Shallow excavations through variable rockmasses in urban centers present significant design challenges, whether considering small diameter tunnels for utilities or large span underground caverns. In designing shallow excavations in urban environs, it is especially critical to minimize the impact of the excavation on surface. In small diameter projects, minimal surface disturbance is often achieved by the employment of TBMs as the excavation method. While reducing the risk of surface subsidence due to displacements in front of the face, TBM progress is sensitive to variable ground conditions and the TBM design must be appropriately matched to the expected geology. Sufficient understanding of the geology and development of geological models are critical in the selection of an appropriate TBM and cutting tools. In this study, recent projects in Calgary, AB are used to highlight the challenges faced with using TBMs through sedimentary rock with distinct, variable units. In larger scale projects, long term excavation stability is critical in the reduction of surface disturbance. Due to the low confining stresses, structural failure is often the primary failure mode in shallow excavations, especially within fractured, heterogeneous rockmasses. In these cases, numerical methods are often used in excavation design. The ability of numerical methods to capture the expected failure modes of shallow excavations through layered rockmasses is explored, with an emphasis placed on the ability of support elements to reduce shear slip for increased stability. Passive bolt models are analysed using both 2D and 3D numerical models to adequately capture the behaviour of a passive support system in shear. The shortcomings of some current support models are discussed, and modifications are suggested. / Thesis (Master, Geological Sciences & Geological Engineering) -- Queen's University, 2012-09-25 20:56:52.083

This thesis presents numerical and analytical investigation of the geomechanics underlying longwall mining. It was tried out to study the disturbances induced by longwall mining in nearby rocks and their influence on the stability of the gates, pillars and main tunnels of longwall mines. The thesis consists of two major parts: numerical and analytical investigations. The study site is the Svea Nord coalmine, Svalbard, Norway. A novel algorithm was proposed for numerical simulation of the longwall mining process. In the proposed algorithm progressive cave-in and fracturing of the roof strata, consolidation of the cave-in materials and stress changes are simulated in detail. In order to outline the caved-in roof rocks a criterion based on maximum principal strain (in tension) was used. The critical tensile strain of roof cave-in was determined through back-calculation of the surface subsidence above a longwall panel at the mine. The results of the simulations were then used to analyse stress changes induced by longwall mining and the stability of gates. The simulations revealed that the stability of the gates and the loading to the rock bolts are closely related to the width of the chain pillars. With slender pillars, shear displacements along weak interlayers and bedding planes result in heavy loading to the rock bolts. Therefore, the locations of weakness zones should be taken into account in rock bolt design. The developed algorithm was implemented to study the loading and stability of the barrier pillar of the mine. The barrier pillars protect the main tunnels and border area of the mine from disturbances induced by longwall mining in the panels. The simulations show that the stresses in the barrier pillars fluctuate up and down during mining because of periodic cave-in events behind the longwall face. A failure zone of about 12 m exists in the wall of the barrier pillars. A large portion of the barrier pillar is still intact and is, thus, capable of protecting the border area. The results of the detailed simulations of longwall mining via the developed algorithm were, also, implemented in a large-scale numerical model. The model consists of all of the longwall panels and the border area of the mine. It is intended that the coal in the border area on the other side of the longwall panels will be mined after completion of the longwall mining. There is concern about how the longwall mining affects the stress state in the border area and how stress changes would affect future mining in the border area. A failure zone of about 20 m developed in the wall of the main tunnels on the side of the border area after all the longwall panels were mined out. The stress state in the remaining portion of the border area remains unchanged. Therefore, it will be possible to mine the border area in the future. In order to investigate the roof strata cave-in mechanism in detail a discontinuous numerical simulation of roof cave-in process was conducted by UDEC code. The block size in the roof strata and the mechanical parameters of the discontinuities were obtained through back-calculations. The back-calculations were conducted with a statistical method, Design of Experiment (DOE). Numerical simulations revealed that jointed voussoir beams formed in the roof strata before the first cave-in. Beam bending results in stress fluctuations in the roof strata. The maximum deflection of a roof stratum at the study site before the first cave-in is about 70% of the stratum thickness. The simulations and field measurements show no periodic weighting on the longwall shields in this mine. Numerical sensitivity analyses show, however, that periodic weighting may occur in strong roof strata. Roof strata with a high Young’s modulus and large joint spacing are not suitable for longwall mining. The maximum sustainable deflection of the roof strata before cave-in depends upon the horizontal in-situ stress state. It slightly increases with the in-situ horizontal stress in the stratum beams, but the horizontal stress would increase the possibility of rock-crushing in deflected roof beams. The implemented numerical method would be useful in assessment of the cavability of the roof strata and in selection of longwall shields with adequate load capacity. As shown through discontinuous numerical simulations, the roof strata above the underground opening constructed in the stratified rocks form voussoir beams. The stability of those beams is the major concern in the study of the gate stability and roof cave-in assessment in the longwall panels. Two different analytical methods were developed for cases with and without the in-situ horizontal stress acting along the beams. In the analytical model for the beams without horizontal stress a bilinear shape was assumed for the compression arch generated within the voussoir beams. The stability of the compression arch is governed by the energy method. The model requires an iterative procedure for convergence, and an algorithm was proposed for it. The analytical method was verified with numerical simulations by means of a discrete element code, UDEC. For the beams subjected to in-situ horizontal stress, the classic beam theory was employed to drive the analytical solution for it. The superposition method was used to obtain bending/deflection equations of the beam. The validity of both the assumptions and the developed method were, also, investigated by numerical simulations. The developed analytical method revealed that high Young’s modulus of a beam rock increases the stability of the beams against buckling but it causes higher stress within the compression arch which increases the probability of crushing failures in the beam abutments and midspan. In-situ horizontal stress along beams increases their stability against buckling and abutment sliding failure, but it raises the possibility of crushing failure at the abutments and the midspan.

The effect of corrosion on the performance of rock support and reinforcement in Australian underground mines has not been widely researched and is generally not well understood. This is despite the number of safety concerns and operational difficulties created by corrosion in reducing the capacity and life expectancy of ground support. This thesis aims to investigate corrosion and relate how the environmental conditions in Australian underground hard rock mines impact on the service life of rock support and primarily rock reinforcement. Environmental characterisation of underground environments was completed at a number of mine sites located across Australia. This provided an improved understanding of the environmental conditions in Australian underground hard rock mines. Long-term testing on the impact of corrosion on the load bearing capacity of reinforcement and support under controlled experimental conditions was conducted in simulated underground environments. Rock reinforcement elements were examined in-situ by means of overcoring of the installed reinforcement and surrounding rock mass. Laboratory testing of the core determined changes in load transfer properties due to corrosion damage. These investigations provided an excellent understanding of the corrosion processes and mechanisms at work. Corrosion rates for a range of underground environments were established through the direct exposure and evaluation of metallic coupons in underground in-situ and simulated environments. / It was found that the study of corrosion is challenging due to the time required to gather meaningful data. In particular, the wide range of materials that comprise ground support systems means that it is impossible to examine all the possible combinations of variables and their potential influence on the observed levels of corrosion and measured corrosion rates. Despite these challenges, the systematic investigation has resulted in new corrosivity classifications for both groundwater and atmospheric driven corrosion processes for various reinforcement and support systems used in the Australian underground mining industry. Previous corrosivity classifications were not found applicable. Furthermore, these new corrosivity classifications are simpler than previous classifications and corrosion rates may be predicted from readily obtained measurements of ground water dissolved oxygen and atmospheric relative humidity. Different types of reinforcement and surface support systems have been rated with respect to their corrosion resistance and estimates have been made for the expected service life for various rates of corrosion.

Swelling clay is an important and challenging geological feature in tunnelling projects.The behaviour of swelling clay is influenced by many internal and external factors.This, combined with a lack of experience and knowledge has resulted in challengeswhen accounting for the swelling phenomena. In addition to this, there is a lack ofunderstanding for the calculation of rock support when swelling clay is present. Themain objectives of this thesis were to compare laboratory methods used to measure theswelling potential, identify possible uncertainties when performing oedometer tests,and to provide a model for design of rock support for single zones with swelling clay.To support these objectives, a literature review has been conducted and laboratory testshave been performed to determine swelling properties using clay samples collectedfrom a tunnel failure in Sweden. In addition, a model has been derived for design ofrock support when swelling clay is present in single zones of different widths. Theresults from the laboratory tests showed that the clay sample could be characterizedas active respectively moderately active clay. The results of the model for design of rocksupport showed that the width of the swelling zone and the swelling pressure have asignificant impact on the shotcrete thickness, rebar spacing and reaction forces at therock bolts. However, future research for design of rock support is required as well asdevelopment of a more detailed standard for testing of swelling clay.

Vid tunneldrivning inom infrastrukturprojekt i Sverige är den mest förekommande bergbulten för permanenta förstärkningar den konventionella kamstålsbulten. En ingjuten bult utan förspänning. Montaget sker genom att med handkraft pressa in bulten i ett cementfyllt borrhål. Detta ger upphov till spill av cement och stor fysisk ansträngning, när tusentals bultar monteras i varje projekt. Den permanenta bergssäkringen uppnås först när den omslutande cementen har härdat. Detta medför ett moment i bergförstärkningsskedet där entreprenören måste säkra tunneln med temporära driftbultar.   För att undvika temporärbultning kan en förspänd ingjuten bult användas som permanent bult. Exemplet på en sådan bult är PC-bulten. Vid montage ger den en omedelbar driftsäkring och en permanent förstärkning genom injektering via det inre röret. Bulten blir då helt ingjuten av cement.  Syftet med studien är att jämföra PC-bulten med kamstålsbulten. Studien undersöker den förmodade nyttan i belastningsergonomi och ekonomi av att använda PC-bulten som permanent förstärkningsbult. Genom att utföra en styrd observation på ett studiebesök där tunneldrivning med bergförstärkning skedde har de båda bultarnas montagearbete jämförts ur belastningsergonomisk synpunkt. För att utreda skillnader i arbetsvolym och tidsåtgång utfördes beräkningar på ett sammanställt bultprotokoll. Beräkningarna visar vilken förstärkningsbult som medför den lägsta kostnaden för bultförstärkning.  Av resultatet i studien framgår det att det finns en stor potential i ett framtida användande av PC-bulten som permanent bult vid tunnelförstärkning. Potentialen ligger i en klart förbättrad arbetsmiljö i avseendet belastningsergonomi och i den tidsvinst framtida projekt kan nyttja. / The conventional rebar bolt, a grouted bolt with no pre-stress, is the most commonly used rock reinforcement bolt for tunneling in Swedish infrastructural projects. The assembly takes place by manually pushing the bolt into a borehole filled with grout. This gives rise to spillage and great physical exertion when thousands of bolts are mounted in each project.   Permanent safety is only achieved when the surrounding cement has cured. This entails a moment in the rock reinforcement stage where the contractor must secure the tunnel with temporary operating bolts.  To avoid temporary bolting, a pre-stressed bolt can be used as a permanent bolt. An example of such a bolt is the PC-bolt. When assembled, it provides immediate operational safety and permanent reinforcement by injection through the inner pipe. The bolt is then completely sealed with grout.  The purpose of the study is to compare the PC-bolt with the rebar bolt. The study explores the supposed benefit of economics and load ergonomics of using the PC-bolt as a permanent reinforcement bolt. By conducting a controlled observation on a field study to a site where tunneling with rock reinforcement occurred. The assembly of the two bolts has been compared out of a load ergonomic view point. In order to investigate differences in work volume and time, calculations were made on a compiled bolt protocol. The calculations show which reinforcement bolt results in the lowest cost of the assembly.  The result of the study shows that there is a great potential in the future use of the PC-bolt as a permanent bolt for tunneling. The potential lies in a clearly improved work environment regarding ergonomic loads and more time efficient installation that can be utilized in future projects.

Prior to an underground excavation a site investigation is carried out. This includes reviewing and analysing existing data, field data collected through outcrop mapping, drill core logging and geophysical investigations. These data sources are combined and used to characterise, quantify and classify the rock mass for the tunnel design process and excavation method selection. Despite the best approaches used in a site investigation, it cannot reveal the required level of detail. Such gaps in information might become significant during the actual construction stage. This can lead to; for example, over-break due to unfavourable geological conditions. Even more so, an underestimation of the rock mass properties can lead to unplanned stoppages and tunnel rehabilitation. On-the-other-hand, the excavation method itself, in this case, drill and blast, can also cause severe damage to the rock mass. This can result in over-break and reduction of the strength and quality of the remaining rock mass. Both of these attributes pose risks for the tunnel during excavation and after project delivery. Blast damage encompasses over-break and the Excavation Damage Zone (EDZ). In the latter irreversible changes occur within the remaining rock mass inside this zone, which are physically manifested as blast fractures. In this thesis, a number of methods to determine blast damage have been investigated in two ramp tunnels of the Stockholm bypass. Herein, a comparison between the most common methods for blast damage investigation employed nowadays is performed. This comparison can be used to select the most suitable methods for blast damage investigation in tunnelling, based on the environment and the available resources. In this thesis Ground Penetrating Radar, core logging (for fractures) and P-wave velocity measurements were applied to determine the extent of the blast damage. Furthermore, the study of the two tunnels in the Stockholm bypass shows a significant overestimation of the actual rock mass quality during the site investigation. In order to gain a more accurate picture of the rock mass quality, Measurement While Drilling (MWD) technology was applied. The technology was investigated for rock mass quality prediction, quantifying the extent of blast damage, as well as to investigate the potential to forecast the required rock support. MWD data was collected from both grout and blast holes. These data sets were used to determine rock quality indices e.g. Fracture Indication and Hardness Indicator calculated by the MWD parameters. The Fracture Index was then compared with the installed rock support at the measurement location. Lastly, the extent of the damage is investigated by evaluating if the MWD parameters could forecast the extent of the EDZ. The study clearly shows the capability of MWD data to predict the rock mass characteristics, e.g. fractures and other zones of weakness. This study demonstrated that there is a correlation between the Fracture Index (MWD) and the Q-value, a parameter widely used to determine the required rock support. The study also shows a correlation between the extent of the blast damage zone, MWD data, design and excavation parameters (for example tunnel cross section and charge concentration).

The Garpenberg mine, owned by the Boliden Mining group, has established a trial area at Dammsjön orebody in order to examine the possibility of increasing the productivity of the mine. The mine uses the rill mining method with a current rill height of 15 m. In order to increase the productivity, the mine is examining the possibility of increasing the height of the rill. The trial area is located at 882 m depth surrounded by dolomite on the hangingwall and quartzitic rock on the footwall side. Rock support arches have been installed, in addition to the regular support pattern, to test their effectiveness on stabilizing the ground around the drifts. The arches have been installed in every 6 m and every 3 m in different parts of the test area. Rock samples from the trial area were brought to the university laboratory for testing. The data gathered from the laboratory tests along with the data from the monitoring of the trial area were used to develop a calibrated numerical model. A three-dimensional (3-D) model was therefore created, by using the FLAC3D numerical code. After the calibration of the model a parametric study was conducted for different rill heights and different arch spacing to investigate the performance of the arches. Specifically, the case of no arch installation along with the cases of an installed arch every 6 m and 3 m were tested, for the rill heights of 15 m, 20 m, 25 m and 30 m. The study concluded that the arches assisted in reducing the ground convergence in the production drift. The results also showed that the total height of the rill bench yields regardless of its height. After the yielding, the rockmass can no longer support itself and caves under its own weight. The larger the rill height, the larger the volume of loose rock that has to be supported and thus, higher the convergence. Furthermore, it was also observed that, significant amount of convergence in the production drift occurred during the drifting of the top drive and less during the stoping of the rill bench. This indicates that, the timely installation of the arches is an important criterion for their performance.