Search Documents

By Glenn Norris, Tyler Buckingham

"Gold recovery gains in the order of one to two percent have been realized in 2009 due to changes in the operation of the grinding circuits at Barrick Gold Corporation’s Hemlo - Williams Mill. The initiative known as “split milling” involves milling low grade open pit ore in one grinding circuit and higher grade underground ore in the other grinding circuit under different conditions.The philosophy of split milling involves processing higher grade underground ore at a lower throughput rate and a finer grind while milling lower grade open pit ore at a higher throughput rate and a coarser grind. Bench scale tests showed that the recovery relationship for underground ore sources was much more grind sensitive than that of open pit ore. This relationship proved to hold in the plant. The gold recovery gain due to finer grinding of the higher grade ore is much greater than the recovery loss due to coarser grinding on the lower grade ore. Since the majority of gold is being recovered at a higher rate, this results in higher overall plant gold recovery when compared to the traditional operation philosophy of blending the ore sources.The challenges for achieving the maximum benefits of split milling were two fold; the need to increase throughput on the open pit ore milling circuit and a means of grinding finer on the underground circuit. The former was met with an increase in SAG (Semi-Autogenous Grinding) mill ball charge, a basic “drill to mill” blasting campaign and changes to the primary gyratory crusher. The latter was met with a reduced SAG mill ball charge and a higher milling density. The resulting benefits in additional gold recovered are in the order of $3-5 million per annum.INTRODUCTIONThe Williams Mill processes ore from three different ore sources, the David Bell Mine (DB), the Williams Underground Mine (WUG) and the Williams Open Pit (OP). All three ore sources are located within the Hemlo camp in Northwestern Ontario adjacent to the Trans-Canada Highway. The Williams Mill was built in the mid 1980’s at which time it was designed to process 3,000 dry metric tonnes per day through a 22’ SAG mill in series with a 16’ ball mill. In 1988 a Mill expansion doubled the mill capacity by adding a second identical grinding circuit. A photo of the current grinding circuit can be seen in Figure 1. Today the Williams Mill processes an average of 8,000 tonnes per day of which approximately 60% is from the Williams Open Pit."

Jan 1, 2010

By L. J. Heaslip, J. M. Toguri, T. R. Meadowcroft, R. L. Addinall

"""Pilot Plant Operations"" is a final-year core course given to the Metallurgy and Materials Science students at the University of Toronto. This course involves a laboratory study of an industrial project at an industrial site. Details of this program are discussed and reviewed after eight years of experience. IntroductionEducators in the field of Metallurgy and Materials Science have long agreed that ""hands-on"" experience is an important component in the education of metallurgical engineers. This philosophy is reflected in the undergraduate curricula of most metallurgical schools. The major problems encountered with laboratory-associated courses are the requirements for equipment, space and funds for both purchase and maintenance of equipment. For example, consider an exercise in crushing and grinding treatment in a mineral processing laboratory. Such operations require the handling of tons of material and often equipment of large size. The problems of storage of ore, methods of handling and maintenance of equipment are a few of the concerns facing the department. Considering that metallurgical departments at universities are, in general, relatively small, the support to purchase and maintain largescale equipment becomes an annual budgetary concern. In reality, universities today are not geared to operate mini-pilot plant equipment, either in terms of space or personnel.For the above reasons, in 1972an undergraduate core course called ""Pilot Plant Operations"" was introduced to the fourth year curriculum in Metallurgy and Materials Science at the University of Toronto. This course consists of I to 2 full weeks (8 hrs per day) of pilot plant laboratory study at an industrial site on a project of mutual interest with the industrial organization whose facility is being used. The first 5 years of the program were carried out at the pilot plant facility at the Ontario Research Foundation; Stelco's Research Centre in Burlington, Ontario, was used for the last 3 years. The nature of the projects undertaken during the 8-year period is shown in Table 1."

Jan 1, 1980

By Sunil Koppalkar, Liming Huang

"Since late 1980s the research group led by late Professor Andre Robert Laplante at McGill University undertook fundamental studies and applied research of the Knelson Concentrator (KC) and developed a sustainable method to correctly estimate gravity recoverable gold (GRG). In the early time a concept of ‘free gold’ was used, it was defined as the gold that could be separated and recovered from gold ore by amalgamation. Due to health and safety hazards, its practical use had declined. Around 1994, a methodology using a 3-in Knelson Concentrator to measure gravity recoverable gold was established at McGill University. Through more than 10 years’ improvement and optimization, this technique had become a standard method to evaluate GRG for gold ores, gold plant streams and platinum group metals. Until 2006, about 200 GRG tests had been carried out at McGill University for the ores from all over the world. Recently, a modified version of GRG test (short test) with the advantages of reduced processing time and lower cost had been developed for the mundane needs of gold industry.INTRODUCTIONGold gravity recovery has evolved significantly over the past twenty-five years, mainly because of the advent of the Knelson Concentrator (Knelson 1988, 1990, 1992, Laplante 2000). Now, a series of Knelson Concentrators for different applications are manufactured and used in gold industry all over the world (Knelson Concentrators 2004).In 1980s the Knelson Concentrator gained acceptance from mining industry as an effective device for the recovery of gold from grinding circuit. Knelson Concentrator became the obvious choice because of its design features: a relatively long processing time without concentrate discharge (compared to some other centrifuges, such as Falcon Concentrator and Chinese Yunxi Centrifuge at that time), large throughput capacity, treating a wide size range of material, yielding a high grade concentrate with small mass that could easily be smelted. As more and more Knelson Concentrators were accepted in gold industry, mineral processing engineers faced with challenges: what are the separation mechanisms of the Knelson Concentrator? How much gold can be extracted for an ore if a Knelson Concentrator is installed? For answering these questions, a laboratory scale Knelson Concentrator was created."

Jan 1, 2007

By P. P. Andrieux

The recovery of the Zone G crown pillar, located between the bottom of the open pit and the top of the underground workings at the Raglan Katinniq Mine in Nunavik, started in the summer of 2008 and should be completed by the end of 2009. This article describes the rock mechanics-based approach implemented to design the recovery of this valuable block of ore. The stereophotography-based ShapeMetriX3D technique was first used to provide a comprehensive assessment of the various joint sets that comprised the pillar. Then, the statistical distributions obtained for the dip, dip direction, spacing and persistence of these sets were used in conjunction with the geometry of the large-scale faults present in the area to build and run a series of explicit rigid numerical models. Each of these models had a slightly different jointing geometry, randomly constructed from these statistical distributions. Incorporating in a properly formulated numerical model the variability of some of the input parameters has allowed the engineering team to assess in a rational and reliable manner the likelihood of success of the proposed mining method and sequence.

May 1, 2009

?? Intro ?? Mass blast principle ?? LaRonde evolution toward mass blast ?? Overview ?? Mining method ?? Longitudinal ?? LaRonde evolution (2002 to today) ?? Drill and blast design ?? Consumptions and statistics ?? Mucking strategies ?? Case study ?? Examples ?? Conclusion ?? As underground mines extends to greater depths, few questions must be asked: ?? How to: keep operations safe eep operation effective reduce the cost maintain and control the dilution ?? Geomechanic method and practices; significant advances have been observed in the field of: ?? ground support ?? mining sequencing ?? Drilling and blasting; research and development has been less extensive, probably due to the following characteristics: ?? Drill and blast are mainly production-oriented ?? Drilling and blasting are more oriented toward productivity ?? Technology relies heavily on suppliers ?? Highly stressed ground conditions ?? generally help the explosives to break the rock for fragmentation ?? intensify over-breaks or blast-induced

Feb 1, 2008

By D. Simon

This paper studies the effect of binder type and content on the early age development of cemented paste backfill (CPB). The early age development of the mixtures is analyzed using scanning electron microscopy (SEM) and energy dispersive spectroscopy (EDX). These tests show that the early age development of CPB is sensitive to changes in the paste composition. The properties of the CPB depend on the cement content and hydration time. The quantity of hydration products increases with hydration time for all of the mixtures containing cement. However, the high cement content pastes (at least 50% PC) exhibit complete coverage of the PC grains and mine tailings by C-S-H, while the low cement content paste shows scattered coverage of the PC grains and mine tailings and a higher void ratio.

May 1, 2007

By Keith R. Vining

Solid fuel, usually coke or anthracite, is commonly added to ground iron ore fines in the pelletising process to reduce the amount of burner fuel required and the maximum firing temperature, while maintaining pellet strength and the productivity of the pelletising plant. However, the use of coke and coal contribute to the production of greenhouse gases (eg carbon dioxide) and are a source of NOx and SOx emissions. Utilising wood biomass / wood char is one methodology CSIRO Minerals is investigating as an alternative to coke and coal, as it is a renewable energy source that is very reactive, has a low ash content and the potential to reduce emissions of greenhouse gases from the iron and steel-making process. Previous work has established that charcoal can effectively replace coke in the sintering process (Lovel et al, 2005), and the following experimental work aims to demonstrate that it can also provide an adequate replacement for coke and coal in the pelletising process. The paper presents the results of an experimental testwork program comparing the behaviour of fired pellets containing traditional solid fuels and charcoal. The physical properties of the pellets are compared in terms of their compressive strength, tumble and abrasion indices and the reduction properties. A qualitative assessment of the pellet microstructures is also presented.

May 1, 2007

"The measurement of error in assays collected as part of a mineral exploration program or mining operation historically has been undertaken in a variety of ways. Different parameters have been used to describe the magnitude of relative error, and each of these parameters is related to the standard measure of relative error, the coefficient of variation. Calculation of the coefficient of variation can be undertaken in a variety of ways; however, only one produces unbiased estimates of measurement error: the root mean square coefficient of variation calculated from the individual coefficients of variation.Thompson and Howarth’s error analysis approach has also been used to describe measurement error. However, because this approach utilizes a regression line to describe error, it provides a substantially different measure of error than the root mean square coefficient of variation. Furthermore, because regression is used, Thompson and Howarth’s results should only be used for estimating error in individual samples, and not for describing the average error in a data set. As a result, Thompson and Howarth’s results should not be used to determine the magnitudes of component errors introduced during geochemical sampling, preparation, and analysis.Finally, the standard error on the coefficient of variation is derived, and it is shown that very poor estimates of relative error are obtained from duplicate data. As a result, geoscientists seeking to determine the average relative error in a data set should use a very large number of duplicate samples to make this estimate, particularly if the average relative error is large."

Jan 1, 2007

By J. Châteauneuf, R. Dupéré, D. Mrabet

"In gold processing plants, it is common to start gold cyanidation in the grinding circuit by simple cyanide addition or as a result of recycling residual cyanide. In such cases, advantages such as a decrease in gold lockup and reduced leach time have been noticed. The prevailing conditions in grinding circuits such as intense mixing, particle breakage and friction or high temperature are quite different from those found in tanks and so is the resulting cyanidation behaviour. To gain a better understanding of the cyanidation behaviour in mills and of the pertinence of cyanide addition in grinding circuits, lab tests were conducted using an industrial ore from Canadian Malartic mine. In this study, the effects of different parameters on cyanidation in grinding circuits were examined; parameters such as such as temperature, dissolved oxygen, cyanide concentration and pH. The results showed that the kinetic benefits of high temperature in the mill can overcome the low dissolved oxygen concentration thus leading to an overall gold dissolution kinetics improvement for temperatures up to 50°C. However, gold leach in the studied grinding circuit at lab scale caused an increase in cyanide consumption with a higher generation of iron cyanide and SCN-. These results were confirmed by tests in the Canadian Malartic plant where the cyanide addition was moved from the SAG mill to the leaching tanks resulting in a 20% savings in cyanide consumption.IntroductionCyanidation in grinding circuits can benefit gold processing by decreasing gold lockup and by reducing leach time. However, other aspects regarding gold and cyanicide dissolution come into play in the conditions prevailing in grinding circuits. Indeed, high turbulence, high temperature, recycling load, friction and particle breakage make cyanidation behaviour in grinding circuits different from in-tank cyanidation (Marsden and House, 2006). In several cases, cyanide in grinding is limited to a very low level or completely avoided such as in circuits with gravimetric separation, for ores requiring oxidative pretreatment or for preg-robbing ores. In other cases, water management imposes a certain recycling of residual cyanide. This problem is generally addressed by a case-by-case approach. However, the chemistry of cyanidation in mill circuits is worth investigating in order to assess the advantages and disadvantages."

Jan 1, 2016

By E. Pribonic

Failure of a mine material hoist can cause injuries, fatalities and significant infrastructure damage, along with an associated loss of production. Principal causes of hoist accidents are overwind and underwind. Past and present means of arresting hoisting equipment are marginally effective and utilize destructive systems that are not reusable following an arrest. The use of permanent magnet eddy current arrestor technology is proposed; the history and fundamental science of the technology are presented, followed by real world examples and calculations of performance for failsafe, friction-free, instantly reusable eddy current arrestors as applied to stopping a 30-ton ore skip traveling at 16 metres per second. A comparison of existing arrestor types with the eddy current arrestor is included and limitations of the ECA are also discussed.

Jan 1, 2008

By Turgut Yalcin

"Recent years have seen a resurgence of interest in flotation cell design, and one of the areas where attention is being focused is flotation in centrifugal fields which aims at increasing the flotation rates by forced bubble-particle contact.A new cycloflotation device was developed at Laurentian University and tested on the nickel ore of Falconbridge Limited in Sudbury. In comparison with conventional flotation, cycloflotation yielded at least seven times faster flotation rates in terms of the mass of solids floated per unit time per unit flotation volume, while producing concentrate grades similar to those of conventional flotation. These results, which were obtained with a bench-scale ""cyclofloat cell,"" suggest that cycloflotation has the potential of setting a new direction in flotation technology.IntroductionFroth flotation is no doubt the dominant separation technique in the mineral processing industry. Since its inception early in this century, much research effort has been and is still being expended on this process. Until recent years, emphasis seems to have been placed on the chemical aspects of flotation, but there is now an increased awareness for the need to improve flotation machine performance and more and more attention is being directed towards this area. In a 1982 article on ""Flotation Machines"" (Young 1982) it is stated that "" ... by now the design of the machines used would have reached a high level of perfection, ... In fact, this is not the case ... "".The process of flotation in a flotation device may be considered to take place in several distinct stages as follows:"

Jan 1, 1994

By Andy Fortin, Jeet Basi, Simon Hille

Goldcorp’s Eleonore mine is situated in Ell Lake area in the northeastern part of the Opinaca reservoir in the James Bay region of Quebec. The design of the 7000 t/d process plant includes crushing, grinding, bulk sulphide flotation, separate concentrate and tails leaching and detoxification circuits and residue handling. Detoxified flotation tails are thickened and filtered. A portion of the filtered tails are blended with the thickened detoxified sulphide concentration for paste backfill. The remaining filtered tails are trucked to the waste management facility. This presentation will describe the challenges faced with the commissioning and subsequent ramp-up of the flotation tails filter plant and the successes achieved by the multi-discipline team at Eleonore.

Jan 1, 2016

By Bryan M. Maybee

The scheduling of underground mining activities is a complex task, in particular because of the requirement to satisfy a large number of constraints. A Schedule Optimization Tool (SOT) for underground mine planning is described, and its use for investigating strategic decisions is illustrated by means of a case study. The main components of SOT include: a genetic algorithm, which searches for optimal schedules; guidance, the application of heuristic rules to guide the genetic algorithm; and sliding, which implements just-in-time development. For the case study, SOT is employed to generate schedules at various ore capacities, accounting for the appropriate financials in each schedule evaluation. The case study finds that preliminary scheduling performed before invoking SOT, intended to provide a good starting point for the search, does not ultimately lead to higher value schedules. Two key uses for SOT are identified. A mine planner can save time by using SOT to produce high-value scheduling alternatives. This process ensures a broader investigation of the value that the deposit can yield. Also, SOT can be used to validate strategic and tactical decisions, by scheduling under alternative scenarios. Both uses produce high value schedules conforming to all imposed resource constraints.

May 1, 2009

By Tad S. Golosinski

The paper describes Online Master of Engineering Program at the University of Missouri-Rolla and discusses the experience gained during the first three years of its offering. This innovative educational program is designed primarily for mature students with an engineering or science background who want to acquire or upgrade their knowledge of mining engineering. Students can specialize in several different areas of mining, such as surface mining, underground mining, mine environment and the like. All courses constituting the program are taken online, through the Internet, without the need to attend the UMR campus in person.

May 1, 2004

By Hamid Mahmoudabadi

In the present paper, the performance of four approaches based on neural networks and geostatistical method for grade estimation are compared and analyzed their performance to find a proper method. Four methods, Multi Layer Perceptron neural network (MLP), Radial Basis Function (RBF), General Regression Neural Network (GRNN) and Ordinary Kriging (OK) are selected for ore grade estimation of variable Fe in an iron deposit located in Iran. For each method, two diagnostic statistics are calculated: Mean square error (MSE) and correlation coefficient. The results show the superiority of the General Regression Neural Network.

Oct 1, 2009

By T. R. Wildeman

"At the Big Five wetland, which was built to treat drainage from a metal mine, encouraging results have attributed to sulfate reduction. Cu and Zn were reduced to below detection limits of 0.50 mg/L, Fe was reduced by at least 90% to below 3 mg/L, and the pH was raised from below 3 to above 6. The study showed that treatment was primarily through microbial processes within the substrate and that plants were not needed to insure success.Constructed wetlands can be taken through design and proofing stages, and into operations scale in a meaningful time frame for on-line acid drainage remediation. In staged design, laboratory experiments are initiated to determine reasonable substrate materials for remediation of a particular drainage. When a reasonable substrate is found, bench scale experiments are initiated as the next stage in the design. In these studies, removal efficiency, permeability, and flow are analyzed to identify candidate substrate which will support sulfate reduction and be hydraulically reasonable. Using this design method at one site, an acid drainage of pH 2.5 was raised to a pH of above 7. Total heavy metal concentrations in this drainage exceeded 900 mg/L. In the pilot constructed wetland Cu, Zn, and Fe were reduced by 99%.Determining how well these anaerobic cells operate in the winter is done at the pilot plant site in Idaho Springs, CO. It's been found that over the last four winters, those cells that emphasize anaerobic removal within the substrate have operated continuously and effectively the year round. Winter operation is aided by the fact that the water temperature of the acid drainage is always between 14 and 16° C, and the cell can be readily insulated from the cold. Activity of the sulfate reducing bacteria may decrease during the winter. However, it appears that organic complexation of metals helps to maintain low concentrations of metals in the effluent."

Jan 1, 1992

By W. Bullen

A hypothetical economic model has been developed for diamond projects in the Canadian Arctic using data from recent exploration and mining projects. The model reflects economic conditions applicable to the diamond mining industry in the Canadian Arctic at the present time. The economic model captures pre-production expenditures, production data, and post-production costs. Before- and after-tax cash flows are derived from the data along with the relevant economic criteria. From the corporate perspective, the model provides guidance concerning the minimum size and/or in situ value per tonne requirements of targets being sought in order to ensure economic success and predict exploration and development expenditures. From the government perspective, the model provides an indication of income and mining tax revenues that can be expected from a particular diamond project. The model shows, on average, diamond projects in the Canadian Arctic are expensive to explore for and develop, but are economically robust under the present conditions. They generate great economic return for mining companies and significant amounts of tax revenue for governments. The fact that diamonds can be transported directly from the minesite by air makes them particularly attractive targets for mining companies that operate in the Canadian Arctic.

Jan 1, 2006

By Turgut Yalcin

"High-grade concentrates are achieved in flotation columns primarily due to the use of thick froth layers and wash water, both of which facilitate the drainage of hydrophilic minerals from the froth. However, conventional columns employ tall, non-agitated recovery sections for bubble-particle attachment that require high headroom and special gas spargers as well as imposing particle size limitations. These disadvantages can be all alleviated by carrying out the recovery part of the process in a mechanical or pneumatic cell which generate froth more conveniently and efficiently.This paper describes the combining of mechanical and column flotation in a single unit and discusses the influence of various parameters on the performance of the system.INTRODUCTIONIn a flotation process, some of the minerals lifted up to the froth phase may drain back to the pulp due to various reasons such as weak attachment of particles to air bubbles, bursting of air bubbles, agitation of froth, etc. Weakly floating and entrained hydrophilic minerals drain at a faster rate than the strongly floating minerals, and effective exploitation of such differential drainage can lead to cleaner concentrates.For a given mineral, if the rate of flotation from pulp to froth is k1 (min·1) and the rate of drainage from froth to pulp is k2 (min""1), then the net (overall) flotation rate of the mineral, k (min-1 ), will be [Lynch et al., 1981 ]:"

Jan 1, 1997

On the 26th of June 2000 a hydraulic fill retaining barricade failed at the Bronzewing mine in Western Australia. The barricade failure resulted in the release of 19,000 m3 of saturated hydraulic fill into the mine and caused the deaths of three miners. Following this tragic accident, the practices and procedures of mines using hydraulic fills came under intense scrutiny by regulatory bodies and mining companies. At this time, the use of hydraulic fill at Normandy?s zinc/copper Scuddles mine was suspended and waste rock became the sole backfill material in use underground. By 2003, development of the adjacent Gossan Hill mine had progressed to a point where a self supporting fill material was required to enable full recovery of high grade ore. Cemented hydraulic fill (CHF) was identified as the optimal fill type and work began on designing appropriate management procedures and controls to ensure the safe implementation. This work included definition of the fill properties, plant design and construction, reticulation system design, design of stope barricades and drainage systems and fill placement control and monitoring practices. CHF placement began on 16th July 2004 and has continued to date. Significant operational experience has since been gained in areas including plant performance, fill product quality assurance, barricade construction, stope drainage and in-situ performance.

May 1, 2007

By C. H. Lee, K. Adham, D. Lemieux, J. P. Mai, A. M. S. Hussein, S. Francey

Copper concentrates contain an increasing amount of arsenic, as the older sources of cleaner ores are depleted. At the same time, the market-determined price penalties on arsenic rich concentrates are rising due to limited arsenic treatment capacity at smelters and increasingly stringent environmental emissions limits, necessitating arsenic removal from the copper concentrates. Heating the concentrate to about 700°C can remove significant amounts of arsenic (as elemental and/or sulfide); however, it is important to achieve that removal economically and with due consideration to post-roasting arsenic (toxic waste) treatment. A comparison is made among different roasting options (kilns, multi-hearths and fluid beds) to qualify their potential application to arsenic roasting of copper concentrates. Both direct and indirect methods of heating as well as inert, air and oxygen atmosphere options are considered. Technology comparisons are made with respect to the roaster’s operation, its off-gas treatment and arsenic stabilization. INTRODUCTION In the last twenty years, the arsenic content in globally produced copper concentrates have risen steadily due to the depletion of cleaner ores and the increase in complex concentrate production. Smelter contracts typically limit arsenic to less than 0.5wt%, with a penalty paid per unit of additional arsenic and other contaminants such as antimony. While penalties can vary, they can be on the order of $70/dry metric tonne of concentrate per 1% of arsenic. Strategies used by miners to avoid the price penalties have been to blend clean and complex concentrates, and to remove arsenic from complex concentrates to achieve smelter limits. Arsenic can be removed from copper concentrates by the heat-induced decomposition of the predominant enargite mineral (Cu3AsS4) at ~700°C. Roasting at such elevated temperature is a commercially proven option, but still suffers from process, cost and environmental challenges. After roasting, the released arsenic sulfide fumes can be separated from most of the entrained concentrate dust in cyclones. Then, the As2S3 laden off-gas can be oxidized to generate a still volatile oxide (As2O3), which can be further separated from the residual concentrate dust in a hot electrostatic separator. Next, the off-gas can be cooled with air or water, to either deposit the arsenic oxide in a dry or dissolved state. Dissolved arsenic acid can be oxidized to As(V) and reacted with iron to yield a stabilized ferrihydrite precipitate or scorodite that can be safely disposed. In the case of dry arsenic trioxide, it can be mixed with reactants (e.g. iron) and fluxes before thermal stabilization in a glass matrix, thus eliminating the need for costly oxidation.

Jan 1, 2019