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By Braaksma M, Dudley K

The Girilambone Copper Mine is located in central New South Wales and has been operational since May 1993. Girilambone is operated by the Girilambone Copper Company (GCC) which is a joint venture between Straits Resources (60 per cent) and Nord Resources (40 per cent). The mine complex consists of an open cut pit, ore crushing and stacking, heap leaching, solvent extraction (SX) and electrowinning (EW) allowing for the recovery of 14 600 tpa of LME grade A cathode copper. The leaching and SX-EW plant have been described previously by Rear, Taylor and Busby (1994) and Miller (1994).

Jan 1, 1996

By Kym Runge, Emmy Manlapig

"The objective of an industrial flotation circuit is to separate valuable minerals from non-valuable minerals utilising differences in their flotation recovery rate. Regrinding and reagent addition are particle modification processes which are employed within these circuits to maximise the flotation selectivity of one mineral with respect to the others present in the ore feeding the process.Batch test nodal analysis, or the ability to compare floatability through mathematical combination of laboratory batch flotation test results, is a powerful technique for assessing the effectiveness of particle modification processes on ore floatability in industrial flotation circuits.In this paper, the use of nodal analysis is demonstrated within an industrial flotation circuit. It is used to show that regrinding and staged reagent addition within Teck Cominco’s Red Dog zinc retreatment circuit is beneficial to pulp flotation selectivity. Optimisation of these processes could potentially improve the zinc recovery within the circuit. Lime addition, which is periodically employed to depress pyrite in the circuit, was found to depress sphalerite in preference to pyrite and to result in poorer metallurgical performance in the circuit.INTRODUCTIONThe objective of an industrial flotation circuit is to separate valuable minerals from non-valuable minerals utilising differences in their flotation recovery rate.Different minerals can only be effectively separated using flotation if they are sufficiently liberated from each other (i.e. do not co-exist within the same particles) and have a sufficient difference in surface hydrophobicity. Size reduction processes and chemical addition are utilised before and within a flotation circuit to achieve sufficient particle liberation and to maximise differences in mineral hydrophobicity. In effect, the objective of these processes is to maximise the flotation selectivity of one mineral with respect to the others present in the ore feeding the process."

Jan 1, 2004

By Drinkwater DJ, Watsford RMS, Johnson NW, Munro PD

This paper describes the use of modelling and simulation techniques in the analysis and optimisation of grinding circuits at the Lead/Zinc Concentrator of Mount Isa Mines Limited.The objectives of the studies were to: -1. improve primary grinding circuit performance by reducing the production of -0.006 mm galena fines.2. maximise the fineness of grind in the second existing grinding power and pumping capacitie The study has shown that quantitative models can operations. The models described plant operation,,and simulated the effect of process changes on plant performanc The use of automated scanning electron microscope technology tcliberation characteristics of the major minerals in the Mount isa Lead/Li ore is also described.

Jan 1, 1985

By M. D. Vilarinho, W. Xuan, F. Vasconcelos Guimarães

"This article summarises the efforts to reduce the energy consumption in grinding in a magnetite ore beneficiation plant. ArcelorMittal Kryvyi Rih ore beneficiation plant produces magnetite concentrate at 65% Fe and 8% SiO2, of which the D80 is approximately 44µm after three stages of grinding with ball mills and magnetic separation. Characterisation and laboratory scale bench tests performed on samples from the beneficiation plant indicate that the industrial flowsheet could be simplified. After the first stage of grinding and magnetic separation, a significant size fraction of the intermediate concentrate is already well liberated. Therefore, it can be directly concentrated in a second stage magnetic separation without further grinding. If a screen of 150µm size can be implemented after the first stage of grinding-magnetic separation, only the oversize of screening, representing 35% of the material, requires secondary and tertiary grinding and magnetic separation. The combined concentrate from the two fractions already meets the industrial specifications. This new route can provide the same final product, with fewer mills; it saves around 30% in energy consumption only in the grinding step without changing the main structures of the processing plant.INTRODUCTION Comminution, including crushing and grinding, is responsible for at least 40% of total energy usage in mining and mineral processing. This paper shows an example to reduce grinding cost, by changing the flow sheet design strategy in ArcelorMittal Kryvyi Rih (AMKR).AMKR is one of the most important mining and steel complexes in Krivoy Rog basin in Ukraine, which has been studied over 100 years because of mining of iron ores. The iron ores are associated to Banded Iron Formation of the Saksagan suite, with an average iron grade of around 36 percent Fe. The total thickness of the Krivoy Rog Supergroup reaches 5-6 km, whereas that of the iron-ore Saksagan Group comprises up to 1400 m. High-grade iron ore extends to more than 2,000 m below present surface"

Jan 1, 2018

By Misra A

After review of the status of research on optirnisation of retreat longwall faces it is argued that an acceptable optimisation model must be based upon realistic cost data that take into account not only the face costs, but the total panel costs. These include the cost of development, face cost, removal and installation cost. The determined optimum length of a retreat longwall face is based upon a realistic model and actual costs.Parameters that influence the optimum face length have been varied over wide limits to allow determination of the sensitivity of the model and the effect of the parameters on the optimum face length. It is concluded that for medium production rates and medium system availability, the optimum face length of a retreat longwall face under Australian conditions lies between 180 and 220 metres.

Jan 1, 1995

By B Levasseur, C Olsen, E Ghasemian Langeroudi, F Larachi, F Safizadeh, C Gagnon

Gold recovery by the cyanidation process is often accompanied by the leaching of other species. This leads to the depletion of the level of free cyanide and oxygen available for gold cyanidation. The largest proportion of the consumed cyanide is diverted by the cyanicides (eg S, Cu, Zn, Fe, etc) into non-valuable complexes.COREMÆs Extractive Metallurgy team has been conducting a research program on the improvement of the gold cyanidation process. As a part of this program, the current study aims to investigate the effect of certain operating parameters on the reactions involved during preoxidation and cyanidation. Flotation tailings and single mineral samples such as pyrite, sphalerite and chalcopyrite, along with their different combinations, were analysed to highlight the influence of ore mineralogy on the operating conditions. The effects of different kinds of oxidants (oxygen, hydrogen peroxide, ferricyanide and ozone) and chemical additives (ammonium sulfate) were investigated.Experiments conducted in alkaline solutions reveal that mineralogy is a key parameter since pyrite, chalcopyrite and sphalerite react quite differently during preoxidation and cyanidation. A significant increase in sulfur dissolution, especially on pyrite, was observed by applying preoxidation. H2O2 preoxidation showed a positive effect on increasing sulfate formation, a desirable by-product from the cyanide consumption point of view. Preoxidation of synthetic ores with ferricyanide resulted in boosting thiosulfate formation. Tests conducted on flotation tailings from a polymetallic mine revealed that a short preoxidation increases the dissolution of the ore leading to higher gold recovery and lower cyanide consumption. Regarding the formation of thiocyanate, the most undesirable by-product of the cyanidation, ferricyanide and ozone were found particularly efficient to hinder their formation. Finally, the addition of ammonium sulfate during the cyanidation process showed some promising results in terms of both gold dissolution and cyanide consumption.CITATION:Ghasemian Langeroudi, E, Levasseur, B, Safizadeh, F, Olsen, C, Larachi, F and Gagnon, C, 2013. Optimisation of operating parameters for cyanidation of sulfide-bearing gold flotation tailings in order to prevent the impact of cyanicides, in Proceedings World Gold 2013 , pp 111-120 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By Z. Saw, C. Poveda, N. L. Mai

Optimising the throughput of a mining network comprising all equipment from shovels to semi-autogenous grinding (SAG) mills is a complicated task, yet the potential benefits can be significant. Key modelling challenges stem from the sheer number of possible routes which material can travel, different equipment operating under different constraints, and the rate of SAG mills being calculated through a nonlinear function of the blended materials’ rock attributes. Nevertheless, this is an ideal onapplication area of mathematical programming, a branch of operations research. This paper describes an application of mixed integer linear programming (MILP) to optimise the mining equipment network of one of the world’s largest copper mines.The model takes as input a year-long monthly mine schedule of shovels in different pits, pushbacks, sequences and faces which produce material to be moved through a network of crushers, conveyors, silos, distributors, stockpiles and SAG mills, ultimately outputting a daily operational plan for all equipment that maximises the throughput of the network. All operational constraints and business rules are considered, including three quality attributes of copper ore. Comparing model results with actual data shows an improvement of 4.47% in network throughput tonnage. Additionally, although the model is solved using a free, open-source solver, optimal solutions are still achieved within a practical timeframe.

Sep 1, 2021

By S. K. Sharma, Sanket Bacchuwar, Prabal Kumar Agrwal, G. V. Rao

"Iron ore processing industries play a vital role in the Indian economy. At present Iron ore of around +64% Fe is used for production of Iron through blast furnace. To produce +64% Fe from low/medium grade ores, crushing & beneficiation is compulsory. But during these processes, large amount of slime is generated. It is estimated that 18% -25% of tailing is generated during the processing of iron ore. The iron ore slime is discarded because of its particle size and chemical composition not being suitable as feed to the blast furnace. Although the fines/slimes can be agglomerated to larger particle, the alumina to silica ratio which is normally more than one pose serious problem during sintering process and subsequent smelting in blast furnace. The alumina content in the slime needs to be reduced for it to be used as the feed for blast furnace. The sample assayed 61.40% Fe, 7.68% SiO2, and 1.52% Al2O3, LOI 2.63% and all other elements were found in critical limit. A series of 27 tests were conducted by using a spiral for optimisation of process parameters to produce a pellet grade concentrate from the stacked slimes obtained from Kirandul. This paper deals with the optimisation studies of various parameters of a spiral concentrator and helps to utilize the fines from tailing ponds.INTRODUCTIONReserves of high-grade iron ores are diminishing all over the world at an alarming rate. As a result, alternative ways of augmenting and conserving precious and non - renewable natural resources are being seriously considered. One such attractive option is the recovery of metal values from slimes / tailings. Iron ore tailings containing around 48 - 60% Fe are generated from the iron ore washing plants and are disposed into tailing ponds without any further utility. These tailings in the form of slimes are not suitable in iron and steel making due to the presence of higher amount of gangue constituents. Several beneficiation techniques have been tried from time to time to reduce the gangue so that the beneficiated products could be effectively used."

Jan 1, 2016

By J. A. Botin, K. H. Singh

INTRODUCTION The Alfredo underground mine is one of the three mining operations of Rio Tinto Minera, S.A., all near the community of Rio Tinto in Huelva province of southwest Spain. The other two operations are Corta Atalaya open pit (Pyrite ore) and Cerro Colorado open pit (Copper ore 1. Underground mining of Alfredo Mine for pyrites started in the nineteenth century and continued without interruption until 1979 when it was discontinued due to economical reasons. In 1972, a new project was started for the mining of copper bearing ores on the footwall of the pyrite deposit. This copper ore is locally named “Cloritas” and it forms a nearly vertical and tabular shaped orebody which forms an envelope at the footwall of the San Dionisio pyrite orebody. Mining of the Cloritas commenced in 1975 using cut and fill stoping. In 1977, the original project was completely revised which led to the introduction of a bulk mining system to achieve a planned production of 3,000 tonnes of ore per day. This paper covers the economic and planning aspects of the mine and discusses the evolution of the mining method for the Cloritas orebody. A GENERAL DESCRIPTION OF THE CLORITAS OREBODY The “Masa San Dionisio” (figure 1 ) is a massive sulphide orebody located in the town of Rio Tinto, about 90 kilometres west of Sevilla, Spain. Extensive vein-type sulphide mineralization exists in the volcanic host rocks of the Masa San Dionisio. Within this area of stockwork veins, there is a higher concentration of copper in the form of chalcopyrite in a zone adjacent to and paralleling the contact with the massive sulphides. This copper envelope is locally referred to as the "Cloritas Orebody” and the ore within it is called "Cloritas”. The main zone of the Cloritas orebody, (figure 11, is tabular and nearly vertical having a length of 375-400 m east-west, a thickness of 20-50 m north-south, and a vertical extension of 500 m from the surface. In addition to this area, there is a second less important low grade area called zone 2A located to the north of the main zone below 26 Level which, in some sections, is not fully separated from the main zone. MINING METHODS Mining in Rio Tinto area can be traced back to the Phoenecians and the Romans. Modern day mining started with the demand for sulphuric acid during the last half of the nineteenth century. By 1971, more than 33 million tonnes of pyrite had been mined out of Alfredo Mine, (San Dionisio Orebody), using labor intensive drift and fill methods and also by open pit method. During the last decade, underground mining of the pyrite had become uneconomic due to lower prices and a slackening demand for pyrite, and the underground mining operations for pyrite were finally stopped in 1979. In 1971, a major change in the long range mining plans for Masa San Dionisio was established. Between 1971 and 1975, underground mining of pyrites was progressively reduced while open pit extraction was increased in

Jan 1, 1981

By J Fut=

The rotary drilling of rocks is widespread technology in mining and in tunnelling. The paper describes the results of research by aim to find relationships between optimal drilling and sound signal of drilling machine. The optimisation criterion has been formulated as the minimisation of specific energy with maximal effective speed of drilling in process of the rock disintegration. The research in this field has been made on experimental drilling stand. Our monitoring system has measured following variables: the pressing force of core drill, the revolutions of drilling head, the length of drilling, the time, the power, and the acoustic signals. For measurement of acoustic signals have been used very sensitively sound meter MEDIATOR 2238. From this measurements have been processed acoustic spectrums. Then we are found relationships between optimal control of drilling and composition of acoustic spectrum. On base of experimental results, we can formulate the optimal control following: æTo achieve the minimum difference between the equivalent sound level at the load and idle modes only at the representative frequency (f = 1000 Hz)Æ. The representative frequency dependences on type of the rock New method based on these relationships enables the optimal control of drilling machine by using measured acoustic signals only.

Jan 1, 2003

By J. Griesel, P. H. F. Bouwer, K. Bisaka

Transalloys produces 160 000 tons of SiMn per annum from five AC submerged-arc furnaces. The lime-ore melt from two AC open-arc furnaces is contacted with SiMn to produce 50 000 tons of medium carbon FeMn per annum. Metallics in the slag are reclaimed in a jigging plant, which has been in operation since 1999. The flow of materials between the jigging plant and furnaces has been optimised. The amount of recycle slag in the feed of the SiMn furnaces has been decreased by more than 30% which still provides satisfactory stability of the furnace operation. The jigging plant has absorbed all the additional slag generated, to recover SiMn metal. A review of the raw materials used and operating parameters for SiMn production was also undertaken. Slag to metal ratios were decreased by more than 30% by reducing the slag recycle, removing magnesite from the recipe and introducing high-grade ore. Electrode penetration was improved by reducing coke levels and by utilising more reactive coal. The pitch circle diameter (PCD) was reduced to get a better operating stability on furnace no 5 using the MINSTRAL Control System. Energy consumption on the furnaces improved by more than 15% while SiMn production increased as a result by more than 18%.

Jan 1, 2004

By I. Demers

Doyon mine, Quebec, Canada evaluates the possibility of desulphurizing their tailings to produce material to be used in as cover over the acid generating tailings disposal site. Laboratory tests were conducted and showed that a single-layer cover made of desulphurized tailings placed over reactive tailings can prevent the generation of acid mine drainage. Modelling with the software Vadose/W was undertaken to optimize the design parameters, namely cover thickness, water table position, and residual sulphide content of desulphurized tailings following laboratory work. It was found that the optimal configuration includes a 1 m desulphurized tailings cover containing 0.8%S, with the water table kept at the interface between the reactive tailings and the cover. To maintain the cover performance to reduce oxygen flux towards acid-generating tailings even in the event of a long dry period, a sand layer is suggested to protect the cover from excessive evaporation. It is possible to reach the environmental objective chosen for Doyon mine (10 mol/m2/year) with the protective sand layer, even in extremely dry conditions.

Jan 1, 2008

By A R. Bonani, M P. Pe±a

The optimisation of slope angles, that is to say, increasing them, compared to the original design is always the subject of analysis of stability for geotechnical engineers and economic analysis for planning engineers. Questions to be answered include: What is the acceptable risk? What factor of safety? What probability of failure or reliability? The Rosario Open Pit has a history of failure, poor quality rocks and high pore pressures. It was decided to deepen the pit below the programmed design to test the practicality of different design approaches, which if successful, could be used to improve the economics of mining at Collahuasi. The revised slope design has factors of safety in the range of 1.0 to 1.2, probabilities of failure of ten to 38 per cent, and increased slope angles of between 3¦ and 5¦ with respect to the original design. The pit deepening with these design slopes will deliver an extra two million tonnes of ore which will have a significant positive impact on mining economics.

Jan 1, 2007

By S. Ranganathan

Production of ferro-chromium in the submerged-arc furnace is highly energy-intensive. Several process rates have been attempted to reduce the specific energy consumption in the production of this alloy. An exhaustive understanding of the thermo-chemical features of the production process can be used to reduce the specific power consumption significantly. The knowledge can be used to reduce the energy drain through undesired reactions in the process can be minimised. Plant trials have shown that specific energy consumption up to 100kWh/tonne of hot metal can be obtained by adopting this strategy. The process requires modification in operation practice and does not require any capital investment unlike other processes adopted for reducing power consumption. The principle of thermo-chemistry involved in this process development and results of plant trials are discussed in detail.

Jan 1, 2007

By Hagen G, Uhlig H

In the Federal Republic of Germany the thermal environment in underground coal mines has deteriorated progressively over the past years. The number and capacity of refrigeration plants have increased correspondingly. In recent years high capacity spray coolers have been used in addition to conventional cooling coils.

Jan 1, 1988

Optimisation of the Advanced Automation System at Inmet Mining's PyhSsalmi Mine, Finland

Sep 13, 2010

By J Bower

The JORC Code does not specify a definition for ælevel of confidenceÆ. The common approach in the mining industry is to use a certain drill spacing or number of data samples to classify resources by the confidence categories. Another approach includes application of the quantitative geostatistical criteria for non-subjective and consistent definition of the resource categories. This approach, which has become increasingly popular in the mining industry, provides a quantitative measure of the resource estimation uncertainty. In this study, uncertainties associated with the different drilling grids at the Weipa bauxite deposit (Queensland, Australia) have been modelled using conditional simulation methodology. Criteria used at Weipa for defining resource categories are as follows: mineralisation can be classified as a Measured Resource if tonnage and grade of its volume representing three months of production can be estimated with a precision error not exceeding ¦10 per cent (at 95 per cent probability, ie two standard deviations); and mineralisation can be classified as an Indicated Resource if grade and tonnage of its volume representing annual production can be estimated with a precision error not exceeding ¦10 per cent (at 95 per cent probability). Several different grids have been created by subsampling the drill hole database. The studied variables include Al2O3, SiO2, thickness of the bauxite seam and location of the footwall contact. The variables have been modelled by sequential Gaussian simulation. Based on this study the following drill spacings have been recommended: Measured Resources û 150 m by 150 m, Indicated Resources û 300 m by 300 m.

Jan 1, 2009

By B Cook, M Karakus, J Light, E K. Chanda

The aim of this research is to develop a model through which the appropriate paste recipe parameters, for example solid content and cement content, can be derived given the stope specific requirements such as stope dimensions, required strength and curing time. A large number of samples and uniaxial compressive strength (UCS) tests were conducted at the Angas Zinc Mine (AZM) paste plant and testing facility. Samples were varied in tailings solids content, cement content and sand content. The results of this testing were modelled to determine the UCS relationship to each of the parameters. The output of this modelling was an easy-to-use program that outputs the design mixture given the stope dimensions and required stope backfill-curing time. This allowed flexibility within the mining schedule for curing times of cemented paste backfill stopes. The optimised schedule using the stope specific mix design, when compared to the standard 14-day curing time used previously, resulted in a smoother weekly mine production schedule. For the modelled 12-week period (one quarter) the 14-day curing time would result in an eight-day shutdown of the mill. The cost for each shutdown is in the order of $125 000/d. Due to the unique nature of each mine’s tailings, this exact outcome would not be directly applicable to another paste fill site. However, the process in this report could be followed for another site to achieve a system similar to the output of this project. This would ultimately lead to a more optimised backfill and stoping cycle and a more optimised mine.CITATION:Chanda, E K, Karakus, M, Light, J, Cook, B and Harris, A, 2014. Optimisation of the paste fill cycle at Angas Zinc Mine, Strathalbyn, South Australia, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 259–264 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

Optimisation of the Performance of Flotation Circuits Using a Genetic Algorithm Oriented by Process-Based Rules

Sep 13, 2010

By C M. Ramotsabi, F Bornman, W Erasmus

Spiral concentrators are gravity based separation devices used for the preparation of coal, iron and heavy mineral ores (Wills, 1992). Gravity concentration remains the main separation method for fine iron ore and is used extensively for treating many minerals. Gravity concentration methods separate minerals of different specific gravity by their relative movement in response to gravity and one or more other forces, the latter often being the resistance to motion offered by a viscous fluid such as water or air (Wills, 2006).The main operating variables for a spiral concentrator are the feed rate, the slurry concentration, the wash water addition and the position of the splitters used to separate the tailings, middlings and concentrate streams (Sadeghi, Bazin and Renaud, 2014). Wash water is an important control variable used to wash away entrapped light minerals from the concentrate stream (Burt, 1984; Bouchard, 2001). Sadeghi et al concluded the spiral operation to be sensitive to the wash water flow addition and resultant economic trade-off. Increasing the wash water flow rate increases the concentrate grade at the expense of the iron recovery.The paper addresses the effects of dripping wash water onto the concentrate stream of a spiral processing fine iron ore. The dripping wash water system offers a low flow alternative and could possibly provide a good balance between grade and recovery of valuable minerals without using excessive water. The Multotec SC20HC/7 WW spiral was used for the experimental test work. The spiral was tested at varying solids concentrations; the wash water rate and the splitter positions were kept constant.The use of wash water at 30 per cent, 35 per cent and 40 per cent solid concentration improved the Fe grade in the concentrate by about 1.5 per cent, six per cent and eight per cent respectively. The additional SiO2 rejected by the wash water is approximately three per cent, nine per cent, and 11 per cent at 30 per cent, 35 per cent and 40 per cent solid concentration respectively. Addition of wash water is shown to have minimum effect on grade at 30 per cent solid concentration, and Fe upgrades obtained with and without wash water were found to be similar. Notable upgrades were observed at 35 and 40 per cent solid concentration with wash water. The results obtained suggest that the additional SiO2 rejection by wash water increases with an increase in solid concentration. This suggests that the benefit of using the drip system is notable at higher solid concentration. It is believed that at lower solid concentration SiO2 particles are reasonably free to move around and are possibly less entrapped than at high solid concentration.The Fe grade/recovery and solid concentration relationship with wash water showed a non-linear correlation. The Fe grade increases with an increase in solid concentration up to a somewhat optimum point and then drops as solid concentration increases further. The recovery of Fe on the other hand follows a converse relationship with solid concentration, the recovery decreases from 30 per cent to 35 per cent solid concentration and then increases again at 40 per cent solid concentration. The increase in recovery at higher solid concentration could possibly be explained by the fact that the mass yield to concentrate also increases with solid concentration. Since recovery is a function of both grade and mass yield, the higher recovery at 40 per cent solid concentration is likely due to the high mass yield obtainable at 40 per cent solid concentration.The drip system could therefore possibly offer a safe guarding against grade and solid concentration variations without using excessive water, and possibly ensure that a product within specification is maintained.CITATION:Ramotsabi, C M, Erasmus, W and Bornman, F, 2015. Optimisation of wash water on an iron ore spiral to improve grade, in Proceedings Iron Ore 2015, pp 175–184 (The Australasian Institute of Mining and Metallurgy: Me

Jul 13, 2015