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Rock wedge stability is a common concern in underground excavations. Since the wedge stability influences the support design, it is important to use a wedge stability analysis method that can capture factors like excavation geometry, joint parameters and properties, rock mass properties, rock cover, and stress field as accurately as possible. This thesis compared the Limit Equilibrium Method (LEM) and the Discrete Element Method (DEM) using the software UnWedge (LEM) and 3DEC (DEM). The objectives included studying how factors like excavation and joint geometry, stress field, rock cover, and rock mass properties could be considered by the methods and how that affected the wedge stability and support design. Nine different analysis cases were defined with the aim of capturing wedge stability analysis parameters representable for both the civil engineering and mining sector. The study found that 3DEC allows for more accurate modelling of excavation and joint geometry, considering joint density and full 3D geometries, while UnWedge has limitations in creating intersecting tunnels and does not consider joint density. Only 3DEC, with stress redistribution and a plastic material model, captures rock mass failure mechanisms other than wedge failure. Based on the study, it is recommended to use DEM for situations where the excavation geometry cannot be assumed as two-dimensional with constant cross-section, and, when stress-induced rock mass failure is expected, use DEM with stress redistribution. It is also recommended to use information about joint lengths if available and apply engineering judgement when studying results of wedge volume and support force.

Roof falls and accumulation of dangerous gasses are the most common hazards in any underground coal mine. Different mechanisms can jeopardize the stability of the roof in underground excavations and successful roof control can only be obtained if the failure mechanism is identified and understood properly. The presence of discontinuities, the inherent variability of the rock mass and discontinuity properties, and the uncertainties associated with directions and magnitudes of the in-situ stress makes the rock engineering problems challenging. The numerical modeling can assist the ground control engineers in designing and evaluating the stability of the underground excavations. If extensive geological and geotechnical data are available, then detailed predictions of deformation, stress and stability can be accomplished by performing numerical modeling. If not, still the numerical modeling can be used to perform parametric studies to gain insight into the possible ranges of responses of a system due to likely ranges of various parameters. The parametric studies can help to identify the key parameters and their impact on stability of underground excavations. The priorities of the material testing and site investigation can be set based on the selected key parameters from parametric studies. An underground coal mine in western Pennsylvania is selected as a case study mine to investigate the underlying causes of roof falls at this mine. The immediate roof at the case study mine consists of laminated silty shale, shale, or sandstone that changes from area to area, and the floor is shale or soft fireclay. This study was mainly focused in the stability analysis of the roofs with the laminated silty shale rock type, where the majority of roof falls had taken place in the roof with this type of roof material. Extensive laboratory tests were performed on the core samples obtained from the case study mine to estimate the intact rock and discontinuity properties of the materials that occur in large extent at the selected interest area of the case study mine. In this research, the three-dimensional distinct element method was used to investigate the stability of the roof in an underground room-and-pillar coal mine. The implemented technique was able to accurately capture the failure of the major discontinuities and rock masses which consist of intact rock and minor discontinuities. In order to accurately replicate the post failure behavior of the rock layers in the immediate roof area, the strain-softening material constitutive law was applied to this region. Extensive numerical parametric studies were conducted to investigate the effect of different parameters such as the variation of immediate roof rock mass strength properties, variation of discontinuity mechanical properties, orientations and magnitudes of the horizontal in-situ stresses, and the size of pillars and excavations on stability of the excavations. The distribution of post failure cohesion along with other measures such as accumulated plastic shear strain, distribution of Z-displacements at the roofline, failure state (joint slip and tensile failure) and displacement (normal and shear displacements) of discontinuities were used to accurately assess the roof stability in this case study. The research conducted in this dissertation showed that the bedding planes play an important role on the behavior of roof in underground excavations. Therefore, an appropriate numerical modeling technique which incorporates the effect of discontinuities should be employed to simulate the realistic behavior of the discontinuous rock masses such as the layered materials in roof strata of the underground coal mines. The three-dimensional distinct element method used in this research showed the clear superiority of this technique over the continuum based methods.

The objective of this master’s thesis is to evaluate the use of Bonded Block Modelling (BBM) in 3DEC software combined with hybrid rock bolts, for three different cases. These cases included the laboratory rock bolt case, the shearing case and the blocky rock mass case. 3DEC is a Distinct Element Method (DEM) numerical software which can be used to simulate both continuum and discontinuum media in 3D. The Bonded Block Model in 3DEC can be used to simulate a rock mass as bonded polyhedral elements. The BBM is a relatively new numerical modelling technique. Earlier studies have focused mainly on laboratory test cases and less on field scale studies. The laboratory rock bolt test was introduced by Hoek and the main idea was to describe the way that rock bolts work. Four different rock bolt spacing designs were simulated and one unsupported model, in order to validate Hoek’s results. The diameter of the blocks was 15 cm while the zones were modelled with length of 5 cm. The tunnel on the shearing case was excavated at the depth of 1500 m. For the stress field, the in-situ stresses of Kiirunavaara mine were considered. The tunnel on the blocky case was excavated at the depth of 30 m and a gravitational stress field was assumed. The shearing model as well as the blocky model, were simulated on a quasi-3D model. The zone length for both cases was 0.1 m. In both cases, a discontinuum non-BBM was modelled first and then, a discontinuum BBM with different rock UCS values was simulated. The discontinuum BBM on the shearing case was simulated for rock UCS of 200, 100, and 50 MPa, while on the blocky case, it was simulated for rock UCS of 50 MPa. The Mohr – Coulomb constitutive model was selected for all three modelling cases. The conclusions of this work were the following: –       The laboratory rock bolt model validated the results of Hoek. Hoek suggested that rock bolt spacing less than three times the average rock piece diameter would be sufficient to produce positive results. The stabilization of the rock pieces as well as the forming of the compression zone were achieved when this equation was satisfied. The geometry of the stabilized material as well as the compression zone, were also correct. –       The discontinuum BBM on the shearing case with intact rock UCS of 200 MPa, produced similar results as the discontinuum non-BBM. This indicates that BBM can be applied for these cases and produce reliable results. The displacement of the fault was expected to be higher than the resulting values. The discontinuum BBM with reduced rock strength (100 MPa and 50 MPa) resulted in rock mass fragmentation. However, the fragmented rock pieces did not detach from the rock mass as the displacement values were not high enough.   –       The discontinuum BBM on the blocky case with intact rock UCS of 50 MPa, produced similar results as the discontinuum non-BBM. There were two discontinuities that affected the smooth transition of the displacement/stress results on the different blocks. The fragmentation of the rock mass due to the existence of the discontinuities did not produce any further rock mass movements.   –       The interaction between rock mass and rock bolts was evident in any modelling case. For the laboratory rock bolt model, the hybrid bolts design was vital for producing correct results. For the shearing model, the hybrid bolts were subjected to shearing movements due to fault movements. In the blocky model, the bolts in the roof of the tunnel were subjected to axial displacements, due to the existence of blocks. The recommendations for further work were the following: –       The hybrid bolts in the laboratory rock bolt test were pretensioned only in the beginning of the computation phase. In reality, the tensioned bolts act at every moment and not only in the beginning. However, it would be interesting to see if the results are similar with continuously tensioned hybrid bolts. It is anticipated that the constantly tensioned hybrid bolts should be able to keep the compressive zones with high values throughout the whole cycling process. Thus, it is suggested for future modellers that this case could be modelled with continuously tensioned hybrid bolts. –       The installation of rock bolts in the shear case as well as in the blocky case, was at the exact same time as the tunnel was excavated.  This is not realistic fact because it is impossible to install the rock bolts exactly the same time as the tunnel excavated. Thus, it is suggested that those two cases could be modelled in the future with more focus on the stress relaxation factor.

According to the Mine Health and Safety Administration (MSHA), between 2006 and 2016, the underground stone mining industry had the highest fatality rate in 4 out of 10 years, compared to any other type of mining in the United States. Additionally, the National Institute for Occupational Safety and Health (NIOSH) stated that structurally controlled instability is a predominant failure mechanism in underground limestone mines. This type of instability occurs when the different discontinuity sets intercept with each other forming rock blocks that displace inwards the tunnel as the excavation takes place, posing a great hazard for miners and overall mine planning. In recent years, Terrestrial laser scanning (TLS) has been used for mapping and characterizing fractures present in a rock mass. TLS is a technology that allows to generate a three-dimensional multimillion point cloud of a scanned area. In addition to this, the advances in computing power throughout the past years, have allowed numerical modeling codes to represent more realistically the behavior of a fractured rock masses. This work presents and implements a methodology that integrates laser scanning technology along with Discrete Element Modeling as tools for characterizing, preventing, and managing structurally controlled instability that may affect large-opening underground mines. The stability of an underground limestone mine that extracts a dipping ore body with a room and pillar (and eventual stoping) mining method is analyzed with this approach. While this methodology is proposed based on a specific case study that does not meet the requirements to be designed with current NIOSH published guidelines, this process proposes a general methodology that can be applied in any mine experiencing similar failure mechanisms, considering site-specific conditions. The aim of this study is to ensure the safety of mine workers and to reduce accidents that arise from ground control issues. The results obtained from this methodology allowed us to generate Probability Density Functions to estimate the probability of rock fall in the excavations. These models were also validated by comparing the numerical model results with those obtained from the laser scans. / M.S. / According to the Mine Health and Safety Administration (MSHA), between 2006 and 2016, the underground stone mining industry had the highest fatality rate in 4 out of 10 years, compared to any other type of mining in the United States. Additionally, the National Institute for Occupational Safety and Health (NIOSH) stated that structurally controlled instability is one of the main causes of rock falls in underground limestone mines. This type of instability occurs when the fractures present in the rock mass intercept each other forming rock blocks that displace into the tunnel as the excavation takes place and poses a great hazard for miners. In recent years, Terrestrial laser scanning (TLS) has been used for mapping and characterizing fractures present in a rock mass. TLS is a technology that allows to generate a three-dimensional multimillion point cloud of a scanned area. In addition to this, the advances in computing power throughout the past years, have allowed simulation softwares such as the Discrete Element Model (DEM) to represent more realistically the behavior of a fractured rock mass under excavation. The aim of this work was to develop and evaluate a methodology that could complement already exisiting design guidelines that may not apply to all kind of underground mines. The presented methodology evaluates rock failure due to presence of discontinuites, through the integration of TLS with DEM and considers site specific conditions. An area of a case study mine was assessed with this methodology, where several laser scans were performed. Information extracted from this laser scans was used to simulate the response of the rock mass under excavation by running Discrete Element Numerical Models. Results from these models allowed us to estimate the probability of rock failure in the analized areas. These, rock block failure probability estimations provide engineers a tool for characterizing, preventing, and managing structurally controlled instability, and ultimately improving workers safety.

Die Rotliegend-Vulkanite aus dem Norddeutschen Becken (NDB) sind in jüngster Zeit stärker in den Fokus geothermischer Betrachtung gerückt. Ein wichtiger Meilenstein war das Forschungsprojekt Groß Schönebeck, bei dem wider Erwarten die Vulkanite als sekundärer Zielhorizont die besten Voraussetzungen für eine geothermische Erschließung boten. Über die hydraulisch/mechanischen Eigenschaften der Vulkanite im Untergrund ist jedoch kaum etwas bekannt und auch die Oberflächenäquivalente sind hinsichtlich geothermisch relevanter Parameter weitgehend unerforscht. Aus den positiven Erfahrungen des Standorts Schönebeck entstand die Motivation einer umfangreichen Analyse der Rotliegend-Vulkanite mit Blick auf eine tiefengeothermische Nutzung. Es wurden thermische, felsmechanische und petrophysikalische Untersuchungen von sieben Oberflächenäquivalenten durchgeführt; drei der Oberflächengesteine sowie zwei Tiefenbohrungen wurden ferner hinsichtlich auftretender Kluftmuster analysiert. Die Daten fungieren als Eingangsparameter für hydraulische sowie hydromechanische numerische Modellierungen zur Potenzialabschätzung und zum Prozessverständnis. Die thermische Analyse der Gesteine ergab eine hohe Wärmeleitfähigkeit für die quarzreichen und dichten Vulkanitvarietäten. Durch die Wärmekapazität und die Reservoirtemperatur wurde das technische Strompotenzial für die Eruptionsstadien ermittelt. Das größte Potenzial liegt im explosiven Ignimbritstadium und im Post-Ignimbritstadium und wird auf einen Wert geschätzt, der allein dem 20-fachen des deutschen Jahresstromverbrauchs entspricht. Regional betrachtet ist das größte Potenzial bei Standortwahl im zentralen östlichen NDB zu erwarten. Die untersuchten Vulkanite sind überwiegend dicht und erfordern Stimulationsmaßnahmen für eine erfolgreiche Erschließung. Auch die stärker porösen Tuffe erreichen nicht die erforderliche Matrixpermeabilität für einen Porenleiter. Triaxiale Druckversuche unter in-situ Spannungsbedingungen haben jedoch gezeigt, dass es nur bedingt möglich ist, Risse im intakten Gestein zu erzeugen. Man ist folglich auf eine gestörte Kruste, also Klüfte im Gestein angewiesen. Sowohl die Oberflächengesteine als auch die Vulkanite im Untergrund sind nachweislich geklüftet. Das tektonische Grundmuster beschreibt Klüfte, die NW-SE bis NNW-SSE sowie NE-SW bis NNE-SSW orientiert sind und dabei steil einfallen. Die Scherfestigkeitskriterien der Kluftflächen liegen deutlich unterhalb derer für das intakte Gestein, so dass die Bedingung für eine Aktivierung der Klüfte im Spannungsfeld des NDB positiv bewertet wird. Die Kluftdaten wurden zum Zwecke numerischer Modellierungen in diskrete Kluftnetzwerkmodelle überführt. Hydraulische Modellierungen ergaben eine bevorzugte Fließrichtung in NW-SE. Die mit der Tiefe zunehmende Kluftschließung führt zu einer Durchlässigkeit, die für eine geothermische Nutzung nicht ausreichend ist, das Gestein muss hydraulisch stimuliert werden. Eine Stimulation der Kluftflächen zur Steigerung der Fließrate wurde mittels hydromechanischer Modellierungen erfolgreich dargestellt. Die wichtigsten Kriterien für eine erfolgreiche Stimulation sind die Geometrie des Kluftsystems und die Orientierung des Spannungsfelds. Aufgrund der überwiegend vertikalen Kluftflächen im Vulkanit und der hohen Vertikalspannung im tiefengeothermischen Reservoir wird eine Erschließung über das Multiriss-Konzept empfohlen. Durch den in der vorliegenden Arbeit dargestellten methodischen Ansatz kann mittels repräsentativer Eingangsparameter für einen Standort entsprechend der notwendige Injektionsdruck sowie die Art und Intensität der Verformung der Kluftflächen für eine hydraulische Stimulation prognostiziert werden.

The stability of the high-walls is one of the major concerns for open pit mines. Among the various factors affecting the stability of high-walls, blast vibrations can be an important one. In general, worldwide the established respective government regulations and industry standards are used as guidance to determine the maximum recommended levels of the peak particle velocity and frequency from the blast to avoid any effects on the structures around the mining project. However, most of the regulations are meant for buildings or houses and do not concern high-walls. This thesis investigates the response of high-walls under the effects of vibrations from mine blasting. In this research, the relationship between the high-wall response, the geometry of the slope, the frequency and the amplitude, of the ground vibration produced by blasting, is explored using numerical models in 3DEC. The numerical models were calibrated initially with data collected using seismographs installed in a surface mine operation and recording vibrations produced by an underground mine drill and blast operation. Once the calibration was accomplished, a parametric study was developed to explore the relationships between various parameters under study and its impact on the stability of high-walls.

Rock slope stability is considered as one of the most important fields in rock engineering. Developments of computation facilities and increase in application of sophisticated mathematical concepts in engineering problems have also affected the methods of slope stability analysis. In recent years, the numerical modeling methods have extensively applied instead of limit equilibrium methods. Also, the probabilistic methods are considered in rock slope designs to quantify the uncertainties of input effecting variables. In this research, a probabilistic-numerical approach was developed by integration of three dimensional Distinct Element Method (DEM) and probabilistic approach to analyze the stability of discontinuous rock slopes. Barton models have been used to model the behavior of rock discontinuities and the shear strain was considered as failure indicator of discontinuities. The proposed methodology was applied to a rock slope in Amasya, Turkey where the Joint Roughness Coefficient (JRC) was considered as the main random variable. The effect of basic friction angle and cohesion of joints infilling material and its strength reduction due to weathering were included in the analysis. In the slope the shearing behavior of fourteen discontinuities and the failure probability of each block were investigated, and the corresponding Reliability Index (&beta / ) was derived for each of the discontinuities.

The aim of the EU BIOMOre Project is to investigate the potential to extract copper from Sandstone formations in the North-Sudetic Trough which lies along the border between Poland and Germany. A new mining concept called bioleaching shall be applied in thin and very low permeable copper mineralization zones (order of 0.2 mD). Briefly, bioleaching process is the injection of a lixiviant (sulphur acid containing ferric iron) and then extraction of a pregnant leach solution through boreholes at the ground surface. This concept requires another special technique which is called hydraulic stimulation. Cracks along a wellbore are generated by pumping large quantity of fluid under high pressure into a cased section of rock during a hydraulic stimulation. This work at hand focuses on the geotechnical methods and scientific-engineering approaches used for extracting copper from very thin mineralization zones. The geological setting with faults and in situ stress state of the exploration zone is generated using measurements, visualised by 3D CAD model (RHINO), and computed via the Discrete Element code 3DEC. The preliminary drilling (stacked dual lateral wellbore) and stimulation design (plug-and-perf completion) are selected based on comprehensive literature survey and industry-based consultancy. In order to calibrate the calculated stress state in 3D, candidate sites for the hypothetical drilling-stimulation are detected using 2D GIS map (QGIS) at CAD model (RHINO). Trend of calculated stresses is in good agreement with the measured ones (σH > σv > σh). The final decision of selecting a drilling-stimulation site is made by using both GIS map and 3D CAD model. A hypothetical drilling-stimulation can be performed up to the depth of 1564 m in the Rotliegend & Grauliegend Sandstone with shale, which is overlain by (Zechstein) Limestone. During a possible stimulation, limestone’s integrity as a caprock and as a stress barrier is of great importance in addition to connect two lateral wellbores for facilitating flow of lixiviant. The preliminary geometrical design of stimulation is set with the cluster spacing (distance between fractures) of 20 m. Subsequent to final cost estimation of selected preliminary drilling-stimulation design, it is decided to use pinpoint (1,200,000 Euro) instead of plug-and-perf completion (2,345,300 Euro) since it is more economical. A possible drilling operation is anticipated to cost approximately 9,000,000 Euro. The 3D in situ stress model is calibrated before transferring of stress state into the sub-model which is used to optimise the selected stimulation design. The results of the last (DEM) sub-model are employed to reduce costs, to enhance the connection between branches of wellbores for bioleaching and to hinder possible penetration of fractures into the caprock. The preliminary geometrical design of stimulation is then modified based on these calculation results while increasing the cluster spacing from 20 m to 40 m. This is performed due to high stress-shadows (alteration of the stresses between fractures in a stimulation) encountered at the preliminary calculations. Results showed that, after the 80 seconds injection duration of water with 0.16 m3/sec into the sandstone, two wellbore laterals are expected to be connected by three generated cracks. They exhibit average aperture and transmissivity of 4.1 mm and 5.8 . 10-8 m2/sec, respectively. Furthermore, fracture initiation pressure ranges between 30 – 35 MPa at the drilling depth. The conclusions can be drawn that through the assessment of 3D CAD, GIS, and numerical DEM modelling methods, approximately 49% of cost reduction can be achieved by employing pinpoint instead of plug-and-perf completion. That is an important proof of the systematically approach for a stimulation planning wherein all necessary phases such as in situ stress estimation, modelling and cost assessment should have been considered. This work can be considered as a milestone for studies of stimulation designs which has been newly initiated in the EU-Region as a promising method for efficiency considering unconventional ore extraction. Moreover, this dissertation revealed again the emerging importance of integrated geotechnical information systems analogous to BIM (Building Information Systems).:LIST OF FIGURES LIST OF TABLES NOMENCLATURE ABSTRACT ZUSAMMENFASSUNG ACKNOWLEDGEMENTS 1. OUTLINE AND OBJECTIVE OF THE DISSERTATION 2. STATE OF THE ART 2.1. INTRODUCTION TO STIMULATION TECHNOLOGIES, EQUIPMENT AND DESIGNS 2.1.1. Technical instruments and frac-materials 2.1.2. Wellbore completion designs 2.1.3. Location and orientation of a wellbore 2.1.4. Fracture placement designs 2.1.5. Summary and conclusions 2.2. MEASUREMENT AND MODELLING OF UNDERGROUND STRESS FIELD 3. DETERMINATION AND MODELLING OF IN SITU STRESS FIELD IN THE NORTH SUDETIC TROUGH 3.1. GEOLOGICAL SETTING OF THE MODELLED REGION 3.2. SIMULATION OF THE IN SITU STRESS FIELD 3.2.1. Determination of the stress regime by measurements 3.2.2. Stepwise procedure of the stress field modelling 3D CAD assisted structural model of geological setting 3D DEM model for stress field simulations 2D GIS maps used for detection of drilling-stimulation sites 4. DRILLING AND WELLBORE DESIGN CALCULATIONS WITH COST ESTIMATION 4.1. DESIGN CALCULATIONS AND TECHNICAL REQUIREMENTS OF DRILLING AND WELLBORE 4.2. ECONOMICAL EVALUATION OF THE SELECTED DRILLING AND WELLBORE DESIGN 5. MODELLING OF THE HYDRAULIC STIMULATION AT THE SELECTED DRILLING SITE IN SANDSTONE 5.1. FINAL CALIBRATION OF THE 3D STRESS FIELD MODELS 5.2. DISCRETE ELEMENT MODELLING OF THE STIMULATION DESIGN AT THE SELECTED DRILLING SITE 5.3. DESIGN OPTIMIZATION STUDY OF THE STIMULATION MODEL AND FINAL COST ESTIMATION 6. SUMMARY AND CONCLUSIONS REFERENCES APPENDIX-A APPENDIX-B APPENDIX-C