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%.     

Effect of grinding parameters on the rheology of pyrite–heptane slurry in a laboratory stirred media mill

The influences of wet ultra-fine grinding parameters on the rheological behavior of pyrite–heptane slurry in a laboratory stirred media mill were investigated with solid concentration, dispersant dosage, grinding time and carbon numbers of organic acid as dispersant. The results reveal that when the solids concentration is increased from 64 wt% to 79 wt%, the rheological behavior of slurry with 1 wt% of stearic acid transforms from Bingham characteristic to the pseudoplastic one with a yield stress. The Casson model fits well for the experimental data. And the apparent viscosity of the pyrite–heptane slurry increases exponentially with increase solid concentration at a given shear rate. The increase of viscosity is propitious to reduce the particle size of pyrite. When the solid concentration is 64 wt%, stearic acid is superior to octadecanol for the reduction of the slurry viscosity, and the slurry with 1 wt% of stearic acid possesses the best flowability. The extrapolated Bingham yield stress with dispersant almost stays constant when the dosage is over 2 wt%. Besides, the rheological behavior and particle size are also related to the grinding time and carbon numbers of organic acids as dispersant.     

Value adding limestone to filler grade through an ultra-fine grinding process in jet mill for use in plastic industries

In this study, ultra-fine grinding of limestone was carried out in jet mill using four levels of classifier rotational speed and grinding pressure. The holdup amount was determined during the grinding process, while the feed rate was kept constant at 8 kg/h. The ground product was characterized for its particle size and shape. In addition, the mechanochemical effect on the ground product was characterized through XRD. The particles size of the ground product ranged from 2.21 μm to 7.29 μm, demonstrating various particle shapes such as cubical, angular, and elongated. The degree of crystallinity of the ground product ranged from 54.5% to 93.7%. Afterwards, the ground product was incorporated as filler in polypropylene (PP), and its performance was characterized for mechanical properties. After conducting the test work, we find that the PP filled with ground limestone exhibited excellent thermal and mechanical properties. The composite flexural modulus, impact strength, tensile strength, and elongation at break were 2.1 GPa, 42 kJ/m², 22.75 MPa, and 21%, respectively, when loaded up to 20%. It likewise exhibited CTE value of 57.2 ppm/°C.     

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.     

Density based separations in the Reflux Classifier with an air–sand dense–medium and vibration

The Reflux Classifier is a device consisting of parallel inclined channels above a fluidized bed. Water-based versions of the system have been successfully employed in industry for gravity separation of ?2 + 0.25 mm coal and mineral matter. In this study an air-fluidized system was investigated using a single 2 m long inclined channel with 100 mm wide channels and 20 mm perpendicular spacing. Sand (?355 + 125 micron) was used as a dense–medium and vibration at two distinct levels was used to improve fluidization stability. Tracer particles of ?6.35 + 1.00 mm nominal diameter and 1300 to 2400 kg/m³ density were used to study the effects of the vibration energy and vibration direction on the separation efficiency. The device was able to separate coal particles from a maximum of 8 mm down to 1 mm on the basis of density; hence this method has potential for industrial application. Results were analysed using a simple 2-parameter dispersion–convection model. Dense particles had negative slip velocities, low-density particles had positive slip velocities and the slip velocities were proportional to particle diameter.     

Determination of the milling parameters of a platinum group minerals ore to optimize product size distribution for flotation purposes

Most concentrators desire to operate under optimal design configuration that guarantees high mineral recovery and low operational costs. The optimal design configurations are determined through studying the material to be milled in a laboratory mill under standard conditions. This is achieved through determining the selection and breakage function parameters and applying the mathematical simulation of the grinding process in order to optimize the size reduction process. The desired particle size is determined by the downstream processes, in our case, flotation. To this end, three mono-size classes feeds 850–600 μm, 600–425 μm and 425–300 μm of a platinum ore were ground using three different ball sizes (10, 20 and 30 mm) in a laboratory mill for the grinding times 0.5, 1, 2, 4, 8, 15 and 30 min. The data collected was used to determine breakage and some of the selection function parameters. The remaining parameters were back-calculated within the population balance model framework. The parameters were then used to obtain the product size distribution (PSD) that was later compared with the experimentally measured one. The milling kinetics for the desired size class for flotation was also simulated.There was a good match between the predicted and the experimentally measured PSD. The results of the milling done for further 60, 90, 120 and 240 min to validate the simulated milling kinetics from the determined parameters also showed good match between the simulated and the experimental one. This further confirms the validity of the determined parameters. From this, it becomes possible to determine the grinding conditions for optimal flotation.     

Formulating and optimising the compressive strength of controlled low-strength materials containing mine tailings by mixture design and response surface methods

Controlled low-strength materials (CLSM), like other cement-based backfill materials, are typically formulated by trial-and-error methods to yield the desired product characteristics. This paper presents the use of mixture design and response surface methods as tools to optimise formulations of CLSM to achieve desirable mechanical integrity with a minimum amount of statistically-sound experiments; while minimising the amount of cement and maximising the amount of by-products used. Statistical combinations of three-component mixtures were formulated to investigate the unconfined compressive strength (UCS) of CLSM comprising: Portland cement, fly ash and mine flotation tailings from a Ni–Cu ore. The data is analysed using the response surface method (using a mixture design of a constrained triangular surface) and ANOVA. Optimum formulations are simulated using a desirability function set at lower (1.0 MPa), target (2.0 MPa) and upper (3.0 MPa) UCS values after 28 days curing. All mix combinations had a constant spread diameter of 229 ± 10 mm, the standard workability for conventional CLSM. Results are compared to conventional CLSM incorporating silica sand in the place of the tailings. A significant quantity of tailings (up to 80 wt% solids) and low quantity of cement (up to 5 wt% solids) produced CLSM with UCS within the 2 MPa target value of excavatability. UCS of CLSM is a function of the individual component proportions, and the mixture design approach can be an important tool to help develop and optimise formulations of cement-based materials consisting of several components.     

Formation of jarosite during Fe2+ oxidation by Acidithiobacillus ferrooxidans

Jarosite precipitation is a very important phenomenon that is observed in many bacterial cultures. In many applications involving Acidithiobacillus ferrooxidans, like coal desulphurization and bioleaching, it is crucial to minimize jarosite formation in order to increase efficiency. The formation of jarosite during the oxidation of ferrous iron by free suspended cells of A. ferrooxidans was studied. The process was studied as a function of time, pH and temperature. The main parameter affecting the jarosite formation was pH. Several experiments yielded results showing oxidation rates as high as 0.181–0.194 g/L h, with low jarosite precipitation of 0.0125–0.0209 g at conditions of pH 1.6–1.7 with an operating temperature of 35 °C.     

Comparison of energy efficiency between ball mills and stirred mills in coarse grinding

Stirred mills are primarily used for fine and ultra-fine grinding. They dominate these grinding applications because greater stress intensity can be delivered in stirred mills and they can achieve better energy efficiency than ball mills in fine and ultra-fine grinding. Investigations were conducted on whether the greater performance of stirred mills over ball mills in fine grinding can be extended to coarse grinding applications. Four different laboratory ball mills and stirred mills have been tested to grind seven ore samples with feed sizes ranging from 3.35 mm to 150 μm. A case study on full scale operations of a 2.6 MW IsaMill replacing the existing 4 MW regrind ball mill at Kumtor Gold Mine in Kyrgyzstan is also included. This paper summarizes the major findings from these investigations.     

Liberated 0–10 μm particles from sulphide ores,their production and separation—Recent developments and future needs

The general types of industrial behaviour of liberated valuable sulphide minerals in the 0–10 μm size fraction are described and tests to explore on-site at a concentrator the general nature of any deficient behaviour are discussed. In addition, the general types of behaviour of the gangue minerals in this size fraction are also described in terms of their mechanisms for recovery.These behaviours are also discussed in terms of the type of grinding device and grinding environment in which they were produced and the target grinding product size. Major recent improvements in the industrial capabilities of grinding devices are discussed.Present difficulties in characterization of particles in the 0–10 μm range are described and desirable improvements for the future are outlined. Further, desirable improvements for industrial processing of valuable minerals in this size range are discussed along with some remedial actions.     

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.     

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.     

Determination of the concentrations of calcium and magnesium released from fluid inclusions of sphalerite and quartz

This study investigates the morphology of fluid inclusions trapped within a natural sphalerite and its closely intergrown quartz, and determines the total concentrations of calcium (Ca) and magnesium (Mg) released from the inclusions after grinding. The results indicated that numerous fluid inclusions with sizes from few microns to dozens of microns exist in the sphalerite and quartz. The inclusions contain abundant Ca and Mg. However, the concentrations of Ca and Mg in the inclusions of quartz are significantly lower than those of sphalerite. The inclusions were broken during the grinding process, releasing Ca and Mg to the solution; these released concentrations increased with the increase of grinding time. The maximum concentrations of Ca and Mg released from the fluid inclusions were 61.19 ppm and 5.23 ppm for sphalerite, and 3.28 ppm and 0.21 ppm for quartz, respectively. This study provides new understanding for the source of Ca and Mg in flotation pulp.     

An analysis of the energy split for grinding coal/calcite mixture in a ball-and-race mill

Interactions among components in the heterogeneous grinding would change energy consumed characteristics of components if compared with those in the single-component breakage. In this paper, energy split phenomenon for the coarse grinding of super clean anthracite coal (SCAC)/calcite mixture of 2.8–2 mm in the ball-and-race mill is investigated. Before the analysis of experimental results, accuracy of energy split function in terms of time-dependent breakage rate is first discussed. Energy consumed characteristics of grinding in the ball mill and ball-and-race mill are also compared. Breakage model of product t₁₀ (yield of progenies in −0.237 mm) vs specific energy is used to describe the energy-size reduction of the single-component and multi-component grinding. Interaction between components is reflected by the comparison of specific energy of components in mixture and single breakage to yield the same product t₁₀. Based on the energy balance, energy split factors of components in different time and mixed conditions are first determined. This parameter shows no change with time. Calcite increases the grinding efficiency of SCAC significantly, with the energy split factor for SCAC ranging from 0.68 to 0.73, which means less specific energy is consumed by SCAC to yield the same t₁₀ if compared with the single breakage. As the volumetric ratio of calcite increases in mixture, grinding energy efficiency decreases and energy split factor of calcite increases from 1.70 to 1.83. Soft material reduces the grinding energy efficiency of hard one in the multi-component breakage.     

Effect of mechano-chemical activation on bioleaching of Indian Ocean nodules by a fungus

The effect of mechano-chemical activation of Indian sea nodules, while recording the zeta potential, particle size distribution and surface area, on the bio-dissolution of metals by Aspergillus niger has been investigated. Activation is a term used to indicate what takes place when increasing grinding time does not result in significant change in particle size but rather results in the accumulation of energy that may lead to the development of lattice defects within the particles that can aid biological attack. It was observed that the mechano-chemical activation improved the bio-dissolution of metals such as copper, nickel and cobalt from the sea nodules at initial pH in the range 4.0–5.0. With 10 min milling of particles of ?75 μm size, 86% material was reduced to ?10 μm size with a change in zeta potential from ?18 to ?34 mV. Above 95% copper, nickel and cobalt each was leached out in 15 days time from the nodules activated for 10 min at 5% (w/v) PD and 35 °C temperature with initial pH of 4.5; the biorecovery being almost similar when the material was activated for 30 min. In the case of nodules without activation, ?89% metal bioleaching was achieved in 25 days time at an initial pH of 4.5 under this condition. The mechano-chemical activation of sea nodules has thus influenced the bio-dissolution process, while providing a wider pH range available for processing of nodules with the involvement of organic acids such as oxalic and citric generated from the fungus.     

Modeling breakage rates in mills with impact energy spectra and ultra fast load cell data

A new era in modeling particle size distribution in grinding mills started at the beginning of 2000s. A direct estimation of breakage parameters became possible via computation of collision energy by discrete element method (DEM) and material breakage data.The material breakage data can be obtained for primary modes of breakage. In this study, impact and abrasion are assumed to be the primary modes of particle breakage, which are readily studied in the laboratory. The impact breakage mode is studied in a drop-weight apparatus and in a specialized device known as the ultra fast load cell. The abrasion mode of breakage is studied in a laboratory scale ball mill. Next, the particle breakage versus energy data is converted into breakage rates via impact energy spectra of the grinding mill computed by a DEM code. The fundamental material breakage information is converted into energy based breakage distribution function.The verification of the modeling concepts is shown for a 90 cm laboratory scale ball mill. In the batch mill, approximately a 10 kg mass of limestone in the 30 mm size is ground with around 100 kg of 50 mm steel ball charge. The breakage rate and the breakage distribution functions constitute the parameters of the energy based batch population balance model. It is shown that accurate particle size distribution predictions are possible with this modeling approach for different grinding regimes.     

Ultrafine coal cleaning using spiral concentrators

The conventional circuit fine coal circuit typically incorporates spiral concentrators to treat nominal 1 × 0.15 mm material. Pilot and in-plant studies have been performed to determine the operating parameter values needed to achieve optimum separation performances when extending the lower limit of a conventional spiral concentrator to 44 μm. Based on experimental and empirical data, a feed solids concentration of about 12% by weight is required with a feed rate of around 60 l/min per start. Under these conditions, 60% of the ash-forming minerals and 48% of the sulfur was rejected from the 210 × 44 μm size fraction of a given coal source.     

The effects of chamber diameter and stirrer design on dry horizontal stirred mill performance

This paper focussed on investigating the effects of chamber diameter and stirrer design on cement grinding performance of a horizontal type dry stirred mill. Within the scope, pilot scale test works were undertaken with two different chamber diameters (20.4 cm and 26.4 cm) having the same length and three different stirrer designs (wing, cross and disc) having the same diameter (16 cm). The chamber diameter tests were performed at the same stirrer design, media size and media filling. The studies concluded that, the use of larger chamber improved the grinding efficiency since 31.8% and 35.8% less energy was consumed than the smaller mill at the RR_d50 of 1.41 and 1.66 respectively. This behaviour of the larger mill can be attributed to the increased gap distance between the chamber wall and stirrer edge. With regards to stirrer design, the statistical evaluations, grinding results and temperature measurements all indicated that the disc design of stirrer ground the particles more effectively at high energy levels (>40 kW h/t). The use of the disc design reduced the energy consumption by 21% (at RR_d50 of 3.5). This was attributed to dissipation of energy as heat since the temperature measured for the wing and cross types were higher than the disc type.     

Bioleaching of six nickel sulphide ores with differing mineralogies in stirred-tank reactors at 30 °C

A bioleaching study was conducted with six nickel sulphide ores from different geographical locations across Canada. Mineralogical and chemical examination revealed considerable variability between the samples, particularly in the silicate phases. The ores contain 0.3–1% nickel, primarily in pentlandite and secondarily in pyrrhotite. Copper is present primarily in chalcopyrite, and cobalt in pentlandite. The ores were subjected to the same crushing and grinding procedure, and bioleached under the same conditions for 3 weeks with a mixed culture of iron- and sulphur-oxidizing bacteria. Stirred-tank experiments with finely ground ore (−147 μm) at 30 °C were conducted to assess the effect of pH (2–5) and the impact of the bacteria. Nickel extraction from pentlandite and pyrrhotite during bioleaching at pH 2 and 3 was generally good (49–86% after 3 weeks), and cobalt extraction tracked nickel extraction over most conditions. All six ores showed a similar response to a change in pH; an increase in pH from 2 to 3 resulted in approximately the same nickel and cobalt extraction (within statistical error), and a statistically significant reduction in sulphuric acid consumption, dissolved iron, and magnesium extraction.     

A simple heatsink for planetary mills

This paper proposes the use of a copper disk as a simple heatsink for planetary mills in order to limit or slow down the temperature increase during prolonged grinding. A series of grinding experiments have been performed in which the surface temperature of different parts of the grinding bowl has been determined as a function of grinding time. Three different experimental series were conducted: a reference series without the heatsink, and two series where the heatsink had been given the respective initial temperatures of 25 °C and ?25 °C. As demonstrated in this paper, the proposed concept represents a cheap but effective way of limiting or slowing down the temperature increase occurring in the grinding bowl. The effect is of a magnitude that can be of considerable practical importance, and the concept can be employed on most planetary mills where the grinding bowl holder is designed to accommodate different bowl sizes.