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The Louvicourt Mine is located 600 km north of Montreal in the province of Quebec, Canada. The owners are Aur Resources Inc (30 per cent), mine operator, Novicourt Inc (45 per cent) and Teck Corporation (25 per cent). It is an underground mechanised operation with a total work of 290 employees. Since 1994, the mine production is 11 120 456 tonnes at 3.60 per cent Cu, 1.51 per cent Zn, 26.07 g/mt Ag and 0.96 g/mt Au. At a rate of 4300 tonnes per day 365 days per year. Louvicourt Mine is a copper operation since 80 per cent of our revenue is from copper. The mining method is transverse blasthole stoping, mining primary and secondary stopes. The stope dimension is 15 m wide by 30 m high and 10 m to 80 m long for primary stopes. The secondary stopes are limited at 25 m long. Close to 100 000 metres of production drilling per year is required for 115 mm and 165 mm diameter holes. Backfilling is done with pastefill from the start of the operation in 1994. The binder ratio is 4.5 per cent in primary and two per cent in secondary stopes (80 per cent slag and 20 per cent cement). Development requirement to maintain production was 5000 m from 1996, 1997, 1998, 4200 in year 1999, 3200 m in year 2000 and an average of 2500 m in 2001 and 2002. Typical 6.0 m + 4.3 m, 4.9 m + 4.6 m and 4.5 m + 4.6 m headings. In addition more than 1500 metres of development will be excavated in/under paste backfill before 2004. A total of 60 mobile units are required among which for lateral development (three two boom electric hydraulic jumbos û AC Boomer 322, four electric hydraulic roofbolters û Maclean MEM-928 and AC Boltec PL350 and two anfo loader truck û Dux) and for production (four ITH drills û MTI, one electric hydraulic longhole drill û AC Simba H1357, eight LHD û AC-Wagner ST8-B, 1 LHD ST-1000 and two haulage 30 tons trucks û Dux. The 1 January 2002 reserves are 4.1 M tonnes grading 3.1 per cent Cu and 1.9 per cent Zn at an NSR (Net Smelter Revenue) of $US 53. The operating mining costs were at an average of $US 23 from 1996 to year 2000 and $US 25 in 2001. The operating costs ranged from a low of $US 0.41 per pound of copper in 1999 to a high of $US 0.49 per pound of copper in 1998.

Jan 1, 2002

Diamonds occur in N.S.W., Victoria and north west Tasmania. Diamonds from different areas appear to have their own characteristics, lending weight to the suggestion that three or four different sources (different in location and age) have produced the diamonds found in alluvials at quite a number of locations in Eastern Australia. Explorers for kimberlites in Victoria must overcome a number of field difficulties. The writer's approach attempted to find exploration targets by literature search, noting unusual garnet occurrences, and by relating records of diamond occurrences to described mining activity. While the records are biased towards gold, the study did lead to a pipe, located well away from known diamond-bearing alluvials. Core from one test hole has shown this intrusive to include an alkaline ultramafic breccia with many kimberlitic affinities - a breccia containing mantle-derived minerals, but not a true kimberlite. (A detailed rock description is appended to this paper, but is not the purpose of the paper.) This result enhances the possibility that kimberlites may occur in Victoria, on or near large scale faults or grabens. The depth of fracture penetration may have been different

Jan 1, 1979

By J Butler, N Patel, S Narendranathan, S Rastogi

The Yallourn Mine (YM) commenced mining of brown coal in the 1920s under operation by the State Electricity Commission of Victoria (SECV), prior to it becoming privatised in 1996. The present owners of the YM, Energy Australia Yallourn (EAY), acquired the site in 2000 and intend to continue operations until scheduled closure in 2028. Brown coal extracted from the YM is fed into the Yallourn Power Station which supplies approximately 22 per cent of Victoria’s electricity. Water within the YM is collected and stored in the Township Field Fire Service Pond. Water levels within the Fire Service Pond (FSP) may fluctuate considerably depending on the requirement for storage at the site. With due consideration to the stability performance of mine batters adjoining the FSP, the authors performed slope stability assessments to evaluate the risk and consequence of batter instability during the potential pond filling as it approaches the ‘critical pool’. The critical pond is defined as the level at which the driving forces (destabilising forces) are at their greatest and the resisting forces (lake forces, friction) are at their lowest. This assessment was undertaken using the Mine Geotechnical Risk Index (MGRI) approach.

Aug 2, 2023

By R Price, B H. Miller

"The rise of autonomous haul truck fleets may have been initially driven by the rapid escalation of the cost of equipment operators, but the actual improvements reported by the major iron ore miners far exceeds the simple reduction in workforce. Utilising cost models developed for feasibility studies and mining contract modelling, the limits of cost improvements due to simple labour reductions are compared to those achieved by the early autonomous fleet adopters and through traditional process improvement techniques such as fleet class upgrades or hot-seating of equipment. As autonomous fleet implementation matures, it is likely that the gains seen by current adopters will seem minor compared to the limits generally governed by the limit of equipment utilisation. This ultimate limit is explored through time utilisation models, which were also developed for use in feasibility studies and mining contract modelling.Citation:Miller, B H and Price, R, 2016. Autonomous Haul System Savings - It's Not about Less Truckies but Better Truckies!, in Proceedings Ninth AusIMM Open Pit Operators' Conference 2016, pp 2–10 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 15, 2016

By Hunter W. H, Wilson I. P, Craig R. W

In September 1988 the Imperial Smelting zinc - lead blast furnace at Avonmouth England, was uprated from its then 52 week annual rated capacity of 90 000 tonnes/year of furnace zinc to 110 000 tonnes. In order to provide sufficient extra feed the sinter plant was likewise uprated in April 1989 by increasing the blowing rate of the sinter/acid plants from 100 000 to 120 000 Nm3/h, effectively a 20 per cent increase. In the 12 months after the uprate zinc production was only 85 000 tonnes. Two years after the uprate zinc production had increased to 100 000 tonnes, but it was not until January 1992 that a rolling 12-month zinc production of 110 000 tonnes was consistently achieved. This paper describes the problems encountered with the uprate and how they have been or are being overcome. It deals chronologically with events starting with the restart of the ISF at the end of 1988 and covers the following issues and their resolution: 1. Initial shortages of sinter due to an offset in uprate timing. 2. Greater difficulties in operating the larger furnace particularly on start-up, the greater tendency for iron production and serious slagging problems. 3. Coincidental timing of deleterious trends in slag chemistry. 4. Refractory problems in the furnace shaft and condenser. 5. Shortages of sinter due to commissioning problems with the wet gas and sinter tip end fans, and difficulties in obtaining sufficient sinter hardness. 6. Unrealised full benefits of the sinter-acid uprate due to a late change in design to dispense with sinter gas recirculation, but offset by successful use of oxygen. In a planned maintenance stop in September 1992 further improvements to the furnace design were carried out to resolve the refractory problems, and with improvements to the sinter plant the furnace zinc output capacity is planned to rise to a rate of 120 000 tonnes/year within two years.

Jan 1, 1993

By M Astorga, J Jackson, J Gaunt

The Ban Houayxai mine (BHX) is a relatively low-grade, low-cost, open pit gold-silver deposit in Laos operated by Phu Bia Mining, a subsidiary of PanAust. The ore production rate at BHX is 4.5 Mt/a, with direct tipping to a semi-autogenous grinding/ball mill circuit with a carbon-in-leach process plant. Approximately 100 000 oz of gold is produced per annum.The operation is located in mountainous terrain with minimal run-of-mine stockpiling. This results in limited capacity for blending from stockpile, with the mill instead reliant upon direct feed ex-pit. Since commissioning in 2012, the plant has seen significant variation in milling rates due to variability in the feed properties of the oxide, transition and primary ores. As part of an ongoing continuous improvement program, an integrated approach was initiated, focusing on maintaining and enhancing production in the future as the proportion of harder primary ore increases. This will specifically focus on direct blending from the loading face using the ore properties and blast fragmentation to maintain mill throughput.This approach was based on the concept of physical assets management, commencing with improving information and knowledge of the condition of the orebody through modelling the characteristics, variability and performance of the feed for processes relevant to throughput. These models are used to support both mine and process plant production planning.The key ore feed characteristics and parameters modelled for the life-of-mine are blastability index, powder factor, crushability or impact resistance and grindability. These predictive spatial models were based on the data from diamond drill holes used in resource definition and geotechnical drilling programs by integrating geotechnical, geological, geochemical and metallurgical data.At the early stages of implementation, the models are being utilised for ore blending decisions to provide guidance and support for budgeting, long-term mine planning, blast design for mill feed and providing the mill with an expectation of performance.CITATION:Jackson, J, Gaunt, J and Astorga, M, 2014. Predicting mill ore feed variability using integrated geotechnical/geometallurgical models, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014, pp 165–176 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By H Utama, hi, A Suw

PT. Indo Muro Kencana (PT.IMK) is gold and silver mine located in Central Kalimantan province, Indonesia. Mining activities in PT.IMK is conducted in Bantian pit area by applying surface mining method and using combination of excavator and hauler (articulated dump truck and rigid haul truck) as the transportation for ore to the crushing plant and waste material to the waste dump. Current blasting practice is ore-waste blasting. Ore-waste blasting is the term where the ore and waste rock are blasted in the same blasting process. Blasting is conducted by using nonel initiation system. By multiple pits operation, monthly powder factor is varied from 0.23 kg/t to 0.27 kg/t. Emulsion blend with density 1.15 g/cc is used as bulk explosive. At this stage, the concept of geoblast is proposed as an engineering tool for Drill and Blast engineering practice in PT.IMK. In addition, there is a need to maintain rock fragmentation by blasting, to maintain monthly powder factor, and to maintain excavator’s productivity. Next, the concept of scaled depth of burial (Chiappetta and Treleven, 1997) and the concept of blastability index (Lilly, 1986) are proposed as geoblast’s tools to improve blasting practice in PT.IMK. It is expected that rock fragmentation by blasting, explosive consumption and excavator’s productivity will be maintained by integration of geoblast and drill and blast engineering. Furthermore, the objective of this paper is to integrate geoblast and drill and blast engineering in order to create an engineering guideline in PT.IMK. Additionally, blasting trials with air decking application is proposed in order to support the geoblast’s improvement. It is planned that these blasting trials will be conducted in 2022. In conclusion, it can be seen that integration of geoblast and D&B engineering in PT.IMK is relevant to improve current drill and blast current practice and applicable to maintain rock fragmentation by blasting.

Nov 29, 2022

By G Gopinath, R Balachander, H S. Venkatesh

This paper describes in detail the controlled blasting adopted and the measurements undertaken for the successful excavation of large twin tunnels under a shallow cover for a six-lane highway project in India. The tunnels are D-shaped, each 15 m in diameter, and the cover between the tunnels and the existing railway track was about 14 m. These twin tunnel tubes were characterised by two separate three-lane carriageways for each direction of traffic, and were planned with about 25 m of reinforced cement concrete (RCC) portals at the entrance and exit of the tunnels. The major challenge of the project was the blasting of tunnels in rock below an operating railway line. As the tunnels approached the area below the railway line, the ground vibrations were monitored vertically above the blasting faces and on the railway track sleepers. Vibration intensities between those generated from tunnel blasting and those from the movement of the trains were compared, and it was found that the vibrations from blasting were lower than those from the moving trains. Though the permissible level of ground vibration based on the Indian Standard (IS) code happens to be 70 mm/s, the maximum charge per delay was calculated restricting this limit to 50 mm/s. This ensured that the vibrations in the vicinity of the railway tracks never exceeded the prescribed limit of 70 mm/s. The surface ground settlement before and after the blasts were also measured using an optical levelling instrument, which confirmed that there was no significant impact on the surface movements due to blasting. Extensometer readings inside the tunnels showed that there was no deformation in the tunnel. These twin tunnels are now operational and are providing faster connectivity with safety.CITATION:Gopinath, G, Venkatesh, H S and Balachander, R, 2015. Controlled blasting for twin tunnels with a shallow cover under an operating railway line, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 801–808 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By H-S Yang, H Park, S-J Kim, Y-H Ko, H-S Kim, S-H Cho, J-K Kim

An investigation was undertaken to devise a new method for reducing dust spread during blasting demolition utilising water tube and detonating cord. Modelling of smooth particle hydrodynamics (SPH) with AUTODYN software was initially completed to determine the expected diffusing behaviour of the water tube. The actual behaviour was then captured with high speed videos during each experiment, and each result analysed in detail. Water mist diffusion was generated and analysed by changing three key variables. In sequence, these included changing the location of detonating cord in the tube, the charge quantity and the tube diameter.As a result of experiments and simulation, optimum design for ideal diffusion behaviour was suggested and each behaviour of water mist was classified by diffusing shape of water tube as the design criteria.CITATION:Ko, Y-H, Kim, J-K, Kim, S-J, Park, H, Kim, H-S, Cho, S-H and Yang, H-S, 2015. Mist diffusion experiment and evaluation of water tube for blasting demolition dust control, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 809–814 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By J V. Mehrabani, A R. Sayadi

Adequate estimation of equipment costs is a key factor in a feasibility study and in evaluation of design alternatives for mining and mineral processing projects. Inaccurate early cost estimation could lead to the elimination of investment benefits. However, this estimation is complex and time consuming because of data accessibility problems and consequently access to a rapid, up-to-date and accurate tool is desirable for carrying out cost estimations.This paper describes the development of models for estimating capital and operating costs for a jaw crusher, which is used as the primary crusher for a wide variety of mineral beneficiation plants, applying single and multivariate regression analyses, together with the design of experiment approach, based on principal component analysis. Explanatory variables include gape size, length of feed opening and power as numeric (quantitative) variables and toggle type (single or double) and the existence of extra heavy duty construction as categorical (qualitative) variables. The performance of each model has been investigated and comparing the estimated R-squared (R2) of each model, the best was selected and its validity examined. The cost models include the capital and operating costs along with the operating cost components. The results indicate that the suggested best cost model is sufficiently accurate for use in feasibility studies.

Sep 26, 2011

By A M. Barrett, P J. Chapman

As filtering technology improves and there is an increased focus on environmental impacts, the concept of managing fine coal rejects by filtering, mixing with coarse rejects and dry stacking them is being considered more seriously. In Queensland, there appears to be a desire from regulators to move away from wet rejects facilities towards dry-stacked facilities, resulting in a change in the risk profile for the facility.A coalminer has a site with many constraints, including limited area that is not planned to be undermined and nearby rivers that are prone to flooding. The design of a dry-stacked facility that meets the operational requirements and environmental compliance criteria is challenging in subtropical climatic settings, where there is a need to consider wet seasons with high-volume, sustained rainfall events. The impact of the rainfall is multifaceted, requiring careful surface water management to separate clean and contact water, adequate foundation preparation to limit seepage to groundwater, management of stack stability given a potentially high phreatic surface and the avoidance of potential flood zones. These design constraints were compounded by a limited area available for rejects storage that would not be undermined.This paper presents a summary of the conceptual design of a dry rejects facility, including consideration of regulatory and client requirements, provision for expansion while also considering progressive rehabilitation, diversion of clean water as much as practicable during operations and siting constraints. A step-by-step concept is presented that addresses each constraint during the life of the facility.CITATION:Barrett, A M and Chapman, P J, 2015. Conceptual design of a dry-stacked coal rejects facility in a subtropical climate, in Proceedings Tailings and Mine Waste Management for the 21st Century , pp 27–34 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 27, 2015

By A E. Annels, G F. Johansen, B W. Cuffley

Nuggety gold-quartz reefs are characterised by localised, high-grades and in many cases coarse free-gold. Geometrical and grade complexities are often observed. The high-grade regions are generally erratic and have a low continuity; though may make the deposit economic. Viable grades are generally contained within discrete oreshoots, which are surrounded by barren/low-grade material. Drilling is an effective measure of geological continuity, however grade distribution can only be reliably obtained from underground development and bulk sampling. In general terms, only an inferred resource is likely to be estimated from surface drill data alone. Underground development, further drilling and probably bulk sampling will be required to delineate inferred and indicated resources. Closely spaced development with bulk sampling is likely to be the best way to determine a probable reserve. The resource evaluation process must driven by a clear geological model that should attempt to understand the controls to reef geometry and grade distribution. Geological interpretation has traditionally been undertaken in two-dimensions, though computer software is now being used to produce three-dimensional models. Grade interpolation has generally been undertaken using classical methods, typically weighted averages, zone of influence weighting and polygonal/sectional blocks. Tonnages can generally be calculated from diamond drill and development information with a reasonable degree of confidence. Grade is much more difficult to define with confidence because of its highly erratic and discontinuous nature. To account for variability it is recommended that a best estimate be reported within a grade range, as this more accurately reflects the uncertainties involved. Reporting should be undertaken within the framework of the JORC code, but should involve special consideration of grade reliability.

Jan 1, 2000

The last decade has seen a very rapid development in TBM tunnelling. Some trends are purely technical or economical and others relate more to preferences of different interests. Power levels have more than doubled and thrust levels too. Shield machines are developed to be able to handle harder rock and the open TBMs are made more feasible for `difficult' rock. What is right and what is wrong is more a matter of opinion, but one thing is certain: TBMs now are a useful tool and no longer an experimental work-shop for the enthusiasts.This paper highlights some recent developments and indicates some probable future ones. A few features that have increased the open TBMs' ability to cope with ever harder rock and also to overcome problems with poor rock are discussed in detail, using the Atlas Copco JARVA Mk 27 -the most powerful TBM built so far - as a reference.

Jan 1, 1993

By T Day, I G. Robertson, G J. Lyman

"Various approaches are used in the industry for establishing procedures (sample mass estimation and method of subsampling) for sampling and analysis of gold ores. The key factors that determine the accuracy of the results that can be expected for a given set of procedures are widely misunderstood.Recent developments in sampling theory show that the estimation of the accuracy that can be expected for a given gold ore as a function of sample mass is actually a simple matter. For a given sample mass, the accuracy that will be achieved depends only on the grade of the ore and the range of masses of the grains of the gold in the ore. This, of course, demands that a reasonable estimate of the distribution or spread of gold grain masses in the ore be known. But this is something that should be known if the gold is to be efficiently recovered in the processing plant.This paper provides a simple explanation of the real facts about sampling of gold, either in the field or in the plant, and provides results for a number of cases of combinations of fine and coarse gold in an ore. The prerequisites for accurate gold sampling are:correct sampling with no loss of sample or contaminationsample preparation procedures that recognise the impact of the gold grain mass spreadpractical analysis of the sample masses required to provide acceptable accuracy.Issues of corporate governance are also discussed.Citation:Lyman, G J, Robertson, I G and Day, T, 2016. The Simple Facts about Sampling Gold Ores, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 159–172 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

Gas dynamic phenomena, particularly instantaneous outbursts of coal and gas are problems experienced in gassy, highly stressed coal, both usually the result of depth. Diff- erent mining geometries present different opportunities for seam gas to escape, starting from the virgin condition unaffected by mining and ending with the mining of the coal. The mining of the coal varies from entering a virgin seam cross-measure, heading out into virgin seam, advancing longwall faces into virgin seam and extracting developed blocks from large longwall retreating blocks to small pillars. Each method has its particular gass- iness patterns and particular stress patterns, both changing to some extent with time, and changing as mining progresses. Geological structural anomalies are the cause of gas and stress anomalies, causing in turn greater gas and stress manifestations. Pre-drainage of gas from the working seam not only degasifies the coal but causes it to shrink and so may affect stress as well as gas patterns. Unfortunately prior degasification is not all benefit, and there are some d13- advantages to its use. Also although it is possible to degasify places ahead, gassiness will return if drainage holes are shut down or sealed up and other deleterious aspects may be induced. Research should continue into the problem of instantaneous outbursts of coal and Gas.

Jan 1, 1986

By A G. Sheard

Fire safety requirements inside tunnels have become more demanding as tunnel designers consider the impact of potentially larger fires. Tunnel ventilation equipment is required to clear hot gases in the event of a fire. This is vital to maintain smoke free escape routes for people trapped in the tunnel and to facilitate safe access for emergency services. As the fire becomes larger, the temperature of clearing hot gases increases. Traditionally, engineers have taken into consideration gas temperatures of 250°C when designing tunnels. Because of larger fires, engineers must now design equipment to clear hot gas up to 400°C.Increasing the temperature of hot gases to 400°C requires tunnel ventilation equipment that goes beyond historic limits. As temperature rises, material strength reduces, consequently mechanically constraining fan aerodynamic performance. Developing tunnel ventilation equipment suitable for clearing hot gases at 400°C requires a fan technology that is more typical of industrial turbomachinery. The standards EN 12101-3:2002 and ISO 21927-3:2006 defi ne the necessary design criteria and testing methodologies to verify that a range of tunnel ventilation equipment is capable of extracting hot gases.This paper describes the process by which the authors certified in accordance with EN 12101-3 and International Organization for Standardization (ISO) 21927-3 a range of powered ventilators suitable for clearing smoke and hot gases at 400°C for two hours. Proving the high temperature capability of the new range provided a challenge as a consequence of the large physical sizes and high motor powers involved. Additionally, the need to install vertically, as well as horizontally, imposed high thrust loads on motor bearings, further complicating both mechanical design and testing. The paper presents the methods by which the authors overcame these challenges.

Mar 8, 2011

By D Brereton, C J. Moran

In this era of rising focus on the social and environmental costs and benefits of mining operations, the valuation methodologies used to assess such impacts become increasingly important. Within an industry dominated by an engineering culture that places emphasis on net present value (NPV) calculations using high discount rates, difficulties often arise when attempting to justify projects with longer-term, harder to quantify benefits. This is particularly so when these benefits fall in the social and environmental domains and involve other stakeholders. In this paper, we summarise the results of a recent literature review on this subject, which focused on the valuation of æBeyond ComplianceÆ initiatives in the mining sector. We discuss a range of approaches including the quantification of hidden internal costs, financial methods of incorporating risk into calculations, the integration of quantitative and qualitative information, and also the valuation of æexternalitiesÆ or costs and benefits borne by others. We then use a series of examples relating to the management of water in the minerals industry to illustrate how some of these approaches could be applied. There are high-level commitments from both individual companies and groups such as the ICMM to incorporate sustainable development criteria into decision-making processes within the industry. We conclude the paper by arguing that, to meet these commitments, there is a need for the industry to adopt a broader range of valuation methodologies than appears to be the case at present.

Jan 1, 2006

By Q Wang

Net present value (NPV) has been the main criterion in mining project evaluation and production plan optimisation. It is a common practice in NPV calculations to take into account only the tangible cost items in the corporate accounting system. The costs incurred due to damage to the local ecosystem and environment are left out and treated as externalities. Such practice on the one hand underestimates the total cost of a mining operation and, on the other hand, ignores the difference in ecological/environmental impacts between different mine plan alternatives. This paper attempts to measure the ecological costs of mining and to incorporate such costs into the evaluation scheme in mining project evaluation and mine planning. The ecological costs of mining are estimated based on the areas and types of land (eg forest land, grazing land, crop land) damaged by mining activities and the carbon emission from fossil fuel consumption. For the land areas (and their associated ecosystems) directly damaged by mining activities, the ecological costs are measured by the prevention and restoration costs, the lost values of eco-services of the damaged land, including bio-product production, soil erosion control, water run-off control, oxygen release, air pollutant absorption and formation of soil nutrients. The ecological cost of carbon emission from the energy consumption of a mining operation is measured by the cost of cultivating the equivalent forest land required to uptake the emitted carbon. A case study on a large-scale open pit mining project is presented to demonstrate the influence of ecological costs on the mine planning outcome.

Sep 26, 2011

By J Crosbie, M Lord, R Bitencourt, P Defferrard

The MMG Century mine, which is the largest zinc/lead mine in Australia, began operating in 1999 and ceased mining from the open pit in July 2015, with processing completed in early 2016. Closure plans were developed from 2006, and detailed planning for closure commenced in early 2011 with the appointment of a dedicated closure planning resource. The site will continue to be managed during the active and passive closure phases to complete rehabilitation commitments and demonstrate that closure completion criteria have been achieved.Completion of the mining/processing phase of the project provides an opportunity to reflect on successes, missed opportunities and lessons learnt with respect to planning and execution of mine closure. This paper highlights examples of success at Century and also identifies key learnings and areas for improvement in progressive closure outcomes. Specific examples examine mine planning and execution involving bulk movement of waste rock and construction of permanent waste encapsulation landforms. The paper also provides recommendations for other sites that are planning for closure.Decisions made during the operational phase of the project life can result in a significant increase in closure liability and reduce the likelihood of achieving lease relinquishment. Likewise, it is of paramount importance to assess the consequences and thoroughly understand the risks associated with changes or deferral of closure plans, progressive rehabilitation targets and closure related key performance indicators.CITATION:Lord, M, Bitencourt, R, Crosbie, J and Defferrard, P, 2016. Closure planning lessons Learnt from the Century mine, in Proceedings International Mine Management Conference, pp 207-216 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 22, 2016

By C Boyle

Brockman 2 mine is located in the Pilbara of Western Australia and is operated by Rio Tinto Iron Ore. Mineralisation is hosted by synclines where there is structural complexity to focus fluid flow. Mineralisation is hosted by the Brockman Iron Formation. Mining is on 10 m benches and blastholes spaced at 6 - 8 m apart are geologically logged, sampled by shovel cuts, and assayed using X-ray fluorescence for ten chemical variables. Blast blocks are digitised by mine geologists to assign destinations of material within blasts. Resource evaluation drilling is on a spacing of 50 m × 50 m across known mineralisation and consists of recent reverse circulation drilling and older open hole percussion drilling. Sampling of resource evaluation drilling is by rotary splitters or riffle splitters, and assaying is by X-ray fluorescence for ten chemical variables. Ordinary kriging of resource evaluation produces a long-term resource model. This study examines whether mine planning would be improved by using a blasthole estimate to extrapolate grades into deeper benches and merging the blasthole estimate with the long-term model made using resource evaluation drilling. The two estimates are merged into one model using an inverse weighting based on kriging variance. However, reconciliation results in a test bench show the original estimate made using resource evaluation drilling more closely matches mine geology block outs than does the merged model. This is due to a local trend of more high grade and higher iron grades of high grade in the area. The blasthole estimate of deeper benches must use extrapolation and it is diffi cult to take local trends into account. In contrast, the estimation using resource evaluation drilling is mainly interpolation and is better able to handle local trends.

Jul 11, 2011