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By Julie Shang, Tim Evans, Michael McDonald, Yu Guo, Bingfeng Xue

Effective chemical grouting represents a technical and economic solution to many challenges related to soil stabilization and blocking the flow of water. Chemical grouts are facing increasing demands in durability, environmental impact and value. To better meet these challenges, a novel and cost effective process was developed to produce a sodium silicate that increases the ratio (i.e. silica to alkali) beyond what has been used in traditional silicate-based grouts. The move to a lower sodium form of silicate was anticipated to have two desirable performance enhancements; a reduction in leachable sodium and improvement in the durability of the silicate grout. An in-depth study was performed at the University of Western Ontario Geotechnical Research Centre to quantify improvement in durability as well as other performance properties such as compressive strength. The study began by first modelling the gelation time and reaction kinetics of the high ratio sodium silicate over a range of silicate concentrations using two different setting agents as well as with/without a hardener. As expected, reducing the level of sodium produced a much more water resistant grout. These same formulations also showed a four order magnitude reduction in permeability. Somewhat surprising, test results showed the higher ratio silicate having much higher compressive strength. At the time of this abstract, the high ratio silicate was moving to the trial stage for geotechnical applications while the petroleum industry was further along with several successful field trials for water blocking as well as cement remediation. This paper will review the chemistry, testing and features of this next generation form of sodium silicate. Introduction Sodium silicate based grouts have been used for decades for soil stabilization and water control. Over the years, several silicate-based systems have been tailored for specific application conditions to maximize performance and reduce cost. Past research has also focused on the selection of setting agents and hardeners for silicate grouts. A constant in the development effort has been the grade of sodium silicate. The vast majority of chemical grouts have been formulated using a ~3.2 ratio sodium silicate. As a very brief overview on the meaning of “ratio” and why it is so important to performance properties. The ratio describes the proportion of sand to alkali or SiO2:Na2O used to make different grades of sodium silicate. Comparing a 2.2 ratio to a 3.2 ratio, it is easier to polymerize and set the 3.2 ratio. Upon setting and forming a silica gel, the system has higher compressive strength values and improved durability with a reduced level of syneresis (i.e. shrinkage). This leads to the question if 3.2 is better than 2.2 why not go to 4.2 or higher?

Jan 1, 2018

By C. I. Walker

"Understanding the influence of slurry particle size on wear is essential to maximising the life of centrifugal pumps operating in mineral processing plants and for hydrotransport duties. Particle size influences not only the wear rate but also the wear pattern in pumps. Complicating the situation is the different relative wear rate for different pump parts depending on particle size. A simple Slurry Wear Type (SWT) number based on particle size and size distribution is proposed as a means for categorising slurry pump wear patterns and optimising pump application and material selection. INTRODUCTION The wear of centrifugal slurry pumps in mineral processing and hydrotransport can be quite severe and is one of the major cost concerns of plant operators. Some typical worn slurry pump parts made from high chrome white iron are shown in Figure 1. Erosion wear rates of up to 2 mm/day are not uncommon [1].Centrifugal slurry pumps are required to handle solid-liquid mixtures with a very wide range of particles extending from sub-micron size clay through to 200mm and larger cobbles. Particle size distribution can also vary widely from uniform and narrow (such as graded sand) through to very broad (such as grinding circuit slurries). Slurry particles may also undergo size change due to attrition during the pumping process. The major wearing parts in a slurry pump are the impeller, side liners and casing liner. These are shown with reference to a Warman® MC centrifugal pump in Figure 2. While many particle properties (including size, shape, hardness, density and concentration) influence pump wear, this work will consider only the effect of particle size as it is the major determinant of wear characteristics. Particle size influences the wear pattern, material performance and relative component wear rate in centrifugal slurry pumps. SLURRY WEAR TYPE (SWT) BASED ON PARTICLE SIZE Soil and grain size classification systems abound. Many have different limits to the ranges for a particular classification, albeit not too"

Jan 1, 2018

By Kazutoshi Haga, Ryo Kawarabuki, Atsushi Shibayama, Masato Kosugi, Masanobu Kawata, Refilwe S. Magwaneng, Kohei Mitsuhashi, Altansukh Batnasan

"The recovery of copper and iron from carbonaceous sulfide ore is of economic and environmental interest despite its processing difficulty. Hydrometallurgical methods can be used to concentrate sulfide metals of these hard-to treat ores in a simple and eco-friendly manner. In this study, high pressure acid leaching process was comprehensively studied to dissolve metals from an ore with organic carbon and carbonaceous gangue material using autoclave equipment. The sample used for the experiment was reduced to an average particle size of 75 µm after undergoing flotation to reduce carbonaceous gangue minerals. A comparative study on two leaching reagents of sulfuric acid and iron sulfate and their dosage was used to observe dissolution of copper and iron. Other parameters of temperature (90-170 0C), solid-liquid ratio (1:10, 1:5, 1:3 and 1:1) were investigated through kinetic studies to access the influence of organic carbon and carbonaceous material on recovery of copper and iron. Materials balancing analysis was used to evaluate the efficiency of metal dissolution. The effects of parameters on leaching rate and concentration was demonstrated. Results obtained from the experiment indicated low concentrations of copper and iron in pregnant leached solutions, which was attributed to acid consumption by organic carbon and carbonaceous material in feed. However, the high efficiency rates of 92.4% and 86.3 % for copper and iron respectively in sulfuric acid can be concluded that leaching of concentrates of carbonaceous ore is possible provided system control on type of acid, pH and temperature. The purpose of this research is important in understanding mechanism involved in dissolution of copper and iron in concentrates of carbonaceous ore and marks an improvement in leaching of highly contaminated ores."

Jan 1, 2018

By Pengfei Wang, Guorui Feng

With increase of cover depth, dynamic incidents, such as large gateroad convergences and coal bursts, are frequently encountered when mining deep longwall panels. It would be more adverse due to large inclination of the coal bed. Because of the shearer and armored face conveyor (AFC) sliding downhill, toppling of shields, disconnection of AFC to the stage loader are related to it. Tangshan coal mine is a typical deep, inclined coal mine situated in the east of China. The split-level longwall panel layout is being adopted for mitigating problems above, where tailgate entries of longwall panels are driven below gob edge and along the floor, while headgate entries are driven along the roof. Wire mesh and steel sets have been used empirically for support for tailgates for about a decade, which reduces the tunneling speed. In view of this fact, the paper serves to investigate the proper adjustment of the tailgate locations and to analyze the resulting stress and yield conditions according to support strategy. First, field measurements and observations were carried out. Stress data distributed within surrounding rock mass were obtained, especially within the roof since the roof consists of only a coal sheet with a thickness of 2–4 m. Roof pressure of the tailgate represents the pressure at the gob edge since it is below the gob edge. The convergence data were also collected. Numerical modelling using FLAC3D™ was conducted and validated by side abutment stress data. The double-yield constitutive model was used to simulate the gob. The finalized tailgate layout and design were determined through comparison of three scenarios. The research in this paper can help engineers choose the proper tailgate layout and support strategies for coal mines that are going to adopt a split-level longwall panel layout for mining deep, inclined longwall panels.

Jan 1, 2018

By B. McFadzean, S. Geldenhuys

"Dynamic froth stability tests have found widespread application to quantify froth performance. However, little to no attention has been given to scale parameters in the performance of these measurements. One of the most frequently changing scale parameters in the laboratory scale is the column diameter as this directly influences the amount of material needed for testing. Four different column diameters (50, 100, 200 and 300 mm) were used for this study. Column diameter experiments were done on an industrial scale by means of manual tracking of froth growth versus time. An increase in measured dynamic stability was seen with increasing froth column diameter, which reached a limiting value at large column sizes. It is thought that this behaviour is controlled by the differences in drainage rates of the bulk froth compared to froth in contact with the wall. The column diameter required to represent full-scale froth stability is proposed. A ratio of column surface area to total bubble surface area was used to correct for different column diameters.INTRODUCTIONThe froth phase in flotation contributes significantly to the overall performance of the process (Zanin, et al., 2009). Numerous parameters have been investigated to quantify froth performance, most notably froth recovery (Feteris, et al., 1987; Seaman, et al., 2004; Vera, et al., 1999). Experimentally gathering data to determine froth recovery is challenging and for this reason froth stability has been highlighted as a possible alternative.Froth/foam stability is defined as the time of persistence of the froth/foam and is affected by numerous factors, such as: drainage of the liquid in the lamellae (gravitational and capillary), viscosity at the bubble surface, thickness of the electrical double layer and gas diffusion through liquid films caused by pressure differences between bubbles (Aktas, et al., 2008; Beneventi, et al., 2001). It is well documented in both froth and foam literature that froth/foam stability will heavily influence the performance and appearance of the froth/foam phase (Ata, 2012; Farrokhpay, 2011; Weaire, et al., 2005). Experimentally this has been explicitly proven as an increase in froth stability, characterised by water recovery, usually results in an increase in both valuable mineral and solids recovery (Wiese, et al., 2011; Banford, et al., 1998)."

Jan 1, 2018

By D. Endara, E. De la Torre, S. Echeverria

"The study was conducted with the sulfide mineral from the province of Zamora Chinchipe. It was characterized physical, chemical and mineralogical, obtaining between its main results a copper concentration of head ore of 0,58 %. The mineral was homogenized and taken to direct closed circuit: rollers mill - vibratory sieve (mesh # 16). Flotation was carried out in two stages: the collective and selective where in each one were performed 6 water recirculation as suggested by the technique used to relate information of a batch flotation to a continuous circuit called cyclic flotation tests (LCT). In the collective flotation (bulk) 2 Kg of mineral were grinded with 10 Kg of balls for 30 minutes and 62.5% of solids, the pulp was taken to the flotation process with solids percentages of 33.3% and pH 7, it was conditioned with 45 g / t of KAX 1%, 50 g/t of Aerofloat 208 and 20 g /t of Pine oil. The concentrate and tailings bulk obtained were filtered separately for the recovery of water and it was recirculated to the next grinding of 2kg mineral. In the selective flotation 1Kg of bulk concentrate was carried to regrind for 40 min, 62.5% percent of solids, 10 g of lime and 5 g of tannin FLOATAN T5, the pulp was taken to the roughing-collection-refining flotation circuit controlling the pH between 10.5 and 11 where conditioning (60 g/t of Aerofloat 208 and 30 g/t of Pine oil) and adding lime to the circuit and FLOATAN according to the procedure. The four fractions obtained from the circuit were filtered; the water was recirculated to the regrind process of the bulk concentrate. In the collective flotation process tends to stability in the fourth recirculating water to obtain a copper content of 4.5% and a recovery of 87%, recirculated water represents 69.8% of all water used in the 6 recycles. In the selective flotation circuit the fifth recirculating of water is the optimal where obtaining a copper content of 13% and a recovery of 72%, recirculated water in the selective flotation circuit represents 95% of the total water used in the 6 recycles."

Jan 1, 2018

By M. J. Raffaldi, L. A. Sandbak, K. K. Dehn, S. N. Warren, J. B. Seymour

"Confidence in the design and stability of an undercut cemented rockfill (CRF) span requires a thorough understanding of its emplaced material properties. The unconfined compressive strength (UCS) of cast 6×12-inch QA/QC cylinders is typically used as an index test to track the performance of a backfill plant. However, these strengths need to be adjusted for oversize aggregate removal and sample size effects if used to estimate the emplaced UCS. UCS testing of 6×12, 12×24, and 18×36-inch cast CRF cylinders was used to quantify the relationship between readily available 6×12-inch QA/QC strength data and the emplaced UCS. Test results indicate that the UCS of 18×36-inch cylinders is roughly 40% of the 6×12-inch UCS for similarly compacted samples. Mine QA/QC data indicate that the UCS of CRF is highly dependent on the degree of compaction as measured by bulk density. Based on these results, knowledge of the emplaced CRF density compared to 6×12-inch QA/QC density is also necessary to estimate the emplaced UCS. The scale effect reported in this study is more pronounced, compared to current understanding of backfill properties discussed in the literature, and underscores the need for further site-specific testing. INTRODUCTION The National Institute for Occupational Safety and Health (NIOSH) Spokane Mining Research Division (SRMD), in cooperation with Barrick Gold Corporation, is conducting research with the goal of determining a relationship between 6-inch QA/QC cylinder strengths and the in-place strength of cemented rockfill (CRF). CRF cylinders with dimensions (diameter × length) of 6×12, 12×24, and 18×36 inches were cast onsite at the Turquoise Ridge Joint Venture (TRJV) and Cortez Hills Underground (CHUG) mines using CRF mixed at the mines’ respective batch plants. UCS tests were performed after 28 days of curing to estimate the size correction factor for each mine’s CRF. This paper presents an overview of the mix designs, sample collections, testing methods, test results, and an estimate of the UCS size correction factor for CRF at these mines. BACKGROUND Underhand cut and fill mining methods utilizing cemented backfill are common in underground metal mines in the United States. These methods are advantageous for controlling weak or highly stressed rock mass conditions by creating an engineered material that can be placed in openings, forming a beam that provides a safe, stable back for the next undercut. Increased use of underhand methods in the last 30 years has greatly improved safety in many cut-and-fill mines. However, with the continual push for more efficient and cost effective mining, there is a need to thoroughly understand the geotechnical properties of cemented backfill to assist in designing safe, stable openings while optimizing costs and production schedules."

Jan 1, 2018

By S. J. Neethling

"Liberation is a key driver in all mineral separation processes as it limits the maximum possible grade for a given recovery. In flotation, this is further complicated by the fact that it is surface exposure of the floatable minerals that determines the ultimate performance. Liberation, grade and surface exposure are commonly quantified using Scanning Electron Microscopy coupled to Energy Dispersive X-ray spectroscopy (SEM/EDX) analysis of polished sections. The intrinsically 2D nature of this technique can result in significant sampling errors and stereological effects that can affect the quantification of the ore’s textural characteristics. X-ray micro-Tomography (XMT) is an imaging method that can non-invasively and non-destructively delineate ore fragments in 3D, thus providing an alternative method that eliminates the need for stereological corrections and readily provides surface exposure. A methodology and automated algorithm were designed for extracting this information from images of closely packed particles, thus allowing samples containing a large number of particles to be assessed.The main drawback of XMT is that, unlike SEM/EDX, it cannot directly measure the mineralogy of the sample, instead producing a 3D X-ray attenuation map. We therefore also present an algorithm for calibrating the thresholding of the XMT images based on SEM/EDX images of sections through the same sample, thus allowing the liberation analysis to be carried out on a 3D mineral map in which the uncertainty in the mineral assignment is small and statistically quantified.The methodology was tested on low grade porphyry copper ore as this is both an industrially relevant and traditionally difficult system to quantify using XMT due to the similarity in the X-ray attenuation of the sulphide species, especially the pyrite and chalcopyrite. As each 3D volume imaged contain 100-1000s of grains, large data sets can be readily produced. By dividing these particles into classes based on both their surface exposure and grade, the extent to which there is preferential breakage of the particles can be assessed—an important consideration if sufficient surface liberation for good flotation performance is to be achieved at coarser particle sizes. It is also shown that erroneous conclusions can be drawn when basing the analysis on 2D sections through the particles."

Jan 1, 2018

By S. V. Volkova, S. V. Mamonov, T. P. Dresvyakyna

"The purpose of the work is complex processing of sulphide-magnetite ores with the separation of iron into iron concentrate, copper and noble metals into copper concentrate.The work shows the research results of the material composition of the sulfide-magnetite ore in the South part of the Peschanskoye deposit. According to the research results it is found that ore contains 0.59% of copper, 66% of which is related to chalcopyrite; 39.9% of iron, 81% of which is magnetite and 3.87% is sulfur, 42% of which is associated with pyrrhotite, 65% of which is a monoclinic modification. According to the results of the mineralogical analysis, it is found that the main iron-containing mineral in the ore under study is magnetite, the content of which is 45%. In addition to magnetite, iron is found in carbonates, pyroxenes, amphiboles, chalcopyrite, pyrrhotite and pyrite. The main copper mineral in the sample is chalcopyrite; covellite, chalcosine, bornite have a subordinate value. The total content of copper minerals in the sample is about 1%. The processing technology provides the three-stage grinding of the ore to a size of 65% of the class-71 µm; a cycle of copper flotation; a cycle of pyrrhotite flotation; wet magnetic separation of flotation hosts. The following optimal parameters are established for the ore under study: the alkalinity of the slurry is 250-280 g/m3 of free calcium oxide in the liquid phase of the slurry; the duration of the main copper flotation is 20 min; the consumption of butyl potassium xanthate is 20 g/t; the consumption of the foam generator T-92 is 15 g/t; the consumption of copper sulfate for the activation of pyrrhotite is 300 g/t; the magnetic field strength is 1500 Oersteds.A complex flotation-magnetic technology for the enrichment of sulfide-magnetite ore has been developed in the laboratory in order to produce a copper concentrate containing 21.6% of copper with 91.7% extraction and an iron concentrate with 66.5% of iron mass fraction and 1.2% of sulfur with an extraction of iron into the concentrate of 87.9%."

Jan 1, 2018

By A. D. Ramirez, L. E. Gutierrez, M. A. Garces

"Scarcity of freshwater in mining operations in arid zones in different parts of the world has led to the mining companies to think about using seawater as an alternative source in mineral processing operations. Examples of this situation are Minera Centinela (AMSA) and Minera Sierra Gorda (KGHM), both plants located in northern Chile. Previous studies showed that molybdenite flotation is depressed at high pH values in seawater as a result of the action of hydroxy-complexes of magnesium and calcium that attach to the faces of the molybdenite particles, reducing its hydrophobicity and as a result flotation recovery. In this work, a study of the interactions between molybdenite and kaolinite is presented. The effect of sodium hexametaphosphate (SHMP) dispersant on the recovery of molybdenite in the presence of kaolinite is studied in different aqueous media and it is correlated with induction time data. The results allow to conclude that kaolinite has a stronger depressing effect on molybdenite in seawater compared to its effect in freshwater. Magnesium and calcium are the cations that have the strongest effect in enhancing the interactions between molybdenite and kaolinite. SHMP reduces the negative effect of kaolinite on the floatability of molybdenite both in seawater and freshwater. The induction time data correlate well with the micro-flotation results.INTRODUCTION The need for water in mining processes in arid and semi-arid zones has led to the search for new alternatives for this resource, among which seawater has gained an important space. In Chile, most of the mining operations are located in the northern part of the country, an area that has limited availability of freshwater for the processing of copper and molybdenum sulfides (Castro and Laskowski, 2011, Castro et al., 2012 a). Seawater is an aqueous solution with high concentrations of ions such as Na+, K+, Mg2+, Ca2+, So42-, Cl-, and other less concentrated species which generate important effects in the flotation process as they modify the properties of the solid/liquid, and liquid/gas interfaces, and also increase the consumption of some important reagents used to depress pyrite such as lime. It has also been proposed that flotation of copper sulfides is negatively affected in seawater because of the reduction of oxygen solubility compared to freshwater (Millero et al., 2002; Duan and Sun, 2003). In contrast, it has been shown that the presence of ions in solution in the process of flotation of natural hydrophobic resources improve recovery as result of the compression of the electric double layer (Laskowski, 1974; Klassen and Mokrousov, 1963, Laskowski and Castro, 2008."

Jan 1, 2018

By Demosthenes (Dennis) Efstratiadis

After European Council’s Directive 93/15/EEC of 5 April 1993 on the harmonization of the provisions relating to the placing on the market and supervision of explosives for civil uses, the so called ”Intra-Community Transfer of Civil Explosives” has been possible between users in the member States of the European Union. In line with this new approach, the placing on the market and the supervision of explosives for civilian use are subject to essential safety requirements. The explosives market in Italy is regulated by the ”day consumption of explosives” that imposes the user to have on site just the strict quantities of explosives and accessories that are necessary for the daily drill and blast activities. Any eventually advanced explosives’ quantities should be destroyed, either returned to the provider’s magazine until the 24:00 hours of the same day.

Jan 1, 2018

By P. Fawell, C. Castillo, C. Ihle

"The need to reduce fresh water consumption has pushed many Chilean mining companies to utilize seawater obtained from distant locations. At the same time, increasing amounts of fine particles in copper tailings raise new operational challenges, in particular, the thickening and transport of tailings featuring significant fine fractions, which are predominantly clays. In the present paper, we study the role of salinity – using seawater and NaCl – on the flocculation efficiency of copper sulphide tailings to which kaolinite was added. To mimic the water recovery stage in thickeners, flocculation has been induced in jar tests and the corresponding aggregate size distribution has been estimated from chord length distribution data obtained using a focused beam reflectance measurement (FBRM) probe, at various mean turbulence intensities. Results show that for constant flocculant dosing, a measurable influence of the salt on the resulting aggregation exists. For each case, resulting optimal turbulence levels are also reported, a particularly interesting feature when comparing flocculation efficiency of the coarse and fine fraction. INTRODUCTION The concentration of copper minerals through comminution and flotation is intensive in the use of fresh water. The need to reduce the consumption of fresh water and the environmental standards associated with increasingly demanding mining operations have driven improvements in the techniques of recovering water from tailings. The challenges are even greater when required to treat minerals with high clay contents. In Chile, kaolinite is one of the most important clays to consider due to its abundance in the tailings from many operations. The most common and highest throughput unit operation in the mining industry for the recovery of water from tailings is gravity thickening, with sedimentation of fine particles assisted by flocculation. Flocculation is the process of forming aggregates by contacting the suspended particles with a flocculant. The flocculant is typically a water-soluble synthetic long-chain polymer, where copolymers of acrylamide and sodium acrylate are the most common types used in the Chilean mining industry. The aggregation mechanism includes the following key steps: (1) flocculant distribution through the suspended particles, (2) flocculant adsorption onto particle surfaces, and (3) collisions between particles leading to polymer bridging to form aggregates, but concurrently (4) breakage of aggregates becomes a factor under prolonged shear (Gregory, 1988)."

Jan 1, 2018

By M. Paymardi, A. Foroutan, T. P. Oladele, M. Abbaszadeh

"Reducing the amount of zinc and iron in lead concentrate is one of the important aspects of processing in lead and zinc flotation plants which results in improvement of lead concentrate quality and prevents zinc loss. In mineral flotation industries, lignin sulfonates are known for their depressant properties for both non-metallic and sulfide minerals such as pyrite and pyrrhotite.This paper investigates the effect of lignin sulfonate on flotation of galena, sphalerite and pyrite in sulfidic lead and zinc bearing ores. Batch flotation tests were carried out on a high pre-activated zinc samples. The feed contains 0.9% lead, 2.3% zinc and 7.5% iron. The dosage of all reagents such as sodium cyanide, zinc sulfate, ethyl xanthate and methyl isobutyl carbinol (MIBC) were kept constant for lead flotation step except calcium lignin sulfonate (CLS) that was varied between 0 and 80 gr/ton.Results for lead rougher concentrate showed that the grade of lead increased by 2.7% while the recovery decreased around 4 % at the highest dosage of CLS. Furthermore, zinc recovery decreased by 6.3% with no significant change in grade. Iron recovery in lead rougher concentrate decreased by 2.3 % while the grade changed around 1%. In sphalerite flotation step, CLS dosage was changed in the range of 0 to 80 gr/ton while the dosage of other reagents such as copper sulfate, lime, amil xanthate and MIBC were kept constant. In this step, no decreasing in grade and recovery of zinc was observed, while iron recovery decreased. It was suggested that CLS has a depression effect on sphalerite in conjunction with other zinc depressants such as zinc sulfate and sodium cyanide. CLS was further used in plant with non-pre-activated zinc minerals. Similar trends were observed in plant.INTRODUCTIONlead-zinc sulfide minerals play an important role in the production of lead and zinc metals in the mineral processing industry (Luo, Feng et al. 2016). Due to complexity of the lead-zinc ores mineralogy, in some situations, their flotation properties vary significantly from one deposit to another one even in itself (Bulatovic 2007)."

Jan 1, 2018

By Martin Schrier

This paper is an overview of a granite quarry cleanup near Kirkland Lake, Ontario, Canada. A combination of local geology and previous subpar blasting techniques led to the quarry needing significant rehabilitation before it could proceed with general operation. Advanced technologies were implemented into the blasting process, including laser profiling, bore-hole tracking, and geometric timing. The contract consisted of two floor shots and two 25,000 tonne (27,558 US ton) production shots.

Jan 1, 2018

By Shazad Hosein, Bill Birch, Catherine Johnson, Toby White, Liam Bermingham

Air overpressure (AOP) from blasting can give rise to a significant number of blasting complaints from residents living adjacent to blasting operations from mines, quarries or civil engineering projects. High amplitudes of AOP are usually associated with the lateral movement of a vertical free face, but it can also be caused by the upward lifting and the venting of gases through the surface of the blast area.

Jan 1, 2018

By N. Nikolaeva, A. Romashev, T. Aleksandrova

Energetic characteristics identify energy consumption for destruction of prills and kinetic characteristics rate formation of finished size grade in the course of destruction and identify crushing and grinding equipment capacity. And if screen sizing or classification processes are applied in the cycle of ore preparation, then the use of characteristics of separation to size – separating characteristics – is required for theoretical description. Separating characteristics have a lower profile for ore preparation but, nevertheless, they have to be taken into account especially for balance and design calculations of crushing and grinding circuits. In most cases there are impregnations of minerals which vary widely in energy consumption and destruction kinetics and thus polyminerality must be considered as one of priority factor. Raw materials which passed through crushing and grinding processes can be assessed numerically by functions of fractional composition of two types: differential distribution function y ( €) and assay function B( c ) . A set of experiments on grinding of crushed raw materials at different time (up to various sizes) was conducted to analyze effect of grinding size of carbonaceous overburden on function y € and B ( c) as well as to justify the possibility of preconcentration of raw materials at the stage of ore preparation. Analysis of metals distribution per size grades showed that metals are concentrated in size grades of less than 0.315 mm (which is equivalent of cut point for this type of raw material), however, there is no overgrinding and oversludging of material, and enriched product for its further processing can be excluded with the help of classification. The stage of size segregation was tested as a method of preconcentration of raw materials for optimization of ore preparation and further separation.

Jan 1, 2018

By Magnus Evertsson, Haijie Li, Erik Hulthen, Mats Lindqvist

"The static and dynamic models for autogenous (AG) and semi-autogenous (SAG) mills have been studied for decades. Though the effects of mixing soft and hard ores in a mill and their interactions are not widely investigated. To develop a general numerical model for AG/SAG mills, a dynamic model has been introduced, and its capability to simulate multi component ore breakage has been examined and compared with previous publications. The simulation results show that the dynamic model can cover a range of milling conditions and is able to simulate complex scenarios with soft and hard ores. The dynamic model discussed here could provide substantial potential for the controlling and optimization of the comminution circuits. INTRODUCTION In the last decades, efforts have been devoted to studying the comminution circuit due to their high energy consumption. The energy directly spent in minerals comminution was about 2% globally in the year 2000 (Maurice et al, 2003). As one important part of the comminution process, AG/SAG mills are widely used for primary grinding in minerals processing. AG/SAG mill models include static and dynamic models (Kojovic et al, 2012; Yu et al, 2014; Yu et al, 2017; Liu and Spencer, 2004; Salazar et al, 2009). Many static models have been well validated over the past decades, but they are all based on steady state assumptions. Even though plants experience a steady-state condition under certain circumstances, it can be inaccurate to simulate the system by static models under all circumstances (Asbjörnsson et al, 2013). The advantages of a dynamic model are apparent that it provides more information and insights both inside the mill and through interactions with other machines in terms of pebble crushers, conveyor belts, screens and so on. In many mines, the multi-component grinding is inevitable, and the ratio of hard and soft components in the mill strongly affects mills operation and performance. Bueno (Bueno et al, 2010; Bueno et al, 2013) studied the effects of blending hard and soft ores in AG/SAG mills and used them to update the JKSimMet toolkit which was a steady state plant simulation software. However, due to the complexity of the particle interactions such as multi-component breakage function, appearance function, in mill transportation, there has been no dynamic AG/SAG mill models developed yet for simulating multi-component."

Jan 1, 2018

By P. Asadollahi, M. Invernizzi

"This paper describes the innovative full-face conventional tunnel excavation technology, with its advantages in terms of savings and safety. The full-face approach has been developed throughout the experiences gained in tunnel construction. The innovative design approach is focused primarily on the deformation response of the ground acting first and foremost ahead of the face, which is analyzed, predicted and controlled by pre-confinement of the core. The advantage of this method is particularly evident in difficult grounds. This paper will present the outstanding benefits given by the adoption of the full-face approach in two projects, the “Sochi” and the “Visnove” tunnels. THEORY AND DESIGN ASPECTS OF THE FULL-FACE EXCAVATION The main challenges in tunnel and underground construction industry are gathering the design data from the ground that is being excavated and foreseeing the critical stage during construction, where the materials are in its most critical configuration. These challenges shall be addressed during excavation when the lining and the excavation disturbance have not yet reach the equilibrium. The in situ stresses in the ground are redistributed (forming an “arch-effect”) around the opening, with zones of increased stresses around the walls of the excavation (Figure 1). The distribution of the stresses can be close or far from the excavation profile, depending upon having a ground capable of withstanding the deviated stresses well or unable to withstand them, respectively. In the latter case, the ground will plasticize and deform in proportion to the volume of ground involved in the plasticization phenomenon. When the ground completely cannot withstand the stresses redistribution, it will cause the collapse of the excavation. While in the first case, the arch effect occurs naturally, in the second case, it does occur naturally only when the ground is stabilized properly. In the case of collapse, the arch-effect must be produced artificially operating on the ground prior excavation. The design should be based on the understanding of if and how the arch-effect can be mobilized during tunnel excavation and ensure that is formed by calibrating excavation and stabilization operations appropriately. The arch-effect is function of the medium that is being excavated, the action taken to excavate and the reaction expected following the excavation."

Jan 1, 2018

By Mohammad Mainul Hoque, Linhan Ge, Ai Wang, Geoffrey Evans, Subhasish Mitra

"Dispersed phase interactions involving bubbles and particles in the turbulent environment play a significant role in the flotation process. The conventional collision models only account for the deterministic rather than the stochastic behavior of turbulenct flow. What the bubble-particle collisions would be like remains an open problem due to the complex nature of turbulent flow. To address this research question, the present study aims to quantify the bubble-particle collision efficiency incorporating turbulence using the computational method. Multiple solid particles were released to interact with a single stationary rigid bubble (non-deformable interface). Effect of particles Stokes number, turbulent intensity (TI) and integral length scale on the bubble-particle collision efficiency (Pc) was examined. Large Eddy Simulation method (LES) was used to compute the continuous phase. Fluctuating velocity required at the inlet boundary condition was derived from Gaussian energy spectrum to generate the upstream turbulence. Discrete Phase Model (DPM) is used to simulate particle trajectory. The characteristics of turbulent flow were analyzed through Energy spectrum and energy dissipation rate. The critical initial releasing radius of the annulus, where Pc equals zero, reaches up to 1.5(Rb+Rp) under high turbulent intensity (TI) 20.48% compared to 0.23(Rb+Rp) calculated from the traditional Pc model. Pc decreases as the distance of initial releasing annulus between bubble centre increases. Pc increases as particle Stokes number increase up to 0.66. The collision efficiency for particles of St less than 0.38 is higher under TI 11.78% compared to TI 4%.INTRODUCTION To increase mineral recovery, interactions between bubbles and particles are essential. All flotation devices achieve it through introducing some level of turbulence into the system by increasing specific energy dissipation rate. Figure 1 shows three different flotation equipment, traditional floatation cell, Concorde cell and cyclonic-static micro-bubble flotation column (FCSMC), with different turbulence generation mechanisms – impeller, down-flow liquid jet and pump recirculation, respectively."

Jan 1, 2018

By V. F. Petrov, O. D. Khmelnitskaya, V. M. Lodeischikov

"The processing of copper gold ores is one of the major issues of gold-mining industry in the beginning of 21st century. In this case, application of ammoniacal cyanidation process for this type of mineral raw materials deserves special attention. Irgiredmet Institute has carried out a wide range of studies on ammoniacal cyanidation process and evaluated the possible commercial use of this process for Russian operating plants. Copper gold-bearing ore from a current Russian deposit was studied. The content of gold and copper in the ore is 9.4 g/t and 0.4%, respectively. Based on the results of preliminary test work, a flow sheet was developed. It included gravity concentration followed by cyanidation of the gravity tailings using CIP. The flow sheet provides sufficient total gold recovery (91.5%). However, this technology has disadvantages, such as high specific consumption of reagents: 4.5 kg/t NaCN and 13.5 kg/t Ca(OCl)2 the ore and the increased gold grade in the liquid phase of the tails slurry (0.2 g/t). In order to improve cyanidation process parameters, test work on gravity tails processing using ammoniacal cyanidation was carried out. After the test work was completed, ammoniacal cyanidation flow sheet for copper gold-bearing ores was developed. Leaching conditions were optimized using recycled process solutions. Recommendations on gold recovery from the solutions and tails detoxification are given. The proposed flow sheet provides 91.5% gold recovery. In this case, compared to the conventional option, the consumption of cyanide decreases by 3-4 times.INTRODUCTION The technology of precious metals recovery from gold-bearing ores is based on cyanidation. This accounts for up to 90% gold recovery and the significant portion of silver (Lodeischikov, 1999). However, there exists the category of so called refractory ores. They are not amenable to direct cyanidation (without using additional expensive processes), or they are characterized by low indices of metals recovery and high consumption of reagents. The latter also include copper gold ores. Their deposits can be found extensively both worldwide and in Russian Federation. A feature of these ores is that they contain not only gold but also copper. It is considered not only as a valuable component but also as a harmful mineral impurity that hinders gold cyanidation, gold precipitation from the solutions by cementation onto zinc dust and adsorption onto activated carbon (CIL, CIP-processes), detoxification of cyanide bearing effluents."

Jan 1, 2018