Slope instability in open pit mining environments present significant safety hazards. The quality and quantity of geotechnical data often expands over time and may give rise to an increase in reliability and a corresponding reduction in uncertainty of input parameters. In this paper, the geotechnical model is built on data obtained from four different consultants over 15 years, spanning from conceptual study to design. Stability conditions are investigated through limit equilibrium method and compared to the numerical analysis using finite difference method. Three critical profiles, based on areas of known concern, are analysed. Kinematically admissible joint orientations are incorporated as ubiquitous joint models and materials are modelled based on the Mohr-Coulomb failure criterion. Limit equilibrium method results revealed that profile A is the most critical slope, with a significant probability of planar and wedge failure at set of benches. Safety factors for overall slope indicate planar failure of profile B, although stable, remains below the acceptance criteria for the overall slope angle, which opted for numerical analysis. Profile C was deemed stable and no further analyses were required. There was a good agreement between methods of analysis, in terms of safety factors and failure surfaces. Finite Difference Method computed lower safety factors to the point of critical stability. INTRODUCTION Mining operations in open pit mines produce progressively deeper pits. These structures account for a large portion of the world's mineral production (Wyllie and Mah, 2004). With the widening scope of mineral applications, their demand is driven by an ever-growing population (Lusty and Gunn, 2015). The United Nations (2015) projected a global population increase of more than 1 billion within the next 15 years; an increase from approximately 7.3 billion in the year 2015 to 8.5 billion. The imminent future thus carries the mineral industry into even more precarious environments in order to accommodate this demand (Lusty and Gunn, 2015). With increasing depths comes increasing risks of slope failure and it is thus essential to rigorously manage the hazard associated with rock slope stability. Fortunately, the advancement and increased utilization of computational tools now allows mine personnel to make improved informed decisions (Hochbaum and Chen, 2015). With increasing depths and associated stresses, Stacey et al . (2003) emphasized the importance of numerical stress analysis methods and Hoek, Rippere, and Stacey (2000) promoted the undertaking of numerical modelling, particularly for more complex slopes or lithologies where limit equilibrium analyses are often too simplistic.

Time-dependent closure data in deep hard-rock mines appears to be a useful diagnostic measure of rock behaviour. Understanding this behaviour may lead to enhanced design criteria and modelling tools. In this paper we investigate the use of a time-dependent limit equilibrium model to simulate historical closure profiles collected in the South African mining industry. Earlier work indicated that a viscoelastic model is not suitable to replicate the spatial behaviour of the closure recorded underground. The time-dependent limit equilibrium model available in the TEXAN code appears to be a useful alternative as it can explicitly simulate the on-reef time-dependent failure of the reef. A key finding in this paper is that the model gives a much better qualitative agreement with the underground measurements. For both the model and actual data, the rate of time-dependent closure decreases into the back area. Calibration of the constitutive model nevertheless led to some unexpected difficulties, and the element size plays a significant role. It was also noted that the simulated closure is complex as it reflects the combined result of a number of elements failing at different times. The closure rate does not decay according to a simple exponential model. In conclusion, explicit simulation of the fracture zone in the face appears to be a better approach to simulate the time-dependent behaviour in deep hard-rock stopes. The calibration of the limit equilibrium model is very difficult, however, and further work is required. INTRODUCTION Several research studies have been conducted to investigate the use of continuous stope closure as a diagnostic measure of rock mass behaviour in the deep gold mines of South Africa ( e.g . Malan, 1995; Napier and Malan, 1997). The rock mass undergoes significant time-dependent deformation in some geotechnical areas, and the closure data is useful for identifying different geotechnical areas and areas prone to face-bursting (Malan et al ., 2007). The value of continuous closure measurements was also demonstrated for the platinum industry. An example is shown in Figure 1 to illustrate the significant ‘creep’ component recorded in some of the areas. For the gold mines, it was proposed that the data may be useful to identify remnants that may be safely extracted. A difficulty faced in these early studies was that no numerical tool could simulate the time-dependent rock mass behaviour on a stope-wide scale. It was hypothesised that the time-dependent behaviour in the hard rock gold mines is caused by timedependent fracturing and other inelastic processes, such as gradual slip on parting planes. In the platinum mines, the face fracturing is less intense and the behaviour is most likely dominated by the time-dependent failure of the crush pillars. The commonly used elastic modelling programs cannot simulate this behaviour and the simulated convergence is simply a function of the mine geometry, depth, and elastic constants.

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.

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.

As projects advance to deeper levels, rockbursts occur more frequently. The rockburst failure mode is particularly problematic, as the rock mass fails abruptly, releasing large amounts of energy. Laboratory tests and observations in tunnels and mines with high overburden have shown that certain rock types have a high potential for storing elastic energy and are hence more prone to rockbursting. In this research we studied the influence of a rock's heterogeneity at grain scale on its rockburst propensity. Several artificial sample sets consisting of a very fine-grained fibreless ultra-high performance concrete (UHPC) and a constant volumetric fraction of different coarse rock grains as aggregate were produced and tested. The added rock grains were chosen based primarily on their differing stiffness characteristics. One sample set, consisting of pure UHPC, acted as reference for the matrix. The findings of this study confirm that the material's heterogeneity at grain scale has a major impact on the failure behaviour. With increasing stiffness heterogeneity between the ‘rock matrix’ and admixed grains the failure mode becomes more ductile and a lower uniaxial compressive strength is reached. Hence, the material is less prone to rockbursts as less elastic strain energy is developed prior to failure. The results agree well with various rockburst parameters, which were additionally evaluated and compared. The findings allow a better understanding of the underlying mechanism of rockbursts and demonstrate the usefulness of petrographic information in rockburst risk analysis. INTRODUCTION Predicting the behaviour of the rock mass is of utmost importance in tunnelling and mining. Only if the potential failure mechanisms of rock and rock mass are well understood can failure be foreseen and avoided and the consequences reduced. Rockbursts are particularly problematic, as the rock mass fails abruptly, rapidly releasing great amounts of energy. Laboratory tests and observations in tunnels and mines with deep overburden have shown that certain rock types have a high potential to store elastic energy. Hence, they are more prone to rockbursting. The amount of stored energy also determines the magnitude of the event.

Parameters needed for rock support design are discussed. Underground loading conditions and the methodology of rock support are covered first, followed by the parameters related to rock support design; that is, the factor of safety, the maximum allowable displacement, the ultimate displacement of support elements and the released strain energy, and their application in support design. The factor of safety needs to be calculated in different ways depending on whether the loading condition is gravitational, squeezing, or a dynamic rockburst. INTRODUCTION The major instability issue related to high in-situ stresses is either rockbursting in hard and strong rock or time-dependent squeezing in soft and weak rock. The ‘driving force’ behind both rockbursting and rock squeezing is the strain energy released from the rock mass after excavation. In the case of rockbursting, the energy released in a seismic event has to be absorbed by the rock support system in order to avoid rock collapse. Therefore, it requires a dynamic support system capable of absorbing a good amount of energy prior to failure. This concept was proposed as early as in the 1960s, and has been well acknowledged by others to date (Cook and Ortlepp, 1968; Ortlepp, 1969, 1992). A good number of energy-absorbing rockbolts have been invented (Ortlepp Bornman, and Erasmus, 2001; Charette and Plouffe, 2007; Varden , 2008; Li, 2010; Cai and Champaigne, 2010; Wu and Oldsen, 2010; Galler, Geschwandtner, and Doucet, 2011; He et al ., 2014; Darlington 2014). On the other hand, there exists a maximum allowable displacement from the point of view of engineering practice. Therefore, two parameters, the amount of energy to be absorbed and the maximum allowable displacement, need to be taken into account in dynamic rock support design. In addition, the displacement capacity of support elements, such as rockbolts, needs to be also taken into account. A satisfactory dynamic rock support system should be capable of absorbing a good amount of energy, but should also be capable of constraining the rock displacement within the maximum allowable limit. In principle, a strong support system is desired so that it does not need to displace much to absorb a given amount of released energy. In this paper, the methodology of rock support is discussed first and then the parameters used for rock support design, such as the factor of safety and the maximum allowable displacement, are presented for the following three types of rock conditions: rockfall under gravity, rock squeezing, and rockburst.

In mining, the design of optimum pillar strength has benefits ranging from increased extraction ratio (economic benefit) to the stability of the hanging-wall (safety benefit). Pillar strength is traditionally determined by empirical calculations and based on square, or near-square, pillar shapes. This study focuses on the preliminary results of investigation aimed at determining a pillar shape that could provide optimum strength. Pillars of different surface shapes (constant area) and constant height were compared. Four pillar shapes were chosen for analysis: square, triangular, hexagonal, and circular. For the physical models, a concrete mix was used to compare the pillar strengths to an engineering-designed strength parameter. These physically modelled pillars were loaded using a 5000 kN servohydraulic control console and load deformation graphs for each shape were obtained. The preliminary results indicate that the hexagon pillar has the highest peak strength of 76% normalized to the cube strength, while exhibiting stiff support properties. The triangular shaped pillar had 67% normalized strength. The triangle was found to have the highest energy absorption capability compared to the other shapes. It is noted that shapes such as the circle were not considered since they would be impractical. To ensure reliable results, numerical modelling was used along with a theoretical design criterion. The criteria to assess the pillar shapes were based on effective width and hydraulic radius. INTRODUCTION Significant research has been done on the effect of pillar shape, with many researchers contributing to this subject. With width-to-height ratio being considered a shape factor by researchers, the UCS of a rock mass is also emphasized as fundamental. Likewise, a considerable amount of research is focused on pillar design methodologies. However, little attention has been given to the influence of cross-sectional area shape on pillar properties, which is an important aspect in pillar design. The majority of the research that has been done to date has focused on the influence of width to height ratio on pillar behaviour and pillar strength.

Release of the in situ state of stress is often the most relevant action during the excavation of underground structures in rock masses, such as powerhouse caverns for hydroelectric schemes, which are often excavated at great depth and have large dimensions. Design of the supports, namely using numerical models, requires an estimate of the magnitude and direction of the principal stresses. However, a large number of factors influence the in situ stress field and its characterisation is a difficult task. Besides, the number of tests carried out for determination of the in situ stresses in rock masses is usually small, due to cost and time constraints, and all the testing methods have limitations inherent to their nature. The paper presents methodologies for analysis of the stress field obtained from the results of in situ stress measurements using three methods: overcoring, hydraulic fracturing and flat jacks. These methodologies integrate all stress measurements and use numerical models of the rock mass that represent the ground topography, the lithology, and the geometry of the underground openings, so that the most likely stress field in the zone of interest for the design is obtained. Instances of application to three underground powerhouses are presented. INTRODUCTION Powerhouse caverns are underground structures of relevance in the design of hydroelectric projects. In power schemes with high water head in mountainous regions, the power conduit is often built underground. Thesual reasons for this are the better quality of the rock mass at depth and the higher in situ stresses, which enable the construction of unlined pressure tunnels and shafts, with economic advantages. The increase in the number of reversible, pump-storage projects, where the hydraulic power units (pump turbines) must be located considerably below the downstream water level, also calls for the construction of large underground powerhouses at great depth, which often exceeds 500 m.

Crustal stress, and especially its orientation, is one of the key parameters in geotechnical mechanics and engineering. Because of the complexity of traditional crustal stress testing methods, especially in deep holes, the determination of crustal stress orientation is difficult, which is hindering the development of crustal stress testing technology. Therefore, the quick, effective, and accurate measurement of crustal stress in deep boreholes has important practical significance. This paper presents a method for the fast, effective, and accurate determination of deep crustal stress orientation by using panoramic stereopair imaging technology and stressed borehole geometric shape. Firstly, borehole panoramic stereopair images of the borehole wall are acquired using a biconical mirror, and the geometric features of the wall, are characterized by constructing a digital elevation model (DEM) of the borehole wall. The borehole geometric shape under crustal stress state is then analysed. In general, the circular borehole will become elliptical in shape. This paper presents the standard equation of the ellipse and establishes the relationship between the elliptical shape parameters and crustal stress. Lastly, from this relationship and the borehole geometric shape obtained by stereopair imaging, the orientation of crustal stress is determined utilizing the positioning capability of this technology. Simulations show that the borehole geometries and orientations are in agreement with the actual crustal stress orientations without the help of petrophysical parameters. This is a direct method of crustal stress orientation assessment based on the analysis of borehole geometric shape obtained by using the latest borehole stereopair imaging technology. This constitutes a new way of measuring deep crustal stress, which overcomes technological problems and promote the development of crustal stress measurement technology.

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.

Simulating pressure changes and production processes in natural fractured formations has been commonly performed by the classic dual-porosity model. Geomechanics factors can be crucial for pressure change and production as both formation deformation and locally induced stresses can contribute to the pressure change thus production significantly, particularly in low-permeability and fractured formations. Multiphase flow in both the matrix and fractured system can significantly affect the production and pressure results, particularly near a wellbore or a hydraulic fracture. The key parts on developing this model are to come up an efficient algorithm and define the parameters characterizing the geomechanics coupling to the flow on top of the explicit coupling to the saturation, where upstream weighting is applied. The proposed algorithm coupled the pressures from both system with the corresponding volumetric strain and subsequently explicitly coupled to the saturation. Petrov-Galerkin finite methods are used where a parabolic approximation verse a linear one is implemented respectively for the displacement and pressures/saturation. INTRODUCTION Flow in naturally fractured reservoirs is normally characterized and modelled by the well-known sugar cubic model initiated by Barenblatt et al , (1960) and proposed by Warren and Root (1963). Extensive studies follow by mixture theory focusing on stress-sensitivity and loosely geomechanics coupling (Wilson and Aifantis 1982; Beskos and Aifantis, 1986; Huyakorn and Pinder, 1983; Elworth and Bai, 1992; Bai, Meng, and Roegiers, 1999) and by an alternatively conventional reservoir-geomechanics coupling model following Biot's theory (Biot, 1941; Duguid and Lee, 1977; Vallianppan and Khalili- Naghadeh, 1990; Khalili-Naghadh and Valianppan, 1996, 1997; Chen and Teufel, 1997). Following Vallianppan and Khalili-Naghadeh (1996) and Chen and Teufel (1997), an extended Biot theory coupling the fluid flow to deformation is proposed and implemented. Those critical parameters between stresses and pore pressure and the deformation coupled between pore volume, bulk volume, porous matrix, solid particle, and fractured volume are highlighted. Studies on dual porosity model have been extended into two phase flows (Nair, Absouleimen, and Zaman, 2005; Bai, Meng, and Roegiers, 1999) by a mixture theory. Following Wang and Chen (2001), in which only a single poroelastic model is proposed, a two-phase flow model with extended Biot formulation will be the focus in this paper.

The ‘A’ formation in the Abu Gharadig basin, northern Western Desert of Egypt is a carbonate reservoir characterized as a low-permeability soft chalk with medium porosity and low-quality natural fractures. Owing to these characteristics, achieving economical gas production rates and efficiently developing the reservoir have become increasingly challenging. To increase production from the field, a full development plan was initiated employing drilling of horizontal wells with multistage fracturing stimulation. An integrated and detailed programme of laboratory core testing was conducted to gain an in-depth understanding of this reservoir's rock mechanics behaviour and to help optimize the hydraulic fracturing and completion designs. Particularly, elastic properties and principal in-situ stresses were acquired to provide the following: Static and dynamic mechanical property information for correlating well log data Calibration to the mechanical earth models constructed from sonic-derived mechanical properties to help provide realistic deformation parameters for hydraulic fracturing design purposes. Strength information for developing a failure locus for the material. Mechanical properties testing was conducted on selected core samples and integrated logs from two offset vertical wells. The laboratory testing programme consisted of the following: Indirect tensile strength testing (Brazilian method) at room temperature. Unconfined compression strength (UCS) testing Single-stage triaxial compression testing with concurrent ultrasonic velocity measurements. All laboratory data, including procedures, tabulated results, stress-strain plots, computerized tomography scan images, and post-test sample images are provided for analysis and visualization of the failure mode. Understanding the reservoir geomechanical behaviour and the mechanical parameters that have a critical impact on hydraulic fracturing propagation facilitated improved decision-making in terms of fracturing design and optimization. Hydraulic fracturing simulation software was used to model the propagation of hydraulic fractures based on the integrated reservoir mechanical properties. Different scenarios of perforation clusters were run to model propagation of multiple horizontal fractures and predict the changes in stress anisotropy in the neighbourhood of the fractures.

Large underground unlined rock caverns for storage of crude oil or liquid hydrocarbons are very often designed to operate with a gaseous phase maintained to a pressure range above atmospheric pressure (>1 bar abs.). To maintain the pressure of the gaseous phase in a cavern above atmospheric pressure it is necessary to inject a compensation gas (gas in) during the cavern emptying phase. The reverse operation (gas out) must be carried out during product filling to avoid overpressurization in the cavern. The management of this compensation gas requires a dedicated industrial unit for production, recovery, and treatment. In case of large underground caverns with net capacity of several million tons, the construction and operation costs for such a unit become prohibitive. To limit the need for compensation gas, the permissible pressure range in operation must be extended. Whereas the maximum gas pressure is mostly limited by the tightness issue, the minimum pressure is mainly governed by stability aspects. In this paper, the possibility of lowering the minimum pressure in operation below the atmospheric pressure is investigated, focusing on the cavern stability aspects. During construction at atmospheric pressure, for the large cavern section and rock mass quality expected, long-term stability can be achieved by conventional spraying of shotcrete and rockbolt installation. During operation, minimum pressure in the storage cavern close to vacuum is investigated and its impact on reinforcement need is reviewed using results of two distinct approaches: homogeneous continuum as well as wedge stability. The particularities regarding the required reinforcement for caverns operated at atmospheric pressure are also investigated. INTRODUCTION Underground rock caverns excavated for the storage of crude oil have capacities of several millions of cubic metres. These caverns are typically excavated in hard rocks, allowing the excavation of large section supported with rock bolts and shotcrete. To reduce the construction cost of a vapour recovery unit and the associated cost of gas compensation during operation of underground rock caverns for storing liquid hydrocarbons, the minimum operating pressure anticipated in the cavern is extended below the atmospheric pressure. As the lower pressure is applied around the cavern wall during operation only, it is important to distinguish between the reinforcement needs during construction and during operation to optimize the construction schedule of the project. Indeed, only the support required to achieve stability during construction remains on the critical path, whereas the installation of the support needed to achieve stability in operation can be delayed and rejected from the critical path.

In the last two decades the demand for primitive stress measurements has increased in the Southern African mining industry. The main reason for this is recognition that the natural complexity and variability of stress in the Earth's crust renders elementary primitive stress estimates inaccurate for engineering purposes. Consequently , primitive stress measurements in situ are essential for a viable primitive stress model of a mine. This paper summarises and reviews experience gained during the last decade when using the CSIRO HI cell (Australian Commonwealth Scientific and Industrial Research Organisation Hollow Inclusion cell) and the BXCCBO strain cell (BX-sized Compact Conical-ended Borehole Overcore cell) to measure in situ stress in Southern Africa. These two strain cells have proven their reliability, but stress measurements remain complex and the risk of measurement failure can be high. The authors describe techniques to control and minimise the risk of measurement failure. There are no reliable benchmarks for measured primitive stresses; therefore the results obtained must be checked qualitatively against major geological trends at the measurement site, and against a background of consistent primitive stress data from the Southern African Stress Database. The paper concludes that the risk of measurement failure can be reduced to an acceptably low level to assure the acquisition of reliable and useful stress measurement data. INTRODUCTION Rock mass structure, rock strength, mining induced stress, and primitive stress are the four most important factors that determine the stability of a mine. Measurements of intact rock strength, together with structures such as joints, faults, folds, intrusions, layering, and any other geological features, help to determine rock mass strength and integrity. The strength of the intact rock material is determined from laboratory testing of intact samples. The frequency and properties of discontinuities, such as joints and faults, are less well known because they are inferred from geotechnical information, and sometimes laboratory measurement data. Modern numerical models can take account of rock mass properties as they are estimated or measured from geotechnical data, and the mining-induced stresses estimated accordingly from an appropriate numerical model of the mine layout. The factor that remains least understood–and least measured–of all is the primitive stress before mining. Without knowledge of the primitive stress from a substantive model backed by measurement data, a complete engineering model of the mine in the rock mass, and hence effective mine design, is not possible.

Most of the lead-zinc mines in the state of Rajasthan, India currently employ the blast-hole stoping method with backfilling. One of these mines, Rajpura Dariba mine, has reached a depth of about 700 m. In this mine the footwall and hangingwall of the orebody contain sheared regions, represented by narrow 0.1–1.2 m zones of crushing, brecciation, and gouging, which are fairly persistent along the strike and dip, and are found close to the orebody contacts. Extraction of ore from these lodes is challenging at greater depth. Major geotechnical issues encountered during stoping include maintaining footwall drives and crosscuts, especially when they are closer to the stoping areas and crossing a shear zone, premature collapse of hangingwall rock prior to mucking of ore, and instability on account of blast vibration. In this paper we discuss the findings from three-dimensional numerical modelling studies conducted using FLAC3D to arrive at a feasible stoping sequence for maximum ore recovery. The sequence finally suggested involves an upward mining sequence using primary and secondary stoping configurations with sill pillars and vertical (rib) pillars. The influence of stoping, as seen in numerical models, extends about 25 m into the footwall; this is corroborated by field observations, where the footwall drive within 25 m of the stope experienced severe spalling and roadway closure. The influence of shear zones was found to be crucial in locating the drivages and designing the stoping sequence at greater depths. INTRODUCTION Lead-zinc mining in the state of Rajasthan, India has been conducted since ancient times. Rajpura-Dariba mine is located at the southern extremity of the Rajpura-Dariba Bethunmi metallogenic belt in Rajsamand district, Rajasthan. The orebodies at Rajpura Dariba are designated as the Main Lode (South and North) and East Lode as shown in Figure 1. The Main Lode extends over a strike length of 1700 m and is separated into South Lode and North Lode by a barren stretch of 300 m. The South Lode (L1), striking north-south and dips 60° to 70° towards the east, and has a strike length of 500 m. The North Lode has a strike length of 900 m. It strikes N-S and dips 70° to 75° towards the east. The average widths of South, North, and East lodes are about 24 m, 18 m, and 18 m respectively and tend to decrease with increasing depth.

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.

A knowledge of the geomechanical properties of the rock mass is a critical requirement for improving almost all mining processes, from geotechnical design, drilling and blasting through to mineral processing. Current methods used to obtain this data such as mapping, core drilling, chip sampling, and geophysical techniques are expensive, yield limited results, are time-consuming, and valuable time is lost in the planning process during the analysis of the data. Estimating rock properties on-line and in real time simply by analysing the response of key drilling parameters to the rock mass is not new, and the technique has been used successfully for many years in the petroleum industry. However, until recently, attempts to obtain reliable and repeatable correlations between these parameters and rock properties in mining have proven to be only partially successful. This paper describes recent advances in the technology and its application at two of the largest open pit mines in South Africa. The data density and quality, and speed with which the rock property data is generated, offers tremendous potential for increasing the efficiency of mining operations by improving numerous short-range planning decisions. Most importantly, the technology provides information at a far lower cost, at greater density and, in some cases, more accurately than conventional methods such as core logging and mapping. The paper describes the geotechnical and other related improvements that are possible from using rock property data generated by the many large production drill rigs that have been equipped at these two mines during the past four years. INTRODUCTION In the mine-to-mill process the nature of the rock mass has a critical impact on almost all the operations, from drilling through blasting, loading, transportation, crushing, grinding, concentration, and waste disposal. However, crucial mine planning decisions are usually made based on a very limited knowledge of the rock mass prior to blasting. Any techniques that are able to provide additional information on the condition and content of the rock mass ahead of mining would therefore be highly desirable and realize considerable value through the optimization of these processes. For example, detailed knowledge of ore boundaries acquired during normal percussion drilling activities would offer substantial mine planning advantages. Drilling and blasting is another example that would clearly benefit from advance information on rock mass properties. The traditional means of obtaining such information is through exploratory drilling, analysis of the rock cores recovered, and geological/geotechnical face mapping. The disadvantages of these methods are clear; disruption to mining activities, time taken to obtain and analyse the samples, additional cost of drilling, loss of core in fractured material, damage to core making an assessment of fracturing difficult, and very sparse sample of data, among many others. Various methods have been used over the years in an attempt to obtain this information as far ahead of mining as possible, with varying degrees of success.

Roof failures in underground coal mines remain a major threat to safe working conditions and can lead to devastating consequences such as a significant impact on the production of coal. The Magdalena Colliery in Dundee, South Africa currently mines two seams; the upper Alfred seam and the lower Gus seam of the Vryheid Formation. These seams are separated by a planar cross-laminated sandstone, here referred to as the parting, which ranges in thickness from 2 cm to 11 m. The parting forms the floor of the upper Alfred seam and the roof of the lower Gus seam. Roof failures are a major cause for concern where the parting forms the roof of the mine. A study of the mineralogy and geotechnical properties of the parting has been undertaken in order to investigate its contribution to the frequent occurrence of roof failures. Thin sections of the parting were subjected to detailed petrographic analysis and X-ray diffraction analysis was undertaken in order to investigate the presence of any clay minerals within the parting. The geotechnical test work includes an impact splitting test, point load test, and Schmidt hammer rebound test on representative samples of the parting. The results classify the parting as a medium- to high-strength rock. However, water and the anisotropic nature of the parting were found to significantly affect its strength. High horizontal stresses associated with the mining method may be of great concern with respect to roof failures in areas where they are concentrated parallel to the laminations. INTRODUCTION Coal mining has played a vital role in the development of energy production in South Africa as coal accounts for 73% of the country's primary energy. In spite of advances in renewable energy technologies, coal will remain the major energy source in the near future due to its abundance and low cost (Jeffrey, 2005). Buffalo Coal currently operates two collieries in Dundee, South Africa; the Magdalena Colliery and the Aviemore Colliery, which are hosted within the Klip River coalfield. The Aviemore Colliery extracts high-grade anthracite coal while the Magdalena Colliery extracts high-grade bituminous coal (Muller et al., 2013). The focus of this study was the Magdalena Colliery, which has been mining bituminous coal since 2005.

In this paper, we describe an integrated remote sensing approach for the collection of geomechanical data to be used as input for continuum, discontinuum, and hybrid numerical analyses. Ground-based and aerial remote sensing techniques, including terrestrial digital photogrammetry (TDP), terrestrial laser scanning (TLS), structure-from-motion photogrammetry (SfM), and terrestrial infrared thermography (IRT) may be used for collecting rock mass data appropriate for input into varied numerical modelling approaches. To demonstrate our suggested approach, we have used the 1965 Hope Slide, British Columbia, Canada. We present the mapping of rock discontinuities for numerical modelling using a hierarchical geological structure order. Large-scale geological structures which were identified and mapped on the pre-failure and present-day topography are used in a preliminary analysis of the rock slope to investigate their influence on kinematic freedom and in bounding keyblocks. Detailed geomechanical mapping is performed on three-dimensional TDP models. IRT data is used to characterize surface water seepage. Unmanned aerial vehicle (UAV) SfM imagery of the landslide debris was used to analyse the block size distribution. Preliminary numerical discontinuum 3D-DEM modelling based on this data and assigned mechanical properties shows that with detailed planning and systematic field data collection techniques, the geological engineer can obtain the data necessary to reduce both model and parameter uncertainty and allow more reliable and realistic numerical slope simulations. INTRODUCTION The stability of high natural and engineered rock slopes is becoming increasingly important. Population growth has resulted in the development of major infrastructure in mountainous areas and in the excavation of deeper open pit mines to meet the demands for resources. The importance of structural and lithological features in providing kinematic freedom for large volume rock slope failures has necessitated significant improvements in rock slope characterization. Additionally, improved characterization of the processes that induce instability of high rock slopes, such as brittle fracture, reduction in shear strength along discontinuities, and changes in kinematic constraint, is of paramount importance in risk assessment and the design of mitigation and stabilization measures both for natural and engineered slopes.

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.