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By Gülay Bulut, Ferihan Göktepe

Pyrite is the most abundant and widespread sulphide mineral and its flotation has been investigated by many researchers. It is well known that, pyrite is floatable under acidic conditions and recovery drops with increase in alkalinity due to changes in solubility of iron hydroxides on the pyrite surface. Most of the works also revealed out that pyrite displays variable behaviour in flotation circuits. This behaviour mainly related with origin and mineralogical composition of ore, liberation size, electrochemical behavior of minerals constituting ore, the trace and minor element content of the mineral, texture, application variables belonging to flotation circuits. In this study, copper activation effect was examined on different origin of pyrites. The experiments were performed with five high grade samples of pyrite from the different mining districts of Turkey. The pyrite samples were different from each other with respect to either their physical or chemical properties; they were different physical appearances and contained different associated minerals. In the micro-flotation tests, 1x10-4 mole/L copper sulphate and 1x10-4 potassium ethyl xanthate were used as flotation reagents. First of all pyrite samples were floated using only xanthate at different pH. Results showed that best floatation results obtained at acidic pH range for all pyrite samples. When copper ions were used to activate, all pyrite samples floated at neutral and basic pH values, too but depending on the samples, recovery and pH ranges were slightly changed.

Jan 1, 2014

By A. N. Zagorodnyaya, Ye. G. Bochevskaya, E. A. Sargelova, Z. S. Abisheva, A. S. Sharipova

"In Kazakhstan, there is a unique sulphide copper deposit, in the ores containing the natural isotope of osmium-187 (more than 90 %). In the pyrometallurgical processing of concentrates for the target metal, osmium is mainly sublimated, with wet cleaning of metallurgical gases (dust and gas mixture) 40 % of it together with rhenium passes into washing solutions (wash acid). On the basis of this acid, the production of ammonium perrhenate was organized at the enterprise of the RSE ""Zhezkazganredmet"" and for the first time in the world practice - the isotope of osmium-187. However, the degree of its recovery from solutions does not exceed 8 % of the amount supplied with the charge or 25 % of the amount that was received for extraction. But the bulk of it (~ 75 %) remains in the raffinate, discharged to sewage treatment plants (wastewater). The studies presented in this work are aimed at solving the problem of osmium isotope-187 extracting from wastewater, which are a potential source of increasing metal production in Kazakhstan. The aim of this work is the development and scientific justification of osmium deposition regimes from wastewater using such reagents as thiosulfate and sodium sulphide. The influence of various physicochemical factors on the precipitation of the osmium was studied and optimal conditions for its maximum transition to sediment were chosen. An original method for obtaining osmium concentrate in the presence of sodium thiosulfate with the addition of potassium iodide from a new type of wastewater - man-made ammonia solutions has been developed. These solutions were combined mother liquors of solid-phase stripping operations and dissolution of pulp of crude ammonium perrhenate. On the basis of the obtained results, a settling technology for extraction of the osmium from man-made ammonia rhenium-containing solutions was introduced at RSE ""Zhezkazganredmet"". It should be noted that the degree of extraction of osmium, depending on its concentration in solutions entering on the settlement, and taking into account the addition of sulfuric acid, will vary from 60 to 90 %. The recovery of osmium during the test period was 58 %."

Jan 1, 2018

By A. Wieniewski

The observed progress in technology of non-ferrous ores enrichment and condition of machines and equipment in the ores concentrators on the one hand have brought a possibility and on the other resulted in necessity for introduction of modernization actions. Positive results of investigations into influence of chemical modification of selected products of enrichment process on copper ores enrichment results gave possibility for implementation of chemical-flotation technologies of ores enrichment. The substantial development in flotation machines structure has contributed to modernization of flotation process resulting in increase of enrichment efficiency. The paper presents the idea and results of chemical modification of difficult in enrichment semi-products in the technology of copper ores enrichment and the results obtained after application of the modernised technology. The structure of large-volume flotation machines used in various stages of flotation and current situation in modernization of flotation process in copper ores concentrators are also described. The paper presents industrial-scale results of flotation after application of modified flotation reagents of collector group. Introduction of modifiers into the widely used xanthates resulted in increase of copper recovery in concentrate.

Jan 1, 2014

By Kilian Neubert, Wotruba. Hermann, Lars Weitkämper

Metallurgical slags are produced in large quantities all around the world as a result of smelting, refining or alloying processes applied to metals, ores and recycling material. The slag phase serves different purposes in metallurgical processes. It prevents heat losses and protects the metal phases from contaminations and absorbs gangue and extracted impurities. Due to achievable metallurgical efficiencies certain amounts of valuable metal contents always remain in the residual slag. Recycling of these metallurgical slags has several positive effects, such as optimized ecological and economical compliance of the producing industry. If recycled metal contents can be reintroduced into existing processes, required inputs of raw material can be lowered. High metal prices enable the economic recovery of metal contents in metallurgical slags by applying different non-metallurgical processing methods. Conventional processing of metallurgical slags utilizes processes such as hand picking, selective comminution, screening and magnetic separation. Those state of the art processing techniques have limitations in terms of achievable metal recoveries and required energy demand. The approach for an optimized processing of metallurgical slags targets enhanced metal recoveries in combination with lower specific treatment costs. Processes such as coarse and ultrafine density separation, as well as sensor-based sorting as a pre-concentration stage are used to enable an efficient recovery of metal contents in slags. By applying these methods a completely dry working process is possible. This paper reviews the commonly used methods and describes approaches for innovative processing of metallurgical slags to enhance metal recoveries. Focus of this paper is laid on sensor-based sorting as well as on processes for dry density separation.

Jan 1, 2014

By Sanja Miskovic, Manjunath Basavarajappa

Numerical investigation of fluid flow characteristics and mixing efficiency of different impeller designs, based on the mixing characteristics of a passive scalar/tracer, is conducted using CFD for lab scale flotation cells of diameter D and height H equal to 300 mm. Fine computational mesh containing close to 1.6 million cells is used to obtain fully resolved steady state flow field close to the impeller. Unsteady scalar mixing simulations are performed on a relative coarse grid consisting of approximately 450,000 cells. Based on detailed numerical simulations, flows are fully characterized for four impeller designs at two off-bottom clearances C of 50 and 25 mm (D/6 and D/12). Passive scalar is injected from the cell bottom and the time required to attain uniform concentration distribution inside the cell volume is taken as the mixing time by monitoring scalar concentration at four discrete locations. The effect of varying agitation speed and off-bottom clearance on mean flow characteristics is also investigated. The scalar mixing results reveal regions of low mixing efficiency and are used to understand mixing process in greater detail. Through the previous research work, it was found that the main flow pattern undergoes transition from double-loop to single-loop configuration with decrease in off-bottom clearance below a certain critical height for stirred tanks using flotation impellers in the absence of stator. The inclusion of stator is found to alter the mean flow behaviour by preventing the jet from transitioning to axial type flow for lower clearance. Power numbers for different impeller designs are calculated based on the torque generated on impeller blades, and also from the total turbulent kinetic energy dissipation in the overall fluid volume. Power number is found to exhibit a significant dependence on off-bottom clearance based on the CFD results.

Jan 1, 2014

By David I. Verrelli, Boris Albijanic

"Froth flotation is an exceedingly complex physicochemical process. The convenience of distilling much of the complexity of the bubble-particle interactions into a single parameter has led to the continuing popularity of the classical ‘induction time’ to quantify the threshold for bubble-particle attachment to occur. Despite this popularity and the simplicity of the concept, there is no single universal method of evaluating the induction period. In this paper, we begin with a critical review of the available techniques for estimating the induction period. These are: back-calculation from experimental (micro) flotation tests; pushing a particle toward a stationary bubble (or vice versa) using an atomic force microscope (AFM); pushing a bubble toward a stationary bed of particles in the ‘Induction Timer’; pushing a bubble toward a stationary solid surface using the ‘integrated thin film drainage apparatus’ (ITFDA); and dropping particles onto a submerged stationary bubble using the ‘Milli-Timer’ device. Each one of these methods has advantages and disadvantages, and the best choice depends on the application. In the experimental section, we present quantitative comparison of the induction periods estimated using two different techniques, namely the Induction Timer and the Milli-Timer. The same particles were tested in each device, under the same conditions. It was found that by tuning the operation of the particle pick-up device, similar estimates of induction period could be obtained to the estimates made by direct observation with the Milli-Timer. In the former device a bubble is driven toward a particle bed at a controlled rate, whereas in the latter a particle’s motion is governed by the hydrodynamics. The potential to match these presents an intriguing prospect for better understanding the bubble-particle interaction, and the possibility to ‘calibrate’ the simpler Induction Timer against direct observations."

Jan 1, 2014

By Glasswell K. Nkonde, Fredrick Chileshe, Steven Simukanga

Improving food security by increasing crop output is a primary objective for most countries in the sub- Saharan region. Most soils in the region are deficient in phosphorus an important nutrient after nitrogen in plant growth. Phosphates are therefore used together with nitrogen and potassium based fertilizers in improving crop production. Zambia like many other countries in the sub-Saharan region imports phosphate fertilizers for use by its farmers. However, indigenous phosphate deposits of igneous origin exist in Zambia of which the main ones are Chilembwe in Petauke district, Mumbwa sugarloaf in Mumbwa district, and Nkombwa Hill in Isoka district. If commercially exploited, these deposits could become a ready source of inexpensive phosphate for the local farming communities in the country. Previous research has shown that partially acidulated phosphate rock (PAPR) produced using Chilembwe phosphate rock is just as agronomically effective as the imported superphosphate fertilizers on food crops in most of the Zambian soils. PAPR is produced by reacting finely ground phosphate rock with only a fraction (usually 50%) of the stoichiometric quantity of the sulphuric acid required to fully convert the apatite in the rock to monocalcium phosphate thus making the process cheaper in terms of acid consumption than the poduction of single superphosphate where the apatite in the rock has to be fully converted to monocalcium phophate. The factors that affect the percentage of available P2O5 in the PAPR produced include particle size of phosphate rock, degree of acidulation, concentration of acid, process temperature profile, residence time and drying temperature. These factors were studied in a 25kg/hr PAPR pilot-plant in the School of Mines at the University of Zambia using Chilembwe phosphate rock. The PAPR pilot-plant consisted of a rotary drum granulator as the main reactor, a rotary drum dryer and a triple deck vibrating screen. Ground phosphate rock was reacted with sulphuric acid and water in the granulator to form a granular product which entered the drier. The dry product was screened into three size fractions. The 1 – 4 mm size fraction was the final product whilst the –1mm size fraction was recycled. The +4mm size fraction was sent to a crusher while the crushed PAPR was sent back to the screens.

Jan 1, 2003

By A. D. Shuloyakov, G. A. Finkelstein, V. I. Kurets, A. K. Vedin

"In disintegrating solid ores with mineral values disseminated in them, one of the principal tasks is to selectively separate valuable mineral particles from the matrix in which they reside, striving at the same time to keep their natural shape and size as intact as possible.The conventional mechanical disintegration methods do not meet this most important requirement in full. The other major objective here is to reduce electric power consumption during the fragmentation stage, since up to 65 per cent of the total power consumption at concentrators processing ores of non-ferrous, ferrous, rare, and precious metals is spent on disintegration steps.Numerous experiments conducted on dozens of ore types (tin, tungsten, and many others, as well as on rock containing diamonds, rubies, emeralds, etc.) have invariably demonstrated by far superior integrity of mineral values as compared with that obtained by mechanical as well as other types of disintegration.However, the process remains as power-intensive as ever which is due to the imperfect nature of the electro-technical devices that convert the network electric current into high-voltage electric impulses responsible for the disintegration process.The maximum capacity attained in disintegrating the hardest materials from 40 mm to 2 mm size in a single operating chamber with four high-voltage electrodes does not exceed 1.0-1.5 t/h. Attempts at increasing the number of electrodes and the capacity of the operating chamber have failed so far because of high overall dimensions and mass of the electrical part of the installation supplying the chamber with pulsating current, as well as of the commutation requirements. At the same time, a number of overseas operations have reported a significant breakthrough in their attempts to reduce dozens of times the overall dimensions and mass of the discharge impulse generating installation (EIS), which enables us to contemplate to increase the capacity of the EIS at least ten times in the nearest future. Yet, even with the currently achieved capacity of the EIS, these installations have found wide use in geological surveying, precious and semi-precious stones production, in grinding super hard materials, etc., since they provide spectacular selectivity of disintegration and thus the integrity of valuable mineral crystals."

Jan 1, 2003

By A. N. Zagorodnyaya, T. N. Bukurov, Z. S. Abisheva, A. S. Sharipova

"This work presents the results of the development of a flowchart for the processing of copper production dusts resulting in the production of of lead sulphate, tribasic lead sulfate (TBLS), ammonium perrhenate, zinc, copper-cadmium cake.Dust having the following composition (wt. %): Pb 39.3, Zn 6.5, Cu 5.5, Stotal 10.8, As 1.05, SiO2 0.6, Cd 0.8, Bi 0.04, Ag 0.02, Re 0.03 was used in the work. Chemical and X-ray analyses indicated that lead, zinc and copper are generally present in dusts in the form of sulfates, various amounts of oxides, sulfides, silicates; sulfur (according to chemical analysis data) is present in the form of sulfates (~ 90%) and in the form of sulfides (~ 10%).Studies have been made on two-stage stripping of copper production dusts containing lead, zinc, copper, cadmium, rhenium in the form of various compounds, at first with soda solution, then with nitric acid solution. It has been indicated that the combination of the operations allows the recovery of 98% rhenium, 82% lead, 92% copper, and 96% zinc and cadmium each in the solution. Furthermore, at the stage of dust leaching with soda, 80% of rhenium and 0.01% of nonferrous metals each are recovered in the solution. Remaining quantities of all metals are recovered in nitric acid solution up to maximum values. Such behaviour of metals is caused by metal phase content in dusts and properties being formed during leaching of compounds. At the stage of dust leaching with soda solutions, carbonates with various compositions are formed due to the exchange reaction between nonferrous metal sulfates and sodium carbonate. The conversion of PbSO4 into PbCO3 is apparently occurs in the solid phase and the conversion of readily watersoluble zinc, copper, and cadmium sulfates into that of carbonate salts occurs in the liquid phase. Rhenium passes into solution due to the presence of rhenium greater quantity in dust in the form of readily soluble perrhenate. Some portion of the perrhenate is, probable, blocked by lead sulfate and liberated in the conversion of PbSO4 into PbCO3."

Jan 1, 2003

By Alan Bax, John D. Lewins

"Platinum Australia Limited (“PLA”) is currently undertaking a Feasibility Study into the development of the Panton Platinum Palladium Project located in the Kimberley region of Western Australia. The Panton Intrusive is a layered, differentiated mafic to ultramafic structurally complex body that has undergone numerous folding and faulting events. This has resulted in a southwesterly plunging synclinal structure some 10 kilometres in length and 3 kilometres wide.High grade PGM mineralisation is hosted within a number of stratiform chromitite reefs within the ultramafic sequence. The focus of the Feasibility Study is on the Top Reef which is situated approximately 150 metres below the mafic/ultramafic interface. The average true width of this reef is approximately 1.2 metres and the reef has a defined resource is 10.1 million tonnes at an average grade of 6.1 g/t PGM + Au.The Palladium to Platinum ratio of the Top Reef is approximately 1.15:1 and the reef also contains an average of 0.4 g/t Au and 0.3 g/t other PGM’s. The Panton ore is considered refractory in nature and standard flotation techniques achieve a PGM + Au recovery of less than 65% with a concentrate grade of less than 100 g/t PGM + Au. As part of the Feasibility Study Platinum Australia in conjunction with their major shareholder Lonmin Plc developed a new metallurgical process which has achieved an overall PGM + Au recovery of approximately 80% into a high grade PGM concentrate which is suitable for feeding direct to a PGM refinery.THE PANTON PROCESSThe Panton Process incorporates a two-stage Milling and Flotation circuit with no cleaning stages to produce a low grade, high recovery flotation concentrate. Typically the concentrate mass constitutes 25% of the feed and has a PGM + Au grade of 25 g/t.This low-grade concentrate is subjected to calcination in a rotary kiln or fluid bed at a temperature of approximately 400 0C for a period of 60 minutes. This step is designed to render the platinum minerals soluble in cyanide and both the temperature and the duration of calcination are critical to the efficiency of the leaching process."

Jan 1, 2003

By S. K. Kawatra, T. C. Eisele

"Coal-fired electrical power generation plants produce 56% of all the electricity in the U. S., burning approximately 860 million tons of coal annually. This results in production of 58 million tons of fly ash; 19 million tons of bottom ash and boiler slag, and 23 million tons of desulfurization sludges. These are a considerable disposal problem, and it would therefore be very valuable to be able to sell these residues and byproducts as useful materials, rather than paying to dispose of them in landfills. However, utilization rates are currently very low, below 32% even for the most thoroughly commercialized combustion byproduct (fly-ash). This poor utilization rate is a result of the high impurity content of combustion byproducts, which make them unsuitable for existing markets. The experiments described in this paper were therefore carried out to study the application of mineral processing techniques to upgrading combustion byproducts.Experiments were conducted to upgrade and purify low-grade combustion products to produce highgrade, readily saleable materials. A process was developed for removing unburned carbon from coal fly-ash by froth flotation, using reagents originally developed for flotation of oxidized coals. This technique could reduce the carbon content of fly-ashes containing up to 11.1% carbon to less than 2%. The froth product was sufficiently carbon-rich that the recovered carbon has value as fuel, or as low-grade activated charcoal.Fly ash was also examined as a binder for producing iron ore pellets, primarily as a substitute for bentonite. It was found that Class F fly ash in combination with an activator (calcium hydroxide) was an effective pellet binder at dosages comparable to the currently-used dosages of bentonite clay. While the flyash worked very well alone, tests in combination with bentonite showed that fly-ash/bentonite mixtures had significantly worse binding performance than either pure bentonite or pure fly-ash-based binder. This was determined to be due to the fact that bentonite and fly ash have profoundly different binding mechanisms, with fly-ash undergoing a pozzolanic chemical reaction while bentonite binds through a physical attachment process."

Jan 1, 2003

By Bertil I. Pålsson, Sture V. Persson, Kent T. Tano

"Autogenous (AG) and semi-autogenous (SAG) mills often operate in an unstable state because of the difficulty to balance the rate of replenishment of large ore particles from the feed with the consumption in the charge. This has led to an increased interest in obtaining an accurate and direct measurement of mill load and the behaviour of the mill charge. A problem in attempting to measure those parameters is getting a sensor that is both robust and sensitive enough for continuous on-line measurement.Techniques that measure the force acting on the lifter when it hits the charge inside the mill have got an increased interest recently because of the ability to combine it with DEM modeling. This paper presents a method for the measurement of the apparent filling level in a pilot ball mill equipped with the Metso Continuous Charge Measurement system (CCM). In this technique a strain gauge sensor is mounted on a flat steel spring which is fitted into a recess underneath a lifter bar. In addition, a conductivity probe, the Boliden Grinding Measurement System (GMS), is mounted in another lifter bar with the purpose of comparing the two measurements. To visualise the behaviour of the charge a 2D DEM modelling technique is used with Metso HFS software.The pilot mill is a grate-discharge mill, equipped with 12 rubber lifters of square size 0.1 m, face angle 45 degrees. Steel balls in the size range 10-30 mm and density 8000 kg/m3, are used in the experiments. Test material is a hematite pellet feed, d50 around 45um, chosen to get stable grinding conditions with respect to feed size variations.The experiment is meant as a first test of the equipment’s performance, and to study how the sensors respond to changes in some operating variables. To get an understanding of, and distinguish between correlation and causality, experiments are performed according to a full factorial design with two levels, and in addition two replicates of the experiments at a center-point in the charge level setting."

Jan 1, 2003

By Stephen Gay, Dewetia Latti

"When improving a plant a mineral processing engineer would like to know the mineral composition distribution. In order to assess this distribution in particles, as compared to the corresponding distribution in particle sections, a procedure needs to be applied to the observed particle sections. Such a procedure is called a stereological correction. There exists many stereological correction procedures, but one group of such procedures is based on using those stereological equations that do not require making assumptions regarding particle shape or texture.In particular Davy derived an equation to estimate the covariance of mineral composition distributions for particles from both particle sections and linear intercepts. Furthermore, this covariance can also be used to measure the variance of the composition distribution of each mineral.This paper shall present results verifying this equation. The experimental method was based on obtaining particle information from serially-sectioned particles. Particle samples selected for serial sectioning displayed a range of natural textures including liberated, binary and composite particles. The exposed surface of the polished sections were analysed using QEM*SEM. This procedure was repeated until all of the grains have been polished away. The covariance of mineral distributions were calculated from the serial sections and compared with that calculated using Davy’s equation."

Jan 1, 2003

By Wu Tao, Lu Shijie, Xia Xiaoou, Zhang Ming

The impeller and stator are the most important components that can significantly influence the performance of flotation cells directly. The performance of the impeller and stator of KYF flotation cell is revealed by CFD (Computational Fluid Dynamic) simulation in this paper. A method of obtaining the circulation volume is proposed. The circulation volume is proportional to third power of impeller diameter. And the circulation volume increases linearly with the increasing of the speed. CFD simulation can draw the conclusions that there is a proportional relationship between power consumption, third power of speed and fifth power of impeller diameter. CFD prediction results agree with the theoretical formula. The difference between the forward and backward impeller is analysed in the paper. The circulation volume of the backward impeller is obviously larger than that of the forward impeller under the same condition, but power consumption of the backward impeller is lower. It is demonstrated that the impeller and stator system of KYF flotation cell is an efficient and reasonable stirring mechanism. CFD simulation could be reliably used in flotation cell design and optimization.

Jan 1, 2014

By Wei Sun, Xue-feng Cao, Lei Sun, Run-qing Liu

The flotation performance of kaolinite in laboratory micro-flotation was investigated with 3- (dodecyloxy) propan-1-amine (ON12) as the collector. Compared with dode cylamine (DDA), which is commonly used as the bauxite reverse flotation collector, ON12 shows stronger collecting ability for kaolinite than using the same amounts of DDA, when experimental pH value reaches more than 6. With the addition of ON12, the isoelectric point of kaolinite shifts to 6.1 from 3.6, which suggests a strong adsorption between kaolinite and ON12. Infrared spectra analysis indicates that the type of ON12 adsorbed on the kaolinite surface is mainly of electrostatic adsorption. Molecular dynamics simulation further indicate that electrostatic adsorption of ON12 cation on kaolinite (001) surface based on electrostatic interaction between NH4+ and O atoms. ON12 will be suitable for bauxite reverse flotation with the efficiency depressors of the diaspora.

Jan 1, 2014

By David De Klerk, Maximilian Richter, Thirunavukkarasu Pathmathas, Indresan Govender

We consider the average power dissipation distribution for dense granular flows of an idealised mono-sized charge (5 mm glass beads) in a 300 mm diameter tumbling mill operated in the cascading-to-cataracting Froude regime. By combining the inherent frictional nature of particles within a dense flow regime and noting that tumbling mills are typically characterised by collisional and turbulent stresses, we derive the power dissipation distribution for athermal boundary conditions. The input data to the model (velocity, solids concentration, pressure and flow depth) are obtained directly from non-invasive measurement using a nuclear imaging technique: Positron Emission Particle tracking (PEPT). The power dissipation due shear stresses is shown to be the dominant mode of dissipation in the tumbling mill with up to two orders of magnitude difference between pressure induced dissipation and shear.

Jan 1, 2014

By Li Maolin, Cui Rui, Wang Guanghui

"As the stirrer speed directly affects the motion of grinding media, and hence the interactions with material particles to be ground, the rotation speed of a stirred mill is well recognized as one of the most important factors affecting the grinding performance. In this work, the effect of stirrer speed on grinding performance in terms of grinding rate and grinding energy efficiency was quantitatively investigated. Batch wet grinding tests of quartz powder at different stirrer speeds were performed, using a self-made horizontal stirred mill with alumina balls as grinding media. Grinding time, operating power and the sizes of ground product were recorded to analyse the grinding performance. The test results showed that grinding processes followed the first order grinding kinetics. The grinding rate constant increased with the increasing stirrer speed. Based on the grinding kinetics, the grinding energy equation was established. The energy utilization coefficient in the energy equation can be used as an indicator of energy efficiency. The energy efficiency decreased with the increasing stirrer speed. Quantitatively, with the stirrer tip speed increasing from 4.25m/s to 9.35m/s, the grinding rate constant was improved by 441.47%, whereas the energy utilization coefficient was decreased by 39.35%. The grinding rate constant and energy utilization coefficient can be used to evaluate the grinding performance when the grinding conditions need to be adjusted. The adjustment of grinding conditons would be optimal when grinding rate get significant improved with little decrease of energy efficiency."

Jan 1, 2014

By Manisha Singh, Manikanchan Sarkar

"Presently the global steel production is mainly based on iron ore sourced from high grade hematite deposits. India has large deposits of good quality hematite ore but with the prevailing practice, such deposits are depleting fast and getting exhausted. However, apart from such high grade iron ores, India is also endowed with large reserves of low grade iron ore, such as BHQ (Banded Hematite Quartzite), BHJ (Banded Hematite Jasper), BMQ (Banded Magnetite Quartzite) and Magnetite. In the present situation, it has become necessary to focus attention on upgradation of these low grade ores to produce useable concentrate. It augurs well that already substantial investigation work has been carried out in this respect and a few plants have been set up in India and abroad. One important aspect in this regard is the process route options with optimum yield and grade. Some of these low grade ores, particularly magnetite, are found to contain high phosphorous level, removal of which poses a major challenge. Since the grade is low, substantial upgradation necessitates fine grinding for appropriate liberation. Fine grinding is energy intensive and needs attention to achieve economy in this regard. Large quantity of tailings is generated in the process, wet disposal of which gives rise to a lot of hassles and hence, possibility of dry disposal of the same needs to be explored. The experience of the authors, with regard to aforesaid aspects which has been gained in course of carrying out relevant assignments is presented in this paper."

Jan 1, 2014

By Michel Brissette, Shaheer Sanuri, Veroljub Mihajlovic

"The main purpose of geometallurgy is to improve the mine planning by mapping hardness and throughput into the mine block model using geostatistics to eliminate the large production fluctuations caused by the ore hardness variability coming from different parts of the pit. However, not all grinding testwork available commercially can be applied to geometallurgy because of large amount of sample required and the associated cost per test. Geometallurgy is now being used in designing grinding mill but as the mining project progress, the accuracy of engineering must improve from 30-35% during the scoping study to 10-15% at the bankable feasibility stage and less than 5% for Engineering Procurement and Construction Management. For SAG milling, the SAG Variability Test (SVT) is an abbreviated version of the standard SAGDesign test developed for this geometallurgical need. The same amount of ore than the original test is ground for three cycles into an open batch SAG test and then the SAG pinion energy is predicted to the standard end point of 80% passing 1.7 mm. For ball milling, a Bond Variability Test (BVT) was also developed using the SAG ground ore from the SVT. The Bond ball mill grindability (Gpb), corresponding to the average of the net grams per revolution from the last three cycles of the Bond test, is also predicted from the third cycle. The relative error for both the predicted SAG ore hardness (WSVT) and Bond Work Index on SAG Variability ground ore (Sv-BVI), in kWh/t, averaged 5.5% and 3.8% respectively compared to the SAGDesign testwork values and are within 7% and 10% respectively from plant benchmark and within 7% from the samples sent to all partner laboratories. Both tests can be continued to measure the true hardness necessary for mill sizing. This article discusses the development of these two new tests, SVT and BVT, their potential use and accuracies in any geometallurgical program."

Jan 1, 2014

By J. Gebhardt, D. McBride, M. Cross, S. Gilbert

Heap leaching of ores containing chalcopyrite offers a low cost option to treat low grade complex copper sulfide ore bodies. However, the slow leach kinetics of chalcopyrite at ambient conditions typically results in unacceptably low recoveries. As the mineralogy of new ore deposits becomes more complex and the effects of gangue chemistry exert a greater influence on copper sulfide leach behavior, the control of leach reaction chemistry becomes more critical. Computational fluid dynamics (CFD) technology has been useful in the capture of multi-phase interactions and provides a path to simulate complex fluid flow, gas and liquid phases, through the heap, along with species transport. Coupling the flow models with the reactive mineral dissolution, and the sometimes competing aqueous chemical and solid-liquid reactions, yields a comprehensive model of the heap leach process. The leach kinetics of chalcopyrite dissolution in heap conditions is examined, and the incorporation of chalcopyrite leach kinetics in the CFD model is described. The application of the model to investigate optimization of the heap is discussed.

Jan 1, 2014