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By M P. Sandy, V Louchnikov, V Eremenko, M Orunesu, O Watson

"Surface support is a key component of most ground support systems which are applied as a means of reducing the risk of rockfall occurrence and maintaining serviceability of mine development. Surface support has two primary functions: distributing the applied load throughout the system by providing connection between rock bolts retaining smaller rocks that unravel between rock bolts.Confinement is also provided to the rock mass, which may increase its capacity to resist further degradation or loosening.Surface support has been routinely installed in most underground mines around the world for many years, yet the operators are still not always certain which type is best suited for deformable ground conditions. A number of factors need to be taken into consideration prior to selecting the type of surface support liner, such as its capacity, availability, practicality of installation, manual handling and loading conditions (static or dynamic) to ensure the system’s optimal performance in terms of safety and costs.This paper presents an overview of the surface support types currently used in underground mines in Australia. Shotcrete, weld mesh, high-tensile chainlink mesh, reinforced weld mesh and high energy absorption (HEA) mesh are discussed. A detailed description of each system is provided in terms of capacity, energy absorption, installation cycle and costs. The cost of each system has been normalised to the units of energy absorbed. The information presented in this paper can be used by underground operators as a reference guide when deciding on suitable surface support type for deformable ground conditions.CITATION:Louchnikov, V, Sandy, M P, Watson, O, Orunesu, M and Eremenko, V, 2014. An overview of surface rock support for deformable ground conditions, in Proceedings 12th AusIMM Underground Operators’ Conference 2014, pp 57–66 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 24, 2014

By I Özen, G Chitumbura

A paper identifying anticipated challenges in the extraction of the N3500 Orebody at Glencore Xstrata Copper’s Mount Isa Copper Operations was written in 2007. The paper identified the depth of the deposit, the anticipated high stresses, talc alteration and issues with ground support as factors that may affect production. A number of strategies were planned to cope with these issues. After six years of extraction those strategies are evaluated and additional measures taken are outlined. Strategies to cope with the limited mine life are presented, and methods to overcome additional challenges encountered during production are described.CITATION:Chitumbura, G, Matthews, D, Özen, I and Jackson, S, 2014. One mile underground – mining the N3500 orebody at Mount Isa Mines, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 265–282 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

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

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

Mar 24, 2014

By B E. Hall, J Poniewierski, A J. Hills

Following a review of life-of-asset work at MMG’s Rosebery mine in which a proposal to increase production from 0.8 Mt/a to 1.2 Mt/a was mooted, questions were raised as to the seemingly arbitrary reasoning for the production increase. They included: Did the proposal actually add value? Was the production rate from the mine sustainable? Was it the best identifiable value-adding configuration?A multiple scenario analysis study commenced to model how various cut-off grade values against mill throughput, along with associated capital expenditure, would impact on the potential value of the operation. This was carried out utilising the ‘hill of value’ approach promoted by AMC Consultants.The traditional approach to these studies uses a constant cut-off value (NSR – net smelter return) for the whole mine analysis. This was thought to potentially introduce a large amount of variability to the model as different ore types were thought to have different processing costs and different underground lenses were thought to have different mining costs. With the aim of developing a consistent value measure that would be comparable between the lenses, an alternative measure was suggested for all ore block values in this model, irrespective of lens location.The new measure used was a mine variable value (MVV). This allowed for the variable factors of final metal concentrate production to be removed from the ore value. From a mining point this was the haulage distance from ore source to run-of-mine (ROM). For the mill it was the amount of reagents required, varying based on grade together with the amount of power required for crushing, grinding and tailings handling.From the initial run of the model, preliminary results lead to a conclusion of increased value being gained by either increasing cut-off grade or increased mill throughput rate. The validity of the result was tested using alternative price scenarios, which although changed the absolute value achieved did not significantly change the identified value-adding strategies. The realisation of these value increases identified are however contingent on higher sustainable development rates being able to be achieved mine-wide, higher backfilling rates being able to be achieved and overcoming recognised ventilation constraints. Further detailed studies are being put in place to identify the practicality of overcoming these recognised constraints and achieving the higher-value strategies identified.The ongoing objective of the hill of value model built is to keep the model alive as a life-of-asset planning tool and to continue refining the foundation data, such that various mining strategies can be analysed with increased confidence.CITATION:Hills, A J, Hall, B E and Poniewierski, J, 2014. Mining scenario analysis and mine variable value, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 111–118 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By J Kovacs

Underground development drill and blast practices have generally remained unchanged for decades. Historically, development optimisation has focused mainly on the application of perimeter control products in order to reduce overbreak, and efficiencies in development cycle times to increase advance rates. In today’s environment of ever increasing operating costs and lower ore grades, it has become even more critical to focus on the optimisation of drill and blast in development mining. It is a crucial component in ensuring mining schedules are maintained, but more importantly achieving the objective of getting to the orebody in the shortest possible time.This paper will focus on the development optimisation process used in a trial at Xstrata’s George Fisher Mine in North Queensland. Not only did the trial include the typical optimisation areas of charge loading practices, perimeter control methods, drill design standardisation and quality control processes, it also investigated the significance of delay accuracy with regards to perimeter control and drill design optimisation.One of the key measures of the trial was a comparative analysis between baseline development and development with the staged introduction of optimisation elements using 3D photogrammetry as a validation tool. A detailed cost analysis between the various stages was also a key measurable of the trial.The results of the work undertaken at George Fisher have shown that the accuracy of electronic detonators can produce significant benefits in development blasting. When combined with standard optimisation philosophies a substantial reduction in overbreak, ground support requirements and development cycle times can be achieved, whilst still being able to reduce the overall development operating costs.CITATION:Kovacs, J, 2014. Mine development optimisation – an evolutionary process, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 51–56 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By J Harris, J Li, M Shi, P A. Kirsch, D Sprott

Australia has an enviable global reputation as a leader in health and safety outcomes in the hazardous industry of mining, in part due to an increased application of risk management regarding operational planning and mining practice. In 2010, the Australian Coal Association Research Program (ACARP) initiated a major project, RISKGATE, to build, capture and deliver a national body of knowledge to assist companies with the management of risk in coal mining. The online system contains information for tyres, collisions, fires, isolation, strata underground, ground control open cut, explosions, explosives underground, explosives open cut, manual tasks, slips/trips/falls, outbursts, coal bumps and bursts, human–machine interface, tailings dams, occupational hygiene and inrush. In 2014, RISKGATE is developing content for fitness for work (fatigue, drugs, alcohol, physical well-being, psychosocial well-being).This paper analyses underground mining fatalities in Australia and the United States since 2002 and introduces RISKGATE with a focus on the knowledge base for management of hazards in underground operations. Through capturing coal industry expert operational knowledge, RISKGATE provides a cumulative corporate memory at a time of high personnel turnover within the mining industry. This coal-based body of knowledge is a platform that can be evaluated and implemented for other mining domains and high-risk industries within the energy supply chain.CITATION:Kirsch, P A, Shi, M, Harris, J, Li, J and Sprott, D, 2014. RISKGATE and underground operations, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 11–18 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By A Bennett, S Arndt

Minerals and Metals Group Pty Ltd (Golden Grove Mine) commissioned a comprehensive study of the life of mine using advanced 3D stress modelling. Essential input into the modelling was the proposed mining sequence, itself the culmination of years of planning and geotechnical experience gained in comparing actual performance to predictions. The modelling approach addressed the whole of mine at a scale of 5 km. There were however several magnitudes of scale to address, from the scale of laboratory testing (0.05 m), to drive size (5 m), extraction size (50 m) and stope size (500 m). Scaling was therefore an integral part of the procedure. The constitutive rock properties in terms of peak and post-peak behaviour varied with the scale. At the laboratory scale, the constitutive relationships for all the main rock units were determined in 3D as well as recovering the in situ stress using the rock memory technique, deformation rate analysis (DRA), whereas at the larger scale the properties were varied to match observations.The modelling, using non-linear analysis to account for rock mass yield and peak-residual strength, demonstrated that if the planned extraction sequence and stope geometry had been pursued there was a high probability of large volumes of excess shear stress being released. However, it was not possible to identify whether the excess shear stress (ESS) will release as one large event or many smaller events. The mining approach was to modify the extraction sequence/stope geometry to manage the hazard. This is where the economic advantage can be realised.Alternatively or in combination with further modelling, the volume identified with the predicted high ESS or energy release could be intensively explored for structures having adverse orientation to the stress change and with sufficient areal extent to release the ESS. The microseismic data can be interrogated for the likelihood of a large event to support the hypothesis of a singular release versus multiple releases of smaller events. If the initial planned extraction sequence had to proceed as designed, changes to dimensioned stopes would be recommended for risk mitigation. Mine planning could as an option, modify the sequence and the extraction could then be remodelled.CITATION:Bennett, A, Arndt, S and Dight, P M, 2014. The economic advantages of detailed geotechnical assessments for extraction sequencing, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 243–250 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By B Crowdey, J Wall

The particular underground operation relating to this paper is a gold producer and has an orebody predominately consisting of rhyolite and marcasites pyrite. The ore has been identified as being reactive to nitrates.Operations currently utilise two twin boom Tamrock™ Super-Drill Jumbos for development mining. Capital (waste) headings have a nominal size of 7 m wide × 6 m high, while ore headings are designed at 5 m wide × 5 m high, with the lower reaches of the mine narrowing down to a 4.5 m × 5 m design. Production drilling is currently undertaken with an Atlas-Copco L6 Solo. Due to the reactivity of the ore, inhibited emulsion phase bulk explosives are used for all blasting activities, with electronic detonators utilised in the production sequence.Sublevel shrinkage stoping is utilised when the orebody is stronger than the host rock. After blasting the ore, and during extraction, the surrounding host rock cannot maintain integrity and it ‘collapses’ on top of the ore. Gravity forces this collapsing rock to follow the blasted ore through the drawpoints.The highest risk involved with sublevel shrinkage stoping is the necessity for charge crews to work upon ‘rills’ at the drawpoints, which are occasionally unstable, to charge the next blast ring to be fired. To overcome this issue a loading technique called ‘precharging’ is used whereby two to three rings are charged, but only the first ring is fired. This technique can minimise risks by taking the charge crew away from the rill and brow.As this mine contains hot and reactive ground, strict guidelines from the Australian Explosives Industry and Safety Group Inc (AEISG) are adhered to. These guidelines relate to the sleep time for the precharged rings.This paper describes the issues faced with precharging in hot and reactive ground, the monitoring required and the systems and procedures that were put in place to enable a successful outcome.CITATION:Crowdey, B and Wall, J, 2014. Precharging sublevel shrinkage production rings in hot and reactive ground with electronic blasting systems, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 283–288 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

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

By C Vejrazka, A Wyllie

A common problem faced by underground mining operations around the world is airborne dust sourced from production and auxiliary processes. In most underground mining methods, air is generally directed through a limited number of drives and is thus more effective at carrying dust away from operational areas quickly. In a block caving operation, there are a number of ways to set up ventilation systems. In its simplest form the air is entering the extraction level through access drives and then split over the number of extraction drives—based on the pressure differences—without being regulated. This is the case at Northparkes Mines. Larger operations and panel caves most often utilise designated ventilation levels and ventilation raises with some form of regulation to ensure adequate airflow to all work areas.Northparkes Mines utilise electric loaders as a productive means to transport ore to the run-of-mine (ROM) bin, which have the added advantage of allowing the mine to require less airflow through the extraction drives as dilution of exhaust gases is no longer required. Dust was controlled during the operational period of the previous two block caves by using water spray bars in each of the drawpoints, which were seen as the main source of dust. These drawpoint spray bars only wet the surface of muck piles and require a lot of water to achieve an effect on the overall air quality. The drawpoint sprays can be remotely operated from mine control, whenever a loader is sent to bog a particular drawpoint. A significant amount of airborne dust is also generated during loading as rilling of material occurs, with further dust being generated during tramming and dumping activities as only limited mixing of the water and ore occurs. The water sprays are ineffective at precipitating dust that has already become airborne.When built, the currently operational E48 cave was fitted with drawpoint water sprays. As the cave is in a different orebody with higher silica content, the water sprays were not able to keep airborne dust and thus silica quantities under control efficiently. It was therefore necessary to find more sustainable solutions. The problem was further intensified by the concurrent load-haul-dump (LHD) processes across multiple extraction drives and tipping points, as well as secondary breaking, using non-explosive blasting and development activities occurring nearby.This paper describes the practical approach taken at Northparkes Mines to reduce personal dust exposure on the E48 extraction level. As the extraction level is one of the main contributors to airborne dust at the mine, it is believed that much can be learnt by focusing improvement efforts initially on this part of the mining process and later applying the effective solutions in other parts of the mine.CITATION:Vejrazka, C and Wyllie, A, 2014. Reduction of airborne dust on the E48 extraction level at Northparkes mines, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 227–232 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By M B. Lovitt

At present there exist few engineering calculations and principles to determine exclusion zones for personnel in the vicinity of underground explosive storage areas. This paper is written to encourage discussion and proposes a method to determine the impact of a blast wave through scientific principles. Then, criteria to determine an ‘exclusion zone’ need to be identified.There has been much work on identifying the pressure rise and the energy of explosive detonations underground but little to develop standards in criteria that can be used to identify exclusion zones for useful design of underground infrastructure. The author presents the newest ways of calculating the pressure rise and energy (impulse) of detonations and puts forward some criteria to be used in the identification of personnel exclusion zones to spark peer review and discussion of these criteria towards developing standards for engineering design.Smith and Sapko (2005) used Sach’s scaling laws to develop a relationship for both pressure rise and impulse, which incorporated factors for intersections. Intersections will have a large bearing on the peak pressure and the energy of the blast wave and needs to be incorporated. Silvestrini, Genova and Leon Trujillo (2009) developed the scaling laws with an energy concentration factor to predict blast propagation in partially confined geometries and together the author has proposed equations for determining the peak side pressure and the impulse in a tunnel. Given the weight and type of the explosive, the average cross-section of the tunnel complex and the number and type of intersections encountered, the pressure and impulse can be calculated.The military has carried out many investigations as to the influence of blast pressure and impulse on the human body. Lung damage, ear damage and velocity knock down are three of the criteria standards that are presented as possible methods of determining exclusion zones around an underground explosives magazine. The author suggests that the velocity curves presented by Bowen, Fletcher and Richmond (1968) could be an acceptable method of determining an exclusion zone as it incorporates the side pressure and the energy of a blast wave to determine the ability of the wave to knock down personnel.CITATION:Lovitt, M B, 2014. Exclusion zone calculation for underground explosives storage, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 319–326 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By T Purvis, J Calverd, A-K Greve

The Ernest Henry Mining (EHM) underground mine is situated adjacent to the Great Artesian Basin in North Queensland and within a fracture zone. Underground development commenced in February 2008 and has extended to approximately 1000 vertical metres below the surface.Groundwater has been encountered in numerous locations throughout the mine, predominantly on the eastern side of the orebody and often related to fault structures. Most of the groundwater encountered has been managed with conventional pumping systems; however, on occasions large water inflows have been encountered, which have had a significant impact on the planned development of the mine. Some of the intersections encountered had the potential to compromise the mine from a combination of high temperature (over 45°C), high pressure (3500 kPa) and high flows (50–100 L/s).This paper describes three case studies where significant water inflows were encountered and then controlled by the injection of cementitious grout into the rock mass. The case studies are the 1650 mid-shaft heading, the 1325 shaft access decline and the 1625 to 1475 fresh air rise (FAR) raise bore. Each case study includes a brief description of the event that occurred and the grouting equipment and techniques developed to enable mining to continue.The paper also describes the development of a hydrogeological map of the mine, the probe-hole regime used for high-risk headings and the required processes and systems to inject both cementitious grouts and polyurethane resin into the rock mass.CITATION:Purvis, T, Calverd, J and Greve, A-K, 2014. Management of high-pressure groundwater in underground development at Ernest Henry Mine, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 217–226 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By C Beal

Proper understanding and management of production drilling in sublevel cave (SLC) mines is essential for meeting production targets and for avoiding inefficiencies that could lead to costly bottlenecks. In many SLC mine management models drilling is considered a derivative activity of firing and bogging. Its scheduling is considered to be implied in a firing schedule and given an arbitrary target completion, ‘several weeks’ or ‘a month’, ahead of production firing. Like development, drilling activities are most efficient when conducted separately from other production activities. Unlike development, however, it is not apparent during planning and feasibility how far in advance of firing drilling activities ought to be conducted. This paper presents a method of defining a minimum threshold of drilled stocks above which drilling, firing and bogging operations occur most efficiently. As an example, the drilling activities, procedures, and maintenance at Glencore Xstrata’s Ernest Henry Mine are presented, and the larger scale consequences of improvements to these areas are presented to illustrate their importance in the operation of an SLC mine.CITATION:Beal, C, 2014. Understanding the role of production drilling and drilled stocks in efficient operation of a sublevel cave mine, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 235–242 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By J P. Sykes, A Trench

Mineral commodity prices are subject to volatility over both the short and longer run – driving fluctuating operating margins for mining companies over time. Such market dynamics influence mine operating decisions depending on the interplay of company-specific and mineral market factors. The company factors include corporate financial strength and price-risk management strategy, as well as mining practices and methods at the operation level. The critical mineral market factor is the price of the ex-mine product. This results from underlying microeconomic and macroeconomic root-causes pertaining to mineral supply and demand, in addition to the activities of financial participants within commodity markets.Significant changes in commodity prices of an extended duration, either up or down, remain the principal influence upon mine economics – and therefore on mining practices too. Accurate and precise forecasting of commodity prices (and of relevant exchange rates) continues to prove elusive both to mineral economists and market analysts. Thus, at any given point in time, there are often widely differing price forecasts available for both the relevant commodity and currency. Companies as a result are unclear how to respond in practical operating terms to such price-side uncertainty – other than to manage incremental capital and operating expenditure. Typical company – and industry financier – responses to the difficulty in predicting future mineral commodity prices are to take measures to manage uncertainties in mining project revenues and costs. These operating practices are prudent – yet come at an economic cost: that is, lost revenue and margin opportunity is inevitable in any such system not easily capable of responding to price volatility in the market. The mine of the future will seek to capture this lost economic opportunity. This paper presents a three-stage framework for analysing commodity price forecasts to aide understanding of the commodity price cycle and the periods of high-margin opportunity.CITATION:Trench, A and Sykes, J P, 2014. Perspectives on mineral commodity price cycles and their relevance to underground mining, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 19–32 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

An upgrade to Vale?s existing Clarabelle dual-pinion, semiautogenous (SAG) mill was completed by the power and automation technology company, ABB. The mill is used to reduce the size of the ore at the company?s nickel and copper mines in the Sudbury area. The project was completed on time and on budget with an excellent safety record. Located in Ontario about 400 km (250 miles) north of Toronto, Vale?s Sudbury operations are primarily underground nickel sulfide mines, with integrated mining, milling, smelting and refining operations. One of the characteristics of Clarabelle Mill is to accept nickel-copper ore from various sources and companies and to produce a nickel-copper concentrate and copper concentrate for further processing. The size of the ore is further reduced in SAG mills, rod mills and ball mills, where all milling is done wet.

Mar 1, 2014

By Ben Roderds

The Waneta Expansion Project (WAX) is located near the existing Waneta Dam site at the confluence of the Pend d?Oreille and Columbia rivers, approximately 10 km (6.2 miles) south of Trail, BC, Canada. The WAX owners consist of a partnership between Fortis Inc., Columbia Basin Trust and the Columbia Power Corp. This design-build project was awarded to SNC-Lavalin in late 2010 and is expected to be operational in mid-2015. The WAX consists of a new 335-MW powerhouse, including two new Francis turbine units, each generating approximately 167 MW. The powerhouse is located downstream of the existing Waneta Dam, and will make use of excess water that would otherwise be spilled during the runoff season. Water will be funneled to the turbines through an intake structure and two 10.5- m (34.5-ft) diameter, concrete-lined Penstock Tunnels. A 10-km (6.2-mile) transmission line will connect the powerhouse to existing electrical grids. Redpath/FKCI Waneta Tunnelers (RFK) was subcontracted by the heavy civil contractor, the Aecon/SNC Lavalin (ASL) joint venture, to excavate the two penstock tunnels and install the cast-in-place concrete liner. RFK is a joint venture partnership between J.S. Redpath Ltd., an underground mine contractor located in North Bay, Ontario, and Frontier-Kemper Constructors ULC, a heavy civil tunneling contractor based in Nova Scotia, Canada. Since penstock tunnel excavation was planned using drill-and-blast methods with a significant grade of more than 17 percent, the combined experience of these two contractors was a perfect match.

Mar 1, 2014

By K. V. Raj

As open pit mines continues to grow deeper, managing air pollution can become challenging. During arctic air inversion, the temperature of the air over the mine increases with altitude, and the cold, heavy, surface air is trapped in the pit. This problem is complex, and any solution will require an extensive understanding of the interactions between the aerodynamic movement of air, air inversion, meteorology, pollutant sources, and transport phenomena in the open pit mines. The availability of computational fluid dynamics (CFD) has elucidated the pollutant flow mechanism in open pit mines. Very few CFD studies have been published, especially the Large Eddy Simulation (LES), on the contaminant transport in deep open pit mines under arctic air inversion. The contaminant transport in an open pit mine under arctic inversion is due to a stable or nocturnal boundary layer in micro-meteorological terms. Understanding the physics of the stable boundary layer (SBL) is relevant to air flow and pollutant transport in deep open pit mines. This paper focuses on a Large Eddy Simulation (WMLES as Sub-Grid Scale Model) modeling of the SBL and pollutant transport in a large open pit mine.

Feb 23, 2014

By M. Tanaji

Over the last many decades, a significant amount of research has gone into the domain of slurry transport. However, design engineers still face many challenges with respect to prediction of pressure drop, critical velocity and other design parameters as a function of Solids % and Particle Size Distribution (PSD). The industry requirement is to transfer the slurry at the maximum concentration as possible (above 30% (volume %)) to make slurry transport more economically viable and to reduce water consumption. To facilitate the design and scale-up of slurry transport in pipelines and in process plants, there is a need for a correlation that can predict slurry pressure drops over a wide range of operating conditions and physical properties of different slurries. The objective of this study is to overcome the limited range of applicability and validity of existing correlations and to develop a generalized but more rigorous correlation applicable to a wider range of slurry sys-tems. The existing Wasp et al. (1977) method is based on multi-phase flow modeling approach. This study attempts to modify this approach by considering material-specific values of Durand?s equation co-efficient and by defining flow regimes based on particle Reynolds number. When compared with experimental data, the modified Wasp method proposed in this study predicts the pressure drop for slurry flows more accurately than other available correlations. Also, the pro-posed method requires minimal test/experimental data for a particular slurry system and can be extended over different input conditions. An iterative computer algorithm is developed to calculate the criti-cal settling velocity and pressure drop in a pipe as a function of Solids % and PSD. The solution method can easily be implemented in de-signing slurry pipes, design validation, and studying the different slurry transport scenarios. The modified method can also be extended to accurately predict pressure drops in dynamic pressure flow networks used in commercial process simulators.

Feb 23, 2014

By J. Kou

In this paper the activation behavior of quartz using Ca(OH)2 as an activator in quartz reverse flotation with anionic collector was evaluated and their adsorption behavior was investigated using a variety of techniques. The adsorption of the collector KS-I on the surface of quartz with and without Ca2+ was primarily characterized using Quartz Crystal Microbalance with Dissipation technique (QCM-D), which is a high sensitivity in-situ surface characterization technique. The flotation performance was evaluated using a laboratory mechanical flotation cell at different process parameters such as pH, collector dosage, and Ca(OH)2 dosage. The anionic collector evaluated was a commercial fatty acid KS-I. The QCM-D data showed that the collector adsorbed on quartz more readily with Ca2+ as activator. The mass and structural property of the collector layer adsorbed on quartz surface at 2.16×10-3 mol/L of Ca2+, 0.3 g/L of collector and pH 12.0 have been studied using the technique of QCM-D. The chemisorption of the collector KS-I on the surface of quartz in presence of Ca2+ were also demonstrated by means of zeta potential, QCM-D and FTIR analysis.

Feb 23, 2014

Passive seismic monitoring is a useful tool for measuring wave propagation in rock mass in deep underground mining. About 20000 mining ? induced seismic events of a hard rock mine, are identified and investigated. The travel times of source-receiver pairs allow us to demonstrate temporal change in velocity. It was found that the average P wave velocity is about 6000 m/s in the mining regions. Velocity investigations on the data during about three months around the main shocks show a significant decrease of the average seismic wave velocity after the main shocks. The sudden drop of average velocity change after the main shocks implies the creation of new fractures in the rock mass due to the main shocks. Afterward, the average velocity increased to a normal level after a time duration of several days. A reasonable explanation for the post seismic velocity increase is that opened cracks begin to close due to crack healing.

Feb 23, 2014