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By W. R. Matizamhuka

"The consolidation of nano-sized composite materials presents a challenge using conventional hot pressing methods. Spark plasma sintering (SPS) technology has shown great promise in the successful sintering of nanoreinforced composite materials. This qualitative review seeks to impart knowledge gathered, and progress made over the years on the consolidation of nanocomposite materials using SPS technology. The review is aimed at introducing this technology to the South African science and engineering community. Emphasis is on improving the mechanical properties of structural ceramic nanocomposite materials, which over the years have shown great promise in a wide range of applications, including transport, energy, mining, and the environment. Although success has been achieved within the laboratory for research purposes, there are still great opportunities to commercialize the technology for the production of larger components with more complex shapes.IntroductionSince the pioneering work of Niihara and Nakahira (1991) over a decade ago there have been extensive attempts to harness the potential of nanostructural metals, ceramics, and composites in a number of applications (Niihara and Nakahira, 1991; Selaho et al., 1997; Sekino and Niihara, 1995; Khalil, 2012; Aalund, 2008). It is generally realized that reducing the grain size of materials to nanoscale offers a significant improvement in properties. Attempts to synthesize nanostructured compacts by conventional sintering methods have not succeeded owing to the uncontrollable high rates of grain growth driven by long dwelling times at the sintering temperatures and large powder surface areas (Sergueeva et al., 2009; Gua, Khora, and Chieng, 2004; Tang et al., 2009). Following the development of spark plasma sintering (SPS) technology, it is now possible to process nanocrystalline materials with controlled grain growth. SPS, also referred to as the field-assisted sintering technique (FAST) or pulsed electric current sintering (PECS), was developed with the aim of improving on the well-established hot pressing (HP) technology (Kessel and Hennicke, 2007;"

Jan 1, 2016

By Maryam Sadeghi, Jonathan Roy, Claude Bazin

"The operation of mineral processing equipments is intimately related to the mineral composition of the ore. However unit performances are usually characterized in terms of metal contents or recoveries as this data is easier to obtain than the mineral contents. This paper describes a method that combines the operations of material balancing and estimation of the mineral contents of the ore. The estimated mineral contents can be used for process analysis and diagnosis of operating problems.IntroductionThe ore that is feeding a concentrator is made of minerals that condition the energy requirements for the ore comminution and the performances of the subsequent separation circuit. Indeed a change in the mineral composition of the ore may lead to lower grinding mill throughput as well as lower recovery or grade of the economical products. Despite this relationship is recognized the mineral composition of the ore is seldom measured or estimated to provide regular information for the diagnosis of operating problems. Some methods such as the QEM-SCAN (Whiten, 2008; Lamberg and Vianna, 2006) are available to measure the concentration of the minerals in ore samples but the method is time demanding and plant operators usually base their evaluation of the plant performances on chemical assays of metals or elements since this data is easier to obtain from ore samples.Usually samples of the streams of mineral processing plants are collected on a daily basis to be assayed for strategic elements. The obtained concentrations are balanced to produce coherent data with the material conservation constraint and then used to characterize the day to day operation of the plant in terms of recovery of valuable metals and grades of the marketable concentrates. The behavior of contaminant elements can also be monitored from this data allowing the operators to explain for instance a lower zinc concentrate grade by a more important contamination of iron. However such diagnosis is only partial as the contaminant element can be transported by different minerals. Indeed the iron contamination of a zinc sulfide concentrate may be caused by an inadequate depression of pyrite or due to another iron carrying mineral. A possible way to answer the question is to determine the concentrations of the various minerals in the ore and produced sealable concentrates. Since measuring the mineral contents of the ore could be a demanding task one should consider to estimate this data from the daily chemical assays of the sample collected on the main streams of the circuit. Some authors have already investigated the problem. Whiten(2008) described a method to estimate the mineral contents of an ore sample. Lamberg and Vianna(2006) proposed a method to do a sequential mineral reconstruction followed by a material balance. Few of these methods account for the mass conservation of the reconstructed mineral contents."

Jan 1, 2016

By S. Lee, F. Campero

"Past heap leach design has evaluated hydrodynamic parameters based on a uniform particle size of heap material, but has ignored the equal importance of the instability of fine migration within heap stack in the beginning stages of the irrigation process. The proposed probabilistic model evaluates internal fine erosion and segregation potential of well-graded crushed ores by considering the entire particle size distribution shape and site compaction degree. Using this model, unsaturated flow model parameters were successfully evaluated based on synthesized soil-water characteristic curve (SWCC). Additionally, one-dimensional unsaturated transient flow modeling with Hydrus 1D® showed that the different recovery efficiencies of heap material under unsaturated flow patterns depending on the particle size distribution and compaction degree of heap material. Agglomeration preprocess employed by most heap leach sites can reduce the potential of fine erosion within heap material significantly, but internal stability of crushed ores and hydrodynamic-based analysis of heap material’s pore structure is equally important in the prescreening stage in order to improve efficiency of the heap leach and avoid a costly agglomeration process. INTRODUCTION Accurate estimation of heap leach recovery is a difficult task due to the variety of design parameters that need to be considered - such as geochemistry, hydro-dynamics, and geotechnical design aspects, most notably particle size distribution through the process of agglomeration. After agglomeration techniques were introduced to heap leach design, the influence of particle size distribution and compaction degree in enhanced heap leach recovery has been neglected. The main purpose of agglomeration is to prevent fine segregation and internal erosion during and after heap stacking for well graded crushed ores, which will eventually diminish heap reach recovery. Recent studies show that agglomerated heap materials that had been consolidated at 10 meters below heap surface can release the fines after contact time with lixiviant to the compacted heap material and produce poor drainage performance (Milczarek et al, 2013). It must be noted that engineers and practitioners should investigate feasibility of agglomeration benefit prior to heap leach design because the agglomeration process can incur up to 20 or 30% of project costs. Internal stability of stacked heap material is mainly governed by the particle size distribution of crushed ores. Depending on this stability, segregation and internal erosion of fines can be mitigated by the application of moderate compaction effort during heap construction. Investigation of segregation and internal erosion potential therefore takes precedence for well graded crushed ore since reaction time and contact area with lixivant are greater in finer ores."

Jan 1, 2016

By J. Herhold, S. Plaum

"Increasing worldwide energy demands, coupled with desires for cleaner energy, have led to greater research of renewable energies. Strongly weather-dependent energy sources, e.g. solar and wind energy, are unable to provide a constant control energy to withstand the fluctuation of the extensive power systems during peak hours. The pump storage plant Nant de Drance in Switzerland will not only have the capacity of a nuclear power plant, producing 900 MW, but will also combine the environmental aspect and the demand for controlled energy. Nant de Drance is one of the pioneer projects for upcoming energy challenges and serves as a model for future pump storage plants. This paper focuses on the planning process and the project execution of interdisciplinary large-scale projects such as Nant de Drance, including project related data handling, communication systems and project handling skills. Furthermore, the paper discusses mining excavation in the context of large scale projects and their regional environmental challenges, e.g. high alpine rock mass on the site of Nant de Drance. PROJECT BACKGROUND Project Overview Nant de Drance is a 900 MW pump storage plant located in the Swiss Alps. To provide such a high energy storage, six turbines (reversible Francis type), each with a capacity of 150 MW, were installed. Due to the efficiency factor of 80% the pump storage plant is able to provide an energy production of 2500 GWh per annum. Construction of the pump storage plant, which has been selected as “Major Tunneling Project of the Year” in 2014 by NCE and ITA-AITES, will be completed over 10 years. Therefore, the project has been divided in six major phases outlined below: - Launch: 2008 - Access Gallery: 2010 – 2012 - Excavation KMA, KTR: 2011 – 2014 - Concreting KMA, KTR: 2014 – 2016 - Installation of HEM equipment: 2014 – 2018 - Tests & entry into service: 2017 – 2018 Figure 1 outlines the single process phases of the project. For the construction of the shafts, galleries and caverns, 1.7 million m³ of metamorphic rock with sedimentary origin (mica schist & paragneiss of the massif „Aiguilles Rouges“) were excavated. A tunnel network with a total length of 15 km was created. The machine cavern measures 194 m long, 32 m wide and 52 m tall."

Jan 1, 2016

By Lei Pan, Roe-Hoan Yoon

"Sea water flotation is becoming a reality due to shortages of fresh water. To better understand the mechanisms involved in sea water flotation, we have studied the effects of KCl and MgCl2 on the surface forces present in the wetting films of water formed on hydrophobic surfaces. The measurements have been conducted using the force apparatus for deformable surface (FADS), which is capable of measuring both the surface and hydrodynamic forces responsible for bubble-particle interactions. The results obtained with the gold surfaces hydrophobized with butanethiol (C4SH) show that the wetting films of Ultrapure water do not rupture despite the relatively high water contact angle (??= 80o) and the presence of a long-range hydrophobic force with a decay length (D) of 25 nm in the film. The film is metastable with its equilibrium thickness (he) at 140 nm, where the repulsive double-layer force is equal in magnitude to the capillary force as shown in Figure 1. The film loses its metastability and ruptures in a 10-2 M KCl solution, in which the double-layer force decreases substantially due to the decrease in Debye length (?-1) while the decay length of the hydrophobic force remains as large as 23 nm.The effects of MgCl2 has also been studied using a C4SH-coated gold surface with a water contact angle of 87o. It has been found that in a 10-4 M solution double-layer force decreases due to the decrease in the negative ?-potentials and Debye length, which should help destabilize the wetting film. However, the film remains stable due to the decrease in the decay length of the long-range hydrophobic force from 45 nm in Ultrapure water to 14 nm in the MgCl2 solution. Thus, the results obtained in the present investigation support the view that the difficulties in sea water flotation may be due to the presence of MgCl2 rather than the presence of NaCl in sea water (Ramos, et al., 2013)."

Jan 1, 2016

By Brandt Porter, Lawrence K. Nordell, Alexander Potapor

"Improving comminution efficiencies with better machine geometries and motions, improved nip angles, rock movement control, and voids regulation among rocks are some of the next steps to better machine performance. Recent gains (~30%) using high pressure grinding rolls (HPGRS) and vertical roller mills, in theory and practice, are based on Prof. Klaus Schönert’s academic studies (1979–1996) on rock breakage physics. More recently, University of British Columbia Prof. Bern Klein’s studies (2006–present) via the Ph.D. thesis of Z. Davaanyam (2015) further validate theory with practice from laboratory measurements used in this paper.They demonstrated ideal rock compression breakage, within HPGR and vertical roller mill machines using a piston-die apparatus. Similar rock size range is used. The piston-die confine rock breakage behavior. However, these studies do not extend to larger rocks (> 75 mm), and do not demonstrate the influence of rock voids among bifurcation of rocks during comminution. Further, machines fed with larger rock sizes, such as gyratory, jaw, and cone crushers, are not designed or evaluated for geometry nip angle efficiency.Herein we illustrate a performance gain of a new comminution machine, the conjugate anvil-hammer mill (CAHM), first presented at SAG 2011 (Nordell, 2012). The CAHM machine digests rock sizes 400% larger than HPGR, improves rock grip/nip geometry, better improves rock voids, improves surface wear life > 500%, and improves rock containment and slippage during compression. In all, the CAHM machine shows 22 potential improvements over the HPGR. Most importantly, CAHM shows a 50% kWh/ton improvement compared to the HPGR. Our studies use an advanced Discrete Element Method (DEM) computer program called ROCKY. ROCKY simulates moving boundaries, surface wear mechanics, and incorporates rock breakage physics based on the fundamentals of JKMRC T10 protocol [×] with tuned A × b breakage parameters.We calibrate rock breakage in our DEM code to mimicking both UBC laboratory piston-die and HPGR comminution measurements. We then model CAHM with these known ore and machine properties. Differences are illustrated using DEM techniques on each machine, including key points of when each: use geometry; initiate bifurcation breakage; provide a map of particle size distribution by location; illustrate the importance of controlling voids to minimize agglomeration after breakage; machine performance limits; and kWh/ton benefit."

Jan 1, 2016

By M. Tadie, J. Wiese, K. Corin

The use of potential control in the flotation of ores has been applied in a variety of operations and presents an opportunity for improvement of selectivity and recovery of valuable minerals. Although it is possible to vary the redox potential (Eh) using chemical additives such as sodium hypochlorite or hydrogen peroxide this is more commonly done by varying the pH and dissolved oxygen (DO) concentration. The present study involves an investigation into the effect of varying dissolved oxygen concentration, ionic strength (IS) and pH on the flotation response with the aim of application to a South African lead-zinc ore. In all cases the Eh is measured as a function of the different DO, IS and pH values. The main objective was to achieve selective flotation of galena from sphalerite. The first part of the work was micro-flotation tests conducted on single minerals with galena used as the basis for effect of the factors studied on redox potential and ultimately recovery. The DO was varied by alternating between the use of air and nitrogen as the gas source to the micro-flotation cell. PH was varied between pH 9.2 and pH 11.8 using sodium hydroxide. In the first part of this study a factorial design approach was used to determine the interactive effects of the variables tested, viz. DO, ionic strength and pH, on the Eh and on the recovery. Redox potential was measured using a combination electrode made up of platinum wire and Ag/AgCl reference combination. Recoveries indicated that pH had the most significant effect on recovery and Eh. The influence of redox potential on recovery was found to be of relevance to the condition of high dissolved oxygen concentration and less so at low dissolved oxygen concentrations. It was also noted that that the combined interaction between ionic strength and dissolved oxygen concentration had a positive effect on recoveries, however this was not so with Eh. The chemistry of the ions in solution was observed to play a dominant role on the recovery due to changes in solubility arising to pH modification. The longer term objective of this work is to determine the optimal Eh conditions for the separation of galena from sphalerite.

Jan 1, 2016

By Pierre De Pasquale, Aquilino Antonio Alvarez-Castro

Greenhouse gas emissions are now an essential issue for extractive industries. Actually, lowering of quarry energy consumption and CO2 emissions participates to answer both environmental and financial issues. The “EE-Quarry” European project, supported by the European Commission (EE stands for Energy Efficient), develops a new and highly effective modelling and monitoring Energy Management System technique to improve Energy Efficiency and move to a low CO2 emission rate in the nonmetallic mineral industry.

Jan 1, 2016

By Enguang Zhu, Kangning Zhang, Yang Li

"Due to the use of outdated mining technology or room and pillar mining process in small coal mines, the coal recovery ratio is only 10% - 25%. In many regions of China: the damage area caused by the small coal mines amounted to nearly one hundred square kilometers. Therefore, special mining techniques must be taken to reclaim the wasted resource in disturbed coal areas. This paper focuses on the different mining methods by analyzing the longwall panel layout and abandoned gateroad (AG) distribution in in the abandoned area of Cui Jiazhai Coal Mine in northwestern China. On the basis of three-dimensional geological model, FLAC- 3D numerical simulation was employed. The abutment pressure distribution was simulated when the panel face passed through the disturbed areas. The proper angle of the inclined face was analyzed when the panel face passed through the abandoned gateroads. The results show that the head end of the face should be 13-20m ahead of the tail end. The pillars on both sides of abandoned gateroads had not been damaged at the same time, and no large-area stress concentration occurred above the main roof. Therefore, the coal reserves of disturbed areas can be successfully recovered by using underground longwall mining. INTRODUCTIONDuring the 70-90's, there were a lot of small coal mines in China. Due to their disordered mining, most of the small coal mines left a great many abandoned gateroads (defined as AG in this paper) and a large area of damage coal seam. In these type of coalfields, there exist many mining risk issues, including nonuniform stress in the overburden, flooding and hazardous gases accumulated in the gob. So how to lay out a longwall panel and successfully mining under this type of abandoned small mines is a big issue nowadays in China.As mentioned above, longwall panel is being performed in the damaged coal seams in order to increase recovery of coal reserves (Wang, 2009). However, the abandoned geteroads intersect to each other all around the area. So a series of production issues arose while the panel face is passing through these AG. Gang and Gong (2015) analyzed the roof stability and abutment pressure distribution during the face was passing those AG that were parallel and inclined to the faceline. Xie, et al. (2015) proposed the ground control mechanism by passing the AG in top-coal caving longwall mining. He recommended to stop the face advance until the roof is over, using high strength bolts and cable bolts, and grouting consolidation. This paper studies the different mining methods by analyzing the longwall panel layout and AG distribution in damage coal seams. Theoretical analysis and field observation were conducted to determine the overburden movement when the face was advancing under the AG. The pillar between face and parallel AG changes from elastic to plastic and eventually to the residual supporting status. All these three status was combined with the roof structure to establish the theoretical model to calculate the shield support capacity. Numerical modeling was employed to simulate the pillar stability and determine the abutment pressure distribution as the face inclined to the AG during longwall mining."

Jan 1, 2016

By Pier Francesco Paglini, Gian Luca Menchini, Pier Francesco Boschi

"PROJECTThe “Caltanissetta” Tunnel is part of the ANAS S.p.A (National Autonomous Roads Corporation) $ 1.3 Billion Highway 640 Caltanissetta-Agrigento (Sicily, Italy) Rehabilitation Project, consisting in upgrading of 62 km (40 miles) of highway. The Italian multinational construction company Cooperativa Muratori & Cementisti di Ravenna, is the General Contractor for the design and build contract. The highway S.S. 640 starts from Porto Empedocle, to the intersection with the Motorway A 19 Catania-Palermo.The entire project, double carriageway designed for speed 70-120 km/h and maximum longitudinal slope of 4.2%, is divided into two sections. First Section: just finished and opened to traffic. Second Section: mainly involving the province of Caltanissetta, length about 28 km (18 miles), works started in 2013. This second section includes the most challenging work of all the project, the “Caltanissetta” tunnel consisting of two separate tubes, each with 3.9 km length and inside diameter of 13.50 m, in order to host 2 traffic lines and an emergency line. HORIZONTAL AND VERTICAL ALIGNMENT - GEOLOGY AND HYDROGEOLOGYThe plan alignment of “Caltanissetta” tunnel is formed by two tubes which are located at a minimum mutual distance of 22 m to a maximum of 63 m, with n. 9 pedestrian and n. 4 road cross passages; ventilation shafts are not provided. Overburden over portals area is about 5 m till a maximum of 120 m. Tunnel sections have variable longitudinal slope from 2.6% to 0.5%, transversal slope of about 3.5%. Right tube length is 3,992 m, left tube length is 3,877 m."

Jan 1, 2016

By S. Mostaghel, Fariba Azgomi, Steven Goodman, M. Oliazadeh

"It is known that copper loss to the slag phase can be in the forms of chemical dissolution and mechanical entrainment. The contribution of each mechanism to the total Cu-loss depends on different parameters and may vary significantly. To reduce emissions to the atmosphere and to shorten the converting cycle, most industrial processes are aiming at production of high matte grades. Increased oxygen-to-sulfur potential, required by these processes, results in a highly oxidized slag, which in turn leads to an increased concentration of the chemically dissolved copper in the slag phase. Hence, slag cleaning processes are becoming increasingly important for economic viability of the smelting operations. Hydrometallurgical, pyrometallurgical, and mineral processing routes have been used for copper recovery from slag. The current article reviews a trade-off study evaluating two of the commonly used methods (namely flotation and electric slag cleaning furnace) by comparing the technical pros and cons, operability/maintenance requirements, and capital/operating costs of the operations.INTRODUCTIONLarge volumes of smelter and converter slags are produced by copper smelters around the world. It is estimated that for production of each ton of metal 2.2 tons of slag is produced (Gorai, Jana, & Premchand, 2003). Considering the current pressure on the mining and metals industry for “zero-waste, sustainable and environmentally friendly” extraction of metals, it is absolutely essential for metal producers to improve their efficiencies in order to remain competitive. To meet these objectives, almost all metal producers, from steelmaking plants, ferroalloys, base and precious metals producers, are trying to reduce their metal losses to the slag phase. In particular, copper producers are aiming to produce higher matte grades in their smelting furnaces to reduce the converting time and improve their overall throughput. Copper smelters need to minimize their copper/precious metal losses to the slag phase before final landfilling/use of slag. The current article, therefore, briefly reviews the copper loss to the slag phase, methods of metal recovery from the slag and then presents a trade-off study conducted for a green field flash smelter of the sulfide concentrates. A comparison was made between two of the major process routes for metal recovery from slag, namely slag cleaning furnace and flotation."

Jan 1, 2016

By J. A. du Plessis, C. Dorfling

Selenium, tellurium and other precious metals (OPMs, namely Rh, Ru and Ir) dissolve in the final leaching stage of base metal refineries. Se and Te are typically removed from the copper sulphate leach solution by precipitation with sulphurous acid. Recovery of dissolved OPMs is desirable because it will reduce OPM inventory and losses. Thiourea is an effective OPM precipitation reagent, but yields insufficient Te removal. Copper is known to enhance Te precipitation when using sulphurous acid as precipitation reagent. The primary objective of this project was to evaluate whether simultaneous precipitation of Se, Te and OPMs using thiourea with metallic copper as precipitation reagents is possible. Batch precipitation tests were performed to evaluate the effect of temperature (50 – 70°C), thiourea addition (160 – 430% excess) and copper addition (2 – 8 g/L) on Se, Te and OPM precipitation. Increasing the temperature from 50°C to 70°C resulted in an average increase of 55.6 percentage points for Se precipitation and 13.5 – 23.5 percentage points for OPM precipitation. The copper addition did not have a significant impact on the precipitation behaviour. Increasing the thiourea addition increased the precipitation of Se and Te, especially when the copper addition was low; Rh and Ru precipitation also increased with up to 10 and 12 percentage points, respectively. Despite these observations, the precipitation of Se was significantly faster when no thiourea was added. It is not currently clear why thiourea appears to inhibit the Se precipitation by reaction with metallic copper. In general, Te precipitation was low (less than 30%) at the conditions investigated. Simultaneous precipitation of Se, Te and OPMs at temperatures below 80°C does not appear to be a viable option; a multiple stage operation for sequential removal of Se and Te with copper followed by OPM precipitation with thiourea would be required.

Jan 1, 2016

By B. R. Neufeld, M. G. Heximer, S. J. Roach, W. G. Van Luven, D. G. Reynolds

"Teck Metals Ltd. operates an integrated zinc-lead complex at Trail, British Columbia, Canada. In the zinc plant, zinc concentrates are roasted to produce calcine, which is then processed in the leaching circuit for zinc recovery. Calcine is first dissolved at pH <2.0. This is followed by a precipitation stage using additional calcine that removes iron as a ferric hydroxide precipitate which is recycled to the lead operations. It is during this pH adjustment that impurities, notably arsenic, antimony and germanium, are co-precipitated with ferric hydroxide which is the primary removal phase for these elements. Recent processing conditions have resulted in lower iron inputs to the leaching circuit and this has resulted in insufficient co-precipitation of key impurities. This paper summarizes the influence of iron concentrations on impurity removal using laboratory and plant scale surveys, and emphasizes key specification limits for the primary iron removal circuit.INTRODUCTIONThe Sulphide Leach Plant (SLP) at Teck Metals Ltd., Trail Operations produces a zinc sulphate electrolyte from three different zinc bearing streams:1. Calcine, primarily containing zinc oxide, from the roasters2. Zinc electrolyte from the Oxide Leach Plant (OLP) is produced by leaching zinc oxide fume from the pyrometallurgical lead smelting operations3. Zinc slurry from the Zinc Pressure Leach (ZPL) autoclaveThe feed sources are blended and introduced to the iron removal process to precipitate iron in the form of a hydrolyzed iron precipitate which is recycled to the lead smelting operation. Along with iron, impurities, namely arsenic, antimony, germanium and silica, also co-precipitate; this process acts as the primary removal step for these elements. Following iron removal, the electrolyte is further processed in the zinc dust purification circuits. Cadmium and copper are removed in the cold stage circuit via zinc dust cementation, followed by cobalt cementation in the hot stage circuit."

Jan 1, 2016

By Jan Drzymala, Przemyslaw B. Kowalczuk

"Flotation is possible when the forces operating between a bubble and the attached particle are properly balanced. There are many forces operating in the bubble-particle aggregate. The magnitude, location and direction of the forces depend on the type of process. The simplest arrangement in flotation is when a spherical bubble is attached to a spherical particle and both are immersed in water. In such a system there are four essential forces which shape the stability of the bubble-particle aggregate. They are excess pressure and interfacial solid/liquid, solid/gas and liquid/gas forces. A proper analysis and balance of the size, location and direction of these four principal forces lead to a general equation delineating the stability of the spherical particle–attached bubble system. Specific cross sections of the system and solution of the general equation lead to different relationships including the Young formula as well as the well-known geometrical relation between contact angle, particle and bubble radiuses. The arrangement of forces in the proposed graphical representation of the Young law is different from that presently used in literature for a bubble attached to an unmovable solid particle. Many real solid/liquid/gas systems are more complex due to the presence of additional forces which change the regular curvature of the liquid/gas interface leading to advancing and receding contact angles.INTRODUCTIONThere are numerous papers (Athukorallage, Aulisa, Iyer & Zhang, 2015; Chau, Bruckard, Koh & Nguyen, 2009; Drzymala, 1994; Lubarda, 2012; Pellicer, Manzanares & Mafe, 1995; Wang, Cui, Zhou, Xu, Sun, & Zhu, 2014; Young, 1804) and books (Davies & Rideal, 1963; Defay & Prigogine, 1966; Lee, 2013; Lyklema, 2000) dealing with the solid/liquid/gas system. Many of them concern the three involved phases having the shape of a sphere or spherical cap and the formed contact angle. The contact angle is usually measured through the liquid phase and can be either smaller or greater than 90o. The essential part of these publications is a graphical representation of the system with the Young equation shown as a result of balance of only three forces that is the excess Gibbs (or Helmholz) free energy per unit area of the three involved interfaces (Davies & Rideal, 1963; Defay & Prigogine, 1966; Israelachvili, 1985) (Figs. 1a-c). The Young equation, when the gravity force/buoyancy is neglected, is:"

Jan 1, 2016

By Thibaut Pannetier, Damien Rogat, Catherine Larive, André Regnard, David Chamoley

"INTRODUCTION For several decades, the use of fibre-reinforced sprayed concrete (FRSC) has been becoming increasingly widespread, either as a replacement for or in addition to traditional reinforcements. Macro-synthetic fibres appeared on the market at the same time as metal fibres (in the 1980s), but their use has developed more recently in France, starting in the early 2000s. While their properties are attractive for underground operations, their long-term performance remains the subject of some controversy, particularly with regard to the phenomenon of creep [1-5]. Unfortunately, the many differences of protocol among the studies carried out (materials, metering, spraying equipment and test methods, etc.) mean that, in general, the results cannot be compared. Although several methods have been developed to assess the bending creep behaviour of fibre-reinforced concrete, the majority of them test notched beams resting on isostatic supports. There are currently no standardised test methods adapted to FRSC on hyperstatic supports (as is the case for FRSC used for underground support). The Asquapro association has thus developed a new test specially designed for this type of concrete. Slabs of thickness 10cm, measuring 60 x 60 cm in plan have been sprayed and tested under bending stress according to a test procedure inspired by the standardised test of energy absorption capacity (NF EN 14488-5). An experimental campaign was launched in October 2014 to study the behaviour of seven types of sprayed concrete, manufactured using the same concrete matrix: six of them were reinforced with different metal or macrosynthetic fibres and the seventh with a welded wire mesh. Fourteen slabs were sprayed for each type of concrete and also for the non-reinforced concrete with the same matrix. The types of mechanical performances measured were the compressive strength and energy absorption capacity (at seven and twenty-eight days), in addition to the creep behaviour of these seven types of concrete, placed under stress on the seventh day. This article follows on from that published in the WTC 2015 congress proceedings [6]: the programme and the test methods are briefly described here and the results available at the time of writing are presented and discussed."

Jan 1, 2016

By Lev Filippov, Saeed Farrokhpay

"Flotation process has had less success to recover nickel from laterite ores. Nickel in laterite ores is often finely disseminated in various minerals in very fine size. Therefore, fine grinding should be used to liberate nickel containing particles. Particles resulted from fine grinding are often very fine, and recovery of such particles is generally low using conventional flotation methods. One option to overcome this problem is enhancing the particle-bubble interaction by increasing the particle size and forming flocs. Selective flocculation of fine particle can be therefore the key to enhance the flotation performance of low grade laterite ores.INTRODUCTIONThe majority of the world’s nickel resources occur as laterite ores which are complex, low grade and expensive to treat using conventional smelting and high temperature and/or high pressure autoclave methods (Xu et al., 2013). While about 70% of the nickel resources are contained in laterites, only 40% of the world’s nickel production comes from these ores. Researchers have tried to improve the nickel laterite flotation recovery by using a number of feed preparation techniques. For example, in a study on a laterite ore sample from Australia, flotation tests on a narrow sized fraction of 38-75 µm resulted in only a small increase in the nickel content, from 1.0% to 1.4% (Quast et al., 2015a). In fact, no physical separation technique (including flotation) has been able to dramatically upgrade nickel in laterite ores (Quast et al., 2015b). Thus the challenge of processing low grade nickel laterite ores still exists. This paper aims to explore the possibility of applying selective aggregation in pre-concentrating of valuable minerals in nickel laterite ores."

Jan 1, 2016

By Faruk Oksuz, Dave Schiemann, Matt Trotter, Mike Padilla, Miguel Sanchez, Carmen Scalise

"Chicago’s Tunnel and Reservoir Plan (TARP) is a nearly $4.0 billion and over 30 years long program and arguably the largest and longest combined sewer tunnel and reservoir system in the world. Recently, a massive tunnel construction is underway for connection of the McCook Reservoir to the existing TARP tunnel system. Once completed in two stages in 2017 and 2019, the reservoir will hold 38 billion-liters (10 billion-gallons) of combined sewer overflows (CSO) and flood waters from the city of Chicago and over 20 communities in Cook County, Illinois. McCook Main Tunnel will connect Chicago TARP’s Mainstream Tunnel to the McCook Reservoir. Connecting two 10-meter (33-ft) diameter tunnels at 92 meters (300 feet) underground has not been a simple task for the McCook Tunnel Project participants, the project owner U.S. Army Corps of Engineers (USACE), local sponsor Metropolitan Water Reclamation District (MWRD) of Greater Chicago, designer Black & Veatch, and the contractor, Kiewit Infrastructure Company (Kiewit). Yet alone excavating and lining massive size tunnels, the team is also challenged with live tunnel conditions in TARP Mainstream Tunnel that carries CSO and flood flows from city of Chicago which runs full during rain events and must remain in service at all times. The project is a noteworthy feat for engineering and construction with the upcoming installation of world’s largest underground wheel gates for flow control within the tunnel. In this paper, we describe one of its kind elements of this major tunnel system along with key design and construction considerations, specifically for the live tunnel connection which is a rare occurrence in the tunneling industry. INTRODUCTION The McCook Main Tunnel system consists of a 10 m (33 ft) finished diameter and 490 m (1,600 ft) long hard rock tunnel constructed from a 27.5 m (90 ft) diameter and 92 m (300 ft) deep main gate and tunnel construction access shaft. The gate shaft houses six high head 4.4 m by 9 m (14.5 ft by 29.5 ft) wheel gates to be installed in the bifurcated and steel lined section of the tunnel. The tunnel also includes portal and energy dissipation structures as it daylights into the reservoir."

Jan 1, 2016

By Monique Pharand, Jeet Basi, Simon Hille

"High quality activated carbon is essential in the recovery of gold in Carbon-in-Leach (CIL), Carbon-in-Pulp (CIP), and Carbon-in-Column (CIC) circuits. Goldcorp’s Campbell complex at Red Lake Gold Mines (RLGM) experienced a period of low soluble gold (Au) recoveries during Q3 2013, during which time the processing plant saw higher than typical Au grades in the CIP tails solution stream. A technical review identified that the CIP circuit may be experiencing a carbon-fouling issue. To further understand the degree of fouling, and to troubleshoot the low recovery problem, samples of carbon were collected from the CIP, elution, acid wash, and regeneration circuits, and submitted for characterization test work. This paper summarizes the steps taken to remedy the carbon-fouling problem at RLGM resulting in increased overall Au recovery.Overview of Red Lake Gold MinesRed Lake Gold Mines is located approximately 450 kilometres northeast of Winnipeg, Manitoba, in Ontario’s western boundary region.All mining is carried out through underground operations, utilizing overhand cut and fill, underhand cut and fill, and long-hole mining methods. Ore types include: silica-replaced carbonate veins with free milling gold, siliceous replacement-type mineralization (rich with arsenopyrite) marginal to veins, broad disseminated sulphide mineralization along major shears, and some minor sulphidized chemical sediment-hosted ore. The operation is supported by two milling complexes: Red Lake, and Campbell.Process DescriptionThe Campbell processing plant's main unit operations consist of grinding, gravity concentrating, flotation, autoclaving, leaching, carbon-in-pulp (CIP), carbon elution and reactivation, electro-winning, bullion smelting/refining and cyanide destruction, and concentrate handling. A schematic of the plant flowsheet is shown in the figure below.BackgroundHigh quality activated carbon is essential in the recovery of gold at many CIP/CIL/CIC processing plants. At RLGM, gold-bearing ore is milled and cyanide is added to leach the gold from the solids into solution. The dissolved gold then comes into contact with activated carbon, and adsorbs onto its highly porous surfaces. Gold is then recovered by separating the activated carbon from the main slurry for further refining. As such, the characteristics and quality of the activated carbon have a significant and direct impact on overall gold recovery."

Jan 1, 2016

By John R. Poland, Benjamin Mirabile

"There are significant ground support hazards to both personnel and equipment when longwall equipment must mine through cross-panel entries. Backfilling support strategies are often cost prohibitive. Additionally, it is often very difficult to install backfill material in full contact with the mine roof in backfilled entries. This paper will demonstrate the use of fiber-reinforced concrete crib material as a method for standing support in longwall cut-through entries. Despite conceptions of poor performance in potentially high convergence environments, fiber-reinforced concrete crib material can be used in longwall support environments subject to high abutment loads and convergence. Central to the proper design of these supports is the use of yield material between crib layers and at the mine roof/crib interface. The placement location of cribs within the cut-through entry has a significant effect on the main roof behavior and associated ground conditions in the longwall abutment zones. It is particularly important to engineer the proper load displacement characteristics of the installed crib to support expected ground conditions. The cost of fiber-reinforced crib supports is less than 50% of backfilling methods.INTRODUCTIONThe design and execution of a longwall cut-through is perhaps one of the most difficult and risky endeavors in the field of ground control. The magnitude of abutment loads and potential for ground failure present a significant safety hazard to underground personnel. In addition, there is significant financial risk, with replacement costs of face equipment exceeding $200 million. Faced with these consequences, it is both advisable and common for an engineer to design ground support with a significant margin of safety. This paper outlines a robust support system that possesses a significant margin of safety along with significantly lower costs than most other support systems employed in similar conditions.SITE PARAMETERSThe mine site described in this study operates in the Herrin No. 6 seam in southern Illinois. The seam height is typically 76- 84 inches. Mining height in gateroads and mains is typically over 96 in, and typical mining height on the longwall face is 82 in. The overburden depth is 950-970 feet in the area of the longwall cutthrough."

Jan 1, 2016

"Use of the Bond Work Index to size industrial grinding equipment was widely known. Bond also described its application for measuring grinding circuit efficiency. Bond Work Index efficiency has been widely used by industry, but until now without concise standard methods, or a formal guideline, for doing so. The Bond Efficiency Sub-Committee of the Industrial Comminution Efficiency work group of GMSG has produced such a guideline, which follows. It covers the methods and provides examples of how to calculate the Bond Work Index Efficiency of most industrial grinding circuits. Bond Work Index testing equipment and methods are also provided. The next steps are: (1) to establish standard reference testing materials and one or more laboratories to serve as bases for reference testing; (2) to establish the accuracy of relative circuit Work Index Efficiency determinations; and (3), to develop and provide a public data base of industrial comminution circuit Bond Work Index Efficiencies. OUTLINE Introduction and Background Part 1. Determining the Bond Efficiency of a Grinding Circuit Method Demonstration/Example Calculations Bond Work Index Testing Equipment and Procedures Accuracy of Comparative Circuit WI Efficiency Determinations (under development) List of Symbols Part 2. Data Base of Bond Grinding Circuit Efficiencies (under development) References Annexes: A1. The Bond Crushing WI Test Equipment and Procedure for Bond Efficiency Determinations (Revised March 26, 2015) B1. The Bond Rod Mill WI Test Equipment and Procedure (Revised March 26, 2015) C1. The Bond Ball Mill WI Test Equipment and Procedure (Revised March 26, 2015) Introduction and Background Please see the references and slide presentation “Bond Efficiency SLC (rev3)-print.pdf.” This efficiency determination applies to most brittle materials in their naturally occurring (un-scalped) size distributions being treated in size reduction circuits down to a product sizing of approximately eighty percent passing 70 µm (see below). Unusually shaped materials (e.g. mica) should be regarded with caution. Part 1. Determining the Bond Efficiency of a Grinding Circuit Method In the plant: • Define the circuit for which the WI efficiency is to be determined • Procure samples of the circuit feed and product • Obtain the power draw of the size reduction equipment at the drive pinion(s) • Obtain the circuit throughput rate (dry tonnage) In the laboratory:"

Jan 1, 2016