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By E. Yan, M. Mankosa, E. Wasmund

"Rougher flotation in sulfide ore applications is typically performed using conventional flotation machines. The trend over the last 100 years has been for these cells to become larger in volume with cells reaching sizes of 600 m3 and in some cases even larger. The associated foundation loadings, transport and installation requirements along with building size have also increased. In an economic environment where projects must be executed with dwindling capital supply, bigger is not always better. More recently, industry is being challenged to identify new technology that allows for improved flowsheets and financial returns through efficiency gains. To meet this challenge, a novel high-intensity two-stage flotation system known as the StackCell™ was tested to potentially achieve installation volume and required retention time gains when compared to conventional flotation circuits. The results from on-site pilot scale work conducted on typical copper concentrator rougher flotation feed showed retention time gains on the order of 6 to 9 times when compared with conventional flotation equipment. This finding has the potential of reducing capital demands for rougher flotation circuit designs when incorporating the StackCell. The results from test work will be discussed in this paper showing comparative test data generated with a Denver bench batch test, pilot StackCell and large conventional cells, all receiving the same feed.INTRODUCTION Over the last few years, a number of flotation devices have been described as two-stage or multi-stage flotation units. The idea behind multi-stage flotation devices is that flotation performance can be improved by utilizing cells with multiple specialized units in series, but not all units of the same type, as is practiced in trains of multiple tank cells or columns. Instead, each unit is specially designed to provide the optimized fluid environment for each stage of the sequential flotation process. This idea has been proposed based on a study of the fundamentals of the flotation process, and the realization that the process consists of discrete sequential processes, that are each optimized with dissimilar process conditions (Finch et al, 1995), (Zhou, 1996). Instead of executing each step simultaneously in a homogeneous fluid environment, where the conditions would be a compromise between particle collection and froth recovery, the multi-stage flotation unit operates in a set of segregated process environments, each with a set of conditions optimized for that step."

Jan 1, 2018

By Frédéric Flament, Mary Goldman, Gille Barbery

"Pilot plant testwork on autogenous and semi-autogenous grinding was conducted at the Centre des Recherches Minérales du Québec (CRM) to investigate the effect of ball addition, solids concentration in the feed, specific energy consumption and feed size distribution.A mathematical model that incorporates breakage and surface mechanisms of loss of mass was developed to predict load size distribution of autogenous and semi-autogeneous mills. An empirical model was used to predict product size distribution of autogenous and semi-autogeneous mills.IntroductionAutogenous (AG) and semi-autogenous (SAG) grinding circuits offer potential capital and operating cost savings when compared with conventional grinding. Empirical relationships have been developed to calculate specific power consumption for SAG mills using pilot plant test results (Mular et al., 1989; Barrett and Allan, 1986). Empirical models can be obtained from plant practice data but are applicable only to the particular ore and set of operating conditions used in the model. Grinding mechanisms have to be taken into account in order to develop a model that can be applied to different ores and mills.In AG and SAG, the ore itself acts as grinding media, with the finer sizes being crushed by the coarser ones. Surface mechanisms of loss of mass, in which the inner core of the particle remains intact, also take place. In SAG milling, the effect of balls also has to be taken into account. The role of each one of these mechanisms changes with particle size and the model has to account for such changes."

Jan 1, 1991

By L. A. Jameson, J. Awmack, P. J. Lind

"Peñasquito is a polymetallic deposit containing gold, silver, lead, and zinc. The operation is currently the largest producer of gold in Mexico. Over the next few years, the process flowsheet will be expanded beyond the flotation of lead and zinc to include flotation of pyrite from zinc tails and subsequent leaching of the pyrite concentrate. This paper outlines mineralogical leach characterization that has been carried out to improve the understanding of the complex hosting of gold and silver minerals in the ore. Organic carbon that is present in the ore can have an impact on flotation and leaching. Mineralogical techniques used to characterize this deleterious component of the ore are introduced.INTRODUCTION Goldcorp’s Peñasquito operation is located in the Mexican state of Zacatecas, approximately 200 km from the city of Zacatecas as shown in Figure 1. Peñasquito has been in operation since 2009 and is currently Mexico’s largest gold mine. Mining is carried out by open-pit methods, with oxide ore proceeding to a heap leach facility and sulfide ore treated in the sulfide processing plant.The processing capacity of the sulfide plant is 130,000 tonnes of ore per day and in 2016 production amounted to 450,000 ounces of gold, 17.6 million ounces of silver, 109 million lb of contained lead, and 263 million lb of contained zinc in lead and zinc concentrates. GEOLOGY The Peñasquito deposit is hosted in and adjacent to two breccia pipes; Peñasco and Breccia Azul. The two breccia pipes are branches of a larger intrusive-related complex, which is hosted in the sedimentary Caracol formation as illustrated in Figure 2. The regional geology is dominated by Mesozoic sedimentary rocks which comprise a thick series of marine sediments deposited during the Jurassic and Cretaceous periods. A 2 km thick sequence of carbonaceous and calcareous siltstones and interbedded sandstones is underlain by a 1.5 to 2 km thick limestone sequence (Goldcorp Inc., 2015)."

Jan 1, 2017

By D. Sbarbaro

In many modem concentrators the use of supervisory systems in their grinding circuits has meant more consistent operations and increases in capacity. These systems rely on a series of specific algorithms that uses different pieces of information and drives actions over the process. One of these algorithms is the ball mill overload detection, which plays an important role in any supervisory strategy designed to maximize ore feed rate in wet grinding circuits. This paper introduces different overload detection algorithms, and compares them in terms of the time taken to detect the overload condition. This comparative analysis is based on highly noisy power draw measurements obtained in an industrial plant.

Jan 1, 2007

By W. R. Bruce, G. W. Hodgson, K. A. Clark

"IntroductionANY PLANT processing mined Athabaska oil •sand for the recovery of oil produces, as a waste material, about one ton of sand for every barrel of oil recovered. A disposal problem of major proportions is thus indicated. For a 100,000-barrel per day plant, provision must be made for moving .about 4,000 tons of sand per hour on a continuous basis to a tailings dump at least several thousand feet removed from the plant site.Of the many known ways of conveying solids on .a large scale, the hydraulic pipeline transfer method is becoming increasingly important because of convenience and economy of operation. It has been in use for many years in mill operations at metalliferrous mines for the disposal of tailings and as such was a natural outgrowth of •hydraulic sluicing (3). Pipeline transportation of mill concentrates has been adopted recently ( 1), wherein high-density solids are moved several miles by hydraulic pipeline. The flotation power of the water suspending the solids may be increased 'by addition of magnetite for the handling of ore in pipelines (8). Low-density solids such as coal have received considerable study for low-cost pipe-lining (2) and a successful operation with coal has been carried out in England with indications of considerable savings over other methods of transportation (9). Dredge operations are based on pipeline transportation for the dredged sol-ids, and the scale of pipelining there .has involved pipe sizes as large as 32-inch (6). However, the tendency has been to regard the pumping of solids in water merely as a simple method of transportation and little consideration has been given in the published literature to study of the most economical modes of operating .a pipeline."

Jan 1, 1952

By G. Danko, C. Lu

"High temperature and humidity at the working negatively affect safety and health in deep and hot underground mines. It is necessary to predict during design the dynamic changes in temperature with heat and moisture emitted into the mine ventilation-thermal-humidity (VTH) system from various sources such as auto-compression, wall of strata, mining equipment and workers. Currently available mine ventilation and climate models are either furnished with a simplified, step-change-type (SCT) model for thermal solver (such as in Climsim, VnetPC, Minevent, MFire, etc), or with more advanced models to account for the “thermal history effect” (also known as thermal flywheel effect), such as in Ventsim and MULTIFLUX. The goal is to quantify the modelling error caused by the lack of variable thermal history component in the SCT model. Two models are compared: one with and one without the thermal history component in long, single airway examples with long-term climate simulations for 20 years assuming an input air temperature with daily variation. The examples illustrate the magnitude and the time-lag phase of the variation of the modelling errors with space, time and model conditions. Large errors are shown in the order of +14°C and –7°C errors, persistent for long time periods, up to 20 days in one example.INTRODUCTIONIt is important to predict the heat and humidity distributions during the design of hot and humid underground mines for the safety and health of the workers. The feasibility of extracting a rich deposit in great depth may depend on the extent of ventilation and mine climatization for maintaining an acceptable working environment. The cost of ventilation and cooling may be substantial, up to 40–50% of the total energy cost of mining (Clausen, 2017). The accuracy of model solution is important in view of the low profit margin in mine feasibility for deep and hot mines. Safety and health of the miners, the primary concerns, are therefore inherently connected to finding ever better solutions in simulation methods and ventilation control techniques."

Jan 1, 2019

By M. Tardif, F. L. -Lavoie, M. Ruel

"In mineral processing, it is often very difficult to reliably measure several key operation variables. The harsh environment and high temperature encountered in pyrometallurgy processes lead to some difficulties when selecting the instruments, valves and actuators. If an important variable cannot be measured or manipulated adequately, controlling it can really become a big challenge. Sometimes, it is simply impossible to find the perfect equipment for the task. To help overcome these difficulties, a different approach is required. This approach will not correct the inherent problems related to the instruments and valves, but will cope with these problems. This article will demonstrate how modern process control techniques can be used to solve complex problems while remaining easy to implement. In this article we will described how soft sensors and multivariable fuzzy logic control were implemented at Koniambo Nickel to stabilize the furnace feed and allow for safe and reliable operation of an 80 MW DC arc furnace.INTRODUCTION Process Overview The process plant at Koniambo Nickel SAS is located in the North Province of New Caledonia, 310 km north of Noumea. The nickel-bearing ore is mined in the nearby Koniambo massif. After being screened and crushed, the ore is transported to the processing plant using a 12 km conveyor. After stacking in two blending piles, the ore, containing 2.3% nickel, is fed to the processing plant where it will go through the NST process (Nickel Smelting Technology) before being sent to refining and shotting. This article will focus on the improvement to control of the NST processes. The first steps of this process consist of milling the ore and removing all moisture through calcining (Figure 1)."

Jan 1, 2017

By R. W. Boyle, W. Dyck

"Scientists throughout the world have observed radioactive disequilibrium in the uranium decay series. Radiometric and fluorometric determinations of U in plants from the Goldfields district and in sediments and rocks from the Beaverlodge district exhibit marked radioactive disequilibrium. The eU/U ratio in plants rooted in mineralized ground shifts from greater than unity to less than unity for those rooted in background soils. A similar shift is observed in the transition from rocks to stream sediments and to lake sediments in the Beaverlodge district.Results of a Rn-Ra-U survey of lake bottoms in the Key Lake district show that anomalous hydromorphic U from U mineralization is parallelled by anomalous Rn, even though there is much less Ra in the lake-bottom sediments than one would expect from the amount of U present. IntroductionDisequilibrium in the U decay series has been studied for many years and is observed in almost all surficial environments. Rosholt (1959) measured U, 234Pa, 23O'fh, 226Ra, 222Rn, and 210Pb in a large number of U ores and outlined six types of situations to explain disequilibrium. Davy (1974) outlines the geological significance of radioactive disequilibrium situations with respect to exploration. Levinson and Coetzee (1978) discuss possible mechanisms by which separations of parent and daughter elements are affected and suggest analytical pitfalls that may be encountered in the use of radiometric techniques when disequilibrium has occurred. Generally, it is the oxyphile and easily complexed parent 238U (with its associated isotopes 235 U and 234U) that is leached from rocks and soils at a faster rate than are the daughters, 230Th, 226Ra, 2loPb, and 206Pb. This mobilized U is adsorbed on sediments in streams and lakes, or enters groundwaters, where it is eventually precipitated under reducing conditions, occasionally in oregrade concentrations. This surface leaching and"

Jan 1, 1980

By Marsh A. Cooper

THE THEME of my remarks provides a great deal of latitude. I shall therefore take advantage of the scope it offers to look at our industry across a broad spectrum. My observations reflect a Canadian mining man's varied career as a geologist, mine operator, geological consult ant and company executive in Canada and overseas ... under differing cultures, business environments and political jurisdictions. I was fortunate to come into this industry in the mid-1930's when it was characterized by great enthusiasm, drive and resourcefulness. It was the beginning of a burgeoning era of some 30 years' duration in which the individualist, the entrepreneur and the builder found ample room for expression. As I contemplate the environment in which we now have to operate and try to perceive what lies ahead, I find myself deeply disturbed. Something serious has happened to this industry along the way. It appears that our very success has been the source of our undoing. I believe it is useful to summarize our industry's basic strength at home as a take-off point f or assessing its position in world terms. I hope that I can focus on some of the things that, in my view, must be done for this industry and by this industry to restore Canada's historic role in the world mining community. The Canadian mineral industry provides 140,000 jobs. It pays the second -highest average weekly wages and salaries among the country's major industries. The output per employee in the mining industry is estimated at about three times that for the economy as a whole.

Jan 1, 1977

By J. J. Crowder, A. L. Coulson

Ground support is a significant component of underground mining costs. The rehabilitation of failed support, however, is often much more expensive than the initial installation. This paper investigates whether the right type and length of support can be designed correctly the first time, through forward modelling of the support and the mining process. Back analyses demonstrate how high-quality instrumentation and monitoring can help with the estimation of the post-peak (post-failure) rock mass parameters required for a comprehensive analysis. A hybrid modelling approach combines a three-dimensional mine-wide boundary element model with a non-linear finite element model to allow for the failure process in the rock mass. The back analyses demonstrate that, for the case study site, the post-peak Generalized Hoek-Brown value of mr must be reduced significantly in order to properly simulate displacements observed in instrumentation data. High-quality instrumentation data is the key to post-peak rock mass parameter calibration. The procedures outlined in the paper highlight the need for a new hybrid numerical modelling package that can use the influence of three-dimensional mining-induced stresses to predict displacements in, and potential failure of, rock masses surrounding mining infrastructure in order to radically alter strategic and tactical mine design.

Jan 1, 2007

By E. H. Hall

"Perhaps .I should start this talk on Arsine poisoning, by giving a brief background and history of the Bralorne Mill.The present Bralorne company started milling operations in 1932, with amalgamation and flotation, and operated continuously until 19 61 . In 19 59, it was decided to change over to cyanidation, and to use double filtration rather than counter current. The cyanide portion of this new plant was completed and went on stream in August 1961 utilizing the grinding circuit of the old flotation plant. The new grinding circuit was completed and started operating December 15, 1961.The grinding circuit consists of a 7 1 x 101 Hardinge Rod Mill, and a 7 1 x 101 Hardinge Ball Mill. In series with the rod mill is a 16"" x 24"" Denver Duplex Jig, and two of these jigs are also in parallel with the ball mill. The rod mill product is classified by a Kreb cyclone, the oversize going to the ball mill. The ball mill is in closed circuit with this same cyclone. The jig products are amalgamated daily, and the amalgam barrel tails are returned to the ball mill for regrind. Arsenious oxide (As203) is one of the reagents used in amalgamation. Approximately 39 ozs. per day is used.The cyanide circuit consists of a 481 x 121 thickener, three 24' x 24' agitators, two 141 x 161 drum filters and a Merrill Crowe precipitation unit.The Krebcone overflow is thickened to 60 - 62% solids. Hydrated lime, along with a settling agent is added to the thickener. Some of the overflow water from the thickener is returned for use in the jigs. The balance of this overflow water goes to waste. To the thickened pulp entering No. 1 agitator, barren solution is added, along with sodium cyanide and hydrated lime. Cyanide strength is held at 0. 9 - 1. 0 lbs/ ton solution and lime at 0. 8 lbs/ ton solution. The lime content usually drops to 0. 5 lbs/ ton in No. 3 agitator. The two drum filters are in series and are washed with barren solution. The discharge of No. 3 agitator is fed to No. 1 filter and the discharge of No. 1 filter is repulped in barren solution and feeds No. 2 filter. No. 2 filter discharges to tails, The filtrate is clarified, deaerated and the gold precipitated. One hundred to two hundred tons of barren solution are bled to waste each day. The ore contains 1 % Arsenopyrite and 1. 5% Pyrite."

Jan 1, 1964

By T. Verdel, Y. Gunzburger

Rockfalls are major and essentially unpredictable sources of danger, particularly along transportation routes (roads and railways). Thus, the assessment of their probability of occurrence is a major challenge for risk management. From a qualitative perspective, it is known that rockfalls occur mainly during periods of rain, snowmelt, or freeze-thaw. Nevertheless, from a quantitative perspective, these generally assumed correlations between rockfalls and their possible meteorological triggering events are often difficult to identify because i) rockfalls are too rare for the use of classical statistical analysis techniques and ii) all intensities of triggering factors do not have the same probability. In this study, we propose a new approach to investigate the correlation of rockfalls with rain, freezing periods, and strong temperature variations. This approach is tested on three French rockfall databases, the first of which exhibits an outstandingly high frequency of rockfalls (approximately 950 events over 11 years), whereas the other two databases are more typical (approximately 140 events over 11 years). These databases come from (1) the national highway RN1 on La Réunion Island, (2) a railway in the Bourgogne-Franche Comté region, and (3) a railway in the Auvergne region. Whereas a basic correlation analysis is only able to highlight an already obvious correlation in the case of the “rich” database, the newly suggested method appears to detect correlations even in the “poor” ones. Indeed, the use of this method confirms the positive correlation between rainfalls and rockfalls on La Réunion Island database. This method highlights a correlation between cumulative rainfalls and rockfalls in the Bourgogne region, and it detects a correlation between the daily minimum temperature and rockfalls in the Auvergne region. This new approach is easy to use and also serves to determine the conditional probability of rockfall function of a given meteorological factor. This approach will help to optimize risk management in the studied areas based on their meteorological conditions. Nevertheless, even if some rockfalls are correlated to meteorological factors, other seems to occur independently of these factors, which highlights the unpredictability of rockfall hazard.

Jan 1, 2015

By K. O. J. Sidwell

STRATMAT LIMITED, through its subsidiary, the Strategic Manganese Corporation, controls two adjoining iron-manganese properties five miles west of Woodstock, New Brunswick. One is a staked property of 77 claims traversed by one trunk highway, several secondary roads, and a branch line of the Canadian National railway. The other property, a 20-square-mile Concession immediately to the north-east, is also traversed by an excellent network of roads. Both proper-ties are situated in gently rolling farmland. The deposits were discovered in 1836 by Dr. C. T. Jackson. The first attempt to exploit them was made in 1848 when a blast furnace was erected with a capacity of 7 tons per day. The ore was mined by a series of deep trenches that followed the folds of the enclosing slates. In the late years of operation, bog iron ore from Burton and Maugerville, in Sunbury county, was used in conjunction with the Wood-stock ore. An additional 5-ton furnace was erected in 1863 and mining was carried on intermittently until 1884. No further extensive work was done until 1953. During the period of activity, approximately 70,-000 tons of ore are said to have been smelted ( 1). During these earl y operations the Iron Ore Hill and Moody Hill deposits were partially opened, with the ?bulk of the furnace feed being mined from the Iron Ore Hill workings. The low tenor and apparently limited extent of the deposits, the high. ?content of phosphorus, and the increasing scarcity of fuel, were the chief factors that brought the enterprise to an end.

Jan 1, 1957

By J. Hedderson, J. Starkey, S. Reeves, B. Hilscher

"As ore sorting becomes more common, it is important to quantify its impact on comminution circuits in order to precisely predict its effect on throughput, grind, and mill operating conditions. Since sulphides and other valuable minerals are often softer then the surrounding waste rock, the ore kept by a sorting system is typically softer than the average ore. In order to accurately forecast production, new measurements of hardness need to be performed on the new concentrated mill feed. This paper proposes a methodology to examine this effect quantitatively at an early stage of project development using the SAGDesign test.INTRODUCTION The application of ore sorting in the mining industry is one of the most promising recent technologies with the potential to cause step change in the way concentrators are designed. Most typically, ore sorters are used to sort a given ore flow for grade, either to reject waste from the mill feed, or rescue ore from the waste pile. Ore sorting technologies have been successfully applied at many mining properties in various forms and while the technology has reached or is reaching maturity in other industries, it is still in development in the mining sector. As a result, there is not yet a general consensus on the considerations that need to be made when designing a plant with ore sorting from a feasibility level above and beyond what is commonly performed as best practice. This paper will discuss those considerations and propose a methodology for integrating ore sorting into a comminution circuit at the design phase. Due to the linearly additive nature of the SAGDesign Test results, a “hardness balance” can be developed to estimate the ore hardness which would be measured in any sorting mass split using only two data points. This will allow a design engineer to make more informed decisions as to the most appropriate mass split to select in relation to the incurred operating costs and grinding circuit sizing."

Jan 1, 2019

By D. Lauzon, H. de Ruiter

"In 1988a decision was taken to design and install a computerized mine monitoring and control system. This system had to monitor and control all pumping and ventilation underground. The operation of the mine air heating, using waste heat from the mine water and compressors, was to be part of the system. It ultimately also had to interface with the four existing Canadian General Electric Series 3 PLC's which controlled the No .3 shaft loading pockets. Beside operational control and trouble shooting, the system is a valuable tool in energy management and ultimate savings are estimated at $50,000 per year.The final equipment selection was a General Electric Series Six PLC on surface with a total of 17 Series One Plus and Juniors PLC located underground. Data is transferred through a two-pair shielded cable using modems. Both analog and digital functions are being utilized.The basic monitoring and control system was commissioned during 1989and, due to a higher level of technology today, further improvements have been planned for the future. The system was designed by Walter Dow together with Gordon Champion of D.G. Champion Engineering Ltd. of Toronto. Advice for the ventilation part of the project was obtained from Maurice White, Independent Ventilation Consultant.The Ministry of Energy, Mines and Resources and the Ministry of Energy of Ontario approved this project together with the waste heat recovery project to qualify for a grant of$135,000 under the Enerdemo-Canada program.IntroductionThe Macassa mine is a narrow vein gold mine situated about 2 km west of the town of Kirkland Lake in northeast Ontario. The mine has been in continuous operation since 1933 and a total of about 3.0 million ounces have been recovered out of 6.6 million tons of ore ."

Jan 1, 1991

By Robert M. Vadas

Special considerations are described as they apply to the design of ring gears for high-power grinding mills. AGMA requirements, the importance of the supporting structure, material selection and related field experience are discussed, with special emphasis on available and developing manufacturing technologies.

Jan 1, 1978

By J. Bouchard, A. Clapperton, L. Ebacher, N. Singh

"Canadian Royalties Inc. owns and exploits the Nunavik Nickel project: a copper and nickel deposit in Nunavik, Quebec. Froth flotation is used to produce nickel and copper concentrates. As is the case at other operations, the presence of magnesium is detrimental to the flotation products. In the past year, the exploitation of a new deposit has increased magnesium in the feed. Thus, it has been challenging to meet the required ratio in the final nickel concentrate. In search for a consistent solution for this problematic, an empirical flotation simulator making use of selectivity curves was developed. The circuit was expressed as a system of linear equations, which considered the distribution of copper, nickel and gangue minerals in the different streams. The relative quantities of each mineral are given by the separation efficiency curves of the various banks in the circuit. Three sources of data were used to calibrate the simulator: historical plant data (metallurgical balance), infinite dilution bench-scale flotation tests, and direct sampling of flotation cells. The calibrated separation curves were used to predict the recovery of the different species based on a single mineral recovery. Validation of the simulator showed good fit for the valuable minerals. The simulator was then used to infer the optimal operating conditions of the flotation circuit to optimise magnesium oxide content of the nickel concentrate. Simulation of two re-arrangements of the circuit gave a promising solution with a Fe:MgO ratio 2.42 times higher than the minimum value.INTRODUCTION Canadian Royalties Inc. owns and exploits the Nunavik Nickel project: a copper and nickel deposit in Nunavik, Quebec. Figure 1 shows the location of the mine site. After crushing and grinding, froth flotation is used to produce nickel and copper concentrates. The flotation circuit is designed with multiple stages consisting of bulk copper/nickel flotation, copper/nickel separation, and a separate nickel circuit (bulk float and cleaning)."

Jan 1, 2018

By Michael Boyle, John Zeni

"The Blind shaft boring method is a mechanised process of excavating large diameter vertical shafts downward from the surface. It is a rotary drilling technique eliminating any requirement for personnel to enter the shaft or access underground workings during drilling and lining phases. Excavated material is delivered to surface carried by a flow of circulating fluid through the centre of drill rods. The drill assembly is suspended from a large A-frame rig and powerful winch at ground level providing vertical thrust to a hard rock cutter head. INTRODUCTION The Southern Coalfields Ventilation Shaft was constructed at a mine, 45 km north-west of Wollongong. The 5 m finished diameter (6.2 m drilled) shaft reaches a final drilled depth of 517 m, the deepest, fully lined blind bored vent shaft to date in Australia. Operated by a mining company in the Southern Coalfields, the expanding mine required a new ventilation shaft and associated surface infrastructure to service future operations. An assessment of future ventilation requirements for the coal seam operations identified a need to improve the capacity and reliability of the existing underground ventilation system in order to ensure a safe and efficient underground working environment. Installation of the new upcast ventilation shaft was determined to be the most cost effective and reliable solution to provide adequate ventilation for underground operations. The completed shaft is smooth and fully lined with steel/concrete composite liners that are hydrostatically sealed. The shaft lining is designed to meet the mine ventilation requirements for a service life of 20 years. The shaft intersects the top of the working section of the coal seam at approximately 513 m of depth. This includes the shaft collar depth and shaft bottom profiled intake section.PROJECT METHODOLOGY AND INDUSTRY BEST PRACTICE The shaft was drilled using the blind bore method, an innovative construction technique and an alternative to the traditional methods of sinking deep and large diameter shafts, such as raise boring and conventional drill and blast shaft sinking. The blind bore method allows the shaft to be drilled from the surface down without anyone having to enter the shaft or the mine during the construction process. This is by far the safest method of shaft construction and is especially suitable for difficult ground conditions where there may be unstable or water-charged strata. As this method involves drilling from the surface downward, shafts can be completed ahead of the underground mine development. Thus, allowing underground operations to focus on development and production without the conflicts of managing ventilation priority for underground contractors and mucking out spoil from excavation workings."

Jan 1, 2019

By S. J. Prout, J. M. Lang, D. Lascelles, K. Sarinas

"There is a growing demand for ore characterization methods to support metallurgical development programs from a true geometallurgical perspective. Today, there are multiple automated mineral analysis systems on the market, many of which are based on conventional mineral liberation data. However, liberation data can be limited when assessing the floatability of a mineral to predict its recovery as it doesn’t consider the surface area available for collector absorption and bubble attachment; whereas mineral exposure measurements can. Using the First-Order Rate equation to predict the recovery versus time of variously exposed particles, a model has been developed to predict the recovery of copper sulphides at a particular grind size. This model provides the opportunity for the prediction of copper recovery and grade from multiple, small-scale samples. These predicted values can be imported into resource modeling software, and guide the domaining of geometallurgical units within the ore body. Data has been analyzed from various global copper porphyry systems and compared with actual flotation testwork.IntroductionAs part of the development of a mining project, resource estimation primarily focuses on the definition of the quantity and grade of the ore in the deposit; as the project evolves from preliminary economic assessment through to feasibility, the geological model is usually further populated with hardness and recovery data.The CIM Definition Standards (2011) identifies measured mineral resource as the part of a mineral resource for which quantity, grade or quality, densities, shape, and physical characteristics are estimated with confidence sufficient to allow the application of “modifying factors” to support detailed mine planning and final evaluation of the economic viability of the deposit as an asset. The modifying factors include, but are not restricted to, mining, processing, metallurgical, infrastructure, economic, marketing, legal, environmental, social and governmental factors.With market fluctuations, many investors in the mining industry have become more and more risk averse which has increased the demand for more reliable data to support these modifying factors. This has led to the widespread use of geometallurgy and statistical analysis to better inform the resource estimation. Geometallurgy is used to characterize ore variability and is especially useful for highly variable or strongly zoned deposits, several close deposits that will be milled together, projects requiring complex or new metallurgical approaches, brownfields or sight-of-mine exploration, expansion projects that exploit new, deeper or adjacent reserves, and projects with significant legacy drill core that are being re-evaluated due to new economic circumstances."

Jan 1, 2018

By R. Rumpold, J. Antrekowitsch

"The continued increase of the automobile production leads to a higher platinum, palladium and rhodium (PGM) demand. Due to the high value of the PGMs the recycle of these metals from automotive catalysts is of special interest. Pyrometallurgical processes, which are only focused on the PGMs, represent the state of the art of automotive catalysts recovery. However, hydrometallurgical recycling processes offer additional recovery of other valuable metals and compounds. The method presented in this paper is based on a combination of hydrochloric acid and hydrogen peroxide to dissolve the valuable metal fraction which can be precipitated afterwards.INTRODUCTIONPlatinum and the corresponding metals from the PGM group, especially palladium and rhodium are well known as catalytic active materials. They support chemical reactions which are inhibited. This effect is used to decrease the amount of harmful compounds like hydrocarbons, carbon monoxide or nitrogen oxides in automotive exhaust gases. The generation of these toxic substances is based on incomplete combustion, which in turn is caused by non-ideal conditions in the combustion chamber. Catalytic converters minimize the harmful compounds by supporting the following Reactions 1 – 3 (Votsmeier et al., 2009).CO + ½ O2 ? CO2 (1)CnHm + (n + ¼ m) O2 ? n CO2 + ½ m H2O (2)2 NO + 2 CO ? N2 + 2 CO2 (3)Platinum and palladium are very effective for the oxidation (Reactions 1 and 2), while rhodium supports the decomposition of nitrogen oxides by carbon monoxide (Votsmeier et al., 2009). The classical three way catalyst was a platinum – rhodium combination with a Pt:Rh ratio of 5:1 (Hagelüken et al., 2005). Today the platinum, palladium and rhodium ratio is highly variable and strongly depends on the engine type as well as on the car producer. Diesel engines are operated with a super stoichiometric air addition in comparison to spark ignition systems. Therefore the exhaust gas composition and off gas temperature is different. Based on the high oxygen amount, diesel engines produce less carbon monoxide and hydrocarbons but more nitrogen oxide in contrast to gasoline engines. This leads to various requirements for the catalytic converters."

Jan 1, 2012