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By Joseph J. Mgumbwa

Pillars are any structures left between two or more underground openings. The stability of pillars is normally considered to depend on shape (defined by width to height ratio), rock mass strength, extraction ratio, in situ stresses and the gross structural features (i.e. joints, contacts, faults etc). Recent experience has shown that pillar stability is also controlled by the relative orientation of the orebody with respect to the in situ principal stress. In inclined orebodies, pillars are loaded in compression and shear, thus creating unsymmetrical stress distribution. Failure mechanisms of these pillars are little known. This paper investigates the failure mechanisms of pillars under shear loading. Two dimensional elastic numerical modeling was used to examine the behavior of pillars in different orebody inclinations relative to the major farfield stress orientation. First, the paper reviews pillar design approaches including the empirical pillar design chart and uses numerical modeling to investigate the effect of the major farfield stress orientation on pillar stability. It is concluded that the empirical pillar design chart be used with caution, and that for orebodies under shear loading mine planning and design should take this into consideration since such geometrical orebody/stress relations put mine structures at higher risk of instability.

Oct 1, 2010

By H. R. Bullis

"A 2003 paper discussed the idea that as gold mining companies grow larger it becomes more difficult for them to discover or acquire new gold deposits large enough to replace extracted reserves on an annual basis. Deposits of the size required to replace annual production of very large gold producers (VLGPs) are statistically few. An update in 2005 reviewed VLGPs with respect to production, reserves, and reserve life and observed recent changes in the gold mining industry. Now, more than 12 years later, this paper explores how those VLGPs have fared and how the gold mining industry has changed. KEYWORDS Gold, Gold price, Production, Reserves, Very large gold producer (VLGP) RÉSUMÉ Un article de 2003 examinait le concept selon lequel, à mesure que les sociétés d’exploitation aurifère se développent, il leur devient de plus en plus difficile de découvrir ou d’acquérir de nouveaux gisements aurifères suffisamment grands pour remplacer les réserves qu’elles extraient annuellement. Statistiquement, il existe très peu de gisements de la taille requise pour remplacer la production annuelle des très gros producteurs d’or. Une mise à jour de l’article en 2005 se penchait sur la production, les réserves et la durée de vie des réserves des très gros producteurs d’or ; elle portait sur les récents chan¬gements au sein de l’industrie de l’exploitation aurifère. Aujourd’hui, plus de 12 années plus tard, cet article explore l’évolution et les résultats de ces très gros producteurs d’or, ainsi que les changements qu’a connu l’industrie de l’exploitation aurifère. MOTS CLÉS or, prix de l’or, production, réserves, très gros producteurs d’orINTRODUCTION In 2005, the world’s top five gold producing companies were Barrick (after purchasing Placer Dome; 9.1 million ounces; Barrick Gold Corporation, 2005; Placer Dome, 2005), Newmont (8.2 million ounces; Newmont Mining Corporation, 2005), AngloGold Ashanti (6.2 million ounces; AngloGold Ashanti Limited, 2005), Gold Fields (4.2 million ounces; Gold Fields Limited, 2005), and Harmony Gold (3.0 million ounces; Harmony Gold Mining Company Limited, 2005). During the years leading up to that point, there had been a flurry of activity in the gold sector directed at consolidating production and reserves to create very large gold producers (VLGPs). As defined in Bullis (2005), a VLGP is a mining company whose primary busi¬ness is gold production and that has annual gold production more than 5 million ounces. A stated corporate goal of several of the VLGPs at the time was to become the largest gold producing company in the world."

Jan 1, 2019

By David E. Pearson

"There are many tests that can be used to characterize and classify coals, and describe their quality. One of the easier tests, and most versatile in its application, is petrography. Rank and coal type are precisely defined by petrographic evaluation, and in this paper they are compared with other definitions of rank and coal type. Brief descriptions are given of the various rank tests, and of rheological and coking tests used to describe coking-coal quality in Canada. Similarly, tests used to evaluate thermal coal quality are discussed. Finally, the use of these tests as parameters to classify Canadian coals is examined with reference to four coal classifications.Introduction Among all coals there are two independent variables that are required to evaluate coal quality. These are, coal rank and coal type (or composition), and although coal petrography can be used to precisely define them, other methods are often used to provide values that are, in fact, functions of rank and coal type. For example, volatile matter yield is used as a proxy for rank, and free swelling index is used as a proxy for coal type in the International Classification of Hard Coals by type. Currently, there are a number of methods in use in Canada to define rank. The approach taken here, (obviously the author's bias), is that Romax, the mean maximum reflectance of vitrinite, is both a quick and an accurate measure of rank. It is logical to compare the other techniques employed in the measurement of rank to this favoured method. Coal type is, in theory at least, simply a quantitative measure of the various varieties of coal. In practice, the microscopic entities are assigned to a Reactive or Inert category, and at that point objectivity fades from petrography. Despite this somewhat subjective approach, petrographic composition of coals remains an independent variable of such.importance that its place in the evaluation of coal is essential. In this paper, a variety of me.thods that are used to assess rank and coal type are reviewed and compared; the influence of these variables over other parameters that are used to determine the quality of metallurgical and thermal coals is examined, and their usefulness in coal classifications is discussed."

Jan 1, 1985

By J. C. Escobedo, J. M. Almanza, A. Flores, Cortés. D. A.

Secondary aluminum dross, which has been exposed for several years to atmospheric conditions, was characterized using chemical analysis, X-ray diffraction, optical and scanning electronic microscopy, and energy-dispersive spectroscopy. Based on the characterization, a method to purify the solids is presented. This method starts with the washing of the dross in boiling water to eliminate the soluble compounds like Na and K, followed by the physical separation of the metallic aluminum by milling and screening. Finally, an acid lixiviation process is applied using three different acid solutions, HCl, HNO3, and H2SO4. The best results were obtained using HCl. The main insoluble components of the dross after lixiviation were Al, SiO2, and Al2O3. MgAl2O4, AlN, Si, and Na2Al22O3·2H20 species were also found. Based on the alumina content, the solid product could be suitable as a raw material in the manufacture of castable refractories or metal matrix composites.

Jan 1, 2005

By Rob Morrison, Peter Cameron

"Many products and technologies have been developed from outcomes of research projects carried out by the Julius Kruttschnitt Mineral Research Centre (JKMRC). When these products and technologies are applied to the design and optimisation of blasting, comminution and separation, improvements in profitability can be realised.In 1970 Alban Lynch promoted two practical ideas to improve plant operations:• Ore and equipment properties can be separated allowing experience with one type of ore or equipment to be transferred to other types of ore and equipment. If this experience was encapsulated in a mathematical model, then these properties could be combined and used to simulate (ie. to design and to optimise) complete mineral processing circuits.• Many problems associated with scale-up from laboratory to industrial equipment could be eliminated by experimental work on full-scale equipment in the plant.Many JKMRC research projects have focussed on the development of models of unit operations in mineral processing and their application to process simulation.Models of comminution and classification developed at the JKMRC have been used since the early 1980's. Ore characterisation techniques developed to support these models are licensed world wide through JKTech. The application of comminution modelling techniques to blasting, which optimises the process feed size distribution is used in Mine-to-Mill® projects. The advances in the understanding and modelling of froth flotation over recent years have provided the opportunity for integrated ore characterisation and process optimisation from resource to product.JKMRC project outcomes often provide proven opportunities for the mineral industry to improve performance and productivity. JKTech was established in 1986 to commercialise and market the most promising outcomes and thus ensure their continuing availability to the mineral industry.They include:• Commercial software (JKSimMet, JKSimBlast and JKMetAccount).• Monographs (Comminution and Blasting).• Technologies (Mine-to-Mill® and flotation optimisation).• Hardware products (JKFrothCam, JKJetLift® and the Mineral Liberation Analyser).This paper summarises these technologies and their benefits."

Jan 1, 2003

By R. Jackman, D. Koningen, G. Wilcox

"Gowest Gold Ltd. is in the process of developing the Frankfield East gold deposit located along the largely undeveloped Pipestone fault northeast of Timmins, Ontario. In addition to resource development and environmental permitting work the company has had an ongoing program of metallurgical and engineering studies aimed at evaluating options for processing the Frankfield East ore. Mineralization at the Frankfield project contains arsenopyrite and pyrite as the major sulphide minerals. Bulk flotation of all sulphides results in extremely high gold recovery. However, the gold is almost entirely associated with the arsenopyrite fraction of the ore and is refractory in nature requiring preoxidation treatment. Selective flotation was investigated to recover a gold-arsenopyrite concentrate thereby reducing the mass which would be subjected to further gold processing. As part of this work, a pyrite concentrate low in arsenic content was also recovered that would be suitable as a source of sulphur for a smelter (or other pyrometallurgical facility). Production of separate arsenopyrite and pyrite concentrates is frequently examined but rarely adopted due to the distribution of gold between the two products or the inability to achieve adequate separation. Several flowsheet options were investigated in the study. The best results were achieved following a sequential flotation flowsheet. The flowsheet configuration was key to meeting the targets for the two products. In this paper, the various options investigated are reviewed and the results of the final flowsheet are presented. The overall plan by Gowest was to evaluate all process options to bring the Frankfield East deposit into production. The ability to effectively separate the arsenopyrite and pyrite components in the ore provides Gowest with added opportunities to pursue both short and long term production scenarios."

Jan 1, 2012

By Andrea M. Ouchi

A major focus of ground control research presently being conducted by the Geomechanics Group in the University of British Columbia Mining Engineering Department is to incorporate weak rock masses (RMR76 less than 50) into existing empirical design relationships. An emphasis is being placed upon the updated span design curve/critical span graph originally developed at UBC. The original database has been augmented by mapping of one mine in Montana and of several mines within the Carlin and Getchell Trends of Nevada. In addition, a neural network analysis has been performed and modifications of the span design curves at the lower RMR76 range have been made. The common factor in all the above mentioned mines is a weak back and/or walls. In most cases, the ore zone is the weakest rock unit and must be dealt with to safely extract the mineral bearing rock. This work attempts to provide rock mechanic tools that will enable a mine operator to make economic decisions that will also ensure a safe working environment.

May 1, 2004

By J. T. Chen, S. Winkler, M. J. Worswick, O. Orlov

This paper presents a review of research conducted at the University of Waterloo concerned with the application of aluminum alloy sheet and tube in automotive structures and body panels. The research has focussed on formability prediction and results using a range of damage-based formulations. These models include continuum-level damage approaches and, more recently, a so-called percolation damage approach intended to capture microstructural heterogeneity effects on damage development. This paper presents applications of these damage approaches to simulate formability of experiments considering stretch flange panels. The experiments consider drawbead effects associated with the bend/unbend deformation as the material flows through the drawbead. Recent work on simulating in situ X-ray tomography experiments that capture damage evolution during tensile deformation will also be reviewed.

Jan 1, 2006

By J. E. Dutrizac, J. B. O'Reilly

"Zinc is a relatively abundant element which occurs as both high-grade sulphide ores {sphalerite ZnS) and oxide ores which the ancients collectively termed ""calamine"" (smithsonite ZnC03 and/ or hemimorphite Zn4Si201(0H)2 - H20 and/ or hydrozincite Zn5(C03) 2(0H)6). Although oxidic zinc is readily reduced by carbon, the process becomes thermodynamically viable only at - 1000°C, and at this temperature the zinc is released as a vapour (B.P. 907°C). Casual cooling of the vapour results in its re-oxidation by excess combustion air or even the carbon monoxide product of the reduction reaction. ZnO + C = C09,, + Zn vapom Only fairly rapid quenching of the zinc vapour and the isolation of the reduction products from the combustion gases will yield metallic zinc; these conditions were not produced deliberately until the advent of distillation technology. Until quite recently, the principal use of zinc was in brass manufacture; galvanized iron was not discovered until the 19th century and zinc casting alloys are a relatively recent invention. As will be seen, brass could be produced without recourse to a pure zinc intermediary, and the discovery of brass preceded that of zinc by many centuries. The purpose of this paper is to outline the discovery and early development of both zinc and brass with special reference to the European scene."

Jan 1, 1999

"7.1. Glossary of the Platinum-Group Minerals7.1.1. IntroductionA glossary of the platinum-group minerals (POM) serves as a useful introduction to a chapter on their mineralogy. Earlier glossaries (Cabri, 1972, 1976) were prepared for different purposes and composed in a different manner. As for most mineralogical textbooks, from Dana's System of Mineralogy, 3rd Edition (1850) , to Fleischer's Glossary of Mineral Species (1980), this glossary is primarily a chemical classification, arranged here in two tables. Table 7.1 groups the mineralforming compounds of the six platinum-group elements (POE) according to their essential anion(s) or according to the nonPOE element in intermetallic compounds. Table 7.2 is a similar listing for the native metals and alloys of the POE. Both tables include unidentified POM considered to be bona fide new minerals requiring additional study for complete characterization and recognition (see Chapter 9*). Some information on the chemistry of the POM is given in the sections on X-ray data (7 .2) and ore microscopy (7 .3). Because the overwhelming majority of POM have been characterized by the electron microprobe, a more detailed account of representative analyses is given in Chapter 8**. The total POM listed in Tables 7.1 and 7.2 is about 100, after deleting some which would not qualify as discrete species if found today, but which are still recognized on historical grounds; e.g. iridosmine, osmiridium, platiniridium. Tables 7.1 and 7.2 also include 28 unnamed POM; included because they are considered to be probably new species (Cabri 1981 b). The referral scheme for unnamed minerals is the same as in Chapter 9. There are considerably more palladium minerals known than other POM; the proportion of total species (listed in Tables 7.1 and 7 .2) for the POE is: Pd 48%, Pt 21 o/o , Ru and Ir 9% each, Rh 8% and Os 5%. These proportions are in marked contrast to the estimated cosmic abundance, as compiled for average type C1 chondrite by Naldrett and Duke (1980): Pt 29%, Ru 20%, Pd, Ir and Os 15% each, and Rh 6%."

Jan 1, 1979

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 E. Swartzman

This paper presents a general re-view of the various methods of briquetting coals, with and without the aid of a binder, with special reference to the production of a suitable briquette from Nova Scotia 'fines' for re-blending with industrial coal. From the economic and technical viewpoints, either small stoker-sized briquettes made with about 4 per cent asphalt in standard low-pressure double-roll presses, or briquettes made without binder in high-pressure double-roll presses, appear to be the most reasonable solution at present.

Jan 1, 1959

By Thomas Deschamps

The main objective of this work was to investigate the effect of binder addition on the geochemical properties of surface paste tailings. Nine leaching columns were prepared following different disposal configurations: non-cemented, non-cemented with cemented layers, and cemented .The proportions of cement by total paste weight were between 0 and 1% for a whole column; the maximum amount of cement in a single layer was 2%. The material used was a sulphide-rich tailings coming from a Canadian hard rock polymetallic mine. The columns have been flushed once a week during 30 weeks. The recovered leachates were analyzed for pH, Eh, conductivity, and a wide range of elements. The drainage rates were monitored during the initial disposal as well as after each column flush. The results show that binder addition can improve the whole environmental performances for some of the studied disposal configurations. The disposal configuration consisting of intercalation of three 2% cement paste layers within non-cemented layers (0.66 wt.% of cement for the whole column), and the configuration with 2% cement added only in the two bottom layers (0,44 wt.% of cement for the whole column), showed the best results in terms of water quality and drainage rates.

May 1, 2007

By W. Duyvesteyn, B. Sabacky, T. Spitler, D. Verhulst

"The Altair TiO2 pigment process is based on hydrochloric acid leaching of ilmenite at high-acid concentration. Titanium is transferred by solvent extraction into a high-purity aqueous stream with a high titanium concentration. HCl gas is recovered by pressure-swing distillation and is used to regenerate the leaching solution.A new method of TiO2 recovery in dense film with simultaneous dehydration and hydrolysis provides great flexibility in adjusting the characteristics of the TiO2 pigment product.Modifications of this method lead to a flexible, low-cost, high-tonnage process first for the production of nano-sized TiO2, and later for stabilized zirconia-based nanomaterials and nano-sized titanates.IntroductionProcesses for the production of TiO2 from ore by a hydrochloric acid route have been proposed in the past. Most of these processes dissolve impurities and recover TiO2 as a residue or as a hydrolyzed phase in the original solution (Turner, 1948; Berkovich, 1975; Morris, 1982; Judd, 1986; Chao, 1993). Solvent extraction of titanium after leaching in aqueous HCl has also been studied (Tsuchida et al., 1982; Narita et al., 1983) or at least mentioned as an alternative (Ellis, 1963). A number of small-scale studies on leaching and solvent extraction in both sulphuric and hydrochloric systems have appeared recently (John et al., 1999; Lanyon et al., 1999; Sole et al., 1999. As far as we know, only Kerr-McGee (Morris, 1982) and a plant in India and one in Malaysia use a hydrochloric acid route (20 wt% HCl) to commercially produce synthetic rutile.The Altair process originated in a search for an economical method to treat high-magnesium ilmenite deposits, while avoiding the production of a waste metal salt by-product. The first attempts aimed at producing pure TiO2 by leaching out the impurities. Further work centred on completely dissolving the ilmenite in concentrated hydrochloric acid and reprecipitating TiO2. To decrease the impurity level in the product, solvent extraction of titanium into a purified stream was introduced. It also became clear that to take full advantage of the increase in leaching efficiency provided by concentrated acid, it was necessary to regenerate HCl gas."

Jan 1, 2002

By Paolo Toscano, Dave Hall

"This paper discusses the project history, process design and layout of the Rainy River Gold Project mill. The circuit consists of a 55 x 83 primary crusher, 15 MW SAG mill and 15 MW ball mill. Both mills are twin pinion with LSS and VFD drives. A pebble crusher is located in the process plant with bypass and stockpile options. Gravity concentration followed by 30 hours of leach residence, carousel CIP, carbon recovery, cyanide destruction and tails disposal in a 1,428 hectare tailings management area (TMA).The Rainy River Gold Project is located approximately 65 km northwest of Fort Frances in north western Ontario. The project is scheduled to be commissioned and reach commercial production in mid-2017. Expected cost is $US877M and is comprised of an open pit mine, process plant, tailings management area and supporting infrastructure. The process plant is designed to produce gold/silver Dore’ bars.The project has been developed on the following timeline:• 1967 – First exploration• 1994 – 17 zone discovered• 2004 – 433 Zone discovered• 2005 – Rainy River Resources Ltd. purchases property • 2006 - ODM zone discovered• 2007 - Cap Zone discovered• 2008 - First NI-43-101 Resource completed• 2012 - PEA completed & Intrepid Zone discovered• 2013 - New Gold acquires Rainy River Resources• 2014 - Updated feasibility study completed & Environmental Assessment report submitted• 2015 - Detailed Engineering >95% complete, EA approved and construction commencesThe property is accessed by provincial highway 600 and is within 17 km of 230 kV power line and railway sidings. Key project criteria are summarized as follows:"

Jan 1, 2016

By M. Tarvainen, J. Kallioinen

"Depressed metal prices coupled with increasing production costs have presented a great challenge for operators to control and reduce the costs of production to maintain profitability. Depleting ore reserves, added to the above has forced Outokumpu to take a hard look at its operations, to develop new methods to profitably extend the life of its mines. One of the areas investigated and approved for further development has been the use of larger and more energy efficient equipment. As a result, large flotation machines were developed and today 38 and 60m3 flotation machines are in use.Secondly, ore preconcentration has been developed to avoid processing the part of the ore that has little or no mineral values.Another quite recent development and the topic of this paper, has been the use of coarse flotation before conventional flotation. The main idea is that part of the valuable mineral is separated as early as practical by floating it with specially designed flotation equipment, installed in the grinding classification circuit. The new flotation cell is specially designed for this application and differs from the conventional cells in the tank design, feed arrangement and modified mechanisms in some cases."

Jan 1, 1984

By Galen Bauer

"Up until the mid 1960's, the only available methods for making continuous weight measurements were through the use of mechanical and electro-mechanical weigh scales. It was during this time period that a new concept in weighing was introduced. It was known and is still known as the nuclear belt weigh scale.During the early days of its appearance in the market place, many things were learned about its advantages and disadvantages for weighing various materials. As with many new devices, it had to go through a ""wringing out stage"" before it became accepted by the weighing industry. Even today, even though it is a proven device, there are skeptics who believe the nuclear belt weigh scale will not survive. Let me assure you, it is here to stay.. 'PRINCIPLE OF MEASUREMENTThe nuclear belt weigh scale consists of a ""C"" Frame assembly which houses the radioactive source (Cesiwn 137) and the radiation detector (Ionization chamber). The ""C"" Frame assembly is put in place around the conveyor belt as shown in Figure 1. The radiation source emits a fan shaped beam of radiation which covers the full width of the material being conveyed on the belt. Radiation which has passed through the material on the belt is detected by the ionization chamber which is located in the lower arm of the ""C"" Frame assembly. The ionization chamber changes the radiation energy into an electrical signal which is related to pounds per foot belt loading. This signal is transmitted to a highly stable amplifier which combines the pounds per foot signal with a belt travel signal from a digital tachometer which allowed the scale user to be able to monitor total mass being conveyed. Many so-called conventional type scales use belt speed rather than travel to obtain mass per unit time, and then integrate the signal to obtain their weight signal. This unnecessary step tends to erode the overall accuracy of the weighing system."

Jan 1, 1975

By Pravin Kumar

Goldcorp Inc.?s Red Lake Mine (RLM) is situated in the Red Lake Mining District, a history rich North Western Ontario mining community. Red Lake Mine formerly known as the Dickenson Mine has experienced various degree and type of bursting since the 1960?s. During the 1960?s burst were common in shrinkage stope box hole pillars. After discovering the High Grade Zone (HGZ) between its 30th and 39th levels, the mine has changed its mining method to mechanized cut and fill with cemented paste fill. The rock mass at RLM is mainly massive basalt-ultramafic and behaves in a manner consistent with massive brittle rock. This paper presents the use of empirical and numerical design techniques. The use of the Span versus RMR graph to determine critical stope spans at RLM and the use of numerical design techniques to determine the stability of the sills and post pillars. Selecting of a sill pillar dimension is illustrated using four different failure criteria?s. The four failure criteria are m & s failure criteria, the Pillar Stability Graph, deviatoric stress ???-?3), and induced stress level criteria ?1 /UCS, which are calibrated to RLM?s ground conditions. A wide range of case histories has been studied ranging from stable to highly burst prone ground conditions in order to calibrate the failure criteria to observed conditions and quantified to local micro seismic system.

May 1, 2002

By R. E. Jones

"INTRODUCTIONCominco Limited, one of the free world's largest producers of lead and zinc having majority ownership in Pine Point Mines Ltd., operates a mining and concentrating complex on the south shore of Great Slave Lake in the Northwest Territories. The mine was developed in 1963-64 and commenced operation late in 1964 with the direct shipping of high grade ore. The concentrator started production in November 1965 at a design throughput rate of 5,000 T.P.D.In 1968, the Company acquired the mining rights to a large ore body located to the east of the Pine Point claims and subsequently an expansion program at the concentrator raised the throughput rate to 8,000 T.P.D. The expansion included a third grinding circuit and additional flotation capacity. The flotation capacity took the form of a completely separate circuit from the existing flotation plant. In 1972, the zinc flotation circuits were modified to include two additional stages of cleaning. This change concluded the ""Pre-Consolidation Flotation"" circuit at Pine Point.In considering these modernization programs of the flotation section of the concentrator, the philosophy was to utilize the largest available machine having proven and modern technological changes. With the advent of the large capacity machines available in the late 1970's, it was considered a suitable time to revamp the entire flotation circuit at Pine Point in 1979. The Circuit Consolidation Project was developed to meet this requirement."

Jan 1, 1982

By A. P. E. Dirige

Extensive field study has been performed to assess the capabilities of thin sprayon liners (TSLs) and conventional spray-on support systems for preventing rock and support material damage that often results due to mining-induced rockbursts. Support performance was studied using field scale explosive detonation trials to simulate dynamic failure effects that are known to develop during typical rockburst events. Multiple seismic and high speed photographic monitoring techniques were used to provide detailed information concerning rock motion, surface fracturing, ejected fragment motion, and support liner survivability characteristics. Results of this study have demonstrated that TSL?s and variant layer combinations may be as effective as or better than conventional support materials for mitigating rockburst or like damage in highly stressed mine environments. In this study, field results have been verified using numerical modeling procedures to better understand the support behaviour of TSLs when subjected to highly stressed mine environments and mining-induced rockbursts, a factor that is essential when designing any rock support system. A series of numerical modeling assessments were conducted using FLAC3D modeling of a half circular TSL-lined tunnel, influenced by anisotropic stresses and simulated mining-induced rockbursts, to investigate the support potential of TSLs under prototype conditions. The effect of a mine rockburst was simulated using the three-dimensional dynamic analysis option of FLAC3D. Results of TSL-lined tunnels in highly stressed and mine rockburst conditions indicated exceptional effectiveness of TSLs for suppressing rock deterioration resulting from stresses and rockbursts. The rock support potential of TSLs on tunnel surfaces was compared with that of thin shotcrete linings, and results indicated that thin spray-on lining products currently available may be equally as effective in support as shotcrete materials.

May 1, 2009