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By S. Banisi, S. Zare, M. Mahmoudi, M. Maleki-Moghaddam, M. Yahyaei

As the orebodies become more complex and lower grade, there is an increasing demand from crushing circuits for a finer product. When the product from a crushing circuit does not satisfy the requirements of the downstream processes, the cost of production increases. Thus, enhancing the performance of cone crushers through optimisation of their operating factors becomes important. One of the key factors affecting performance of cone crushers is the filling of the crushing chamber during the operation. The effect of varying the filling of the crushing chamber of a cone crusher was investigated. To assess the crusher performance, parameters such as crusher capacity, power utilisation, and reduction ratio were used. Extensive plant trials were conducted to obtain samples and operational data when the crusher chamber was 50% full (half-full) and 100% full, which is 300 mm above the top of the crusher chamber (super-choked condition). The results indicated a 62% difference in the crushing capacity when the crusher was operated in a super-choked condition compared to the half-full chamber level. In the super-choked condition, the size-specific energy at 12.7 mm was 20 % lower than that of the half-full condition. The Bond ball mill work index (BBWI) of crusher product indicated that the BBWI of the product of the super choked condition is 5.6% less than that of the product of half-full condition. This can be attributed to the generation of micro cracks in particles due to increase in the interparticle breakage when the crusher was operated in the super-choked mode. Keywords: Cone crushers, crushing efficiency, crushing chamber filling

Jan 1, 2020

By Nelson Sánchez

The Los Bronces Development Project (LBDP) was approved in October 2007 by the Anglo American plc Board at a capital cost of some US$1.74B, and is scheduled to be completed by the end of 2010. The project represents a big step forward for the Los Bronces mine since it will increase the treatment rate by 140%, the mining rate by 64% and almost double copper production during the first four years of the project. This economy of scale will also position Los Bronces as a much bigger and efficient mining operation, better prepared to face adverse market scenarios and allowing Anglo to consolidate its position in base metals. The project addressed numerous scope alternatives such as plant capacity vs mine capacity, cut-off grade strategy to maximise value based on the existing flotation and leaching process, cutback strategy, etc. Main infrastructure aspects of the project are as follows: ?Dump area with 2400Mt capacity for waste and low grade material ?Mining equipment: 1 x electric shovel, 12 x 300t trucks, 1 x drill ?Primary crusher: 60? x 89? ?Three new power lines (144MW capacity) ?Grinding plant at 87kt/day nominal capacity (includes a 4.3km tunnel with a regenerative conveyor belt) ?Grinding plant: 1 SAG mill 40? x 25?, 2 ball mills ?Slurry pipeline (52 km from grinding to flotation plant) with five choke stations ?Flotation plant (rougher, cleaning, regrinding circuits and molybdenum flotation). The history of Los Bronces started in 1867 when it was discovered. In 1978 ENAMI (Government company) sold Los Bronces to Exxon for US$112M. The same year a snow avalanche destroyed San Francisco Flotation Plant. In 1979 the plant was rebuilt at 4.9kt/day and from 1979 to 2002 several plant expansions were made up to 54kt/day capacity. In 2002 Anglo acquired Los Bronces for US$1.3B. The 2002 resource and reserve statement showed reserves at 459Mt at 1.02%CuT and resources at 780Mt at 0.80%CuT. The 2007 statement shows reserves at 1,480Mt at 0.67%CuT and resources at 1,644Mt at 0.45%CuT. Defining the size of a plant expansion is the key challenge that any expansionary project faces. The scope of this paper is a general review of the project, describing all the mining techniques that were used to define the best options that maximise value. In this context, a range of plant capacities (from 58 to 230kt/day) combined with a range of mining capacities (from 83 to 166Mt/a) were evaluated in detail, defining for each option a mine plan, Capex and Opex, and cut-off grade strategy for flotation and leaching processes and the associated Net Present Value (NPV) that was used as the selection criteria. The Los Bronces mineral resources potential provides good growth opportunities for the mine beyond LBDP implementation, offering an exciting invitation to continue studying new initiatives that maximise the value of this deposit.

Jan 1, 2008

By E. Cetin

Orebodies that contain more than one economically important mineral are generally evaluated by parametric cut-off grades. This approach often leads to mis-valuation of mineral deposits because cut-off grades are not based on the grades of each individual mineral and, because of the parametric formulation, are only indirectly related to the individual grade distributions. The only realistic approach is a formulation that accounts separately for each component mineral. The grid search method can be used as a means of multi-mineral cut-off grade optimization in this context. This paper describes the use of the grid search method in cut-off grade optimization for multi-mineral deposits. The authors introduce the general concepts of the method and formulate its application to cut-off grade optimization; they describe a software implementation of the method that can accommodate cut-off grade optimization for mineral deposits that contain up to three economic minerals. Keywords cut-off grade, optimization, grid search, multi-mineral deposits.

Aug 1, 2013

By C. P. Smyth

Large volumes of quantitative geochemical data, presenting a formidable interpretation task to geologists, are generated in the course of exploration drilling for new ore deposits. The incorporation of geological expertise into computer programs developed for the interpretation of this data would improve the speed and accuracy with which it is interpreted. A detailed analysis of the geochemical log interpretation problem, together with a survey of relevant research in Artificial Intelligence and petroleum well- logging techniques, confirms, that artificial intelligence, particularly as developed in the fields of expert systems and pattern recognition, provides techniques for implementing such data interpretation programs. In addition, it provides data analysis techniques significantly more powerful than those currently applied to geochemical data, and very appropriate to the hypothesise and test' character of mineral exploration problem-solving. The conversion of raw borehole log data to meaningful higher level geochemical entities, in terms of which geological reasoning may take place, is a critical step in applying these techniques to log data analysis. A method of achieving this conversion as an exercise in signal-to-symbol transformation is presented.1

Jan 1, 1987

By L. A. Sandham

Environmental impact assessment (EIA) in South Africa is administered by the Department of Minerals and Energy (DME) for the mining industry, and the Department of Environmental Affairs and Tourism (DEAT) for all other sectors. Thus far, EIA research has focused on the process under the auspices of DEAT but none has focused on the EIA process that has been mandatory in the South African mining industry since 2004. Using the Lee and Colley(Lee et al.13) review package as a basis, a review model specifically tailored to the needs of the mining industry was applied to review the quality of a sample of 20 environmental impact reports (EIRs)approved by the DME. Results reveal that 85% of the EIRs are of a satisfactory quality. Presentational and descriptive components generally achieved higher quality grades than the analytical components such as impact magnitude and alternatives. The results show that in spite of some areas of weakness, and in spite of being conducted in terms of different legislation, EIR quality in the mining sectors appears to conform to the overall standard of quality of EIRs in other sectors in South Africa, and is also on a par with quality abroad. Hence, despite the criticism that DME is usurping the role of DEAT, it is concluded that EIRs of comparable standard are being produced, and that the quality of EIRs in the mining sector do not provide supporting evidence for this criticism. However, EIR quality is only a single aspect of EIA effectiveness, others including action and implementation of EIA proposals and mitigation measures.

Jan 1, 2008

By W. C. Joughin

A risk evaluation model has been developed, that enables the quantification of expected injuries and economic losses resulting from rockfalls associated with a support system in a given geotechnical environment. This enables the comparison of different support systems based on the expected frequency of injuries and the total cost of the support system, including the expected economic losses. The model takes into consideration the variability in joint orientations and joint strength within the geotechnical environment and the variability and quality of support systems. The quantification of the benefits of other rockfall risk-mitigating strategies, such as the effectiveness of barring or monitoring, are also incorporated in the model. The risk model is described in this paper and the validation of the model and case studies are described in Part II.

Jan 1, 2012

By K. Eckert, A. E. Sommer, S. Heitkam

One highly relevant challenge in flotation is the recovery of fine particles. Due to their low inertia, these particles are mostly pushed aside by rising bubbles. Consequently, they typically exhibit low probability for bubble-particle collision and thus, a poor recovery rate. In this work, the trajectories of fine particles in the vicinity of rising bubbles were investigated. The measurements considered a chain of millimetric bubbles in a rectangular container filled with deionized water. The flow field near the bubbles was measured by tomographic particle image velocimetry (TPIV), employing fluorescent tracer particles of 33μm diameter. Subsequently, trajectories of larger particles and bubble-particle collision events are recorded with 4D particle tracking velocimetry using a high temporal and spatial resolution. The results reproduce the well-known collision of particles on the leading edge of a rising bubble. Additionally, collisions on the tailing edge were observed in cases with a low Stokes number. The TPIV results demonstrate, that the high turbulent kinetic energy in the bubble wake allows particles to divert from the fluid streamlines and collide with the tailing edge of the bubble. The tailing edge collision probability increases with the bubble Reynolds number and decreased particle inertia. Overall, the investigation shows that the collision of fine particles in the bubble wake should be considered for the development of further collision probability models. Furthermore, the importance of turbulence on the fine particle flotation was demonstrated. Keywords: 4D Particle tracking velocimetry (PTV), tomographic particle image velocimetry (TPIV), fine particle flotation, collision probability, bubble-particle interaction, wake induced collision

Jan 1, 2020

By A. J. Feltenstein

A theoretical model for prediction of dissolved gold loss from a rotary vacuum filter was developed as part of a program to implement optimizing control on gold plants The filter was modelled using a modification of the Hagen-Poiseuille equation and the Choudhury and Dahlstrom washing equation. The model has three parameters which must be determined from accumulated data before predictions can be made. These are the cake and cloth resistances and the filter washing efficiency. The model predicts with a fair degree of accuracy the response of the filter to changes in the independent variables and will therefore be a useful tool in the plant optimization study.

Jan 1, 1987

By E. -U. Reuther

To improve the climatic environment underground the German coal mining industry has installed refrigerators with a capacity of 300 MW cooling power. One third of this total is accounted for by central refrigeration plants with more than 5 MW capacity. The central refrigeration comprises three stages: mechanical cooling, transport and distribution of chilled water via pipelines and exchange of heat in the air coolers. Of these three stages only the mechanical generation of chilled water has been resolved satisfactorily. Transport (including distribution) and exchange require further technical research. The Mining Institute of the Aachen University of Technology has done extensive studies on this subject, concentrating on controlling the flow of chilled water according to demand in the mine. Additionally, the favourable exchange in the air coolers is taken into account. The studies are based on a model mine which simulates the conditions in deep coal mines in Germany. The simulation program represents firstly the condition of the mine air at the location of the air being cooled, then the heat exchange in the air cooler itself, and finally the transport of the chilled water from the central refrigerator to the consumer. The program also reproduces the seasonal and operational changes in conditions of mine air as well as different air cooler designs. A regulated distribution of chilled water is also suggested. Finally, estimations of costs can be made as an aid for further planning.

Jan 1, 1987

By W. J. Jordaan

The ability of soils to retain and hold cations is called the cation exchange capacity (CEC). Adsorbed cations, however, can be replaced by other cations through the process of cation exchange. Heavy mineral deposits contain a significant portion of slimes including the very fine clay minerals. In most circuits the slimes fraction hampers efficiency and has to be removed during early stages of the process. Being a nonvaluable process stream, the processing of the slimes is a costly exercise. In thickening applications it might be possible to manipulate the ability of clay minerals to retain and hold positively charged cations in order to optimize flocculation and therefore reduce flocculation costs. The CEC value of the slimes could be a measure of this possibility. To investigate this hypothesis, different slimes samples were characterized for size, chemical composition and clay minerals present. The CEC value of each sample was determined and explained in terms of the mineral phases present. Settling tests were conducted and an attempt was made to explain the CEC characteristics of the clay minerals in terms of settling rate. The scope of this investigation was to test the hypothesis and should not be seen as an in-depth study of the subject matter. A greater number of samples and detailed analyses to identify and quantify minerals and phases present would be required to investigate further the relationships between clay minerals in the deposit, CEC and settling rates. Location and geological setting Exxaro KZN Sands is mining a heavy minerals deposit at Hillendale mine located at Richards Bay, South Africa. The orebody consist of a dune some 8 km inland from the coast to the north of eSikhawini. A geological synopsis undertaken by Botha1 concluded that the Hillendale mining area represents a coastal dune sequence that accumulated north of the Tugela River. The northward movement of sediment by long shore drift currents result in a high sand budget on the beaches north of the river youth and is the dominant contributing factor to the accretion of the coastline and development of parallel dune systems along the coastal zone in this area. Beach swash zone processes and enrichment of the fine sand fraction occurs during wind transport off the beach by ascending parabolic dunes that form the coastal barrier dunes. This process results in heavy mineral concentration on the windward dune face and localized enrichment parallel to the dune forms. The aeolian sedimentary facies processes result in the primary concentration of higher grades in the upper silt-enriched reddish brown horizons of the profile. The synopsis by Botha1 further showed that an additional component of heavy mineral concentration can be ascribed to the post-depositional relative enrichment of heavy minerals due to weathering. Therefore, it is likely that the Hillendale orebody protolith was pedogenically weathered to form the homogeneous, silt-enriched, clayey reddish brown to dark red, palaeosol profile that grades downwards into yellowish brown, less clay-enriched sand. Slimes characterization Samples The Hillendale orebody was divided into geochemically discretized zones based on the concentration of ilmenite and other specific minerals and/or phases. Six composite samples were made up from exploration samples taken from Zone 9, Zone 11 and Zone 12. The red and orange horizons within these composite samples were separated and test samples were prepared to conduct test work and analyses. The test samples were submitted for preparation and analysis to characterize the slimes properties. Each test sample was deslimed at 45 µm representing the slimes fraction (-45 µm) and dried at 40°C. The -45 µm slimes samples were then split to conduct size analysis, chemical analysis, XRD and to determine the CEC. Size analysis The size distribution of the composite samples is shown in Table I. The size analysis shows that there is a difference in size distribution between the orange and red horizons of each zone. The orange horizons contain less 850 µ material as well as less slimes (-45 µm). The red horizons contain yore slimes (-45 µm). The size distribution of the slimes fraction (-45 µm) was determined using Malvern size analysis. The cumulative particle size distribution of the slimes is shown in Figure 1. The red horizons of the different zones show similar size distributions. The orange horizons show some differences. However, there is a difference in size distributions between the red and orange horizons.

Jan 1, 2007

By N. Naude, W. P. Roux, O. Kaba

Ceramic filtration plates are used for dewatering slurries with high solid concentrations and particles in the size range of 30 μm to 150 μm. After extensive operating times, the ceramic filter plates at Minas Rio in Brazil begin to blind due to the precipitation of barium sulphate on the surface of the ceramic filter plate, which reduces the overall efficiency of the filtration process. Further research is needed to identify the barium precipitate phases on the blocked filter plates and its thermodynamic and kinetic behaviour in the system. A crystallography analysis was performed on the sample and the results showed that there are two different sets of precipitates formed on the filter plate: grey crystals which were large and scatted on the filter plate and red micro crystals that covered the entire surface of the filter plate. The red microcrystals on the sample were identified with XRD, which indicated that the sample contained barium sulphate in the mineral form barite. Experiments were conducted to determine the optimum conditions under which barium precipitation will occur. The conditions tested as a possible driving force for barium sulphate precipitation were the effect of temperature, pH and concentration. The pH and temperature ranges were pushed to both the lower and higher boundaries. The effect of temperatures as low as 15oC was studied and pH levels of 7 and upwards. The first nuclei started appearing at a pH of 7 but significant growth is promoted at higher pH levels. The X-Ray diffraction analysis confirmed the phase of the precipitates formed, and the orthorhombic crystallographic structure formed is a characteristic of barium sulphate. Keywords: Iron ore, barium sulphate, temperature, pH, concentration, morphology, and ceramic filter plate

Jan 1, 2020

A set of six design principles for rock engineering was developed by Bieniawski: design objectives; minimum uncertainty; simplicity of design components; state-of-the-art practice; optimization; and constructability. These principles translate into a ten-step design process. Although this methodology was developed for rock engineering, it is applicable to any design process, feasibility study, etc. Diligent use of such a process will ensure that a defensible design or evaluation has been carried out. Recently a ?strategic conversation? process has been identified by Ilbury and Sunter that includes the following ten steps: scope; players; rules of the game; key uncertainties; scenarios; SWOT analysis; options; decisions; measurable outcomes; and meaning of winning. A comparison of these steps with those of the design process shows a remarkable correlation. This is perhaps not surprising since engineering design is a strategic up-front issue, and front end loading in a project is critical to achieve the expected project performance. Based on the close correlation between the strategic conversation and engineering design processes, a circular design process, rather than a linear process, is proposed. The ?circle of design? better reflects the importance of review and monitoring in the two phases of successful design and implementation?defining the design, and executing the design.

Jan 1, 2009

By B. B. Tati

The project report is based on vacation work done by the writer at the Khutala Colliery during December/January 2009?2010. The purpose of this report is to explain and describe the design process in a multi-seam mine design for coal extraction at the Khutala Colliery. The multi-seam design will focus on the No. 2 and No. 4 seams as these are the seams that are currently being mined by underground methods at Khutala Colliery. The seam thicknesses and the parting thickness between the seams provide suitable conditions for designing a multi-seam mine design. An investigation into the pillar design aspects is carried out followed by a numerical modelling analysis of various scenarios. This analysis validates the first principles approach provided by multi-seam design guidelines. The design results from the analysis of the selected pillar design indicate sufficient parting for primitive conditions between seams, thus illuminating the need for superimposing in-panel pillars. This multi-seam mine design confirms the potential for the extraction of coal in a new panel. The design indicates that significant amounts of power station coal are present and suitable for extraction by room and pillar mining methods. The panel block indicates 401 625 tonnes of ROM coal. This block has a life of panel that will be mined just over a 4-month period. This 4-month period is assumed to mined at a production rate of 48 000 tonnes per month. The total capital cost required for the new panel is R48 281 127.05. Keywords Multi-seam mining, panel design, ventilation flow, superimposition.

Jan 1, 2011

By D. E. Kraushaar

Ruhrkohle in West Germany evaluates and compares the net present values of long-term periods from its coal mines by a computer assisted long-term planning system. Two main parts of this system are computer programs for the prediction of: (a) mineable reserves and their distribution in longwall panels as a function of dependent geological variables; (b) daily outputs and costs of longwall panels as a function of dependent geo- logical, layout and operational variables. This paper describes the investigations, which ended in the above computer programs: ? the collection of information from past workings and their organization in databases; ? the statistical evaluation of the databases, and ? the resulting correlation formulas. The correlation formulas vary the mean dependent variables as a function of independent variables which can be indicated with high reliability for planning projects in the long run. The system thus serves the purpose of facilitating long-term planning and excluding biased individual judgement of local planning staffs.

Jan 1, 1987

One of the key issues facing gold producers is the management of cyanide and the implementation of the International Cyanide Management Code. Many operations now use cyanide destruction processes to ensure that the material being discharged to the tailings dam is compliant with regulatory requirements and the International Cyanide Management Code. One of the problems for operations is the quantification of both cyanide species and their reaction products throughout the leach and destruction circuits. Therefore methods have been developed utilizing high performance liquid chromatography (HPLC) to measure free cyanide, cyanide complexes of metals such as copper and iron, and other species formed from reactions in leaching and cyanide destruction, including cyanate, thiocyanate, and sulphur species (eg. sulphite, thiosulphate and sulphate). This enables a complete mass balance for cyanide and sulphur in both leaching and cyanide destruction. This paper outlines the methodology which has been developed and discusses the application of these methods in the fundamental research currently being conducted on the impact of sulphide minerals in cyanidation and cyanide destruction chemistry.

Jan 1, 2009

Acid heap leaching has been extensively applied to the processing of copper oxide ores, with recovery of copper by solvent extraction and electrowinning. Heap leaching has also been extended to the treatment of secondary sulphide copper ores, where oxidative conditions are generated in aerated heaps by the bacterially-assisted oxidation of ferrous iron to ferric iron. The treatment of low-grade chalcopyrite ores has recently been the focus of research, with Mintek playing a prominent role. When processing chalcopyrite ores, more complex column test work is performed, which incorporates the prediction and simulation of temperature behaviour in pyrite-containing heaps. Mintek has devised a trade-marked column leaching apparatus (SmartColumn?) for the simulation of the natural temperature profile development in pyrite-containing heaps. As heap leach applications are becoming more complex, the parameters that need to be built into the design and operation of heap leach plants are growing in number, and the amount of data to be collected and processed, and the number of decisions to be taken daily, are also increasing. Mintek has therefore developed the HeapStar® administrative advisory software, as a guidance system to ensure that the correct protocols are implemented at various stages of the leach cycle.

Jan 1, 2012

By H. F. Fuls

Skorpion Zinc, an integrated zinc mine and refinery located in the south of Namibia, has a production capacity of 150 000 t/a of special high grade (SHG) zinc. The Skorpion zinc oxide resource will be depleted by 2017. Extensive exploration drilling for additional zinc oxides was conducted without success. With the abundant availability of zinc sulphide concentrates regionally (Black Mountain and Gamsberg in the Northern Cape or Rosh Pinah Zinc and new deposits in the Rosh Pinah region) the life of the operation can be extended by processing of zinc sulphide concentrates. The leach kinetics of zinc sulphides is vastly slower than for oxides, preventing the processing of zinc sulphides in the existing process under current conditions, and Skorpion Zinc faces the challenge to identify a suitable zinc sulphide treatment process that can be integrated with the existing plant. Various zinc sulphide processing options were identified through an extensive literature search and non-viable options were eliminated. Mass and energy balances were established for potential options and an economic evaluation (including capital and operating cost estimation) was conducted to identify a suitable process. It is concluded that ferric leaching of sphalerite in a sulphate medium under atmospheric or pressure leach conditions are suitable processes for Skorpion Zinc. These processes present advantages over all other zinc processing options when elemental sulphur is preferred as the final deportment for sulphide sulphur. Both these processes can be integrated into the existing refinery and will be able to process concentrates with a wide composition range.

Jan 1, 2013

"Pressure leaching is an option for copper recovery from chalcopyrite. Leaching takes place at high temperatures and pressures in the presence of an oxidizing agent. This work shows how the performance of pressure leaching circuits may be improved by optimizing the configuration of autoclaves and the heat removal method. The objective was achieved by creating an autoclave model that combines mass balances, energy balances, and population balances. A base case circuit consisting of a single autoclave was chosen, and this base case was compared with options for increasing capacity by adding more autoclaves to the circuit. These capacity increase options include circuits in which the additional autoclave is added in parallel, in series, and in series with thickening between the two autoclaves. The copper extraction, productivity, and cooling requirements for these options are compared. The series circuit with interstage thickening has the highest extraction and productivity; however, the cooling duty in the first autoclave is high. In addition to comparing circuit options, three options for heat removal were investigated: cooling coils, quench water, and flash recycling. The flash recycle option results in the highest copper extraction, and the quench water option the lowest copper extraction. IntroductionBackgroundCopper sulphide ores and concentrates containing chalcopyrite are typically hard to leach under atmospheric conditions. Smelting of these ores may also be a challenge, due to the inability of smelting processes to handle low-grade ores, as well as impurities such as arsenic and antimony. Alternative hydrometallurgical routes to treating chalcopyrite are therefore required. One option is pressure leaching, which is the focus of this work.In the pressure leaching process, sulphide ores and concentrates are leached in autoclaves at temperatures of 200 to 220°C and a pressure of about 3000 kPa (Schlesinger et al., 2011). An oxidizing agent, usually oxygen, is required to oxidize sulphide ions to sulphate ions. The oxidation of sulphur to sulphate is exothermic; consequently, a method of heat removal is necessary in order to maintain the temperature in the autoclave."

Jan 1, 2016

By P. G. Van Der Walt

This paper deals with the reticulation of deslimed tailings from surface to the stope face underground. Reference is made to the experience gained at the pilot plants operating on four of the gold mines administered by Gold Fields of South Africa, as well as at the full-scale system at East Driefontein which was commissioned recently. Mine-wide backfill systems are currently being installed on five other mines administered by Gold Fields. These six full-scale systems are capable of handling more than 270 000 tons of backfill solids per month. The type of reticulation system chosen, the selection of flowrates and the composition of the slurry, as well as the selection of the appropriate pipes and ancillary equipment are discussed. Both theoretical and practical aspects are taken into consideration. Other aspects covered include the design of underground storage dams and communication systems.

Jan 1, 1988

By K. B. Owusu, W. Skinner, R. K. Asamoah, J. Karageorgos, M. Zanin, C. Greet

High Pressure Grinding Rolls (HPGR) are mostly used prior to pelletising processes to increase the specific surface area (SSA) of iron ore concentrate. For expressing SSA, the Blaine index is utilised as one of the most important dimensional properties of iron ore pellet feed. ISO 21283 is a standard test method that explains the manual of Blaine air-permeability apparatus for determination of this parameter. In this standard, the disaggregation method is based on passing the minus of 1 mm samples from the 0.150 mm sieve. The disaggregation process has an effective role on the results of Blaine airpermeability test especially for regrinding HPGR products. In this research, firstly the Blaine index of samples was measured without disaggregation. Following this, disaggregation was done by passing the minus of 1 mm samples through three sieves with different dimensions, 0.150, 0.090 and 0.045 mm, separately, and the Blaine index was calculated for each one. Finally, lab-mill with rubber balls, called Jar mill, was investigated as another option. Screen analysis and image analysis with SEM were performed to evaluate the effects of the sample's disaggregation on the size distribution and agglomeration of samples. The results showed that the disaggregation method with sieve size 0.150 mm, according to ISO 21283, was not applicable for HPGR products because the particles were still stuck together. In addition, the disaggregation method with the sieve size 0.045 mm or with Jar mill offered reasonable results. Keywords: Disaggregation, Blaine index, specific surface area, regrinding HPGR, ISO 21283, Jar mill

Jan 1, 2020