Search Documents

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 F Kock, Y Sulianto, T Trott, B G. Putland, D Morgan

"Northparkes installed and commissioned a secondary crushing circuit in 2010 to debottleneck the semi-autogenous grinding (SAG) mill constrained grinding circuits. Following further improvements to the circuit (inclusion of a double deck vibrating screen), the project successfully delivered finer feed to the SAG mill. The result of this project was an increase in the milled tonnes from 5.6 Mt/a in 2012 to 6.0 Mt/a in 2013.The increased throughput rates introduced new challenges with the circuit becoming power limited. How to best optimise the available comminution power, while delivering the optimal particle size to the flotation circuit was made the key optimisation objective. To achieve this objective, Northparkes initiated a project that included engaging comminution experts. Recommendations included changes to the SAG mill including mill lifter design, increased grinding media charge levels and reduced grinding media size with aim of producing a finer grind size.Northparkes implemented these recommendations along with some other changes to SAG discharge grates and the mill discharge vibrating screen. This achieved a substantial reduction in the flotation feed particle size distribution (approximately 20 microns) by utilising increased average SAG power utilisation, in turn resulting in a two per cent increase in copper recovery.With power utilisation now maximised, the next phase of optimisation has moved to the throughput-grind-recovery relationship and the cost benefit analysis of installing additional grinding power.Citation:Sulianto, Y, Trott, T, Morgan, D, Bell, T, Kock, F and Putland, B, 2016. Optimisation of Installed Mill Power Post-installation of Secondary Crushing Circuit at Northparkes Mines, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 331–338 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

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 D Horacek, I K. Sheppard

Ok Tedi is in a period of transition. By 2008 mill feed will consist of higher levels of high-grade but metallurgically challenging skarn ores. Additionally, the Taranaki pit will start production in 2008 along with the start-up of a new pyrite concentrator. Taranaki ores are very high in sulfur. Sulfur is captured in a pyrite concentrate prior to tailing entering the Ok Tedi River. These and other complex scheduling issues will need to be resolved over a critical three-year period. With these requirements in mind, Ok Tedi contracted AMC Consultants to develop a custom version of the Life of Business Optimisation System (LOBOS¬) to assist medium-term production scheduling at Ok Tedi. LOBOS utilities enable Ok Tedi planning engineers to:import polygon-based reserve data (Blast Masters) from MineSightÆs Interactive Planning Tools (IP Tools); graphically construct block sequencing and material flow constraints; edit constraints defining standard scheduling, processing and financial relationships; define dump and stockpile attributes; and assign capacities to production systems, plants, dumps and stockpiles. LOBOS converts Ok TediÆs production scenario into a mixed integer program (MIP) and submits the resulting model to IlogÆs Cplex MIP Solver. The Solver seeks an optimal production schedule that maximises net revenue per mill hour while meeting all constraints. This paper introduces Ok Tedi and its production requirements and relates these requirements to the facilities developed for an Ok Tedi version of LOBOS.

Jan 1, 2007

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 Fabjanczyk MW, Guy RJ, Tarrant GC

The design of the reinforcement systems required to stabilise mine roadways has developed through the application of field measurement techniques. These techniques have allowed the level of security of a given opening to be determined from an assessment of the loading within the reinforcement, the extent of roof deformation and experience. The design systems developed are now used extensively in a wide variety of geological environment. Recent interpretative techniques have been developed to improve the level of confidence when designing roadway reinforcement. The techniques are particularly applicable for roadways subject to either weak materials or extreme stress levels. A second area of application is where roadway advance rates and hence the timing of reinforcement placement are critical. The newly developed techniques include both analytical and computational models. The validated computational techniques allow a more rational approach to investigation of the effect of varying reinforcement type, timing of placement and density on the level of stability achieved.

Jan 1, 1992

By Winton J. Gale

The design of the reinforcement systems required to stabilise mine roadways has developed through the application of field measurement techniques. These techniques have allowed the level of security of a given opening to be determined from an assessment of the loading within the reinforcement, the extent of roof deformation and experience. The design systems developed are now used extensively in a wide variety of geological environment. Recent interpretative techniques have been developed to improve the level of confidence when designing roadway reinforcement. The techniques are particularly applicable for roadways subject to either weak materials or extreme stress levels. A second area of application is where roadway advance rates and hence the timing of reinforcement placement are critical. The newly developed techniques include both analytical and computational models. The validated computational techniques allow a more rational approach to investigation of the effect of varying reinforcement type, timing of placement and density on the level of stability achieved.

Jan 1, 1992

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

By B P. Wrench, N R. Steward, C Wu

A gold mine in southern China has transitioned from an open pit operation to a fully underground mine. An overhand mining method is used which requires the mined stopes to be filled with backfill having an unconfined compressive strength of at least 0.35MPa. Tailings from the process plant is deposited into separate flotation and Carbon in Leach (CIL) tailings storage facilities. The tailings has a high fines content and the thickener on site could only produce a dilute flotation tailings at a solids content of about 30 wt%. This resulted in two undesirable outcomes: the density of flotation tailings in the Tailings Storage Facility (TSF) was significantly lower than planned so that the TSF could not store the life of mine tailings and that the backfill product had to include a high percentage of manufactured aggregate and increased cement content to achieve the required strength. This paper presents results of a tailings study which identified that the flotation TSF would be capable of storing the remaining life of mine tailings if an amended backfill product made up of 100% of flotation tailings could be introduced. A 100% tailings backfill with the addition of 8% cement was developed by classifying the tailings and using the coarser fraction in the backfill. This backfill allowed 60% of the flotation tailings to be placed underground and only the remaining 40% in the flotation TSF. A new thickener was designed and installed to thicken the tailings and rheological testing of the tailings, using a closed loop pilot plant, provided parameters to upgrade the underground reticulation system to deliver cemented backfill to stopes throughout the mine and to deliver thickened tailings to the TSF. The revised design was implemented and has resulted in considerable cost savings. CITATION:Steward, N R, Wrench, B P and Wu, C, 2018. Optimisation of surface tailings storage and tailings backfill – a case study, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 454–464 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

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

Sep 13, 2010

The ball level and pulp load play a key role in the operation of industrial scale ball mills. These two operating variables have a significant influence on the performance of the ball mill and any attempts to optimize these devices require manipulating either one or both of these variables. There are several models that can be employed to perform simulations to optimize the performance of industrial ball mills. However, most of these models require ball load and pulp load measurements as inputs to calibrate the model. The changes suggested by the simulator also provide operating regimes that involve moving either the ball level or pulp load and sometimes both of these variables. A ball mill model coded in JKSimMet was used in conjunction with the Sensomag to optimize the secondary ball mill at Waterval concentrator in South Africa. The results show that the perfect mixing ball mill model can be used in conjunction with the Sensomag to optimize the performance of the secondary ball mill.

Sep 1, 2012

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