Classical rock support methods for underground engineering emphasize the importance of investigating and controlling the unloading deformation of surrounding rock, such as in the New Austrian Tunnelling Method (NATM) and Norwegian Method of Tunnelling (NMT). Current engineering practice in underground tunnels and caverns under high geostress conditions indicates that stress-induced collapse and rock bursts are the major engineering hazards. Prior to this kind of failure, crack initiation and propagation occurs. Thus, a new support concept, termed the crack-restraining method, has been developed based on the NATM and NMT for underground hard-rock engineering under high geostress conditions. This method includes three key technical points: The length of the rockbolts is slightly greater than the maximum depth of cracking; the optimal supporting time of the rockbolts is decided based on the crack evolution and deformation tendency; shotcrete with high tensile strength is used to ameliorate the stress state of the rock mass' surface. The aim of these supporting efforts is to restrain the subsequent cracking and enhance the stress-bearing capability of the rock mass. Application of this approach to the Baihetan underground cavern (the largest hydraulic cavern in China) and other deep underground tunnels indicated that this crack-restraining support method is appropriate for determining the rockbolt length and supporting time. INTRODUCTION Increasing attention is being given to deep-level excavation and mining due to the gradual exhaustion of shallow mineral resources and recent discoveries of deep mineral resources. Most South Africa gold mines have reached depths of more than 2000m, and the mining depth at Mponeng almost 4350m (Kwiatek, Plenkers, and Dresen, 2011); the Creighton mine in Canada has reached 2400m in depth (Snelling, Godin, and McKinnon, 2013); the Kolar gold mine has progressed to 2400m (Mishra and Panigrahi, 1999). Greater depths are also being attained in civil engineering projects–some traffic tunnels and hydraulic tunnels have been excavated at depths greater than 1000m (Zangerl et al ., 2008, 2010, 2014). All these engineering projects have shown that the unloading behaviour of hard rock under high geostress is very complex. The support scheme for hard rock under the conditions encountered in these projects obviously presents a great technical challenge.

Letšeng Diamond Mine (LDM) is the highest open-pit diamond mine in the world located at 3 100 masl. Two kimberlite pipes, the Main Pipe and the Satellite Pipe, are currently being excavated and are called the Main Pit and Satellite Pit respectively. Two failures of approximately 3 000 and 2 000 t of material occurred in August and September 2016 respectively in the southern portion of the Main Pit at Cut 3 West (MC3W) along a major shear zone that intersects the southern edge of both pits. An investigation was undertaken to determine the causes of failure in order to mitigate and prevent future failure occurrence. The investigation involved a detailed area inspection, geological mapping, rock mass characterisation, discontinuity survey, kinematic analysis and a back-analysis. The field inspection revealed that the shear zone is composed of sheared and friable kimberlite, calcite, pyrite and altered basalt. Rock mass classification confirmed that the shear zone is made up of a weak rock mass with low rock mass rating. It is evident from geological mapping and discontinuity survey analysis that the shear zone consists of a combination of faulting, kimberlite dykes and closely spaced sympathetic joints which are liable to toppling failure. Stereonet based kinematic analysis using Rocscience Inc. DIPS shows only one cluster, which has an average dip and dip direction of 85°/175°. The mode of failure was a combination of toppling failure on the major planes coupled with unravelling of friable material. Back-analysis using RocTopple suggests that failure may have been triggered by groundwater conditions as a result of infiltration of rain water into the relative weak rock mass. INTRODUCTION The stability of a rock slope is governed by the local geological and geotechnical characteristics of the rock mass, the hydrogeological, geomorphological conditions, and other environmental factors such as freeze and thaw cycles (Wyllie and Mah, 2004; Verma et al. , 2011). Structural geology plays a major role in the design and maintenance of an open pit mine as the presence of structural defects affects the mechanical and hydrogeological properties of the rock mass (Bye and Bell, 2001). Therefore, the assessment of geological structures forms an essential part of rock slope stability analysis and slope management.

Crush pillars have been extensively applied on the Merensky Reef horizon since the late 1970's. Once in a crushed state, the residual strength of the pillar provides a local support function and must support the hanging-wall to the height of the highest known parting. The design process of crush pillars is mainly limited to specifying a width to height ratio (w:h) of approximately 2:1. It is also required that a pillar crushes close to the face, while the pillar is being formed. On many mines the crush pillar system is problematic owing to the difficulty of controlling pillar sizes. This is mainly caused by poor drilling and blasting practices. As a result, pillar crushing is not always achieved. Crush pillars are implemented at relatively shallow depth, the pillar dimensions have remained essentially unchanged over many years, and the impact of regional pillars and geological losses contributing to the regional behaviour of the rock mass are overlooked. In many cases the pillar system is the source of seismicity. In this paper, the influence of mining losses (potholes) and the use of sidings are discussed as possible contributors impacting on crush pillar behaviour. A limit equilibrium model implemented in a displacement discontinuity boundary element program is used to demonstrate crush pillar behaviour. The results are compared to the pillar behaviour at an underground investigation site, which supports the preliminary findings. INTRODUCTION Crush pillars have been extensively used in Merensky Reef stopes. The key function of the pillar system is to prevent the occurrence of back-breaks (large-scale collapses in the back area of a stope) occurring as a result of hangingwall separation along parting planes or fractures. One such problematic parting is the Bastard Merensky Reef, situated between 5–45 m above the Merensky Reef. The pillar dimensions are selected such that the pillars should be fractured while being formed at the mining face. This is typically achieved when a pillar is cut at a width to height (w:h) ratio of approximately 2:1 (Ryder and Jager, 2002). Once crushed, the residual strength of the pillar provides the required support function. Factors influencing pillar stress ( i.e . mining depth, pillar width, mining height, percentage extraction) will impact on crush pillar behaviour. Du Plessis and Malan (2015) demonstrated how oversized pillars could potentially result in unpredictable pillar behaviour. The execution of a mining layout can impact on the size of the pillar cut at the mining face. This is demonstrated by the examples and case study presented in this paper. Similarly, the presence of potholes or blocks of unmined ground will influence pillar crushing. While these are common occurrences on most mines using crush pillars, these factors have historically not been associated with poor pillar crushing or pillar seismicity.

Mining in the Carletonville area of the Witwatersrand Basin predates 1934. Owing to the depletion of higher grade ore, the current activities focus increasingly on the extraction of lower grade secondary reefs as well as remnant and pillar extraction. Of particular interest is multi-reef remnant extraction. Numerical modelling was conducted to investigate the rock engineering aspects of remnants being understoped on a secondary reef horizon. An analysis of the stress evolution in the middling between the two reef horizons indicated that a zone of high major and low minor principal stress develops between the two reefs. This indicates a high risk of violent shear failure. Some pillars were nevertheless successfully understoped in the past, and a study was conducted to understand this phenomenon. An ‘extended’ energy release rate concept introduced by Napier and Malan1 proved to be useful for investigating this problem. It was found that bedding planes and lithology play an important part in the stable dissipation of energy in multi-reef remnant geometries. The study indicated that stope convergence and the various energy components are affected by the presence, position, and properties of a bedding plane. The energy solutions are complex and sometimes counterintuitive, and. care should be exercised when modelling specific cases. The modelling was nevertheless valuable for indicating that energy dissipated on weak layers may reduce the risk of violent failure in a multi-reef mining scenario. This was confirmed by an actual case study. INTRODUCTION To extend the lives of the Witwatersrand gold mines, work is being conducted to include a larger number of remnants and pillars in the reserve base of the mines. According to Jager and Ryder (1999), a remnant is an unmined block of ground surrounded by extensive mining. Remnants are typically left because of adverse or difficult mining conditions caused by geological structures, poor ground conditions, low grade, or rockburst damage (Durrheim, 1997). In contrast, pillars are typically the result of rock engineering designs to control energy release rate (ERR) and seismicity. During the early 1900s, remnants formed by the practice of scattered mining layouts had a significant effect on seismicity. Numerous task teams were appointed during this period. As the industry expanded, the severity of rockbursting increased and the research was formalized after the 1960s. Various recommendations were made and many of these are still used as best practice today ( e.g . Jager and Ryder, 1999).

Shallow hard-rock bord-and-pillar mine designs are typically characterized by solid pillar systems and stable span designs. These spans are optimized to maintain stability, which permits safe and economic extraction of the orebody. Reef-parallel structures are known to affect the strength of pillars. In areas where such structures exist, careful considerations need to be given to rock mass properties with respect to pillar design, bord span and type, and support design, depending on the placement of the structure. Part of Unki mine has a reef-parallel structure referred to as the Footwall Fault (FwF), which is located either within or below the mining cut. In addition to the FwF, mining in the upper sections of the mine encountered poor ground conditions, characterized by low-angle joints trending N-S and truncated by E-W high-angle faults, joints, and dykes. These features resulted in frequent falls of ground, which were attributed to incorrect span design, inadequate support design, pillar fracturing, and sidewall failures. This paper presents the bord span review and the corresponding support optimization process, as well as the resultant improvements and benefits to the mining process as applied to this top section of the mine. The results presented here have led to successful mining through the challenging ground conditions. The same conditions had earlier resulted in the closure of one of the uppermost sections of the mine. Crucial to this research is the inclusion of rock mass properties in the bord span re-design, modifications to the design formulae to suit the downrated conditions, and probabilistic numerical modelling for support design. INTRODUCTION Mining in the upper portions of the Unki East shaft has been faced with ground stability challenges, leading to decommissioning of one of the half-levels (1 South). These upper half-levels contain a reefparallel structure referred to as the Footwall Fault (FwF) which is located either within or below the mining cut. Reef-parallel structures are known to affect pillar strength and have been cited as a major contributor to mine collapses. Bimha mine in Zimbabwe and Everest mine in South Africa are typical examples. A combination of the splay faults from the FwF, N-S low-angle joints intercalated by E-W high angle faults, dykes, and sympathetic joints resulted in frequent large falls of ground, which were ascribed to inadequacy of the span and support designs for the prevailing poor ground conditions, which were not envisaged at the mine design phase. Work presented in this paper has been carried out to mitigate the impact of this combination of geological structures and avoid prematurely decommissioning half-levels. The work was conducted when 1 North had intersected poor ground, entailing a high possibility of half-level decommissioning.

A mining operation transitioning from open pit to underground in an area where the mine design does not allow caving through to surface normally requires a crown pillar for protection of the underground workings beneath the pit. Crown pillar stability is maintained to prevent inrushes of water and mud, air blasts, and collapse or large-scale movement of the pit walls or surrounding rock mass. This is important in providing a safe workplace for mine personnel, equipment, and infrastructure, and ensuring achievement of production goals. The Paboase underground mine at Chirano Gold Mines Limited in Ghana is situated beneath an existing and non-operating open pit. A crown pillar is required to separate the pit from the underground mine, where sublevel open stoping method was used to extract the upper mining block. During early studies, the crown pillar was designed based on preliminary data available at that time. With additional geotechnical data available from field mapping, core drilling, and laboratory testing, efforts were made to evaluate the stability of the crown pillar. Empirical and analytical methods were applied to determine the factor of safety for the crown pillars of various thicknesses under different stoping geometries. Numerical modelling was also utilized to evaluate the crown pillar performance under various in-situ stress conditions and mining sequences. In addition, stope stability in the upper mining block was analysed using the updated Matthews' stability graph. This paper presents the mine design, input data, and performance evaluation of the Paboase crown pillar and summarizes the results from the empirical, analytical, and numerical modelling analyses. INTRODUCTION When mining transitions from open pit to underground, it is usually necessary to leave a crown pillar for the underground mine beneath the bottom level of the pit. It is vital that the crown pillar remains stable throughout its planned service life to protect surface land users, and the underground mine and workers from inflows of water, and failures of soil and rock (Carter, 1989). The design of a stable and cost-effective crown pillar is one of the most challenging practical problems faced by mining engineers today. Surface subsidence, to a greater or lesser degree, is an inevitable consequence faced by many underground mining operations. The subsidence resulting from a crown pillar failure has always been a concern for economic and safety reasons.

In order to improve the extraction ratio at a hard-rock mine, the partial post-extraction of strike stabilizing pillars was considered as an option at a concept level of study. An evaluation of the potential revenue based on a maximum extraction of up to 50% of the pillars was carried out, which suggested that the project would indeed be a viable way to increase the total potential revenue. This necessitated a study on the stability of the underground excavations during the pillar extraction process based on a conceptual design. The proposed mining layout and extraction sequence was simulated to determine whether this could be included in the mining plan, and if so, to determine the potential additional extraction that could be incorporated into the mining plan. INTRODUCTION A hard-rock mine in the Bushveld Complex in the Northwest Province of South Africa commissioned a study to review the potential for extraction of the regional stabilizing pillars, the function of which is to support the overburden, as well as to act as barriers (Zipf, 2001). The function of the barrier pillars is primarily to compartmentalize the mine. If a pillar collapse occurs, the barrier pillars need to prevent this collapse from spreading over a very large area. The Bushveld Complex is a basin-like rock mass consisting of layers of igneous rock. The main rock types comprise norite, anorthosite, and pyroxenite. The upper Critical Zone of the Bushveld Complex hosts the platinum-bearing orebodies referred to as the Merensky Reef and the UG2 Reef. Although the Merensky Reef is generally regarded as uniform, large variations occur in reef thickness and orientation (rolling reef). Potholes (steep troughs within the stratigraphy that result in significant thinning of the reef) are generally common throughout the eastern and western limbs of the Complex. A conventional mining method is to be employed for stoping, i. e. drilling and blasting of stope panels, separated by crush pillars. Cleaning of stopes will be done by winches and scrapers, which will feed directly onto a conveyor belt in a permanent strike drive. The mining height was anticipated to range between approximately 1 m and 1.5 m.

When shield tunnel boring machines (TBMs) are used, the outer lining is constructed using pre-fabricated lining segments. In hard rock conditions the required gap between the lining and excavation boundary is backfilled with pea gravel. In this case, designers consider only the backfilled material relevant for the bedding effectiveness. Thus, a sound knowledge of the deformation properties of pea gravel is essential. At present these parameters are determined in non-standardized laboratory or on-site tests. Moreover, no validation of the used parameters is performed in situ . Based on the static load plate test a prototype for the in-situ determination of the deformation properties of pea gravel was designed and manufactured at the Institute of Rock Mechanics and Tunnelling at the Graz University of Technology. The testing device allows for a maximum test stress of 1.50 MPa under the load plate. This research focuses on the development process and the detailed design of the manufactured prototype. The recorded measurement data (force and displacement) allows to calculate the stiffness properties ( i.e. bedding modulus, static deformation modulus, stiffness modulus) of the material tested. Radial openings in the lining segments serve as measurement locations. Mounting of the testing device in the tunnel is possible without any alterations in the lining elements or the lining design. Determination of the stiffness parameters is shown using the test data of one exemplary test series. INTRODUCTION In mechanized driven tunnels, using single or double shield TBMs (S-TBM/DS-TBM), prefabricated lining segments are used as outer lining. Thereby, the load-carrying rings are assembled within the protection of the shield. Due to the construction method, a gap between lining and excavation boundary (i.e. rock mass) exists after a newly installed ring leaves the shield. This so-called ‘annular gap’ needs to be backfilled using a suitable material in order to allow for load transfer. Ideally, the occurring rock loads are evenly distributed. In hard rock conditions, a fine grained and closely graded gravel, termed as pea gravel, is pneumatically injected through radial openings in the lining. The invert area is backfilled using a pea gravel-mortar-mixture.

Understanding the evolution of micro-cracking induced during rock indentation by roller tricone drilling tools is a key aspect of maximizing drilling efficiency. Crack network patterns and interaction between neighbouring cracked zones induced by two adjacent tool bits are a challenge for drilling engineers. In this paper, a methodology based on the cavity expansion model is developed to analyse the indentation of rocks by a blunt spherical indenter. Rock elasto-plasticity behaviour is used to propose an analytical solution for the plastic zone radius and a description of stresses within and outside the plastic zone. A Maple program was written to derive the expressions of plastic zone radius and stresses when adopting either the Mohr-Coulomb criterion or the Hoek- Brown criterion. The two approaches were then compared. All analyses were conducted on two hard rock types: granite and limestone. The first is encountered in geothermal reservoirs and the second in petroleum drilling. The results were also compared with those obtained using an experimental procedure developed in a previous investigation, in which the cracked zone radius is estimated using image analysis to distinguish the indentation microcrack network in the vicinity of the tool bit. INTRODUCTION Rock indentation has been intensively studied for more than three decades, for two main reasons: the requirements of the underground space sectors (storage tunnels, mining activities) and the development of rock excavation technology. Drilling performance depends on the rate of penetration of the tool, which can be predicted by numerous methods, including those based on indentation tests on representative samples. Several phenomena occur during the indentation process. The first phase of deformation is elastic; however, the loading evolution leads inevitably to the development of irreversible deformation, which proceeds to the initiation and propagation of cracks due to the tensile stresses, and then to fragmentation. This behaviour of the rock under indentation has been widely studied (Lawn and Wilshaw, 1975; Kutter and Sanio, 1982; Sanio, 1985; Nelson, Ingraffea, and O'Rourke, 1985; Wijk, 1989; Fowell, 1993, Kou et al. , 1995; Tan, Kou, and Lindquist, 1996; Alhossein and Hood, 1996).

The observational method includes the idea that rock engineering design can be modified as construction progresses, and that identification (confirmation or rejection and updating) of e.g. rock classes can be used in order to choose from predefined design classes. To facilitate rock engineering design, we suggest that geological and hydrogeological reference conditions are used as a basis. Fundamental to a reference condition approach is the identification and sampling of reference sites. Attributes, in this case geological and hydrogeological attributes, are selected and used to define the reference conditions. The reference conditions should be defined and described early on in a project, using relevant engineering parameters, and grouped together into rock classes with similar engineering characteristics. The reference conditions are subsequently used to define e.g. grouting design classes, to adapt to project-specific flow requirements. In this paper we suggest and give examples of geological and hydrogeological reference conditions for grouting, and present an example where reference conditions were used to define the grouting design for a site in crystalline rock, located at a depth of 400 m at the Äspö Hard Rock Laboratory, Sweden. INTRODUCTION Constructions in rock below the groundwater table will influence both flow and groundwater levels. Sealing these constructions using e.g . grouting will improve working conditions and limit the impact on groundwater levels in the surroundings. Successful sealing will decrease the flow into the construction, resulting in a decrease in the pumping capacity required at the site. If the water quality is unacceptable, smaller volumes of water will need to be treated. The aim of this paper is to present a systematic way of handling grouting, starting from geological and hydrogeological reference conditions for crystalline rock, continuing with examples of these reference conditions for a Swedish field site, and concluding with an example from the same site, where one of the reference conditions was used for the grouting design.

Severe instability in an underground hard-rock bord and pillar mine situated on the Great Dyke in Zimbabwe necessitated a thorough revision of the existing design. Local geotechnical conditions characterized by undulating, subhorizontal and intersecting shear structures, variable in thickness and relative location within the hangingwall, orebody (pillars), and footwall, were only partially understood prior to the instability event. As a result, a change in the behaviour of the respective horizons was encountered, with an unfavourable outcome. A revision of the pillar design therefore became necessary. The revised design took a holistic approach, beginning with a re-evaluation of the local rock mass characteristics, a revision of the empirical design, and application of both linear elastic and nonlinear numerical analysis to test and optimize the final outcome. The resulting design is considered to be robust, with the provision of continual evaluation and monitoring of geotechnical conditions concurrent with production activities to test and validate the design. This paper presents a high-level overview of the investigative process that was undertaken to provide the operation with revised, stable pillar design criteria. INTRODUCTION Production at a bord and pillar mine situated on the Great Dyke of Zimbabwe recently suffered a serious setback when geotechnical conditions within the hangingwall, orebody, and footwall gave rise to widespread instability. Unfavourable rock mass behaviour included footwall heave, falls of ground (FoG), and bulking and spalling of pillars. The presence of a shear structure, subhorizontal and undulating in attitude, variable in thickness and continuity, was identified as a primary cause for the deterioration in rock mass conditions. Investigations into the contributing factors that gave rise to the situation were carried out, based on which it was concluded that the design and associated mining practices evidently needed to be revised. A new, robust, workable model was required that would allow the resumption of mining with confidence while at the same time trying to avoid a knee-jerk response of excessive conservativism.

The paper discusses an essentially iterative design and construction process for a rock shaft excavation undertaken to provide a pump chamber sited on the fringes of one of South Africa's primary water storage reservoirs, Vaal Dam, to service a major inter-basin water transfer project. Preliminary design for the facility was based on a limited geotechnical drilling investigation which broadly covered the intended footprint of the intended abstraction works. The project then proceeded rapidly to construction, the outcome of which indicated a significantly different response to rock excavation from that envisaged by the designers. Detailed design for rock support necessitated several phases of additional drilling and geotechnical evaluation in parallel with construction, each of which was designed to inform the ongoing excavation and rock support programme. The experience provides further vindication of the well-established view of geotechnical practitioners, that as a consequence of the variability inherent in geological formations, some more than others, the initial investigation provides only a limited appreciation of circumstances that will be encountered and thus a preliminary guide for design, details of which need to be developed, refined, and updated throughout the construction process. INTRODUCTION The Vaal River Eastern sub-System Augmentation Project (VRESAP) is a South African highveld interbasin water transfer scheme, originating in the Vaal River catchment and feeding onto the watershed and headwaters of the Olifants River basin. Planning for the project was initiated around 2004/2005 by the Trans-Caledon Tunnel Authority (TCTA) on behalf of the Department of Water Affairs (DWA). VRESAP was designed by the Project Engineer, Vaal Pipeline Consultants (VPC), as an emergency water supply to supplement the water resources of the Mpumalanga coalfields, which feed numerous surrounding Eskom power stations and the Sasol petrochemical plants around Secunda. The project comprises an intake works and low-lift pump station facility located at the Vaal Dam reservoir, desilting works, a high-lift pump station – known as the Boschkop Abstraction Works, and roughly 120 km of 1.9 m diameter pipeline terminating at the Knoppiesfontein diversion weir near Secunda. Rock excavation and support work for the intake/low-lift pump station are the subject of this paper.

The onset of yielding for the walls of deep underground excavations in hard rock is significantly less than the laboratory strength of intact rock samples. The wall stability is rather controlled by the microfracturing strength of the rock ( i.e . crack propagation (CD) and initiation (CI) thresholds). Factors such as loading and unloading effect of excavation, glaciation, stress rotation in front of a tunnel face, etc . can contribute to the progressive damage of the rock, influencing the CI and CD thresholds and therefore contributing to the damage intensity around underground openings. The effect of stress rotation and stress fatigue on crack damage thresholds is systematically investigated in this paper through state-of-the-art laboratory testing and monitoring techniques in combination with advanced grain-based numerical modelling. INTRODUCTION The long-term stress path for the near-field rock surrounding deep underground excavations is typically more complex than simple monotonically increasing deviatoric stress conditions. The state of stress for a point in the wall of an underground opening can go through multiple loading and unloading cycles as the excavation advances and moves past the point. For underground structures such as deep geological repositories (DGRs) with ultra-long design life ( e.g . a million years), the stress oscillation can continue after the excavation is completed due to glacial loading/unloading cycles. Microfracturing behaviour of intact rocks can change significantly when they undergo mechanical loading/unloading cycles, in comparison to the crack damage parameters that are measured from the initial state of the rock. Strength degradation through accumulation of crack damage for brittle rocks was examined in detail by Martin (1993), Eberhardt (1998), and Diederichs (2000) through cyclic testing of the Lac du Bonnet granite. Martin (1993) captured the cohesion loss and mobilization of friction resulted from accumulation of crack damage within intact brittle rocks through a series of damage-controlled cyclic tests in which the maximum applied stress was increased in each consecutive load cycle. This was mainly accomplished by the investigation of the critical damage (CD) locus that was defined from the collected strain gauge data. Eberhardt (1998) further advanced the knowledge of the ‘fatigue’ strength of brittle rocks by using the acoustic emission (AE) technique in addition to the measured strain data collected from damage-controlled cyclic tests whereby samples were loaded beyond CD, below the CD threshold, or loaded incrementally to reach failure.

The site for a residential apartment building overlies an abandoned iron mine in granitic gneiss in northern New Jersey. The mine stope is about 137 m (450 long) and dipping over 344m (800 feet) at 450 to 500. As the building footprint straddles the mine site needed remediation. The remediation scheme consisted of compaction grouting a minimum 10 m (30 ft.) depth of the mine stope in rock to establish a buttress for the hanging wall and allow support of the building foundation. The rock strength parameters (friction and cohesion) were established based on Hoek Geologic Strength Index (GSI). The derived strength parameters were used in the wedge analysis to simulate rock cave-in. It was concluded that a cave-in would be unlikely. Verification holes confirmed the effectiveness of grouting. Although post grouting micro gravity survey depicted a few anomalies, no anomalies were found to exist by further drilling and excavation. INTRODUCTION The abandoned Bull Frog Mine is located in Morris County, New Jersey, in an area where several mines were operating in the late 19th century into the 20th century. Figure 1 depicts the location of the site which borders the major interchange of I-80 in northern New Jersey. Figure 2 is extracted from USGS map (Sims and Buddington, 1958) depicts the Bull Frog Mine (No. 22) in relation to other mines in the area. The Mt. Pleasant mine (see No. 21 on Figure 2) strikes north of Bull Frog Mine and extends under Interstate 80 and has caused problems in the past with subsidence under I-80. It is believed that the Mt. Pleasant Mine ends at the shaft discovered just over the northern property line of the site in the NJDOT Right of Way at the northeast terminus of the Bull Frog Mine (See Figure 3). The iron ore was removed by open stoping with some timbers and stulls used (Shea 1977). The iron ore was formed in foliated planes in granitic gneiss rock although in some places the ore did not occupy the entire foliated area. After the ore was removed from such planes, small to large pieces could fall from the hanging wall that eventually may fill the narrow stope cavity and provide support (Shea 1977).

The design of support pillar layouts in shallow South African mines has historically been based on comparisons of pillar strength to pillar loading, providing some acceptable factor of safety. For this evaluation of pillar stability, empirical strength formulae were compared to either tributary area or modelled loading. Empirically founded pillar strength formulae are based on recorded cases of stable and failed pillars in specific mining areas. These methods inherently include area-specific behaviour of both the overburden system that loads pillars and the responding pillars. In applying empirically derived formulae to a project area, it is assumed that the relationship between the loading system and pillar behaviours is similar to that in areas for which the formulae were derived. However, the same pillar strength formulae appear to have been applied on most of the South African gold, platinum, chrome, and manganese hard-rock mines, all of which have substantially different loading environments. Despite this possible incorrect application of empirical formulae, vast mined-out areas supported by pillars so designed have historically remained stable. However, it is not known whether these pillar systems are in a state of unstable equilibrium that will lead to system failure, or so stable that substantial amounts of ore are unnecessarily locked up in the pillars. In this paper, a method for support pillar layout design that promotes the determination of area-specific loading system and pillar behaviour is described, making use of computer-based numerical modelling techniques. The method is then used to evaluate the potential for optimizing the pillar layout at a bord-and-pillar coal operation. The results indicate huge potential for optimization, but also the critical need for substantial rock property data and instrumentation for calibration. INTRODUCTION Historically, support pillar layouts in shallow South African hard-rock and coal mines have generally resulted in stable workings. In these layouts, pillar dimensions were mostly derived using empirical pillar strength relationships, derived from back-analyses of stable and/or failed pillars. Despite the apparent success, pillar behaviour modelling using empirical methods is explained solely by a factor of safety (FoS). The description of pillar behaviour is more complex than just quoting a single number, and should also include reference to issues such as pillar stiffness, foundation failure/punching, yield point, and peak strength, as well as failure mode.

The study sought to establish the applicability of rock mass classification as a primary input to wall control blasting at a hard-rock mine in Botswana. Conventional rules of thumb are used to develop blast designs based on parametric ratios with insufficient consideration of the rock mass factors that influence the achievability of final wall designs. Control of the western highwall of the mine's pit had proved to be challenging in that the designed catchment berms and wall competence were perpetually unachievable from the pit crest to the current mining levels. This exposed the mining operation to safety hazards such as local wallrock failure from damaged crests, frozen toes, and rolling rockfalls from higher mining levels. There was also an effect of increased standoff distances from the western highwall, which reduce the available manoeuvring area on the pit floor and subsequently the factor of extraction that is safely achievable. The study investigated the application of rock mass classification and the blastability index (BI) as a means to improve wall control. This was achieved by establishing zones according to rock type forming the western highwall rock mass, in which the distinguishing rock mass classification factors were categorized and used to establish the suitable wall control designs through a design input tool (DIT). The DIT consolidated rock mass classification methodologies such as the Geological Strength Index (GSI) and the Rock Mass Rating (RMR) and related them to the BI and discontinuities of the rock mass to produce a tool that can be used to develop objective wall control designs. The designs developed by the tool inherently take into account the rock mass characteristic factors at the centre of the rock mass classification methods above and significantly reduce the dependence on rules of thumb. It was found that this approach yields designs with explosive energy measures that are consistent with the rock-breaking effort and the behaviour of discontinuities, all the while remaining biased towards preservation of perimeter wall rock. The work presented in this paper is part of an MSc research study at the University of the Witwatersrand.

The stability of slopes and underground excavations is generally affected by the geological structure of the rock mass, and it is important that this factor is adequately assessed. Whether instability will arise from the presence of the structures depends on the tectonic events that resulted in the formation of the structure, the orientation, characteristics (width, infill, strength, ability to dissipate strain energy, etc .), and the effect of mining-induced stress on the structure, which either enhances or reduces stability. Three case studies are presented where the influence of geological structures resulted in major instability in Southern African and Canadian underground mines. One case resulted in changes to the mining layout and support system, one resulted in a support system change, and the third resulted in the collapse of the mine, necessitating redevelopment and sterilization of ore. CASE STUDIES 1 AND 2: PLATINUM AND DIAMOND MINES Introduction A platinum mine in Southern Africa and a diamond mine in North America experienced similar instability associated with a weak (shear) layer in either the pillar foundations or within the pillars themselves. In the one instance, the instability initially manifested as footwall heave, unravelling of the pillar walls, and eventually in the form of a pillar run that resulted in a collapse of the underground workings. In the second instance, the initial instability comprised slabbing of the pillar walls below the weak layer, and footwall heave was observed in another less-frequently accessed roadway. Unfortunately, the long-term effect of this failure mode could not be assessed as the mine has since closed due to economic reasons. Platinum Mine Study Mechanized mining was taking place on a room and pillar layout, with the only footwall development comprising a conveyor belt decline shaft which was used to convey ore from underground to surface storage bins. The mining height was approximately 2.25 m and pillar dimensions varied according to ground class (typically classed according to the NGI Q-system). Stable pillars, designed to support the load of the overburden, were intended to provide both local and regional stability of the underground workings. Initial instability manifested as footwall heave (Figure 1) adjacent to several pillars in the shallower portion of the mine and unravelling of blocks along joints from the pillar walls (Figure 2).