Search Documents

By P. K. Frempong, S. D. Butt

"A case study for the application of a novel geostatistical technique for resource estimation of a narrow steeply dipping, gold-silver mineralized quartz vein deposit is presented. The technique is novel in two respects. First, mineral grades are kriged directly rather via the accumulation method more often applied to vein deposits and, second, a modified 2D ordinary kriging algorithm is used that allows the calculation of a single grade and variance for irregular shaped blocks of varying sizes. The objective of this investigation was to evaluate the suitability of this novel geostatistical technique for resource estimation of the Poura mine deposit and, if appropriate, to lay a firm foundation for future geostatistical work at the mine.The reserve estimation concentrated on the remnant ore blocks and pillars in both underground and surface workings. The data, which was collected in 2D coordinates with grade and vein thickness values, was first analyzed statistically and used to compute experimental semi-variograms. Appropriate semivariogram models were fitted and various aspects of the estimation technique were evaluated using cross-validation error analysis, with acceptable results. A block kriging computer program called Krige2D was developed to implement the modified 2D kriging algorithm and calculations. Production data were not available to evaluate the computed block resources, but they were compared with a previous resource data determined using an arithmetic approach. The geostatistical technique estimated slightly higher grades but lower overall tonnages as compared to the earlier arithmetic estimates."

Jan 1, 2004

A case study for the application of a novel geostatistical technique for resource estimation of a narrow steeply dipping, gold-silver mineralized quartz vein deposit is presented. The technique is novel in two respects. First, mineral grades are kriged directly rather via the accumulation method more often applied to vein deposits and, second, a modified 2D ordinary kriging algorithm is used that allows the calculation of a single grade and variance for irregular shaped blocks of varying sizes. The objective of this investigation was to evaluate the suitability of this novel geostatistical technique for resource estimation of the Poura mine deposit and, if appropriate, to lay a firm foundation for future geostatistical work at the mine. The reserve estimation concentrated on the remnant ore blocks and pillars in both underground and surface workings. The data, which was collected in 2D coordinates with grade and vein thickness values, was first analyzed statistically and used to compute experimental semi-variograms. Appropriate semi- variogram models were fitted and various aspects of the estimation technique were evaluated using cross-validation error analysis, with acceptable results. A block kriging computer program called Krige2D was developed to implement the modified 2D kriging algorithm and calculations. Production data were not available to evaluate the computed block resources, but they were compared with a previous resource data determined using an arithmetic approach. The geostatistical technique estimated slightly higher grades but lower overall tonnages as compared to the earlier arithmetic estimates.

Feb 1, 2004

By E. Henry, C. Mayer, H. Rott

Hangingwall deformation monitoring at Kiruna mine is a major concern for Luossavaara- Kiirunavaara AB (LKAB) since underground production is deepening toward the town of Kiruna. In the quest for more effective monitoring techniques, LKAB decided to evaluate the ability of satellite-borne synthetic aperture radar (SAR) interferometry to monitor ground subsidence by generating continuous deformation maps from satellite radar images. While adopted by other geoscience-related industries, SAR interferometry is still relatively untested in the mining industry. This paper describes the basic principles of SAR interferometry and describes how subsidence is monitored at Kiruna mine. The encouraging results are discussed as well as the benefits (continuous deformation mapping, low cost, fast, etc.) and limitations (sensitivity to snow and vegetation cover) of the technique.

Jan 1, 2004

By Chris Larsen, Carole Fortin, Dave Scott, Curtis Deredin, Ian Orford, Jean Guy Gauthier, Andy Trusiak, Lyne Côté, Mike Cooper

"Upon successful commissioning of the Stage Grinding Project (Cormier et. al., 2002), process improvements predicted through pilot plant testing were obtained in several flotation circuits, resulting in higher grades and recoveries for zinc and lead concentrates. Subsequent to establishing new baseline operating conditions from the Stage Grind Project implementation, additional opportunities were identified, and through the application of Six Sigma methodologies further gains were made. Six Sigma improvement projects were carried out in the copper-lead, zinc, lead upgrading and bulk flotation circuits. The overall performance increase resulting from incremental improvement projects has been substantial.Most remarkably, given the operational maturity up to the late 1980’s, the Brunswick concentrator team has continued to deliver a steady stream of major improvements over the past fifteen years. While senior managers and their programs have changed several times, there have been some critical constants over the long run. Common elements have been a commitment to team work, a commitment to recruit and maintain technical personnel, a structured business improvement framework supported organizationally by senior management, and focus on a few priority initiatives rather than trying to fix everything at once.The paper briefly discusses the longer term setting and then endeavors to illustrate, with examples, how improvements are presently achieved by applying the Six Sigma quality program.INTRODUCTIONThe Noranda Inc. Brunswick Mine, located near Bathurst, New Brunswick, is one of the world’s larger lead/zinc mines. Milling of ore began in 1964 from the No. 12 orebody, with an expansion in 1966 to accommodate ore from the No. 6 mine. In excess of 100 million tonnes have been milled to date. The concentrator employs differential flotation to produce lead, zinc, copper, and lead/zinc bulk concentrates. The ore is massive sulphide, with approximately 20% valuable minerals, 60% pyrite and 20% silicates."

Jan 1, 2005

By S. Cossio, J. L. Noyola, R. Espinosa

"Minera Penmont’s La Herradura is a joint venture between Industrias Peñoles, S.A. de C.V. and Newmont Mining Corporation. The mine is located in the northern Mexican state of Sonora. La Herradura consists of an open pit mine with a two-stage crushing plant and heap leaching facility. Gold is recovered from the pregnant solution using the Merrill-Crowe process; production began in July 1998. La Herradura produces about 130 000 oz gold per year and its treatment capacity is approximately 22 000 t/d. Ore mineralization occurs in a shear zone with quartz and native gold. Gold distribution is quite heterogeneous due to coarse visible gold, which in some cases ranges from 100 to 600 microns.Sampling and assaying protocols were established during exploration and prior to start-up of operations, based on nomographs developed from early heterogeneity tests. However, as the coarse gold problem became noticeable, it was necessary to develop new sampling protocols to have an acceptable relative standard deviation to control the fundamental error within ± 15%.An experimental program was developed and conducted on a selected ore block consisting of 195 blastholes; regular ore grade control procedures were applied to both sets of data to evaluate the difference in results between the protocol established prior to start-up of operations (existing) and the new protocol (leaching). In the existing protocol, sub-sampling is done through riffle splitters, and coning and quartering. Fire assaying was used to estimate the gold content on a 50 g sample. In the leaching protocol, a 12 000 g sample of bottle-rolled cyanide was leached, and the leach residue was assayed on a 50 g sample using fire assaying, and a weighted average total gold content was calculated. Currently, a new alternative method is being evaluated, where sub-sampling is done using rotary splitters, and fire assay on a 50 g sample; in this rotary sampling protocol and for comparison purposes, gold assays were also repeated using the leaching protocol.Test results were compared in three groups: < 0.3 g/t (below cut-off), 0.3 to 0.8 g/t, and >0.8 g/t. Overall, results show that the protocol established prior to start-up of operations underestimates the gold assays. Detailed technical results, alternative options, and their implications are discussed in the paper."

Jan 1, 2004

"The Key Lake mill was commissioned in 1983 to process the two Key Lake ore deposits. With the depletion of these ore bodies in 1999, the plant was converted to mill the exceptionally rich McArthur River deposit located seventy eight kilometers northeast of the Key Lake mine site. This paper describes in detail the Key Lake milling process. The mill consists of a grinding circuit, a leachkounter current decantation circuit, a solvent extraction circuit, a yellowcake precipitatiodcalciner circuit, an ammonium sulphate crystallization circuit, and a waste treatment circuit. The paper also describes process changes to handle McArthur River ore, including the ore receiving stationLOCATIONThe Key Lake mill is located 78 kilometres south of the McArthur River mine. It is located at latitude 57' N in northern Saskatchewan, within the southern boundary of the Athabasca formation. The nearest town is Pine House, located 240 kilometres to the south. Access to the mine is by an all weather road and by a private airstrip.HISTORYAND DESCRIPTION OF OPERATIONThe Key Lake mill was commissioned in October 1983 to process ore from the Gaertner and Deilmann ore bodies. The mill was designed to produce twelve million pounds of uranium oxide (U308) a year from a feed grade of 2.4% U3O8.Following grinding, the uranium is leached in a two stage leach circuit using oxygen and sulphuric acid. An eight-thickener counter current decantation (CCD) circuit is used to separate the uranium bearing solution from the solid waste residue. The uranium is further purified in a solvent extraction circuit and then precipitated using ammonia. It is then calcined at 840' Celsius, and packaged into drums for shipping to refineries. There is a crystallization circuit to recover ammonium sulphate produced in the process and a waste treatment circuit."

Jan 1, 2001

By H. Agnerian

"A survey of 925 mineral property transactions worldwide was carried out between 1987 and 2004. These mineral properties with reported Mineral Resources include 457 gold properties, 322 base metal properties, and five platinum group element (PGE) properties. Based on constant Canadian dollars, the results of the survey indicate that:• In general, unit metal values in the ground (e.g. $/oz Au or $/lb Cu) have kept pace with the changes in the respective metal prices during the period of the survey.• A general decrease in unit gold values in the ground has occurred from about $57/oz Au in 1987 to about $11/oz Au in 2002 for Canadian mineral properties, with the exception of a peak of $58/oz Au in 1996.• A similar trend for mineral properties has occurred in the United States, but with lower unit values, in the range from $1/oz Au to $10/oz Au during the last three years of the survey. • Unit metal values in the ground for gold properties in Canada were higher than for those in the rest of the world.• Unit metal values in the ground for copper properties in Canada were in the mid-range compared with other countries in the rest of the world; the United States and Australia were higher and Latin America, Asia, Africa, and Europe were lower than in Canada.• Unit metal values in the ground for zinc properties in Canada were also in the mid-range compared with other countries in the rest of the world; Latin America and Europe were higher and the United States, Australia, and Asia were lower than in Canada.• Unit metal values for nickel properties in Canada were highest compared with other countries in the rest of the world, followed by the United States, Australia, Africa, and Asia.• Gold values in the ground in Canada have ranged from 6% to 14% (with an average of approximately 10%) of the gold price.• The unit values of base metals in the ground have been in the order of 2% of the quoted metal price, during the period of the survey."

Jan 1, 2006

By Tom C. Palangio

The mining industry has come to rely on imaging devices to monitor and record all kinds of activities. We use them to measure, improve the process, ensure security, increase production and automate machinery. Cameras can be installed in locations most people wouldn?t want to live and be our ?eyes? to record the events at that location. This paper discusses digital imaging, the equipment, methods, techniques and information that can be learned from digital devices used in mining operations. The guy who said ?a picture is worth a thousand words? probably understated it. Images are the perfect way to document conditions before, after and even during mining activities. Today?s wide choice of cameras, camcorders and other digital devices allow us to record most activities and events. The technology is advancing at a high rate. Digital cameras and camcorders are marvels of technology with high resolution optics, compact size, automatic time and date stamp, great battery life, shutters to freeze motion and computers to auto focus and eliminate jitter (image stabilization). Add to this the economical price and ease of use and everyone in this industry should be aware of the possibilities.

Oct 1, 2010

By A. R. Barnes, N. O. Lotter

Recoverable economic copper sulphide minerals such as chalcopyrite, bornite, chalcocite and covellite often occur together in varying proportions in the major copper-bearing ores, and have individual flotation requirements and characteristics. Pyrite also occurs in these ores to varying extents as the sulphide gangue, and is problematic because of its natural tendency to float quickly and easily. In a bulk sulphide float, selectivity against pyrite is desirable, particularly if it does not host other paymetals such as gold or silver. At the same time it is a requirement to float all of the copper sulphides despite their electrochemical differences. The electrochemistry and semiconductor properties of these minerals are reviewed, together with implications for flotation with and without collector addition. Mixed collector systems for the improved flotation of these sulphides are proposed as a solution. The use of xanthate and dithiophosphate in the collector suite allows the co-existence of dixanthogen and free dithiophosphate radical because the latter has a higher redox potential requirement than xanthate to oxidise to the dithiolate. Because some of these minerals require dixanthogen, and others, free thiolate, to generate surface hydrophobicity, a bulk flotation of all the species becomes possible in the overlapping area of Eh and pH between the two dithiolate equilibrium lines on the Pourbaix diagram. The arsenic-signature copper minerals are added to the study, since many copper operations encounter arsenic as a penalty element in the saleable concentrate. It is shown that the addition of arsenic to the copper and iron sulphides alters the semiconductor and electrochemistry properties, and in turn, its flotation characteristics. The degree of mineral association and liberation between these minerals can be a complicating factor due to textural associations, and should also be considered in the process as a next step.

Jan 1, 2016

By S. Banisi, M. Pourkani, A. Kargar, D. Hamidi, M. Sarvi

"Flotation circuit of the Sarcheshmeh copper mine, which processes 40,000 t/d of ore at 0.9% Cu, consists of rougher, cleaning and recleaning stages. To study the stage recovery of copper, one of the eight identical flotation banks was selected for sampling. It was observed that 17% of copper goes to the tailings and the lowest copper recovery was for particles coarser than 105 microns. The third five-cell unit of this bank not only did not recover appreciable amount of copper but also decreased the concentrate grade. The grade of the last five cells was lower than the feed grade suggesting that mixing of the concentrates of the three units was not beneficial. To alleviate the problem, the concentrate of the first four-cell unit was directed to the cleaners feed box (i.e., by- passing the regrind mill). This increased the overall copper recovery by 1.7 %. Also rather than the current practice of adding the reagents only at the feed tank (i.e., 100-0-0%), various combinations of stage addition were examined. It was found that 75-25-0% addition regime could increase the total copper recovery by 1.3%. At the recleaning stage in order to increase the final copper concentrate grade, a froth washing system was used. To ease the maintenance of the system, a shower type washing system was installed. The optimum wash water superficial flowrate was found to be 0.06 cm/s. Continuos sampling of the recleaners with and without wash water over a period of two months showed an increase of 1.5% in the final copper concentrate grade at constant overall copper recovery.INTRODUCTIONThe Sarcheshmeh copper ore body that may rank as third or fourth largest in the world contains 1 billion tonnes averaging 0.9% copper and 0.035% molybdenum. It is located in southeastern Iran and has been processing 40 000 t/d of ore since 1982. The mine produces 100 000 tonnes of copper and 2200 tonnes of molybednite concentrate per year.After three stages of crushing, the ore with a F80 (80% passing size) of 12.7 mm is fed to eight parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of rougher flotation circuit is ground; rougher tails constitute the main component of the plant final tailings. A concentrate with an average grade of 31% Cu (chalcocite and chalcopyrite) is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83% and 87%, depending on operating conditions and ore type."

Jan 1, 2001

By Jeff R. Reinson

The Diavik Diamond Mine is producing diamonds from two orebodies, the A154 South and A154 North kimberlite pipes, discovered beneath the waters of Lac de Gras. Innovative dike technology with appropriate water collection and treatment systems allowed Diavik to start open pit mining of the ore bodies in 2002. To allow planning, engineering, and implementation of sufficient water handling infrastructure, reasonably accurate estimates of groundwater inflow quality and quantity over the life of mine are required. Prior to mining, a numerical model using the continuum approach was developed for regional and site wide water management. After approximately two years of mining, groundwater inflows to the pit were found to exceed those predicted by the base case hydrogeologic model. The most likely explanation was flow through Dewey?s Fault zone. As this fault was not explicitly accounted for in the model, a detailed hydrogeologic investigation was conducted to evaluate its geometry and hydraulic properties. A multi-disciplined approach was used and involved core logging, fluid logging, geophysical logging, packer testing, a pumping test, and water quality sample collection. The results of the field investigation were used to update the conceptual and numerical models and to develop a Discrete Fracture Network (DFN) model to address small scale problems such as flow into future underground drifts. The updated numerical hydrogeologic models predict that up to 70 per cent of the groundwater inflow to the A154 N/S open pit is through the relatively narrow fractured rock zone associated with Dewey?s Fault. Consistent with the continuum model, the DFN model predicts that inflow to a drift excavated in Dewey?s Fault will be approximately 10 times the inflow to a drift excavated in competent rock. The DFN model provides a better understanding of the distribution of these inflows on a local scale over the length of the drift.

May 1, 2006

By A. V. Corlett

THE use of the diamond-drill hole as a means of breaking clown ore and waste is not new. Years ago, the method was adopted in mines in northern Ontario and in the western United States to reach out to bad pillars, or to remnants of ore left on walls, the winning of which by any other method might be dangerous or too expensive. The practice is still in vogue and has generally proved a success. Of the known mineral deposits in Canada and elsewhere, many are of very low grade. Some are marginal and not even worthy of being classed as &apos;ore&apos; until some means is found and proved of exploiting them economically. Other conditions being favourable, if a mining method could be found that would give a large &apos;bulk break&apos; with a sufficiently low cost of development and a low cast of mining per ton, mineral deposits of this latter type might be brought into the classification of ore. We know that any block-craving system whereby the ore may be undercut and allowed to crush and cave will give a very high breakage and low costs. In some cases, however, the physical characteristics of the ore do not lend themselves to this method. In such ore-bodies, the alternative system of &apos;bulk-breaking&apos; might be aided by long diamond-drill blast holes. We could obtain this &apos;bulk breaking&apos; by the use of long diamond-drill holes in conjunction with a sub-level system using extremely thick benches, or possibly by the adaption of modified open-pit methods to underground mining. ln fact, if conditions are favourable, the idea would be simply ta open up a slot or free face and break into it on a larger scale with longer holes. The use of deeper benches-that is, using each level as a sub-level and thereby eliminating much development footage, slashing, setting up, and mucking time-would naturally be expected to lead to cheaper mining.

Jan 1, 1940

By D. B. Felsher

"Lakeshore Gold’s Bell Creek Mill is located in Timmins, Ontario. The mill has been in operation since 2009 and has undergone several small expansions since then. In 2012 - 2013 the first major mill expansion was commissioned raising the mill capacity from 2,000tpd to 3,150tpd. This was done through the installation of a new ROM ore truck dumping facility, jaw crusher, ore storage dome, SAG mill, leach tank, CIL tank, and electrowinning cell. This paper will cover the equipment installed to raise the tonnage as well as the commissioning of the new equipment including problems encountered and how these problems were solved.IntroductionLake Shore Gold acquired the Bell Creek Mill, along with the Bell Creek Mine, from the Porcupine Gold Mines Joint Venture, a joint venture between Goldcorp Canada Ltd. and Kinross Gold Corporation, in December 2007.The Bell Creek mill, which processed ore from the Bell Creek Mine has consistently achieved recoveries of over 95%, is a conventional gold mill circuit, involving crushing and two stage ball mill grinding, gravity and leaching, followed by Carbon-In-Pulp (CIP) process for gold recovery. The facility had been placed on care and maintenance in 2002. At the beginning of 2009, the mill had been refurbished with an initial capacity of 800 tonnes per day and was subsequently refurbished to a capacity of 1,500 tonnes per day near the end of the third quarter of that year. The Mill was then expanded in 2010 to a capacity of 2,000 tonnes per day.In late 2011, the Company commenced a 50% expansion of its milling facility, from a processing capacity of 2,000 tonnes per day to 3,150 tonnes per day. Phase 1 of the expansion, which mainly involved enhancements to the solution circuit (“back end”) and increased the mill’s capacity to 2,500 tonnes per day, was completed at the end of 2012. Commissioning of Phase 2 of the expansion, to a capacity of 3,150 tonnes per day, commenced at the end of July 2013."

Jan 1, 2015

By Mike Turner

The Otter-Juan mine was closed by WMC Resources? Kambalda Nickel Operations (WMC-KNO) in 1999 due to the combined effects of low Nickel prices and high mining costs. GBF Mining Pty Ltd purchased Otter-Juan Mine from WMC in early 2001 as part of WMC?s disinvestment from the nickel mines in the Kambalda area. Goldfields Mine Management (GMM), (a wholly owned subsidiary of GBF Pty Ltd) re-opened Otter-Juan in May 2001. The Otter-Juan mine is one of the oldest mines in the Kambalda region and is now mining at depths below 1160mbs. Stresses are very high, rock mass deformation considerable, and the mine has a history of serious seismic activity. Orebody widths in the current mining area range from 2 to 2.5m, with dips from 25° to 40°. On taking over the operation, GMM converted the mining method from the previous mechanised drift and fill method to a slot-rise and strip method using hand held airlegs. During the latter years of production under the control of WMC, production averaged around 12000 tonnes per month, with grades of 2.7% and orebody extraction of only 55 to 60%. In the last three years of operation under the control of GMM the mine has produced an average of 12000 tonnes per month of ore at an average grade of 4.2% and with an orebody extraction of 70 to 75%. The mine is now highly profitable with at least another 3 years of reserves remaining at depth. The main benefits of the current method include reduced dilution and increased extraction. In the relatively poor ground conditions experienced this has only been possible using airleg mining, with excavations designed to suit conditions, support to suit rockmass deformations, and with improved extraction sequencing. Other benefits have included improved operating conditions, reduced falls of ground incidents and a dramatic reduction in seismic activity. These achievements have only been possible by stepping back in time and using traditional airleg stoping methods to extract the maximum amount of the orebody with minimum dilution. This paper includes details of the mine history, geology, mining method, seismicity, and support.

May 1, 2004

By Paul Gauthier

Aujourd?hui, avec un faible prix de l?or sur le marché, les opérateurs miniers doivent revoir leur façon de faire. La méthode de minage «Eureka» permet de réduire les coûts de production et d?accélérer la mise en production des chantiers. Les principales étapes du cycle de minage et les problématiques associées à l?utilisation de cette méthode à la Mine Bousquet y sont décrites.

Oct 1, 2001

By E. V. Manlapig, K. R. Rahal, J-P Franzidis

"The Floatability Characterisation Test Rig or FCTR is a highly instrumented, mobile flotation pilot plant built by Baker Process (now EIMCO) for the AMIRA P9 Project for use in flotation plant modelling and simulation studies. With the Mintek Floatstar system controlling the level in every cell, and the facility for full stream sampling on all streams, the FCTR is a unique tool for producing data of sufficient quality for flotation plant modelling. The plant was commissioned at Lonmin's Karee and Eastern Platinum concentrators in 1999/2000, and completed a number of test work programs to investigate the effect of power input and regrinding on metallurgical performance. Lonmin's experience with the FCTR has caused it to build two of its own to undertake further projects on its various mines. This paper describes the design of the FCTR and the features that make it a unique research tool, and the benefits derived by Lonmin through their experience with the unit.INTRODUCTIONAs part of the Australian Mineral Industries Research Association (AMIRA) P9 project, a new model for flotation has been developed, in which the overall recovery, R, of the jth floatability class of the ith size class of a particular mineral, is given as (Savassi, 1998):"

Jan 1, 2002

By M. Vallée

"RÉSUMÉDurant les 10 dernières années, les améliorations dans le contrôle des limites des sautages et les arpentages après l’exploitation ont fourni des informations préalables et post facto sur les limites minières réalisées par rapport à la planification, plus particulièrement pour les grands chantiers exploités par les méthodes chambres vides et sous-niveaux. Cependant, peu de recherches ont été faites pour améliorer la délimitation du minerai, l’estimation des ressources et des réserves, les études de faisabilité et la récupération minière dans d’autres types de gisements. Il faudra d’abord expliciter les impacts des lacunes, en quantité et en qualité, de l’information géologique et de l’échantillonnage sur le rendement de l’extraction. Il faudra également remplacer la notion de « dilution générale » par celles de « dilution » et de « perte de minerai ». Les impacts des lacunes observées sur l’exploitation des gisements d’or filoniens sont démontrés par une simulation basée sur des résultats d’études antérieures à la mine Thompson d’Inco, des impacts financiers de densités d’information moindre et par des analyses des valeurs actualisées nettes (VAN) de deux cas types.Deux gisements similaires, de 3 m et 6 m de largeur et exploités par chambres vides sont étudiés, en prenant pour cas de base la délimitation sur maille de 7,5 m (25 pi.) et en étudiant les impacts des limites moins précises supportées par une maille de 15 m (50 pi.), en assumant que ces imprécisions moyennes de +/- 0,62 m se transforment soit en surestimation (dilution), soit en sous-estimation (pertes de minerai).Les impacts d’information moins précise ont été beaucoup plus forts sur le gisement plus étroit : autant la dilution que les pertes de métal ont diminué les VAN du gisement de 3 m de 39,27 % et 45,41 % respectivement, tandis que les VAN du gisement de 6 m ont été diminuées que de 12,84 % et 11,55 %.Ces impacts des valeurs actualisées nettes diminuées justifient et exigent des efforts plus systématiques pour s’assurer que les travaux de mise en valeur et les études de faisabilité soient réalisés de façon requise pour optimiser l’exploitation et offrir aux investisseurs des rendements compétitifs par rapport à ceux des autres secteurs industriels. ABSTRACTDuring the past 10 years, research work to improve the survey of stope limits after mining supplied useful information, particularly for large open and sub-level stopes, regarding waste dilution caused by differences between final production limits and planned stope limits.However, the efficiency of this information to improve the overall efficiency of the mining process is limited, given that the impacts of inaccurate ore limits leading to dilution (sensu stricto) and ore losses have not been considered adequately. Geological/sampling information should be optimized in terms of quantity and/or quality for the deposit characteristics and the constraints of the mining method used. As industry practice regarding the definition and recording of dilution is flexible and as ore losses frequently are poorly documented, the use of these terms is reviewed.Simulations were carried out on two typical narrow gold deposits, based on an earlier study at Inco’s Thompson mine regarding the impacts of wide to narrow delineation grids using recorded variations of vein geometry. Vertical to sub-vertical deposits of 3 m and 6 m widths mined by long-hole open stopes were considered. Deposits delineated by 7.5 m (25 ft) grids were taken as a base case and the impacts of less accurate limits in conjunction with long hole mining based only on a 15 m (50 ft) delineation grid. The impacts of overestimation (dilution) and underestimation (metal losses) were considered separately, using the average error of +/-0.62 m recorded in the Inco study.The impact of less accurate information was much stronger on the narrower deposit: both dilution and ore losses diminished the net present value of the 3 m wide deposit"

Jan 1, 2005

By J. M. Houston

"In October 1982, Rio Algom Limited, purchased a tin prospect near East Kemptville, Nova Scotia, from Shell Canada Resources. This purchase led to the development of the first substantial tin mine in North America. This paper deals with the drillingand blasting operations at Rio Algom, in particular the use of emulsion explosives technology.This presentation describes what emulsion explosives and Heavy AN/FO are, and the support system used to deliver these products to the blast site.HistoryThe East Kemptville mine site was discovered by Shell Canada Resources (SCR) in July 1978. SCR started diamond drilling on the tin prospect in March 1979. After 8220 m of BQ drilling, it was estimated a reserve of 56 000 000 tonnes of 0.165% Sn was present.The property, which is located 54 km northeast of Yarmouth, Nova Scotia (Fig. 1), showed, after a feasibility study that, given an ore treatment rate of 9000 tonnes per day, a mine life of 17 years could be expected.GeologyThe tin deposit of the East Kemptville Mine, occurs in granitic rocks at their contact with metasedimentary rocks. The mineralization is in the form of cassiterite with associated sulphides of copper and zinc. The tin occurs in a stockwork of tin bearing Greisen alteration zones.East Kemptville MineThe mine planned an open pit operation with the first five years being mined at a rate of 14 000 tonnes per day, of which 5000 tonnes will be waste and low grade ore. Eventually, the total mining rate will be 21 000 tonnes per day, including 12 000 tonnes of waste and low grade ore. Figure 2 shows the general layout of the mine site.East Kemptville Chooses Emulsion ExplosivesIn December 1984, East Kemptville Mine awarded the Explosives Division of C-I-L Inc. a contract to supply explosives to the borehole. The system chosen gives East Kemptville the on-the-bench choice of regular AN/Fa, or a high density, Heavy AN /Fa product called SuperAN· or a water resistant emulsion explosive called MAGNAFRAC· for holes that cannot be dewatered.Emulsion ExplosivesAn emulsion is an intimate mixture of two liquids that do not dissolve in each other. Emulsions have contributed to our daily lives in such products as insecticides, salad dressings and cosmetics."

Jan 1, 1987

By R. F. Hochstein, A. J. Neale, B. C. Flintoff

"Process control technology holds very significant potential benefits for many mineral and coal processing plants. The impact can be expected in terms of performance improvements (e.g. increased throughput, better quality control, higher availability), through reduced operating costs (e.g. lower reagent consumptions, manpower reductions) and in the form of intangibles (e.g. reduced vulnerability to manpower turnover, increased productivity in the user group). However, the installation of a process control system will represent considerable capital expenditure and will generate an operating cost. No management team would approve the funding for such a project without some reasonable measure(s) of the profitability of a proposed system. This paper describes a methodology for addressing the justification problem. Of particular interest is the approach to quantifying performance improvements, frequently found to be the predominant benefit of process control.INTROPUCTIONProcess control has become widely recognized as an essential component of any processing operation. It has direct impact on productivity, process efficiency and product marketability. The reported pay backs associated with successful installations underscore this observation. As a consequence, many operators who have little in the way of process automation are looking to initiate control projects, while those who currently enjoy some level of automation are looking to upgrade existing control systems (= process + instrumentation + hardware + software + operators + support). While the emphasis here would appear to be on retrofit process control projects, new plants, or those currently in the design stage, will usually find themselves in one of these categories shortly after startup 1."

Jan 1, 1990

By Mohamed Ourriban, Stéphanie Somot, Éric Proulx, Jean-Sébastien Marois, Marjolaine Drouin, Paul Blatter, Ahmed Bouajila, Claude Gagnon

"The mining industry has to face increasing restrictions concerning the use of fresh water in Québec and throughout the world. Increasing recirculated water usage could lead to a better control of the quality and the quantity of water supplied to the mineral processing units. Furthermore, it could reduce the volume of waste water to be managed and treated before its release to the environment. However, process water may have a negative impact on mineral processing operations, thereby limiting the incentive for sustainable development initiatives. As a first step into selecting and testing remediation strategies, one of which being blending local waters, a mapping of process water quality (inorganic and organic fractions) allowed to locate zones in an industrial case study where fresh water is introduced into the circuit and to detect dissolution and precipitation zones. A complementary hydro-geochemical modeling study on water blends was also performed showing encouraging results towards the objectives of the project.INTRODUCTION The industry has to face increasing restrictions on the use of fresh water in Québec (Quebec Government, 2009) and throughout the world (Brown, 2003; UNESCO, 2006). The cost for the use of fresh water and for water management is highly variable and case-dependent (from 0.2 to up to 7 million dollars/year according to available private estimations, excluding the capital expenditure). It may be, or may become, significant for any plant; thus, risk assessment is of concern here. Using recirculated water could lead to a better control of the quality and the quantity of water supplied to the mineral processing units. Furthermore, it could reduce the volume of waste water to be managed and treated before its release to the environment. However, process water may have some positive and negative impacts on mineral processing operations (Janssen, 2003; Johnson, 2003; Rao and Finch, 1989). Negative impacts could limit the incentive for sustainable development initiatives. Water management strategies at mineral processing sites are supported both by a strategic plan and a water management plan. The strategic plan includes legal aspects, investment costs, potential benefits and various risk levels (Corder and Moran, 2006; Côte et al, 2006; Evans et al., 2006; MacIntosh and Merritt, 2003; McLellan et al., 2008; Merritt and MacIntosh, 2006; Robinson and Keddie, 2006; van Beers et al., 2007a; van Beers et al., 2007b; van Berkel, 2007). In practice, the strategic plan evaluates technical possibilities and costs of various scenarios, such as for example: bringing fresh water to the plant or recirculating waters, or using blends of both. Two important factors to consider in order to enhance the future production efficiency and to minimise risks are: the durability and the stability of the resource, both from quality and quantity points of view."

Jan 1, 2012