In rock tunnels in regions with colder climates, the load-bearing structure, including the rock and the reinforcing elements, is exposed to repeated destructive freezing and thawing cycles during the winter. If water accumulates in cracks or in the interface between rock and shotcrete, frost shattering may occur. If there is adequate adhesion between the rock and shotcrete, degradation of the shotcrete as a reinforcement element due to frost shattering should not present a problem. However, if adhesion is poor, a small void will form between the rock and the shotcrete where water can accumulate. If the water in these voids is subjected to freeze-thaw cycles, ice will develop, thus exerting pressure on the interface and causing the shotcrete to crack and degrade. In tunnel sections with complex water conditions, for example, relatively water-bearing open joints and weak zones, the adhesion of the shotcrete and its stability and reinforcing effect may be strongly affected when exposed to freezing temperatures. This article describes a laboratory study that comprised freeze-thaw tests on shotcreterock panels with the objective of studying how water migration affects the growth of ice and the ice pressure in the shotcrete-rock interface to better understand the degradation of the reinforcing effect of shotcrete

Despite extensive grouting efforts to prevent water from leaking into tunnels, water seepages remain. When exposed to freezing temperatures, ice formations occur. During the winter, the Swedish Transport Administration's railway tunnels are affected by major problems caused by ice, such as icicles from roof and walls, ice loads on installations, ice-covered tracks and roads, etc. To ensure safety and prevent traffic disruptions, many tunnels require extensive maintenance. Improved knowledge about frost penetration in tunnels is required to reduce maintenance of the tunnels. Frost insulated drain mats are often used at leakage spots to prevent ice formation along the tunnels. To find out which parts of a tunnel are exposed to freezing temperatures, the University of Gävle and the Royal Institute of Technology in Stockholm conducted a laboratory model test on behalf of the Swedish National Rail Administration (now the Swedish Transport Administration). The laboratory model test aimed to find a method to determine the expected temperature conditions along a tunnel to decide which parts of the tunnel require frost insulation to protect the drainage system from freezing and prevent ice formation. To evaluate the laboratory model test, the Swedish Transport Administration in collaboration with Luleå University of Technology have performed field surveys in two Swedish railway tunnels. The field measurements involved monitoring temperatures in air, rock surfaces and rock mass, as well as measuring wind direction, wind and air velocity and air pressure. The measurements in the tunnels show that the frost penetrates further into the tunnels than was expected from the laboratory model test, which was based on a completely uninsulated tunnel. Frost insulated drains do not only prevent the cold air from reaching the rock mass, but also prevent the rock from emitting geothermal heat that warms up the cold tunnel air. Consequently, the frost penetrates further into the tunnel than it would do if the heat from the rock mass was allowed to warm up the outside air on its way into the tunnel. The number of frost insulated drains and how much of the tunnel walls and roof are covered thereby affect the length of the frost penetration.

During the winter season, ice causes major problems in many Swedish railway tunnels. Ice, rock and shotcrete in the roof and on the walls may come loose and fall down, installations and cables can break due to ice loads and the tracks can become covered with ice. To maintain safety and prevent traffic disturbances, many tunnels require frequent maintenance. The removal of ice, loose rock and shotcrete is expensive and potentially risky work for the maintenance workers. To reduce maintenance costs, it is important to improve our knowledge of frost penetration inside tunnels and investigate the effect of ice pressure and frost shattering on loadbearing constructions. The aim of this investigation was to gather information about the problems caused by water leakage and its effect on the degradation of a rock tunnel when subjected to freezing temperatures. There are many factors that determine whether frost or ice formations will appear in tunnels. To collect information on ice formation problems, field observations were undertaken in five of Sweden’s railway tunnels between autumn 2004 and summer 2005. For one of the tunnels, follow-up observations also took place in March during the years 2005, 2006 and 2007.

Even though extensive pre-grouting is carried out during the construction of tunnels, certain leakages and drips remain. These remaining leakages are remedied by a combination of post-injection and drainage measures with, for example, frost insulated drain mats, whose function is to prevent the cold tunnel air from reaching a leakage spot and causing water to freeze. Despite these measures, some water may still enter the tunnels and cause problems during winter with ice formations and frost shattering. Icicles, ice pillars and ice-covered roads and railway tracks require constant maintenance. If ice occurs in the fracture network close to the tunnel contour or in the interface between the rock and shotcrete, it can cause degradation of the load-bearing capacity of the tunnel and fall-outs of both materials. In tunnel sections with water leakage problems it is common to protect the load-bearing structure from freezing with insulated drainage systems. To determine where along the tunnel efforts must be made to prevent ice formation, the temperature conditions of tunnels must be investigated. This article presents parts of the results from field measurements in two Swedish railway tunnels. The measurements involves monitoring of air and rock temperatures, air pressure and air velocity.

The sublevel caving used in Kiirunavaara mine induces failure and subsidence of the hangingwall. Two sections of the mine were studied by means of numerical analyses. Numerical models were developed using finite element and discrete element codes. The former was applied to calculate the location of new failure surfaces in the hangingwall and to estimate the break angle when mining advances downwards. The latter was used to analyse the displacement path of the caved rock during draw and to determine its effect on the stability of the hangingwall and footwall. The models were calibrated using displacement monitoring data. The finite element analyses indicated that the break angle is almost constant for deeper mining levels but may change if the geometry of the orebody changes. The discrete element model showed the formation of a stationary zone along the footwall that reduces the magnitude of the shear forces during draw, increasing its stability.

Numeriska analyser av bergmassans beteende under fyllningsdammar för vattenkraftproduktion.

Sweden has more than 190 large hydropower dams, of which about 50 are pure embankment dams and over 100 are concrete/embankment dams. This paper presents results from conceptual analyses of the response of typical Swedish rock mass to the construction of a hydropower embankment dam and its first stages of operation. The aim is to identify locations and magnitudes of displacements that are occurring in the rock foundation and grout curtain after construction of the dam, the first filling of its water reservoir, and after one seasonal variation of the water table. Coupled hydro-mechanical analysis was conducted using the two-dimensional distinct element program UDEC. Series of the simulations have been performed and the results show that the first filling of the reservoir and variation of water table induce largest magnitudes of displacement, with the greatest values obtained from the two models with high differential horizontal stresses and smallest spacing of sub-vertical fractures. These results may help identifying the condition of the dam foundation and contribute to the development of proper maintenance measures, which guarantee the safety and functionality of the dam. Additionally, newly developed dams may use these results for the estimation of the possible response of the rock foundation to the construction

The rockfall risk due to mining-induced seismicity reduces by installing appropriate rock support to absorb the kinetic energy from a seismic event, which is calculated by assuming the mass of ejected rock and its ejection velocity. Estimation of ejection velocity is normally based on scaling laws that do not consider the effect of the excavation free-surface and existing fractures near the excavation free-surface. Field monitoring studies have shown that the peak particle velocity on the free-surface can be much larger than the velocity in deep solid rock. The interaction between the fractures and the free-surface under incident S-wave is investigated by using a two-dimensional UDEC model with fractured zone characterized as one, two, three and four sets of parallel fractures with varied intersecting angles. The results show that wave amplification factor varies according to the incident wave angle, the number of fracture sets and fracture spacing.

When rock support is designed in a seismically active underground mine, it is important tochoose the right ejection velocity and calculate corresponding kinetic energy. Field monitoringand back-analyses have shown that ejection velocity of the order of 10 m/s and higher can resultfrom seismic events of moderate magnitude. Such velocities are much higher than those predictedusing peak particle velocity (PPV) obtained from scaling laws. Many researches have reportedthe amplification of particle velocity near excavation surface. Velocity amplification of P-wavetravelling through fractured rock near a free surface was recently studied. The amplification ofseismic waves on the skin of excavation is of interest in case of large seismic events. Seismic eventswith large magnitude are often associated with slip along weaknesses or shear fracturing of intactrock, which according to observations radiate much stronger S-wave as compared to P-wave.In this paper, velocity amplification of S-wave was investigated by modelling the dynamicinteraction between fractured rock and a free surface using a 2D discontinuum-based numericalprogram, UDEC (Universal Distinct Element Code). A 1D model with a fractured zone wasused to represent the fractured rock in this investigation. It is found that the shear stress ratio,wave frequency, fracture stifness, fracture spacing and thickness of fractured zone afect thevelocity amplification, in which the shear stress ratio is the most crucial factor influencing wavetransmission. The results have proved that the interaction of the seismic wave and multiplefractures near the free surface strongly influences the ground motion.

As the underground mining extends gradually towards depth, more and more seismic events induced by fault slip occur and cause great damages, which have become a severe potential threat to mining safety. In view of the plane strain problems, through the three dimensional discrete model established, comparison and analysis was carried out between the equivalent calculation of plane strain in 3D model and a 2D discrete model. The results have shown that the research model developed to simulate the propagation of seismic wave in 3D is feasible and applicable. The study of the effect of radiation pattern on seismic propagation revealed and tested the direction of P-and S-wave propagation, which presents high consistency to double couple model of the fault slip. On this basis, the comparison with the design scaling law formulas proposed by Kaiser and associates finds that the existing design scaling law does not totally satisfy the demand of practical engineering. Numerical calculation and analysis with the three dimensional discrete model can further optimize the support parameters, provide better service for system design of mining support, and ensure the safety and high efficiency in mining.

Field monitoring carried out in a deep tunnel of the Dingji coal mine in China confirmed the zonal disintegration phenomenon by using the borehole TV. Based on field monitoring, an analogical model test was conducted to research the fracture shape and forming conditions of the rock mass in the Dingji mine. To perform the model test, an analogical material and optical sensor were developed independently. Through the test, the occurrence of zonal disintegration was confirmed and the forming process was monitored. The fracture pattern of zonal disintegration was determined, and the radii of the fractured zones were found to fulfil the relationship of geometric progression. The displacement laws of surrounding rocks during zonal disintegration were obtained and found to be non-monotonic. The test results are in agreement with the field-monitoring results. Through a theory analysis based on fracture mechanics, the mechanism of zonal disintegration was revealed. The fracture zones occur as circles concentric to the cavern periphery, which is the “false face”. Each fracture zone ruptures at the elastic–plastic boundary of surrounding rocks and then coalesces into a circle. The geometric progression ratio was determined; it is related to the mechanical parameters and ground stress of the surrounding rocks and calculated as follows: And the mechanism of the non-monotonic displacement law is revealed; the continuous formation of the “false face” causes the geostress redistribution and crack opening.

Knowledge of the rock mass strength is important for the design of all types of underground excavations. A frequently applied approach for estimation of the rock mass strength is through an empirical failure criterion, often in conjunction with rock mass classification/characterisation systems. This paper presents a review of existing methods to estimate the rock mass strength using empirical failure criteria and classification/characterisation systems—in this study, commonly denoted as estimation methods. A literature review of existing methods is presented, after which a set of methods were selected for further studies. The selected methods were used in three case studies, to investigate their robustness and quantitatively compare the advantages and disadvantages of each method. A Round Robin test was used in two of the cases. The case studies revealed that the N, Yudhbir-RMR76, RMi, Q-, and Hoek–Brown-GSI methods, appeared to yield a reasonable agreement with the measured strengths. These methods are thus considered the best candidates for realistic strength estimation, provided that care is taken when choosing values for each of the included parameters in each method. This study has also clearly shown the limits of presently available strength estimation methods for rock masses and further work is required to develop more precise, practical, and easy-to-use methods for determining the rock mass strength. This should be based on the mechanical behaviour and characteristics of the rock mass, which implies that parameters that consider the strength of intact rock, block size and shape, joint strength, and physical scale, are required.

The vibrations generated by a moving train in a tunnel will radiate into the surrounding ground which, in densely populated areas, will reach nearby buildings and its residents. Analyses are commonly made where the aim is to estimate the ground-borne noise and vibrations levels that may occur in nearby buildings. A common assumption is to treat the rock mass as an isotropic, homogeneous, and linear elastic material. Thus, the influence of discontinuities on the propagation of waves is not considered in the analyses. Within this study, numerical simulations were performed to study the propagation of low-frequency waves through a rock mass near a tunnel. A single period sinusoidal wave was applied as dynamic source on the floor of the tunnel. Observation points were located on the ground surface and around the tunnel. The influence on wave propagation from overburden, position of a discontinuity in relation to the tunnel, and normal and shear stiffness of the discontinuity, was studied by using the Universal Distinct Element Code (UDEC). The results show that increasing overburden reduces the vibration levels on the ground surface. Furthermore, the influence of the normal and shear stiffness of discontinuities depends on where the horizontal discontinuity is positioned in relation to the tunnel. If the horizontal discontinuity is positioned above the dynamic source (e.g. above tunnel or in the tunnel wall) the vibration levels on the ground surface will be reduced but if the horizontal discontinuity is located below the dynamic source (e.g. below the tunnel) the vibration levels on the ground surface will be enhanced. In our analyses, discontinuities only have an impact on the wave propagation if the normal and shear stiffness of is ≤10 GPa/m.

Mining operations at the Kristineberg Mine approaches depths of more than 1300 m. Severe ground conditions are consequently expected due to the combination of weak rock formations and relatively high in-situ stresses. The required support techniques, however, are not unique and depend on the ground deformation characteristics. On the other hand, the complex geology of the mining zone induces different types and magnitudes of failures. As a result, to improve the support performance, the possible ground conditions which may arise from varieties of geological parameters at the mine need to be characterized. Among different geological parameters at the mine, the location of the rock formations is seen to be an important factor, influencing the deformation behavior of the rock mass. The aim of this study is to increase the understanding of how the rock mass behavior changes due to the location of the rock formations. The study is performed through numerical modeling of an unsupported stope using FLAC2D; the required input data are determined based on the field observations and the corresponding literatures. Then, the modeled deformation behavior of the stope is studied through “Parameter Analysis”. The results, which show the sensitivity of the ground deformation to the location of the surrounding rock formations, can be used in future support design analysis.

Numerical modeling techniques have been applied in many mining and civil engineering projects. Traditionally, deterministic methods have been used frequently for the estimation of design or input parameters for numerical modeling. Whereas, it is known that the effect of variability and uncertainty sourced from the complex and variable nature of rock cannot be considered by deterministic approaches using single or mean value. In this paper, the authors tried to apply a probabilistic approach to consider the uncertainties and variability in rock properties. This is to make more a realistic assessment of design parameters of rock masses around an instrumented test drift in Malmberget Mine within the content of the “Rock mass - Rock support interaction project” conducted at the Division of Mining and Geotechnical Engineering, Lulea University of Technology. To calculate the design parameters GSI of rock mass, UCS and mi constant of the intact rock are considered as random variables. For each of these random variables ranges were specified depending on the laboratory and field information. Using Monte Carlo simulation method a possible range of each of necessary strength and deformability properties were obtained and presented. The assessed values can be used as preliminary input parameters and considered as basis for further numerical modeling calibration studies.

This paper presents a probabilistic approach for optimizing stope design methodology whiletaking into consideration the variability in the rock mass properties. For this study, a complex orebody in aCanadian mine was used. Because of the variability in the rock mass properties of the orebody, it was notpossible to determine precisely, the values of geotechnical design input parameters and hence the need toutilize a probabilistic approach. Point Estimate Method (PEM), a probabilistic tool, was incorporated intonumerical analysis using FLAC3D to study the deformation magnitudes of various stope geometries todetermine the optimal stope geometry with a minimum ground control problem. Results obtained for thedistribution of the wall deformations and the floor heaves for each option of the stope geometry werecompared to select the best geometry to achieve the optimum stability condition. The methodologypresented in this study can be helpful in the process of underground mine planning and optimization incomplex orebody.

Представлен вероятностный подход для оптимизации методики расчета параметров забоя с учетом изменчивости свойств горного массива на примере сложного рудного месторождения в Канаде. Для точного определения значений входящих геотехнических параметров использован вероятностный метод PEM (метод точечной оценки) в совокупности с численным анализом и программным обеспечением FLAC^3D. Определены деформации выработок с различной геометрией и оптимальные параметры выработки для минимизации проблем управления состоянием массива. Полученные результаты по распределению деформаций стенок и почвы выработок анализировались для каждого варианта с целью выбора их наилучшей геометрии и достижения оптимальной устойчивости

Stability analyses of underground rock excavations are often performed using traditional deterministic methods. In deterministic methods the mean or characteristics values of the input parameters are used for the analyses. These method neglect the inherent variability of the rock mass properties in the analyses and the results could be misleading. Therefore, for a realistic stability analyses probabilistic methods, which consider the inherent variability of the rock mass properties, are considered appropriate. A number of probabilistic methods, each based on different theories and assumptions have been developed for the analysis of geotechnical problems. Geotechnical engineers must therefore choose appropriate probabilistic method to achieve a specific objective while taking into account simplicity, accuracy and time efficiency. In this study finite difference method was combined with five different probabilistic methods to analyze the stability of an underground rock excavation. The probabilistic methods considered were the Point Estimate Method (PEM), the Response Surface Method (RSM), the Artificial Neural Network (ANN), the Monte Carlos Simulation (MCS), and the Strength Classification Method (SCM). The results and the relative merits of the methods were compared. Also the general advantages of the probabilistic method over the deterministic method were discussed. Though the methods presented in this study are not exhaustive, the results of this study will assist in the choice of appropriate probabilistic methods for the analysis of underground rock excavations. 

This paper presents a probabilistic approach for modelling complex rock masses wherethe intrinsic properties are highly variable. For this study a complex orebody in aCanadian mine is used. The mechanical properties of the host rock and the ore in thismine are found to be intrinsically variable with high contrast between their mechanicalproperties. It is apparent that the use of traditional deterministic methods to study thebehaviour of the open stopes is not appropriate for this mine. Hence, in this study aprobabilistic approach is adopted which allows the propagation of the variability of theinput parameters in the numerical modelling. Three different approaches were used toanalyze the stability of the open stopes based on the distribution of the different materialproperties of the rock mass. The results of the analysis using the three methods werecompared and the limitations and the potentials of each of the methods were discussed.The study provides insight into the significance of the rock mass property variability inthe numerical modelling of open stope stability and different ways that it could beincorporated into the modelling.

The rock mass is intrinsically variable in its physical and mechanical properties which makes it complex. This complexity is evident from the spatial random distribution of the properties from any site characterization program. The precise values for these properties are never known in most cases hence most geotechnical mine designs are based on fixed or discrete values of rock mass properties for stability analysis. This traditional deterministic approach neither reflects the inherent variability nor the uncertainty in the rock mass properties. Therefore, it is desirable to utilize a probabilistic approach which provides a range of possible results based on the variability in the rock mass properties. Understanding the effect of this random distribution and variability of the properties on stope stability is essential for more realistic mine design. In this study, a series of numerical analyses using the explicit finite difference element code FLAC, have been conducted to study the effect of the random distribution and variability of rock mass properties on the stope stability. The rock mass in the FLAC model is represented by different material properties randomly distributed to each zone. In order to compare the results, fixed average values of the material properties were also used for the FLAC model in another simulation. The results clearly indicate that rock mass property variability does affect the stope stability and that a deterministic approach to stope stability analysis could lead to conservative results.

A probabilistic approach is presented for the analysis of open stopestability. The approach considers the inherent variability and uncertainty whichare typical of rock mass properties. In this study, a series of numerical analyseswere performed using FLAC to study the stability of open stopes while takinginto account the variability in the rock mass properties. The rock mass wasdivided into six strength classes: three classes for the host rock and threeclasses for the massive sulphide ore. Each class was randomly distributed to theelements in the FLAC model. The host rock-to-massive sulphide ore ratio isenvisaged to have a strong influence on the stope. To verify this, three casesof different ore percentages were considered and the results compared.The results, which were presented as Probability Density Functions (PDFs),indicate that the zones of low stiffness show high range of displacements andthat the increase in the percentage of the ore significantly affects the stability ofthe stopes.

Stability analyses of any excavations within the rock mass require reliable geotechnical input parameters such as in situ stress field, rock mass strength and deformation modulus. These parameters are intrinsically uncertain and their precise values are never known, hence, their variability must be properly accounted for in the stability analyses. Traditional deterministic approaches do not quantitatively consider these uncertainties and variability in the input parameters. To incorporate these variability and uncertainties stochastic approaches are generally used. In this study, a stochastic assessment of pillar stability using Artificial Neural Network (ANN) is presented. The variability and uncertainty in the rock mass properties at the Laisvall mine were quantified and the probability density function of the deformation modulus of the rock mass was determined using probabilistic approach. The variability of the in situ stress was also considered. The random values of the deformation modulus and the horizontal in situ stresses were used as input parameters in the FLAC3D numerical simulations to determine the axial strain in the pillar. ANN model was developed to approximate an implicit relationship between the deformation modulus, horizontal in situ stresses and the axial strain occurring in pillar due to mining activities. The closed-form relationship generated from the trained ANN model, together with the maximum strain that the pillar can withstand was used to assess the stability of the pillar in terms of reliability index and probability of failure. The results from this study indicate that, the thickness of the overburden and pillar dimension have a substantial effect on the probability of failure and reliability index. Also shown is the significant influence of coefficient of variation (COV) of the random variables on the pillar stability. The approach presented in this study can be used to determine the optimal pillar dimensions based on the minimum acceptable risk of pillar failure

The anisotropy and stress-dependency of the strength and deformability of rock joints were investigated experimentally through shear tests of concrete replicas of natural rock joints. 50 concrete replicas of two natural granite joints were sheared on a servo-controlled hydraulic test frame under different magnitudes of normal stress and in different shear directions. Significant anisotropy in both the friction angle and shear stiffness of concrete replicas of joints were observed and both of them change with the variation of normal stress. The rates of dilatancy and contraction of joint samples appear to be different during cyclic shear tests, depending very much on the initial conditions of sample surfaces. Empirical relations were then postulated to describe these newly found properties of joint samples and used in developing new constitutive models for rock joints

A conceptual model for the behaviour of rock joints during cyclic shear and under constant normal stresses was proposed according to results from shear tests with 50 concrete replicas of rock joints. The shear strength and deformability of joint samples were found to be both anisotropic and stress dependent. Based on these experimental results, a two-dimensional constitutive model was developed for rock joints undergoing monotonic or cyclic loading sequences. The joint model was formulated in the framework of non-associated plasticity, coupled with empirical relations representing the surface roughness degradation, appearance of peak and residual shear stresses, different rates of dilatancy and contraction, variable normal stiffness with normal deformation, and dependence of shear strength and deformability on the normal stress. The second law of thermodynamics was represented by an inequality and used to restrict the values of some of the material parameters in the joint model. The new joint model was implemented into a two-dimensional Distinct Element Method Code, UDEC, and its predictions agreed well with some well-known test results.

The Norra Alliansen orebody of the Malmberget sublevel caving mine consists of iron ore interspersed with biotite schist and granitic inclusions. The schist is squeezed between the ore and the host rock and in direct contact with the ore along the majority of the length of the footwall. The schist exhibits high deformation when exposed to stress. SMART cable bolt roof deformation measurements are re-analyzed to draw conclusions regarding the patterns of deformation in the mine. Each bolt’s head is placed at the origin of a spherical coordinate system and the radius and inclination angle between the bolt and every production blast occurring during the bolt’s recorded lifetime are calculated. The deformation experienced by each instrument is investigated by comparing the long-term recorded movements with the developed geometric variables. Patterns of deformation magnitude and rate are found with respect to production-blast distance and inclination angle, instrument location, rock quality designation, and likely mining-induced stresses. Results show that deformation magnitude tends to be higher when driven by production blasting occurring on the production level above the instrumentation when accounting for the effects of distance, but average deformation magnitude for very-near production blasts tends to be higher than that for production blasts occurring directly above the instrument. Correlations also exist between the measured RQD, estimated rockmass parameters, and the measured deformation. Empirical evidence allows the identification of six scenarios which account for 91% of the recorded high-deformation-rate events. These scenarios help determine which production activities are most likely to cause high deformation rates.

The aim of our study is to explore the possibility of using the Kaiser effect method to assess damage in rock induced by blasting. Cores were drilled from a hard rock tunnel. The specimens prepared from the cores were uniaxially compressed in the laboratory and the AE from the specimens was recorded. The AE onset stress was taken as a measure of the damage in the rock induced by blasting. The results showed that the AE onset stress decreased with the distance from the tunnel floor. The extent of disturbance due to blasting could be estimated using the curve of the AE onset stress.

A constitutive model for brittle rocks has been developed based on the analysis of microcracks. The macroscopic deformation of a rock is decomposed into three components: response of the rock matrix, closure of open cracks and deformation due to the fracture of microcracks. The main contribution of this approach is the prediction of macroscopic deformation induced by microcracking. The basic model units are sliding cracks. Wing cracks are initiated at tips of sliding cracks and propagated in the major loading direction. The relationships between the compressive stress, the length of the growing crack and deformation due to the fracture of cracks are established. A deformation solution to an elliptic crack is employed to describe the closure effect of an open crack under compression. The total deformation caused by the closure and fracture of cracks is obtained by the sum of the components of individual cracks. In this model, the nonlinear behavior of deformation results either from the closure of open cracks at low stress levels, or from the fracture of microcracks at high stress levels. Hysteresis is captured under cyclic loading. Comparisons of the model simulations with experimental data are presented and show good agreement in the prefailure stage.

The Kaiser effect is a measure of damage developed in rock. It was investigated using 61 core specimens of eight types of rocks under uniaxial cyclic loading. The effects of the delay time between subsequent loading cycles and the holding time on the Kaiser effect were examined in granite specimens. The characteristics of AE occurring during unloading and holding were also studied. The experimental results showed that most of the rocks, with the exception of some iron ores, showed an obvious Kaiser effect before the load was very near the level of the strength. The delay and holding times did not strongly influence the Kaiser effect. Breakdown of the Kaiser effect was expressed by the felicity ratio that may be taken as a measure of the quality of rock. The onset of the continuously increasing AE in the first loading cycle can be taken as a measure of the damage in rock. The mechanism of the Kaiser effect was studied with the aid of a micromechanical model. An expression for the damage surface in the stress space was derived from the model. The theoretical expression was compared with the results of the Kaiser effect tests, which showed a satisfactory consistence.

Fracture is the main reason for the non-linear behaviour of hard rocks. The fracture mechanics of rock is studied in this article by analysis of the fracture process under compression. A constitutive model that describes the relationship between the macro deformation of rock and the micro fracture within rock is developed. The propagation of microcracks, the non-linearity of deformation, the loading-and-unloading hysteresis and the variation of the apparent Young's modulus and Poisson's ratio are studied using the developed model. The model simulations demonstrate that: (1) the fracture toughness, initial crack length, crack density, and Young's modulus are four crucially important parameters that affect the deformation behaviour of rock; (2) the elastic parameters (E and v) of the rock matrix should be measured in triaxial tests. If they are measured in uniaxial tests, the upper straight unloading portion of the stress-strain curve is suggested to be used for the purpose, unless the closure effect of open cracks will be included in the estimations. In addition (3), the slope of the reloading stress-strain curve is a measure of the damage in material.

The influence of Poisson"s ratio on overcoring was investigated using data from 47 tests. In these tests both the CSIRO HI triaxial cell and the triaxial cell developed by Swedish State Power Board were used. Boreholes drilled parallel to the major principal stress resulted in higher Poisson"s ratio compared to boreholes drilled at an angle to major principal stress. Numerical modelling by FLAC was used to simulate overcoring and biaxial testing. When the major principal stress is acting perpendicular to the borehole, the extensional strain is larger on the inner surface of the core during overcoring than when parallel to the borehole. Simulations of a biaxial test of a virgin core show extensional strain, which may be larger than critical extensional strain of the rock type, indicating non-elastic behaviour.

Studying the deterioration of shotcrete due to freezing and thawing is important for improvement of the understanding of the failure mechanisms/debonding of shotcrete in cold regions. Water leakage in a tunnel leads to ice growth during freezing temperature and ultimately creates favorable environment for fallouts of shotcrete and rock. Repeated freezing and thawing of shotcrete lead to development of new micro cracks and propagation of pre-existing micro cracks. In this study, test panels of granite with dimension 800 x 800 x 80 mm covered with 50-mm thick shotcrete were subjected to freezing and thawing action in a controlled environment. The initiation and the development of freeze-induced micro cracks in shotcrete-rock interface were studied by continuously monitoring acoustic emissions (AE) and temperature. The clustering of the AE events during freezing and thawing indicates that micro cracks appeared in the shotcrete-rock interface and caused adhesion failure. The larger number of AE events in the panels, with access to water during freezing, confirmed that water contributes to material deterioration and also reduces the adhesive strength. The test results showed that most of the acoustic emission occurred during the freezing cycle and the number of acoustic emission events did not increase with the successive increase of the number of freezing and thawing cycles.

Road and railway tunnels in cold regions are often affected by problems related to water leakage and freezing temperatures. Water leakage in a tunnel leads to ice growth when the temperature goes below freezing and creates favourable environment for fallouts of shotcrete and rock. This paper presents results and observations from laboratory freezing – thawing experiments on rock blocks covered with shotcrete and focuses on the degradation of the shotcrete-rock interface due to ice growth.. The initiation and the development of freeze-induced micro cracks in shotcrete-rock interface were studied by continuously monitoring acoustic emissions (AE) and temperature. The clustering of the AE events during freezing and thawing indicates that micro cracks appeared in the shotcrete-rock interface and caused adhesion failure. The larger number of AE events in the panels, with access to water during freezing, confirmed that water contributes to material deterioration and also reduces the adhesive strength.

Tailings dam failures have been occurring in recent years. Many of these failures have caused human casualties, destruction of property, and damage to environment and huge economic loss to the mining industry. The monitoring of the dam is essential to know the existing state of the dams and to ensure the safety of the dam over its life time. The present study has been conducted to test the applicability of electrical resistivity and self-potential (SP), for detecting anomalous seepage through mine tailings dams in Sweden and monitoring the physical condition of the dam. This study has demonstrated the potential of using geoelectrical methods for monitoring the conditions of the tailing dams related to seepage.

The static state of stress at the brow in a sub-level caving mine is, due to stress re-distribution, almost uniaxial (major principal stress perpendicular to the cross cut). Since large amounts of explosives are detonated in each production round, the impact of stress waves on the brow can be significant. An extensive failure mapping programme in the Kiirunavaara mine showed that many of the failures close to the brow were structurally controlled. Furthermore, the area of damaged shotcrete was extensive when plain shotcrete was used. At brows supported by fibre reinforced shotcrete, damage in the roof was observed within a horizontal distance of about 3 m from the drawpoint. To study the behaviour of roof wedges supported by shotcrete and subjected to blast-induced vibrations a single-degree-of-freedom (SDOF) model was developed. The model consists of a shotcrete layer and a rock wedge. Vibration measurements showed that maximum particle velocity was approximately 1.2 m/s. The acceleration record showing the largest magnitude was used as the load in the dynamic analyses. The analyses showed that a wedge can be ejected by a dynamic load even if the static safety factor was >10. Furthermore, the non-linear response of the wedges was in most of the cases greater when the wedge was supported both by the joints and the shotcrete layer compared to the case when the wedge was only supported by shotcrete. A conclusion from the analyses is that it is difficult to predict the dynamic response from static calculations. To provide a safe working environment close to the drawpoint, the rock support must sustain the impact of stress waves from production blasting. To support rock wedges subjected to dynamic load the support must be able to consume the energy imposed on the wedges from blasting. The non-linear numerical analyses showed that reinforced shotcrete has the necessary bearing capacity to support the wedges formed in the roof of the cross cut close to the brow. This was in fair agreement with the failure mapping. The single-degree-of-freedom (SDOF) model can be used to study the response of an arbitrarily shaped rock wedge supported by shotcrete as long as the movement of the wedge can be idealised by a pure translation and the dimensions of the wedge are small compared to the length of the incident wave. Analyses showed that 2D wedges can be used to judge whether symmetric or non-symmetric 3D wedges in a uniaxial stress field (which occurs close to the brow) are stable or not when they are subjected to waves induced by blasting.

The shotcrete-rock interaction is very complex and is influenced by a number of factors. The influence of the following factors was investigated by a series of numerical analyses: the surface roughness of the opening, the rock strength and Young's modulus, the discontinuities, the extent and properties of the excavated disturbed zone, the mechanical properties of the interface between shotcrete and rock, and the thickness of the shotcrete lining and the rock bolts. The study was carried out as a sensitivity analysis. The results showed that the rock strength and the surface roughness had significant impact on the number of failures at the rock-shotcrete interface and in the shotcrete lining. Furthermore, the behaviour of the lining is sensitive to small amplitudes of the surface roughness. In all the cases investigated, a high interface strength was favourable. The results indicate that if a thick shotcrete lining is dependent on the bond strength. The benefit of using a thicker lining can be doubtful. The analyses showed that for an uneven surface the extent of the EDZ had a minor effect on the behaviour of the shotcrete lining. Furthermore, if rock bolts were installed at the apex of the protrusion instead of at the depression, the number of failures decreased both at the interface and in the lining

Shotcrete is widely used as rock support in mines and in civil engineering projects. In many cases the adhesion between the shotcrete and the rock is one of the most important properties. The objective of this study was to: (i) identify typical failures of shotcrete in a mining environment, (ii) measure the growth of the adhesion strength of shotcrete with respect to age, (iii) investigate the influence of surface treatment (scaling and cleaning) on the adhesion strength of shotcrete and (iv) investigate the occurrence of shrinkage cracks in the shotcrete and failures in the interface between the shotcrete and the rock. The failure mapping showed that most of the observed failures of shotcrete are in areas with a thin shotcrete layer ( less than or equal 20 mm) together with a low adhesion strength. Furthermore, the measured average thickness of shotcrete (40 mm) was almost the same as the ordered thickness, the problem is the wide scatter. In 18% of the thickness measurements, the shotcrete thickness was less than or equal to 20 mm. If the shotcrete lining in the mine has the same thickness distribution as during the tests, the supporting ability of as much as 18% of the shotcrete lining can be considered as more or less negligible. No differences in free shrinkage were observed between plain and steel fibre reinforced shotcrete. Relatively large areas with lack of contact between shotcrete and wall were observed in the shrinkage tests. This indicates that restrained shrinkage can destroy the bond between the shotcrete and rock. The results from the restrained shrinkage tests show the importance of the bond at the interface. If bond failure occurs, the ability to distribute and control the crack width ceases. The growth of the adhesion strength on a sandblasted concrete wall and the growth of the compressive strength showed a correlation. However, the scatter in the results indicates that further investigations have to be done in order to establish a reliable relation between the growth in adhesion and compressive strength with age. The results from the field tests showed that the adhesion strength was significantly higher on rock surfaces that had been water jet-scaled (with 22 MPa water pressure) than those treated by mechanical scaling followed by cleaning of the rock surface (water pressure 0.7 MPa).

The Kiirunavaara mine is a large scale sub level caving (SLC) mine located near the city of Kiruna in northern Sweden. It is owned and operated by LKAB (Luossavaara-Kiirunavaara AB). The mine produces approximately 28 million tonnes of iron ore annually. Over the last 30 years the mine has experienced a slow but progressive fracturing and movement in the footwall rock mass induced by the SLC operations. The footwall contact which assumes a “slope-like” geometry is partially supported by the caved material from the hangingwall. However, since the late 1980s damage has been observed on the footwall crest as well as within the footwall. Progressive rock mass movement in the footwall is indicated by surface subsidence and visual observations underground. The extent of the damage has traditionally been estimated using empirical relations. Most of the current long term underground infrastructure within the footwall is located at a considerable distance from the ore contact. However, for new developments on deeper levels it is imperative to predict the future extent of the damage volume. Approximating the position of the damage boundary in the footwall at the current state of mining would assist in predicting the extent and characteristics of the damage volume as the mine deepens. LKAB and LTU (Lulea University of Technology) have therefore initiated a joint research project to study the long term stability of the footwall at the Kiirunavaara mine. This paper constitutes part of the work in this research.The paper describes a damage mapping campaign and subsequent analysis of the Kiirunavaara mine footwall to approximate the outer boundary of the damage. The footwall was systematically mapped on 6 levels between 320 and 800 m. The mapping results were then used to interpolate damage lines on the respective levels. The damage lines were used to construct a continuous damage surface between the studied levels. Existing records of damage mapping, monitoring and predictions were reviewed and compared to the results from the current campaign. The new results show that, the outer damage surface appears to remain stationary on the upper levels while new damage was observed on the deeper levels. At levels above 740 m the damage is judged to be mainly controlled by movements along natural discontinuities. At levels below 740 m the majority of the damage seems to be stress induced.

The performance of two different percussive rock drills has been studied theoretically by simulating the drilling process. The two drills studied correspond to an idealized drill and a more complex, prototype drill. The efficiency, the maximum stress (maximum tensile stress) and the minimum stress (maximum compressive stress) have been determined during the simulations. Also, the stability of the drill string with respect to buckling has been estimated as a function of a constant applied thrust. Two cases of applied thrust were simulated. They correspond to a constant applied thrust and an applied thrust with time dependent harmonic variations. The results show that for low frequency variations in thrust, the performance of the rock drill is almost the same as for a constant applied thrust. It was assumed in the remainder of this study that the applied thrust can be treated as a constant force. The results obtained for a constant applied thrust are summarized.