Rorie Gilligan

Near-complete (>99%) dissolution of lithium and cobalt was achieved by the leaching of black mass from spent (end-of-life) lithium-ion batteries (LiBs) using 4 M H 2 SO 4 or HCl at 60 • C. Raising the temperature to 90 • C did not increase the overall extraction of lithium or cobalt, but it increased the rate of extraction. At 60 • C, 2 M H 2 SO 4 or 2 M HCl performed similarly to the 4 M H 2 SO 4 /HCl solution, although extractions were lower using 1 M H 2 SO 4 or HCl (~95% and 98%, respectively). High extractions were also observed by leaching in low pulp density (15 g/L) at 60 • C with 2 M CH 2 ClCOOH. Leaching was much slower with hydrogen peroxide reductant concentrations below 0.5 mol/L, with cobalt extractions of 90–95% after 3 h. Pulp densities of up to 250 g/L were tested when leaching with 4 M H 2 SO 4 or HCl, with the stoichiometric limit estimated for each test based on the metal content of the black mass. Extractions were consistently high, above 95% for Li/Ni/Mn/Cu with a pulp density of 150 g/L, dropping sharply above this point because of insufficient remaining acid in the solution in the later stages of leaching. The final component of the test work used leaching parameters identified in the previous experiments as producing the largest extractions, and just sulphuric acid. A seven-stage semi-continuous sulphuric acid leach at 60 • C of black mass from LiBs that had undergone an oxidising roast (2h in a tube furnace at 500 • C under flowing air) to remove binder material resulted in high (93%) extraction of cobalt and near total (98–100%) extractions of lithium, nickel, manganese, and copper. Higher cobalt extraction (>98%) was expected, but a refractory spinel-type cobalt oxide, Co 3 O 4 , was generated during the oxidising roast as a result of inefficient aeration, which reduced the extraction efficiency.

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.

Brannerite, UTi2O6, is the most common of the refractory uranium minerals and requires leaching under intense conditions to effectively extract the uranium contained within. Brannerite is often found together with apatite in metasomatic uranium deposits. Apatite and other phosphate gangue minerals can inhibit uranium leaching by generating phosphate ions which interfere with the reactions between acidic ferric sulphate and uranium minerals. As part of a detailed fundamental study on the leaching reaction mechanisms for brannerite, tests under a range of selected leach conditions were carried out with the addition of 10 g/L fluorapatite with the goal of identifying conditions where the negative effects of phosphate are reduced. Leaching was carried out for 5 hours with 0.05 mol/L Fe3+ as Fe2(SO4)3 with 0.25–1.00 mol/L H2SO4 and at temperatures between 25 and 96 °C. A single test was performed with 0.05 mol/L Fe3+ as FeCl3 and 1.00 mol/L HCl. In the sulphate system, the effect of phosphate was weakest at the highest acid concentration (1.00 mol/L H2SO4). In the chloride system, phosphate did not suppress uranium extraction, suggesting that HCl leaching could be a viable alternative for the leaching of high-phosphate refractory uranium ores.

The commercialisation of vanadium redox flow batteries for large scale electric energy storage and power grid stabilisation is expected to increase the global demand for vanadium in the coming years. Currently most of the vanadium is used in the production of steel alloys and this amount is expected to remain consistent in the years to come. Much of the new demand is expected to come in the form of electrolytes for the application of vanadium to energy storage in the vanadium redox flow batteries and in the form of ultrapure vanadium salts for use a precursor reagents in the production of cathodes for lithium ion batteries.

Unlike other metals such as copper, nickel or zinc, vanadium does not form concentrated deposits. Owing to the similarities between the V3+ and Fe3+ cations, vanadium is often found as a minor component of iron minerals. The vanadium mineral coulsonite, FeV2O4 forms series with chromite, FeCr2O4 and magnetite, Fe3O4. Most of the vanadium is produced from titanomagnetites either directly from titanomagnetite ores/concentrates or indirectly from the slag left from smelting titanomagnetite ores. Titanomagnetite ores are associated with mafic igneous rock and have been found in large quantities in Russia, China, South Africa and other countries. Vanadium can also be produced from vanadiferous sandstone, shale and vanadate deposits, though these are of lesser industrial importance compared to titanomagnetites.

Vanadium is usually extracted from ores, concentrates and slags by roasting with sodium carbonate or another sodium salt to convert vanadium into water soluble sodium vanadates. This review summarises the established and proven processes for vanadium production as well as some newer processes which have yet to be commercialised.

In this study, synthetic zirconolite samples with a target composition Ca0.75Ce0.25ZrTi2O7, prepared using two different methods, were used to study the stability of zirconolite for nuclear waste immobilisation. Particular focus was on plutonium, with cerium used as a substitute. The testing of destabilisation was conducted under conditions previously applied to other highly refractory uranium minerals that have been considered for safe storage of nuclear waste, brannerite and betafite. Acid (HCl, H2SO4) leaching for up to 5 h and alkaline (NaHCO₃, Na2CO3) leaching for up to 24 h was done to enable comparison with brannerite leached under the same conditions. Ferric ion was added as an oxidant. Under these conditions, the synthetic zirconolite dissolved much slower than brannerite and betafite. While the most intense conditions were observed previously to result in near complete dissolution of brannerite in under 5 h, zirconolite was not observed to undergo significant attack over this timescale. Fine zirconolite dissolved faster than the coarse material, indicating that dissolution rate is related to surface area. This data and the long term stability of zirconolite indicate that it is a good material for long-term sequestration of radioisotopes. Besides its long term durability in the disposal environment, a wasteform for fissile material immobilisation must demonstrate proliferation resistance such that the fissile elements cannot be retrieved by leaching of the wasteform. This study, in conjunction with the previous studies on brannerite and betafite leaching, strongly indicates that the addition of depleted uranium to the wasteform, to avert long term criticality events, is detrimental to proliferation resistance. Given the demonstrated durability of zirconolite, long term criticality risks in the disposal environment seem a remote possibility, which supports its selection, above brannerite or betafite, as the optimal wasteform for the disposition of nuclear waste, including of surplus plutonium.

The project, part of a major collaboration between the Minerals Research Institute of Western Australia (MRIWA), Murdoch University and sponsors Lithium Australia and Venus Metals Corp, sought to develop new technology for battery grade lithium production. The project also aimed to capture valuable byproducts from the extraction of high-grade lithium carbonate from certain groupings of minerals, known as micas.

The acidic ferric sulphate system is the industrial standard for the acid leaching of uranium ores. Alternatives such as ferric chloride/hydrochloric acid have been investigated, but never used industrially. Selected ferric sulphate leaching experiments from previous work were repeated in acidic ferric chloride media to evaluate the effectiveness of this alternative system. Leaching experiments were run for 5 h and the effects of acid concentrations between 0.25 and 2.00 mol/L and temperatures between 25 and 96 °C were studied. A single experiment for each system was repeated with the addition of 10 g/L fluorapatite, known to inhibit uranium leaching in ferric sulphate media.

Ferric chloride media was less effective than ferric sulphate media for the extraction of uranium from brannerite at same temperature and acidity, except at high (2.00 mol/L) acid concentrations. Acidity had a significantly greater effect on the rate of dissolution in the chloride media. The rate of the dissolution reaction was first order with respect to acid concentration in the chloride system, compared to an order of approximately 0.5 with respect to acid concentration in the ferric sulphate system. This suggests that dissolution in chloride media takes place via a different mechanism in which the acid has a greater role in the leaching reaction.

Comparing the results from the chloride and ferric sulphate leaching suggests that the formation of stable uranyl complexes is a key component in the leaching process, as uranyl sulphate complexes are stronger than uranyl chloride complexes. Interestingly, the alternative system was less susceptible to interferences from phosphates, and therefore may be a viable alternative when dealing with high-phosphate refractory uranium ores.

There are several metasomatic uranium deposits in the area around Mount Isa in Queensland, containing a total of 56,400 t of uranium. Many of these ores are refractory in nature, meaning that relatively high leach temperatures (> 75 °C) and reagent dosages (> 50 g/L H2SO4) are needed to effectively extract the uranium from them. Also, these ores are hosted in alkaline rock which means that acid leaching is unlikely to be economical.

While refractory uranium ores are not typically leached in alkaline media, previous work has shown that uranium can be extracted from brannerite in alkaline media, albeit slowly. The same leaching conditions previously shown to be effective for the extraction of uranium from brannerite were repeated with a sample of refractory uranium ore from a deposit near Mount Isa in north western Queensland. Mineralogical analysis with a Tescan Integrated Mineralogical Analyser (TIMA) showed that the uranium was present as brannerite (51%) and coffinite (49%).

The ore was leached in sodium carbonate media for 24 h. Leach temperatures of 50, 70 and 90 °C, and sparging with oxygen, air and nitrogen were tested. The effect of adding ferricyanide chemical oxidant was also tested. Similar initial uranium extraction rates were observed for the Mount Isa uranium ore compared with Sierra Albarrana brannerite leached under the same conditions in earlier work. The final extractions were lower however, due to the fine-grained nature of the uranium mineralisation in the ore.

TIMA analysis on a residue produced under the most intense set of leaching conditions (90 °C, O2 sparged) showed that the extent of coffinite dissolution was greater than that of brannerite dissolution during leaching. Likewise, the TIMA analyses showed that the extent of apparent uranium liberation decreased during leaching, due to poorly liberated particles remaining undissolved. These results indicate that alkaline leaching could be a viable method for the processing of these long overlooked ores provided that the uranium is sufficiently liberated.

Refractory uranium ores containing uranium as multiple oxides typically require leaching at elevated temperatures (> 75 °C) and acid concentrations (> 50 g/L H2SO4) in order to effectively extract the uranium. Some uranium ores contain large amounts of acid-soluble gangue, such as carbonates or soluble silicates and when the carbonate content exceeds approximately 8%, the acid consumption by gangue may render acid leaching uneconomical. In these situations, leaching in alkaline-carbonate is a potential option. The advantages of this approach include improved selectivity for uranium over other elements, reduced gangue dissolution and the ability to recycle the lixiviant. One disadvantage is slower rates of leaching and hence the need for finer grinding. However, when a uranium ore contains significant amounts of acid-soluble gangue with refractory uranium mineralisation (a double refractory ore), the process options are more limited. Refractory uranium minerals, such as brannerite, are often reported as slow to dissolve in alkaline carbonate media. If a double refractory uranium ore is to be processed effectively, it is necessary to understand the behaviour of minerals like brannerite in the alkaline leaching system. In the present study, a sample of brannerite was leached in a sodium carbonate solution (1 mol/L total carbonate) for 24 h at temperatures from 50 to 90 °C. The effect of potassium ferricyanide (25 mmol/L) added as an oxidant was also examined. All residues were characterised by XRD and SEM-EDX techniques, and both uranium and titanium dissolution rates were monitored. Uranium extraction reached 83% after 24 h of leaching. The leaching rate showed a high dependence on temperature with an activation energy of 45 kJ/mol. The residue was pitted, similar to what has been observed previously after leaching in acidic media. At 80 °C and 90 °C, the titanium re-precipitated within the pits, potentially slowing the dissolution process.

Brannerite (UTi2O6) is the most important uranium mineral after uraninite and coffinite, and the most common refractory uranium mineral. As the more easily leachable uranium ores are becoming exhausted, it is necessary to process the complex and refractory ores in order to meet the growing demand for uranium as an energy source. This typically requires either more intense leaching conditions or a better-designed process based on sound understanding of feed mineralogy and reaction chemistry. The present study was carried out to provide information that will enable the development of a more effective processing strategy for the extraction of uranium from ores containing brannerite. The leaching behaviour of brannerite in sulphate media under moderate temperature conditions was investigated and compared with its relative leachability in alternative acid and alkaline systems. The feed and the leached residues were characterized by XRD and SEM-EDX techniques. Brannerite dissolutions of up to 95% after 5 hours of leaching in ferric sulphate media, up to 89% in ferric chloride media under similar conditions, and up to 82% in 24 hours in sodium carbonate media were obtained. Since alkaline leaching was considered promising for acid-consuming ores, leaching was repeated with a high-carbonate brannerite-bearing ore, with comparable extractions. Mineralogical characterization showed that altered and amorphous regions are a regular feature of brannerite, and that pitting is typically observed on the surface of the leached grains. The leaching results, coupled with mineralogical data, showed that the uranium and titanium in brannerite generally dissolve congruently, with faster dissolution of the altered and amorphous regions in the brannerite grains than of the crystalline regions. We conclude that the extent of brannerite alteration is a key factor in process selection, along with the grade, liberation size, and gangue mineralogy.