Gamini Senanayake

PARK7 mRNA encodes DJ-1 protein, which functions as a protective agent against oxidative stress and cell damage within the brain cells. Mutations in the mRNA can lead to reduced production of DJ-1 and initiate brain diseases such as Parkinson’s disease. Transport of appropriate mRNA to damaged brain cells may provide a suitable treatment. Mesoporous silica nanoparticles (MSNPs), particularly pore-expanded and dye-labeled varieties, are regarded as potential carriers for large therapeutic agents such as mRNA. This study explored the influence of alterations in reaction conditions on the structural characteristics of MSNPs to produce nanoparticles with favorable characteristics for delivering large therapeutic agents to target sites. One-stage and two-stage procedures were compared for the introduction of 3-aminopropyltriethoxysilane (APTES) and APTES-dye adduct, in conjunction with two different surfactants, cetyltrimethylammonium bromide (CTAB) and cetyltrimethylammonium chloride (CTAC). Analysis of the MSNPs shows that the two-stage method using CTAB as a surfactant produced amine-functionalized, dye-labelled particles with smaller overall size and better uniformity than the one-stage approach. However, due to their small pore size (<10 nm), these particles were unable to encapsulate the PARK7 mRNA (926 nucleotides). The one-stage method via CTAC produced MSNPs with large (150 nm), broad pore distribution (10–20 nm), and high aggregation, limiting their suitability for brain-targeted gene delivery. In comparison, the two-stage method using CTAC yielded well-ordered MSNPs with an optimal size (80 nm) and pore diameters (15–20 nm), enabling effective encapsulation of the large PARK7 mRNA and offering strong potential for future brain gene therapy studies.

The sulfuric acid bake followed by water leaching is now well-established technology used for processing of ores and concentrates containing rare earth phosphate mineralization. Gangue minerals can have a significant impact on the process. In this work, we tease out the impact of the presence of iron minerals on the monazite acid bake process, focusing particularly on magnetite. Sulfuric acid baking and leaching of a magnetite sample alone and a magnetite/monazite mixture at bake temperatures ranging between 200 and 800 °C was conducted. The presence of magnetite with monazite had a significant impact on the extraction of rare earth elements and impurities from monazite and these were explained by the phases formed in the bake.

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.

Activated carbon (AC) products are valuable for adsorption applications due to their high internal pore volume, surface area, surface functionality and strong adsorbent properties, which allow the capture of various ions and impurities in many applications, including precious metals recovery, water treatment, and gas purification. This review found that most of the world's AC is derived from bituminous and sub-bituminous coal, but low-rank coals are more suitable for AC production due to their higher reactivity to activation and increased porosity compared to high-rank coals. Coal-based AC is preferred for adsorption due to its surface functionality, porous structure, lower cost, suitability for single-use without reactivation, and the potential for regeneration. A positive correlation between carbonisation temperature and BET (Brunauer-Emmett-Teller) surface area and the impact of KOH and NaOH percentages on AC yield and surface area is identified. AC production, quality and yield are affected by coal type, rank, particle size, chemical composition, and process conditions such as activation method, activation temperature, and pre- and post-treatment steps. Properties of AC, such as surface area and iodine number, are higher in chemically activated carbon than in thermally activated carbon. The sustainable use of ilmenite/iron oxide activation yields AC suitable for water treatment due to its high iodine content. Low-rank coals, including those with high ash content, could be utilised with green collectors or gravity separation to minimise ash content.

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.

Sulfuric acid baking of monazite ((Ce,La,Th)PO4) bearing ores is one of the major processes used in commercial production of rare earth elements. In hydrothermal vein type rare earth deposits, apatite is often found together with monazite. Rare earth enriched fluorapatite ores may also contain significant amounts of rare earths hosted in monazite. An understanding of the effect of apatite on the sulfuric acid baking of monazite is therefore important for the development of effective sulfuric acid based treatment methods for such ores. In this work, the addition of a natural hydroxyapatite (Ca5(PO4)3(OH)) sample to a monazite acid bake followed by acid leach was investigated using a combination of chemical analyses, SEM-EDS, XRD and TG-DSC. Hydroxyapatite addition significantly decreased the dissolution of rare earth elements from monazite in the leach after baking at temperatures above 300 °C. For a bake temperature of 500 °C, the rare earth dissolution in the leach dropped from 80% for monazite alone, to 30% for a 1:1 (w/w) addition of hydroxyapatite. This decrease in rare earth leaching was attributed to the formation of an insoluble thorium and rare earth bearing polyphosphate. The rare earth elements were incorporated into this polyphosphate phase in preference to calcium. At 800 °C, monazite was re-formed, causing a further reduction in rare earth extraction, while simultaneous formation of calcium pyrophosphate (Ca2P2O7) led to an increase in calcium and phosphorus dissolution. The detrimental effect of apatite could be partially overcome by the addition of goethite. Addition of goethite to the acid bake of a monazite/apatite mixture at 500 °C improved the total rare earth dissolution from 29% to 85% in the subsequent leach. Results also demonstrated that the order of reactivity, in terms of formation of polyphosphates is Fe > REE > Ca.
•Identified negative impact of apatite on rare earth leaching after acid bake at >300 °C.•Phases responsible for altered leaching behaviour characterised.•Demonstrated ability of added goethite to offset negative impact of apatite on monazite acid bake.

Monazite, a rare earth phosphate mineral, is the second most important primary source of rare earth elements (REEs). Current technologies for processing monazite ore using sulfuric acid are primarily focused on REEs recovery. However, these technologies result in the loss of phosphorus in waste streams. Therefore, there is a need for efficient processing methods that could recover both REEs and phosphorus from monazite ore. This study presents a method for recovering both REEs and phosphorus as potential feed material for iron phosphate battery precursor from monazite ore by sulfuric acid baking with the addition of sulfate salts. The leaching efficiency of REEs and phosphorus varied depending on the additive used, with the highest efficiencies observed for the ferric sulfate system. As the temperature increased, the leaching efficiency of REEs and phosphorus decreased when baking with no additive. However, the addition of ferric sulfate salt to the baking reactants improved leaching efficiency of REEs and favourably enriched P in the residue for subsequent processing. The XRD confirmed the successful constraining of P and Fe in the residue while more than 95% REEs were selectively leached. The results suggest that this method could be a promising alternative to conventional methods for processing monazite ore. An integrated flowsheet was proposed to produce a marketable rare earth oxide product of over 99% purity.

Thermochemical splitting of a mixture of CO2 and water into syngas (and methane) remains a viable approach toward an industrial-scale treatment of CO2 emission. However, most deployed catalysts encompass expensive metallic ingredients such as Pd or Pt. In an integrated experimental-modelling approach, this work reports a high conversion of CO2 over alternative and cost-effective configurations. Ceria-based catalysts are proven to be effective in numerous catalytic processes owing to the facile switch in the Ce redox cycle. Evolution of H2 and CO from thermochemical splitting of water and carbon dioxide is an important process in the production of syngas and in fuel cell applications. In this study, NbOx ceria catalysts are prepared, characterized, and applied for production of syngas via the thermochemical splitting of CO2 and gaseous water. It was observed that presence of Nb in the ceria matrix up to 12 wt%, increases the reaction yield at higher temperature (up to 79% CO2 conversion with 80% selectivity to syngas at 600 ºC). Moreover, the catalytic reaction was found to display a higher selectivity towards syngas over the ceria supported catalysts. The attained conversion is higher than other ceria-supported catalysts such as Rh-CeO2, CeO2-ZrO2, and V2O5-CeO2. Mapped-out reaction pathways by DFT calculations portray accessible routes into syngas. Results provided herein should be useful to optimize a continuous process for the valorization of CO2 into syngas over relatively affordable a catalytic formulation.

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.