Stripping conditions to prevent the accumulation of rare earth elements and iron on the organic phase in the solvent extraction circuit at Skorpion Zinc

The concentrations of the rare earth elements (particularly Y, Yb, Er and Sc) in the zinc-stripped organic phase streams in the solvent extraction process at Skorpion Zinc mine have increased gradually over the past four years. Iron is the only other impurity present in notable quantities in the organic phase after washing and scrubbing prior to zinc stripping. This project aimed to evaluate the effects that rare earth elements and iron in the organic phase have on the zinc solvent extraction process and to subsequently find appropriate stripping conditions for the removal of these elements from the zinc-stripped organic phase.Results obtained by performing laboratory scale batch tests indicated that the viscosity of the organic phase doubled and the phase disengagement time increased from 100 s to 700 s when the total rare earth elements and iron concentration in the organic phase was increased from 3100 mg/L to 6350 mg/L. The zinc loading capacity of the organic phase after two extractions furthermore decreased by a value between 1 g/L and 3 g/L, depending on the composition of the pregnant leach solution. The stripping of low concentrations of rare earth elements and iron from 40% di-2-ethylhexyl phosphoric acid (D2EHPA) diluted in kerosene was evaluated using two different stripping agents (H₂SO₄ and HCl) with concentrations between 1 M and 7 M, organic-to-aqueous ratios between 0.5 and 4, and temperatures between 30 °C and 50 °C. The highest stripping percentages were achieved at acid concentrations greater than 5 M, organic-to-aqueous ratios of less than 0.5 and high temperatures.     

Research on the recycling of valuable metals in spent Al2O3-based catalyst

A new technology was developed to recover multiple valuable elements in the spent Al₂O₃-based catalyst by X-ray phase analysis and exploratory experiments. The experiment results showed: In the condition of roasting temperature of 750 °C and roasting time of 30 min, mol ratio of Na₂O: Al₂O₃ 1.2, the leaching rate of alumina, vanadium and molybdenum in the spent catalyst is 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by barium hydroxide and barium aluminate, the precipitation rate of vanadium and molybdenum is 94.8% and 92.6%. Al(OH)₃ is prepared from sodium aluminate solution with carbonation decomposition process, and the purity of Al₂O₃ is 99.9% after calcinations, the recovery of alumina can reach 90.6% in the whole process. The Ni–Co concentrate was leached by sulfuric acid, a nickel recovery of 98.2% and over 98.5% cobalt recovery was obtained respectively under the experimental condition of 30% (w/w) H₂SO₄, 80 °C, reaction time 4 h, liquid:solid ratio (8:1) by weight, stirring rate of 800 rpm.     

Study on sulphidization roasting and flotation of cervantite

A new technology, sulphidization roasting of antimony mineral cervantite with elemental sulfur followed by froth flotation is reported in this paper. The effects of roasting temperature and time, sulfur to antimony molar ratio on the properties of treated product and its flotation behavior were studied. Optimum roasting conditions are: roasting temperature 723 K; roasting time 30 min; and sulfur to antimony molar ratio of 1.5. Under these conditions, the mineral phase changed from cervantite to stibnite as expected. The flotation recovery of the sulphidized cervantite is over 90%. A flotation concentrate grading 21.04% Sb with a recovery of 77.15% is achieved by sulphidization roasting and flotation from a feed grading 1.11% Sb in which cervantite is the main antimony mineral.     

A novel method to recover zinc and iron from zinc leaching residue

A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe₃O₄ under CO, CO₂ and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO₂ ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR.     

Dissolution of platinum in catalyst materials using hydrochloric acid: A new method based on the use of complex oxides

Platinum group metals (PGMs) are used widely in various applications, including as environmental catalysts. Since PGMs are rare and expensive, they are recycled after being recovered. Currently, PGMs are recovered, after a pyrometallurgical step that upgrades the concentration, by dissolving in strong acids containing toxic oxidizing agents like aqua regia. To avoid the use of such toxic agents, we have proposed a new route to dissolve PGMs in hydrochloric acid (HCl) using complex oxides. In the present study, we used this new process to dissolve Pt in supported metal catalysts. Complex oxides of Pt were prepared by calcining mixtures of Pt/Al₂O₃ and alkali metal salts at 600–800 °C in air. These were then dissolved in 12 M HCl. The results showed that the Pt in the calcined samples dissolved readily in HCl and that the Pt solubility was nearly 100% under the appropriate preparation and dissolution conditions. We also confirmed that the new process is suitable for leaching PGMs from spent automotive catalysts.     

Flotation of rare earth minerals from silicate–hematite ore using tall oil fatty acid collector

The flotation of rare earth (RE) minerals (i.e. xenotime, monazite-(Nd), RE carbonate mineral) from an ore consisting mainly of silicate minerals (i.e. primary silicate minerals and nontronite clay) and hematite was investigated using tall oil fatty acids (Aero 704, Sylfat FA2) as collector. The RE minerals are enriched with Fe. The effects of tall oil fatty acid dosage, pH, temperature, and conventional depressants (sodium lignin sulfonate, sodium metasilicate, sodium fluoride, sodium metasilicate and sodium fluoride, and soluble starch) were determined at grinding size of P80 = 63 μm. At this grinding size, the grain size of the RE minerals ranges from 2 to 40 μm, percentage liberation is 9–22%, and percentage association with nontronite and quartz is 30–35%. Results indicated that Sylfat FA2 at 22450 g/t concentration was the more efficient tall oil fatty acid collector at natural pH (pH 7) to basic pH (pH 10.0–11.5). Flotation at the room temperature (25 °C) gave higher selectivity than 40 °C temperature flotation. The results on the effect of depressants showed similar selectivity curves against the gangues SiO₂, Al₂O₃, and Fe₂O₃ suggesting that the chemical selectivity of the depressants has been limited by the incomplete liberation of the RE minerals in the feed sample. High recoveries at 76–84% (Y + Nd + Ce)₂O₃ but still low (Y + Nd + Ce)₂O₃ grade at 2.1% in the froth were obtained at flotation conditions of 63 μm, 25 °C, pH 10.5, 1,875 g/ton sodium metasilicate and 525 g/ton sodium fluoride or 250 g/ton soluble starch as depressant for the silicates and hematite, and 22,450 g/t Sylfat FA2 as collector for the RE minerals (initial (Y + Nd + Ce)₂O₃ feed grade = 0.77%). The recoveries of gangue SiO₂, Al₂O₃, and Fe₂O₃ in the froth were low at 25–30%, 30–37%, and 30–36%, respectively. The mineralogical analysis of a high grade froth and its corresponding tailing product showed that the RE minerals have been concentrated in the froth while the primary silicate minerals and hematite have been relatively concentrated in the tailing. However, the clay minerals, primary silicate minerals, and hematite still occupy the bulk content of the froth. This suggests that incomplete liberation of the RE minerals led to the poor grade result, supporting likewise the selectivity curve results by the different depressants. This study showed that liberation is important in achieving selective separation.     

Chemical and mineral transformation of a low grade goethite ore by dehydroxylation,reduction roasting and magnetic separation

The utilization of abundant low grade goethite (α − FeOOH) ores is potentially important to many countries in the world, especially Australia. These ores contain many detrimental impurities and are difficult to upgrade to make suitable concentrates for the blast furnace. In this paper, chemical and mineral transformations of a goethite ore were studied by dehydroxylation, reduction roasting in CO and CO₂ gas mixtures, and magnetic separation. The goethite sample was taken from a reject stream at an iron ore mine from the Pilbara region, Western Australia. The roasting temperature range investigated was 400–700 °C. Chemical and mineralogical analysis was conducted using XRF, XRD, optical microscope, EPMA, and SEM. Magnetic separation was conducted using a Davis tube tester and a high intensity magnetic separator.The results show that reduction roasting can remove moisture and impurities but does not significantly change the Fe content in the feed. However, reduction roasting transforms goethite to hematite and eventually maghemite which can be recovered by magnetic separation, allowing upgrading. Further studies are needed to optimize the reduction roasting and correlate it with the magnetic separation to maximize the efficiency of iron upgrading.     

Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

Leaching of chalcopyrite (CuFeS2) with an imidazolium-based ionic liquid in the presence of chloride

The effects of independent variables such as, temperature, concentration of ionic liquid (1-butyl-3-methyl-imidazolium hydrogen sulphate, [bmim][HSO₄]), chloride and sulphuric acid on copper extraction from chalcopyrite (CuFeS₂) ore were studied by surface optimization methodology. The Central Composite Face approach and a quadratic model were applied to the experimental design. The optimal copper extraction conditions given by the above methodology were 20% (v/v) of [bmim][HSO₄] in water, 100 g L⁻¹ chloride, and 90 °C. The concentration of chloride and the temperature together exert a synergistic effect in enhancing chalcopyrite dissolution. Experimental data were fitted by multiple regression analysis to a quadratic equation and analyzed statistically. A model was developed for predicting copper extraction from CuFeS₂ ore with variables such as Cl⁻, [bmim][HSO₄], H₂SO₄ concentrations and temperature in the range studied. The activation energy was calculated to be 60.4 kJ/mol (temperature range 30–90 °C), indicative of chemical control of the reaction and [bmim][HSO₄] acts as an acid in the reaction.     

Selective separation of fine albite from feldspathic slime containing colored minerals (Fe-Min) by batch scale dissolved air flotation (DAF)

In this study, the separation of feldspar minerals (albite) from slimes containing feldspar and iron containing minerals (Fe-Min) was studied using dissolved air flotation (DAF) technique whereby bubbles less than 100 μm in size are produced. Before the flotation experiments with slimes, single flotation experiments with albite and Fe-Min were carried out using DAF in order to obtain optimum flotation conditions for the selective separation of feldspar from the slimes. Flotation experiments were performed with anionic collectors; BD-15 (commercial collector) and Na-oleat. The two methods of reagent conditioning were tested on the flotation performance; traditional conditioning and charged bubble technique. In addition, the effect of pH, flotation time, rising time, and drainage time which influence the selective separation in the DAF system were studied in detail. Overall, the flotation results indicated that the separation of albite from Fe-Min can be achieved with DAF at 5 min of rising time and 5 min of drainage time. Interestingly, these results also showed that the conditioning of the particles with the charged bubbles increased the flotation recovery of Fe-Min compared to the traditional conditioning. Furthermore, the flotation tests with the feldspathic slime sample were carried out under the optimum conditions obtained from the systematic studies using the single minerals. The charged bubble technique produced an albite concentrate assaying 0.33% Fe₂O₃ + TiO₂ and 11.07% Na₂O + K₂O from a slime feed consisting of 1.06% Fe₂O₃ + TiO₂ and 10.36% Na₂O + K₂O.     

Metals recovery from galvanic sludge by sulfate roasting and thiosulfate leaching

The purpose of this work is the selective recovery of Au, Ag, Cu, and Zn from two types of galvanic sludge using a mixed process of sulfate roasting and sodium thiosulfate leaching. In the experiments, the sludge was mixed with a sulfate promoter (sulfur, iron sulfate, or pyrite) and treated by pyrometallurgical processes at temperatures up to 750 °C. At this stage, this agent is thermally oxidized, turning the furnace atmosphere into a reducing one and the metallic oxides into water-soluble sulfates. Afterward, the sulfates can be treated by leaching with water for recovery of Ag, Cu, and Zn. The gold does not form sulfates in this reaction and was recovered through a second leaching stage using sodium thiosulfate, an effective reagent and less harmful to the environment than cyanide. Different parameters such as the sulfate promoter that achieves the highest recovery of metals, the proportion of galvanic sludge to sulfating agent, the temperature, the heating time in the oven, and the leaching time were evaluated. Additionally, a comparison of gold recovery using cyanide versus sodium thiosulfate was performed. The configuration that showed the best metal recovery included a 1:0.4 ratio of sludge to sulfur, an oven temperature of 550 °C, a roasting time of 90 min, and a water leaching time of 15 min. Using these parameters, recovery rates of 80% of the silver, 63% of the copper, and 73% of the Zn were obtained. The sodium thiosulfate leaching resulted in a recovery of 77% of the Au, close to the values obtained using cyanide.     

Solvent extraction of tetravalent hafnium from acidic chloride solutions using 2-ethyl hexyl phosphonic acid mono-2-ethyl hexyl ester (PC-88A)

Solvent extraction of Hf(IV) from acidic chloride solutions has been carried out with PC-88A as an extractant. Increase of acid concentration decreases the percentage extraction of metal indicating the ion exchange type mechanism. The plot of logD vs log[extractant], M is linear with slope 1.8 indicating the association of two moles of extractant with the extracted metal species. Plot of logD vs log[H⁺] gave a straight line with a negative slope of ∼2 indicating the exchange of two moles of hydrogen ions for every mole of Hf(IV). The effect of Cl⁻ ion concentration at constant concentration of [H⁺] did not show any change in D values. Addition of sodium salts enhanced the percentage extraction of metal and follows the order NaSCN > NaCl > NaNO₃ ∼ Na₂SO₄. Stripping of metal from the loaded organic (LO) with different acids indicated sulphuric acid as the best stripping agent. Regeneration and recycling capacity of PC-88A, temperature, extraction behavior of associated elements was studied.     

Oxidation of Fe(II) in a synthetic nickel laterite leach liquor with SO2/air

Under specific controlled conditions, the addition of SO₂ to oxygen or air produces the peroxy-monosulphate free radical in solution, which is a stronger oxidant than oxygen alone. In this study, the practical strategies required to optimise the oxidation of Fe(II) with SO₂/air was investigated at 75 °C as part of a process to remove iron as Fe(III) oxides from a synthetic nickel laterite high pressure acid leach solution containing 5 g/L Fe(II), 1 g/L Fe(III), 8 g/L Ni, 30 g/L Mg in sea water at pH about 2. The rate of Fe(II) oxidation was optimised in the pH range of 1.2–2.0 with respect to SO₂/air ratio and gas flow rates for minimum production of H₂SO₄ and maximum utilisation of SO₂. In order to minimise the air flow rates into the reactor vessel, the maximum rate of SO₂ addition that could be employed with air was established whilst maintaining oxidising conditions. The results provide strategies for commercial applications of the SO₂/air oxidising system and indicate important factors for reactor design.     

Solvent extraction and separation of palladium(II) and platinum(IV) from hydrochloric acid medium with dibutyl sulfoxide

The solvent extraction and separation performances of Pd(II) and Pt(IV) from hydrochloric acid solutions were investigated using dibutyl sulfoxide (DBSO) diluted in kerosene. Pd(II) was strongly extracted by a lower concentration DBSO in a lower concentration hydrochloric acid solution while the reverse was obtained for Pt(IV) extraction. Based on independent extraction and separation experiments of Pd(II) and Pt(IV), the separation parameters of Pd(II) and Pt(IV), including dibutyl sulfoxide concentration, contact time of aqueous and organic phases, organic/aqueous (O/A) phase ratio and H⁺ concentration of aqueous phase, were studied in detail, and the optimal separation parameters were obtained and summarized as the following: dibutyl sulfoxide concentration 0.6–1.2 mol dm^?3, organic/aqueous (O/A) phase ratio 0.6–1.0, H⁺ concentration of aqueous phase 1.0–1.5 mol dm^?3 and contact time of two phases 5 min. The as-prepared separation parameters were corroborated by the extraction and separation from a synthetic stock solution containing Pd(II), Pt(IV) as well as several common impurities like Fe(II), Cu(II) and Ni(II). The results revealed that Pd(II) could be separated efficiently from Pt(IV) with a high separation coefficient of Pd(II) an Pt(IV) (2.7 × 10⁴) by predominantly controlling dibutyl sulfoxide and hydrochloric acid concentrations. The extraction saturation capacity of Pd(II) was determined from 1.0 mol dm^?3 HCl solution with 3 mol dm^?3 dibutyl sulfoxide and its experimental value exceeded 14 g dm^?3 under the experimental conditions.Stripping of Pd(II) from loaded organic phase was performed using a mixed aqueous solution containing NH₄Cl and ammonia solutes. Pd(II) (99.2%) was stripped using the stripping solution containing 3% (m/v) NH₄Cl and 5 mol dm^?3 ammonia, respectively.     

Process development for the recovery of europium and yttrium from computer monitor screens

This work describes the development of a process for the recovery of Eu and Y from cathode ray tubes (CRTs) of discarded computer monitors with the proposition of a flow sheet for the metals dissolution. Amongst other elements, europium and yttrium are presented in the CRTs in quantities – 0.73 w/w% of Eu and 13.4 w/w% of Y – that make their recovery worthwhile. The process developed is comprised of the sample acid digestion with concentrated sulphuric acid followed by water dynamic leaching at room temperature. In the CRTs, yttrium is present as oxysulphide (Y₂O₂S) and europium is an associated element – Y₂O₂S:Eu³⁺ (red phosphor compound). During the sulphuric acid digestion, oxysulphide is converted into a trivalent Eu and Y sulphate, in solid form, with the liberation of H₂S. In the second step, metals are leached from the solid produced in the acid digestion step by dynamic leaching with water. This study indicates that a proportion of 1250 g of acid per kg of the sample is enough to convert Eu and Y oxysulphide into sulphate. After 15 min of acid digestion and 1.0 h of water leaching, a pregnant sulphuric liquor containing 17 g L⁻¹ Y and 0.71 g L⁻¹ Eu was obtained indicating yield recovery of Eu and Y of 96% and 98%, respectively. Both steps (acid digestion and water leaching) may be performed at room temperature.     

The role of citric acid in the flotation separation of rare earth from the silicates

The Nechalacho project is the most advanced large heavy rare earth elements (HREE) project outside of China. Open circuit and locked cycle flotation tests along with pilot plant testing of rare earth elements (REE) concentration from the host rocks are accomplished with collectors of alkyl phosphates and the modifier of citric acid. In this study, the function of citric acid in the separation of rare metals against silicates is investigated by a combination of micro-flotation tests and time of flight secondary ion mass spectrometry (ToF-SIMS) surface chemical analysis. It was observed that there was little effect of citric acid on the REE recovery in the micro-flotation tests conditioned with de-ionized water (DIW). To evaluate the flotation response with excess secondary ions in the pulp, micro-flotation tests were performed to look at changes in recovery as a result of adding Al ions and the subsequent presence of citric acid. The results from three micro-flotation tests (DIW, DIW with the addition of 100 mg/L Al and DIW + 100 mg/L Al and 500 g/t citric acid) revealed that the addition of Al ions led to a decrease of REE grade, a lower REE minerals recovery and/or an unexpected promotion of silicates to the concentrate. Citric acid reduced the negative effect generated by the Al ions in the flotation, which was shown by an improvement in REE grade. ToF-SIMS surface analysis of undifferentiated grains from the tests with and without citric acid revealed that grains reporting to the concentrate are doing so in response to collector attachment in combination with having more secondary Al on their surface. Citric acid may partially form aqueous soluble metal–ligand complexes resulting in less Al ions on the grains surface, which were rejected to the tailings. Citric acid also may form chelation competing for adsorption on gangue minerals, resulting in a diminished effectiveness of the activation site.     

An environmentally-friendly technology of vanadium extraction from stone coal

A novel technology characterized by higher recovery of vanadium and which was environmentally-friendly was developed to recover vanadium from stone coal. Vanadium in stone coal could be leached by NaOH solution after roasting stone coal at 850 °C for 3 h. H₂SO₄, Mg(NO₃)₂ and ammonia were employed, respectively, in two steps to remove the impurities of Si and Al from the leach liquor. After extracting vanadium from the leach liquor with 10 vol% N₂₃₅, 20 vol% secondary octyl alcohol and 70 vol% sulfonated kerosene, 1.5 mol/L NaOH was used as a stripping agent to strip vanadium from extracting solution. Adding 80 g/L NH₄NO₃ to the stripping solution at 30–40 °C and pH 7.5, vanadium could be crystallized as ammonium metavanadate. Roasting ammonium metavanadate at 540 °C for 1 h, the purity of V₂O₅ met the standard specification. The total recovery of vanadium reached 67.39%, which was higher than the classical technology.     

Copper leaching from chalcopyrite concentrate in Cu(II)/Fe(III) chloride system

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.     

Amenability for processing of oolitic iron ore concentrate for phosphorus removal

Beneficiation routes aimed at dephosphorisation of oolitic gravity magnetic concentrate and involving a combination of roasting, re-grinding, magnetic separation and water and acid leaching are investigated. Roasting was carried out at 900 °C for 1 h without or with lime or sodium hydroxide as roasting additives. When additives were used, cement phases of Si–Al–Na–Ca–O type were detected as well as the mineral giuseppettite. During the thermal process sodium silicate is liquefied and the newly formed phases coat the oolites and penetrate inside the cracks. Energy Dispersive Spectroscopy analysis has indicated that the zone surrounding the oolites consists of Na, Al and Si phases with part of phosphorus being captured there. As a result of the alkaline roasting, goethite is partly transformed to magnetite and this reduction is reinforced with an increase in sodium hydroxide dosage. Investigation of redistribution of phosphorous shows that it could be only partly separated if leaching is not accompanied by re-grinding and physical separation. The recommended dosage of the reductive agent for the final flowsheet is 8 mass% ratio to concentrate. Grinding to a mean size of 0.040 mm, with water and acid leaching and double magnetic separation creates conditions to obtain a high-quality iron concentrate with 65.97% Fe and recovery of 92.43%, with simultaneous decrease in the phosphorus content from 0.71% to 0.05%.     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).