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By E. De Souza

Two comprehensive surveys, which focus upon backfill practices and the economics of backfilling in underground mines in Canada, have been conducted. The two surveys were designed to assess operational practices and costs associated with the wide scope of filling operations that must be met by individual mines, ranging from raw material feed to the preparation plant, through underground backfill transportation, to in-stope product placement. This paper presents the current state of backfill usage in Canada, identifies trends in minefill practices and provides mine operators with an updated database of information upon which they might compare their own practices and costs with present trends throughout the industry.

May 1, 2003

By Patrick F. R. Murphy

An ultra-low profile LHD has been developed by EJC Mining Equipment of Burlington, Ontario for narrow-reef mining applications. This paper discusses the challenges encountered during the evolution of the machine's design and subsequent field testing. Continuing research and development focuses on optimizing machine performance and controllability, reducing emissions, and enhancing maintainability. Details will be given on field trial experience and ongoing test work being carried out at the manufacturing facility.

May 1, 2004

By T. G. Joseph, K. Barron

A strain-softening constitutive relationship is introduced that describes the post-failure stress-strain curve for rock regardless of whether the rock is intact or broken. The core of this relationship is an “apparent friction” concept that relates the post-failure apparent cohesive and frictional characteristics of the rock to the strain. The relationship can be determined for a given rock via a series of triaxial strength tests establishing the peak strength and corresponding strain for a range of confining pressures, together with the determination of the base angle of friction via a simple tilt test. The postulates and assumptions used in deriving this relationship are verified from results of tri-axial tests on both intact and broken rock. This strain-softening relationship is then applied to coal. Having determined the post-failure stress-strain curve for coal specimens, this information can then be used in a finite element program, such as Phase 2, to examine the post-failure characteristics of coal pillars. Using example data, the strength of coal pillars of different width/height ratios are estimated and compared with well-known empirical pillar size-strength relationships. In addition, the post-failure stress-strain behaviour of a coal pillar is compared with an empirical relationship reported in the literature.

Jan 1, 2003

By M. Pourkani, M. Karga, Samid Banis, Andre Laplante, G. Langari-Zadeh

"The Sarcheshmeh mine processes 40,000 t/d of ore grading 0.9% Cu and 0.035% Mo. Grinding is effected in single stage ball mills using 80 mm forged steel balls, with an average consumption of 850 g/t. One of its 5 m diameter, 8 m length ball mills was emptied to measure the ball size distribution. The broken balls were first separated from the total charge, and the remaining balls classified into five size intervals of +7O mm, -70 +60 mm, -60+50 mm, -50+40 mm, -40 mm.The charge was composed of 280,534 balls (excluding broken balls) with a total mass of 305,180 kg. It was found that 4.0 % of the balls were broken; the size distribution of the charge was also markedly different from the desired one, in particular for the small-sized balls. For instance, 9% of the balls were expected to measure 2.54 to 3.18 cm in diameter, but none was found. On the other hand, the amount of top size balls was 30% higher than expected. The radial hardness of balls, which was determined by cutting the balls in halves, showed that the hardness sharply drops when the balls become smaller than 65 mm. The unexpected ball size distribution was attributed to the rapid change in the balls radial hardness. Based on the measured ball size distribution, a new ball make-up charge which uses 80% of 80 mm and 20% of 60 mm balls instead of 100% of 80 mm was proposed and implemented. The effect of the new ball make-up procedure in terms of ball consumption and grinding efficiency is being investigated."

Jan 1, 1999

By B. K. Gorain, A. W. Stradling

"The traditional metallurgical scale-up techniques for flotation plants are based on lumped flotation rate constants, dimensionless hydrodynamic numbers and the use of safety factors. Once the plant has been designed, the flotation principles that govern scale-up play almost no role in the optimization of the circuit to meet its metallurgical targets. This deficiency can potentially lead to a plant design incapable of meeting its targets, especially when circuits incorporate large flotation cells. This paper demonstrates the application of the k = PSbRf model in providing a suitable framework for flotation modelling, scale-up and circuit optimization, thus presenting an integrated approach to the problem. Studies were conducted at Teck Cominco’s Red Dog concentrator to characterize cell hydrodynamics, gas dispersion properties, froth flow behaviour, entrainment and ore floatability to obtain the key data necessary to develop a model for the flotation circuit. The findings of these studies are presented in this paper, which provide valuable insights into the modelling and optimization of the Red Dog flotation circuit. IntroductionThe traditional methodology for flotation scale-up and plant design is based on the ruleof- thumb with safety factors and dimensionless hydrodynamic group similitude considerations. With the recent trend toward larger flotation cells with volumes of 150 m3 and greater in concentrators, the problem of scale-up has become even more acute. The scale-up factors that may have been tolerated on smaller cells are not applicable to larger cells. The transportation of froth in these large cells adds further complexity to the problem. Arbiter et al. (1980) foresaw this problem and pointed out that if scale-up were attempted by rule-ofthumb and trial-and-error, the probability of failure would then increase with scale. This problem is confounded by the fact that the design and scale-up considerations used for a new flotation circuit play almost no role in optimization of the circuit to meet its metallurgical target. This results in a plant design, which may be incapable of meeting its metallurgical targets. There have been several instances where new concentrators have required significant circuit modifications and additional capacity to achieve design targets."

Jan 1, 2003

By William Pitman

Ground instability and subsequent rock falls are a major concern in underground mines. As a result, several mining companies are conducting risk assessments of the ground support installed in their mines. Risk assessment is now commonly used to provide mine operators with a systematic process to assess the strategic risks that could lead to injury to personnel or loss of production, and evaluate possible risk control measures. Risk assessments can be carried out to identify major hazards, assess the likelihood of occurrence and consequences of each hazard and estimate the level of risk. This process allows assessment of possible remediation actions to minimize the risk and likelihood of occurrence or mitigation of the consequences to a level that is reasonable and practical. This paper presents two practical mine applications of qualitative risk assessment of ground support to highlight the benefits of this process. Yearly or bi-yearly risk assessment of mine excavations can assist mine operators in prioritizing resources and areas that require support remediation. In addition, it provides a mechanism for proactive response to reduce potential risks associated with personnel safety, equipment safety and environmental impact. It also provides a baseline against which changes in ground condition can be compared and quantified.

May 1, 2004

By E. J. Fuller, J. C. Monette

"Gold and silver cyanide complexes are presently desorbed from activated carbon by means of hot caustic elutions. These solutions may or may not contain additional sodium cyanide. There are several parameters which affect the rate of stripping. Some of these include pH, ionic strength, and temperature. (1) Stripping is conducted in recirculating (ZADRA) or once through (AARL) circuits.This paper describes the development of a new chemical eluant which significantly increases the rate of stripping of both gold and silver cyanide complexes from activated carbon. Two case history applications are also described.INTRODUCTIONAdditional cyanide is used in strip circuits in order to prevent its depletion which could lead to precipitation of base metals or lead to a change in the nature of the gold cyanide complex to one which does not strip as easily. (2)The strip circuits that are able to function efficiently without the use of cyanide are mostly those where the carbon is acid-cleaned after stripping.Once the minimum cyanide requirement to maintain the complexes in their proper state are fullfilled, any futher incremental increase in the level of cyanide will have no bearing on the rate of stripping. (3)DEVELOPMENT OF THE NEW ELUANTGeneral DiscussionVarious carboxylated materials were investigated including octanoic acid, polycarboxylic acid (various molecular weights), salicylic acid, polymaleic acid, acetic acid. benzoic acid, terephthalic acid, 4, bi-phenyl carboxylic acid, phosphoric acid and stearic acid.The initial work was done on medium activiiy, non acid-cleaned carbon. The silver to gold ratio on the carbon varied greatly.Two different blanks were used. The first was an elution using a 2% caustic and 0.4% cyanide solution while the second was a 2% caustic elution."

Jan 1, 1992

By P. Massi

"Wholly owned by the Coniagas Mining Company, the oldest continuously listed mining company on the Toronto Stock Exchange, St. Andrew Goldfields is located on Hwy. 101 approximately 30 miles east of downtown Timmins, Ontario. Geographically situated on the Porcupine-Destor Fault, the mine and mill in Stock township is positioned in the heart of Northeastern Ontario's goldmining communities.In July 1988, a 600 ft. shaft, hoist, and compressors were already in place when construction started on a 500 T.P.D. capacity C.I.P. mill and crushing plant.By June 15, 1989 the mill was commissioned and attained the design capacity by August of that same year. Since then, changes to the circuit have increased throughput to 900 T.P.D. from two significantly different ore sources.Meeting the Ministry of the Environment's guidelines is an ongoing commitment utilizing the Inca S02 process in line between the tailings pond and secondary polishing pond.INTRODUCTIONExploration on what is now known as St Andrew Goldfields' Stock Township Mine was carried out ln the early l960 1 s by Hollinger Mines Ltd. The property was acquired by Quebec Sturgeon River Mines 1n 1973 and the main shaft was started ill 1979.When St Andrew Goldfields was incorporated 1n 1983, exploration and development commenced to identify the ore body. The company decided 1n 1987 to bring the Stock township property into production after an extensive feasibility study and testwork at Lakefield Research in Toronto. Leslie Engineering was contracted to design, engineer and procure al 1. equipment from the coarse ore bin to the tailings distribution system and backfill plant"

Jan 1, 1992

By Francis Chartrand

The Kwyjibo deposit is an iron oxide Cu-REE-F-Mo-Au mineralization of Mesoproterozoic age located in the Grenville Province approximately 120 km NE of Sept-Îles, Quebec (Figure 1). Occurring along the SE margin of the Canatiche granitic complex, the deposit is composed of several polymetallic showings that are related to different textural facies of magnetite. The mineralization process at Kwyjibo is complex, and is divided into Early Granitophile, Magnetite, Main Polymetallic and Specular Hematite stages (Cayer et al., submitted for publication). The Main Polymetallic stage, comprising chalcopyrite, pyrite, fluorite, molybdenite and REE-bearing minerals, is superimposed on pre-existing magnetiterich rocks. Discovered by SOQUEM Inc. in 1993, the Kwyjibo deposit shares many characteristics with other hydrothermal polymetallic iron oxide deposits around the world. Foose and McLelland (1995) suggested that this class of mineralization could occur in the Grenville Province. However, only a few polymetallic occurrences are known to be associated with low-Ti iron oxide deposits in the Grenville. Kwyjibo is one of the most significant examples of a polymetallic iron oxide deposit to have been found in the Grenville Province to-date.

May 1, 2003

By Norman R. Ball

"Many people in the mining and metallurgical industries were brought up to embrace Henry Ford's pronouncement that ""History is bunk"" but it is not generally known that the same Henry Ford created and paid for one of the finest technological museums in the Western World. I am aware that many people feel that industry history is something you ask aging executives to do or that history is a clip-andpaste compendium of pictures re-discovered every 25 years or on the occasion of issuing a press release on plant expansion. Moreover, there are many who fear historians as somewhat deranged individuals interested only in corruption, injustice and sensational labour strife. Good history lies somewhere between the controlled P.R. handout and the hysterical attack. Writing good history is a difficult creative act based on the application of experience, insight, wide knowledge and careful reasoning to available evidence. To date, the Canadian mining and metallurgical industries have achieved a great deal but left too little evidence and it is this lack of evidence which must be corrected before it is too late. It is the purpose of this article to stimulate thoughts with regard to archives, one of the more important vehicles for the preservation of such evidence."

Jan 1, 1999

By Jean-François Lapointe

Many advances have been made concerning the automation of heavy equipment teleoperators such as those found in mining. Complete automation, however, is not achieved yet and the operation of these vehicles still relies on the presence of a human operator. Fortunately, a lot of research has been done in order to improve the performance of these human-machine systems. This paper presents the characteristics of different control modes and input devices used for the control interface of those telemanipulators. We describe the benefits and drawbacks of each particular control interface on the overall system performance, based on a review of research in the field. We conclude with some recommendations and a discussion of future research needs, given the new potential advances offered by virtual environment technology and embedded computing.

May 1, 2001

By B. McLaughlin

Agnico-Eagle started sinking the Penna Shaft in 1995 with a hoisting plant capable of hoisting 1800 t/d from a 2240 m deep shaft. Exploration success led to the installation of a hoisting plant capable of extracting 4500 t of ore and 1800 t of waste per day from a depth of 2210 m. The introduction of a new rope selection factor, permitting a safety factor of 4.0, represented the breakthrough required for a cost-effective project. This paper describes the process taken to introduce the new rope selection factor and its impact on the original plant.

Jan 1, 2004

By Thomas M. Barczak

A critical safety component for all underground mining is intrinsic and standing support systems. Specifically, tabular hardrock mining exploits a wide variety of deposits that require various methodologies to safely mine the ore material. Several new support systems have been developed in recent years for hard rock applications which have been modified or applied in a variety of ground control applications. These include prestressing devices, improved cribs, and free standing supports. Various standing support and prop-type systems have been designed for hard rock applications with seismic loading conditions to accept prestressing. The prestressing, using water-filled cells, creates an active load upon installation and is considered essential to maintain support during and after the blasting of the mine faces. Heavy seismic activity is also present in these mines and the prestressing units can provide some energy absorption capability to help preserve the integrity of the support. This paper presents an overview of these innovations in support technology.

May 1, 2007

By D. A. Cormode

"INTRODUCTIONPine Point Mines, operated by Cominco Ltd., is located approximately 55 miles east of Hay River, N.W.T. by the southern shore of Great Slave Lake.The milling operation commenced in the fall of 1965 at a design capacity of 5,000 tpd. An addition which came on stream late in 1968 raised the nominal capacity to 8,000 tpd. Currently the milling rate is 11,000 tpd.The lead and zinc sulphides occur in dolomite host rock. Typical feed grades at present are 2% lead, 5.5% zinc and 5 % iron. When dolomite is present even in relatively minor amounts in the zinc concentrate, it presents serious treatment problems and causes inefficient operation of electrolytic zinc refineries.By 1972, it had become apparent that magnesium rejection in the form of dolomite was deteriorating because of increased plant throughput of less amenable ore. This paper will briefly outline the Pine Point zinc flotation circuit and will discuss the two phase program for reducing the magnesium level in the zinc concentrate. The program included changes and alterations to the flotation circuit and the construction of an acid leaching plant."

Jan 1, 1977

By Andrew Bamber

The introduction of pre-concentration technology into the underground narrow-vein mining scenario has the potential to change the entire economics of utilizing bulk-mining methods in these situations. An idealized integrated underground mining and processing system, comprising elements of bulk mining, underground preconcentration and backfill systems is presented. Through simulation and a systems engineering approach, the implications of enabling the utilization of bulk mining methods in a narrow vein deposit through the adoption of this technology is considered. Parameters considered include orebody characteristics, ore mineralogy, grade, metal value, mining rate and cost.

Oct 1, 2004

By J. Dionne

"Next to the actual mine of Les Explorations Muscocho Ltée in Montauban, Quebéc, you could find a total of 364 000 tons of flotation tailings assaying 1.4 Au g/t et 82 g Ag/t to the south of the municipality. This tailings pile is what is left of previous operations. Preliminary leaching tests on these tailings in laboratory columns gave interesting results which led to tests in larger columns in 1984 and to a pilot test on a 750 ton heap in Montauban during the summer of 1985.This paper briefly describes the columns tests performed at the Centre de Recherches minérales and the heap leaching test designed and conducted by SOQUEM in Montauban. It also presents the results obtained.RESUMEPrès de la mine actuelle de Les Explorations Muscocho Ltée à Montauban, Quebéc, se trouvent, au pied du village, 364 000 tonnes de rejets de flottation titrant environ 1,4 g Au/t et 82 g Ag/t. Ce tonnage de rejets représente ce qui reste de l'opération des anciennes mines de Montauban.Des essais préliminaires de lixiviation en colonne sur ces rejets ont donné des résultats encourageants. Ils ont été suivis en 1984 par des essais en colonnes plus volumineuses et par un essai pilote à Montauban en 1985 sur une pile de rejets de 750 tonnes.Ce te x te donne une brève de script ion des essais en colonne effectues au Centre de Recherches minérales et de l’essai pilote planifie et exécute à Montauban par SOQUEM. Il présente les résultats obtenus."

Jan 1, 1987

By Karina Forster

Rock mass characterisation is a critical component of effective stope design. Unexpected excessive stope dilution or uneconomic stope sizes can be the result of rock mass characterisation that is not representative of actual stope conditions. Comprehensive characterisation of the rock forming the stope hanging wall, back, ends and footwall is best done by a rock mechanics or ground control expert, however, this is time consuming and is often limited by available access. There are two general approaches for characterising a stope hanging wall for dilution estimating. Detailed classification and mapping data can be collected in local areas where good exposure is available. Personnel, exposure and time are usually not available to do a thorough analysis for all stopes. The stope hanging walls are then broken into lenses or domains and similar conditions are assumed for each area. The second approach consists of using readily available data such as RQD. Other rock mass properties which influence hanging wall behaviour are estimated based on back analysing a few case histories. Stope design is then based on the variability of the one parameter, such as RQD, that is consistently measured for the rock mass surrounding the ore body. This paper presents a third approach which looks, in depth, at the interpretation of geological core logging data for rock mass characterisation. Several rock mass properties that influence stope design, such as joint alteration, rock mass strength, joint spacing and orientations, have been linked to data collected by the mine geologists. This drill hole logging data was calibrated against underground mapping and used to back analyse the dilution prediction for mined-out stopes.

May 1, 2006

By Godfrey MacDonald

"The East West Caribou deposit with an underground mine, concentrator and ancillaries is located 30 km west of Bathurst, New Brunswick. The Restigouche deposit, an open pit mine is 30 km west of East West Caribou. The ore will be trucked from the Restigouche site to the East West Caribou minesite.History:The East West Caribou ore body was discovered in 1955 and the subsequent metallurgical programme focused on producing a bulk sulphide concentrate. However, in 1966 a Mr. P.G. Kihlstadt of Anaconda did float a Copper-Lead concentrate and a Zinc concentrate. He also reported that as the grind became finer the recoveries improved. In 1968 the Mines Branch in Ottawa produced a Lead concentrate of 30 percent at a 60.5 percent recovery and a 52 percent Zinc concentrate at an 83.4 percent recovery.Laboratory testwork proceed the 1988 start up of our current concentrator and underground mine facilities. The concentrator was designed to produce a Lead ""rich"" concentrate and a Zinc ""rich"" concentrate which were combined to make a marketable Bulk concentrate. In 1990 the operation was suspended because of declining base metal prices, increasing smelter treatment charges, increasing operating costs and operating difficulties.The 1997 operating plan is to produce a Lead concentrate and a Zinc concentrate.The Restigouche ore body was discovered in 1989. The subsequent testwork done at Lakefield Research recommended the production of a Copper, a Lead and a Zinc concentrate. The ore reserve could not support a stand alone concentrator so the project was put on hold.Laboratory testwork was carried out at Lakefield Research to confirm compatibility of the East West Caribou ore and the Restigouche ore in their processing."

Jan 1, 1997

By H. E. Pawson

"COMMON METHODS OF ELIMINATION: The four methods most frequently used to dispose of cyanide wastes are:1. Oxidation2. Dilution with water or with cyanide free wastes3. Acidification and dilution of resulting gas with air4. Alkaline ChlorinationMost operators practice a combination of the first two methods, allowing oxidation and dilution of cyanide waste in tailings ponds. Then, pond effluent is introduced to a stream for further dilution andoxidation which is expected to take care of the cyanide. This method oftreatment does work to a certain degree. Atmospheric oxidation in tailings ponds and in receiving streams chemically breaks down the cyanide radical to form carbon dioxide and nitrogen.2(CN) + 202 2 CO2 + N2The monetary advantage to this method is obvious to the operator but too frequently though, climatic conditions are ignored. Length of winter, and reduced flow of the receiving stream in summer, adversely restricts the result of this combined method. For those of us who rely mainly on dilution it would be wise to remember that 70 million pounds of water are required to dispose of one pound of sodium cyanide (1).The third method outlined involves addition of acid in a closed vessel and venting resulting gas to atmosphere with or without air dilution by mechanical means. Cost, safety and some form of secondary treatment to complete destruction of cyanide remaining in the solution usually make this method prohibitive.The alkaline chlorination method is the one most commonly accepted for treating cyanide bearing wastes since it is felt that cyanide is completely eliminated in this manner."

Jan 1, 1973

By M. Pekguleryuz, S. A. Argyropoulos

"The recovery of additions in light metals is an important parameter influencing consistency of the finished product, which ultimately means cost. Recovery is a function of many parameters such as assimilation speed, addition density, addition practice, liquid metal processing conditions, etc. In liquid metal processing operations, it is desirable to have the maximum possible recovery at the shortest possible time.In this paper a new index will be presented, namely, the rate of recovery (RR). This index quantifies the addition recovery and introduces the time dimension in the recovery increase. As such, the RR equips the liquid metal practitioner with an accurate picture of addition assimilation in liquid metals. Results will be presented on how this index is applied to a variety of additions in liquid aluminum. IntroductionThe data on addition assimilation kinetics offer a valuable parameter describing how fast a given addition enters the metal bath. These data may be expressed as gr m-2 s-1 of additive for a set of experimental conditions which might include bath temperature and degree of bath agitation among others. However, the liquid metal processing practitioner is more interested in the recovery of additions. In other words, the point of interest is the percentage of the used addition entering the metal bath. Ideally, the addition recovery can reach the 100% level. However, in practice, the addition recovery never or rarely reaches that level. Many factors can contribute to this. Some of these factors can be slow assimilation kinetics, addition density, addition practice, turbulence of the melt, etc. In some cases, sloppy addition practices can contribute to low recoveries, or even worse, to erratic recoveries.In this paper a new parameter is introduced. This parameter relates the maximum addition recovery observed over the time needed to achieve this maximum recovery."

Jan 1, 2002