The behaviour of the liner-rock interface is an important property which determines the support characteristic of polymer-based rock support systems. The interaction between the liner and the rock substrate was defined using a traction-separation description of a cohesive zone model in ABAQUS. A cohesive zone model was developed in this study to simulate a double-sided shear test. The shear behaviour observed in the double-sided shear test of a polymer-plaster assembly was characterised by two stages, elastic behaviour of the shear bond before the peak stress and the gradual failure of the shear bond in the post-peak region. The initiation of shear bond failure was defined using a stress-based criterion. The post-peak behaviour was simulated by introducing a damage variable. An exponential damage law was considered to model the evolution of the damage variable from the initiation to complete failure of the shear bond. The parameters for defining the cohesive law were taken from experiments carried out at the University of Wollongong. The damage variable to define the damage evolution law was determined using back analysis. A parametric study was carried out with different values of the damage variable and the results obtained were compared against the experimental result. Finally, a three-dimensional numerical model was developed using the calibrated parameters. The failure behaviour of the interface in the double-sided shear test was simulated accurately, and evolution of the exponential damage variable was also explained. The results observed in the numerical model were found to be in agreement with the experimental result.

Abstract Two dimensional plain strain distinct element method has been used to analyze a curved slope under static and dynamic loading condition. Most vulnerable cross profile of the slope has been selected through kinematic analysis and slope mass rating specially devised for curved rock slopes. Static loading has been induced in the model to simulate conditions of the prevailing rock mass. Again dynamic loading has been performed according to the expected peak ground acceleration in the area in order to simulate the slope mass condition during earthquake. According to kinematic result and slope mass rating, the slope is more vulnerable towards 70° face direction. Static analysis results an unstable portion at the base of the slope. The whole slope is unstable under dynamic loading and is expected to threat the region with the risk of rock fall as well as failure of slope. 1. Introduction Failures in rock slopes are controlled by both rock material and discontinuities. In highly jointed rock mass with weak rock material, failure surface cuts through intact rock but in rock masses, with less number of joints and strong rock material, failure is controlled mainly by joint planes [1]. With the presence of joints, rock mass responses better to dynamic and static loading [2]. There are many numerical codes available to simulate dynamic loading in both continuous and jointed rock mass. Among them continuum method is well-known among geotechnical fraternity [3–5]. But despite of its astounding capability to handle any complex problem with ease and without any previous assumptions, this method lacks its compatibility in representing discontinuous medium [6]. A discontinuous medium is defined as an assemblage of discrete blocks separated by interfaces. Though continuum methods use special techniques (e.g. joint elements in FEM), they are only applicable for less jointed rock mass where little displacement is expected [7, 8]. To deal with inherent discontinuities in the rock mass and possible rupture in the intact rock, discontinuum methods prove to be best till date. Discontinuum methods are based on dividing rigid and/or deformable blocks by deformable interfaces. Among other discontinuum methods, distinct element method has an advantage of adopting both rigid and deformable blocks along with deformable contacts [9]. This method being explicit, directly solves the equations of motion by time marching scheme [10].

Abstract This paper deals with the study of stoping parameters vis-a-vis rock mass behavior in an underground manganese mine situated in central India. Strain bars are installed at various locations in the mine as part of rock mechanics instrumentation programme to monitor rock mass movement. The effect of change in stope height from 30 to 45 m with varying stope widths up to 20 m is evaluated through empirical approaches and numerical modeling. The instrumentation monitoring data is analyzed and results are corroborated with those from empirical and numerical modeling approaches. It is observed that an increase in stope height from 30 to 45 m do not adversely affect the stability of stope, cross-cuts and haulage roadway at this mine. 1. Introduction In the case of manganese mine reported in this paper, haulage roadways and cross-cut drifts serve a crucial purpose of transporting blasted ore from the working stopes. The haulage roadways and cross-cut drifts are driven in the hangwall rock mass at this mine. The mine management plans to increase stope interval from 30 to 45 m between 12 th Level and 13.5 th Level (Fig. 1). The mine management in association with CSIR-CIMFR, India has conducted the rock mass deformation monitoring through rock mechanics instrumentation programme. The programme involves installation and monitoring of strain bars at various cross-cuts in 10 th , 11 th and 12 th levels with a view to study the stability. The numerical modeling studies are conducted to assess the influence of various stoping parameters in terms of the strength factors, stresses and displacement developed around the stope excavation boundary. Parametric study is conducted through different numerical models to study the response of excavation geometry on the mining-induced stresses and displacements, strength factors (SF) within and around excavation boundary as well as surrounding rock mass etc. Instrumentation monitoring data is obtained and analyzed for evaluation of rock mass deformation. The results so obtained are compared with those obtained from the numerical modelling studies. The ultimate aim of this study is to achieve safer stoping environment with an increase in stope interval from 30 to 45 m to enhance production and productivity with safety.

ABSTRACT: Mine pillars are in-situ rock left after mining to ensure mine safety. The determination of pillar sizes dictates the cost of mining and mine safety. In this study, the rock fracturing progress analysis program, RFPA, was adopted to simulate pillars of various sizes to understand their failure mechanisms and patterns. The simulation results indicate that with increasing pillar size, the failure mode changes from tensile splitting to shear failure of the pillar with failure initiating from the pillar boundaries. Failure first occurs on the pillar surface and gradually develops toward the core of the pillar. Finally, shear failure occurs in the pillar. With increasing dip angle of the pillar, stress concentration first appears at the four corners, leading to micro-cracking. As the stress in the pillar continuously increases, micro-cracks gradually coalesce. Subsequently, spalling occurs and failure ensures at the pillar core. 1 INTRODUCTION In underground mining a pillar can be defined as "the in situ rock between two or more underground openings" (Coates 1981). As major structural elements, they serve to support the rock mass for safe and economic extraction of the orebody. Hence, it is important for the determination of pillar sizes to ensure acceptable cost of mining and safety. In spite of much research conducted for pillars design, pillar failure still poses safety issues and occasionally put lives at risk (Mortazavi et al. 2009). With the increase of mining depths, pillar failure becomes more frequent and critical due to the further increase in in situ stresses (Cook 1976). The failure process in pillars is a progressive as it starts with initiation cracks which then propagate with time and may eventually lead to failure (Fig. 1). The fracture process of a squat pillar begins with the initiation of local failure at random locations, and progresses through extension of these failed locations, coalescence of the extended sites, and slabbing or spalling of the rib sides, until development of large shear fractures ultimately occurs (Fang & Harrison 2002). On the hand slender pillars may fail suddenly or by gradual yielding depending on the amount of the imposed load, usually by tensile splitting.

ABSTRACT: Rock fracture mechanics has been widely applied to various fields of rock and civil engineering. But most researches covered mode I behavior, though mode II is dominant in rock engineering. Currently, there is only one ISRM suggested method for mode II toughness of rock. New mode II toughness test method was developed considering 1) application of confining pressure, 2) easiness of notch creation, 3) utilization of existing equipment, 4) simple test procedure. The stress intensity factors were determined by 3D finite element method considering two types of loading conditions. Tests were carried out using MTS rock test system with a loading rate of 0.002 mm/sec. Mean value of mode II fracture toughness of granite showed 2.33 MPa m 0.5 . Mode I toughness of the same granite was 1.12 MPa m 0.5 , determined by Brazilian disk test and K IIC /K IC = 2.08. The smooth fracture surface also supported mode II behavior of SCC method. 1 INTRODUCTION Rock fracture mechanics has been widely applied to various fields of rock and civil engineering. Until now, the research on rock fracture toughness has mainly focused on mode I fracture, though mode II or mixed mode is dominant in rock engineering. Currently, four methods are suggested to determine the mode I fracture toughness of rock by ISRM (Franklin et al. 1988; Fowell 1995; Kuruppu et al. 2013). Therefore, it is strongly required to establish a reliable fracture criterion for predicting mode II fracture and to develop an effective method for testing mode II toughness of rock (Rao et al. 2003). Various methods had been suggested for mode II toughness, but most of them used squared pillar, which requires extensive machining work (Whittaker et al. 1992). Recently PTST(Punch Through Shear Test) method (Backers et al. 2002) was registered as an ISRM suggest method. PTST method uses borehole core and confining pressure can be applied easily. One of the dis-advantages is the difficulty in notch making. This can be a hard work in case of subcritical crack growth test for it requires lots of specimens.

ABSTRACT: Since the 1999 Chi-chi earthquake and the 1999 Kocaeli earthquake damaged many important structures due to surface rupture, as well as a strong motion by the fault rupture, displacement and inclination in ground surface become the significant issues. In this study, we conducted fault rupture simulation using two and three dimensional finite element method for the 2014 Kamishiro Fault Nagano prefecture earthquake, which is a thrust fault type earthquake with an observed surface rupture of 9 km in length, and confirmed the applicability of numerical method and conditions, such as initial stress distribution, modeling of the fault plane and constitutive law, by comparing results with the observed ground motions and displacement. As a result, the acceleration response could not be simulated due to mesh size and constitutive law of fault plane. The displacement of 2D-FEM becomes larger than the actual displacement while the 3D-FEM yields results, which are in good agreement with the actual displacement. 1 INTRODUCTION Since the 1999 Chi-chi earthquake and the 1999 Kocaeli earthquake damaged many important structures due to surface rupture, as well as a strong motion by the fault rupture, displacement and inclination in ground surface become the significant issues. Generally strong motions are estimated by Green's function method and fault displacement is estimated based on geological surveys. However an earthquake occurs by rupture of earthquake source fault. When the displacement is large, it will come up to the ground surface as a surface rupture. Therefore, ideal analytical model is able to simulate a fault rupture process and estimate displacement and strong motion at the same time. Fault rupture simulations by Finite Difference Method (FDM), Finite Element Method (FEM) and Boundary Element Method (BEM) are carried out. However, they have not become practical as analytical results greatly vary according to initial stress conditions and modeling of fault rupture. In this study, we conducted fault rupture simulation using two and three dimensional finite element method (2D-FEM, 3D-FEM) for the 2014 Kamishiro Fault Nagano prefecture earthquake (M w 6.3), which is a thrust fault type earthquake with an observed surface rupture of 9 km in length. The analytical method and modeling of fault is dynamic response analysis considering fault rupture process proposed by Toki & Miura (1985) and Toki & Sawada (1988). We confirmed the applicability of numerical method and conditions, such as initial stress distribution, modeling of the fault plane and constitutive law, by comparison with the observed surface ground motions and displacement.

ABSTRACT: Fluid injection into the subsurface may trigger seismic events when shear stress acting on a fracture/fault plane exceeds its shear strength. During injection tests at the CO 2 storage pilot at Longyearbyen (LYB), Svalbard, seismic- and non-seismic slipevents were experienced. Two approaches are used to explain the seismic versus non-seismic events occurring at the site. An analytical approach, based on Mohr-Coulomb failure criteria, was employed to evaluate the slip tendency of faults. It shows that vertical faults are more likely to slip at deeper intervals where microseismic events were observed. At shallow injection intervals, where no seismicity was observed, horizontal faults are more likely to slip. The second approach, a velocity stepping direct shear test, showed that the friction coefficient of LYB shale (clay-rich) increases with increasing velocity for bedding parallel faults. This experiment shows that the shale present at shallow interval is prone to velocity-strengthening behavior, which may explain the non-seismic failure during injection tests. 1 INTRODUCTION Capturing carbon dioxide from source points such as power plants and refineries and storing in the subsurface formations is suggested as one of the main measures to mitigate atmospheric greenhouse gases. The Longyearbyen (LYB) CO2 storage pilot is a potential site for CO 2 storage on Svalbard, Norway. Longyearbyen is a small city with a population of about 2000 people in the polar wilderness of central Spitzbergen-the main island of the Svalbard archipelago at the northwestern margin of the Barents Sea Shelf (Braathen et al . 2012). The city has a 10 MW coal combusting power plant, which emits about 64,000 tons of CO 2 annually (Ogata et al. 2012). The LYB pilot project was initiated by the University Centre on Svalbard (UNIS) to explore feasibility of the site to store the produced CO 2 and convert the Longyearbyen to a carbon neutral community. The site is located 5 km east of the Longyearbyen city and eight boreholes (Fig. 1) have been drilled to study properties of the overburden, cap rock and reservoir.

ABSTRACT: The primary aim of this paper is to review a recently proposed shear behaviour model of infilled rock joints in which joint surface characteristics and the properties of the joint and infill materials has been taken into account. The empirical model has also finally validated based on a few experimental data from the literature. The results of that study showed an acceptable confidence level for the model and reported that the new model successfully describes the observed shear behaviour of natural infilled rock joints. However, further experimental investigations were required to be undertaken to modify the proposed model and increase the field applicability. In addition, although some steps have been taken by the authors to replicate and conduct tests on natural joint profiles, there are still limitations caused by the narrow range of JRC examined in that proposed model. Therefore, to provide a large spectrum, the authors started to investigate several evaluations on other good quality laboratory shear tests and large scale rock joint surfaces from a radwaste repository's underground caverns in Hungary with granitic host rock. Based on these investigations, which are currently under execution, the model will become more accurate and quantitative (especially the roughness profiling). 1 INTRODUCTION Over many years, fine sediments resulting from weathering and other surface processes could subsequently ingress to rock joints, reducing the overall shear strength of the joint surface (Barton 1974 & Hoek 1983). Rock joints that are naturally filled with fine materials are likely to be the weakest elements in a rock mass and can have a dominant influence on its shear behaviour due to of the low frictional properties of the infill (Ladanyi et al. 1977). The most effect of filling material is to separate the discontinuity walls and thereby reduce intact rock contact, but shear strength will also be influenced by the nature of the filling material itself and the characteristics of the wall-fill interfaces. Because of the lack of reliable and realistic theoretical or empirical relations and the difficulties in obtaining and testing representative samples, engineers generally rely on judgment, often considering the shear strength of the infill itself to be conservative. In critical cases, in situ tests may be carried out to provide site specific design criteria, but invariably amount of testing that can be undertaken precludes the establishment of fundamental relations. During the past 30 years much more information has become available on the shear behavior of joints infilled with soil material. Several models have been proposed to predict the shear strength of infilled joints under both constant normal load (CNL) and constant normal stiffness (CNS) boundary conditions, considering the ratio of infill thickness (t) to the height of the joint wall asperity (a), i.e., t/a ratio (Ladanyi et al. 1977 & Leichnitz 1985). The experimental researches to date have tended to focus on modelled joints or replicas rather than natural rock joints. In other words, although some works have been done on the effect of infill material by using natural rock joints, they have not proposed prediction models and it can be noted that most of the previous models in the literature have been developed based on the laboratory tests over the simulated artificial joints and not on the real rock joints (for simplicity and reproducibility reasons). In addition, there have been no shear behaviour models which consider differences in infill type within the rock joint. However, this research will give an account of application of a statistical analysis on a series of data obtained from a complete testing program on natural infilled rock joints, taking into account three different material filling the rock joint. The output of this analysis has consequently been conducted to propose an empirical model. The tests in this research have been done in constant normal load (CNL) conditions. In general, the CNL condition is more realistic for shearing planar interfaces where normal stress applied to the shear plane remains relatively constant, and the problem of surface (shallow) slope stability. The development of shear resistance is a function of constant normal stiffness (CNS), and based on the opinion of some researchers (e.g. Oliveira et al. 2009 & Indraratna et al. 2012 & 2013), the use of CNL test results leads to underestimated shear strengths because the surrounding rock freely allows the joint to shear without restricting the dilation or there is no dilation during the shearing process, thereby keeping normal stress constant during shearing process. To evaluate the behaviour of rock joints under conditions more commonly encountered in underground excavations, it is necessary to simulate the stiffness of the rock mass normal to the direction of shearing (Leichnitz 1985 & Morris 2003). However, this condition is only existent in very deep situations, and the consideration of CNL conditions would, therefore, be adequate for an available model and the tests and resulted model proposed in this paper is therefore suitable to be utilized in shallow rock structures and slope stability applications.

Abstract On 27 December 2007 a M w 1.9 seismic event occurred in the Mponeng gold mine, South Africa. The seismic event's hypocentre, as well as the fore- and aftershocks were located with the JAGUARS acoustic emission array, placed at mining level 116. Thus, hypocentre depth (3509 m) and focal mechanism are very well constrained. Since no mining activity took place more than two days before the event, dynamic triggering due to blasting is ruled out. We investigate the hypothesis that static stress transfer due to excavation of the gold reef induced the event. To estimate these mining-induced stress changes, we set up a small scale (450×300×310m 3 ) high resolution 3D geomechanical-numerical model, containing most of the seismic events detected. The model is made of the four different rock units present in the mine: quartzite (footwall), hard lava (hanging wall), conglomerate (gold reef), and diorite (dykes). The virgin in-situ stress, determined in nearby TauTona mine, is used to implement initial conditions. Elastic rock properties are taken from laboratory measurements. For the numerical solution, we are using the finite element method, with a discretised mesh of ~1 million elements. From the computed 3D stress tensor we estimate the Coulomb failure stress change (ΔCFS) on the fault plane of the event in two ways: The first with shear stresses resolved in rake direction (ΔCFS slip ) and the second using the maximum shear stress (ΔCFS max ) on the plane. The ΔCFS slip values are negative, indicating that, according to the model, mining activity brought slip in the resolved direction on the resolved fault plane further away from failure. Considering model uncertainties with respect to fault plane geometry, rake vector, and elastic parameters, this result remains stable. In contrary, the ΔCFS max value is positive, indicating that, given the modelled stress is correct, a high potential for triggering a rupture is available.

Abstract In this paper, Perzyna-type viscoplastic rate equations are integrated for a time step by considering the step as stress-driven. Depending on how the increment is imposed (constant, linear etc.), different strategies arise. The secant compliance is obtained by truncated expansion of the yield function. The viscoplastic model can be applied to materials exhibiting rate-dependent behavior, but it can also be used to recover an inviscid elastoplasticity solution when stationary conditions are reached. Within this framework, a viscoplastic relaxation iterative strategy is developed, relating the iterations with the fictitious time steps. Some examples of application are presented in the context of the Finite Element Method with zero-thickness interface elements for slope and stability problems with discontinuities. 1. Introduction Viscoplasticity has been widely used for engineering materials with (physical) time-dependent behaviour over a treshold stress level (Cormeau, 1975, Zienkiewicz and Cormeau, 1974), or in the context of viscoplastic relaxation strategies to obtain the stationary solution of an inviscid problem via a fictitious (non-physical) pseudo-time (Underwood, 1983, Zhang and Yu, 1989). In either case, the rate-type infinitesimal viscoplastic formulation requires a time integration strategy to a) discretize time in increments b) evaluate a linearized relation between stress and strain increments for each time step and, possibly, some residual force calculation and iterative strategy. Typically, the algorithms are based on the initial stress scheme used in FEs, in which the strain increments are prescribed to the constitutive equations. A variety of such algorithms has been proposed since the original constant stiffness and constant stress procedures (Hughes and Taylor, 1978, Peirce et al., 1984, Peric, 1993, Simo, 1991) to more recent and sophisticated contributions (Alfano et al., 2001, Betegon et al., 2006, Lorefice et al., 2008, Ponthot, 2002). In contrast, stress driven schemes are not that common (Caballero et al., 2009, Zienkiewicz and Cormeau, 1974). Their implementation can be numerically advantageous (explicit integration of the constitutive equations and simple coding) but their stability is strongly related to the size of the time step and, as a consequence, if a small time step is required an enormous amount of iterations are required.

Abstract Shear and deformation behaviour of infill rock joints under constant normal stiffness conditions have been studied much less despite their natural occurrence, because of difficulty in physical, experimental, analytical and numerical modelling due to increased number of variable parameters. Due to unavailability of reliable and realistic results or models and the difficulties in laboratory testing of representative samples, engineers generally rely on judgment, often considering the shear strength of the infill itself as shear strength of rock joints. This assumption not only leads to uneconomical but also sometimes unsafe design of underground structures, slopes, rock socketed piles and foundations. In the present work influence of thickness of infill material on shear behaviour of joints are studied by conducting tests on the modelled rock joint of plaster of Paris having the asperity 30°–30° under constant normal load (CNL) and constant normal stiffness (CNS) boundary conditions at different initial normal stress ranging from 0.10 MPa to 2.04 MPa. The thickness of infill (t), varied from 0, 5, 7 and 10mm over an asperity height (a) of 5mm, which makes the t/a ratio as 0, 1, 1.4 and 2. The test results indicate that shear strength decreases with increase in t/a ratio for both CNL and CNS conditions, bur reduction in shear strength is more for CNL than CNS condition for a given initial normal stress. 1. Introduction 1.1 General The shear behaviour of rock mass is different than that of intact rock. Because it contains discontinuities of the geological origin in the form of joints and faults, which completely dominate the shearing and sliding behaviour of in situ rock mass under given stress conditions. Rock joints are seldom clean they are normally filled up with varying thickness of the fine sediments resulting from weathering of the joint walls of parent rocks (gouge) blended with transported sediments. Most of the infill materials transported by the water are either cohesive i.e. clay, silty clay or frictional i.e. silty, silty sand. These infill form the weakest planes, from where sliding or shearing of the rock joints initiate and reduce the overall shear strength of the joint. The shear behaviour of the rock joint can be investigated in the laboratory under constant normal load (CNL) and varying normal load i.e. constant normal stiffness (CNS) boundary conditions. The shearing under CNL boundary conditions is suitable for situations where the surrounding rock freely allows the joint to shear without restricting the dilation, thereby keeping normal stress constant during shearing process. This mode of shearing is only applicable for planar rock joints and non-reinforced rock slopes where normal load is constant during the shearing process. The conventional direct shear apparatus is used to investigate the shear behaviour of rock joints under CNL boundary conditions.

Abstract High-level radioactive waste repositories are to be designed to ensure long-term stability, and it is necessary to consider various effects at deep depth in various aspects. Especially, many uncertainties are involved in characterizing the behavior of deep rock mass in the designing process and interactions among groundwater flow, increased temperature induced by radioactive wastes and stress disturbance due to excavation should be considered. In this study, to investigate the shear characteristics of a saw-cut smooth surface of crystalline rock under various thermal-hydro-mechanical conditions, multistage triaxial tests were conducted on granitic rock. After testing, shear characteristics of the saw-cut surfaces were analyzed based on Mohr- Coulomb failure criterion. From the experimental results, it was observed that the shear characteristics of the granitic rock were sensitive to confining pressure and water pressure but not to temperature variation. Numerical analysis was conducted as well by COMSOL providing with a good agreement with experimental results. 1 Introduction High-level radioactive waste disposal facilities are usually designed to be stable for an extended period of time as long as 10,000 years. To ensure the mechanical, hydrological and chemical stability of the facilities, long-term performance is assessed with respect to many different variables and their coupled effects. It is well-known that the international cooperative project DECOVALEX has been accumulating lots of information and knowledge of coupled processes in fractured rock masses. In addition many independent studies were carried out to investigate the T-H-M-C coupled effects in numerical and experimental methods. They include the researches by Lockner et al. (1982), Elliott & Brown(1988), Zhao & Brown(1992),Yoon & Jeon (2004) and Wanne & Young (2008). However, few studies were carried out on the shearing behavior of a shear surface of crystalline rock under various thermal (T), hydro (H) and mechanical (M) conditions. In this study, a series of laboratory scale experiment was carried out to investigate the behavior of smooth shear surface of a crystalline rock under triaxial compression to simulate the stress condition at depth, different water pressure and temperature conditions to simulate the situation at the vicinity of a canister of high-level radioactive waste. The test results were analyzed using Mohr-Coulomb failure criterion. In addition, numerical analysis was carried out using COMSOL Multi physics to simulate the test.

Abstract Hydroelectric power projects are pollution free and renewable sources of energy. These projects involve the construction of dams across the flowing stream ofwater. Construction of dams in rock regions requires stability analysis of the foundation and the supporting rocks at the banks of the reservoir area. Stability of the concrete foundation over the rocky strata and stability of the banks depend upon the shear strength of the concreterock and rock-rock interfaces, respectively. Therefore, evaluation of shear strength parameters (cohesion and friction angle) at rock-rock and concrete-rock interface is the most important primary step for stability analysis and design of dams. Determination of shear strength parameters involves the in-situ tests in drifts on the dam axis site and interpretation of test results. In view of the above, site-specific shear strength parameter interpretation is carried out for Amochu hydroelectric power project site, Bhutan. This project envisages construction of 175m high concrete gravity dam across river Amochu. The existing rock type on both left and right banks of the site is observed to be weathered Phyllites. From the analysis of 20 field tests in 2 drifts, it has been found that for rock-rock interface shear strength values are approximately 50% higher at the right bank as compared to the left bank. On the other hand, for concrete-rock interface, this variation is around 10% only. Therefore, seeing the large variation of estimated shear strength in rock-rock interface, it is proposed to use individual bank stability analysis whereas, marginal variation in concrete-rock interface suggests the use of the combined analysis of the data from both the banks for foundation stability analysis. 1 Introduction The design of foundations of dams in rocky strata includes determination of bearing capacity, settlement analysis and sliding/slope stability analysis. Bearing capacity and settlement analysis involve the ability of the rock foundation to support the imposed loads without shear failure and without excessive settlements. Sliding stability analysis involves the ability of the rock foundation or slope to resist the imposed loads without shearing or sliding. Both analyses must be coordinated and satisfied in the design. For performing theses analyses shear strength parameters viz. cohesion and friction-angle are used as input parameters. There are two major categories of analysis, a) Analysis at the concrete foundation and rock joint interface and b) Analysis at the rock-rock interface. Henceforth, determination of shear strength parameters is required for the concrete over rock (C/R) interface as well as rock over rock (R/R) interface on the project site. Former is for stability analysis of concrete dam resting over rocky strata while later is concerned with the ability of the surrounding rocks to hold the reservoir.

Abstract It is fundamental to understand the process of crack initiation and propagation of intact rock to clarify the coupled hydro-mechanical properties due to the stress change in underground. We have been working on the study of coupled shear-flow properties on sedimentary rock using pumice tuff, Japan. The changes of shear stress, normal stress, flow rate and process of fracturing on sample surface have been investigated by using a new type coupled shear-flow test apparatus. In this study, we designed a numerical model for the coupled shear-flow test using a discrete element code. We report the comparison between the experimental results and the simulation results. The porosity changes of model showed good agreement with the flow rate behavior of the test. 1 Introduction New fractures due to underground excavation or fault cause mechanical and hydrological problems in underground (Tsang et al., 2005, Mitchell and Faulkner, 2009). The fractures are generated through a variety of mechanisms with various scales (Bossart et al., 2002, Scholz et al., 1993). For the research of underground disposal or geological evaluation, it is necessary to investigate the coupled hydro-mechanical behavior of rock (Olsson and Barton, 2001, Li et al., 2008). We conducted an experimental study on coupled shearflow properties of sedimentary rock, and designed a numerical model for the coupled shear-flow test using a three-dimensional simulation tool, PFC3D (Particle Flow Code 3D , Itasca Inc.). The purpose of this study is to examine the performance characteristics of the numerical model. The test system and model are briefly introduced, and the results are compared focusing on stress, permeability and crack propagation. 2 Particle Flow Code 3D PFC3D is classified as a discrete element code (Cundall and Strack, 1979, Potyondy and Cundall, 2004) based on the definition in the review by (Cundall and Hart, 1992). It allows finite displacements and rotations of discrete bodies (including complete detachment), and recognizes new contacts automatically as the calculation progresses. PFC3D can be considered as a simplified implementation of the DEM because of the restriction to rigid spherical particles.

Abstract: Recent regulations introduced a limit state approach to geotechnical design by using representative values of the actions and of the strength parameters, partial safety factors that affect them, and by including safety margins in the calculation models. Moreover, Eurocode 7 stresses the importance of making use of test results for establishing ground parameters, and so rock joint shear tests are set to play a relevant role in the assessment of the shear strength required in the design of important projects, such as concrete dams, large bridge foundations, slopes, or underground excavations. In this keynote, joint shear tests are described, along with a presentation of their equipments and different procedures. Test results and calculations for the assessment of relevant shear strength parameters are illustrated, and several topics regarding sampling and variability are discussed. Opportunity is taken to present a practical apparatus allowing to perform simple shear tests (push tests) under very low normal stresses with advantages over tilt or pull tests. 1 INTRODUCTION Certain projects, such as concrete dams, rock slopes, or relatively shallow under-ground works, require the design of geotechnical works in rock masses where stresses are low when compared with the intact rock strength. In these cases, stability is structurally controlled by the shear strength of kinematically unfavourable rock discontinuities (joints, bedding planes, shear zones, faults, and cleavage or foliation planes). The analysis of this type of limit state requires the estimation of the shear strength of the rock joints, which is usually done by means of shear tests (Goodman 1976; Hoek & Brown 1977, Muralha 2007). Though Eurocode 7 is intended to be applied mainly to common civil engineering works, it establishes a comprehensive framework for the design of any kind of structure, such as underground caverns, tunnels, slopes, and dam or large bridge foundations.

Abstract: The paper describes a laboratory testing programme focussed on determining the effects of suction and roughness on the mechanical behaviour of rock joints. Tests were conducted under varying suction, normal stress and asperity roughness angle. Test results were analyzed by means of a hyperbolic yield surface. Constitutive parameters (apparent cohesion and residual strength) were found dependent on roughness as well as suction. Mathematical expressions for calculating the strength parameters were proposed. These developments allowed the formulation of a novel elasto-plastic constitutive law of rock joints. This law was implemented as an interface joint element into the finite element program Code-Bright. Finally, numerical simulations of tests were carried out. 1 INTRODUCTION The effect of suction on the mechanical behaviour of rock joints has not been reported in the literature to the author's knowledge. Since the prevailing suction has a very significant effect on rock strength (Oldecop & Alonso, 2001) it was anticipated that rock joint behaviour should be affected. This was the motivation for the performance of a laboratory testing programme concentrated on the mechanical behaviour of rock joints subjected to direct shear under suction control. Suction was controlled by a vapour equilibrium technique (Fredlund & Rahardjo, 1993; Romero, 2001). Artificially prepared joints of Lilla claystone were tested. Joint roughness of varying intensity was created by carving the surfaces in contact in such a manner that rock ridges of different tip angles were formed. These angles varied between 0° (smooth joint) to 45° (very rough joint profile). The geometric profiles of the two surfaces in contact were initially positioned in a "matching" situation. It takes into account the effect of suction on the strength parame-ters and the degradation of rock joints. The performance of the model is checked against the recorded shear stress-relative displacements.

ABSTRACT: This paper describes the validity of the Multiple Yield Model to the earthquake response analysis of the discontinuous rock by two examples of the actual collapsed rock slope caused during the Mid Niigata Prefecture Earthquake in 2004 (M 6.8) and the large scale and/or very important facility like as a nuclear power station on rock. Results show that geometrical and mechanical conditions of joints have strong influence on the seismic response. 1 INTRODUCTION The earthquake response analysis of rock mass is generally performed as elastic body. However, rock masses are essentially discontinuous, which have a lot of joint, fissure and others. The mechanical behavior of rock mass is strongly controlled by the mechanical property and geometrical conditions of discontinuities such as orientation, dip angle, spacing, number of set. The dynamic response analysis of the equivalent continuous rock with discontinuities which are depended on confining stress and have strong nonlinearity, have not been practicable due to use the direct integration. But we could be thought that the constitutive equations of joint are improved from the results of dynamic cyclic direct shear test and the technique of calculation ability has been developed rapidly, therefore dynamic analysis becomes now a practicable stage. We describe the results of two cases of earthquake response analytical results by the Multiple Yield Model (hereafter, it is abbreviated as MYM) (Sasaki et al. 1994). 3 EXAMPLES OF EARTHQUAKE RESPONSE ANALYSIS BY MYM 3.1 The collapsed slope analysis in Mid Niigata prefecture Earthquake (M 6.8), Japan, in 2004 Due to the earthquake, a large number of collapses and land-slides occurred at slope of folded mountain range which consists of the Neogene sedimentary rocks. Fig.1 shows the geological and analytical conditions at the collapsed site from focal distance about 8Kmused in this study.

ABSTRACT: Mechanical rock properties affect borehole pressure and permeability during production. The objective of this paper is to estimate and predict potential borehole failures and form a geomechanical model detailing the most likely areas to fail under drilling, completion or borehole collapse stress. A model of the fracture patterns of the area will be made and compared with the structural geology of the zone. This project uses data found in petrophysical logs for the study of geomechanical properties, later assisting in geomechanical correlations throughout the reservoir. The elastic and mechanical constants ofYoung modulus (E), Poisson ratio ( v ), Shear modulus (G), and Bulk modulus (K) were derived to determine weak and strong zones. The fracture initiations were then predicted and compared for each weak and strong zone using the Mohr-Coulomb criterion. It was also found that an increase of vertical stress (σv) induced due to drilling causes failure by 200MPa in the strong zones. The results also show how the reduction in the σmin (minimum horizontal stress) between 0.47MPa and 25MPa causes normal failure. Geomechanical failure risk can be linked to the thickness and pattern of the fractured zone. The fractured zone thickness is larger in the northern part of the field compared to the south-east. In conclusion, geomechanical failure risk related to fracture is reduced to the south of the structure. 1 INTRODUCTION This paper focuses on a carbonate reservoir of the Middle East. As the work is being done on a limestone reservoir it is predicted to encounter mostly brittle and fractured rocks that are classified as strong zones (SZ); while more ductile rocks are in zones called weak zones (WZ). Therefore knowing the zone properties, as well as the stresses required to cause failure can allow more secure drilling operations in the future.

ABSTRACT: This paper presents the results on the size effect of a tensile fracture in granite on the shear behavior determined by direct shear tests with a normal stress of 10MPa. Normal and shear stresses were unloaded at designated shear displacements, and the change in the surface topography was measured by using a non-contact surface profile measurement system with a laser profilometer, to determine the evolution of surface damage and aperture during shear. The standard deviation of the initial aperture of the sheared fracture significantly increases with both shear displacement and fracture size, resulting in an increase in the non-linearity of the closure curve with shear displacement. Furthermore, while the shear stiffness of the fracture in the initial stage decreases with fracture size, that in the residual stage increases with fracture size. In addition, it was shown that the damage zones are enlarged and localized with shear displacement and tend to form perpendicular to the shear displacement. 1 INTRODUCTION Over the past decade, considerable efforts have been made to better understand the mechanical behavior of rock fractures since it governs that of rock mass. Fractures in rock mass range in scale from a microfissure to a large fault system. Therefore, the size effect on the mechanical behavior of a fracture has to be considered when estimating the mean aperture and the hydraulic conductivity of a fracture under stress. Giwelli et al. (2009) have shown that the closure of a fracture at a given normal stress significantly increases with fracture size. This effect can be attributed to an increase in the standard deviation (SD) of the aperture with fracture size, since the SD of the aperture gives a measure for the matedness of the fracture. Accordingly, fracture size may affect the shear behavior of a fracture.

ABSTRACT: In order to evaluate the applicability of Acoustic Emission (AE) as an indicator of instability of active joints in dam structures, an extensive feasibility study was done on 40 joint samples. To this end, direct shear tests were conducted under different conditions (in various normal loads and displacement rates) and different joint characteristics (with various roughness and bonding percentages) and AE signals were acquired using attached sensors to the samples. In this study two methods were used to verify the effect of normal load as one of the most important parameters affecting shear behavior and consequently generated AE signals during direct shear test of rock joints. In the first method several samples were tested under normal stress of 0.5, 1 and 2MPa respectively. In the second method, normal stress was changed from 0.5 to 2MPa in the residual section of the same sample. The results showed that when normal load increases, the AE signals show a more significant peak in maximum shear stress point. Increasing and decreasing normal load during residual section make significant changes in corresponding AE signals. 1 INTRODUCTION Previous researches have addressed application of AE for monitoring rock joints both in laboratory (Moradian et al. 2008, Hong & Seokwon 2004, Li & Nordlund 1990) and in site (Sasao et al. 2003, Shiotani et al. 2001). Among several parameters affecting shear strength of rock joints, the magnitude of the normal stress has the most important role (Barton & Choubey 1977). Any change in this parameter makes a significant change in shear behavior and, consequently, in the generated AE signals (Li & Nordlund 1990, Filimonov 2005). In order to apply AE technique for stability monitoring of rock joints successfully, it is necessary to understand the effect of normal stress onAE activity in rock joints.