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By A. K. Bakshi

Slurry rheology affects autogenous grinding performance by changing the rock's breakage characteristics. In this study pilot scale batch autogenous grinding was carried out to observe the effect of viscosity on the breakage characteristics of the rocks (the media), and on the overall specific energy consumption. Clay/fines content, and temperature of the initial charge, were varied to alter the viscosity of the slurry. Results showed that at higher slurry viscosity, rocks became more difficult to break, resulting in an overall increase in the specific energy consumption of the mill. Also, it was observed that viscosity had a significant effect on the amount of critical size material present in the grinding circuit, where the critical size material is the material which is too large to be effectively ground in the mill, and too fine to act as grinding media.

Jan 1, 1996

By P. G. Corser

A gold and silver open-pit mine located in northwestern Chile contracted ACZ in 1991-92 to assess the stability of their spent ore pile and to determine the cause and magnitude of cyanide contamination associated with leaching operations. The mine is a conventional open-pit located at an elevation of approximately 4,800 meters (m.) (15,744 feet). A spent ore pile is located approximately 2,000 m. southwest of the mine and approximately 1,000 m. northeast of the active pad and plant facilities. The pile was constructed on a gentle northwest slope dipping 12 to 13 degrees. The pile is bordered by a small river to the northwest and by a larger river to the northeast. Both rivers eventually flow into Bolivia. Heap leaching and processing commenced at the mine in 1988. Since 1988, nearly 6.0 million tons of ore at an average grade of 1.96 grams/ton (0.07 ounces/ton) of gold and 52 grams/ton (1.83 ounces/ton) of silver have been treated by cyanide heap leach methods with zinc (Merrill Crowe) precipitation. In approximately three years of operation, the mine has yielded more than 225,000 ounces of gold and over 1.5 million ounces of silver. Currently ore is agglomerated and hauled to the leach pad by truck and is top loaded to produce a 5-meter lift height. Four contiguous reusable (on/off) leaching pads were constructed with an asphalt layer above a carefully compacted sub-base of High Density Polyethylene (HDPE) liner and screened fines. Leach solution of pH 10-11 containing 1.2 to 1.5 gram/liter of cyanide is applied to the heap at a rate of 10 liter/hour/square meter for a total active application period of 30 days. Leached (spent) ore is transferred to the spent ore pile located approximately 1,000 meters to the northeast. The pile currently contains approximately 6 million tons of spent ore. The pile liner, which has been constructed over a prepared sub-grade consists of three geosynthetic layers (geotextile/flexible membrane liner/geotextile). Long term objectives include washing the pile to recover an additional 3 to 4 percent gold and silver.

Jan 1, 1993

By Ramazan Comakli, Cemal Balci, Nuh Bilgin, Hanifi Copur, Can Polat, Deniz Tumac

"Application of mechanical excavators for rock excavation in both civil construction and mining engineering fields has increased significantly in recent years. Prediction of the excavation performance of any mechanical excavators – tunnel boring machines, roadheaders, impact hammers, continuous miners and shearers for any geological formation is one of the main concerns in determining the economics of a mechanized mining and/or tunneling operation. A new portable linear rock cutting (PLCM) device was developed in the Mining Engineering Department of Istanbul Technical University and supported by The Scientific and Technological Research Council of Turkey (TUBITAK-112M859). The PLCM is used for cutting small core or block rock samples by a real-life mini disc cutter. The results from the PLCM are compared with the results obtained from the full scale rock cutting tests and theoretical model. It is found that there is reliable correlation between test results obtained from both cutting tests and theoretical model. 1. INTRODUCTION Mechanical excavators are currently the only alternatives to drill and blast due to higher production rate, increased safety, uniform muck size, less roof support, easier ventilation, continuous rather than cyclical operation, and possibilities for automatic control. The most of large civil tunneling and some mining projects are now carried out by mechanical excavators rather than drill and blast methods. Mechanical excavators can be categorized into two main groups: full-face and partial-face machines. Full-face machines such as tunnel boring machines excavate the face in one mode at the size of the cutterhead. Due to their cutting mode, the shape of the face profile by the full-face machines can only be circular. As for partial-face machines, the cutterhead is attached at the end of a boom. The face profile excavated by the partial-face machines can be from circular to horseshoe shape due to the cutting mode of the boom."

Jan 1, 2016

By Tony B. Szwilski, K. Unrug

INTRODUCTION Since the introduction of longwall coal mining systems, reliable predictions of roof support requirements at the face and the roof caveability in the gob area have been major contributors to the success of the longwall face by means of safety and face productivity. That is, the successful operation of a longwall is well-served by good regular caving of the gob. In recent times, with the advent of increased coal face mechanization and capital investment, the evaluation of the feasibility of longwall mining has become essential. In the United States, where a face unit for a face length of about 150 metres may cost around $6 million, a quantitative system that would assist in selecting the best equipment for the coal would prove extremely beneficial. There have been many such varied investigations in the past few years. A vast -amount of work has been invested principally by West Germany [3, 7, 8, 91 , Poland [I, 101 , United Kingdom [4, 5, lo], France [4], and South Africa [4]. It is the object of this paper to review this previous research work and give an outline that may be used as guidelines for long- wall evaluation. Essentially, for the initial design phase for a longwall system, account must be taken of the geometry of the coal panel, its position relative to the surface (depth), direction of the coal cleat system, faults and in-situ stresses and other neighboring workings in the same coal seam and other seams both above and below. The advance of the coal panel disturbs the stress field and in doing so creates an abutment zone around the face. At this stage the condition of homogeneity of the surrounding rock mass does not apply. The magnitude of the rock strata pressure may exceed the ultimate strength of the rocks and create zones of clastic, plastic and elastic material, the limits of which are difficult to determine. The collapse of the gob and resulting expansion of the displaced waste roof rock, create a statically undefined support complicating an already unclear rock structure. For example, it is often a point of debate whether, indeed, the vertical strata pressure develops to regain the original magnitude of virgin stress. Under these post-failure circumstances an analytical approach based on simplified and idealized assumptions can only be considered a very general approximation. Theoretical concepts, however, may be adequately supplemented by empirical methods that are based on numerous underground observations and surveys [3, 4, 7, 8]. These methods employ the use of rock indices that very handily correspond to the competence of the roof rock mass. Scale model work [3, 41 , using equivalent materials representing the surrounding rock mass within the coal seam zone, have given valuable qualitative and sometimes quantitative data in addressing the behaviour of the rock mass around longwall panels. To characterize the state of art of caveability predictioqthe principal concepts will be reviewed with special emphasis placed on practicality for the planning of longwall operation. It has been the general observation of the authors that many of the literature surveys published in English often neglect reference to other international mining literature. It is with this point in mind that an overall review has been written including mining concepts identified with the principal coal mining countries. THEORETICAL CONCEPTS Arching Theory For the development of this theory, assumption has been made that an extracted height, h, in a coal seam results in an extracted panel of width, w . Also, the coal seam is horizontal and the rock strata of both the roof and floor are homogeneous. The primary (virgin) stress distribution will change in the roof as well as the floor. As a result of lack of support caused by the excavation (in the gob area) an additional load is developed due to the redistribution of the load around the sides of the panel, referred to as the abutment, R, where: [ ]

Jan 1, 1982

By Sandip K. Mitra, John E. Dorn

The Basinki-modified Seeger equation for deformation of metals by means of thermally activated intersection of dislocations was used to ascertain the various factors responsible for the strain hardening of poly crystalline aluminum and a series of a solid solutions of magnesium in aluminum. The analyses were based principally on the rate of change of strain rate with stress and the Cottrell-Stokes ratio. From measurements of these quantities on a series of strain-hardened states, the force displacement diagram for intersection, the density of dislocations in the entanglements, and the contributions of the athermal local stress and long-range stress interactions were deduced. Poly crystalline aluminum was observed to have the same force-displacement diagram for intersection as single aluminum crystals, and the corresponding diagram for the magnesium alloys differed in a manner that suggested that small amounts of magnesium may decrease slightly the stacking fault energy of aluminum. Although poly crystalline aluminum had about the same initial density of dislocations as the single crystals, the density of dislocations increased much more rapidly with strain in the polycrystals. This factor and the higher back stresses generated in the poly crystalline aluminum were the major factors responsible for the higher strength of the poly crystalline aluminum, As the magnesium content of the alloys was increased, the initial density of the dislocations also increased. Most of the alloy strengthening arose from this factor; alloying effects due to short-range order strengthening, Suzuki locking, Cottrell pinning, and so forth, were estimated to be quite small. The gveater rate of strain hardening in the alloys was deduced to be due primarily to the greater increase in the density of the dislocations with strain as the magnesium content was increased. It is the purpose of the present investigation to shed additional light on the average plastic behavior of a grain in a polycrystalline aggregate, by analyses, based on dislocation theory, of appropriate experimental test results. These deductions when compared with those from single-crystal data will then serve to illustrate the major items of difference between the plastic behavior of single crystals and the average behavior of a grain in a

Jan 1, 1963

By D. W. H. Su

Following a roof fall in the B12 headgate of a CONSOL Pennsylvania Coal Company (CCPC) underground coal mine in February 1998, CONSOL R&D and Exploration initiated an underground roof geology reconnaissance program along the B13 and B14 developments to define potential roof instability areas and to recommend appropriate supplemental headgate support. This reconnaissance program revealed the presence of a thick, massive Pittsburgh sandstone channel overlying the B13 and B14 panels. Surface and underground core holes confirmed the close proximity and the very massive, nonlaminated nature of the sandstone. To mitigate anticipated adverse conditions at the B 13 and B 14 longwall faces due to delayed caving of the massive sandstone behind the shields, a task force consisting of Operations, Exploration, and R&D personnel was formed in May 1998. Several operational adjustments were made by mine management, including repairing damaged lemniscate links, maintaining immediate forward shield support, and implementing measures recommended by the task force intended to boost the shield set load density at the face. After reviewing available geotcchnical information, the task force also recommended fracing of the massive sandstone over the B13 and B14 panels to enhance caving characteristics and further reduce the bending-induced stresses in the sandstone near the face. Horizontal fracs were created at a horizon about 25 feet above the Pittsburgh seam main bench over the two panels in June and November 1998, respectively. To evaluate the effect of sandstone fracing and operational adjustments on the bending-induced stresses in the sandstone, three finite element models were constructed and analyzed. Results from the finite element analyses indicate that one horizontal frac, 25 feet above the Pittsburgh seam, would reduce the bending-induced stresses in the sandstone by approximately 50 percent, maintaining them at a level below the ultimate strength of the sandstone. Operational adjustments to increase the support set load density and support stiffness could reduce the bending-induced stresses in the sandstone by a maximum of an additional 10 percent. Shield pressure monitoring at the B14 midface confirmed the presence of a lower roof pressure zone inby and outby the 1314 frac holes. This lower face pressure significantly contributed to improving the B13 and B14 longwall productivity and safety. The total production downtime due to bad top at the B 12 face within the interval of channel influence was 30,600 minutes compared to about 3,500 minutes each for the B13 and B14 faces. The average longwall advance per machine shift for the B13 and B14 faces within the channel was about 70 percent higher than that for the B 12 face.

Jan 1, 2001

By Panagiotis (Takis) Katsabanis, Lyall Workman, Jack Eloranta

The reduction of metallic ores to concentrate consumes a great deal of electrical energy. In the past two decades, a number of case studies have demonstrated a beneficial relationship between powder factor and grindability. However, changes in mill performance cannot be attributed to changes in feed size alone. Blasting energy and the distribution of that energy appears to affect subsequent comminution beyond crushing. This experience has prompted researchers to attempt to duplicate and quantify the mechanisms involved. To date, published reports indicate a measure of success. However, the blast engineer and the mill engineer still lack the tools and basic understanding necessary for full drill to mill optimization.

Jan 1, 2015

The introduction and subsequent ascendancy of autogenous and semi-autogenous milling for comminution circuits has undoubtedly lead to many economic advantages. From a process viewpoint, however, at least one downside has been the sensitivity of these machines to process input variations. Foremost among these is ore competence, though feed size comes a close second. This sensitivity is due to the reliance of ag/sag mills on the feed ore for their grinding media competence and size. Autogenous (ag) mills are the most sensitive in this respect with semi-autogenous (sag) mills being increasingly less so as the ball charge is increased. Increasingly designers and operators are recognising the importance of feed size and the benefits of manipulating it to improve performance. This has been achieved by installing additional crushing/screening capacity ahead of the ag/sag, better operation of the primary crusher, better control of segregation cased by stockpiles and bins and by changing blast design/practices to modify the ROM size distribution. This paper describes the influence of feed size distribution on ag/sag mill performance with reference to a range of plant data. It particularly concentrates on how changing blasting practice to modify ag/sag feed size can have a considerable overall benefit.

Jan 1, 2003

By Sanjeeva Latchireddi

"Deteriorating ore grade and high demand of metals has tremendously influenced the unit capacity of SAG mills. Pebble circuits have been added to the existing SAG circuits to increase their unit capacity. To make the pebble circuits justifiable, the grate openings are increased almost to 4.5 inches to draw large quantity of pebbles. The inherent transportation problems associated with SAG mills in closed circuit with pebble fraction are more complicated than the single stage SAG mill operation, because the coarse pebbles behave significantly different to that of slurry in pulp lifters. In addition, operation of mills at higher speed, to take advantage of precise design of high release angled shell lifters significantly reduces the discharge efficiency of the conventional pulp dischargers. The presence of a sizable quantity of pebbles in the pulp lifter limits the space available and reduces the flow gradient through the grate thus increases the load inside the mill and hence the mill power-draw. This paper analyses all the problems associated in discharge of slurry and pebbles in SAG mills with pebble circuits. The Outokumpu patented Turbo Pulp Lifter (TPLTM) aims at eliminating the inherent material transport problems to ensure good grinding conditions. Recent installation of TPLTM in a 26-ft SAG mill has proven these claims by reducing the specific power consumption while increasing the mill capacity.INTRODUCTIONThe conventional wisdom is to look at the semi-autogenous grinding (SAG) mill as just another grinding mill in which ball charge, mill filling, and mill speed are varied to attain maximum throughput. Liner design for the SAG was also approached more from a wear perspective rather than a metallurgical and hydraulic-flow perspective. The extensive research work carried out by Latchireddi and Morrell (1997,2003a and 2003b) has demonstrated to think of the SAG as more of a pump in which the grates and pulp discharger end (“impeller”) were designed to efficiently discharge the product sized particles smaller than the grate size with minimal recirculation out of the mill.The rock load in the mill is essentially depends on ore characteristics and the discharge rate of broken particles. The discharge rate of the product material depends on how efficiently the discharge pump (grate and pulp lifters) is operating. Similar to the impeller design affecting the pump capacity, the pulp lifter design affects the discharge capacity (or mill throughput) of the SAG/AG grinding mills. The general arrangement of AG/SAG mills discharge system is shown in Figure 1."

Jan 1, 2007

By William J. Conrad, Ryan J. Molka, Erik C. Westman

"In order to study pillar and overburden response to retreat mining, a ground control program was conducted at a Central Appalachian mine. The program consisted of several monitoring methods including a seismic monitoring system, borehole pressure cells in the pillars, and time-lapse photogrammetry of the pillar ribs. Two parallel geophone arrays were installed, one on each side of the panel with the sensors mounted 3m (IO ft) into the roof. A total of fourteen geophones recorded more than 5,000 events during the panel retreat. A MIDAS datalogger was used to record pressure from borehole pressure cells (BPCs) located in two adjacent pillars that were not mined during retreat. A series of photographs were taken of the pillars that had the BPCs as the face approached so that deformation of the entire rib could be monitored using photogrammetry.Results showed that pillar stability and cave development were as expected. The BPCs showed an increase in loading when the face was 115 m (380 ft) inby and a clear onset of the forward abutment at 30 m (100 ft). The photogrammetry results displayed pillar deformation corresponding to the increased loading. The microseismic monitoring results showed the overburden caving inby the face, again as expected. The significance of these results lies in two points, I) we can quantify the safe manner in which this mine is conducting retreating operations, and 2) we can use volumetric technologies (photogrammetry and microseismic) to monitor entire volumes of the mine in addition to the traditional point-location geotechnical measurements (BPCs).INTRODUCTION AND BACKGROUNDExcavation of underground openings is a difficult job in a challenging environment. The rigor of the process is compounded by the lack of a method that allows quantification of changes within the rock mass as excavation progresses. Statistics kept by the Department of Labor, Mine Safety and Health Administration show that 16% of fatalities and lost time incidents in the underground mining industry are due to unexpected rock mass failure (MSHA, 2015). A recent example is the Crandall Canyon, Utah, coal mine accident caused by the vertical collapse of a longwall coal mine on August 6, 2007 (Dreger, Ford, Walter, 2008)."

Jan 1, 2016

By A. Saki, P. Gebski, W. Ying

Industrial smelting furnaces are subjected to significant structural changes and deterioration caused by the conditions under which they operate. The current methods for the structural assessment of furnaces include visual inspections, analysis of thermal data, acousto ultrasonic-echo (AU-E) of the refractory lining and surveying of the key structural elements. Acoustic Emission (AE) non-destructive testing (NDT) technique has proven its capabilities for continuous monitoring of industrial installations and structures. Recently, Hatch Ltd. has been successful in applying acoustic emission for monitoring furnace tapblocks, and expanding the scope of the acoustic emission monitoring to the structural integrity of the entire furnace. The capability of detecting emissions related to the fracture development in the furnace shell, movement of the refractory, electrode arcing, combined with efficient source location algorithms makes an acoustic emission system a comprehensive tool for the real-time furnace integrity monitoring. In this paper the concept of such state of the art system for structural integrity assessment is presented. The initial results of the acoustic emission monitoring together with data analysis techniques are discussed.

Jan 1, 2011

By I Gray

Other than being reasonably certain of the average vertical stress being lithostatic, there can be little certainty about stress in the ground. The stresses and strains in rock are a function of gravitational effects, external load, thermal changes, diagenesis and their mechanical properties including elastic, creep and post failure behaviours. It is highly desirable to measure stress where mining is going to take place, even if the virgin stress condition has already been changed by mining. The concept of some far field stress value that is more or less constant is usually invalid and can be dangerous to use in mine design. Measuring stress is not simple. The basic processes are broken into two categories, one where the rock will not fail around the borehole in which the stress is being measured and one where the rock has failed or will fail. Because of the complexity and cost in obtaining a single, precise stress measurement it is highly desirable to have some system to be able to obtain an idea of the uniformity, or otherwise of the stress situation between precise measurements. Table 1 shows the methods for measuring stress. The first five methods work on the basis of some fairly direct measurement of mechanical rock behaviour while the last three are attempts to get some idea of stress out of core long after it is removed from the host rock. While these latter methods appear to tick all the boxes for applicability there is a great deal of uncertainty in the results that they deliver. This paper reviews the methods for measuring stress and presents two new methods. These are a three dimensional overcore system to provide precise stress measurements in rocks that will not fail around the borehole and a core ovality system. The latter can be used to provide information on the stress difference orthogonal to the core. It is particularly useful because it can be used every metre of core length.

Nov 29, 2022

By F. E. du Plessis, P. De Waal

Through the development of innovative in-line mineral quantification sensors, the possibility has arisen to apply modern control theory to implement automatic process control for electrostatic high tension (“HT”) roll separators. A pilot-scale set-up was constructed to enable continuous tests to be carried out. A typical dry mill feed was continuously pre-heated and separated with a modern lab scale HT roll machine. The mineralogy as well as mass-flow of the non-conductor stream was quantified in-line, and the combined streams were re-circulated in an endless loop by a combination of vibration feeders and a bucket elevator. A fractional factorial experiment was executed to determine the most significant control parameters of the HT roll separator. Then a multi-input, multi-output dynamic model of the HT electrostatic separator was extracted by the application of standard system identification techniques for automatic control. The most important control variable to ensure a stable product grade could be identified. This allowed a robust single-input, single-output controller to be implemented. The process variable and set-point was mineral grade and the control variable was electrode voltage. Roll speed was manually adjusted while the mineral grade was kept on set-point by the control system. The roll speed, where maximum mass flow and therefore maximum yield was produced on-grade, could therefore be rapidly determined. Atmosphere humidity and ore temperature served as disturbances. Some other adjustable parameters were kept constant during the experiments, namely feed mineralogy, splitter settings and feed rate. Delivery of maximum yield, with a controlled grade, can improve production while reducing the burden of complex manual tuning. Manual tuning previously involved coping with multiple parameters with uncertain cross coupling influences on the end result. Even the end result was typically only known after completion of mineral assays. Automatic grade stabilisation and yield optimisation for a HT roll separator bank requires in-line measurement of mineralogy and mass flow, as well as the implementation of a robust process control system.

Jan 1, 2005

By R. Lasmanis

"Washington State's diverse geological setting is reflected by its several types of porphyry Cu-Mo deposits. All are associated with calc-alkaline intrusions, and they range in age from Jurassic to Miocene.The largest is the Mount Tolman Mo-Cu deposit located in the Omineca Belt. It contains 2.18 billion tonnes of material averaging 0.093% MoS2 and 0.09% Cu. There are intermediate-size Mesozoic deposits in the Intermontane Belt. The Cascade volcanic arc hosts small- to medium-Tertiary deposits. Among these, of particular interest is the Sunrise breccia pipe, which has been explored over a vertical distance of 1000 m. It contains reserves of 64.6 million tonnes averaging 0.319% Cu and 0.071% MoS2 with significant tungsten values. The Glacier Peak deposits have potential resources of 1.7 billion tonnes averaging 0.334% Cu and more than 0.015% MoS2 and also contain tungsten. The Margaret deposit can be classified as a porphyry Cu-Au type; it has reserves of 200 million tonnes averaging 0.434% Cu, 0.0175% MoS2> 0.27 g!t Au and 1.92 g!t Ag. Drill intercepts of the North Fork deposit also show significant gold values.Mineralization and hydrothermal alteration of these deposits are typical of porphyry deposits, although some differences can be noted. The Mount Tolman molybdenum deposit contains topaz alteration zones and is transitional between a plutonic type and a Climax-type molybdenum system. Other deposits, such as the Kelsey, Mazama, North Fork, and Margaret, are low-sulphide systems that contain significant amounts of magnetite."

Jan 1, 1995

By I M. van Waveren, F Hasibuan, M J. Crow

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. Two Hg-Au occurrences west of Bangko, Sumatra occur within the West Sumatra Permian Volcanic Plutonic Arc. The Melipun Hg-Au occurrence was exploited in the past and mineralisation occurs within a hydrothermally altered cupola of a buried intrusion. The Salak Hg-Au occurrence occurs within the outcrop of the Dusunbaru pluton, likely within the remnants of a caldera in which hydrothermal mineralisation was developed. Both occurrences formed late in the history of the volcanic arc, probably between the Triassic and early Jurassic. The Melipun occurrence may have potential at depth, but that at Salak is small and lies within the area of a proposed Geopark.CITATION:Crow, M J, van Waveren, I M and Hasibuan, F, 2015. Two Hg-Au occurrences in the west Sumatra Permian volcanic-plutonic arc west of Bangko in Sumatra, Indonesia , in Proceedings PACRIM 2015 Congress, pp 573–578 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By HaiWen Luo, Bin Hu

A ultra-high strength medium Mn steel with good ductility was designed and manufactured via the combination of warm rolling and V alloying, which produced not only the extensive precipitation of fine VC particles in both ferrite and austenite but also the bimodal size distribution of retained austenite (RA) grains. A substantial increase of yield strength up to 650 MPa was achieved due to the addition of V without deteriorating ductility. The coarse RA grains firstly transformed to martensite during yielding at the much higher stress threshold, followed by a significant work hardening due to dislocations multiplication, twinning-induced-plasticity (TWIP) and transformation-induced-plasticity (TRIP) in the ultrafine RA grains. The best combination of strength and ductility included 1.5 GPa ultimate tensile strength and 28% total elongation. This was achieved due to the sustainable TRIP and TWIP effects in the later straining stage resulting from the ultrafine austenite grains, the latter were reversely transformed from the recrystallized ferrite grains.

Oct 1, 2019

By Dan White

Stress measurements taken at Climax, El Teniente, Henderson, Lakeshore and Urad mines indicate that the deposits cited are generally subject to an inclined stress field that can be highly anisotropic. The horizontal stresses vary significantly between deposits and exceed the measured vertical stresses at Henderson and El Teniente mines. Experience at Henderson, Lakeshore and Urad mines is used to suggest a relationship between cavability and the premining stress. In poorly fractured, high strength rock masses, high stresses can assist the caving process. In well fractured, low strength rock masses, caving is ensured by breakage along fracture planes. The relationship between ground weight and the premining stress is also investigated. Ground weight problems appear to be most related to the competence of the ground rather than the magnitude of the premining stress as supported by experience at Climax and El Teniente mines. Recent developments in block caving favor the extraction of poorly fractured, high stressed rock. Accompanying the development of these deposits are new problems such as rock bursts. These pose particular problems to mine designers in that no experience currently exists in combatting rock burst problems at block caving mines.

Jan 1, 1984

By Louis A. Panek

The practice of roof bolting or rock bolting to stabilize rock surrounding underground excavations has increased tremendously within a few years. This widespread application has led to many questions concerning the mechanics of bolting and has pointed up the need for a quantitative method of designing bolting systems. To answer these questions a research program was initiated to determine the principles of roof bolting. The first problem chosen for investigation was the use of bolts to reinforce bedded roof, since the great majority of bolts are used under these conditions. The attack employed is a combination of theoretical and experimental stress analysis. Centrifugal testing of models was chosen as the experimental method because it is particularly suited to the study of stratified roof. This report describes the models, the experimental method, and the model-prototype relations. By using dimensional analysis, a general expression for a roof-bolting design formula in terms of the structural variables has been obtained. Certain model-prototype similarity requirements have been relaxed, thereby minimizing the number of variables to be studied and extending the applicability of quantitative test data to a large class of full size structures.

Jan 1, 1956

By L Neindorf

The development and application of paste fill, cement slurry aggregate fill and shotcrete bulkheads at Mount Isa Mines is described. The progressive changes in filling style in the Mount Isa copper mines from a dry fill based cemented rock fill to tailings based paste fill and slurry fills is explained. The development of shotcrete arched bulkheads and engineered drainage system is described. Details on how Mount Isa Mines has achieved approximately 30 per cent cost reduction in backfill operations through innovations are given. The development of a rocky paste fill is explained.

Jan 1, 2005

By K. Woodward, Y. Potvin, B. McFadyen, M. Grenon

Open stoping has become a popular mining method in hard rock mines, not only due to the safety of the method as a non-entry approach, but also because of the high extraction rate and low costs. At mine sites, stope performance is evaluated by calculating stope overbreak using the stability chart. However, limitations of the stability chart regarding the precision of the predictions, nonconsideration of factors such as the influence of blasting, and the exclusion of underbreak have led to non-optimal designs. The capabilities of today's computers have increased the amount of data being collected and the power of models being built. This article presents a step towards a new stope design approach where stope overbreak and underbreak are measured and georeferenced using octrees at an approximately cubic metre resolution and predicted using multivariate statistical models (partial least square, linear discriminant analysis, and random forest). Results show that overbreak and underbreak location along the design surface and their magnitude are predicted with good precision using a random forest model. These predictions are used to build the expected geometry of the open stope. The resolution of the data and the use of multivariate analysis has enabled the prediction of variation in stope performance along the design surface, going well beyond the simple qualitative per stope face prediction provided by a traditional stability chart approach.

Jun 5, 2023