Atmospheric acid leaching of siliceous goethitic Ni laterite ore: Effect of solid loading and temperature

In this study, atmospheric acid leaching behaviour of siliceous goethitic nickel (Ni) laterite ore is investigated. Specifically, the effect of −200 μm feed solid loading (30 vs. 45 wt.%) and temperature (70 vs. 90 °C) on leach kinetics, acid consumption capacity and Ni and cobalt (Co) extraction was studied under isothermal, batch (4 h) leaching conditions at pH 1. Incongruent leaching was observed for constituent elements reflecting slow but steady release of value (Ni and Co) and some of gangue metals such as Fe, Mg and Al accompanied by faster and sharp release of Na and Si. Higher temperature and lower pulp solid loading, both led to a 40–50% increase in overall Ni and/or Co extraction and higher acid consumption. At 70 °C and 45 wt.% solid loading, Ni/Co extraction after 4 h was the lowest (∼14/16%) whilst the highest extraction (∼67/56%) was observed at 90 °C and 30 wt.% solid loading. Temperature appeared to have dramatic influence on Ni/Co and other impurity metals’ extractions revealing the chemical reaction controlled nature of the leaching. Higher solid loading and longer leaching time also both slowed down the leach kinetics. A two-stage chemical reactions-controlled leaching mechanism involving a faster initial leaching kinetics followed by a slower leaching at lower rate constants and higher activation energies was established for release of Ni, Co, Fe and Mg. The mechanism reflects the fast leaching of reactive host mineral phases (e.g., clays and Mg–silicates) during first 30 min followed by slow leaching of more refractory mineral phases (e.g., goethite and quartz) during the rest of leaching period. The findings provide a greater understanding for enhanced atmospheric acid leaching process of siliceous goethitic laterite ores.     

Kinetics of high-sulphur and high-arsenic refractory gold concentrate oxidation by dilute nitric acid under mild conditions

On one hand, high-sulphur and high-arsenic refractory gold concentrate (HGC) leads to regional ecological damage. On the other hand, it contains a lot of valuable elements. So utilization of it can bring social and environmental benefits. In this paper, the kinetics of HGC oxidation by dilute nitric acid under mild conditions was investigated. The effects of particle size (50–335 μm), reaction temperature (25–85 °C), initial acid concentration (10–30 wt.%) and stirring speed (400–800 rpm) on the iron extraction rate (C_r) were determined. It is obvious that C_r increases with the rise of initial nitric acid concentration, reaction time and stirring speed, but decreases with the increase of particle size. Oxidation kinetics indicates that the rate of reaction is diffusion controlled. The activation energies were determined to be 10.70 kJ/mol in the 10% HNO₃ and 12.25 kJ/mol in the 25% HNO₃.     

Paste backfill of high-sulphide mill tailings using alkali-activated blast furnace slag: Effect of activator nature,concentration and slag properties

The effect of activator type, concentration and slag composition on the strength and stability properties of paste backfill (CPB) of high-sulphide tailings using alkali-activated slag (AAS) as binder (7 wt.%) were investigated in this study. Acidic and neutral (AS–NS) slags were activated with liquid sodium silicate (LSS) and sodium hydroxide (SH) at 6–10 wt.% concentrations. Ordinary Portland cement (OPC) results were used for comparison. The strength development was found to remarkably improve with increasing the concentration from 6 to 8 wt.%. Further increase in concentration did not enhance the strength. SH was determined to produce higher early-age strength whilst LSS produced higher long-term strengths as an indication of slag selectivity for activators. More extensive gypsum formation was observed at lower concentrations in SEM/EDS studies. An increase in Na₂O concentration raised the activator consumption. High concentrations also led to poorly crystallized C–S–H gel, loose structure and drying shrinkage cracks especially in NS–SH samples. A reduction in total porosity up to 20% was obtained in AAS samples compared to OPC. Amorphous structure, chemical modulus ratio and/or basicity index (BI) values were seen to control the pozzolanic reactivity, and therefore, the alkali-activation and hardening process.     

Recycling of waste pickle acid by precipitation of metal fluoride hydrates

Stainless steel is pickled in mixed acid solutions (1–3 M HNO₃ and 0.5–4 M HF). The spent solution is usually neutralized with lime, and in Sweden about 18,000 tons/yr of metal hydroxide sludge is disposed as landfill waste. We are developing a cost-saving and environmentally friendly process, involving crystallization of β-FeF₃ · 3H₂O, where the metal content is recovered and the acid is recycled.Iron has been successfully separated from spent pickle bath solutions by precipitation of β-FeF₃ · 3H₂O in a continuous crystallizer (10 L scale) where the solution is concentrated by nanofiltration. The crystal growth rate of β-FeF₃ · 3H₂O has been determined in industrial pickle bath solutions at 50 °C and the results have been compared to previous measurements in pure HF/HNO₃ solutions prepared in the laboratory. The growth rate of β-FeF₃ · 3H₂O crystals at 50 °C is in the order of 10⁻¹¹ m/s in both industrial and pure acid mixtures.     

Pulse electrodeposition as a new approach in electrowinning of high purity cobalt from WC–Co scraps. Part I: The effect of frequency and duty cycle

The high price of cobalt is the reason for the necessity of its recovery from recycled cutting tools, catalysts and so forth. In this research, the effect of frequency and duty cycle in pulse electrowinning of cobalt from WC–Co scraps has been studied. The scraps was first dissolved in a solution containing 1 N hydrochloric acid and 0.1 M citric acid in anodic potential in the range of 0.05–0.3 V_Ag/AgCl so that a solution with 22 g L⁻¹ cobalt was acquired. Electrodeposition was performed on a stainless steel cathode in frequencies of 30, 50, 70, 100 and 135 Hz; duty cycles of 10%, 30% and 50%; and peak current density of 250 A m⁻²; and electrolyte temperature of 60 °C. The surface of the deposits was observed with SEM and its purity was measured using the ICP method. The studies showed that increased frequency reduces the crystal size and the purity of the deposit increases up to 99.1 wt% under the same conditions. Increasing the duty cycle raises the cobalt crystal size and the highest purity of deposit is obtained at a duty cycle of 30%. The optimum conditions in terms of economics and purity involve electrodeposition at a frequency of 70 Hz and duty cycle of 30%, where deposits with a purity of 94.13 wt% are obtainable.     

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.     

Bioleaching of a low-grade copper ore,linking leach chemistry and microbiology

Three largely-independent studies were undertaken on the same heap leach system during the period of transition from processing oxidised ores to sulfide ores: monitoring of heap solutions for microorganisms, analysis of samples from a spent heap, and column tests. Microbial cell numbers and diversity were monitored in process water samples from the transition heap over a four-year period. Cell numbers remained low throughout, 1–30 × 10⁴ cells mL⁻¹, possibly reflecting growth inhibition by the high element concentrations in process water. High iron, magnesium and aluminium concentrations in spent heap pregnant leach solution (PLS) are attributed to siderite and clinochlore dissolution and would be expected to impact on microbial growth. Planktonic cell numbers in a column leachate declined rapidly by two orders of magnitude when concentrations of ferric ion and sulfate exceeded 30 and 75 g L⁻¹, respectively. Nevertheless, a variety of bacterial strains closely related to Acidithiobacillus (At.) ferrooxidans, At. caldus, Leptospirillum (L.) ferriphilum, Acidimicrobium (Am.) ferrooxidans, Acidiphilium (Ap.) cryptum, an Alicyclobacillus-related strain and Sulfobacillus (S.) thermosulfidooxidans, and the archaeon Ferroplasma (F.) acidiphilum were isolated, mainly from the more acidic intermediate leach solutions (ILS). Overall, the results obtained from the use of culture-dependent and culture-independent methods of community analysis were complementary and consistent. The majority of identified genera and species were present in both the process water samples from the operating heap and the solutions and ore samples from the spent heap. In the spent heap, distinct populations dominated different sample types. Leptospirillum- and Acidithiobacillus-like strains dominated PLS samples and Leptospirillum also dominated seven of eight spent ore samples and all of the heap sediment samples, making it the primary iron(II) oxidising species.     

Selective sulphidation and flotation of nickel from a nickeliferous laterite ore

The sulphidation of a nickeliferous lateritic ore was studied at temperatures between 450 and 1100 °C and for sulphur additions of 25–1000 kg of sulphur per tonne of ore. The experiments demonstrated that the nickel could be selectively sulphidized to form a nickel–iron sulphide. It was found that both the grade and the sulphidation degree largely depended upon the temperature and the sulphur additions, with temperatures above 550 °C exhibiting the highest nickel sulphidation degrees and grades. A DTA/TGA with mass spectrometer was used to further elucidate the nature of the phase transformations that occurred upon heating of the ore in the presence of sulphur.It was found that at low temperatures, the Fe–Ni–S phase was submicron in nature and heating to temperatures between 1050 and 1100 °C allowed for the growth of the particles, due to the increased sulphide mobility associated with the formation of a liquid sulphide matte phase, containing dissolved oxygen. Flotation studies conducted on 60 g samples showed that the sulphides responded to flotation with maximum grades of up to 6–7 wt.% nickel being achieved. Recoveries were approximately 50% on a sulphide basis and it was determined that the low nickel grades were due to the entrainment of magnetite fines.     

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.     

A novel low pH sulfidogenic bioreactor using activated sludge as carbon source to treat acid mine drainage (AMD) and recovery metal sulfides: Pilot scale study

In this work a pilot scale sulfidogenic bioreactor was used to treat acid mine drainage (AMD) from Zijinshang copper mine. In this process, S²⁻ produced in the Up-flow Anaerobic Sludge Bed (UASB) reactor were recycled in the two precipitation tanks for copper and iron precipitation, activated sludge from local waste water treatment plant was used as the carbon source. The reactor were steady operated in acid condition (with no pH control) for 4 month, AMD with a copper concentration of 100–120 mg/L, iron concentration of 170–200 mg/L, sulfate concentration of 2000–2500 mg/L and pH of 2.34–2.56, were feeding into the reactor under a feed rate of 1 m³/days and HRT of 3 days, copper and iron removal were 60.95%, 97.83% respectively. The precipitant in the precipitation tank containing 15.7% Cu and 22.66% Fe, thus indicating a recovery possibility of copper by pyrometallurgy process. From these results we can conclude that an SRB process would be a viable method of treating Zijinshan AMD.     

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.     

Suitability of bentonite-paste tailings mixtures as engineering barrier material for mine waste containment facilities

This study investigates the feasibility of using bentonite-paste tailings (BPT) as a barrier (liner, cover) material for mine waste containment facilities. Improvements of the hydraulic properties are realized by using compaction to densify the paste tailings in the first stage and mixing an additive-like natural bentonite to paste tailings to further reduce the voids that control hydraulic conductivity in the second stage. A significant decrease in hydraulic conductivity is observed with these transformations. Values that are as low as 1 × 10^?9 and 4 × 10^?9 cm/s are obtained in 8% and 4% bentonite and BPT, respectively. Based on these satisfactory values, additional investigations are conducted to evaluate the freeze–thaw and wet–drying performance of BPT up to 5 and 6 cycles, respectively. The results show that negligible to acceptable changes in hydraulic conductivity occur. None of the changes reach one order of magnitude. As a final step, a cost analysis is undertaken to evaluate the economical benefits that can be obtained from such new material. When compared to conventional compacted clay–bentonite barrier or sand–bentonite with 12% bentonite concentration, it is found that the 4% BPT is less expensive by 66%. The results place this recycled BPT material as a promising candidate for barrier design while reducing the amount of waste to be managed and the cost of surface tailings management.     

Geopolymerization of coal fly ash in the presence of electric arc furnace dust

The results of synthesis of fly ash based geopolymers and the experimental investigations of the geopolymerization of fly ash in the presence of eclectic arc furnace dust were presented. The electric arc furnace dust (EAFD) is a hazardous solid waste formed during remelting processes of iron and steel scrap in an electric arc furnace. In this work, the possibility of immobilization of Zn (as a most abundant toxic metal in EAFD) from EAFD and the influence of EAFD addition and alkaline dosage on the mechanical and microstructural properties of fly ash based geopolymers was investigated. Our result show that the best mechanical properties of FA and FE geopolymers were obtained using 10 M NaOH and 10 wt.% EAFD. The X-ray phase analysis confirmed the presence of aluminosilicate and oxide phases in investigated samples of FA, EAFD, as well as FA and FE geopolymers. SEM–EDS results have revealed the presence of Zn in an amorphous aluminosilicate phase. This material is considered as mesoporous and the change of porosity is greatly dependent on the alkaline dosage. The efficiency of immobilization of Zn from EAFD was evaluated using TCLP, EN 12457-2 and EPA Method 1313 leaching tests. Results of TCLP and EN 12457-2 tests have shown that the increase of alkaline dosage leads to the more effective immobilization of Zn and other toxic metals present in EAFD. On the other hand, the results of EPA method 1313-pH dependent leaching test have indicated that FE geopolymers are relatively stable in neutral and alkali environment while in a pH range 2–5.5, solubility of Zn and Pb is a concern for this type of materials.     

Sample requirements for HPGR testing procedure

Green field projects demand relatively large amounts of sample from drill-cores. Besides chemical analysis, samples are required for mineralogy and liberation characterization, physical characterization, concentration tests and a number of tests for crushing and grinding parameters. If the project’s process route includes a possible HPGR grinding stage, lab-scale tests for scale-up and variability analysis are required. HPGR grinding characterization can be carried out in a small diameter roll HPGR, such as the LABWAL. Some commercial labs recommend 20–30 kg samples for steady-state tests, but this is a rather large sample from the point of view of a green field project campaign. The question that is being assessed here is how much sample is really required? Surely, the more material that is available the more reliable will be the test results. However, when the sample mass size is reduced, what is the impact on the data that is produced? In this work, six phlogopitite samples weighing 20 kg were tested in the LABWAL HPGR using six initial hydraulic pressures, 10, 20, 30, 40, 50 and 60 BAR. The specific capacity and specific power factors were determined, as well as the critical angle of nip, and the critical gap. Size distributions were measured and size-mass balance parameters were determined for the Austin model under the range of grinding pressures that were produced. With the data, a hypothetical industrial HPGR for the phlogopitite was designed for a standard capacity of 100 t/h and operating at 2 N/mm² specific grinding force. The work was then repeated using 10 kg samples and 5 kg samples. Results show that, under the conditions that were chosen, samples weighing 5 kg are sufficient for characterization in the LABWAL HPGR.     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

Valorization of steel slag by a combined carbonation and granulation treatment

This work reports the results of a combined accelerated carbonation and wet granulation treatment applied to Basic Oxygen Furnace (BOF) steel slag with the aim of producing secondary aggregates for civil engineering applications and of storing CO₂ in a solid and thermodynamically stable form. The tests were carried out in a laboratory scale granulation device equipped with a lid and CO₂ feeding system. In each test, humidified slag (liquid/solid ratio of 0.12 l/kg) was treated for reaction times varying between 30 and 120 min under either atmospheric air or 100% CO₂. Under both conditions, the particle size of the treatment product was observed to increase progressively with reaction time; specifically, the d₅₀ values obtained for the products of the combined granulation and carbonation treatment increased from 0.4 mm to 4 mm after 30 min and to 10 mm after 120 min. Significant CO₂ uptake values (between 120 and 144 g CO₂/kg) were measured even after short reaction times for granules with diameters below 10 mm and for the coarser particle size fractions after reaction times of 90 min. The density, mineralogical composition and leaching behavior of the obtained granules were also investigated, showing that the combined granulation–carbonation process may be a promising option for BOF slag valorization, particularly in terms of decreasing the Ca hydroxide content of the slag. Another interesting finding was that the leaching behavior of the product of the combined treatment appeared to be significantly modified with respect to that of the untreated slag only for coarse uncrushed granules, an indication that the carbonation reaction occurs mainly on the outer layer of the formed granules.     

Acid leaching and rheological behaviour of a siliceous goethitic nickel laterite ore: Influence of particle size and temperature

This study investigates the isothermal, batch, H₂SO₄ acid leaching behaviour of siliceous goethitic (SG) nickel (Ni) laterite ore and its links to pulp rheology. Specifically, the effect of feed ore particle size (−0.2 vs −2.0 mm), leaching temperature (70 vs 95 °C) and pulp rheology on Ni and pay metal, cobalt (Co) extraction kinetics and yield was studied for 4 h on 40 wt.% solid dispersions at pH 1. The leaching behaviour was distinctly incongruent, reflecting the disproportionate proliferation of major gangue mineral’s constituent elements (e.g., Fe, Al, Mg, Na, Si) alongside Ni and Co in the pregnant leach solution. At 70 °C, Ni/Co extraction rates were notably lower (<20%) in contrast with 95 °C where a significant increase in Ni/Co extraction to 78/77% and 74/77%, respectively, for the −0.2 and −2.0 mm feeds occurred. The slurries displayed a non-Newtonian, shear thinning Bingham plastic rheological behaviour of which the viscosity and shear yield stress increased markedly in the course of 4 h leaching. The pulp viscosity and shear yield stress were greater at lower temperature than at higher temperature and they were also greater in slurries with finer than coarser feed particles. The dynamic pulp rheology, however, had no marked effect on the overall Ni/Co extraction rates. Whilst the feed ore particle size had no remarkable impact on overall Ni/Co extraction, it led to noticeably higher acid consumption and enhanced slurry rheology in the finer sized ore. The mechanism of leaching the SG ore followed a two-stage, first order chemical reaction-controlled shrinking core model, the kinetics of which gave higher rate constants and lower activation energies for the release of Ni, Co, Fe and Mg in the first stage. A faster leaching process involving more reactive minerals during the first 30 min is envisaged to be followed by leaching of the more refractory minerals.     

Review of rate constants for thiosulphate leaching of gold from ores,concentrates and flat surfaces: Effect of host minerals and pH

A review of literature data for different types of sulphide concentrates and gold ores has been carried out to examine the impact of host minerals and pH upon gold leaching. Analysis of initial rate data over the first 30–60 min of gold leaching from sulphide concentrates or silicate ores over a range of ammonia, thiosulphate, and copper(II) concentrations, pH (9–10.5) and temperatures up to 70 °C shows the applicability of a shrinking sphere kinetic model with an apparent rate constant of the order k_ss = 10⁻⁶–10⁻³ s⁻¹. The dependence of apparent rate constant on pH and initial concentrations of copper(II) and thiosulphate is used to determine a rate constant k_Au(ρr)⁻¹ of the order 1.0 × 10⁻⁴–7.4 × 10⁻⁴ s⁻¹ for the leaching of gold over the temperature range 25–50 °C (ρ = molar density of gold, r = particle radius). These values are in reasonable agreement with rate constants based on electrochemical and chemical dissolution of flat gold surfaces: k_Au = 1.7 × 10⁻⁴–4.2 × 10⁻⁴ mol m⁻² s⁻¹ over the temperature range 25–30 °C. The discrepancies reflect differences in surface roughness, particle size and the effect of host minerals.     

α → γ → β-phase transformation of spodumene with hybrid microwave and conventional furnaces

Spodumene is the most important lithium containing hard rock mineral. In order to extract lithium from spodumene concentrate by leaching, the crystal structure of spodumene must be converted from the natural monoclinic α-form to the tetragonal β-form. The technical possibilities to generate the heat to the conversion process of spodumene concentrate via microwaves were studied. The heat treatment experiments were carried out with a domestic microwave furnace (700 W) with silicon carbide susceptor and with the conventional resistance heated furnace as a reference. In the microwave furnace the phase transformation of spodumene began after 110 s of heating and samples were converted almost completely to β-spodumene after 170 s. Partial melting of gangue minerals was observed in samples after 170 s of heating. Heating in microwave furnace for 170 s corresponded with the heating of approximately 480–600 s at temperature of 1100 °C in the conventional furnace. In addition to α- and β-forms an intermediate phase, hexagonal γ-spodumene, was identified from samples heated with both furnaces. The conversion of spodumene samples was verified with X-ray diffraction (XRD) and with field emission scanning electron microscope (FESEM).     

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.