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By G du Toit, P L. Gillot, M Swanepoel

Randgold Resources’ Loulo-Gounkoto Complex is one of the largest gold mining operations in Mali, West Africa. The mine, which is situated on the western border with Senegal, treats more than 4 Mt/a of ore through a conventional three-stage crushing and ball milling circuit followed by a carbon-in-leach plant. Ore supply is achieved by a combination of open pit and underground mining, with the majority of the ore supply sourced from the Gara and Yalea underground mines situated approximately 2 km and 6 km respectively from the gold plant.In 2011, Randgold Resources commissioned Golder Associates to perform an evaluation study for providing structural backfill to both the Gara and Yalea underground mines using the current gold plant tailings. The test work performed as part of the evaluation study concluded the need for a tailings cycloning and dewatering facility capable of producing slurry at a solids concentration of 65 per cent with a particle size distribution that had minimum 25 per cent passing 20 µm and less than 5 per cent of -5 µm material. Tailings desliming was required to ensure a consistent feed quality to the paste preparation plants and to ensure that the system would be capable of handling variation in the feed composition from the gold plant.WorleyParsons, previously TWP Projects, was contracted by Randgold Resources to complete the engineering design for the slurry preparation plant, whilst Golder Associates was retained to complete the engineering design of the two paste preparation plants.This paper discusses the test work, design considerations and engineering design of the slurry preparation plant, termed the ‘intermediate plant’, as well as all ancillary facilities required for the delivery of the deslimed slurry to the paste preparation plants located at the two underground mines.CITATION:Swanepoel, M, du Toit, G and Gillot, P L, 2014. Case study – engineering design for a slurry preparation plant for a paste application in West Africa, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 477–486 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

The Teutonic Bore deposit lies in a major greenstone belt within the Archaean Yilgarn craton of Western Australia. It occurs towards the base of a suite of tholeiitic basalts, some 100 m above a felsic volcanic pile of rhyolitic composition. Sediments comprise a minor, albeit important, percentage of the mine stratigraphy.The deposit consists of a single, steeply dipping lens of massive sulphides totalling 1.4 million tonnes grading 4.16 per cent Cu, 16.4 per cent Zn, 1.22 per cent Pb and 203 g/t Ag. In addition, cross cutting "stringer" mineralisation in the footwall contributes a further 0.75 million tonnes at 2.38 per cent Cu, 1.92 per cent Zn and 52 g/t Ag to the mining reserves. Principal primary sulphides in massive and disseminated mineralisation are sphalerite, pyrite, chalcopyrite and galena, while minor amounts of arsenopyrite and argentiferous tetrahedrite are also present. Weathering has affected the primary mineralogy down to at least 90 m, resulting in a strongly leached oxide profile to about 85 m, overlying a thin, erratic zone of chalcocite supergene enrichment.Lateral metal zoning is marked both in massive sulphide and stringer mineralisation, with Cu/Zn ratios decreasing away from the centre of the orebody. Alteration is well developed in association with footwall stringer mineralisation, changing from a central Fe-Mg carbonate, chlorite and silica alteration zone below the massive sulphides to an outer halo of sericitisation. The spatial distribution of these minerals is patchy. Geochemically, the footwall alteration zone is characterised by central Cao, Na20 and Sr depletion (with less regular enrichment of SiO2 , Fe and MgO), surrounded by a peripheral zone of K20 enrichment.Evidence supports a classic volcanic-exhalative origin, involving submarine discharge on the flanks of a rhyolitic dome and accumulation of the massive sulphides in a local palaeotopographic depression. The sulphides overlie their feeder zone as shown by alteration and stringer mineralisation. Tholeiitic wallrocks are thought to be genetically unrelated to the mineralising solutions, although their presence as a cap over the felsic volcanics may have permitted the build up of a circulating metalliferous brine system prior to sulphide deposition.

Jan 1, 1984

Attention is drawn to the largest eruptions with dense rock equivalent (D.R.E.) volumes approx-imating 50 km3. These have been strongly clustered during the last 40 000 years, with 11 recognised in the period 35 000 to 22 000 years ago, and 10 in about the last 9000 years, but none in the period 22 000 to 9000 years ago. With the latest such eruption, at Tambora, only 175 years ago, it is likely that we remain in a period of increased activity. During the last 100 000 years, roughly the period during which modern Homo sapiens has existed as a species, there has only been one eruption an order of magnitude larger (>_ 500 km3 D.R.E.). This took place at Lake Toba, Sumatra, about 75 000 years ago, and is estimated to have deposited ash to a thickness of a centimetre or so in parts of Ethiopia, Somaliland, Saudi Arabia and Iran, and to a thickness of a few millimetres in the eastern Mediterranean. The average interval between global eruptions of this magnitude is probably about 40 000 to 50 000 years. Insufficient work on the Quaternary history of volcanoes has been carried out to suggest where the next such eruption might occur. At Yellowstone, which has been well studied, the average interval between very large explosive eruptions is about 700 000 ¦ 120 000 years.As a prerequisite for magnitude-frequency analysis of eruptions, and hazard and risk impact assessments, it is important to classify past eruptions in a consistent manner. In an attempt to find a universally applicable method to estimate order of magnitude D.R.E. volumes from the distribution of airfall tephra, an empirical correction has been derived for the elongation of tephra dispersal fans due to wind. Dispersion distances for various thicknesses of tephra, recalculated for windless conditions, have been obtained from a study of about 40 tephra isopach maps for which D.R.E. volumes have been calculated, and a range of distances derived, equivalent to D.R.E. volumes from 106 m3

Jan 1, 1990

By G Tsatouhas, R Newell, K Y. Wong

A sulfide ion selective electrode was used to control sulfidiser addition in the flotation testing of a synthetic and a naturally occurring ore containing malachite in order to compare the behaviour of sodium sulfide and calcium polysulfide. Both slug and controlled potential sulfidisation methods of addition were used. Under defined optimum CPS conditions, calcium polysulfide was a more effective sulfidising agent than sodium sulfide on the basis of both capital and operating costs. Surface and solution studies were used to compare the mechanism of each sulfidiser. For calcium polysulfide, XPS showed that polysulfide ions are adsorbed, whereas for sodium sulfide it is the hydrosulfide ions that are adsorbed on the mineral surface. For calcium polysulfide, the CuSx layer was believed to form on the malachite surface, which promoted the flotation of the malachite, by xanthate. Studies of surface morphology of the sulfidised malachite showed that a fairly uniform sulfidised layer is formed with both reagents. The use of sodium sulfide resulted in a thinner initial sulfidised layer compared to polysulfide at equivalent initial sulfur concentration.

Jan 1, 2005

By A Dal-Corobbo

Traditional acid and metalliferous drainage (AMD) wastewater treatment focuses on the iron co-precipitation/adsorption process as this is a very effective method of reducing heavy metals, including copper, lead, zinc, arsenic, etc, to parts per billion (ppb) levels in AMD wastewater streams. One of the drawbacks to this method however is the large volume of sludge produced and subsequent disposal costs for the mine site concerned.A proprietary metals treatment system called high density sludge (denoted HDS) which significantly reduces the quantity of waste generated is described in this paper. This process generates a dewatered sludge that contains in the order of 50 to 55 per cent solids. This sludge is denser than conventionally dewatered sludge, which typically contains only 20 to 30 per cent solids. This advanced process has a wide range of applications and can be easily retrofitted into existing AMD treatment systems. It has been successfully installed at several mine sites to treat AMD wastewaters. This paper discusses the high density sludge process, its benefits and provides data and operating results from pilot test installations.CITATION:Dal-Corobbo, A, 2013. High-density sludge process for treating metal bearing acid and metalliferous drainage wastewaters, in Proceedings Water in Mining 2013 , pp 217-220 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

The Douglas mineral sands deposit is located 75 km north of Hamilton in western Victoria, Australia. The deposit was discovered in the late 1990s by Basin Minerals Limited and purchased by Iluka Resources Ltd in 2002. Prestrip mining commenced in December 2004, with ore mining commencing in August 2005. Douglas is the first of several projects being implemented by Iluka Resources Ltd in the Murray Basin, with all heavy mineral concentrate planned to be treated at the companyÆs mineral separation plant in Hamilton. The mineralisation at Douglas lies along the south-eastern margin of the Murray Basin. It is hosted by marine Pliocene Loxton-Parilla Sands and consists of three discrete deposits; Bondi Main, Bondi West and Bondi East. The mineralised sediments dip at a low angle to the west and lie unconformably beneath a veneer of top soil (50 - 100 mm) and subsoil (up to 600 mm) and between three to 12 m of overburden. The barren overburden, characterised by high clay content (<53 ¦m material) generally provides a stark contrast to the sandy ore, which is between 1 m (parts of Bondi East) and 15 m (Bondi Main) thick. Both overburden and ore are removed utilising standard dry mining techniques using both scraper-bulldozer and truck-excavator combinations. All ore is delivered to run-of-mine (ROM) stockpiles before being reclaimed by front end loader into the processing plant. Grade control at Douglas is based around visual dig control aided by high precision survey, which enables accurate reference to the resource drilling and block model data. Vertical excavator faces, inclined scraper faces plus regular test pits and trenches provide opportunities for in pit mapping and sampling. This data is used to refine not only the resource estimate for Douglas but also the drilling, interpretation and resource modelling methods used for similar deposits across the basin. The foundations of the geological reconciliation process at Douglas are the in pit surveys undertaken each time the mining fleet has completed building a ROM stockpile. These surveys and the subsequent interrogation of resource models enable robust estimations of the heavy mineral content of the stockpiles, which are subsequently used to predict the daily input grades into the plant during their reclamation. Reconciliation of this daily data is then undertaken at the end of each month against metallurgical production data and plant balances. Learnings and results from the reconciliation process are used to further refine subsequent resource models and modelling methods.

Jan 1, 2009

By M Spleit

A long-term production schedule for the LabMag iron ore deposit in northern Quebec, Canada is derived using stochastic integer programming. The optimisation formulation maximises the schedule’s net present value while simultaneously managing the risk of deviations in production tonnages and qualities by considering stochastic simulations of the orebody instead of a single deterministic model. The formulation also smooths and minimises haul truck requirements and ensures that as mining progresses, space is created within the mined out pit for the return of waste material.CITATION:Spleit, M, 2014. Production schedule optimisation – meeting targets by hedging against geological risk while addressing environmental and equipment concerns, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 397–404 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By M A. Rhamdhani, K Avarmaa, A Putera, J Chen, G Brooks

Silicon-based photovoltaic cells are currently the most dominant types applied to convert solar energy to electricity. In solar-grade silicon (SG-Si), the control of boron (B) and phosphorus (P) is very critical. Generally, these elements are difficult to remove into acceptable sub-ppm levels from lower grade Si. This study investigates slag treatment as a process for Si refining, focusing on the behaviour of P and B. The experiments were conducted by equilibrating Si-B-P alloys with 43 wt per cent SiO2 – 47 wt per cent CaO – 10 wt per cent Al2O3 slag under inert atmosphere at 1500°C. Different crucible materials were used (Al2O3, SiO2 and MgO), which determined the primary saturation phase for the system and modified the final slag composition accordingly. The samples were equilibrated and analysed with microanalysis techniques (EPMA (electron probe microanalysis) and LA-ICP-MS (laser ablation inductively coupled plasma mass spectrometry)). New standards were prepared and applied in the LA-ICP-MS analyses. Experimental results show that silver is essentially fully dissolved in silicon, as well as most phosphorus with LP ( = wt per cent P in slag/wt per cent P in Si) between 0.1–0.3. The distribution coefficient of boron was measured to be between 1 and 2. The crucible material used had an influence on the phase compositions and slight influence on the distribution coefficients of boron and phosphorus. The experimental results provided new accurate data when considering applicability of slag treatment as a technique for Si refining and Si PV cell recycling.

Jun 19, 2024

By E C. Holloway

"The financial valuation of mining assets has a number of fundamental differences to the valuation of many other asset groupings. One such difference is that typically, the greater proportion of both the cost and revenue sides of the cash flow are inherently governed by systemic market forces.As a result, the range of outcomes for the valuation of each mining asset is more unpredictable, with higher levels of risk and uncertainty. In contrast, the management function of many other non-mining businesses or industry groups can control at least a portion of the cash flow by maintaining the margin or controlling the sales price. The management teams of mining businesses and operations typically have limited options, and are thus price-takers for both costs and revenues. As mining is an inherently cyclical industry, large, long-life operations can exist through several price cycles, making valuation both subjective and challenging.The outcomes of both static- and stochastic-based valuations, as well as the internal probabilities of the associated cash flow statements, are significantly impacted when the correlations between the appropriate parameters are quantified, modelled and incorporated into the valuation framework. The appropriate parameters for correlation modelling primarily include those that are governed by systemic market forces. Applying this process results in a different risk profile for the asset by allowing the risks resulting from systemic market movements to be more accurately quantified and then objectively minimised. This process also reduces the level of residual uncertainty in the valuation outcome.This paper presents a valuation methodology to pragmatically incorporate this level of valuation sophistication and statistically quantify the associated increase in the level of confidence of the valuation outcomes reported.CITATION:Holloway, E C, 2016. Correlated valuation methodology, in Proceedings Project Evaluation 2016, pp 181–200 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By K G. Harding, E L. Haggard, C M. Sheridan

Mining is a water intensive sector, contributing to declining water resource quality around the world. Improvements to current mine water management practices can help reduce the amount of water utilised by the mining sector and the impact on water resources. Creating a water account can help companies understand water use in their operations and identify areas where water reduction is possible. Water accounting methods include the Water Footprint Network method and the WaterMiner tool.The Water Footprint Network method is used to determine the amount of direct and indirect water used by an organisation either in a process or to manufacture a product. The total water footprint is divided into the blue, green and grey water footprints. Blue water is the amount of water abstracted from surface and ground water resources. Green water is the amount of rain water utilised. Grey water is the amount of freshwater required to assimilate the waste stream to the natural resource and still conform to ambient water quality standards.The WaterMiner tool is a program available via the internet, which can be used to model water use of mineral processing sites. The program reports the amount of raw, worked and treated water used by the process and recommends flows that can be added to reduce the amount of fresh water required by the process.To gain information on mine water management practices in South Africa, a water account for a platinum mine was developed using the Water Footprint Network method and WaterMiner tool. Based on the results, suggestions to reduce the amount of fresh water utilised are made.CITATION:Haggard, E L, Sheridan, C M and Harding, K G, 2013. Water footprint for a South African platinum mine, in Proceedings Water in Mining 2013 , pp 15-20 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

By F H. Henrfquez, A G. Oroz

Design and operating conditions of large size mechanical flotation cells were evaluated by comparing them with actual operating conditions in a plant. The objective was to determine the time scale-up factor, typically based on empirical rules. Experiments were conducted on the rougher flotation circuit at Minera Escondida Ltd. The circuit consisted of self-aerated mechanical cells of 160 m3, arranged in six parallel banks with nine cells each. The rougher circuit flotation kinetics were evaluated from direct sampling and local mass balances around each cell of the bank. Adjusted overall mass balances were also developed. This information was used to fit different kinetic flotation models, and it was found that the rectangular distribution function was the most appropriate to describe the distributed rate constant for industrial operation. Then, a rougher flotation simulator was developed to describe the actual operation in terms of the operating variables (mass flowrate, solid percentage, feed grade) and the actual volumetric flowrate entering to each cell. In this study feed pulp samples were taken in parallel from the rougher circuit and were simultaneously floated in the laboratory. The kinetic behaviour was then modelled at a laboratory batch scale in order to determine the time scale-up factor between laboratory batch flotation data and industrial size flotation. The time scale-up factor observed for large sized cells, 160 m3, was found to be reasonably similar to those previously determined for self-aerated mechanical cells, but of lower size, operating at similar recoveries. In addition, the relative effect of mixing, between laboratory batch and an industrial flotation bank, was quantified by the parameter, separating the impact that kinetic and mixing changes have on the time scale-up factor. In general, the rougher flotation operation was found to reach the predicted metallurgical target, and the optimal separability criterion was also respected. The diagnostic generates information about the internal state of the process and helps to identify potential improvements for design, operation and control of the circuit.

Jan 1, 2005

In 1985 BMR assessed Australia's economic demonstrated resources of mineral sands as follows: rutile 8.0 Mt, ilmenite 41.4 Mt, zircon 11.5 Mt, and monazite 229 000 tonnes. These resources are located along the east coast of Australia, in the southwest corner of Western Australia, and in the Eneabba region, north of Perth, WA. However, over 40% of Australia's rutile and zircon resources, particularly those on the east coast, are presently not available for mining because of environmental considerations. As well, east coast ilmenite has been considered unsuitable for use by sulphate route pigment producers because of its relatively high chrome content. World demonstrated resources of rutile and zircon are estimated by BMR as 29 Mt and 38 Mt respectively. Demonstrated resources of ilmenite amount to about 1 billion tonnes. These occur predominantly in hard rock deposits in Norway, Canada, USSR and China, as well as alluvial deposits in Australia, South Africa and India. Quantitative data for world monazite resources is sparse. However, world production of rare earth oxides is now mainly from bastnaesite, a hard rock resource. Australia is presently the world's principal producer of rutile and zircon, and also provides a major share of the world's ilmenite. With the expected increase in Australia's rate of output of these commodities, combined with restrictions on mining along parts of the east coast, its position could diminish substantially over the next 25 years.

Jan 1, 1986

Sublevel open stoping (S.L.O.S.) was introduced at Rosebery in 1972 in the wider ore blocks below 14 level, largely replacing cut and fill methods which had been in use up till that time. During 1973/74 the method produced 10% of primary stoping tonnes, and by 1981/82 it increased to 65%. Because of the increasing reliance placed on sublevel open stoping for primary stoping production, every effort was made in both planning and operating to: - maximise ore recovery and reduce dilution - increase productivity - reduce costs.

Jan 1, 1982

By Reichman J. P

Extensive use is made of a small scale computer (IBM 1130), in the calculations of ore reserves for Western Mining Corporation's nickel mine at Kambalda, W.A. All surface drilling data such as assays, density and location information, are stored on disk files at the Company's Technical Computing Centre at Perth, W.A. The computer is used to search through the data, select significant ore intersections, calculate the area of influence of each intersection and perform the necessary calculations to arrive at an estimate of the ore reserve. Comparing the computer-assisted ore reserve calculation with an entirely manual operation, it is estimated that a saving of 40 percent in time is achieved together with

Jan 1, 1973

By C Muñoz, N Ortega, J Vergara, D Avilés

Mass blasts have been the subject of extensive studies because of their potential benefits in mining operations. However, it has not been possible to advance due to the paradigm that states that the blasts near the designed final wall or to critical structures have to be low tonnage. Another reason is the absence of specialised simulation programs to more precisely determine the influence of the energy released by blasting to the rock mass. There are some studies that demonstrate the possibility of massive blasting that maintain the design of the final slope (Farias and Santander, 2014). It should be noted that these studies develop the concept that the size of the blast should not increase the peak vibration in the rock mass.The study was carried out in partnership with Los Colorados Mining Company. Analysis, reconciliation and wall control for two types of blasts were performed:blasting production: approximately 850 000 t, with special wall control considerationsblasting of approximately 950 000 t over the entire width of the bench.Monitoring systems were used in the near field and far field using examination of holes in order to obtain information to assess the impact on the rock mass and then to compare these to conventional blasting combining production and contour blasting.The results show that for a large blasted volume, production and/or those generated from toe to crest on the bench did not produce more impact in the rock mass than the conventional blasting.The conclusions drawn from this study are:an improvement in the excavation rate of 7500 t/h that is higher than the maximum rates normally obtainedthe peak vibrations of the P wave generated by both blasts did not exceed the peak particle velocity criteria (Vergara and Muñoz, 2014)holes examination indicating no difference in their entire length, in pre- and post-blast reviewover US$2 000 000 savings generated in the year (estimated by the client) (Avilés, 2014).CITATION:Vergara, J, Muñoz, C, Ortega, N and Avilés, D, 2015. Benefits related to applications of a mass blast in open cut mining, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 625–632 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By G Lumley

Gold miners have been underperforming the price of gold for a number of years. So why has the æmarketÆ undervalued our gold miners relative to what they have been mining? The low valuation ascribed to gold mining companies is largely attributable to their profit performance. In the period since 2006 the price of gold in A$ has risen about $900/oz, but the real cash margin (operating cash costs + cost of finding and/or acquiring deposits + the cost of bringing them into production) has only risen about A$100. Rising operating costs are a major part of this reduction in real margin. Energy and labour costs have risen more than the inflation rates across a number of countries as have taxes, royalties, etc. There are however, more significant contributors to the rising cost base. Firstly, in 2006 the average gold grade was 2.0 g/t; today it is 1.3 g/t. Secondly, strip ratios (for surface mines) have increased by 19 per cent during this time. Open cut mining equipment unit output across the world has dropped by 26 per cent during 2006 - 2012. Finally, costs have risen by 55 per cent during this time period. Consequently, there is now 3.8 times more mining cost required to produce an ounce of gold from surface mines compared with 2006. The decline in gold grade looks to be a permanent trend as does increasing strip ratio if output is to be maintained. Costs will continue to rise although cost control will be necessary. Reversing the decline in equipment performance and closing the gap between median and best practice provides the hope for a sustainable reversal of cost increases. The current gap between best practice and median performance has grown to 112 per cent for mining trucks, 168 per cent for excavators and 169 per cent for front end loaders. Closing this gap is the potential for a more profitable future. For most gold mining companies improving mining efficiency is not an option; it is an imperative.CITATION:Lumley, G, 2013. How many gold mines will inefficiency kill?, in Proceedings World Gold 2013 , pp 267-276 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By X Zhang, Y Zhu, P E. Ngoepe, P P. Mkhonto, W Xiong, L Han, L Lu

The triazine molecules have not been given much attention and they are potential collectors that may positively impact the mineral processing industry. The adsorption of triazine as collectors such as sodium 2,6-dithio-4-butylamino-1,3,5-triazine (SDTBAT) on pyrite form a basis in design of flotation collectors that may be applicable in a wide range of sulfide minerals. The SDTBAT collector was compared with the sodium normal butyl xanthate (SNBX) and sodium normal butyl dithiocarbamate (SNBDTC). All collectors were adsorbed in their neutral form containing sodium on the pyrite (100) surface. The computational calculations found that the adsorption energies followed the order: SDTBAT (-300.35 kJ/mol) > SNBX (-186.10 kJ/mol) > SNBDTC (-163.78 kJ/mol). It was observed that the SNBX and SNBDTC adsorbed through the S atoms, while the SDTBAT adsorbed through the S atoms and N atoms onto Fe atoms. It is clear that the SDTBAT had a strong binding energy which is attributed to the presence of the triazine group that is electron depositing to the active S atoms compared to the SNBX which possessed the oxygen that is electron withdrawing. Moreover, the SDTBAT collector was found to bind stronger than the xanthate. This was complimented by microflotation tests that showed fast floating and higher recoveries of pyrite when using the SDTBAT collector than the xanthate and DTC. This clearly demonstrated that the SDTBAT is a potential collector to replace the xanthate collector due to its high selectivity and flotation power.

Aug 24, 2022

By Aspinall T. O, Loch R. J, Bell L. C

Design of reshaped spoil surfaces is an important pan of rehabilitation of spoil piles resulting from open-cut coal mining. To achieve the optimum design, it is necessary to make accurate predictions of soil losses from spoil piles. Predictions can be based on data gathered using laboratory and field rainfall simulators and field plots. These data are used to derive input parameters for erosion prediction models. Model input parameters include rainfall erosivity, spoil erodibility, sediment size distribution, runoff, soil loss and peak runoff rate. Model outputs can be used in design and include predicted sediment size distribution and predicted soil loss.The traditional method of gathering input data is monitoring field plots. This method is very expensive, must be done on a long-term basis and is affected by irregular rainfall. Long-term field plots are necessary to obtain data that can be applied to empirical prediction models such as the Universal Soil Loss Equation (USLE).

Jan 1, 1992

By S Agdaca

Leightons is the sole contractor at the Wahana Coal Mine Project (part of the Bayan Group), a multi-seam thermal deposit consisting of 13 coal seams with thicknesses ranging from 20 cm to 500 cm. The average dip of the deposit is 18 - 20°. However the deposit is heavily faulted in certain domains with anticlinal and synclinal structures present causing seams to dip up to 90°. Average thickness of coal inter-burdens range from less than 2 to 12 m. The minimum coal thickness being mined is 20 cm. Coal production is direct mine mouth feed and there is no coal processing plant currently used.Conventional blasting between the highly variable coal horizons in the early phases of the mine life was suboptimal from a production perspective. Drilling short holes with a reduced drill pattern, and on occasions with sloping benches, resulted in low drilling and explosives loading efficiency.A high frequency of small size blasts fired up to two times daily was required with each blast consisting of four up to five small blast areas. Each blast averaged approximately 36 000 bcm thus causing a low utilisation rate of the mining fleet from frequent blast delays and the inability to deploy larger Liebherr 9350 excavators on bench. It was identified that increasing the production targets required a more efficient drilling and blasting process. Furthermore, due to the highly variable nature of the deposit, the existing coal model was determined to not be able to accurately predict the coal seam location prior to blasthole drilling.In August 2009, Leighton Contractors Indonesia and Orica Mining Services with the full support of the Wahana team, initiated through seam blasting trials using electronic blasting systems. Further trials were conducted in November 2009. Trial evaluations indicated that the benefits from decreased lost time events due to blasting, an increased vertical advance rate per blast, and increased overall blast size could be attained while not damaging or diluting coal.Full-scale?through seam blasting has been ongoing since February 2010 for the upper eight coal seams to extract coal seam thicknesses ranging from 20 - 100 cm. This paper will discuss the parameters of the through seam blasting process in Wahana and the resulting benefits to the overall mine production.

Nov 8, 2011

By F K. Altuntov

The study presented in this paper is based on the Kiirunavaara underground mine located in the northern part of Sweden. The deposit is mined using the sublevel caving (SLC) mining method. This paper briefly reviews the alternative SLC layouts, as well as presents challenges for the application. In Kiirunavaara currently, the traditional SLC design is employed, however, continuous improvement of material transportation systems used in mines generates a need for investigation of the current and future production area layouts in sublevel caving mines. When identifying the potential benefits of the alternative production area layouts, the production targets and effects must be further investigated to understand the benefits and potential effects. In this study, several different layouts have been investigated and evaluated. The designs include different development configurations and orepass infrastructure moved further away from the orebody and still require further investigation.

Mar 29, 2023