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"Hecla Limited has been investigating the potential benefits of using a Spendrup/Compact Filter Technic (CFT) diesel filtration system to remove diesel particulate matter (DPM) from the mine atmosphere to optimize use of the available air. The National Institute for Occupational Safety and Health (NIOSH) has evaluated the scrubber efficiency for dust and DPM which is to be presented in another report. This paper reports on a new mining application of a carbon particulate sensor’s ability to continuously monitor long-term mine DPM levels in conjunction with the scrubber evaluation. Magee Scientific is the manufacturer of the Aethalometer®, a continuous black carbon monitor which has been used extensively in worldwide ambient air quality monitoring networks. Two Aethalometers were used at the mine to assess their potential to continuously track underground DPM levels. In this first underground mine application, the Aethalometers were used to continually monitor black carbon (BC) in the intake and exhaust of the CFT scrubber system. Results showed that an off-the-shelf Model AE33 Aethalometer could continually track underground BC levels for extended periods of time over the course of the six-week trial. Data at the scrubber intake over 14 days of testing showed that DPM filtration would be required less than half of the time to maintain BC levels below 50 µg/m3. If the Aethalometer BC data output were “smart coupled” to control a 200-HP in-mine scrubber, significant electrical cost savings would be realized. INTRODUCTION Mine ventilation airflow quantities are calculated to maintain a safe and healthy environment for miners. Dilution of diesel particulate matter (DPM) is often a key factor used to determine the required quantities of ventilation air (Bugarski et al. 2011). DPM emissions and exposure levels are highly variable depending on mine schedules and activities (Janisko & Noll 2010); however, lacking technology to continuously monitor levels, mine ventilation systems are often over-designed in order to handle worst-case scenarios to ensure that DPM levels are maintained at or below regulatory limits. As mines go deeper it becomes more difficult to supply air, and the development of new shafts for air is not economically feasible (Basu et al. 2008). Cleaning and reuse of available air then becomes an attractive option to maintain good air quality."

Jan 1, 2016

By L. Xingwana

"The purpose of this study is to understand how ore loss and dilution affect the mine call factor, with the aim of subsequently improving the quality of ore mined and fed to the mill. It was necessary to know the relationship between ore mined and delivered for processing in order for the mine to operate efficiently. A review of the literature suggested the mine-to-mill integration approach as a standard intervention to address declining mine call factors. A case study was undertaken at a gold mine in the Free State on the influence of the fragment size distribution on ‘free gold’, the effect of ore loss and dilution on the mill efficiency, and time tracking of ore movement from source to the plant. The study revealed that fragmentation, underground accumulation of ore, and dilution have a significant influence on the mine call factor and mine output. It was difficult to quantify the total ore loss due to the time lag between mining and milling, which is extremely variable, when estimating the mine process inventory. However, it was possible to estimate the total dilution with some degree of precision by analysing the daily blast and head grades. It is therefore recommended that mines accept that some gold loss and dilution will always occur during blasting and transport of broken ore, and this loss must be minimized.IntroductionThe mining of high gold grades is obviously very beneficial for the economic viability of an operation, but optimum recovery of the product is required to maximize profits (Chapman and Bindoff, 2010, p. 13). Although integrating these two requirements is very difficult, due to the various activities occurring between mining and processing, it is indispensable. The main driver of this integration is the impact that different ore types and characteristics have on head grade. According to Isokangas et al. (2012, p. 871), mining and processing operations involve a series of well-defined sequential stages that are complementary in an efficiently operating mine, with the performance of one operation affecting that of another. An attempt to optimize each function separately without considering the whole extraction process will not result in the optimum efficiency of the entire mine. Several researchers have developed a proven methodology for revealing these inefficiencies (Dance et al., 2006, p. 1)."

Jan 1, 2016

By Per Hedvall, Jan-Eric Ståhl

"The main objective of this study is to introduce the new Sandvik’s production cost-model within Crushing & Screening (C&S) technology. So far, there have been different ways of determining different production costs in the industry, however, the previously published and presented methods have not been very efficient in calculating the costs per product and metric ton. Accordingly, the current study has been planned and started in collaboration with Lund University to provide an accurate and easy to implement cost estimation program to determine the costs for crushing and screening plants worldwide.The proposed model has been adapted to fit the C&S processes of different rocks/ores fractions. The result is presented as cost per ton of the end product(s).Within this paper the cost estimation results for 3 different crushing and screening plants are presented based on the information obtained from the following sources: a) documents at Sandvik SRP, b) Sandvik’s crushing & screening internal/external machines/plants data and information, c) literature studies, including the use of Sandvik’s internal/external simulation and software programs, d), visiting the crushing and screening plants ,e) interviews, information systems’ data, and, f) testing and observations by ourselves.In order to improve the level of accuracy and the validity of the attained results, simplistic Monte-Carlo simulations have been used, where needed. The achieved results, so far, have proven the reliability of the model. The calculated cost estimations have shown an accuracy over 80% when they compared with real cost data obtained from plant practices.The current cost calculation model, will offer the possibility of:•,Calculating the production costs in proposed and/or existing C&S plant.•,Making detailed cost analyses for the end product(s) in proposed and/or existing C&S plants.•,Determining the profitability of each finished product when it’s required.•,Cost optimization by potentially cutting the cost for part of C&S plant with an existing plant practice"

Jan 1, 2016

By S. Amirshenava, S. Nadimi

"Premature and unplanned mine closure can result in significant adverse impacts on the environment and community and need to be managed appropriately. Risk management techniques can reduce the negative impacts and risks of early mine closure. Risk Assessment 3D Model, is an efficient tool to quantify risk. The aim of this paper is to identify the mine closure risks in Choghart iron ore mine, and calculate the overall mine closure risk level. The result shows that 51.4% of assessed risks have low risk level and the overall risk level of early mine closure is low. Keywords: mine closure, risk assessment 3D model, Choghart iron ore mine INTRODUCTION The modern mining life-cycle can be divided into three main stages; exploration, mining (opening and production) and post mining (mine closure, and reclamation). Mine closure is the critical component of the modern mining life-cycle and it refers to the period of time when the operational stage of a mine has permanently ceases (Heikkinen et al, 2008). Planned closure occurs due to the depletion of ore reserves, but sometimes it is premature. Laurance (2006) shows that only 25% of mine closure in the past 30 years was due to resource depletion and in 75% of the cases the mines were closed for the other reasons. Ghanbari et al (2010) demonstrated that 48.6 % of copper mine closure in the past 100 years was due to resource depletion and 51.4 % of copper mines have been closed due to other reasons. The main reasons of premature mine closure are as follow (Laurance, 2006; Ghanbari et al, 2010; Maiti, 2013): • Economic, (low commodity prices or high costs) • Geological, (over-estimation of grade and tonnage) • Geotechnical (adverse geotechnical conditions) • Technical (equipment failure) • Regulatory, due to safety or environmental breaches • Social or community pressures • Closure of downstream industry or markets • Flooding or inrush Minaei Mobtaker and Osanloo (2013) developed a mine closure prediction model based on the data collected from closed iron ore mines in the past 100 years. They showed that economic factors, exhaustion of ore reserve and geological conditions are the main reasons of iron ore mine closure. They treated mine closure as a decision making problem in order to predict the causes of mine closure, including the uncertainties of some mine plan parameter such as iron ore price, production and consumption."

Jan 1, 2016

By Matthias Neuenschwander, Andreas Theiler, Patrick Steinle, John Singer, Martin Vogel, Diego Martini, Ömer Ündül, Markus Weh

"Quartz Dust is a major health hazard in underground works, wherever the ground contains quartz and depending on the chosen method of excavation. A research project initiated by the Swiss Tunnelling Society and realized by the Swiss Federal Institut of Technology, Zurich, established a correlation between the potential of quartz dust generation during excavation and well known geotechnical characteristics – the Cerchar Abrasivity Index CAI and the Uniaxial Compressive Strength UCS - of ten typical Swiss rocks. The results are at the center of a new guideline by the Swiss Society of Engineers and Architects on the requirements for early project stages regarding ventilation and logistics in function of the quartz dust related risk of the expected ground type of the project. The approach presented could be applied worldwide, contributing to more effective health protection as well as to better cost stability in underground works. CONTEXT Respirable quartz dust (aerodynamic diameter of particles < 10 micron) is universally recognized as a major workplace health hazard. Chronic bronchitis, silicosis and lung cancer can be caused by excessive exposure to quartz dust, as has been proved by many medical studies. Therefore, the legislations of many nations specify a maximum workplace concentration (Maximale Arbeitsplatzkonzentration MAK, Occupational Exposure Level OEL). In Switzerland, for example, the Threshold Limit Value TLV for respirable Quartz or Crystalline Silica imposed by the Federal Committee on Exposure Levels (Grenzwertkommission) is 0.15mg/m3 air. Some nations refrain from establishing a OEL because of the carcinogenic nature of quartz."

Jan 1, 2016

By Morgan M. Sears, lhsan B. Tulu, Ted M. Klemetti

"Why do some room and pillar retreat panels encounter abnormal conditions? What factors deserve the most consideration during the planning and execution phases of mining and what can be done to mitigate those abnormal conditions when they are encountered? To help answer these questions, and to determine some of the relevant factors influencing the conditions of room and pillar (R&P) retreat mining entries, four consecutive R&P retreat panels were evaluated. This evaluation was intended to reinforce the influence of topographic changes, depth of cover, multiple-seam interactions, geological conditions, and mining geometry.This paper details observations made in four consecutive R&P retreat panels and the data collected from an instrumentation site during retreat mining. The primary focus was on the differences observed among the four panels and within the panels themselves. The instrumentation study was initially planned to evaluate the interactions between primary and secondary support, but produced rather interesting results relating to the loading encountered under the current mining conditions. In addition to the observation and instrumentation, numerical modeling was performed to evaluate the stress conditions. Both the LaModel 3.0 and Rocscience Phase2 programs were used to evaluate these four panels. The results of both models indicated a drastic reduction in the vertical stresses experienced in these panels due to the full extraction mining in overlying seams when compared to the full overburden load. Both models showed a higher level of stress associated with the outside entries of the panels. These results agree quite well with the observations and instrumentation studies performed at the mine.These efforts provided two overarching conclusions concerning R&P retreat mine planning and execution. The first was that there are four areas that should not be overlooked during R&P retreat mining: topographic relief, multiple-seam stress relief, stress concentrations near the gob edge, and geologic changes in the immediate roof. The second is that in order to successfully retreat an R&P panel, a three-phased approach to the design and analysis of the panel should be conducted: the planning phase, evaluation phase, and monitoring phase."

Jan 1, 2016

By Magnus Evertsson, Johannes Quist

HPGRs have since introduced in the 1980s (Kellerwessel, 1993; Schönert, 1988), now become integral comminution units at several minerals processing plants around the world. Even though a lot of the initial issues concerning performance, availability and wear have been addressed and solved there are still phenomena and aspects which can be improved and understood further. Schönert and Sander (2002) presented experimental results and a mechanistic model including the internal force dynamics between the rollers and the compacted bed. They derived expressions for the pressure distribution involving most of the influencing terms and effects. However, in order to solve the differential equations several of the terms vanish by assumptions. The Schönert and Sander model reflects the experimental data in many aspects but the predicted shear force distinctly deviates from the experimental results at the angular position of the peak pressure level. This finding suggests that the particle to particle interaction and breakage process in the different zones seen in Figure 1 are complex and difficult to model using analytical models. A numerical modelling approach using DEM does not require the same amount of assumptions and simplifications of boundary conditions. Hence such modelling may provide further knowledge and vital understanding. In this paper the breakage process is simulated in order to investigate the shear slip behaviour and how it relates to different operating variables. A laboratory scale HPGR corresponding to the experimental setup used by Schönert (2002) have been modelled in EDEM (DEM-Solutions). The bonded particle model (BPM) is used for modelling quartz particles (Potyondy, 2004). The breakage model strength parameters are calibrated against interparticle breakage experiments in an Instron 400RD compression device. The contact model parameters are calibrated using calibration methods developed by Quist (2015). The floating roller dynamics are modelled using a spring-damper system. Results present insights regarding the interparticle slip behaviour as well as the frictional tangential force distribution on the rollers. The simulated roller pressure distribution shows good correspondence with previously reported findings in the literature.

Jan 1, 2016

By Michael Murphy, Khaled M. Mohamed, Berk Tulu

"Failures of coal ribs are major hazards in underground coal mines. For the period of 2010 to 2014, rib falls were the leading cause of groundfall fatalities and accounted for 35% of all groundfall-related fatalities. Currently, the National Institute for Occupational Safety and Health (NIOSH) is conducting a research project to develop an engineering-based design methodology for rib control in underground coal mines to reduce rib fall accidents and fatalities. Numerical models can be used as part of the design methodology provided the models are appropriately calibrated and verified.This paper presents a coal rib model to simulate the mechanism of coal deformation and fracturing that can lead to stress-driven rib falls. The proposed coal rib model was calibrated against published data from an instrumented rib site at the West Cliff Mine in Australia. Fracturing of the coal rib was defined by critical VALUES (of plastic shear strain of the coal material and rib displacement. A critical plastic shear strain of 2.8% and a sudden increase of rib displacement of 30 to 45 mm were sufficient to develop a fracture plane in the rib model. Accordingly, a discontinuous rib deformation was achieved and it compared favorably against field monitoring data.INTRODUCTIONCoal ribs are the vertical walls of coal formed when entries are developed in a coal bed. Failures of coal ribs are a major hazard in underground coal mines. Furthermore, failing ribs can contribute indirectly to roof and floor instability by increasing opening widths across intersections and entryways. Over the last 18 years, there has been a continual occurrence of injuries and fatalities in underground coal mines due to rib falls. For the period of 2010 to 2014, rib falls (excluding longwall face failure and bursts) were the leading cause of ground fall fatalities and accounted for 35% of all groundfall-related fatalities (MSHA, 2015)."

Jan 1, 2016

By V. van der Merwe, C. W. Siyasiya

"A petrochemical company experiences premature thermal fatigue failure of the casting moulds used in catalyst production. The aim of the project was to find an alternative alloy that would outperform the current low-alloy cast steel used for the moulds. Based on their thermo-fatigue properties, 3CR12 ferritic stainless steel and H11 tool steel were chosen for testing and comparison with the currently used BS3100 B7 cast steel. Samples of each material were subjected to temperature cycling in a Gleeble 1500TM thermo-mechanical processing simulator, followed by surface analyses. The main parameters derived from the test work were the total true strain, the hot strength of the materials, and the number of cycles to failure. Additionally, the coefficient of thermal expansion for each material was measured using a Bähr dilatometer. H11 tool steel yielded the best performance by way of having the fewest surface cracks, the lowest total true strain per cycle, the most cycles to failure, the highest hot strength, and the lowest coefficient of thermal expansion.IntroductionThe Catalyst Manufacturing Plant (CMP) is a dusty environment with temperatures higher than ambient. Molten mill scale (magnetite, Fe3O4) is tapped from the electric arc furnace into the casting moulds at temperatures close to 1600°C. After 10 seconds the magnetite is quenched with a water spray for 50 seconds to a minimum temperature of 400°C. Sometimes there is residual water in the pans from the cooling fog process (water sprays) in the empty mould. In a typical plant, the factory produces 40–44 batches a day containing between 150–300 moulds, depending on the time required for the specific batch (Buchholz, 2014). The typical batch in the CMP usually weighs about 6 t. The thickness of the tap, speed of the casting belt, and tilting angle all affect the number of moulds per batch.One of the main problems faced during the molten mill scale casting is the premature failure of the casting moulds. Cracks initiate on the surface of the mould, which is in contact with the hot catalyst, and propagate into the interior of the mould. Cracks are generally concentrated in the middle area of the mould where the catalyst makes first contact with the mould, and which experiences the highest temperature fluctuations. This suggests that the cracks are caused by thermal fatigue during the casting process."

Jan 1, 2016

"A carefully planned laboratory test program is the foundation for the successful development of a sulfide flotation process. In this paper, the development of a copper-molybdenum flotation process is described. However, the concepts discussed can be applied to most sulfide flotation systems and to many industrial minerals projects as well. Each step is examined, beginning with sample procurement and covering sample preparation, flotation test preparation grinding, scoping flotation tests for reagent scheme selection, rougher flotation parameter evaluation and cleaner circuit evaluation, through to locked-cycle testing and pilot plant design and operation. Ore variability and modeling to determine the response of various ore types to the optimum conditions established is also examined.IntroductionA carefully planned laboratory test program is essential for the successful development of any sulfide flotation process. These programs usually begin with a series of batch tests on several ore type composites to evaluate appropriate reagent schemes and delineate optimum rougher flotation conditions, such as primary grind size, flotation pulp density, flotation pulp pH and retention time. Cleaner parameters such as regrind size requirements, cleaner pH, pulp density and number of stages are then defined. Semicontinuous locked-cycle tests are performed using the parameters established in the batch tests to determine the effect of middling streams recirculation on flotation performance. The data generated and the experience gained in the locked-cycle tests may be used to design and operate a continuous pilot plant if relatively large amounts of concentrate are required for further process development, such as optimization of byproduct molybdenum production from bulk copper-molybdenum concentrate.A comprehensive ore variability program is undertaken to determine the response of various ore types to the optimum conditions established. These ore variability programs often test hundreds of samples in standardized batch rougher and cleaner tests to ensure that the orebody is well understood.This article describes the development of a copper-molybdenum flotation process, but the concepts discussed can be applied to most sulfide flotation systems and many industrial minerals projects. The block flow diagram in Fig. 1 provides an overview of the development process."

Jan 1, 2016

By Ruan Chi, Zhenyue Zhang, Zhengyan He

In order to understand what governs the leaching of exchangeable rare earth (RE) and impurities and enhance the RE leaching process, the effects of the concentration of ammonium sulfate, pH and leaching temperature on the migration of RE and aluminum (Al) during the leaching of weathered crust elution-deposited rare earth ore was studied. The elution curves for ammonium ion (NH4 +) were also studied to reduce the consumption of the leaching agent. It was found that a high concentration of leaching agent can enhance the leaching process of RE and Al, reduce the breakthrough volume of NH4 + and more rapidly obtain the equilibrium of NH4 +. However, excessive leaching agent can both increase the cost and the burden on the environment. There is almost no variation on the elution curves of RE and Al in the pH range of 4-8 and control of the pH of the leaching agent can be omitted. It is better for the leaching work to be done in summer since higher temperatures can enhance the RE leaching process and improve the use ratio of ammonium. Using 0.1 mol/L ammonium sulfate for leaching 250 g RE ore, 95.42% RE extraction can be obtained when the volume of leachate is 102.6 mL. At this moment, only 18.17% of NH4 + is used to exchange RE and 49.22% of NH4 + is left in the leachate that can be reused as leaching agent by replenishing ammonium sulfate. The results of our studies can provide theoretical guidance for high efficiency and low consumption exploitation of RE ores.

Jan 1, 2016

By D. Saramak, T. Gawenda, Z. Naziemiec, A. Saramak

"Crushing and grinding operations are burdensome to the environment due to generating large amounts of dust, noise and waste during course of comminution operations. An indirect negative contribution of industrial crushing processes is an increase of carbon dioxide emissions, which is a byproduct in energy production processes. On the other hand comminution processes in ore enrichment technology are essential due to liberation the useful minerals from the gangue, and producing the demanded particle size fractions of products. The article concerns investigation over negative influence of comminution devices on environment in terms of dust and noise emissions. Several various crushing devices were under investigations, both in laboratory and in industrial scale.The obtained results were analyzed in terms throughput and energy consumption achieved in various crushing devices. The test results allow to conclude that an intensity of dust generation for individual types of crushing devices is varied. An additional aim of this article is to link the aspect of mineral processing with its environmental impact and hence an influence of raw materials beneficiation industry on the quality of life of residents in immediate and close vicinity.INTRODUCTIONMineral raw materials beneficiation processes are an integral part of multi-stage processing of coal, iron ore and non-ferrous metals, production of aggregates, cement, or precious metals and stones. One of the common features of all beneficiation operations is their high energy consumption, which is mainly connected with various operations of feed particle size reduction. Comminution processes in ore enrichment technology are essential to expose and liberate the useful minerals from the gangue, while in aggregate production processes – to obtain the demanded particle size fractions of products. Crushing operations are characterized by a high degree of negative impact to the environment and humans, mainly due to:• High electricity consumption, which indirectly contributes to the increased emissions of carbon dioxide into the atmosphere, since CO2 is often one of the byproducts of energy production;• Relatively rapid wear of device working parts, what forces a production of spare parts and involvement of steel industry;• Intense dust pollution generated during crushing and grinding processes. This is one of the factors affecting the air quality and the consequent deterioration of the living conditions of inhabitants near mines and processing plants; and • Excessive noise emissions."

Jan 1, 2016

By Anita Wu

"GENERAL BACKGROUND/CONTEXT The £300 million Bond Street Station Upgrade project will provide essential congestion relief and step-free access to an important interchange and destination station in the heart of the London Underground network. In 2010 London Underground (LU) let a design & build contract to Costain Laing O’Rourke (CoLOR), with the works to be completed in early 2017. CoLOR’s lead designer is Halcrow-Atkins JV (HAT). The station upgrade works comprise the following principal elements: •,A new satellite ticket hall on the North side of Oxford Street •,Two additional escalators and stairs providing congestion relief to the Jubilee line platforms •,Four lift shafts to provide step-free access between the different levels of the station •,Additional stairs to Central line platform level •,Tunnelled connection to the Crossrail station immediately to the south •,Connecting passageways, including ten connections to existing station tunnels. To achieve this, an existing property at 354-358 Oxford Street has been purchased and demolished, and the basement deepened to make provision for the new ticket hall. From this basement two circular shafts have been sunk, providing construction access for the tunnelled passages and lift and escalator shafts. The tunneling is principally in the London Clay stratum, which is widely recognized as a favorable tunneling medium, although the deepest parts of the works extend into the Lambeth Group strata which consist principally of stiff clays at this level but also contain water-bearing sands. The majority of tunneling has been carried out using mechanical excavation and sprayed concrete linings but other techniques have been used in areas where SCL is not practicable.A key challenge for design and construction of the works has been spatial constraint, in particular due to: •,The existing underground station layout •,The Limits of Deviation for the works •,The small construction site at 354-358 Oxford Street •,Overlaying Third Party assets including Listed buildings, sewers and water mains."

Jan 1, 2016

By P. Radziszewski

"The comminution energy curve is a profoundly simple graphical technique that displays detailed operating efficiency information for industrial comminution processes. The database underlying the methodology has recently more than doubled, to now contain a substantial fraction of the international comminution output (1.7 billion tonnes of rock per year) across more than 8 major commodities. This increase has deepened the analytical capabilities and effectiveness for assessing comminution circuit performance in relation to comparable operating mines. This paper will explore how the energy curve methodology can be used to examine the ability of stirred mill technologies to increase efficiency.Specifically, this paper explores how the energy curves methodology can be used to present energy efficiency gains achieved at three mines employing a stirred mill technology under different duties. The first of these cases studies was conducted by Rosa (2014) who compared the performance of two ball mills operating in parallel with a Vertimill in the regrinding circuit at the Samarco Iron project in South East Brazil. The Vertimill was found to achieve the same product specification while consuming 40% less energy. Brissette (2010) investigated the impact of grinding media on the performance of both ball mills and Vertimills. The energy savings achieved varied from 10% to 40% for the ball mills and 30% to 60% for the Vertimills. Lastly, Bergerman (2012) investigated the operating conditions of the Sossego regrind circuit from 2006 to 2011 against design specifications and laboratory testing. The operating work index of this circuit varied from 5.2 kWh/t to 44 kWh/t over this 5 year period. The full suite of energy curves"

Jan 1, 2016

By S. Tuylu, D. Adiguzel, A. Bascetin, O. Ozdemir, U. Cinar

"In recent years, use of mineral processing tailings as a paste backfill has been significantly increased due to the environmental problems occurred during the disposal of the tailings in mine area. The paste material used to fill the stopes excavated is generally consisted of tailings (solid ratio: 70–85%) and cement (3-7%w/w). Meanwhile, the strength of the paste material is completely related to ratios of these materials. Cemented Paste Backfill (CPB) system is used as support depending on the roof pressure and also as an underground disposal method. Therefore, temporary strength value of the paste material gains a great importance. For example, the paste backfill material must compensate the minimum 4 MPa strength value in order to procure the permanent fortification, and the mixture cost must be minimal as well. For this reason, in this study, the Pb-Zn tailings were used to prepare for the paste backfill material using cement in order to perform the uniaxial compressive strength (UCS) tests. Additionally, a statistical experimental design and linear programming techniques was used to find the optimum mixture in terms of strength and cost. The pulp solid ratios (PSR) of the tailings and cement (by the solid) were varied as 75-85% and 3-8%, respectively. The results of this study; the optimum mixture is ~79.5%PSR and 6.8% cement ratio (by the solid) for 4 MPa UCS (28 day) and 16-17 cm slump value determined by this method with statistically reviewed experimental data. INTRODUCTION Fine-grained tailing waste materials resulting from mining activities are an important environmental problem. In terms of environmental protection, the use of tailings with paste backfill is a very attractive approach and a desirable solution to tailings management. Paste backfill, which is used to fill underground openings, is made by primarly mine tailings mixed with hydraulic binders and water. Its main benefits are not only disposal of mine tailings in underground, but also providing a certain strength to ensure a safe working environment. Therefore, it is necessary to use hydraulic binders due to durability of the system (Benzaazoua et al., 2004). Paste backfill, formed from the mix of dewatered plant tailings (75-85 wt. % solids), water and binder (3-8 wt. %) (Brackebusch, 1994; Landriault et al., 2005).Water is used to determine the chemical reactions between tailing and binder materials for strength as well as increasing pumpability. Portland cement and ash can be used as binder materials (Douglas and Malhotra, 1987). The factors which affect strength and stability of paste backfill separate into two groups called interior and exterior factors. Interior and exterior factors determine the strength, stability and deformation features of backfill in short and long terms and also affect the transportation of backfill to underground (Benzaazoua et al., 2004; Erçikdi et al., 2006; Yilmaz et al., 2010; Nasir and Fall, 2010). Material properties of tailings as well as site operations to fill the underground openings are the most important steps for the design purposes. Grain size distribution, chemical, and physical properties of tailings, and mechanical properties of paste fill material are used to understand the suitability of tailings as paste fill material (Ozturk et al., 2014). In addition, Fig. 1. is also proposed to apply a typical design for the operational stages of paste fill applications on the site."

Jan 1, 2016

By Peter Johnson, David Barber, Lachlan Henderson

"The traditional approach to rail tunnel fire safety design is to follow NFPA 130, which assumes a major tunnel fire could occur and requires bi-directional emergency ventilation. There is no regard to the impact of ventilation on fire growth and smoke production. In Europe and other parts of the world, the need for tunnel ventilation systems to provide smoke management in the event of fire is being questioned. The European Commission’s Safety in Railway Tunnels TSI is based on a philosophy that improved running capability (preventing fires and other events stopping the train in the tunnel) and elevated fire performance of materials means the chances of a major fire in a tunnel is highly unlikely, with only small fires expected, and no tunnel ventilation required. This approach is reinforced by the International Tunneling Association which has recently published an Engineering Methodology for Performance Based Fire Safety Design on Underground Rail Systems, which includes a methodology that does not require assessment of tenability in tunnels, but requires a scenario based risk assessment to be undertaken to determine overall fire safety measures. Recent research by the University of Edinburgh has shown that forced airflow from a mechanical ventilation system may promote fire growth and increase smoke volumes much more than with natural ventilation. Whilst the concept of no mechanical ventilation (i.e. natural ventilation) is a new concept for some, for many rail tunnels it is worthy of serious further consideration. RAIL TUNNEL SMOKE MANAGEMENT AND STANDARDS Smoke management using mechanical bi-directional ventilation systems has been a central part of the fire safety design solution under NFPA 130 for rail tunnels since 1982. The current version of the standard, NFPA 130: 2014, states that mechanical ventilation should be fitted to rail tunnels of length greater than 305 meters and an engineering analysis should be undertaken for tunnels of length 61 to 305 meters to determine whether mechanical ventilation should be fitted. However, this approach is not necessarily reflected elsewhere in the world. In Europe in particular, mechanical ventilation systems for smoke control within rail tunnels have taken a much less central role within the standards framework."

Jan 1, 2016

By K. Loutet, A. Guenkel, K. Nikolaisen

"The INCO SO2/air process was patented by INCO in 1985. Well over 80 mining operations worldwide are using this process to oxidize cyanides to cyanates in the tailings from gold, silver and copper mining operations. The sulphur dioxide (SO2) for this process can be supplied either as liquid SO2, SO2 gas or as a sulphite-based salt. The majority of SO2/air operations have used sulphite-based salts as the ‘SO2’ equivalent probably due to low capital cost. Liquid SO2 has also been used frequently, but due to concerns with safety and availability, this option is often not preferred. The solid salt sodium meta-bisulphite (SMBS) is commonly employed, but its cost is an order of magnitude higher than that of sulphur. Due to the high cost of SMBS, the increased capital cost for a sulphur-burning SO2 gas plant (green field or retrofit) usually has a pay-back of less than three years. In addition, a SO2 gas plant offers potential for significant heat recovery from the sulphur burning process. In terms of safety, the SO2 gas plant only requires the storage of relatively benign sulphur and the site inventory of SO2 will be limited to the plant gas volume. Therefore, a SO2 gas plant will not trigger the Canadian Environmental Protection Act (CEPA) E2 Regulations or the US Clean Air Act Section 112(r) accident prevention regulations, as may be the case for liquid SO2.Most existing mining operations using ‘SO2’ are gold operations, but in recent years there has been an increased interest in larger sulphur dioxide gas plants for hydrometallurgical facilities that employ leaching of cobalt/manganese minerals or precipitation of manganese/iron (e.g. Tenke Fungurume and El Boleo).This paper focuses on the production of sulphur dioxide gas generated on site and on demand for the use in hydrometallurgical processes. The capital and operational costs of a SO2 gas plant is compared to conventional process alternatives using sulphite-based salts or liquid SO2."

Jan 1, 2016

"Guideline Augered Cast-In-Place (ACIP) Pile Specification and Commentary11.0 GENERAL1.1 SCOPEThe Pile Contractor shall furnish all labor, materials, tools and equipment necessary for furnishing, installing and testing Augered Cast-in-Place (ACIP) piles as shown on the Contract Drawings and as specified herein.ACIP piles are commonly referred to by several other names. Some of them are trademarks or service marks owned by Foundation Contractors and some are names that try to describe the installation process. The term “cast-inplace” is sometimes replaced with “cast-in-situ,” especially in Europe. Some common names are Auger Cast Piles, Intruded Mortar Piles, Augerpress Piles, Augered Pressure Grouted Piles, AugerPile Foundations, Continuous Flight Auger (CFA) Piles, Grouted Bored Piles, Augered Grout-Injected Piles, Drilled or Augered Uncased Piles, and Uncased Cast-in-Place Concrete Piles. To avoid confusion, ACIP pile will be used exclusively herein.1.2 PILE CONTRACTOR QUALIFICATIONSThe ACIP piles shall be installed by an experienced Pile Contractor who upon request of the Engineer shall submit evidence of successful ACIP pile installation under similar job and subsurface conditions, including a job supervisor who shall have a minimum of five years of method specific experience.The proper installation of the ACIP piles is dependent upon the procedures used by the Pile Contractor’s field representatives and equipment provided. The Pile Contractor’s personnel responsible for the installation should have relevant experience with the auguring and pumping equipment, experience placing piles under similar site conditions (including restricted headroom if appropriate), in similar subsurface conditions, as well as experience handling the special grout mixes, admixtures, etc."

Jan 1, 2016

By James Wonneberg, William Edgerton

"Today’s underground construction megaprojects are tremendous feats of engineering on a massive scale, excavating through ground conditions once thought impossible, using new and innovative contracting practices. But one important truth remains unchanged; projects are built by people. What separates winning projects from the rest? Great people, working in great organizations. In straight-talk fashion this paper sets forth fundamental differences between good project leaders and bad ones, on topics ranging from client relations, design management, and scheduling, to use of the whiteboard, and everything in between. Broadly applicable to all corners of project delivery, this paper will resonate with project leaders from around the world and help identify skills necessary to succeed in the management of underground facilities. A NOTE FROM THE AUTHORS This is a training document, and not really a “paper” per se. We have attempted to tackle some big and potentially complex topics using bite-sized paragraphs, each of which are intended to promote thoughtful introspection and dialog. So please feel free to skip around, or even set this document down while you reflect on your own personal experiences, tactics, and philosophies. We have found that providing examples of “what not to do” is oftentimes more instructive than simply listing “what to do”, hence the Good/Bad format of this document. Please note that “PM” refers loosely to anyone on the project management team not just the “Project Manager”. We also wish to preface this with the following cautionary statements: • The authors do not pretend to be the “Good PM” in all cases. In fact, we’re almost certainly the “Bad PM” for some. We aim to improve on those areas. • It has been said that “The only real training for leadership is leadership.” (Antony Jay) We all started somewhere, and nobody has 20 years of experience on day one. • Converting this text into some sort of prescriptive checklist would be a disservice to the evolution of project management. Our world – and our awareness of it – is constantly changing. We must always be open minded to new ideas and different perspectives on project delivery."

Jan 1, 2016

By T. Guan, G. Fu, X. Zhou, L. Deng

"Currently, tailings water is required to be recycled in minerals processing plants. When the flotation tailings water of complex copper-sulfur-tungsten polymetallic ore is reused to the flotation of copper-sulfur ore, a portion of tungsten would be lost in copper concentrate or sulfur concentrate due to the addition of the collectors used for oxidized ore which are also effective for the sulfide minerals in the process of tungsten flotation. At the same time, the technical index of copper concentrate and sulfur concentrate would be seriously affected by tailings water. Our experimental researches of many years show that the copper concentrate with a grade of 18.36% Cu at a recovery of 76.67% is obtained for the raw ore containing 0.60% Cu, 18.99% S, 0.16% WO3 in small locked cycle test using recycled water, through simple pretreatment of the tailings water from tungsten flotation and adding special regulator CW and new tungsten collector TM. The result is close to the performance without using tailings water. And, in the continuous pilot experiment which was continuously conducted for 4 days with the same technology, the following results was obtained: the copper concentrate with a grade of 21.74% Cu, the recovery was 71.55%; the sulfur concentrate with a grade of 45.16% S, the recovery was 80.10%; the tungsten concentrate with a grade of 42.63% WO3, the recovery was 54.62% relative to raw ore (the recovery relative to scheelite was 76.47%). The throughput of the continuous pilot experiment is 1.0~1.2 t/d, and the raw ore contains 0.56% Cu, 18.80% S, 0.14% WO3 on average.INTRODUCTIONMining industry is one of the most important developing fields, it supplies raw materials for industrial production as well as discharges much waste water. Tailings water is the main source of beneficiation waste water, it would result in severe environmental pollution if discharged directly. Generally, the tailings water would bring undesirable effect on flotation index after simple pretreatment by lime (Kalin, Fyson, & Wheeler, 2006). On the other hand, the treatment of tailings water through ion-exchange (Cui & Liu, 2015) before recycling is of high cost. Generally, in the flotation of Cu-S-W polymetallic ore, when the tailings water from tungsten flotation is treated simply by lime and recycled to the flotation of sulfide minerals, the flotation index of sulfide minerals decreases, and a portion of oxide mineral would be lost in the sulfide concentrate. As a result, the recycle rate of this kind of tailings water is relatively low, some of the tailings water is discharged after simple treatment and then environmental pollution is caused subsequently. One of the effective ways to improve beneficiation performance and increase the recycle rate of tailings water is to conduct research on developing recycle technology for flotation tailings water of complex copper-sulfur-tungsten polymetallic ore."

Jan 1, 2016