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The effect of long term exposure of full size bolts to corrosive environments is presented. A special test rig was used to test four bolts under different loading conditions. Four, X-grade identical profile bolts, each of 21.7 mm core diameters (23.7 mm full diameter) were subjected to prolonged corrosion testing using acid sulphate water. The pH value of the circulated water varied between 3.4 and 4.3. The corrosion exposure test period lasted three and half years. Two bolts were axially loaded to 10 and 20 tonnes force respectively, the third bolt was subjected to a 360 Nm torsion load and the fourth bolt was left unstressed to act as a reference bolt. After the test period ended, the bolts were stripped of their corroded coatings and weighted for weight loss. The diameter of each bolt was subsequently measured, and the loaded bolt samples were first tested non-destructively for tensile cracks and then tested for tensile failure. No cracks were found on post corrosion bolts tested non-destructively. The failure strength reduction on all four post-corroded bolts was significant, varying between 21% and 39%. The onset of corrosion was not confined to the targeted mid-section length of the bolt, however, the severest corrosion occurred at the anchored ends of the bolts.

Aug 1, 2013

By J. Drenda

The model for methane removal from the workings directly behind an advancing longwall as presented in this paper may be used for U-ventilated cross or length-wise systems of longwall advance, starting from the boundaries of the extraction field. In mines with high methane concentrations, Uventilated longwall extraction from the boundary of the extraction field produces a methane risk. The risk is manifested by a higher methane concentration at the end of the longwall, upper gate and incline. The methane in increasing concentrations, which is emitted mainly from the workings with caved areas, appears towards the end of the longwall and in the upper gate. The graphical model of the longwall provided includes both the workspace of the longwall and the caved area. Based upon a model of the rock distribution in the cave, a ventilation model comprising two standard air splits and large number of crosswise air splits has been constructed. Simulations indicate that if the shearer is not present in the longwall, the air distribution in the caved area is symmetrical. The presence of the shearer, both at the bottom and the top of the longwall, causes an asymmetrical air distribution in the cave, that is the asymmetrical flows in the goaf and longwall workspace and in the inflows and outflow between those regions. This phenomenon causes an increase in methane concentration at the end of the wall and at the upper gate. The methane removal method proposed for workings with a caved area directly behind the advancing U-ventilated longwall, starting from the boundary of the field, consists of the removal of methane-air mixtures from the caved area with the boom pipes. These booms may be of several meters long and may be extended if necessary. The insertion of the boom pipes into the cave towards the end of the longwall and connecting them with the general methane removal system creates negative pressure causing a change in the depression field, which is the aerodynamic potential field in the goaf area around the ends of the boom pipes. This will promote the extraction of the methane-air mixture from voids and the goaf cracks, and the area from which methane is being extracted may extend to the side of the cave. This is a positive phenomenon, which could result in the methane flow from the goaf into the airflow in the longwall workspace decreasing or even stopping. A reduction in the methane flow into the longwall workspace with extractions booms will be advantageous for the production continuity. Based on the floor load model for the rocks in the cave and the methane-removal boom load model, the value of the force required to move the boom behind the advancing wall has been calculated. Strength calculations for the pipe were performed using a finite elements method. The analysis results, as well as the comparison of the pipe material effort to the strength parameters of steel, which can serve as piping material (for the assumed load conditions), lead to a belief that the proposed boom construction satisfies the strength requirements.

Aug 1, 2013

By P. Karimi, A. Amini, H. Abdollahi, A. Mirzaeian

Due to detrimental effects of cyanidation process on the environment in gold industry, application of cyanide alternatives such as thiourea are being highly considered in recent projects. In this study, the extraction of gold from poly metal ore, using thiourea was precisely investigated. The representative sample contained 2.5 ppm Au and 160 ppm Ag. The sample was categorized as oxide gold ore, because the main part of its pyrite was changed to iron oxide and hydroxide such as goethite and lepidocrosite. According to mineralogy, the main gangue mineral was quartz. It shows that gold grains are free and interlocked with quartz matrix. Also, the presence of a high amount of silver in free and sulfide forms and also the low content of other sulfide minerals such as base metal sulfides is a remarkable point of these studies. In the gold extraction process with thiourea, there are several parameters that affect the rate of gold dissolution such as concentrations of thiourea as a gold lixiviant and ferric sulfate as an oxidant, pH and size fraction of feed and leaching time. The influence of these parameters on the gold dissolution rate were investigated and optimized. The results showed that the concentration of thiourea is the most influential parameter which could increase the gold recovery to 71% (at 10 g/kg sample). The optimal amount of ferric sulfate was determined as 6 g/kg sample. Also, the optimal levels of size fraction of feed and pH of solution were 88% under 75 micron and 2 as acidity; respectively. Leaching time was 6 hrs. In addition, this sample was leached with cyanide and the optimized values of size fraction of d88=75 microns, pH=10.45, 2500 ppm of sodium cyanide with time leaching of 24 hours were concluded for maximum recovery of gold in solution. Finally, the results of cyanidation tests were compared to the thiourea dissolution after optimizing all affected parameters in cyanidation process.

Jan 1, 2016

By Neil Schofield

"The Pine Creek gold mineralization is associated with saddle reefs and quartz and quartz sulphide veining within the Enterprise Anticline. The anticlinal structure is complex and the mineralization was considered prior to development to be dominantly strata controlled. Gold grades vary up to 160 gin 1.5 m core samples with coefficients of variation in excess of three.The estimation of ore reserves for such a deposit must account for both the geological complexity and the extreme variability of gold grade.This paper describes the structural and variogram analysis of the gold grades and demonstrates the importance of incorporating the geological model in such an analysis. Variogram interpretation is greatly improved and greater insights into the finer geological structure areprovided. Part 2ofthis work presents the application of multigaussian kriging to reserve estimation for deposits with highly variable grade and which consequently demand highly selective mining operations. IntroductionThe Enterprise Gold Mine is situated some 220 km southeast of Darwin in Australia's tropical north. Gold is believed to have been discovered there in 1869 during the construction of the overland telegraph line.From 1894to 1915, some 75 000 ounces of gold were produced from 121 000 tonnes mined mainly by Chinese leaseholders. From 1906 to 1915, government-sponsored drilling led to the sinking of the Enterprise Shaft to 79 m and several drives were developed. Mining ceased because of war.Since late 1980, Goldfields Exploration Pty. Ltd. had explored on several mining tenements near Pine Creek under joint venture with the title holders, Enterprise Mines NL. This work culminated in the development of the current mine which opened on October 29, 1985.Feasibility studies to determine the economic viability of mining were based on ore reserves generated by project geologists using a sectional polygonal approach tuned to the interpreted geological control of the mineralization.The study motivating this two-part paper covers are-estimation of the recoverable reserve s using the geostatistical approach of multigaussian kriging developed at Stanford University, (Verly, 1984)."

Jan 1, 1988

By A. A. Ruitenberg, G. Ruitenberg, J. Chandra

"The Caledonian Belt is mainly underlain by Late Precambrian and possibly Early Palaeozoic rocks, which are in part overlain by younger strata. Most of these rocks are weakly deformed, except for those in some linear . zones, which have been intensely deformed by cataclasis. Mineral deposits related to two distinct metallization episodes have been recognized. The earliest are zinccopper- lead sulphide deposits, which have been concentrated in the axial regions of late-stage cross-folds within the main cataclastic zone. The host rocks are andesitic and dacitic tuffs which show effects of intense talcose alteration, silicification and chloritization. Deposits related to the second episode are mainly composed of copper sulphides with locally abundant tennantite and tetrahedrite, which occur in brecciated and fractured carbonate-quartz veins. These veins postdate the late-stage cross-folds. VLF-EM and high-sensitivity magnetics proved to be very effective in delineating the geology of these deposits."

Jan 1, 1972

By W. Blum

Progress made in improving the creep resistance of Mg alloys is reviewed with the emphasis on microstructural interpretation. Experimental results on the increase of creep strain with time at constant stress a measured for limited times and strains are displayed in a transparent form as creep rate-strain diagrams from which the minimum creep rate stress is safely derived. In addition, data on the thermal stress component a derived from high-precision a-reduction tests are reported. Within a simple statistical dislocation model these data allow one to quantify the relation between dislocation velocity v and a . In combination with structure evolution laws this relation serves as a basis for modeling the creep rate. New alloys have distinctly higher creep resistance than AZ91 where precipitation hardening plays a minor role. This is due to hardening by intermetallic phases which may form a continuous skeleton along the grain boundaries (e.g. Ca-alloyed AZ91) and/or submicron-sized precipitates within the grains (MRI230D, Mg-Zn-Ca alloys).

Jan 1, 2006

By M. P. Tibbs

An autonomous spray application system for a structural Thin Spray-on Liner (TSL) is targeting productivity and safety on underground coal continuous miners by potentially replacing steel mesh, and has broader applications as surface support and confinement in metalliferous mines and civil tunnelling. Polymer chemistry, coating formulation, geotechnical assessment and the autonomous application system are being developed as an integrated package as part of a continuous miner automation project. This polymer-based TSL system has significant benefits over conventional steel mesh skin confinement, including being more automatable by spray application, providing complete skin confinement and active skin reinforcement as a fast-curing composite with substrate, and having effective fire retardant properties for the polymer and the underlying coal. This paper presents a preliminary assessment of distance sensing systems that would enable autonomous adaptation of a nominal spray path to spray a consistent thickness of TSL (nominally 5mm) over an irregular topography. Two alternate sensors are assessed; firstly, a 2D LIDAR mounted on a linear axis is used to produce a 3D surface profile for generating an adapted spray path; and secondly, a robot end-effector-mounted ultrasonic sensor is assessed for wall following techniques to dynamically adapt a nominal spray path.

Aug 1, 2013

By C. A. Petch

Abstract - The High Lake polymetallic deposit is located in the northern Slave province, 40 km south of Coronation Gulf, Nunavut. The main zone, the AB zone, hosts an estimated resource of 3.56 Mt grading 5.0% copper, 1.1% zinc, 0.06% lead, 1.7 g/t gold, and 18.2 g/t silver. Archean felsic metavolcanic and metavolcaniclastic rocks are the dominant rock type and are associated with lesser mafic metavolcanic and carbonate-rich metasedimentary rocks. The sequence strikes north-northeast, dips steeply to the west, and is truncated to the west by a large granodioritic intrusion. The mineralization at the AB zone is comprised of massive and stringer chalcopyrite-sphalerite-pyrrhotitepyrite± galena. At the core of the mineralized zone, intensely chlorite-altered rocks are locally metamorphosed to form distinctive zones of anthophyllite-magnetite-cordierite-altered rock. Peripheral to the mineralized core, alteration consists of sericitization, silicification, and weaker chloritization. Primary morphology and facing direction of the zone are difficult to interpret in part due to poor primary metal zonation of the deposit. However, the irregular shape of the mineralized zone and the erratic zones of zinc enrichment at the fringes are consistent with the interpretation that the zone was formed by the replacement of a porous volcaniclastic pile.

Jan 1, 2004

By Laxman (Lucky) Amaratunga, Hisham Madi

Final disposal of mine wastewater neutralization sludge continues to be one of more pressing problems in the mining industry. The disposal of mining and metallurgical waste products into underground reduces the environment impact and improves the safety of the mining operation and the economics of the enterprise. This current research focuses on an investigation of utilization of mine wastewater neutralization sludge obtained from Kidd Creek Mine site, as a supplementary material for their proposed deep mine pastefill system. Their proposed pastefill consist of combination of rock aggregate in a matrix of imported sand and tailings with binder (Portland cement + Slag cement). Pastefill mixes containing various percentages of the waste sludge 0 (control mix), 1, 2, 3, and 5% by weight as replacement to the fine sand were prepared and tested with different percentages of binder. Hypothesis behind this approach is to study the behavior of sludge as replacement of sand in matrix as well as its ability to act as a binder. Compressive strength, percentage of sludge, binder dosage, slump, water to cement ratio, percentage of solid, and curing time are being observed. The results indicated that sludge can be incorporated into the pastefill and enhance its strength without compromising the pastefill quality and integrity.

Jan 1, 2004

By C. J. Clayton, A. J. Haynes, D. A. Hickson, B. Thiele

The Meadowbank Gold Project, formerly owned by Cumberland Resources Ltd. and currently owned by Agnico Mines Ltd., is currently being developed approximately 70 km north of Baker Lake, Nunavut, in the eastern Canadian Arctic region. The Meadowbank Gold Project currently consists of a series of gold-bearing deposits within relatively close proximity to one another, which are planned to be mined primarily as a truck-and-shovel open pit operation using conventional gold milling and extraction. The mining plan indicates that approximately 20 million tonnes of ore will be mined over a mine life of eight to ten years. The operation will generate approximately 160 million tonnes of mine waste rock and about 20 million tonnes of tailings. The site is located in an area of continuous permafrost with mean annual average air temperature of -11°C and average annual precipitation in the range of 100 to 200 mm. Lakes cover a large portion of the site, including some of the ore deposits. The location of the site and the physical environment pose challenges to development. This paper presents some of the designs considered to effectively manage and utilize the waste products of mining at the project. Note that volumes, tonnages, construction concepts, and data presented in this paper are accurate for the time of writing and are subject to change and modification as the project progresses, additional modelling is carried out, and additional data are obtained.

Jan 1, 2007

By S. Gheibi, H. S. B. Duzgun, N. Dadashzadeh

It is widely recognized that there are various uncertainties in the analysis of rock slope stability owing to inadequate information for site characterization and inherent variability and measurement errors in geological and corresponding parameters. Therefore, reliability-based approaches that allow the systematic and quantitative treatment of these uncertainties have become a topic of increasing interest in rock slope engineering. There has been extensive reliability analysis of rock slope stability in the literature. However, all of these analyses are based on the First-Order Reliability Method (FORM). The probability of failure of a rock slope in the First-Order Reliability Method is provided by a linear approximation of the failure function. The FORM is most commonly used because it is efficient; its accuracy, however, deteriorates when the nonlinearity of limit-state function increases. A better approximation of the failure function can establish a more accurate value of the probability of failure. The Second-Order Reliability Method (SORM) overcomes this drawback with a cost of lower efficiency in terms of the number of function calls and computation expenses. Despite considerable studies of the application of the SORM in other engineering fields, a few attempts have been made in the slope stability studies. In this study, a recently published Second-Order Reliability Method with First-Order Efficiency (SORM-FOE) is implemented on a selected rock slope in Amasya, Turkey. The SORM-FOE overcomes the low efficiency of the SORM by decreasing the number of function calls. The probability of failure of both FORM and SORM-FOE are obtained and the results are compared. It is shown that, in high value of failure probability, SORM’s do not significantly affect the analysis and design. However, in very low failure probability, the discrepancy between FORM and SORM-FOE becomes pronounced. It is also shown that, where the second order methods are required, the SORM-FOE is more efficient than the conventional curvature fitting methods in terms of computational costs.

Jan 1, 2015

By P. J. Deveaux, M. J. Kreczmer

The Tartan Lake deposit 12 km northeast of Flin Flon, Manitoba, was discovered in March of 1984 during the investigation of a VLF anomaly. The discovery hole intersected 16.83 metres averaging 21 gm Au/tonne. Since then, 99 holes, over 21,000 metres, have been drilled outlining 419,491 tonnes grading 11.79 gm Au/tonne in a di stinct group of mineralized lenses referred to as the Main Zone. Three other mineralized zones are known on the property but no reserves have been allowed for these, the South , East and Base Line zones. The Main Zone consists of seven parallel, east-west striking, quartz lenses. The lenses dip steeply to the north and plunge steeply west to northwest. Lens number l strikes for 160m and extends vertically to the 90m level. Within th is lens there is a 60m interval of low grade materi al. Lens number 2 is located about 5-l0m within the foo twall of Lens number I. Near surface this lens stri kes for 40m, at 200m level the strike length increases to 200 m. To date the deepest mineralization on Lens 2 has been intersected at a vert ical depth of 527m. Five smaller lenses are located in the foo twall of Lens 2 , containing gold va lues similar to the two main lenses but erratically distributed. The South Zone, located 100m south of the Main Zone, strikes east-west fo r l00m above the 100m level. The East Zone, found 230m east of the Main Zone, strikes fo r 80m and may possibly be an ex tension of the South Zone. The Base Line Zone is located 275m south of the Main Zone. Sporadic, but very significant gold intersections have been obta ined1n all three zones. Gold-bearing quartz-ca rbonate tourmaline veins are generally confined to a silicous crystal tu ff hori zon. The crystal tuff exhi bits fi ve distinct types of alteration: silicification, sericitization, ca rbonati zation, chlorit ization and pyritization. Silicification has resulted in a perasive addition of silica and the generation of multiple quartz veins. Sericiti zation is restricted to the crystal tuff. All rock units are pervasively carbonatized . Ankerite is abundant in the quartz veins and within the shear zones, particul arly those located along the chilled margin of a nearby gabbro-diorite intrusion. Chloritization is widely spread th roughout the rock units, while pyritization fo rms a broad alteration halo associated with the orebody. Sulphide accesso ry minerals associated with the gold , whi ch fo r the most part appears to be free, are chalcopyrite and pyrite. Generally, the higher grade gold values are related to an increased content of sulphides in the qua rtz veins. The average sulphide content in the deposit is less than 5 %.

Jan 1, 1986

By J. E. Nesset

"This paper discusses the importance of considering residence time distribution (RTD) in the design of flotation and mixing circuits and in the analysis of operating problems. The experiences and successes of Brunswick Mining in this area are covered and include: the development of a practical RTD measurement technique using pulp conductivity, baffle installation to improve the RTD of existing process equipment, and the method in which RTD considerations have been used to design the newest flotation and conditioning operations. IntroductionThe use of residence time distributions (RTDs) in the design and analysis of unit operations is well understood by the chemical engineering profession in while mineral processing engineers have tended to rely on empirical and rule-of-thumb factors. For example, in scale up from laboratory and pilot test results to full size flotation plant design, empirical rule s are often used to establish residence time and number of cells in the bank. Coleman suggests a factor of at least 2, and up to 5 for gold ores, be applied to laboratory determined flotation times. This is further illustrated in Table I, taken from Denver equipment.The number of cells in series to achieve the required retention time while minimizing short-circuiting has generated a variety of conclusions (Table 2). Obviously there is less than complete agreement on the subject. Young notes that a clear difference of opinion has emerged and cites as extremes, (i) the publications of Outokumpum showing unimpaired recovery for its largest machines (38 m3) compared to equivalent retention times for numerous smaller machines, and (ii) the experiences of Eastern European countries where the practice is to install banks of 30 unit s or more, operating in series, often with unit retention times of only a fraction of a minute.The empirical factors used by engineers presumably account for other than tested feed conditions, inefficiencies in large continuous machines, inevitable day-to-day circuit disruptions, gradual equipment deterioration and any other safety factor the engineer feels compelled to incorporate. An understanding of RTD concepts and techniques should help the practicing mineral engineer to develop more soundly based factors. This paper relates the experiences of Brunswick Mining in using RTD techniques in the unit operations of flotation and mixing."

Jan 1, 1988

By G. O. Crowell

This paper gives the history of the celestite deposit, now being mined by Kaiser Celestite Mining Ltd., near Sydney, Nova Scotia. Mining and milling procedures are covered, and the regional and local geology of the area are described, including the origin of the celestite.

Jan 1, 1971

By Garston H. Blackwell

For ageing open pit mining operations where the economic viability of another pushback is questionable, mine life can be extended by increasing the wall slope. If the steepening is started where ore intersects the pit wall, the capital outlay for waste stripping is minimal. Such a venture must be conducted safely, and a basic requirement would include loading equipment capable of working remotely. In the application described, alternatives in the choice of loader included equipment available on site, or the purchase and assembly of used equipment. Only electric cable shovels were in use at the operation, and it was thought that these would be more capable of digging high muck piles at the productivity demanded by an increasingly efficient mill, and in a challenging environment less suited to a rubber tired loader or hydraulic front shovel. A decision was made to convert an older electric cable shovel to remote controlled operation and to include the application of semi-robotic control. The conversion was completed successfully and is described, along with methods developed for remote operation of a shovel, and the application of various sensors for semi-robotic operation. A plan of operations to mine out several completed double benches and make two quadruple benches with wide catchment berms is depicted. On completion of this task, the working benches developed would be wide enough to operate at near full production. Prior to moving the remote controlled shovel back into the pit, a large wall instability became active, eventually closing the operation. The instability could have been safely mined using the remote controlled shovel.

Aug 1, 2013

By James Stenhouse

Abstract The aqueous oxidation of pyrite in caustic solutions under oxygen pressure was investigated by measuring that effect of reaction variable on oxygen consumption by the reactions. From the results of this investigation, a model for the process has been proposed. In this model, sulphur diffuses as a positive ion through an oxide layer which forms on the sulphide particles and is then oxidized to sulphate ion. Iron is oxidized by substitutional diffusion of oxygen atoms into the oxide layer. The effect of most reaction variables on the rate of oxygen consumption can be explained with this model. Introduction Recent investigations (1,2,3) have shown that certain sulphide ores can be readily leached by agitating the ore in the form of a pulp under pressure in the presence of oxygen. Since very little was known of the mechanism of this oxidation, a programme of investigation was inaugurated in 1949 in the laboratories of the Department of Mining and Metallurgy of the University of British Columbia. Financial support for the investigation was received from the National Research Council in the form of student grants and bursaries, from the Canadian Liquid Air Company, Limited, and from the Consolidated Mining and Smelting Company of Canada, Limited. The first objective was to devise a technique to study the primary oxidation reactions in aqueous solutions and to apply the technique to a detailed study of the oxidation of pyrite in caustic solutions.

Jan 1, 1952

By Enzo Angeles, Desmond Nanchengwa, Lee Julian, Julian Ruiseco

Since the early 1980s, surface hard-rock mining companies have searched for a blast movement model to enable grade control optimization efforts. Significant time and money were spent attempting to create blast simulations to aid post-blast grade control. Despite nearly 40 years of effort, practical blast movement models remained elusive until computing power reached levels that make real-time simulation possible. From the mid-1990s, mines have been creating in-situ grade control polygons and translating the polygon nodes horizontally to account for blast movement. In many cases, objects are placed in the blast volume and located post-blast to provide local vectors for these node translations. This method was always a compromise between practicality (2D) and reality (3D), but it was the only viable solution that matched the needs of operations. This study considers 21 blasts and evaluates the incremental value achieved by new processes unlocked through 3D modelling, while investigating challenges, limitations, and opportunities for each method.

Jan 1, 2023

By J. E. Nesset

"This paper discusses the importance of considering residence time distribution (RTD) in the design of flotation and mixing circuits and in the analysis of operating problems. The experiences and successes of Brunswick Mining in this area are covered and include: the development of a practical RTD measurement technique using pulp conductivity, baffle installation to improve the RTD of existing process equipment, and the method in which RTD considerations have been used to design the newest flotation and conditioning operations IntroductionThe use of residence time distributions (RTDs) in the design and analysis of unit operations is well understood by the chemical engineering profession in while mineral processing engineers have tended to rely on empirical and rule-of-thumb factors. For example, in scale up from laboratory and pilot test results to full size flotation plant design, empirical rule s are often used to establish residence time and number of cells in the bank. Coleman suggests a factor of at least 2, and up to 5 for gold ores, be applied to laboratory determined flotation times. This is further illustrated in Table I, taken from Denver equipment. The number of cells in series to achieve the required retention time while minimizing short-circuiting has generated a variety of conclusions (Table 2). Obviously there is less than complete agreement on the subject. Young notes that a clear difference of opinion has emerged and cites as extremes, (i) the publications of Outokumpum showing unimpaired recovery for its largest machines (38 m3) compared to equivalent retention times for numerous smaller machines, and (ii) the experiences of Eastern European countries where the practice is to install banks of 30 unit s or more, operating in series, often with unit retention times of only a fraction of a minute.The empirical factors used by engineers presumably account for other than tested feed conditions, inefficiencies in large continuous machines, inevitable day-to-day circuit disruptions, gradual equipment deterioration and any other safety factor the engineer feels compelled to incorporate. An understanding of RTD concepts and techniques should help the practicing mineral engineer to develop more soundly based factors. Thi s paper relates the experiences of Brunswick Mining in using RTD techniques in the unit operations of flot ation and mixing"

Jan 1, 1988

By V. Shumila, T. Butler

Microseismic monitoring systems listen to a rockmass and record ground vibrations associated with brittle rock fracture. Modern systems record continuously and can identify as many as one thousand small and possibly larger magnitude events occurring each day from induced stress change and redistribution. This paper presents examples of the real time use of microseismic information in mining applications. It shows how alerts can quickly identify large magnitude events or unexpected clusters of seismicity and how seismic re-entry protocol can help to improve safety decisions. Additionally, newly developed advanced analysis techniques applied to mine seismic data sets are shown including the use of clustering methods to see into a cloud of seismic data to identify significant activity and quantify possible planar features. The use of source parameter trend analysis is shown to be a beneficial method for interpreting changes in relative stress levels. Seismic moment magnitude distributions are analyzed using the TGR (Truncated Gutenberg Richter) and TMD (Tapered Moment Distribution) methods to estimate bvalue and expected maximum magnitude. Seismic hazard analysis is shown as a useful way to simplify a dataset and highlight seismic hot spots. The rockmass response to excavation is a complicated spatial and temporal process and microseismic monitoring is a method that can be used to help understand the location and mechanics of failure occurring in the rock ultimately helping to improve safety and productivity at a mine site.

Jan 1, 2015

By J. H. Gray

"IntroductionIn 1980, Fording Coal Limited started conceptual planning studies which led to major changes in the long-range mining plan at their Fording River Operations. These new plans have been completed and are now in operation.To assist personnel in a timely completion of this major revision to the long-range plan, existing in-house drill hole data base and scheduling systems were used and a computerized Mine Modelling and Planning System was purchased. With these tools, plan s were developed which led to completion of initial access to the peak of Eagle Mountain and the start of mining operations in March 1983. As mining area s open up and new geo logical data is obtained, detailed plan s are being extended and revised.With the geology and mine plan s on line, this work can now be done with less effort and plans ca n be adapted to provide flexibility in mining operations. Changes in market conditions can be accommodated by moving mining operations into different seams for different coal types. The computer systems are used to revise pit designs and schedules accordingly."

Jan 1, 1986