Towards understanding nickel converter matte solidification

Nickel converter mattes are complex metallurgical solutions of Ni, Cu, S, Fe and O along with low concentrations of many other elements including Co, Pb and PGEs. Studies on how such complex mixed solutions evolve upon cooling may contribute towards an improved understanding of matte solidification. Liquidus and primary phase equilibria were calculated for Cu–Ni–S ternaries including fixed iron and cobalt concentrations. True liquid matte starting compositions and calculated assays were subsequently superimposed on relevant Cu–Ni–S_FeCo ternary systems. Multiphase cooling equilibria were also calculated for variable Cu–Ni–S–Fe–Co–O matte systems. In addition, actual industrial mattes were analysed using automated mineralogy, electron probe microanalysis and field emission scanning electron microscopy.The effect of the end composition of the ternary systems at fixed iron and cobalt concentrations on the liquidus temperature range has been quantified. The liquidus temperature range is lowered when the fixed iron and cobalt concentration decreases. The solidification pathway of oxygen-free liquid matte has been estimated. Moreover, it has been shown that variations in the starting composition of oxygen-free matte alter the path of solidification towards the eutectic. The examination of multiphase cooling equilibria for variable Cu–Ni–S–Fe–Co–O liquid phase systems provided a quantitative understanding of solidification processes to within ±2.5 °C. The analysed nickel and copper-sulphide phase structures have shown to exhibit chemical non-equilibrium within high and low iron matte. It can be concluded that the present study has provided a coherent insight into nickel converter matte solidification processes.     

Phase stabilities and thermodynamic assessment of the system Cu–Pb–S

An equilibration quenching technique was used to evaluate experimentally the phase relations in the Cu–Pb–S system between 763 and 1273 K. At 763–788 K, the formation of a ternary phase was confirmed. In addition, the thermodynamic properties of the Cu–Pb–S ternary system between 763 and 1573 K were critically assessed and optimized by the CALPHAD (CALculation of PHAse Diagrams) method, using the experimental results from this work and critically evaluated literature data. The obtained thermodynamic dataset allows the thermodynamic properties and phase equilibria in the Cu–Pb–S system to be calculated with good agreement to the experimental results. The effects of different process variables on the lead and copper distributions between the metal and matte can be calculated to improve the process control and economics.     

Geochemical Behavior of Mine Tailings and Waste Rock at the Abandoned Cu–Mo–W Azegour Mine (Occidental High Atlas, Morocco)

The abandoned Azegour mine is located in the High Atlas Mountains of Marrakesh (Morocco), and was mined for Cu, Mo, and W. About 850,000 t of waste rocks and tailings were deposited on the surface and have been exposed to weathering for 40 years. The remaining acid-producing potential (AP), acid-neutralizing potential, and geochemical behavior of the Azegour Cu-and Mo-rich tailings were investigated. The tailings were found to contain 9.6–19 wt% sulfur, mostly as sulfate (gypsum, anhydrite, and jarosite) while the waste rocks contain less (1.25–6.58 wt%) sulfur. The waste rocks and tailings contain 0.21–9.24 wt% Mo and 0.003–2.78 wt% Cu. The gangue is mostly composed of quartz, talc, chlorite, pyrophyllite, actonolite, clinoptilolite, and alusite. Lead, zinc, cobalt, arsenic, titanium, and nickel are also present. The calcium, which is mainly expressed as calcite, gypsum, scheelite, and powellite, is present at higher concentrations in the waste rocks (18–22 wt% Ca) than in the tailings (4.7–8.6 wt% Ca). Static ABA determinations showed that the Azegour mine wastes still have high AP, 38–205 kg CaCO₃/t in the waste rocks and 46–387.7 kg CaCO₃/t in the tailings. This was confirmed in weathering cell tests, where the Azegour tailings leachate had a pH range of 1.98–3.19 and high concentrations of SO₄ (468–45,400 mg/L), Ca (230–675 mg/L), Fe (3–55,900 mg/L), Mn (0.1–1,430 mg/L), and Cu (2.3–9,000 mg/L). The Mo concentrations were high (35 mg/L) during the two first weeks of kinetics tests; W concentrations were below the 0.005 mg/L detection limit.     

Thermodynamic analysis of the Fe–Ni–Co–Mg–Si–O–H–S–C–Cl system for selective sulphidation of a nickeliferous limonitic laterite ore

An increasing percentage of nickel is being extracted from the laterite ores for which there exists considerable potential for new process development. In the current research, a thermodynamic analysis of the Fe–Ni–Co–Mg–Si–O–H–S–C–Cl system has been performed in order to establish the possible sulphidation conditions for the upgrading of a limonitic laterite ore. The study was performed using the equilibrium module of HSC Chemistry® 6.1, and the data available in the literature were utilised to determine the activity coefficients of the various species employed in the calculations. The effects of variables such as temperature, amount of sulphur and carbon and chlorine additions on both the nickel grade and recovery in the monosulphide solid solution (mss) were determined. Alternative sulphidizing agents were also considered. Nickel recoveries of about 85% could be achieved at a grade of 10% nickel in the mss. Limiting factors in the sulphidation process were the formation of iron sulphide (FeS) and the relatively high stability of nickel oxide in the nickel ferrite (NiFe₂O₄), in comparison to the other phases, which resulted in excessive sulphur requirements of over 100 kg of S/tonne of ore. Methods to overcome these restrictions were discussed and the thermodynamic results were compared to the published experimental values.     

Removal of Metals and Acidity from Acid Mine Drainage Using Municipal Wastewater and Activated Sludge

Co-treatment of acid mine drainage (AMD) and municipal wastewater (MWW) using the activated sludge process is an innovative approach to AMD remediation that utilizes the alkalinity of MWW and the adsorptive properties of the wastewater particulates and activated sludge biomass to buffer acidity and remove metals. The capacity of these materials to treat AMD was investigated in batch mode metal removal tests using high-strength synthetic AMD (pH 2.8, Al 120–200 mg/L, Cu 18–30 mg/L, Fe 324–540 mg/L, Mn 18–30 mg/L, and Zn 36–60 mg/L). Using material from a range of MWW treatment plants, the performance of screened and settled MWW, activated sludges with mixed liquor suspended solids (MLSS) concentrations of 2.0 and 4.0 g/L, and return activated sludges with 6.0 and 7.4 g/L MLSS were compared. Similar trends were observed for the MWW and activated sludges, with removal efficiency generally decreasing in the order Al = Cu > Mn > Zn > Fe. Trends in Fe removal using settled MWW and activated sludges were highly variable, with removal <30 %. Using activated sludges, average removal efficiencies for Al, Cu, Mn, and Zn were 10–65 %, 20–60 %, 10–25 %, and 0–20 %, respectively. Sludge solids concentration was an important controlling factor in metal removal, with removal of Al, Cu, Mn, and Zn increasing significantly with solids concentration. Municipal wastewaters had greater neutralization capacities than activated sludges at high AMD loading ratios. Mixing AMD with screened MWW gave the highest removal efficiency for all metals, achieving average removal of 90–100 % for Al, Cu, and Fe, 65–100 % for Zn, and 60–75 % for Mn. These empirical findings are useful for developing process design parameters in co-treatment systems. Utilizing MWW and activated sludge to remediate AMD can potentially reduce materials and energy requirements and associated costs.     

The Potential Use of Phosphatic Limestone Wastes in the Passive Treatment of AMD: A Laboratory Study

In Morocco, there are many sedimentary phosphate mines that produce large quantities of phosphatic limestone wastes (PLW) that contain calcite (46 wt%) and dolomite (16 wt%). These mines are located near contaminated sites, such as the abandoned Kettara pyrrhotite mine. The surface drainage water at the Kettara mine site has a pH of 2.9–4.2 and elevated concentrations of SO₄ (from 47 to 5,000 mg/L) and Fe (from 1 to 1,200 mg/L). The efficiency of PLW was assessed in the laboratory as an alternative alkaline material for passive acidic mine drainage (AMD) treatment. A series of experiments were carried out using a synthetic AMD (pH 3) containing Fe (500 mg/L), SO₄ (3.4 g/L), Ca (220 mg/L), Al (160 mg/L), Mn (20 mg/L), Zn (15 mg/L), Cu (23 mg/L), and trace amounts of Co, Cr, and Ni. Experiments were done in both anoxic and oxic conditions, in batch and column tests, with hydraulic retention times of 24 and 15 h, respectively. The PLW efficiently increased the alkalinity and pH, inducing precipitation of most metals. The neutralizing capacity of PLW prepared at different particle sizes (0.8 mm–0.5 cm, 0.5–1, 1–2, and 2–3 cm) was found to be similar in batch tests. The initial AMD value increased from 3 to 5–6.5 during the batch tests and 6.5–8 in the columns. In batch tests under anoxic and oxic conditions, there was a significant decrease in concentrations of Fe (500–120 mg/L), Al (160–1.7 mg/L), and Cu (23–0.002 mg/L). In the column tests, Al and Cu decreased (177–2.5 and 26–0.002 mg/L, respectively), while Fe decreased less significantly (618–300 mg/L). The availability and low cost of the PLW make its use in passive AMD treatment potentially feasible.     

Interfacial phenomena in the sulfur–polytetrafluoroethylene system under hydrothermal conditions

Sulfur occlusion of pyrite particles during oxidative pressure leaching of mineral sulfide concentrates at temperatures below 160 °C has a deleterious effect on pyrite oxidation kinetics. Existing techniques in mitigating sulfur occlusion involve the use of surfactants or operation at temperatures above 160 °C which are a burden to the process economy. This study looks at the suitability of polytetrafluoroethylene (PTFE) to act as a solid sorbent for the liquid sulfur generated during the pressure leach process. The suitability of PTFE as a solid sorbent was determined by comparing interfacial phenomena in the corresponding sulfur–PTFE–leach and sulfur–mineral–leach systems. The work of adhesion in a sulfur–PTFE system was calculated using the interfacial tension and contact angle measured in the relevant sulfur–PTFE system, which is comparable to published values in the pyrite–sulfur system.     

Gold and arsenic recovery from calcinates of rebellious pyrite–arsenopyrite concentrates

The products of stepped calcination of pyrite–arsenopyrite concentrates contain ferric arsenate phases FeAsO₄, Fe₄As₂O₁₁, Fe₃(AsO₄)₂. The article reports the research data on leaching of calcines in alkaline medium including caustic soda and hydrogen peroxide used as an oxidizer. The experiments prove the efficiency of alkaline leaching with H₂O₂ for dissociation of gold associated with ferric arsenates (II) and (III) and for reduction of arsenic content from 1.2–1.4 to 0.006–0.15%. Preliminary leaching of calcinates enhances gold cyanidation yield from 91–92.9 to 97.3–97.9%.     

Effects of pH,temperature and solids loading on microbial community structure during batch culture on a polymetallic ore

The bioleaching of an organic-rich polymetallic ore was conducted under conditions intended to probe the boundaries of microbial activity using iron and sulphur oxidising microorganisms and heterotrophs enriched from self-heating pyritic coal. Solution chemistry parameters such as rapidly increased ORP and reduction in pH subsequent to inoculation point to the development of active microbial communities. The ease with which communities adapted to the organic-rich ore and the bioleaching systems indicated that the organic compounds were not present in leachates at toxic levels. Overall, extractions obtained in three series of inoculated tests were at 35 °C: 79–96% Zn, 48–82% Cu, 47–55% Ni and 79–86% Co; at 55 °C: 96–97% Zn, 72–80% Cu, 46–50% Ni and 82–83% Co. T-RFLP provided semi-quantitative estimates of species abundance. The greatest microbial complexity was observed with moderate pH and low solids loading. Microbial complexity was reduced significantly by low pH or increased solids loading. Nevertheless, efficient bioleaching was observed over a relatively wide range of operating conditions. Even under the more extreme conditions, the community profile was dominated by combinations of organisms not typically seen in most commercial operations.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.     

Iron redox-equilibria and sulphide capacity of PGM melter-type slags

New measurements have been made on the ferric to ferrous ratio as well as the sulphide capacity for platinum group metals (PGM) melter-type slags. In South Africa, these slags are produced from the smelting of low-grade copper–nickel sulphide ores, Nell [Nell, J., 2004. Melting of platinum group metal concentrates in South Africa. The South African institute of Mining and Metallurgy 104 (7), 423–428]. The typical mass compositions are 5–10% Al₂O₃, 2–15% CaO, 5–30% FeO_x, 15–25% MgO and 40–60% SiO₂ with a molar basicity defined as (CaO + MgO)/SiO₂ of 0.6–1. The industrial furnaces operate at temperatures ranging from 1450 to 1600 °C under fairly reducing conditions (typically a pO₂ close to 10⁻⁸ atm at 1500 °C). The gas–slag equilibrium was studied by subjecting a synthetic slag to controlled atmospheres in a vertical tube-furnace using Ar–CO–CO₂ (–SO₂) gas mixtures. The ratio of ferric to ferrous was determined at 1450 °C for oxygen activities, defined as pCO₂/pCO, ranging from 0.11 to 1.75 by analysing the quenched slags using the standard titration and XRF techniques. The measured Fe³⁺/Fe²⁺ ratio increased from 0.029 to 0.110 with the increasing oxygen activity. Slight non-ideal iron redox behaviour was observed, as has been reported for low alumina and low iron-containing slags. The present results are in good agreement with the trends found in the literature for similar multi-component slag systems (mostly iron bath smelting slags). Sulphide capacity was measured at partial pressures of oxygen and sulphur of approximately 10⁻⁹ and 10⁻³ atm respectively, with total-iron contents of 8.2 and 15.6 wt%, and temperature ranging from 1450 to 1525 °C. The present sulphide capacity data ranged from 10^−4.43 to 10^−3.71. The expected increase in sulphide capacity with increasing temperature was observed, and at a given temperature, the sulphide capacity increased with an increase in iron oxide content.     

Palladium(II) removal from chloride and chloride–nitrate solutions by chelating ion-exchangers containing N-donor atoms

The chelating ion-exchangers of functional iminodiacetate (Amberlite IRC-718), amidoxime (Duolite ES-346) and aminophosphonic (Duolite C-467) groups have been applied for Pd(II) removal from the model chloride (0.1–6.0 M HCl) and chloride–nitrate (0.1–0.9 M HCl and 0.9–0.1 M HNO₃ and 0.1–1.5 M HCl and 1.9–0.5 M HNO₃) solutions. The total ion-exchange capacities as well as recovery factors of Pd(II) were determined by the batch method. The influence of acid concentrations, phase contact time and macrocomponent addition (AlCl₃, CuCl₂, NiCl₂) was studied. The results show that the ion-exchangers of functional amidoxime and iminodiacetate groups can be widely recommended for Pd(II) ion removal from anodic slimes, and used up catalysts, as well as Pd(II) trace analysis due to their high selectivity.     

Acid Rock Drainage Remediation and Element Removal Using a Peat-Humic Agent with Subsequent Thermal Treatment of the Metal–Organic Residue

A novel, alternative method for acid rock drainage (ARD) remediation and metal recovery has been developed that uses a peat-humic agent (PHA) created by mechanical, chemical, and thermobaric treatment of peat from the Krugloe deposit (Novosibirsk region, Russia). The PHA effectively neutralised moderately acidic ARD and removed potential pollutants (e.g. Fe, Al, Zn, Cu, Pb, Cd, Ni, Co, and Hg), forming metal–organic residues. The organic matter can be removed completely from the metal–organic residues by heating them at 450–500 °C. After this treatment, the metal concentrate residues generally contained aggregates (20–350 μm in size), mainly composed of metal oxides and sulphates. Thermal decomposition of the organic matter in the PHA and metal–organic residues is an exothermic process with significant calorific value (9–15 kJ/g).     

Changes in an Australian laterite ore in the process of heat treatment

This paper examines an Australian garnieritic-type ore and changes in phase composition and morphology caused by heating in argon at 400–1000 °C using XRF, XRD, DTA/TG, SEM/EDS and BET analyses. The mineral phases detected by XRD in the original ore include chlorite, talc, hematite and quartz. Traces of iron silicate, Fe–Cr spinel and monoxide phase (predominantly manganese oxide) were observed by EDS. Nickel was detected in chlorite, talc, iron silicate and monoxide phase. Heat treatment at 400–500 °C did not change XRD patterns. At 600 °C, dehydroxylation of the brucitic phase of chlorite occurred. Chlorite was converted into olivine (forsterite) and enstatite at 600–800 °C. Upon heating to 900–1000 °C, talc was also converted into olivine and enstatite. Ni-bearing phases after heat treatment at 800–850 °C were forsterite, enstatite, talc, iron silicate and monoxide.     

Textural,mineralogical and chemical characteristics of copper reverb furnace smelter slag of the Okiep Copper District,South Africa

The Okiep Copper District in South Africa has produced more than 110 million tons at a grade of 1.71% Cu from several small mafic ore bodies. The ore was smelted on site and generated ∼5 mt of slag. During the life of mine attempts to recover copper from the slag by flotation had limited success. After mine closure the challenge of environmental rehabilitation and the possible disposal of the slag, triggered a reinvestigation into the viability of slag as a copper resource. Characterisation of the slag as a contribution to the potential copper recovery is the objective of this study.The slags are hard, vitreous with a matrix of Si–Fe–Al–Mg–Ca glass and laths of Mg–Fe–olivine, Fe–Mg–orthopyroxene and minor Cr-spinel. Copper grade varies between 0.11% and 0.42% with minor nickel, cobalt, molybdenum, zinc and tungsten. All economic elements are hosted by disseminated spheroidal prills which consist mainly of the copper sulphides bornite, chalcocite, covellite and chalcopyrite with exsolved sulphide phases of the minor base metals as well as rhenium and silver. Prills consisting of metallic copper and alloys are minor constituents. Prill diameter is highly variable with most in the 40–60 μm range and the historically poor copper recovery is attributed to the small prill size. Crushing of slag to −45 μm as opposed to the previous −75 μm should significantly increase sulphide liberation and recovery of copper and minor base metal sulphides by conventional flotation.Provided the operation is economically viable, redistribution of the processed slag to environmentally acceptable sites will resolve the present pollution and rehabilitation challenge related to the dumps in the Okiep Copper District. The operation will also have a positive socio-economic impact on this poverty-stricken part of South Africa.     

Interaction mechanism of miscible DDA–Kerosene and fine quartz and its effect on the reverse flotation of magnetic separation concentrate

In this study, a modification of oil assisted flotation processes of quartz particles has been proposed, which is based on introduction of miscible Dodecylamine (DDA)-Kerosene as collector with DDA cationic surfactant coated on kerosene to the hydrophilic quartz particles in the pulp. The property of miscible DDA–Kerosene emulsion was investigated. Due to the adsorption of DDA at kerosene/water interface, a smaller and uniform kerosene emulsion formed. Addition of cationic surfactant to the kerosene emulsion changed the zeta potential value from negative to positive, which resulted in enhancing the adhesion of the oil droplets to negatively charged quartz. The results showed that agglomeration and flotation process can be realized simultaneously with DDA–Kerosene. The agglomeration of fine quartz minerals in the presence of miscible DDA–Kerosene led to the formation of very large compact agglomerates resulting in increasing hydrophobicity of the particles and inducing a higher probability of collision and adhesion to air bubble. Experimental data indicated that miscible DDA–Kerosene had better selectivity and stronger collectability to quartz than DDA–HCl, which can be used as an efficient collector in the reverse flotation of magnetic separation concentrate of TISCO. At the same DDA dosage (60 g/t), separation efficiency got to 18.53% when using DDA–HCl as collector; while a better result was obtained with DDA–Kerosene, the efficiency of separation reached 59.07% which was identical with 120 g/t DDA–HCl.     

Mechanical,hydrological and microstructural assessment of the durability of cemented paste backfill containing alkali-activated slag

The durability of full (FT) and deslimed sulphide-rich tailings (DT) cemented paste backfill (CPB) of alkali-activated slag (AAS; LSS–S: activated slag with aqueous sodium silicate and SH–S: activated slag with sodium hydroxide) was investigated based on the mechanical (strength), hydraulic and microstructural properties. DT and AAS improved the strength gain rate by decreasing the total- and macro porosity. DT and LSS–S configuration had a profound effect on strength development and produced 4.3–9.9-fold strengths than ordinary Portland cement (OPC) samples. DT also favoured the degree of cementation with dense packing matrix, decreased the permeability and thus, limited the adverse effects of acid and sulphate. DT hindered the decalcification of C–S–H leading to higher Ca/Si ratios for OPC while AAS samples exhibited approximately similar ratios over the curing period. SH–S led to drying shrinkage due to the greater amount of threshold pores. LSS–S/DT combination provided better durability performance in case of aggressive environments by improving the microstructure of CPB.     

Zinc removal in anaerobic sulphate-reducing liquid substrate process

Zinc and sulphate removal from synthetic wastewater was investigated by using four laboratory parallel upflow-mode reactors (referred as R1 to R4; R1 contained carriers to retain biomass, whereas R2–R4 were operated as suspended reactors). All reactors were inoculated with anaerobically digested cow manure. R1 and R2 were first fed with glucose- and sulphate-containing feed for 48 days after which all four reactors were fed with wastewater containing 50 mg l⁻¹ of zinc in R1–R3 and 200 mg l⁻¹ in R4 and operated for 96 days. In all reactors, hydraulic retention time, organic loading rate, and sulphate load were 5–6 d, 0.2–0.4 kg COD m⁻³ d⁻¹ and 3.3–3.8 g SO₄ l⁻¹ d⁻¹, respectively, whereas the zinc load in R1–R3 was 0.074–0.077 and in R4 0.282 g Zn l⁻¹ d⁻¹. During the runs, 30–40% of sulphate and over 98% of zinc was removed, and up to 150–200 mg H₂S was produced in all reactors. Effluent pH dropped in all reactors (feed pH 6.5) to 3–5 by the end of the experiment. No significant effects on zinc removal were observed, despite differences in operating conditions and feed. It was only in the latter part of the runs (i.e. between experiment days 120–142) that zinc removal began to fluctuate, showing a negligible decrease in R3 and R4, whereas in R1 and R2 zinc was removed below the limit of detection (<0.01 mg Zn l⁻¹). Qualitative X-ray diffraction analysis of the reactor sludge at the end of the runs indicated that the compounds precipitated were most probably ZnS (Code 05-0566 Sphalerite), suggesting metal removal through sulphide precipitation; this was supported by the fact that sulphate was reduced and zinc removed simultaneously.     

Stirred milling kinetics of siliceous goethitic nickel laterite for selective comminution

The objective of this study is to determine how grinding conditions affect the breakage rate with respect to the sample mass, major elements, and minerals present in siliceous goethitic (SG) nickel laterite. This information is helpful in determining the optimal grinding conditions for selective comminution and nickel upgrade. The kinetics of batch wet grinding of nickel laterites with feed sizes of 2.38–1.68, 1.68–1.18, 1.18–0.85, 0.85–0.6, 0.6–0.42, 0.42–0.3, 0.3–0.21, and 0.21–0.15 mm were determined using a Netzsch LME4 stirred mill under the following conditions: 1000 rpm, 50% charge volume, 150.0 g of solid. The grinding behaviour of the majority of the feed samples was non-first-order due to the fast breakage rate of soft minerals and the low breakage rate of hard minerals in the feed. Therefore, an enrichment of the soft mineral was obtained in the underscreen product by selective grinding. The effect of selective grinding on Ni upgrade was evaluated by looking at grinding time, feed size, and product size. Optimum grinding time with respect to Ni upgrade was 0.25 min for SG nickel laterite samples. Generally, grinding larger particles and/or collecting finer product size yielded better Ni upgrade results. The effect of selective grinding was evaluated by the changes of the major soft and hard minerals for the selected samples. Selective grinding was also examined with respect to the major element weight ratio (e.g. Si/Ni for SG nickel laterite). With respect to Ni upgrade, the best result was achieved from the 1.18–0.85 mm feed on the −400 mesh product after grinding for 0.25 min. The Ni grade increased from 0.73% to 1.30% (upgrade 76.8%), with 14.4% Ni recovery; the Mg grade increased from 1.30% to 3.96% (upgrade 205.6%); the Si grade decreased from 28.7% to 16.2%.     

Mössbauer quantification of pyrrhotite in relation to self-heating

Pyrrhotite (Po) is widely considered the most significant mineral in the self-heating of sulphide ores and concentrates. It is therefore desirable to determine the amount of Po in a sample. This is commonly accomplished through X-ray diffraction (XRD) or quantitative evaluation of materials by scanning electron microscopy (QEM–SEM), neither of which are suited to oxidizing samples such as Po. The sample preparation for these methods often requires heating and drying, which may alter the composition. This paper introduces another quantitative method based on Mössbauer analysis, whose sample preparation is limited, and not as likely to alter the sample. The technique is tested using binary and ternary mixtures containing pyrrhotite with pentlandite (Pn), pyrite (Py) and sphalerite (Sp). A detection limit of 2 wt.% Po is easily obtained in the Po–Pn binary mixtures. The Po is also successfully measured in Pn–Po–Py and Sp–Po–Py mixtures; however, the presence of sphalerite increases the time required for an accurate measurement. The time required to measure Po in complex samples may be optimized through future work.