Search Documents

Control of construction projects is the responsibility of the Project Leaders who have the authority to make decisions. Project Controls cost and scheduling systems provide the Leaders with timely information on which to base their decisions.The Client (Owner) will invest up to three per cent of the project cost on Project Controls. To ensure that this investment is effective, the Owner or Manager must have a clear understanding of the information required from the Project controls systems. If appointing a single Managing Contractor, the Owner or Manager must stipulate what is and what is not required from the Project Controls Systems. There are numerous computer programs available for scheduling and cost control and it is important to select systems which best serve the Owner's needs.

Jan 1, 1993

By J Gould, T Elkington

Mine operation strategies are commonly optimised with a net present value (NPV) maximisation objective. There are well known properties of a typical mining project optimised for NPV, including high production rates and declining cut-off grade schedules. In practice, however, frequently there are operations that diverge from such NPV-maximising properties. Although a range of reasons for this can be at play, one possible explanation is that operators intuitively adjust their strategies to compensate for the flaws inherent to NPV maximisation. In particular, NPV analysis does not reflect the benefits of managerial flexibility inherent in a mining plan. Real options valuation (ROV) is an alternative approach that addresses this issue. We compare the properties of NPV and ROV optimisation solutions generated for the same scenario. Our simple case study demonstrates that the presence of real operating options in the optimisation model not only entails increased project value as expected, but also tends to encourage lower processing capacity (with lower associated capital cost), higher cut-off grades and greater use of stockpiles.CITATION:Elkington, T and Gould, J, 2012. Real option value optimisation, in Proceedings Project Evaluation 2012 , pp 35-40 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 24, 2012

By S H. Algie, E J. Knoblauch, I A. Lovatt, S G. Doig

Mill operating teams are expected to engage in a perpetual quest for process improvement – pushing the boundaries of practical process limits and testing new technology to achieve the next step-change in plant performance. Often the barrier to improvement lies in gaining management approval for process changes.Getting approval for investment in novel technology carries an added degree of difficulty, particularly in risk-averse operations, when it challenges conventional wisdom. A case in point is a new technology for controlling variability in slurry distribution. There is a widely held view that less-than-specified classification efficiency and spraying and roping of individual units in a hydrocyclone cluster, must be accepted as unavoidable. That these are wholly correctable deficiencies resulting from unequal loading of the individual units seems a bold claim. Wouldn’t the high turbulence of pumped slurries overcome all potential for segregation of solids? This paper discusses how the task of developing confidence in a new technology can be approached by a combination of physical modelling at small scale, and full-scale operating experience. Proving efficacy then leads to the next task: demonstrating value. Again, the particular technology serves as an example that could be of value to others developing relatively low-cost technologies. A high-pay-off as an investment may be achieved with improvements in metallurgical performance that are small in relation to normal process variations. This has led to an exploration of how process simulation modelling could be used, in conjunction with production data, to provide a benchmark for the expected change, which could be expressed in monetary terms.The next goal, which is currently being achieved in operating plants, is to realise the greater value now available through the optimisation of cyclone configuration, which becomes practical when the confounding effects of cluster instability are removed through equal loading.CITATION:Doig, S G, Algie, S H, Lovatt, I A and Knoblauch, E J, 2018. Investing in new process improvement technology: reducing the risk of implementation – a way forward, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 571–584 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

Heatings, fires and explosions in New South Wales (NSW) and Queensland coal mines since 1972, are reviewed. Geological and mine settings are outlined. Where available, gas analysis and critical decisions are discussed. This paper updates work by NSW Department of Mineral Resources (1995) and the Queensland Department of Mines and Energy in a report on fires and explosion in Queensland coal mines (Richardson and Ham, 1996). Incident data has important applications in terms of evidence-based risk assessment and identifying key issues to be integrated into competency based training programs. Predictive indicators are reviewed in terms of coal parameters in seams where the incidents occurred.

Jan 1, 2005

By M M. Collisson

The Synthetic Rutile (SR) plant operations, North Capel, produce electrostatic precipitator (ESP) fume and kiln scrubber residue. Typically ESP contains about 18 per cent Ti, 13 per cent Fe, 0.5 per cent Mn, eight per cent Si, six per cent Al, 16 per cent C, 0.9 per cent Zn, 0.6 per cent Pb, 300 ppm As, 300 ppm Cr and 5 ppm Cd. These wastes are held in a storage dam together with neutralised acid effluent (AE) liquor. A project was initiated to treat the fume and produce a benign residue that could be delivered to landfill, reducing dam space storage. Plant AE, containing free acid (9 - 14 g/L), Fe (6 - 7 g/L) and Mn (1 g/L), was used to leach ESP. The degree of extraction achieved for various elements include about 20 per cent Zn, almost 50 per cent As, five per cent Al, 0.4 per cent Pb and 15 per cent Cd. Extraction was most effective at pH The residue from AE leaching still contained Pb. TCLP tests of the leach residue produced Pb extraction significantly higher than the 0.1 ppm stipulated by Class 2 landfill guidelines. Other elements (eg As, Cd, Cr) successfully met the limits. Further leaching of the residue with EDTA extracted up to 65 per cent of Pb content. TCLP tests on treated residue showed Pb leaching potential of Lead was removed from the leach circuit as the sulphate by acidification of EDTA leach liquor, allowing EDTA to be recycled. The study showed that one waste stream can be used to treat another and help reduce plant waste.

Jan 1, 2001

By W Zeng, D Walker, D Cai, E Baafi

"The efficiency of the truck-shovel haulage system is significantly influenced by the truck fleet size. A discrete event simulation model with 3D animation has been developed to estimate the productivity of a truck-shovel network system, shovel efficiency, truck cycle time, truck utilisation and the optimal fleet size. The simulation model considers the dynamic variability of traffic conditions and also the truck as an individual entity. The resulting truck cycle time was found to be realistic and thus providing a better tool for strategic mine planning.Citation:Zeng, W, Baafi, E, Walker, D and Cai, D, 2016. A 3D Discrete Event Simulation Model of a Truck-shovel Network System, in Proceedings Ninth AusIMM Open Pit Operators' Conference 2016, pp 265–273 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 15, 2016

By J Euston

In the 1990s round flotation tanks were introduced as a result of the lower fabrication costs for larger cells. It was soon observed that round tanks held a number of further advantages over the traditional rectangular cells, such as better hydrodynamics, lower costs and reduced footprint per unit volume.In the last twenty years flotation cell sizes have increased almost exponentially and designs of 250 to 300m3 are a well established standard with plans for full scale pilot installations of 500m3. The option to install these large cells gives flexibility in circuit arrangements and allows efficient and economic matching of flotation lines with modern mill designs. The benefits of scale up are well known with reduced numbers of cells required, lower power costs per unit volume, etc. These benefits will be briefly reviewed with specific reference to the installation of 300m3 to 350m3 flotation cells at the Copperton Concentrator of Kennecott Copper in Utah, USA.The sophisticated techniques required to model these ‘supercells’ have led to a critical review of the fundamentals of flotation science and engineering. In addition to outlining the advantages of large flotation cells, this paper will review the science which supports, and occasionally challenges, the ongoing trend to larger flotation cells.

Aug 8, 2011

By J Franz, S Bock, N Hentrich

Artificial ground freezing is the unrivalled method for groundwater control during mine shaft sinking. Moreover, ground freezing may also provide temporary ground support. The fast development of fibre-optic temperature monitoring devices, three-dimensional freeze pipe survey methods and numerical modelling software provide grounds for developing new and reliable tools for ground control during freeze shaft sinking. This paper presents examples of predictive numerical models that are based on back-analysis and calibration of temperature data that covers the complete construction cycle from pre-freezing through to artificial thawing. The back-analyses showed that standard numerical models are not capable of predicting the prevailing thermo-mechanical conditions satisfactorily. This paper describes the development and implementation of new tools for improved consideration of concrete hydration and latent heat, along with the introduction of other algorithms that improve and simplify the efforts of managing complex numerical models. A function is presented that allows for an automated correction of the model’s mesh based on results of three-dimensional freeze pipe surveys. This paper demonstrates that the latter consideration is important, because the often employed assumption of idealised freeze circle geometry in numerical models impacts on the accuracy of such model’s predictions. Subsequently, the verification of these extended numerical models is demonstrated. Due to its flexibility, the FLAC3DTM code was employed for the numerical modelling. This software provides the possibility of implementing additional functions written in internal FISH or Python language into the models. It is concluded that state-of-the-art techniques for freeze hole drilling and surveying, on-site monitoring and advanced numerical analysis provide a significant potential for both the design and the operation of largescale ground freezing measures. The application of the extended numerical models for predictive simulations during construction allows for temporary shaft lining optimisation, for assisting in operational decision making and for compensating temporary loss of monitoring data. Suggestions are made for required future work in order to further improve the developed ground control tools. CITATION:Bock, S, Franz, J and Hentrich, N, 2018. New tools for ground control in freeze shaft sinking, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 393–406 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By P Holtham, K Duffy, B, W Valery, E Tabosa, rade

Pre-concentration has potential to significantly improve Mineral Resource efficiency by removing gangue early in the process, before energy intensive and expensive comminution and fine processing. However, there are numerous options for pre-concentration and the performance and benefits of these are very case specific. To determine the best option for a particular deposit and to successfully integrate it into the overall mining and processing operations requires comprehensive evaluation. This needs to consider orebody suitability, technology capability and economic viability. The in-situ variability of a deposit and how it is influenced by mining methods and the liberation size of gangue are key factors which are generally not well understood. Geostatistical methods provide an insight into orebody variability, and the overall selection of any pre-concentration solution must be an integrated approach between geology, mining and processing. Many pre-concentration evaluations investigate the potential of just one option. Technology selection should start wide and may include: selective mining, pre-screening (with or without selective blasting), sorting (bulk and or particle), coarse gravity separation (e.g. jigging, DMS), coarse particle flotation, magnetic separation, etc. The upgrade performance of these processes is very ore specific and should be determined from testing, with detailed multi-component modelling used to determine the overall process performance. Many of these pre-concentration processes require significant ancillary equipment, feed preparation and materials handling. It is important to realistically estimate all costs to understand overall economics and select the best option. Some examples showing evaluation and comparison of pre-concentration options are presented in this paper. These include assessments of in-situ variability using geostatistical tools, process modelling of pre-concentration options and overall performance, preliminary layouts and cost estimates, and economic evaluations.

Nov 10, 2020

By Y Xie, L O’Connor, S C. Dominy

"Sheeted vein gold deposits are characterised by multiple subparallel veins and free milling gold. In situ heterogeneity of mineralisation results in grade and process parameter variability, which will increase project risk. Measured variability is often enhanced by poorly designed sampling and test work protocols. Protocols that are optimised within the framework of the theory of sampling to suit the ore type, together with quality assurance/quality control systems, will reduce variability and provide fit-for-purpose results. The geometallurgical approach requires spatially distributed samples within a deposit to support variability modelling. Diverse attributes from core logging, mineralogical/textural determination and small-scale tests are used to measure variability. This contribution emphasises early stage geometallurgical intervention during project development. It presents a discussion of representative sampling in the context of the theory of sampling. A case study presents a protocol that exemplifies how data can be gained from a well-designed and planned program to support a resource estimate and prefeasibility study.CITATION:Dominy, S C, O’Connor, L and Xie, Y, 2016. Sampling and Test Work Protocol Development for Geometallurgical Characterisation of a Sheeted Vein Gold Deposit, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 97–112 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

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

By R Varden

Friction rock stabilisers, commonly termed ‘Split Sets’, friction rock bolts, or continuous friction coupled bolts are the most widely used rock bolt in the Australian underground metalliferous industry. Historically, they were used for temporary reinforcement, mainly within ore development. However, their use as primary reinforcement has grown with the demand for higher development rates at lower costs. The introduction of galvanisation has increased their longevity and pattern bolting in conjunction with weld mesh, specifically made for use in development headings, has allowed the use of friction rock stabilisers to increase rapidly with development jumbo’s undertaking the installation. The installation of friction rock stabilisers appears to be deceptively simple, particularly when compared to resin encapsulated bolts. The friction rock stabilisers capacity is a function of the borehole parameters, rock mass quality/strength, the bolt dimensions and steel quality, and lastly and perhaps the most important the operator skill levels. However, as in industry we are still seeing common occurrences of friction rock stabilisers not reaching their design pull test capacity. Pull testing the friction rock stabilisers is used to verify the installation of the design capacity. It involves loading the collar of the bolt with a hydraulic ram until the design load is reached or the bolt slips. Pull tests often show failure rates of 20 per cent or higher. This is despite over 40 years of use and with millions of units installed annually. Such high failure rates would suggest there remain issues with the quality of installation. Ground support quality assurance and quality control (QAQC) programs are needed in combination with operator training to ensure that pull test failure rates are kept as low as practical. However, there are times when despite the QAQC being completed the pull test failure rates increase quite suddenly. When this occurs, it is important to conduct a thorough investigation. These investigations often show there are several factors that are not obvious to personnel installing the bolts. To complete a useful investigation, it must also be done methodically, working through the entire process that includes less obvious factors such as steel quality and the calibration of the hydraulic pull test kit. The aim of the paper is to give mine practitioners an understanding of the complexities of friction rock stabilisers, what QAQC requirements they should be doing on-site, what can go wrong during installation and, if necessary, a process of how to conduct investigations into high friction rock stabiliser pull test failure rates.

Mar 29, 2023

By H S. Welideniya, K Ekanayake

Carbon rich graphite deposits found in Sri Lanka are of inorganic origin and they occur as narrow steeply dipping veins. The vein width of these deposits varies from a few centimetres to metres, with an average vein thickness of 1 m. It is believed that the carbon rich fluid intruded into fracture zones under high pressure and the temperature has crystallised into graphite. This process has also generated many minuscule graphite veins centred on main mineralisation, which has reduced the host rock overall rock mass rating. These discontinuities could also be considered as joints with graphite infill. Graphite mining at Bogala Mines, Sri Lanka is a unique example of narrow vein mining faced with numerous wall stability issues resulting from weak graphite infilled joints.The Bogala vein type graphite orebody has an average vein width of 1 m. Overhand cut-and-fill mining method is used for stoping. The stope consists of two boundary winzes for material handling, access and ventilation. Mining is done on the strike of the vein and the average mining slice height is 2 m. A void to the height of one mined slice is left below the mining level during mining. Waste rock resulting from vein blasting and mine development is dropped into this void as fill rock. Any void height left greater than a single slice height will increase stresses in the stoping zone. This will increase formation of rock blocks and their effective separation from wall rocks. Split miniscule veins occurring around main mineralisation intensifies rock block formation in an environment of mining induced stress release. Consequently, any rock blocks sliding into stopes on graphite infilled joints will behave under constant normal stiffness (CNS) conditions and their shear strength will converge rapidly towards remoulded shear strength of graphite alone. Unlike with general infill deposited in joints, graphite has a very low friction angle and with its slippery nature, a small thickness may affect in great deal to reduce joint shear strength.In addition to immediate filling of excavated voids, controlled vein blasting and spot bolting have improved mine safety. Furthermore, rock bolts connected to each other through iron plates and anchored to anchor bolts installed in non-fractured solid rock are practiced to stabilise rock falls and wedge failures in narrow vein graphite mining. These practices have significantly contributed to safe mining without any rock bursts and catastrophic failures.CITATION:Welideniya, H S and Ekanayake, K, 2012. Challenges faced by narrow vein graphite mining and influence of graphite infill on wall stability, in Proceedings Narrow Vein Mining 2012 , pp 71-76 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 26, 2012

By W D. Erskine, M J. Saynor

Ranger Uranium Mine, surrounded by Kakadu National Park is located in the Alligator Rivers Region (ARR) in the wet-dry monsoonal tropics, 250 km east of Darwin, Northern Territory. A trial landform (TLF), measuring approximately 200 m by 400 m (8 ha) was constructed during late 2008 and early 2009 by Energy Resources of Australia (ERA), adjacent to the northwestern wall of the tailings storage facility at the Ranger Mine. A collaborative research program involving the Environmental Research Institute of the Supervising Scientist (eriss) and ERA, has been underway to measure long-term (five to ten year) rainfall, run-off, sediment and solute losses, seepage and vegetation establishment on the TLF (Saynor et al, 2012; Lowry et al, 2014). This paper reports on bed load losses from four erosion plots built on the TLF. These were achieved by physically isolating approximately 30 × 30 m areas from the surrounding landform on three sides with sidewalls along with an open polyvinyl chloride (PVC) pipe on the downslope side. Each plot was instrumented with a range of sensors. Two plots, 1 and 2 were constructed on waste rock material, while plots 3 and 4 were constructed on waste rock mixed with approximately 30 per cent lateritic material. Plots 1 and 4 had tube stock planted in early 2009, and plots 2 and 3 were initially planted by direct seeding in July 2009 and were infill planted with tube stock in January 2011.CITATION:Saynor, M J and Erskine, W D, 2016. Bed load losses from experimental plots on a rehabilitated uranium mine in northern Australia, in Proceedings Life-of-Mine 2016 Conference , pp 168–171 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 28, 2016

By B Small

CSA Mine, named after the nationalities of the mines first owners being Cornish, Scottish, Australian, is one of Australia’s deepest and highest-grade copper mines. A combination of the modified Avoca method and shrinking was a standard practice where production rings were fired against and bogged to waste fill. This method resulted in dilution and involved rehandling of waste material to refill the bogged stopes in the Western region and the QTS Central. The introduction of pre-charging blastholes using a wireless initiating system (WebGen™) has enabled CSA to design and fire stopes in reverse. This involves the use of a pre-charged Temporary Rib Pillar (TRP) against the waste fill material in a previously mined stope. The temporary pillar reduces dilution and improves conventional bogging as material is blasted towards the drawpoint rather than towards the waste fill material. Additional drill, prep and blast re-work is reduced or completely removed as rings can be charged prior to, in conjunction with, or after, charging of the slot. This is a safer practice as the charge crew spend no time working adjacent to a brow or open void. A TRP trial was conducted in the western orebody at CSA. The risk of exposure to a brow or open void while charging underground, and the bogger rehandling of waste material was successfully eliminated. Once reverse stoping was established in modified avoca stopes at CSA, two TRPs were implemented in challenging ground conditions in the QTS Central region. The TRPs were utilised in a retreating four-way intersection to extract a central diminishing pillar to close the level out. The ore recovered in this stope provided 5 per cent of CSA’s copper for the year. Without pre-charging with a wireless initiating system, two 5 m pillars totalling 6936 t of ore at 7 per cent either side of the stope would not have been recovered. CSA has implemented the practice of reverse stoping in modified avoca stopes using TRP for ten stopes since the successful trial. The development of additional blasting techniques enabled by precharging with wireless initiating systems is currently underway in the QTS North orebody at CSA Mine.

Mar 29, 2023

By Tumilty JA, Douglas WD

With a view to characterising activated carbons with respect to their ability to adsorb gold from cyanide solutions, an attempt was made to establish correlations between the equilibrium capacity constant of the activated carbon for gold (k value) and five other parameters. These parameters were percentage ash content, iodine number, methylene blue value, surface area, and pH respectively. Twenty four different proprietary brand activated carbon products were examined, the correlations being established by means of linear regression analysis. It was shown that a significant correlation only existed between k and pH value. This was interpreted in terms of the presence of oxide complexes on the activated carbon surface. Further test specifications as developed at the Anglo American Research Laboratories are also outlined with a view to the selection of activated carbons suitable for use in' a carbon-in-pulp operation. Methods for assessing both the capacity and the kinetic response of carbons for gold adsorption are presented, while the measurement of attrition resistance, particle size distribution and moisture and solubles content is also outlined. The measurement of the capacity constant for gold adsorption together with attrition resistance are considered to be necessary in the initial selection of a suitable carbon.

Jan 1, 1982

By J F. C L Costa, J C. Koppe, C Diedrich

Currently mineral resources evaluation is firstly expressed in terms of volume. Next the resources are transformed into tonnes by using an average bulk density applied to the whole deposit or, in some other circumstances, various different densities to distinct geological domains. Previous work in uranium, copper and nickel deposits used linear correlation between grades and bulk density to derive the density. Geophysics methods are also another good way to obtain density indirectly.These approaches do not take into account that two samples with the same grade but with different bulk densities can characterise different sample supports, leading to severe bias in grade and tonne estimates. It is important that the systematic measurements of bulk density are included into the estimation process, particularly when density is not correlated with the grades. Additionally, the extra time and costs involved in obtaining the bulk density data have made the mining industry disregard it.This study reports the impact in resource estimates by using a methodology that includes the bulk density variability. It involves the determination of a variability model for the bulk density and another model for the composite that was weighted by density as well as by its length. Next, these variables are estimated by kriging and the final grade estimates are obtained. The tonnages are obtained with the bulk density estimates. The conventional methodology was also applied and both estimates are compared to the reconciliation data from a major copper deposit. The results show a significant difference between the tonnage and grade estimates obtained from these two methodologies.The results show that considering density variability when compositing leads to a more accurate methodology resulting in better grade and tonnage estimates locally and globally.

Sep 26, 2011

By L A. Martinez Tipe, J McKibben, S Barry

"Metal prices are central to a mineral investment decision and have a profound impact on the financial performance of companies in the mining industry. At its most basic level, metal prices are influenced by the disequilibrium between international metal volume supply and demand. That is, the commodity price is a regulator that equates international metal supply and demand where inventories act as a buffer between the emergence of this disequilibrium and its dissolution through a price change.From this viewpoint, mining companies are required to forecast metal and bulk commodity prices when evaluating minerals projects through discounted cash flow (DCF) analysis. While supply-demand models are instructive, many companies use historical prices, spot prices, futures/forward pricing and consensus opinion (or a combination of all of these) to develop price forecasts.This paper discusses the definition, fundamentals and differences between futures/forward and consensus pricing and the appropriateness of each in a DCF model for mine project economic evaluation.Our results demonstrate that great care is required when using either futures/forward or consensus pricing and that the decision to use one rather than the other requires consideration of other important elements in the DCF model (ie the risk-adjusted discount rate).CITATION:Martinez Tipe, L A, McKibben, J and Barry, S, 2016. Metal prices and the differences between forward and consensus pricing – which one is better for use in a discounted cash flow model?, in Proceedings Project Evaluation 2016, pp 149–158 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By D Cork, R Peet

"Methane is a signifi cant greenhouse gas. Methane is released from coal seams during the mining process, both in surface and underground operations. This uncontrolled release is termed æfugitive emissionsÆ and in underground mines is transferred to the atmosphere in ventilation air, known as ventilation air methane (VAM). The global VAM emissions are over 369 Mt of CO2-e/a (Butler, 2011; US EPA, 2010). Coal mines may abate this VAM or consider abatement should there be a price on carbon. These fugitive emissions are heading towards ten per cent of AustraliaÆs emissions by 2020 (Department of Climate Change, 2011).There are safety issues involved in VAM destruction. Due to the nature of coal, and coalextraction, unpredicted methane spikes above the lower-explosive-limit (LEL) can occur in theventilation air; for example outburst events or ventilation failure. This could lead to a mine fi re ifthe mine ventilation air was connected to a VAM abatement plant and the risks are left unmitigated.CorkyÆs is working with the mining industry to directly address these safety issues. CorkyÆs calltheir technology VAM RAB, which stands for ventilation air methane regenerative afterburner(VAM RAB). VAM RAB utilises the principles of a thermal oxidiser.This paper discusses fi ve independent layers of protection (ILPs) to fi rst stop methane reachingthe thermal oxidiser and secondly prevent the fl ame front from burning away from the VAMRAB. The test work completed or contemplated includes safety integrity level (SIL) assignmentto various vents, barriers and door systems to prevent a fl ammable mixture reaching the VAMRAB, proven rapid response methane monitoring sourced from the petrochemical industry, anddefl agration experiments to test the eff ectiveness of the passive geometry and active engineeringcontrols. These experiments and SIL assessments indicate that it is possible to safely connect theVAM RAB to a working coal mine.Further work is required to upscale the pilot safety work. Refi nement to the design is possible;for example improvement to safety duct controls once trialled at demonstration scale plus thermalenergy capture for power generation or other applications to provide heat. Nevertheless, thispaper shows how safety issues can be managed by a multilayer of protection approach usingvents, barriers and doors, and appropriate thermal oxidiser geometry."

Jun 1, 2012

By I T. Ross

In the last 50 years there have been tremendous improvements in equipment used in development on underground mines. Compressed air powered jacklegs have been replaced by high powered electro-hydraulic drifters on sophisticated drill jumbos and load haul dump units have taken over from small shovels and rail cars.Despite the increase in power applied in key parts of the development cycle, average advance rates have fallen and costs have increased. The factor that is most often regarded as the main contributor to slowing advance rates is increased support requirements. Whilst there is no denying that the drive to improve safety has resulted in the requirement for more bolting and the installation of a skin support (mesh or shotcrete), this is not the only factor.Within the industry as a whole, many of the basic principles of development appear to have been lost over time. This paper looks at a historical case study and discusses some of the features of old development cycles that are no longer in use. This highlights some of the basics that appear to have been forgotten. These may be indicators to what may provide improvements in the conventional drill and blast development cycle.CITATION:Ross, I T, 2014. Development – back to basics, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 67–74 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014