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By Ismet Canbulat

The cut out distance in the continuous miner sections in South African collieries is limited to 12 m. However, this distance can be extended to 24 m by the permission of the Department of Mineral and Energy (DME). This paper investigates the risk of roof Calls associated with the extended cut out distances permitted by the DME. Extensive underground experiments have been conducted using sonic probe extensometer under various geological conditions to identify the effect of cutting distance and tinge on the stability of the workings and the performance of the support. The results showed that that the most critical parameter with respect to ground control aspects of extended cut out distance is the bord width, which determines the amount of deformation that will take place in the roof. This was confirmed by the numerical modelling. The results also showed that a significant portion of the maximum deformation in the roof takes place before the ratio of cut out distance to bord width exceeds two. This indicates that once the face is extended more than twice the bord width, the roof will stabilise, assuming that no geological discontinuity is present and that there will be no further geometrical changes in the area as a result of for example the development of an intersection.

Jan 1, 2000

By Ken Mills

This paper describes the first successful use of hydraulic fracturing to induce caving events "on demand" in Australia. Moonee Colliery operate a longwall immediately below a thick conglomerate strata. This strata temporarily bridges across the extracted longwall panel to create a large area of standing goal' When this standing goat' eventually collapses, the windblast generated presents a significant hazard to men working on and around the longwall face. Hydraulic Fracturing has been successfully introduced to take control of the timing of these caving events so as to eliminate the risk of windblast injury, The longwall face area is completely evacuated during the treatment. Water is pumped into an injection point located in the conglomerate strata above the standing goat. A horizontal fracture is generated and grows outward from the in point, separating the conglomerate strata below the fracture horizon_ At some point the strata can no longer span and a goat fall is initiated. After a treatment, mining can be recommenced with the windblast hazard eliminated.

Jan 1, 2000

By Jinsheng Chen

To maintain the stability of longwall tailgate entries, they have to be supported by not only the primary roof support but also the secondary roof support due to the impact of the longwall mining¬induced abutment loads. Depending on the mining condition of a typical coal seam and its distance from the previous panel's gob, the longwall tailgate entry is often reinforced by the secondary roof support in the form of standing supports and/or cable bolts or trusses during longwall retreat, Even though lots of research work has been done to study the behavior of different standing supports in the laboratory. their function and requirements for a typical mining condition are still not fully understood due to lack of field data and correct assumptions about the behavior of the already-bolted entry roof under mining-induced abutment loads. Therefore, the application of the secondary roof support for tailgate entry roof control is still limited to trial and error, which often results in either higher cost for tailgate roof control if the design is too conservative or more production delays due to roof falls in the outby tailgate entry if the design is inadequate. To have a better tailgate roof control with reasonable roof control cost, RAG American Coal, NIOSH, and HeiTech have been conducting a joint research project to monitor actual roof loading, roof deformation, and entry convergence in the tailgate entry of 10 North panel at RAG Emerald mine. The actual roof loading data are used to calibrate and improve the roof control model for longwall tailgate entries, which has been developed by RAG American Coal, and to further reduce the roof control cost for longwall tailgate entries. The authors in this study attempt to discuss the Roof Control Model developed by RAG American Coal. in which some guidelines or important parameters for designing the secondary roof support for longwall tailgate entries are proposed.

Jan 1, 2003

By J. Riefenberg

U.S. Bureau of Mines (USBM) researchers are developing a personal computer-based hazard mapping system for use in underground coal mines. Hazard mapping is rapidly gaining interest as delineating areas of possible instability is becoming more complex. The development of a hazard map is based on the consensus that combined effects of local and regional stresses, roof, floor, and rib quality, geology, faulting, presence of ground water, etc., all contribute to stability of the ground. In order to demonstrate the concept, a hazard map was developed that combined stresses anticipated to occur in a multiseam setting using numerical modeling, and the newly developed Coal Mine Roof Rating (CMRR) where roof quality ratings in untested areas were modeled using geostatistical methods. Ten borescope readings from an underground mine were used to compute CMRR's for the geostatistical analysis. CMRR's for the mine ranged from 57 to 77, and are considered in the moderate to strong range. A hazard map was developed by combining the stress effects of interseam mining, scaled to weigh equally with CMRR's, and CMRR geostatistical analysis results. This hazard map demonstrates a synergistic effect between the mining-induced stress fields and the roof quality ratings, indicating, as expected, a greater hazard index in areas of high stress and lesser roof quality.

Jan 1, 1994

By Bruce K. Hebblewhite

Underground pillar extraction continues to be practiced in Australia in various forms, in spite of the dominance of the longwall method, which accounts for over 90% of underground production. Australia has been a leader in various forms of pillar extraction, particularly since mechanisation was introduced to the industry during the middle of the 20th century. Historical developments have included a range of thick seam mining pillar extraction methods; the widely adopted Wongawilli method; early adoption of the use of mechanized breaker line supports; and a number of more recent partial extraction systems including what is known as rib stripping. Current pillar extraction practice varies from total extraction panels to various forms of partial extraction, driven by either surface subsidence constraints or by underground conditions. This paper reviews a number of different pillar extraction methods and mining layouts, and also highlighting a number of fundamental geotechnical principles which have been found to be critical for safe pillar extraction ? often through lessons learned the hard way, as a result of back analysis of different accident scenarios. These principles include the all-important understanding of the location of the goaf (gob) edge at any point in the operation, and the management of that goaf edge; the impact of geological structures on local and regional stability; the critical importance of extraction lifting sequences; the role of stooks (remnant pillars); and a number of other factors. The paper will also briefly reference the educational and training practices adopted in Australia aimed at the continuous improvement of pillar extraction safety.

Jan 1, 2011

By Klaus Opolony

The German coal industry operates a multi-seam extraction system, with average working depths currently around 1,000 meters. The gate roads serving the longwalls are often used for a second time - a practice which has the advantage of allowing several panels to be extracted with a minimum of roadway drivage work. This preference for re-using gate roads is based on a number of factors: such an arrangement a) prevents the strata pressures in the subsequent panels from being adversely affected by the presence of support pillars, b) helps achieve the target of full mineral extraction and c) allows the roadways to be used for air conditioning and ventilation purposes as part of an advancing longwall system. This paper explains the planning procedure required for a rectangular rockbolted roadway, using as an example a gate road in the 3 meter-thick D2/C seam at Lippe colliery, and describes the subsequent operating experience. Here special attention is focused on modem procedures and techniques for mine roadway planning. The investigation also examines the part played by the mine ventilation system and the available machine technology, which are particularly important parameters for the design of the T-junction areas and support system. Keywords: multiple seam extraction, longwall mining, roadway drivage, roof bolting, roadside pack system, hydraulic frame type support.

Jan 1, 2002

This paper describes the experiences with the application of the Coal Mining Roof Rating (CMRR) System to South African strata conditions. Roof classification studies using the CMRR system were undertaken as part of a SIMRAC investigation into falls of ground in South Mean collieries. Colliery roof rating exercises were undertaking in a range of geotechnical conditions from shallow level coal seams under flat topography to workings with cover depths of greater than 300m A comparison is made between the CMRR and other Rock Mass Rating systems used to classify different roof geotechnical conditions. The need for geological discontinuity orientation, blasting and rockmass stress adjustments to the CMRR is discussed. The effects of acid mine water on strata weatherability and the need to incorporate immersion tests using acid mine water are also detailed. In addition limited correlation is given between CMRR values and other geotechnical factors such as joint trace length, joint shear strength, joint orientation, weatherability and roof damage due to horizontal stress. Taking cognizance of these observations, a concept of stress induced roof damage triggering is discussed.

Jan 1, 2001

By Adam Mackenzie, Dan Payne, Yass-Marie Rutty

"Historically, the Crinum mine has experienced significant falls of ground when longwall production slowed to prepare for recovery in weak roof areas. These conditions continue through the recovery process and result in both safety concerns and delays. When mine plans and exploration revealed that most of the mine's future longwall recoveries were located in weak roof areas, a decision was made to try pre-driven recovery roads as a solution to the problem. After completing eight pre-driven recovery roads with varying degrees of success and numerous lessons, the Crinum North mine now utilizes a modified Pre-driven Recovery Roadway (PDRR) to improve the longwall take-off process in weak roof areas. During mine development, a standard roadway is driven where the final recovery location of each longwall is planned. After the installation of secondary support, the PDRR is backfilled with a cement-flyash mix to provide support to the roof, and confinement to the ribs and floor of the roadway. The method has been refined over the last four years to provide greater strata stability and improved operational and safety performance compared with conventional take-offs at Crinum, and has resulted in a site record of longwall relocation in 11.5 days (pull mesh to picks in coal). This paper describes the evolution of the PDRRs from Crinum East to Crinum North, including lessons from initial attempts and changes to the secondary support regime, the operational approach during the final stages of retreat, the backfill strategy and plans for the future.INTRODUCTIONThe Crinum mine is the underground component of BHP Billiton Mitsubishi Alliance's (BMA) Gregory Crinum mine, located northeast of Emerald, Queensland (Figure 1).The Crinum mine has a history of roof control problems coming into and during longwall recoveries. Weak roof, combined with the slow mining, which is a consequence of preparing for equipment recovery, pulling mesh (which also creates tip-to-face issues), and the lengthy bolt-up process, often results in loss of immediate roof and subsequent roof falls. In fact, approximately 50% of the first 13 longwall recoveries (which were at depths of 180 to 200m) experienced roof fall delays, the longest of which took four months to mine the last pillar and recover the longwall. This recurring problem prompted the mine to investigate options to reduce or eliminate the problem."

Jan 1, 2016

By Neil Styler

This paper presents an analysis of measurements of inter-burden de- formations above six longwall faces. An attempt is made to demonstrate some correlation between the movements at the various sites, and to examine their importance with respect to predicting caving height, disruption of overlying seams, and disruption of aquifers. This analysis demonstrates some significant differences between predicted surface subsidence, and inter-burden deformation. In addition it is shown that the caving height above a longwall face is equal to 8 to 12 times the extraction height, with a zone of fractured rock ex- tending to approximately 50 times the extraction height above the seam.

Jan 1, 1984

By Kanaan Hanna

The integrity factor approach was developed by the Bureau of nines to assess the stability of mine pillars and has been applied primarily to longwall chain pillars. Recently this approach was applied to a room-and-pillar mine in the Illinois Coal Basin. A special configuration of hydraulic borehole pressure cells was installed in a pillar to determine the in situ strength-and- stress relationship of the coal pillar and to verify the applicability of the integrity factor approach. Results indicate that the integrity factor approach is applicable to the assessment of pillar st6b:lity and nay bs a realistic and effective approach for room-and-pillar mine design.

Jan 1, 1988

By Gregory E. Smith

"Tensionable headed roof bolt systems are common in the U.S. mining industry. In the last 15 years or so, they have made a comeback as a hybrid system where a mechanical anchor is used in conjunction with resin. When such a system is used, it is common to use a spherical washer to comply with the requirements of 30 CFR Part 75.204(f)(8). This standard requires an angle-compensating device be used in installations where the bearing plate is offset from the roof bolt axis by an angle greater than five degrees. This paper evaluates system performance of four different traditional angle-compensating device/plate combinations including a new patented extruded plate/extruded washer system that improves overall load capacity and stretch.INTRODUCTIONA large western U.S. underground coal mine was using a fully grouted headed rebar system as their primary support. They had developed their component specifications over many years of working with suppliers to improve and optimize these components to integrate well as a system. As shown in Figure 1, the components included a patented 7 /8"" -diameter Grade 60 headed rebar bolt with a flange capable of full bar load (Walker, 2009) and a patented 8 in x 8 in Grade 3 extruded hole bearing plate manufactured by DieMatic Corporation.The mine decided to pursue a fully grouted tensionable system using a mechanical anchor as the tensioning device. Their primary goal was to eliminate the need for the roof bolt operator to secure (hold) the bolt during resin set time. An additional benefit was the increased initial tension on the bolt. To comply with the requirements of 30 CFR Part 75.204(f)(8), regarding tensioned roof bolts, they were required to include an angle-compensating device because a share of the bolt population would be installed at angles greater than five degrees from the perpendicular to the bearing plate. Also, to improve cycle time and torque/tension consistency, the mine desired to have the washers and anchors assembled at the factory. This required the use of a bearing plate with a center hole large enough to accommodate the 1-3/8-in-diameter mechanical anchor to pass though it during assembly by the roof bolt operator. The mine wanted to continue to use an extruded hole bearing plate in the tensionable system, and EMS was asked to propose a plate and angle-compensating device."

Jan 1, 2016

By Lisa Burke

In underground coal mines, concrete block stoppings are widely used to control mine ventilation. Although stoppings are not intended as roof support, they are subjected to roof to floor convergence, and may fail as a result of this loading. Premature failure of stoppings can significantly affect mine ventilation, limiting the amount of fresh air reaching the working faces and increasing the risk for methane explosions. Softening materials are often used in stoppings to reduce the damaging effects of roof to floor convergence. However, to date, research has not focused on the behavior of stopping construction materials subjected to roof or floor loading, leaving the design process to trial and error. A combination of numerical simulations and large scale physical tests were employed to develop a scientific understanding of stopping performance. The first series of physical models were constructed using standard CMU (concrete masonry unit) blocks and tested in a load frame. The behavior of these walls was simulated with a distinct element model. A key finding was that very small variations in block size and irregularities in block shape significantly affect wall performance. These variations, which are inherent to the type of blocks typically used underground, create stress concentrations that initiate the failure of stopping walls. A second series of models employed concrete stoppings that incorporated woos planks, a shredded wood material, and foam planks like those commonly used underground. Numerical models were again able to match the physical test data. Analysis of the softening layers used in stopping construction indicated that the strength and stiffness of the material relative to that of the wall is crucial in determining the amount of roof to floor convergence the wall can withstand prior to failure. The product of this study is a numerical model that can be used to evaluate the performance of stopping materials and different wall geometries in a controlled environment. Testing was done in a laboratory setting with walls constructed on flat, parallel surfaces and subjected to uniform loading at a constant rate. Parametric studies with the model indicate that in this ideal environment it is possible to control the amount of vertical displacement the stopping can withstand by adjusting the number of softening layers built into the wall and the thickness of the layers. Additionally, under these conditions, the position of softening layers within the wall appears to have little effect on the amount of displacement walls can withstand prior to failure.

Jan 1, 2004

Previously, a variety of risk rating systems were used on different mines. However, these were mostly site specific and geology based systems. The purpose of developing a uniform rating system, to be used on all Anglo Coal open cast mines, is to provide an unbiased, standard and quantifiable assessment of the risk from highwall and lowwall failures. This takes into account relative differences in the importance of hazards, as experienced on each mine as a result of different combinations of geotechnical factors and mining conditions. The system is based on critical geotechnical and other parameters that have been identified by mining personnel and rock engineers. The primary advantage of this risk rating system is that all open cast mines in Anglo Coal use an almost identical system, which enables the user to compare their risk to the risk at neighbouring mines and/or other open cast mines. The system can be adjusted to meet local, mine specific requirements. The implementation of this new system (a computer program that automatically calculates the risk) has been made as practical and easy to use as possible. Therefore, this tool can be used by personnel from other mining disciplines not directly related to rock engineering.

Jan 1, 2004

By Michael Gauna

Ground fall injuries and fatalities in United States underground coal mines have shown a diminishing trend. Advances in bolting and standing support systems have led to improved safety in underground coal mines. However, the Crandall Canyon Mine disaster in 2007, in which 9 individuals were fatally injured as a result of a widespread collapse, emphasized that proper mine design is a critical issue that must be addressed. The Mine Safety and Health Administration (MSHA) has instituted technical review procedures structured to facilitate the implementation of safe mine designs in underground coal mines. The technical review process and the typical information required are outlined in this paper. The range of mining situations that lead to a technical review, common design methods submitted for review, and their relevant guidelines and limitations are also discussed. Additionally, the relationship between design criteria and mine site historical experience is discussed.

Jan 1, 2011

By Keith A. Heasley, Jian Yang

"LaModel uses a laminated overburden boundary-element model and can calculate not only seam-level stresses and displacements but also surface subsidence for thin tabular deposit such as coal seams (Heasley, 1998). Up to this point in time, the material property wizards in LaModel have primarily been designed for calculating accurate stress distributions in single- and multipleseam situations, and for investigating and optimizing pillar sizes and layouts in relation to overburden, abutment, and multipleseam stresses. However, the critical input parameters that will give the most accurate seam-level stress distribution do not necessarily produce the best surface subsidence prediction. The objective of this paper is to develop a methodology for calibrating the critical input parameters in LaModel to optimize surface subsidence prediction.For optimum surface subsidence prediction, it was found that the gob convergence (as defined by the final gob modulus) and the overburden stiffness (as defined by the lamination thickness) were the two most critical parameters that needed to be calibrated (Heasley, 2016). By calibrating numerous panels with various widths and depths, a formula that provides the optimum final gob modulus to use for subsidence prediction was developed. Also, three different empirical formulas for optimizing the lamination thickness as a function of overburden depth and/or the panel width to- depth ratio were determined for the three cases: supercritical panels, and subcritical panels with and without offsets. Further, if the user has measured data for subsidence factor and angle-of draw, the optimum final gob modulus and lamination thickness can be determined from the measured data. These new subsidence prediction formulas are being implemented into new material wizards in LaModel.INTRODUCTIONIn the past few decades, with the increased need for highly productive mining techniques, full extraction mining methods have been increasingly employed in the U.S. coal mining industry. These total extraction mining methods, which include both longwall mining and room-and-pillar retreat mining, normally cause immediate roof caving and the associated surface subsidence. These methods can be highly productive and cost-effective, but the associated surface subsidence can cause damage to surface structures and negatively impact the surface environment (Peng, 2008). Therefore, a reliable program for predicting the surface subsidence associated with full extraction mining is required for appropriately engineering and managing the surface subsidence impacts of full extraction mining."

Jan 1, 2016

By Andrew J. Hyett

The ISRM commission on testing methods has recently presented Suggested Methods for rock stress determination (Kim & Franklin, 1987). This proposes that even if the magnitude of in situ stress is not sufficient to cause significant ground problems, the optimum shape, orientation, and layout of underground structures can be significantly affected by it. In the design of underground excavations, the engineer often requires a knowledge of the average in situ stress tensor in the rock mass. However, attempts to measure a stress state applicable to design have proved frustrating, and a high degree of uncertainty is introduced by the variability in results. In the belief that deduced fluctuations are random in nature, many separate measurements are often made, so that a more reliable stress tensor average, and some indication of the associated variance (see Dyke et al., 1986) can be calculated. Unfortunately, this approach is limited by economic considerations and, as with any sampling procedure, it is imperative to understand the sources of measured variability, in order to design an optimum measurement programme. Cundall(1987) itemised the main causes of stress variability as a. Instrument and measurement errors, and b. Natural fluctuations of stress from point to point, induced by the action of a boundary stress on the internal rock structure. At least four categories exist: i. inhomogeneous rock, bedding, intrusions etc.; ii. differential contraction, creep, or changes in rock properties with time; iii. proximity to faults or discontinuities; iv. cycles of tectonic activity that cause movement on joints. The latter was successfully investigated using the distinct element programme UDEC. In this paper, the contribution of items (iii) and (iv) (i.e. the items related to discontinuities) to the total stress variability will be discussed, for the kind of rock mass shown in Figure 1. Although many authors have concluded that rock joints have an influence on the stress in their vicinity, a more analytical approach is required if the stress distribution associated with different jointed rock masses is to be (i) compared, and (ii) related to routinely measured joint parameters. If this can be achieved then, prior to any attempt to measure stress in a jointed rock mass, a consideration of joint geometry, joint deformability, and the geological kinematics of joints (e.g. the magnitude and direction of previous movements), can place valuable `common sense' constraints on the expected stress system for minimal extra cost.

Jan 1, 1989

By Nielen van der Merwe

The data base of failed pillar cases as used by Salamon and Munro (1967) in their original analysis to determine the strength of coal pillars empirically, has been updated for this study. It is shown that the Vaal Basin and Klip River coal fields exhibit similar behaviour, i.e. failure at high safety factors within a short period of time after their creation. Those collapse cases have been omitted from the new data base. New failures that occurred after 1967 have been added, except for those that occurred in the Vaal Basin and Klip River coal fields. (1967) a technique to minimise the area of overlap between the populations of failed and intact pillar cases, the new data base was re-analysed in conjunction with the original Salamon and Munro (1967) data base of intact pillar cases. It was found that a linear formula reduced the overlap area by 22%. The new formula predicts higher strength for pillars with a width-to-height ratio greater than approximately 2S and a lower strength for smaller pillars. For most current South African coal mining conditions, using the new formula for pillar strength will result in leaving smaller pillars without sacrificing stability. The reason for this is that the Salamon and Munro (1967) formula underestimated the strength of larger pillars and overestimated the strength of smaller pillars. The potential benefit for the South African coal mining industry amounts to saving reserves equal to the total production of two large mines per year.

Jan 1, 2002

By Phillip E. Gramlich

Through extensive practical experience and multiple studies, tensioned roof control systems have proven beneficial in difficult mining conditions. Due to increased mining of less desirable seams, tensioned roof control systems have claimed a progressively larger share of the total number of roof bolts installed in the US coal industry. Each tensioned roof control system has its benefits, but also has problems concerning cost, cycle time, stress concentration, load retention, and protrusion into the mine opening. For example, resin-assisted mechanically anchored bolts (also known as resin point anchor bolts) have a fast cycle time and the potential for high tension loading, but the stress from the mechanical shell is concentrated in a small area and can create concerns with certain roof structures (Zhang and Peng, 2001). Resin point anchor bolts can be difficult to fully grout. The cost of resin point anchor bolts can be an issue; a two-piece resin point anchor bolt has a large number of parts (10) and manufacturing steps (9), adding to its higher cost. Tension rebar bolts can achieve high loads and are easier to fully grout but have longer cycle times. In addition, a high percentage of the installed load can ?bleed off? if the resin mixing time or the ?hold time? before tensioning is short-circuited. Tension rebar bolts can be notorious for protruding into the mine opening, creating an obstruction for man and machine (Figure 1), in particular, with reduced mining heights. A new tensioned roof control system has been developed that doesn?t have moving parts, is cost effective, provides fast cycle times, has a low profile, and is simple t install.

Jan 1, 2014

By Bruce K. Hebblewhite

BHP Billiton Illawarra Coal operates Dendrobium Mine in an area 10-20km west-northwest of Wollongong in the Southern Coalfield of New South Wales, Australia. The mine recently completed longwall mining in Area 3A adjacent to an overhanging rock feature known as Sandy Creek Waterfall, which had been identified as a significant, and potentially sensitive natural surface feature of the environment. This waterfall lies within the Sydney drinking water catchment area, in a surface terrain of variable topography where non-conventional subsidence behaviour such as valley closure, and far-field horizontal movements are known to occur, over and above conventional subsidence responses to mining. As a part of the conditions of approval to mine, Illawarra Coal undertook to protect the waterfall and the section of Sandy Creek immediately upstream of the waterfall from the effects of longwall mining using an innovative management process and an array of very high resolution monitoring systems. This paper outlines the type of non-conventional subsidence effects that can occur; the range of different monitoring regimes adopted in response to such effects in order to develop an understanding of the valley closure phenomenon in proximity to the waterfall and provide suitable early warning data. The paper also describes the management processes that were adopted, incorporating the use of these monitoring regimes in order to successfully protect the waterfall from the valley closure effects of mining four adjacent longwall panels in close proximity to the waterfall while continuing to maximise recovery of the coal resource in the area. The management structure adopted involved a technical committee, a steering committee, and an external independent reviewer. The technical committee was responsible for design of the monitoring systems, interpretation of the results, and making timely recommendations to the company steering committee who were then able to make informed decisions on when to cease mining each adjacent panel. The steering committee comprised Illawarra Coal management and technical personnel. Although the steering committee took advice from the technical committee, all decisions relating to mining were made by the steering committee. An external reviewer was engaged by the steering committee at the end of each longwall panel to review the results, interpretation and management decisions. This management structure was effective in successfully protecting the very sensitive structure of Sandy Creek Waterfall from potential impacts of nearby mining in the midst of active ongoing natural erosion processes.

Jan 1, 2014

Since the initial development and trials of FLEXIBOLT flexible roof bolts, further comparative testing and product development has been completed as part of the changeover to this roof support system at two Australian longwall operations, Ellalong and Angus Place Collieries. Details of the practical implementation of FLEXIBOLTS at both collieries are outlined together with their impact on roadway development rate productivity. At Ellalong where fully encapsulated 2.1 m long bolts are used, the reduction in bolt density is discussed with respect to both support system design and work force acceptance. The impact of shear deformation on bolt loading is described. Bolts longer than the working height have been required at Angus Place to achieve a suitable bolt anchorage horizon. Details of the long hole drilling and the benefits of a two stage grouting procedure are outlined. Improvements in roof conditions which have resulted from the inclusion of FLEXIBOLTS into the overall roof support strategy at Angus Place are described. Their importance to the viability of the Angus Place operation is also highlighted.

Jan 1, 1994