Raiseboring technology has significantly reduced the disturbance caused to the surrounding rock mass compared to historical drilling and blasting techniques. However, the ground through which the shaft is established must be of sufficient quality to allow a free-standing and unsupported excavation. Empirical rock mass classification schemes have been widely and successfully used to quantify the risks attached to a shaft prior to excavation. However, the risk of raisebored excavation failures due to unforeseen ground conditions is still a concern, hence the need for improved stability assessments. This paper outlines a combination of the empirical techniques employed to identify potential geotechnical stability problems in a raisebored ventilation shaft, the numerical approach adopted for determining the effects of jointing, as well as an evaluation of the effectiveness of these methods once the excavation has been established. INTRODUCTION The literature suggests that raises more than 4.0 m in diameter would be considered stable if established in a rock mass of at least fair quality, and that raises up to 4.0 m in diameter would exhibit only isolated incidents of failure when intersecting poor quality ground. The geotechnical risk of a 3.0–4.0 m diameter raisebored ventilation shaft was assessed by applying a combination of empirical techniques supplemented with a numerical assessment of wall stability, due to the importance of this ventilation shaft to the operation. The investigation consisted of the following aspects.

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.

Managing rockbursting conditions in mine development and operational environments is a complex and difficult challenge. The hazard and the associated risks can be managed based on local experience, monitoring, and informed data-rich analysis. On the other hand, blind development for deep tunnelling is being carried out around the world at depths in excess of 2 km and rockbursting has become a common and serious challenge. The rockburst mechanism is predominantly tunnelling-induced dynamic rupture or strain bursting, distinct from remote or mine-generated events impacting mining excavations. Considerations of rock petrology, fabric, mechanical parameters and structure allow an estimate of brittle response. The potential for energy storage and rapid release must be accounted for to understand the burst potential early in the basic design stage for deep tunnels. Failure to do so can result in years of delay and unsafe conditions. In this paper a multistep semi-empirical approach for early strain-burst or dynamic rupture potential along deep tunnel alignments in variable ground is presented. INTRODUCTION Rockbursts are explosive failures of rock which occur when high stress concentrations are induced around underground openings (Hoek 2006) in brittle rock or rock masses with brittle structure. The Canadian Rockburst Handbook (Kaiser, McCreath, and Tannant, 1996) defines a rockburst as: ‘damage to an excavation that occurs in a sudden or violent manner and is associated with a seismic event’. Codelco (2008) uses the following definition of rockburst: ‘Loss of continuity of the production process of the mining operation, caused by the rupture and instant projection of the rock mass, associated with a seismic event.’ It is important to note the evolution, through these three definitions, from a term describing a mechanical phenomenon to a general and then a specific form of damage causing defined consequences. From a hazard and risk management perspective it is important to maintain a separation between the physical process of ‘dynamic rupture’ of the rock around a tunnel (the hazard) and the operational ‘rockburst’ impact (risk). The assessment of the hazard of dynamic rupture is the subject of this paper.

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.

Although the application of destress blasting as an active rockburst mitigation measure is not yet commonplace in hard coal longwall mining, the method is assuming heightened importance due to increases in mining depth and high horizontal stresses in the rock mass. The main goals of destress blasting are the softening of competent rock layers, the reduction of strain energy storage and rock mass stress release, which together contribute to minimizing rockburst occurrence and risk. One such region in which destress blasting in competent rock is applied is the Upper Silesian Coal Basin, mainly in its Czech part. Here a case study of destress blasting is presented and evaluated in terms of rockburst prevention, focusing on a thick coal seam (about 5 m) subject to longwall mining under very unfavourable geomechanical conditions (great depth – 1200 m, competent rigid rocks between coal seams, unfavourable stress field due to long-term mining). Destress blasting stages were carried out in a group of boreholes (4–8; diameter 95 mm) with the total explosive charge ranging from 2100 kg to 3750 kg, detonated regularly at a distance of 43 m to 114 m ahead of the advancing longwall face. Evaluation of destress blasting based on seismic monitoring revealed that the longwall blastings were very successful in terms of rock mass stress release and decreasing the rockburst risk, with only one rockburst occurring on a roadway 210 m from the longwall face. The following longwall section was mined safely without rockburst occurrence. INTRODUCTION In the hard coal reserve of the Upper Silesian Coal Basin (USCB), which is shared by the Czech Republic and Poland, longwall mining is the dominant underground mining method. The Czech part of the USCB, known as the Ostrava–Karvina Coalfield (OKR), lies in the northeastern part of the country (Figure 1). The exhaustion of the upper seams due to the continuation of coal mining activities for around 200 years has shifted activity to a greater depth (> 800 m). Under existing mining and geological conditions in the Karvina sub-basin of the USCB, underground extraction of coal is typically accompanied by rockbursts, the first of which occurred in 1912 (Pelnar 1938). A number of attempts have been made to address rockburst activity in both the Czech ( e.g . Straube et al ., 1972; Holecko et al ., 1999; Takla et al ., 2005; Holub, Rušajová, and Holecko, 2011) and Polish parts of the USCB ( e.g . Dubinski and Konopko, 2000; Drzewiecki and Kabiesz, 2008).

Tunnel design and excavation are complex problems that require in-depth studies and a lot of data. Since the second half of the 20th century, scholars have developed new methods in order to better understand the behaviour of the rock mass and provide appropriate construction solutions. These methods, which are used worldwide, all focus on the importance of monitoring during excavation. This paper describes studies carried out on a new system developed to monitor tunnel deformations during both construction and operation. We developed two different kinds of instrumentation using 3D Modular Underground Monitoring (MEMS)-based sensors, which provide information about convergence of tunnels sections (CIR-Array) and about 3D deformations a few metres inside the rock mass (RAD-Array). This kind of instrumentation is rapid and easy to install, has a high reading frequency (which permits statistical analysis of the data, thus increasing the reliability and efficiency of the system) and features automated storage and presentation of results. Other important aspects are the possibility to extend the monitoring phase beyond construction, and the economy of the system in comparison with other monitoring systems. We carried out a series of laboratory tests reproducing a real-scale tunnel shape by means of a fibreglass rod, to which known deformations were applied. Each deformed shape was measured using different techniques along with the CIR-Array or RAD-Array systems; in particular, photogrammetry and topography, which are two techniques commonly used to investigate tunnel deformations in current practice. We compared the results in order to ascertain the sensitivity and accuracy of each method and evaluate the two systems. The aim of this paper is to collate and analyse the results obtained in order to highlight the potential application of the systems.

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.

This paper focuses on the geotechnical analysis and slope design for the development of the Otjikoto gold mine in Namibia, with emphasis on the characterization and analysis of the sub-surface conditions for the proposed mine to the final pit depths. The geotechnical investigation was based on laboratory testing data, together with data from geotechnical boreholes. The analysis of the sub-surface conditions was assessed using the Laubscher (1990) Mining Rock Mass Classification System. The results were used for limit equilibrium analyses and confirmed using numerical methods. As the focus of the study was based in the classic rock mass classification systems, strength and experience is added to the knowledge base regarding these systems. This allows for a reference point to situations requiring similar solutions. INTRODUCTION Otjikoto gold mine is located approximately 300 km north of Windhoek, Namibia's capital city, between the towns or Otjiwarongo and Otavi (Figure 1). It is owned and operated by B2Gold Corporation and is currently in production. This paper presents the geotechnical feasibility study that was conducted prior to mine construction. Previous geotechnical investigations did not include orientated core drilling, and as such, a major gap was present in the geotechnical model. Data for the study was made available from dedicated orientated geotechnical holes drilled specifically for the project. The paper addresses the geotechnical nature of the intact rock, the joint strengths, and an analysis of bench and inter-ramp (stack) stability as well as overall slope stability.

To assess the stability of large-scale underground caverns within steeply dipping rock strata, a high-resolution microseismic monitoring system was deployed in the right-bank underground powerhouse of the Wudongde hydropower station in southwest China. The progression of damage in the surrounding rock mass, including the initiation, propagation, coalescence, and interaction of rock microfractures, was first demonstrated by analysing the temporal and spatial distribution of microseismic activity. The microseismic monitoring results were validated using conventional monitoring data and field observations. A typical failure of the surrounding rock mass was observed, and data on the evolution of the microseismicity and its corresponding mechanism during the development of this failure was also obtained. The results provide an overall understanding of the progressive damage and failure evolution in large-scale underground openings. INTRODUCTION Layered rock masses are widely distributed on land. During formation of this kind of rock mass, differences in material composition, particle size, colour, and fabric result in rock stratification. Consequently, a set of dominant and parallel discontinuities, called bedding planes, is generated. As can be expected, these bedding planes play a vital role in the stability of engineering projects in layered rocks, and in particular in their modes of failure. Hence, the mechanics of the failure mechanism in stratified rock masses has its own distinctive features. Many large-scale underground openings have been constructed recently in stratified rock masses, due to their universal distribution. Therefore, it is of significance to study the failure modes and mechanism of underground openings in layered rock masses.

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.

Large tunnel and rock cavern spans that have failed for ‘geological’ reasons, or because of design errors, are the main focus of this presentation. The effect of adverse and sometimes unexplored geology will be illustrated. We will need to recognize that pre-investigations might miss some important detail, despite an exceptional frequency of core drilling. Usually, this is of minor importance and does not mean that failure will occur. Being wise after an event, which is always much easier, one might generally question why cross-hole seismics is not performed more frequently to obtain between-borehole information. Prior to construction, the need for this in a particular, but at that time unknown location, cannot of course be foreseen. It is apparent that really large failures seldom result from just one or two oversights, but are caused by a multitude of adverse factors working together. Small failures might be caused by designer or contractor short-cuts, or more likely by failure to log ‘today's’ conditions and react with more support. The really big failures are due to faulty design, and therefore inappropriate support, but sometimes just the multiple effects of some exceptionally adverse and unanticipated ‘geology’, aided by a remarkably adverse location, preventing the natural and needed arching. All or many of the following represent a rich assortment of possible reasons for failure underground: oversimplified design method assumptions that are erroneous, erroneous details of sub-surface topography, ignored details of surface topography, anisotropic low-strength jointing, the presence of unexpected discontinuities with adverse properties, and finally the use of deformable and weak temporary support like lattice girders. It seems that a really massive failure may involve about five or more adverse factors. With sufficient factors involved, fatalities may be a regrettable consequence, besides the huge economic losses. Surprisingly, whether or not a tunnel or cavern reaches this point of ultimate collapse might not depend on real-time interpretation of the instrument readings. This is because reaction to a new and exceptional rate of deformation may be too late when too many unknown adverse factors are already combined. A state of ‘guaranteed failure’ may be reached despite instrumentation warnings. The cavern and tunnel collapses that are generalized in the above paragraphs are specifically from the world of city metro, motorway ring-road tunnels, and hydropower caverns. They will be added to by reference to the largest open-pit slope failure to date, which showed the mechanism of progressive failure and an adverse pit shape, resulting in an absence of large-scale tangential stress. Failure was located in the ‘unstressed’ nose, and comprised 150 Mt of waste rock and ore. The tunnel and cavern failures include two 140 m tunnel collapses, two 35 000 m 3 progressive cavern collapses, and a 15 000 m 3 total collapse. A relative absence of sufficient tangential stress (arching stress) can be blamed in each case, and if the shear strength or designed support are in question, massive failure may result.

In open pit mining, a rock mass model can provide important input for the optimization of production in relation to stability challenges and a safe workplace. In this sense, rock mass classification systems are useful for defining rock mass domains that identify areas of poorer ground conditions that may require more detailed investigation or modifications to mine designs. These methods are widespread among geotechnical engineers as one practical way to assess quality of the rock mass. In open pit mine bench faces with no clear discontinuities, one single joint may dominate the pit slope plane failure mechanism. Analysis of low rock mass quality zones combined with the jointing orientation may provide an improvement in the assessment of instabilities. The aim of this paper is to develop a rock mass model of an open pit mine under active operation, based on extensive field mapping. Through this process, surface maps based on different rock mass classifications (RMR 89 , GSI, and Q-system) systems will be established for the mine. The mine was classified and assessed using RMR 89 . Finally, a three-dimensional rock mass model was developed based on the values obtained from surface and underground mapping. It was hoped that such a model allowed the identification of areas of the mine that are likely to have poorer ground conditions and may require more attention during the design and operational phases of the mine life. INTRODUCTION The planning of an open pit is a matter of determining the most profitable pit limit and the most economical mining sequence for a given orebody. In the exploitation phase, the excavation sequence implies the development of artificial slopes the stability of which is a crucial factor both for safety and economics of mining. Therefore, the assessment of rock slope stability is an essential requirement for the open pit mining industry, not exclusively during the engineering studies, but also throughout the operational life of a project. For this matter, rock mass classification systems are extensively used in order to quantitatively address the quality of the exposed rock mass in open pit slopes. A variety of rock mass classification systems have been proposed in the last 50 years, many of which were originally intended for tunnel design and subsequently adapted or modified for use in slopes. Four systems have gained broad acceptance in the mining industry: the RMR89 (Bieniawski, 1989), the Q-system (Barton, 1974), the MRMR (also known as Laubscher's RMR: Laubscher, 1990), and the GSI (Hoek, 1994). Flores and Karzulovic (2002) stated that the most widely used methods for rock mass classification in underground mines are Laubscher's RMR89, or MRMR, (53%), followed by Barton's Q (26%), and Bieniawski's RMR89 (15%). On the other hand, the most widely used method for rock mass classification in open pit mines is Hoek's GSI (39%), followed by RMR89 (26%) and MRMR (22%). This is shown in Figure 1.

In a previous study, the authors investigated the cable-strapping of pillars in room and pillar mines in order to solve local instability problems, by conducting laboratory tests. Based on experience in an old, small room-and-pillar haematite mine in Spain, the authors suggested a few steel cables placed around mesh-wrapped pillars as an effective means to improve pillar stability and control mine deformation. In the present study, and with the background of available laboratory uniaxial compressive strength tests on cabled samples and empirical knowledge of pillar behaviour, we present the results of numerical modelling of this type of pillar. Using FLAC and UDEC, it has been possible to develop models representing pillar response. These models add to the knowledge of pillar behaviour and improve our control of both standard and cable-strapped pillars, in terms of stress-strain evolution, plastic response, and the energy that the pillars are able to absorb. INTRODUCTION Standard practice, as well as published data on pillar design, showed that neither empirical formulae nor numerical modelling can fine-tune pillar design. However, either of these techniques (or better, both together) can be used as a preliminary approach to pillar design, provided monitoring is then carried out to control pillar performance (Malan and Napier, 2014; Esterhuizen, Dolinar, and Ellenberger , 2011). Rocks and rock masses tend to show variable behaviour and significant scale effects, so in a mine where a large number of pillars are generated in the same formation, a certain amount of variability in strength can always be expected. It is therefore not surprising that, even if by means of traditional approaches (tributary area and empirical formulae) and numerical models a design is reliable, in the mining practice (where pillars do not always have the prescribed dimensions and the rock has a variable nature) stability problems may arise – typically only in a few pillars, unless a very conservative approach was adopted.

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.

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.

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 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 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.