Accurate ore estimation processes are of crucial importance in the mining scenario. Over the last 20 years, one practical approach to improve ore grade estimation has encouraged the need to incorporate the geology of the ore deposit being estimated in the estimation process. The objective of this study is to investigate the impact of the geology on the kriging estimation of the ore grade of a portion of a porphyry copper deposit. Preliminary data analysis demonstrates the need to perform a subsequent variogram modeling and kriging estimation of the ore grade by rock type separation. Global and local estimations were done to assess the influence of the geology on the ore grade estimation at a global and local scales. The results obtained in this study demonstrated that for the portion of the deposit studied the incorporation of the geology does not produce substantial improvement on the ore grade estimation.

The success of rock fragmentation due to blasting depends on many variables, such as rock properties, in-situ fracturing, and blast design. Traditionally, the size distribution of fragmented rock particles has been determined through screen sieving. Modern techniques using video images and computer image processing techniques have the potential for analyzing rock fragmentation accurately and efficiently. A procedure has been developed for analyzing rock fragmentation which uses a high-resolution video camera for capturing images in the field, and specialized computer algorithms for processing these images. First of all, computer algorithms have been developed to delineate the individual rock fragments in the images. Secondly, a set of experiments have been conducted in the laboratory, in which the two dimensional information from the images is correlated with sieve results. Based on these experiments, a set of probabilities have been determined for correctly determining the size and volume of rock fragments from two dimensional images. Using these probabilities along with the particle delineation algorithm, the size distribution for the rock fragments is calculated. The computer algorithms can also combine information from many images to take into account sampling and images taken at different scales.

Rock is a very heterogeneous material, containing structural weakness at all scales. These weaknesses include grain boundaries, pores, and cracks on the small scale, and joints, faults, and bedding planes on the large scale. Nonlinear rock deformation in the low-temperature, low-confinement regime is due primarily to the growth of cracks from these weaknesses and the coalescence of cracks to form macroscopic structural features. Another important aspect of rock deformation and failure is the statistical distribution of weaknesses in the initial microstructure. Borehole breakout is the process by which portions of a borehole wall fracture or spall when subjected to compressive stresses. Studies of borehole breakout in the past twenty years include experiments, field studies, and numerical modeling. With regards to the numerical modeling of borehole breakout, the rock surrounding the borehole is considered as a nonlinear continuum material in most of the previous approaches. Experiments and field studies, however, have shown that the heterogeneous and discontinuous nature of rock has a strong impact on the mechanics of borehole breakout. This dissertation describes a numerical model that has been developed to simulate the damage of rock and the corresponding non-linear stress-strain behavior, and also the progression of borehole breakout in heterogeneous and discontinuous rock by mixed mode crack growth, interaction, and coalescence. The rock is simulated as an elastic material containing a random distribution of cracks. As compressive load is applied, the initial cracks grow, interact, and coalesce to form macroscopic fractures. The numerical model was developed by making a series of modifications to the displacement discontinuity code of Crouch and Starfield (Crouch & Starfield, 1983). The most important modifications include modifying the boundary element for the calculation of stress intensity factors, adding Coulomb friction for closed portions of cracks, adding a crack generator, and adding an algorithm for crack coalescence. The numerical model is used to simulate the non-linear deformation and the progression of breakout in Westerly granite, and the results are realistic.

The objective of this investigation is to experimentally determine the effectiveness of fracture sealing in welded tuff using ordinary portland cement and microfine cement grouts. Fracture grouting will most likely be used to seal fractures intersecting high level nuclear waste repositories. Fractures are potential pathways for the migration of radionuclides. Laboratory experiments have been performed on seventeen tuff cylinders. (1) tension induced cracks, (2) natural and, (3) sawcut surfaces serve as fractures. Prior to grouting, the hydraulic conductivity of the intact rock and that of the fractures themselves are measured under a range of normal stresses. Grouts are injected through axial boreholes at pressures of 0.3 to 4.1 MPa while holding fractures under a constant normal stress. Five grout formulations have been selected. Minor amounts of bentonite (0 to 5 percent by weight) have been added to these grouts to increase stability. Water to cement ratios range from 0.45 to 1.0. Permeameter testing of grouted fractures is used to evaluate the effectiveness of fracture grouting.

The objective of this investigation was to experimentally determine the variation in the flow characteristics of tactite copper ores during the leaching process. The laboratory work involved estimation of porosity and permeability of cores before leaching, and after various intervals of leaching. The tests were carried out using specimens of diameter 1.5 and 4.0 inches, prepared from boulders obtained from the Cyprus Casa Grande mine. A solution containing about 25 grams of sulfuric acid per liter was used as the lixiviant. Porosity was determined using the water saturation method. The permeability was estimated using either helium or water with the specimens stressed triaxially to simulate in situ conditions. The rate of copper recovery from the specimen during the leaching process was also determined by estimating the copper content in the solution used for leaching. The results indicate that the permeability and porosity of the specimens increased with leaching, but both tend to remain constant after most of the soluble material in the core was dissolved. The rate of copper recovery was high initially but dropped as leaching progressed and the copper in the specimen was depleted.

The history of mine backfill shows that in the past, considerable improvements in backfill-reliant mining technology were made when new fill systems were introduced. The present trend in mine backfill technology is towards the use of high-density fill systems. Tight filling and void reduction have also become essential requirements in engineering design of mines to ensure global stability. High-density fills have a low moisture content and are more competent products requiring less binder and time for stabilization, compared to conventional hydraulic backfill. Cemented rockfill and tailings/sand paste fills are two familiar high-density backfill systems in current use. In future there could be a high demand for low porosity high-density fills, as mines go deeper and ground stresses increase. / This study was an original attempt to investigate the characteristics and property of high-density composite fill systems. Composite fills are made up of derivatives of waste rock, tailings, sand and metallurgical by-products. Composite fills represent the future direction of backfill technology. As mines go deeper, the ore could be processed underground and the waste rock and tailings combined together to form a low-porosity competent fill product. The application of composite fill systems will also increase the material available for backfilling, provide more flexibility in backfill mix design and produce competent fill systems for ground support. It will also benefit the underground mine environment through effective utilization of mine wastes. / The fundamental basis of the work required the study and understanding of the characteristics and properties of cemented rockfill and paste backfill. The effects of sand addition to fine tailings as a means of reducing porosity and improving the mechanical properties of the fill product were also investigated. Additionally, a new concept of backfill, namely, Composite-Aggregate Paste (CAP) that consists of a mixture of fine tailings and graded coarse aggregates was introduced and the material properties were investigated.

Cable bolt supports are widely used as ground support in underground hard rock and coal mines. During the last two decades, the cable bolt technology has improved significantly and new types of cable bolts, new grouts and new installation methods and pumps have been introduced and successfully developed. The design of cable bolts is currently attempted by empirical methods based on rockmass characterization. However, until such methods do not take into account all the important parameters affecting the mechanical behaviour of cable bolts namely, the type of cable geometry, grout quality, stiffness of the host medium, the amount of confining pressure and the presence of end anchorage and pretension. / This thesis deals with the numerical modelling of cable bolts considering all of the above mentioned factors. To evaluate the shear bond stiffness of cable bolts, which simulate the interface between the cable bolt and the rockmass, a new empirical model is developed. A database with data of pull-out tests is constructed. Empirical relations were derived between the shear bond stiffness and the variation of confining pressure, the modulus of elasticity of host medium and the water:cement ratio. The relations were derived for three types of cables (standard cable, nutcase cable and garford bulb cable). A new numerical model using the finite element method is developed for the numerical simulation of cable bolts. Together with the empirical model for shear bond stiffness, the numerical model computes the load distribution along the cable bolt and thus can simulate fully grouted, partially grouted, end anchored and tensioned cable bolts. The model is implemented in an existing finite element code (e-z tools). Parametric analyses are performed on the new model, and proved useful in the demonstration of the role of each design parameter. / A new design methodology is proposed to evaluate the load distribution along the cable bolt. Two case studies are presented: a stope hanging-wall support by cable bolt in a hardrock mine and a tailgate secondary supports in a coal mine. A comparison with measurements in the field shows good agreement with computed forces.

A review of automation, robotics and communications technology has established the need for the development of a communications infrastructure capable of supporting future underground hard rock mine automation systems. A series of underground experiments were undertaken at Copper Cliff North Mine to evaluate the design criteria and performance of several communications infrastructures. The work successfully demonstrated the capability of real-time operation of voice, data and stationary video communication as well as surface-to-underground tele-operation of a load-haul-dump machine. This was achieved with a communications system consisting of a broadband bus linked to leaky feeder coaxial cables by means of distributed antenna translators. The success of the trials permitted a strategy for mine automation to be devised. The economic benefits of mine automation were estimated by means of economic models developed for the mine. Projected benefits, evaluated in terms of mining cost reduction, throughput time and quality improvement, were concluded to be significant. As a result of the analysis, future research and development is concluded to be best targeted at improving ore grade, optimizing process productivity and maximizing machine utilization.

The depth at which underground mines operate has been increasing continuously which is particularly true in the case of hard rock mining. The stability issues associated with mining at great depth pose tough challenges to engineers and researchers alike. Long-term mine developments in deep hard rock mines such as haulage drifts need to be functional during the entire life of the mine plan without posing any major stability concerns, which will otherwise hamper the production and other logistics associated with mining operations. High convergence and rockburst hazards are the main problems due to high stress and mining-induced seismicity in deep hard rock mining. In such circumstances, the understanding of drift support behavior under static and dynamic conditions is crucial for mining engineers when dealing with drift stability in deep, hard rock mines. In this thesis, current design methods for selecting drift support systems are reviewed, which are mostly dependent on empirical approaches and are geared towards static support design. Based on this, the current research focuses on ground support analysis under both static and dynamic conditions to understand drift support behavior with respect to nearby mining. Numerical modeling of drift primary and secondary supports is performed by developing two models using the 2-dimensional FLAC code. Axial loads induced in the drift support system under static and dynamic conditions are estimated for the case study hard rock mine in Canada at a depth of 1500 m. The results of numerical modeling are obtained in terms of axial loads in the drift support system, wall damage due to tension under dynamic conditions, and the extent of rock mass yielding around the drift. It is found that mining on the same level is critical to drift stability under static conditions, and rock mass yielding in the south wall of the drift (towards the ore body) extends beyond the bolting horizon once this stage begins. The results also show that by providing secondary support before same level mining commences, drift stability is greatly enhanced. The static model is calibrated through the implementation of an in-situ monitoring program of axial loads induced at the head of the rockbolt. A new load monitoring device called U-cell is successfully used for this purpose. Measured and estimated axial loads are then compared and found to be in good agreement. The preliminary dynamic analysis shows that a peak particle velocity of 2.0 m/s at the periphery of the drift will cause wall damage more than 1.0 m when only primary supports are provided, and around 0.5 m when secondary supports are installed along with the primary ones, and when there is no nearby mining taking place. The effects of lower level and same level mining under dynamic conditions are also examined, and wall damage and rock mass yielding are estimated. The estimation of wall damage depth is crucial in designing dynamic rock supports. It is demonstrated that wall damage due to various levels of ground motion can be estimated by dynamic numerical modeling. Finally, a methodology for the design of dynamic rock supports is presented, which is based on the selection of yielding support type and pattern, the estimation of the ejection velocity, and the volume of wall damage as obtained from dynamic modeling. / La profondeur des mines souterraines a augmenté de manière continue, particulièrement en ce qui concerne les mines en roches dures. Les problèmes de stabilité associés aux mines profondes représentent comme des defis pour les exploitants, comme pour les chercheurs. Les développements miniers à longue durée de vie dans les mines profondes, tels que les galléries de roulage, doivent rester fonctionnels pour toute la durée de l'exploitation, sans poser de soucis majeurs, qui, sinon, nuiraient à la productivité et à l'organisation des opérations minières. Les fortes convergences et le risque de coup de terrain constituent les principaux problèmes dus aux fortes concentrations de contraintes et à la sismicité minière induite dans les mines profondes en roches dures. Dans de telles circonstances, la compréhension du comportement du soutènement des galeries sous l'effet de chargements statiques et dynamiques est essentielle pour les ingénieurs miniers confrontés aux questions de stabilité dans les mines profondes en roches dures. Dans cette thèse, nous exposons les méthodes courantes de dimensionnement du soutènement des galeries, qui reposent principalement sur des approches empiriques et ont pour objectif d'assurer la stabilité sous chargement statique. Sur cette base, la recherche se concentre sur des méthodes de dimensionnement du soutènement sous des charges statiques et dynamiques, dans le but de comprendre le comportement du soutènement des galeries adjacentes aux zones en cours d'exploitation. La modélisation numérique du soutènement primaire et secondaire des galeries est réalisée en développant deux méthodes recourant au code bidimensionnel FLAC. Les charges axiales dans le soutènement des galeries sont estimées sous des sollicitations statiques et dynamiques, dans le cas d'une mine canadienne en roche dure, à 1500 m de profondeur. Les résultats de la modélisation numérique sont présentés en termes de charge axiale dans le soutènement, d'endommagement des parois sous l'effet des tractions induites par les sollicitations dynamiques et de l'extension de la zone rompue autour des galeries. Nous montrons ainsi que l'exploitation sur le même niveau a des conséquences importantes sur la stabilité des galeries en chargement statique, et qu'au niveau du parement sud (i.e. du côté du gisement), la zone rompue s'étend au-delà de la longueur des boulons au début de cette étape. Les résultats montrent aussi que la stabilité de la galerie de roulage est très nettement améliorée si un soutènement secondaire est mis en œuvre lorsque commence l'exploitation sur le même niveau. Le modèle statique est calibré en utilisant des mesures in situ de la charge axiale sur les têtes de boulons. Pour ce faire, un nouveau dispositif de mesure de la charge, appelé « U-cell » a été utilisé avec succès. Les mesures de charge et les résultats de la modélisation sont comparés et sont en bon accord. L'étude dynamique préliminaire montre que des vitesses de points matériels de l'ordre de 2.0 m/s à la périphérie de la galerie de roulage induisent un endommagement au delà de 1.0 m de profondeur lorsque seul le soutènement primaire est mis en œuvre, et au delà de 0.5 m lorsqu'un soutènement secondaire est installé, pour peu qu'il n'y ait pas de zone en exploitation à proximité. Les effets de l'exploitation sur le même niveau et sur un niveau inférieur sont également comparés; l'endommagement des parois et la rupture de massifs rocheux sont estimés. L'estimation de l'endommagement de la paroi est essentielle afin de dimensionner le soutènement dynamique. On montre que l'endommagement de la paroi peut être estimé par modélisation numérique, pour différents niveaux de vitesses du terrain. Pour finir, une méthodologie pour le dimensionnement du soutènement dynamique est présentée; elle est basée sur la sélection du type et de la géométrie du soutènement. La vitesse d'éjection et l'endommagement de la paroi sont estimés par modélisation numérique.

Mine developments such as haulage drifts and cross-cuts are the primary access to the mining blocks of an orebody in multilevel mining systems for tabular ore deposits. Thus, their stability is of utmost importance during the planned period of production or the life of a mine plan. Many Canadian underground mines use longitudinal and transverse stoping with delayed backfill to extract tabular ore deposits. These methods require access to the orebody through a number of sill drives or cross-cuts which link the orezone to the haulage drift hence creating intersections on multiple levels. Mine development instability could lead to serious consequences such as injuries, production delays and higher operational cost. The objective of this research is to develop a hybrid approach in which deterministic numerical modelling is integrated with probabilistic methods to evaluate the stability of mine developments due to nearby mining activity. A case study comprising four consecutive mine levels in a deep underground metal mine in Sudbury, Ontario has been adopted for this study. The stability performance of the haulage drift is assessed using two separate evaluation criteria, namely Mohr-Coulomb yield function and Brittle Shear Failure. Random Monte-Carlo (RMC) technique is then employed in conjunction with Finite difference modelling software FLAC to determine the probability of instability or unsatisfactory performance of the haulage drift with respect to nearby mining sequence. In this study, the haulage drift performance is considered unsatisfactory when the yield zones or brittle shear failure around the haulage drift extend beyond the anchorage limit of the rock support. A comparison of the results from Mohr-Coulomb and Brittle Shear conditions has revealed that Mohr-Coulomb is more conservative from a design point of view. A three-dimensional, elastoplastic, finite difference model (FLAC 3D) is then constructed to simulate the case study mining orezone. The unsatisfactory performance of the intersection is evaluated with respect to mining sequence in terms of the strength-to-stress ratio computed by FLAC3D. Unsatisfactory stability performance is defined by a strength-to-stress ratio that is less than 1.4 and its corresponding extent into the rockmass around the intersection. Due to the large size of the FLAC3D model, the probabilistic simulations are conducted with the Point-Estimate Method (PEM), which requires significantly lesser number of simulations than Random Monte-Carlo (RMC). The results are presented and categorized with respect to probability, instability, and mining stage. In order to validate the numerical model, Multi-point borehole extensometers (MPBX) are installed at selected intersections to monitor the rock deformations as mining activities progress. The monitoring results revealed a lateral shift of the drift walls toward the orebody and much less deformations in the drift back. Finally, a methodology is developed to estimate the geotechnical risk of drift instability by considering the probability of failure and cost of consequence of such failure at an intersection. A 5-level risk index is derived which ranges from low to extreme. The methodology is demonstrated through an intersection from the case study mine, and the risk index is shown to vary with mining sequence. It is shown that the risk-index methodology can be used to confirm the need for enhanced supports, but it can also be used as basis for the comparison alternative mine designs. / Les développements miniers tels que les galeries de roulage et les travers-bancs constituent les accès principaux au gisement lors de l'exploitation d'un gisement tabulaire sur plusieurs niveaux. C'est pourquoi leur stabilité est d'une importance primordiale pendant la période de production ou pendant la planification d'une l'exploitation. De nombreuses mines souterraines canadiennes emploient des méthodes d'abatage par chambres avec remblayage différé pour exploiter les gisements tabulaires. Ces méthodes nécessitent un accès au gisement par de nombreux travers-bancs qui relient le gisement aux galeries de roulage, créant des intersections à de nombreux niveaux. L'instabilité de ces galeries peut conduire à de graves conséquences mettant en jeu la sécurité du personnel, à des retards de production et à des couts d'opération plus importants. L'objectif de cette recherche est de développer une approche hybride, basée sur une modélisation numérique déterministe intégrant des méthodes probabilistes, pour évaluer la stabilité d'une galerie d'avancement en fonction de la proximité de l'activité minière. Nous présentons une application à une mine métallique située à Sudbury, en Ontario, dans laquelle l'exploitation est réalisée sur 4 sous-niveaux.La stabilité d'une galerie de roulage est calculée à partir des 2 critères suivants: critère de plasticité de Mohr-Coulomb et « Brittle Shear Failure ». La méthode de simulation aléatoire de Monte Carlo (RMC) est utilisée conjointement avec le logiciel de différences finies FLAC pour déterminer la probabilité d'instabilité de la galerie de roulage en fonction de la séquence d'exploitation choisie. La stabilité de la galerie de roulage est considérée comme non satisfaisante dès lors que la zone de plasticité autour de la galerie excède la longueur des boulons. Une comparaison entre les critères d'évaluation montre que le critère de plasticité est le plus sécuritaire pour témoigner de l'influence de la séquence d'exploitation. Un modèle élasto-plastique en 3 dimensions, calculé par la méthode des différences finies (FLAC-3D), est crée pour simuler le cas d'application. La performance insatisfaisante d'une intersection est évaluée au moyen du ratio contrainte/résistance. La stabilité non satisfaisante est définie par un ratio inférieur au seuil de 1,4 et par l'étendue de la zone correspondante autour de l'intersection. Du fait de la grande taille du modèle numérique, les simulations probabilistes sont réalisées avec la méthode d'estimation ponctuelle qui nécessite un nombre significativement moins important de calculs que la méthode de Monte-Carlo aléatoire. Les résultats sont présentés et classés selon leurs probabilités, leur degré d'instabilité et l'état d'avancement de la séquence d'exploitation. Des extensomètres de forage à points de mesure multiples (MPBX) sont utilisés pour mesurer les déformations rocheuses d'une intersection au fur et à mesure de l'excavation. Les résultats sont utilisés pour calibrer le modèle FLAC-3D. L'auscultation a montré l'existence d'un déplacement latéral des parois de la galerie de roulage en direction du gisement et une déformation moindre du toit. Le coût des conséquences de la rupture d'une intersection est estimé par le cout de développement d'un contournement. Une échelle de risque à 5 niveaux, allant de « faible » à « extrême », est proposée. Cette échelle de risque est appliquée à une intersection de la mine étudiée et on montre que le niveau de risque dépend de la séquence d'exploitation. On montre également que cette méthodologie peut être mise en œuvre afin de confirmer la nécessité d'un soutènement amélioré. Elle peut aussi servir pour la comparaison entre différentes méthodes d'exploitation.