The challenges for industrial implementation of the process comprise recycling of large amounts of water, avoiding the precipitation of silica that forms rinds, accelerating the hydrolysis of Li-depleted sublayer, concentrating the dilute solutions of Li+ and overcoming hesitance of industry to consider high-pressure treatment.

The conventional processing of fl-spodumene involves decrepitation at about 1100 degrees C, digestion with concen-trated sulfuric acid at 250 degrees C, and several purification stages that identify the process with high energy, feed-stock, and by-product intensity. In addition to the low-value by-product of sodium sulfate (Na2SO4), the disposal cost of another by-product, hydrogen aluminosilicate (HAlSi2O6, aka keatite-HAlSi2O6), drives the research to avoid its formation in lithium refineries or converting this material into marketable products such as zeolites. Two processes are developed to leach lithium from fl-spodumene in a single digestion step using solutions of NaCl, with and without the addition of NaOH, operating at up to 200 degrees C. The optimised processes extract at least 92.9% of Li by varying the leaching time, temperature, liquid to solid mass ratio, NaCl/fl-spodumene mass ratio, solution pH and agitation. The two processes are simple, prevent using concentrated sulfuric acid and preclude forming HAlSi2O6 and Na2SO4 by-products, resulting in significantly lower amounts of impurities in the leach liquor, and high selectivity for lithium extraction. The keatite process functions at circumneutral pH, high NaCl/ fl-spodumene and L/S ratios and generates sodium aluminosilicate (keatite-NaAlSi2O6) as a by-product. The analcime process requires alkaline pH to produce analcime (NaAlSi2O6 center dot H2O) as a by-product. The study dem-onstrates a shift in the mechanism between circumneutral and alkaline conditions from ion exchange to dissolution-precipitation (i.e., recrystallisation). In an analogy to keatite and analcime processes for NaCl, we also investigate the extraction of lithium from fl-spodumene with KCl with and without KOH but find both re-actions to be ineffective in recovering Li. The analcime process offers a cleaner and less hazardous alternative to the conventional sulfuric acid process while maintaining a high lithium extraction effectiveness, rapid reaction rates and high concentration of Li in the leach.

Heat-treatment of serpentine minerals generates structural amorphicity and increases reactivity during subsequent mineral carbonation, a strategy for large-scale sequestration of CO2. This study employs thermal analyses (TGA-DSC) in conjunction with in-situ synchrotron powder X-ray diffraction (PXRD) to record concurrent mass loss, heat flow, and mineralogical changes during thermal treatment of antigorite. Isoconversional kinetic modelling demonstrates that thermal decomposition of antigorite is a complex multi-step reaction, with activation energies (Eα) varying between 290 and 515 kJ mol−1. We identify three intermediate phases forming during antigorite dehydroxylation, a semi-crystalline chlorite-like phase (γ-metaserpentine) showing an additional reaction pathway for the decomposition of Al2O3-rich antigorite into pyrope, and two distinct amorphous components (α and β-metaserpentine) which convert into forsterite and enstatite at higher temperature, respectively. The combination of isoconversional kinetics with in-situ synchrotron PXRD illustrates, for the first time, that local crystal structure changes, related to intermediate phase and forsterite formation, are responsible for the steep increase in activation energy above 650 °C and only 49% dehydroxylation can be achieved prior to this increase. This suggests that the high thermal stability of Al2O3-rich antigorite would severely limit Mg extraction during application of mineral carbonation under flue gas conditions.

With the rising demand for lithium-ion batteries, mining of lithium minerals and their refining to lithium chemicals has grown rapidly over the last years. At the same time, environmental impact and sustainability are increasingly important for lithium extraction from mineral sources to remain viable in the future. The conventional extraction of lithium from spodumene via calcination, acid roasting, leaching, purification and crystallisation achieves high lithium recoveries and lithium hydroxide purities but is also energy- and chemical-intense and produces large amounts of low-value by-products. Large proportions of spodumene ore are discarded as waste during mining and processing because of the strict quality specifications for spodumene concentrate imposed by the calcination step. Furthermore, melting and clinker formation during calcination can degrade lithium recovery from spodumene that has been partly altered to muscovite and cookeite. The presentation will illustrate strategies that address the above challenges, including improved understanding of calcination and the effect of alteration and low grades on lithium recovery, as well as an alternative extraction route that results in marketable by-products.

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 residual remnant NaOH for leaching.

In this study we present the first Mg isotope data that record the fate of Mg during mineralisation of atmospheric CO2 in ultramafic mine tailings. At the Woodsreef Asbestos Mine, New South Wales, Australia, weathering of ultramafic mine waste sequesters significant amounts of CO2 in hydromagnesite [Mg5(CO3)4(OH)2·4H2O]. Mineralisation of CO2 in above-ground, sub-aerially stored tailings is driven by the infiltration of rainwater dissolving Mg from bedrock minerals present in the tailings. Hydromagnesite, forming on the surface of the tailings, has lower δ26Mg (δ26MgHmgs = −1.48 ± 0.02‰) than the serpentinised harzburgite bedrock (δ26MgSerpentinite = −0.10 ± 0.06‰), the bulk tailings (δ26MgBulk tailings = −0.29 ± 0.03‰) and weathered tailings containing authigenic clay minerals (δ26MgWeathered tailings = +0.28 ± 0.06‰). Dripwater (δ26MgDripwater = −1.79 ± 0.02‰) and co-existing hydromagnesite (δ26MgHmgs = −2.01 ± 0.09‰), forming in a tunnel within the tailings, and nodular bedrock magnesite [MgCO3] (δ26MgMgs = −3.26 ± 0.10‰) have lower δ26Mg than surficial fluid (δ26Mg = −0.36‰) and hydromagnesite. Complete dissolution of source minerals, or formation of Mg-poor products during weathering, is expected to transfer Mg into solution without significant alteration of the Mg isotopic composition. Aqueous geochemical data and modelling of saturation indices, along with Rayleigh distillation and mixing calculations, indicate that the 26Mg-depletion in the drip water, relative to surficial water, is the result of brucite dissolution and/or precipitation of secondary Mg-bearing silicates and cannot be assigned to bedrock magnesite dissolution. Our results show that the main mineral sources of Mg in the tailings (silicate, oxide/hydroxide and carbonate minerals) are isotopically distinct and that the Mg isotopic composition of fluids and thus of the precipitating hydromagnesite reflects both isotopic composition of source minerals and precipitation of Mg-rich secondary phases. The consistent enrichment and depletion of 26Mg in secondary silicates and carbonates, respectively, underpins the use of the presented hydromagnesite and fluid Mg isotopic compositions as a tracer of Mg sources and pathways during CO2 mineralisation in ultramafic rocks.