Search

Rock joints will undergo a sequence of cyclic shearing loadings during a seismic event. However, the effect of cyclic shear loading on the energy-absorbing rock bolts has never been studied before. Laboratory shear experiments were carried out to study the shear behaviour of rock joints reinforced by the energy-absorbing rock bolts under cyclic loading condition. The results illustrated that the support effect of the energy-absorbing rock bolts was very small after the first cycles in the cyclic shear experiments. In the case of small cyclic distances, the shear resistance of the energy-absorbing rock bolts will gradually recover after the shear displacement has exceeded the cyclic distance in the subsequent shear experiment after 5 cycles. In the case of large cyclic distances, no recovery of shear resistance was found in the subsequent shear experiment, indicating that the energy-absorbing rock bolts had completely lost its supporting role after cyclic shear loading. A new index of shear energy loss ratio (SELR) was proposed to evaluate the shear behaviour of energy-absorbing rock bolt and rock joint under cyclic shear loading condition. The results showed that the SELR of rock joints was commonly less than 20%. However, the SELR of rock bolts could reach nearly 100% when the cyclic distance was larger than 8 mm. When the cyclic distance was 4 mm or 6 mm, the SELR of the fully encapsulated rock bolts almost reached 100%. However, the SELR of the energy-absorbing rock bolts were located in the range of 50–80% for the same condition. The results indicated that the shear behaviour of a rock bolt inserted in a rock joint was strongly influenced by cyclic shear loading. The shear performance of the energy-absorbing rock bolts was better than the fully encapsulated rock bolts under cyclic shear loading conditions.

The present study focuses on the load transmission law along the rock bolt and the efficacy for different anchored conditions. Shear displacement along the joint face is considered on the basis of traditional rock bolt stress calculation method with regard to joint open displacement. Combining the elastic foundation beam model and the rock bolt pull-out model, an analytical solution for the stress distribution along the rock bolt through a single joint face is derived. The performance of the proposed analytical solution is compared with an experimental result and a previous study. In addition, joint displacement, joint location, and dip angle of the joint face are set as variables to further explore the corresponding anchored effect. The results indicate that larger joint displacement along the joint face tend to cause a distinct increase of the shear and compressive stress along the rock bolt. When the joint face moves inward, the interaction force between the rock bolt and the surrounding rock mass decreases, suggesting that the strength of the rock bolt is not taken full use. Moreover, for a certain joint displacement and joint location, larger anchored angle between the rock bolt and the joint face contributes to better reinforcement effect.

Steel tube is used to simulate rock mass in laboratory impact tests of rock bolts. The steel tube must be equivalent to the borehole in the rock mass in radial stiffness in order to obtain valid test results. The stiffness of the steel tube is directly associated with the thickness of the steel tube. The importance of the tube thickness has not been paid enough attention in the laboratory tests of rock bolts at present. This technical note derives a formula for calculating the required thickness of the steel tube used for laboratory tests of rock bolts. It also provides a formula for calculating the stiffness of the steel tube under plane stress which is the realistic loading condition of rock bolts in the field. The analysis reveals that the grout in the tube hole does not play a role in the calculation of the thickness of the steel tube. The required thickness of the steel tube is solely dependent on the Young's moduli and Poisson's ratios of the tube steel and the rock mass to be simulated.

A high spatial resolution distributed optical fiber strain sensing technology is demonstrated to advance the assessment of support element behaviour during in-situ pull tests. A technique to instrument typical tendon support elements with a fiber optic sensor is discussed and was trialed at an underground salt mine through a series of pull tests on instrumented rebar elements. It is shown that a continuous strain profile can be measured along the length of a support element under pull test load, which, in turn, allows the interfacial shear stress distribution and deformation of the support element to be resolved. In comparison to load and displacement measurements solely at the support element head, which is what is traditionally reported for in-situ pull tests, the optical fiber strain sensor is found to provide significantly more insight into the mechanistic behaviour of the support element by readily measuring highly variable strain distributions and by quantifying the load development length of the support element. The sensor arrangement allowed the conventional handling and installation procedures to be followed without any implication to the integrity of the sensor. Accordingly, an instrumented reinforcement element can be installed with minimal interruption to on-going mining/construction procedures at the given project.

We have proposed a theoretical model for the analysis of bolt anchorage effects through cross-section uniformity and elastic theory. When the surrounding rock and soil are fixed, then the slip resistance forces directly proportional to the displacement of the bolt. An equilibrium differential equation for the wholly bolt has been established. The axis stress expression has been presented from the combined boundary conditions. It is a monotonic function which consists of negative index and minus index function of the distance, away from the force. The effects of friction proportional coefficient and radius have been analyzed as well. For local stripping bolt, a different model having separate parts has also been analyzed. We have also deduced the axis stress distribution function. Moreover, the curves of different separate conditions were compared which conclude that there is a terrace for the separate part. The longer the separate part, the quicker the axis stress decreased. Finally, an experiment has been performed which show the distribution of axis stress has the same trend with the strain gauge record. The theory solution of axis stress is also proved to be reasonable.

Analysis of mechanical behaviour of rock mass reinforced by fully grouted rock bolt is introduced based on the enriched finite element method (EFEM). A solid element intersected by a rock bolt along any arbitrary direction is defined as ‘enriched’ element and it has additional degrees of freedom at each node for estimating displacements and stresses in bolt rod. Numerical procedures of EFEM are developed to form enriched stiffness matrix using constitutive relations of rock mass, properties of bolt rod and grout material and orientation of the bolt. Decoupling of rock bolt and elasto-plastic behaviour of rock mass has been incorporated into the EFEM procedures. Released displacement of rock mass prior to bolt installation has also been considered in the procedures to provide realistic solution. The results of this method are verified with an analytical pull-out test model. In addition, a numerical example of a bolted tunnel is provided to demonstrate the efficacy of the proposed method for practical applications.

A continuous dynamic model for the axial vibration of a rock bolt system is presented. The model comprises three sections: the fixed length, bonded into the rock, the free length, which is not coupled to the rock, and the protruding length, which extends beyond the rock. The head assembly is modelled as a discrete mass and a spring, and a further discrete mass is included, representing a testing device that can be attached to the protruding end. Each section is modelled as a continuous elastic rod governed by the wave equation, with suitable compatibility conditions applied between the sections and boundary conditions, which also account for the effect of the discrete components, applied at the ends. Solutions in non-dimensional form are substituted into the boundary conditions to allow the natural frequencies to be calculated, and it is shown that two possible solutions for the mode shapes can be used for the fixed length—an exponential solution or the classical sinusoidal solution—depending on the stiffness of the grout relative to that of the bar. The conditions for which the two solutions are valid are developed, and changes in the frequency ratio with changes in length ratio, and the stiffness ratios of the grout and the anchor head relative to the stiffness of the fixed length of the anchorage are examined. Generally, the state of a bolt after installation is unknown and this does not provide proper assurance of the safety of the structure for which the bolts are used. The model provides a viable tool for helping to assess the condition of the bolt by using the natural frequencies associated with areas of the bolt of particular interest, e.g. the free length. The results show how the changes in the stiffness and/or length ratios affect the dynamics associated with fixed length of the bolt and the quality of the bonding installation. A case study is presented showing how the model can be used effectively to interpret real data.

An analytical model for rock bolts has been developed based on an improved Shear–Lag Model. The development of the model is based on the description of the interaction behavior of the rock bolt, the grout medium and the rock mass. On the basis of the model, the coupling and decoupling behavior of the rock bolt in pullout tests, uniform deformation the rock mass and intersecting joints are analyzed. The pullout test characteristic is described by the proposed model, and a back analysis method is proposed to calculate the shear strength of the interface media. For the rock bolts in a deformed rock mass, the influence of the installation time of the rock bolt has been taken into account, and the theoretical prediction is verified by the measured data. According to the proposed model, the position of the neutral point is not only related to the length of the rock bolt and the radius of a tunnel, but also is strongly influenced by the mechanical properties of the rock mass. Analysis of the joints intersecting the rock bolt shows that there may be more than one neutral point on the rock bolt, and the prediction of the simplified model is consistent with the pullout model. By using this model, a method is proposed to analyze the interaction behavior of the rock bolt and the surrounding rock mass, with a way of evaluating the supporting performance quantitatively.

An interaction model is proposed to describe the interaction between the energy-absorbing rock bolt and the rock mass. Based on the plane-strain axial symmetry assumption and the incremental theory of plasticity, the equilibrium equations and compatibility equations of rock mass, as well as the response of the energy-absorbing rock bolt are deduced theoretically. The proposed method was programmed in a Visual Basic environment, and a semi-analytical solution for the coupling model was achieved. The reinforcement mechanism of the energy-absorbing rock bolt in conventional tunneling is clearly demonstrated through an illustrative case study. The reinforcement effect of the energy-absorbing rock bolt under different conditions was estimated quantitatively, and its mechanical work transfer ability is presented. In addition, the validity of the proposed method was verified through numerical simulations. Finally, a number of derivative cases were investigated to reveal the influence of the bolt and rock properties on the reinforcement effect and the bolt work transferred on the rock mass. In the case of higher in-situ stress or low-strength rock mass, the support effect of the energy-absorbing rock bolt is significantly improved, and the bolt absorbs more energy. Increasing the bolt installation density could always be helpful for the stabilization of the surrounding rock mass. However, additional rock-bolt length could hardly affect ground reinforcement because the bolt section embedded in the elastic region of the rock mass could barely help to constrain the elastic displacement release. The bolt should be installed no later than the stage of critical inner pressure, namely when the plastic region occurs., International Journal of Rock Mechanics and Mining Sciences, 103, pp.168-178; 2018

Longwall mining is the most widely used mining method in underground coal mining. The stability of gateways is crucial for longwall mining as they provide many essential services such as transportation of excavated coal, ventilation of longwall face. High in situ stresses , mining induced stresses due to advancing longwall face, and poor rock mass quality leads to large deformation jeopardizing the stability of the gateway in Guobei coal mine. In this paper, the global-local modeling approach is used for analysing the stability of the gateway in the mine. In this approach, a mine scale 3D global model was constructed to extract the stresses acting on the gateway. The extracted stresses from global model were implemented into 2D local model of the gateway for detailed analysis. The used approach was verified by the field measurements taken from the modeled gateway. In an attempt to deal with large mine induced displacements, a support system composed of rock bolt, cable bolt and shotcrete was proposed. The verified model was used to evaluate the performance of proposed support system. It was concluded that the proposed support system can significantly reduce the displacements. The verified model facilitates the performance evolution of other alternative support systems considered by the mine management. The presented approach can be applicable for other deep underground coal mines, experiencing large gateway convergence.

Some mountain tunnels suffer from perennial deformation induced by overburden and confining stresses. Conventionally rock bolts are being used as a countermeasure . Oftentimes the conventional rock bolts fail to deliver their optimum performance, owing to several potential causes. Such as: the ground around the tunnel is too plasticized while transmitting shear forces along with the solidified mortar; and the rock bolts are poorly anchored with grout at the rear end. Considering these facts, we develop a bonding of rock bolt, named CIB-bolt, a combined mechanism deals with ground Improvement, to improve the plasticized ground while preventing further deformation of tunnel and to achieve positive bonding of rock bolts. To confirm the effectiveness of the developed rock bolt which was utilized in the experiment model, a real field application and numerical analysis also were conducted. Observed results indicate that, the developed rock bolt exhibits excellent performance in transmitting axial stresses towards the hard layers of rock. This mechanism can prevent the yielding or plasticization of adjacent ground and control the deformations of tunnel lining .

Rockburst is one of the leading disasters in mining and rock engineering causing serious casualties and property losses. A major engineering challenge for underground roadways in mines experiencing significant seismicity is the performance of the support system. Applicability of energy-absorbing support system was studied basing on a novel facility, and the support performance was compared with that of the ordinary rock bolt support under the same simulated geology conditions. The results demonstrate that the modelled roadway still kept in good condition after several successive impacts with increasing energy level, suggesting that the roadway retained its resistance to dynamic loads at that time. The bursting failure occurred suddenly when the dynamic loading energy reached a certain critical level. Significant differences of acceleration signals from two roadway sections with different support types appeared, indicating that the energy-absorbing support could greatly change the shock behaviour of the roadway under dynamic loading. A very significant decrease in acceleration amplitudes was produced by the good shock resistance of the roadway section supported with additional energy-absorbing elements to perfect absorbing shock energy of the roadway structure. During the bursting failure, roadway section with traditional rock bolt support was severely damaged and almost closed, while the front roadway section supported by additional energy-absorbing elements remained in relative good condition. The combination of energy-absorbing elements and rock bolt support provides a very efficient and ductile support type in extreme burst-prone conditions.

A new analytical model that represents the behavior of reinforced ground near a circular underground opening in a homogeneous, uniform stress field has been developed by considering the interaction behaviors between the grouted rock bolt and ground. Axial force distribution along the rock bolt has been analyzed at first. The coupling and decoupling behaviors of the interface between the ground and rock bolt are taken into account in the model. Thereafter, the mechanical behaviors of the reinforced mass are also evaluated according to the elasto-plasticity concept and the strain softening characteristic of ground. It highlights the influence of the different bolting patterns on the extent yield zone and tunnel deformation. The derivation of the analytical model and the influence of bolts on the stress and displacement field near an opening are illustrated in this paper. The theoretical predictions are verified with the measurement data in field, e.g. a Shinkansen tunnel named as Tawara Zaga Tunnel, which is planned to be open in 2022. Moreover, the applications of the proposed approach to tunnel design are discussed further according to the rock bolting effects of the standard supporting patterns of Japan Highway Public Corporation (JH) in different ground conditions.

An enriched finite element procedure is proposed for formulating interaction of rock bolts with multiple intersecting/ non-intersecting joints in a rock mass. The numerical method involves analysis of reinforced rock mass having arbitrary joint dip angle, bolt angle and joint-bolt-grout properties in extended finite element (XFEM) platform. Shifted Heaviside and junction enrichment is incorporated for the intersecting jointed element. The paper presents procedures and formulation of stiffness matrices of a bolt crossing multiple joints (BCMJ) which may or may not intersect each other. Hence an element may contain two intersecting joints and a bolt. In that case for a 2D element, each node is enhanced to eleven degrees of freedom, 2 for rock displacements, 2 each for displacements of rock joints, 2 for displacements of the junction of joint intersection and 3 for displacements and rotation of bolt. The procedure also allows traction to be applied to a side of an enhanced element and to be transformed as nodal forces at enhanced degrees of freedom. The paper considers elastoplastic behaviour of joint surface and bolt-grout interface using Mohr-Coulomb criterion . For yielding of bolt rod unified tensile and bending criterion is adopted. The proposed method is implemented using 3-noded triangular elements and can be upgraded for higher order elements. The efficacy of the proposed procedure is evaluated by analyzing behaviour of a bolt reinforced jointed sample under direct shear test . The method is also applied to analyze displacements and stresses around a circular tunnel made in rock mass having multiple intersecting joints at different orientations, reinforced by bolts along the periphery of the tunnel. The results in terms of displacements and stresses are compared with those of unbolted rock mass model. The proposed method shows that effect of an individual joint, bolt and their interaction can be analyzed to understand the stability of excavation under given loading conditions and can be applied in wide variety of applications.

In this paper, the impact dynamics of rock bolt with passive friction control is modeled. Numerical investigations are performed for the bolts with frictional energy absorber under various dynamic loads. The results are compared with those from fully grouted rigid bolts. The studies indicate that without the frictional energy absorber, the bolts are highly susceptible to failure under dynamic impact loads, while the bolts with the fictional energy absorber significantly reduce this risk. The studies also reveal that the bolt sliding may occur at the combined effect of vibration frequency and velocity magnitude. Sliding will only initiate above certain threshold value defined by vibration frequency and velocity magnitude. Furthermore, the simulation results show that under certain impact forces when rock mass subject to only elastic deformation, bolt sliding may occur causing the detachment of bolt head and rock face, which may incapacitate the rock bolt in supporting the rock excavation. The investigation signifies that further development and appropriate design of the rock bolt system with a frictional energy absorber are critical in maintaining stability under dynamic loading.

In hard rock masses, rock destruction is often manifested in the form of the local failure of the blocks formed due to fracturing. In this paper, we compare the application of block theory and three-dimensional (3D) block-cutting analysis to underground caverns. Using block theory, we can determine both the type of removable block and key block for different excavation surfaces. We can then identify ‘support-required’ blocks from the key blocks by drawing and analyzing the shapes of the maximum likely blocks (i.e. the largest blocks likely to be formed when the fractures are extended indefinitely in a definite-sized cavern). Subsequently, 3D block-cutting analysis is used to identify all the spatial blocks cut by all the finite-sized fractures in the rock mass region. Based on the geometrical information gathered about each spatial block, we carry out an analysis of the progressive failure of the block system. The simulation results allow us to explore the statistical laws governing the distribution of progressive failure blocks. This is a very important aspect which can be used to determine the spacing and length of the rock bolts required, or to check the design of the intended rock support regime. Overall, if we can understand the shortcomings of classic block theory (due to, for example, the assumptions made) and properly utilize the benefits gained from 3D block-cutting analysis, we can deal with more complicated issues. Thus, we hope to gain a more intensive and extensive understanding of block failure analysis.

Numerical modelling of fully grouted rock bolts loaded in tension is presented by implementing a non-linear bond–slip relationship of bolt–grout interface into a commercial finite difference rock mechanics code. The proposed model shows a close match with the experimental results and analytical predictions in terms of load–displacement relationship, rock bolt axial force distribution and interfacial shear stress distribution. A detailed debonding failure process is also presented, which includes five successive states: elastic state, elastic-softening state, elastic-softening-debonding state, softening-debonding state and debonding state.

Babolak water conveyance in north of Iran on Babolak river is under construction. Tunnel support consists of lattice girder, rock bolt, and shotcrete. According to application of classification systems RMR and Q, rock mass in the path of primitive part of the tunnel which is excavated in Marn is evaluated weak and very weak. In first part of the paper the Closure parameters of Panet equation associated with face advance and time effect of the tunnel in six different stations are determined using optimizing Levenberg–Marquardt and Trust-Region back analysis techniques on convergence instrumentation data. In second part of the paper, Burger׳s creep parameters are determined using instrumentation results in 0+603 station and Panet equation results in a back analysis performed utilizing FLAC3D. Predicted convergence results by the Panet equation are in good agreement with numerical analyses.

The interaction of the rebar bolt profile with the surrounding medium is investigated analytically. The grout/bolt interface failure is categorised into dilational slip and parallel shear failure based on experimental observations. For each failure mode, an analytical expression for the failure is derived after introducing Mohr–Coulomb׳s failure criterion. The theoretical predictions are compared to, and showed good agreement with, experimental data. After introducing the concept of the most vulnerable slip angle in dilational slip failure, the role of the rebar profile is identified quantitatively with respect to bond failure. It is found that the optimal profile geometry is related to the mechanical properties of the grout and the confining pressure when failure occurs. Therefore, bolt profile studies must take into account the mechanical properties of the grout and confine pressure. In addition, conclusions drawn from laboratory pull out tests should be evaluated properly considering both the testing and in-situ conditions before any application. A case study has been included to demonstrate the application of the developed theory. In a coal mine roadway primary reinforcement design, rock bolts have to be selected out of the four most popular types currently used in the Australian mining industry. The advantages and disadvantages of each bolt are discussed analytically for each case, providing a way for engineers to select the optimal bolt in their application.

The cross-hole seismic technique was used to identify the extent of the Excavation Damage Zone (EDZ) is in a test tunnel and in a tunnel under construction. For the cross-hole test, a new electro-mechanical impact-based transmitter was developed and tested. The transmitter was specially designed to generate appropriate impact energy and fit into a 45-mm rock bolt borehole. In the test tunnel, 79P-wave rays were measured. The interval between the boreholes was 1.5 m. From the measured P-wave velocity, 2D tomography was constructed and the average velocity gradient curves at the offsets of 0.21, 0.57, 0.93, and 1.29 m were drawn. Given that the wave velocity gradient curves show spatial variation due to the quality of the rock mass, upper and lower bounds that delimit the P-wave velocity were drawn. The extent of the EDZ was determined as the depth at which the average velocity is lower than the lower bound. By applying this approach, the depth of the EDZ at the test tunnel was determined to be about 1.1 m from the tunnel surface. The cross-hole seismic measurement was compared with the condition of the extracted rock cores. A gray-scale image of the rock core was compared to the P-wave velocity profile. According to the image, disintegrated rock cores were present up to a depth of 1.1 m. From the seismic data and a visual inspection, a similar result was found. The same approach was applied in a tunnel under construction. From the site, the observed depth of the EDZ was about 0.9 m. In addition, because of the developed transmitter can generate S-waves, P- and S-wave velocities were measured and compared. The results show that S-waves can be successfully applied to identify an EDZ. The S-wave profile shows less scatter compared to a P-wave measurement.

This paper presents a novel numerical approach to simulate the longwall mining in detail, aiming to investigate the stability of gates and the loading process to rock bolts. The study site is the Svea Nord coalmine in Svalbard. In the proposed approach, progressive cave-in and fracturing of the roof strata, consolidation of the cave-in materials and stress changes are simulated in detail. The results of the simulations are used to analyse the loading process of the rock bolts in the gates. The numerical model is calibrated with the measurement data. The simulations reveal 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 results in heavy loading to the rock bolts. Therefore, the locations of weakness zones should be taken into account for rock bolt design. The proposed numerical approach can be used to find an optimal combination of the dimension of the chain pillars and the rock reinforcement in the gates.

A doubly enriched finite element (DEFE) procedure has been developed by introducing strong displacement discontinuities in the form of rock joints and reinforcement affect in the form of rock bolts. The proposed procedure extends the concept of extended finite element method (XFEM) and enriches the nodes of an element for determining displacements and stresses of bolt rod as well as displacement jumps and tractions of rock joint. The paper elaborates on the theoretical development of this procedure and derives stiffness matrix of a DEFE using variational approach. The stick-slip behaviour of joint, yielding of rock mass as well as grout material have also been implemented in the DEFE procedure. The paper also provides verifications of this procedure by solving two known examples, (i) direct shear test performed on a bolted joint sample-experimental verification, and (ii) reinforcement of a joint located in the vicinity of a circular tunnel-analytical verification.

1. IntroductionThe purpose of rock support and reinforcement is toensure excavations remain safe and open for their intendedlife span. The effectiveness of a reinforcement strategy isimportant for two main reasons: these being safety topersonnel and equipment, and achievement of the mosteconomical access to extract ore. For a particular rockmass, a stabilization scheme capable of matching theexpected behaviour is selected based on an assessment ofthe likely failure modes predicted from the interaction ofan excavation (geometry and purposes), the network ofgeological discontinuities, weathering and the loadingconditions from stress and blast damage [1–4].In most underground mines, the primary form ofexcavation stabilization is provided by a pattern of rockbolts installed within the rock mass. This is complementedwith the use of passive support, such as that provided bymesh or shotcrete, in order to provide surface restraint atthe exposed excavation boundaries.The reinforcement controls the overall excavationstability through keying, arching or composite beamreinforcing actions [5], while the mesh and/or shotcretesupports the small loose pieces of rock that may detachbetween the rock bolt plates [6,7].In general, a stabilization scheme cannot be selectedwithout consideration of the ground support drillingequipment available at a particular mine site. A modern,optimal strategy would consist of mechanized installationof reinforcement and support in a single pass in order toincrease productivity and reduce exposure of personnel andequipment during installation.2. Mechanized resin anchored bolt installationOver the last decade or so, jumbo-installed, 45mmdiameter galvanized and black steel friction bolt stabilizershave become the preferred form of reinforcement inunderground hard rock mining in Australia [8,9]. Thishas been mainly driven by a desire to achieve fastdevelopment rates and low costs in order to allow theextraction of low-grade orebodies. In more recent years, asthe mining operations are getting deeper and the rockmasses are becoming highly stressed, other reinforcementschemes such as fully encapsulated resin anchored bars arebeing considered as an alternative to friction bolts for long-term reinforcement [10–13].The typical bolts being used in the underground hardrock mines have been modified from the bolts used in thecoal mining industry. The modifications have beennecessary due to the need to drill larger hole diameterswith the type of equipment used in the hard rockmetaliferous mines. The modification is mainly in the formof paddles or the use of a spring welded on to the endsection of the bolts. Fig. 1 shows the anchor sections for a24mm Posimix bolt with spring arrangement and a 27mmSecura bolt showing a paddle arrangement. The Posimixwire is 3mm in diameter and has a length of 500mm. Thepaddle width is 29.2mm and they have been sheared intothe end of the bolt for the purpose of mixing resin.Nevertheless, the introduction of mechanized resinanchored bolting using jumbos has been difficult toimplement economically due to the high cost of resintransport and storage: this requires the use of refrigerated

Although the design of underground structures is becoming increasingly dependent on modeling-based analysis approaches, ground control systems continue to be designed primarily using a site-specific trial and error approach. This discrepancy is related to the difficulty of reproducing rock-support interaction behavior in its entirety using commonly applied modeling methods. To help bridge this gap, this study has conducted a comparison of continuum and discontinuum models with a focus on their ability to realistically capture the support effect generated by rock bolts with or without wire mesh. It was found that continuum models can adequately replicate the damage processes and deformation of reinforced rockmasses, but show a negligible change in ground behavior when support elements are explicitly considered within the models. We propose a ‘strain-continuity’ hypothesis where the inability of the modeling method to allow discrete block separation after failure is thought to be responsible for this unrealistic behavior. Two discontinuum bonded-block modeling approaches (Voronoi Tessellation and Trigon) were subsequently tested; realistic behavioral differences between the supported and unsupported model conditions that are consistent with empirical data were exhibited by the Voronoi model. Finally, a conceptual framework has been presented that combines field data from the literature with the modeling results to provide an understanding of how support responses in the continuum and discontinuum models compare to reality.

Rock bolts have been widely used to reinforce the jointed rock mass. Modelling the mechanical shear behaviors of the bolted rock joints are difficult due to the complex interactions between bolts and rock joints. The applied pretension forces combined with the axial loads developed in the bolt act as the normal forces which are applied to the rock joints. However, these combined normal forces are not considered in the existing analytical models. An analytical model is proposed in this study to predict the shear behavior of the bolted rock joints, by taking into account the pretension forces, the axial forces developed in the bolt, the interfacial bond stress between the bolt and grout, and dowel shear loads acting transversely to the bolt axis. The proposed analytical model is able to provide complete curves of the dowel shear loads, axial loads, and the global shear loads as a function of the joint shear displacement. The analytically predicted axial load vs shear displacement curves and the global shear load vs shear displacement curves are verified by available experimental tests. The validation shows that the proposed model has the capacity to predict the global shear load evolution as well as the axial load evolution. The factors such as the pretension forces, the bolt inclination angles, the concrete strength and the rock joint friction are successfully accounted for in the analytical model.

Short encapsulated pullout tests are frequently used in the laboratory and in the field to examine the bond performance of rock bolting. Because of the restriction of sample conditions and of testing apparatus, push testing is usually used instead in the laboratory. Rock bolts have a better performance in push testing and this is a thought to be attributed to Poisson's ratio effect of the steel rebar. In this paper Poisson's ratio effect in short encapsulated push/pull tests is evaluated analytically. The percentage difference of axial load in push and pull tests is calculated based on the failure mode analysis and the corresponding mechanical models of rock bolting mechanisms. Analytical results show that Poisson's ratio effect is closely related to the radial stiffness and embedded length. The theoretical predictions were compared with experimental data and showed good agreement, and so can be used as a guideline in laboratory testing design.

This paper explores the use of dynamic testing of rock bolts, currently used for collar load estimation as a means for indirect measurement of the effective installed bolt length, in order to determine potentially broken bolts. Both experimental and numerical studies were undertaken. A number of rock bolts of different total length were installed in a concrete block. These bolts were assessed using non-destructive testing techniques commonly used to assess prestress load. Initially a system based on the application of an impulse load (GRANIT) was used, followed by a periodic (swept sine) loading using a magneto-strictive shaker. A dynamic model, developed previously, was used to assist in interpretation of the results obtained from the experiments. Analysis of the data obtained with the two testing methods is undertaken and the results compared with those from the dynamic model. The results suggest that both methods give similar responses but also that both excitation methods need to be altered to allow effective fixed/total lengths to be estimated. The numerical model used in this study confirms the findings from the experiments and gives a frequency range which, if excited can provide data which can be related to effective total bolt length. This length could be potentially different from the original design due to either incorrect installation or bolt breakage.

The Ghatghar pumped storage hydro-electric project is under construction in the Thane district of Maharastra state in India for the generation of 250MW of electricity. The scheme consists of two underground caverns, a machine hall, a transformer hall and intersections. The interconnections between the caverns and connections to the surface, as well as to the reservoir, are made through a network of tunnels. The rock mass is predominantly of basaltic formation. The joints are mostly tight and undulating with generally rough to very rough contact surfaces. The rock mass characterization was conducted using Bieniawski’s RMR (1979) and Barton’s Q. The associated values vary from 66 to 81 and from 5.9 to 25, respectively—which indicate very good rock mass conditions. The support requirements for these structures were estimated from empirical approaches and numerical modeling. The support system consists of 6m long fully grouted rock bolts at 2m grid spacing with 50mm thick SFRS for the two caverns and 7m long rock bolts around the intersections of all openings. An extensive instrumentation scheme was implemented in these structures as a means to evaluate the efficacy of the support system and to have detailed knowledge of the rock mass behavior during and after construction. The instrumentation consisted of: tape extensometers for measuring convergence; multi-point and single point borehole extensometers for radial displacements; load cells for rock load and strain meters for rock strains. These instruments were embedded during the process of cavern excavation. The rock mass deformations around the cavern openings, as indicated by the MPBXs (Fig. 1), the convergence measured by the tape extensometer, and the rock bolt tension measured by the load cells were too low to be significant and collectively indicate that the roof and the walls of both the caverns and of the bus ducts and passageways are stable, indicating that the support system is adequate. It was, however, suggested that the vulnerable wedges should be identified on account of the jointing; these would need to be stitched to the rock mass and the wedge stability be monitored by site specific arrangements. This paper presents in detail the instrumentation scheme implemented at this project plus the analysis of the monitored data, highlights the stability status of the caverns, and recommends the future strategy for monitoring.

The research outlined in this paper is aimed at the development of a portable non-destructive testing instrument for evaluating the condition of rock bolts. In applications such as coal mine roof reinforcement, the opportunities for rock bolt inspection are currently limited to destructive techniques such as the pull-out test. It is proposed that guided ultrasonic waves can be used to solve this inspection problem, using a pulse-echo test carried out from the free end of the bolt. Suitable test frequencies have been identified through the use of modelling software, and successful laboratory and site trials have been undertaken. The research has shown that the proposed approach is capable of determining the bolt length, and of identifying major defects such as necking, deformation, and loss of resin encapsulation.

1. IntroductionRock bolts have been extensively used for rock reinforcement inunderground coal mine roadways around the world, especially inthe United States and Australia, where rock bolts were initiallyused in coal mines [1,2]. In Chinese coal mines, rock bolts wereinitially used to reinforce roadways driven in rock layers ratherthan coal seams. Early types of rock bolts include mechanically-anchored bolts, grouted dowels, point-anchored bolts with resinor cement capsules and split-set bolts which have low strengthand can only be used for simple geological conditions. During1996–1997, a so-called “high-strength rock bolting technology”was introduced from Australia to Chinese coal mines. This tech-nology employs high-strength resin-anchored rock bolts andpretension cables installed in small diameter (28–32 mm) bore-holes. These have been used in many coal districts and thereinforcement effect is generally significant. With continuousimprovements and modifications, rock bolting has become themain reinforcement pattern in Chinese coal mines with around70% of roadways being reinforced by rock bolts.In the last decade, mining depth has gradually increased inChinese coal mines. At depth, geological conditions become com-plex and both the in-situ stresses and mining-induced stresses arehigh. Certain issues have occurred when using conventional highstrength rock bolts to reinforce deep roadways suffering from highstresses. The deformation of these roadways could not be effec-tively controlled. Severe failure and damage to rock bolts, bearingplates, straps, and meshes were commonly observed. To solvethese issues, a so-called “high-pretension, extreme-strength rockbolting technology” was developed [3–5]. This technologyintroduced two major improvements compared to the previoushigh strength rock bolting technology. Firstly the strength of therock bolts has been considerably increased (from 400 to 600 MPaof yield strength and from 600 to 800 MPa of ultimate strength).The diameter of cables has been increased from 15.2 to 22 mm,resulting in a significant increase (from 260 to 560 kN) in theultimate tensile capacity. The other major improvement is that theimportance of high pretension is specially highlighted [6,7]. Boththe rock bolts and cables are designed to be installed with highpretension. This high-pretension, extreme-strength rock boltingtechnology has been implemented in deep roadways in theXinwen, Huainan, Lu'an, and Jincheng coal districts and successfulsuppression of roadway deformation has been achieved [8].But a new issue has occurred on the high-pretension, extreme-strength rock bolts. When employing them to reinforce roadwaysunder severe stress conditions, fractured bolts are frequentlyobserved in the roadways. Bolts that fail in the thread section falldown on the roadway floor with plate and domed washer. Boltsfailed in the middle section usually exhibit loose platens and canbe easily pulled out by hand. The zone anchored by the fracturedbolts is a local weakness point from which large-scale failure mayinitiates, resulting in potential safety issues. Understanding of thepatterns and mechanisms of fractured bolt is of great importance,and has been studied by many researchers. Ferrero [9] carried outexperimental measurements and numerical modeling to studyrock bolt failure and recognized two failure mechanism types.When a bolt is installed in a strong and stiff rock material, failuremay occur at the joint intersection due to shear and tensilestresses. When a bolt is installed in a weaker rock, two symmetricplastic hinges might occur in the bar, leading to a failure if eitherthe tensile strength or the ultimate elongation of the reinforce-ment is reached. Li [10] stated that bolt fracture results from thecombined tension and shear loads based on the field monitoringdata of bolt fracture states in a highly stressed rock mass. He statedContents lists available at SciVerse ScienceDirectjournal homepage:www.elsevier.com/locate/ijrmms

Fiber-optically instrumented rock strain and temperature strips (FROSTS) are a new method for monitoring in situ strain and temperature in intact rock masses. FROSTS are an alternative to other techniques such as borehole extensometers or instrumented rock bolts for monitoring strain around underground openings. FROSTS have six fiber-optic strain and temperature gages installed at 30 cm intervals along specially designed strips of 304-stainless steel and are embedded into the rock mass. FROSTS are modified Rock Strain Strips (ROSS). If pretensioned prior to installation, the FROSTS can measure both shortening and elongation. Two FROSTS were pretensioned and installed on the 4100 level (~1250 m below the surface) of the Sanford Underground Research Facility (SURF) in Lead, SD. Data from the initial monitoring of these FROSTS record strain in the rock mass due to the closure of microcracks and fractures created during drilling for installation and then stabilization of the FROSTS sensors. The fiber-optic temperature sensors on the FROSTS also record a detailed thermal gradient of increasing temperature with depth into the rock mass. Laboratory experiments show that the FROSTS accurately measure strain, are more compliant than the rock mass they are embedded into, and thoroughly couple to the rock mass to ensure accurate readings. FROSTS provide an accurate and detailed record of in situ strain and temperature in rock masses and are useful for the structural health monitoring of underground spaces.

A numerical modelling strategy has been developed in order to quantify the magnitude of induced stresses at the boundaries of production level and undercut level drifts for various in situ stress environments and undercut scenarios. The results of the stress modelling were in line with qualitative experiential guidelines and a limited number of induced stress measurements documented from caving sites. A number of stress charts were developed which quantify the maximum boundary stresses in drift roofs for varying in situ stress regimes, depths and undercut scenarios. This enabled many of the experiential guidelines to be quantified and bounded. A limited number of case histories of support and support performance in cave mine drifts were compared to support recommendations using the NGI classification system, The stress charts were used to estimate the Stress Reduction Factor for this system. The back-analyses suggested that the NGI classification system might be able to give preliminary estimates of support requirements in caving mines with modifications relating to rock bolt length and the support of production level intersections. (C) 2002 Elsevier Science Ltd. All rights reserved.

The in-situ rock structural weaknesses, referred to herein as geo-structural defects, such as naturally induced micro-cracks, are extremely responsive to tensile stresses. Flexural toppling failure occurs by tensile stress caused by the moment due to the weight of the inclined superimposed cantilever-like rock columns. Hence, geo-structural defects that may naturally exist in rock columns are modeled by a series of cracks in maximum tensile stress plane. The magnitude and location of the maximum tensile stress in rock columns with potential flexural toppling failure are determined. Then, the minimum factor of safety for rock columns are computed by means of principles of solid and fracture mechanics, independently. Next, a new equation is proposed to determine the length of critical crack in such rock columns. It has been shown that if the length of natural crack is smaller than the length of critical crack, then the result based on solid mechanics approach is more appropriate; otherwise, the result obtained based on the principles of fracture mechanics is more acceptable. Subsequently, for stabilization of the prescribed rock slopes, some new analytical relationships are suggested for determination the length and diameter of the required fully grouted rock bolts. Finally, for quick design of rock slopes against flexural toppling failure, a graphical approach along with some design curves are presented by which an admissible inclination of such rock slopes and or length of all required fully grouted rock bolts are determined. In addition, a case study has been used for practical verification of the proposed approaches.

An energy-absorbing rock support device, called a D bolt, has been recently developed to counteract both burst-prone and squeezing rock conditions that occur during underground excavation. The bolt is a smooth steel bar with a number of anchors along its length. The anchors are firmly fixed within a borehole using either cement grout or resin, while the smooth sections of the bolt between the anchors may freely deform in response to rock dilation. Failure of one section does not affect the reinforcement performance of the other sections. The bolt is designed to fully use both the strength and the deformation capacity of the bolt material along the entire length. The bolt has large load-bearing and deformation capacities. Static pull tests and dynamic drop tests show that the bolt length elongates by 14–20% at a load level equal to the strength of the bolt material, thereby absorbing a large amount of energy. The impact average load of a 20 mm D bolt is 200–230 kN, with only a small portion of the load transferred to the bolt plate. The cumulative dynamic energy absorption of the bolt is measured to be 47 kJ/m. D bolts were tested in three deep mines. Filed measurements show that D bolts are loaded less than rebar bolts. This paper presents the layout and principle of the D bolt, and corresponding results from static, dynamic, and field tests.

To ensure safety in underground excavations, it is important that the support systems used are capable of resisting the dynamic loads produced, for example, by rock bursts. In this paper, a dynamic simulation model for a cone bolt is proposed based on an experimental study. Drop weight tests were performed on resin-based cone bolts. These experiments revealed that the bolt has two energy absorption mechanisms: sliding in the resin and plastic deformation. To simulate this behaviour, a two degrees-of-freedom lumped-mass model is proposed. Experimentally, the proportions of sliding and plastic deformation were found to vary significantly from one test to another. To account for this variability, two methods are proposed to determine the value of the parameters governing the sliding of the bolt in the resin, whereas a dynamic force–elongation model is used to simulate the plastic deformation. Comparing the results of a simulation to experimental data proved that the constitutive elements of the model are appropriate to simulate the dynamic response of the cone bolt.

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.

Experiments were conducted to study the behavior of guided waves in free and grouted rock bolts. Ultrasonic waves with frequencies from 25 to 100 kHz were used as excitation inputs. Tests were first conducted on free bolts to help understand the behavior of guided waves in non-grouted bolts. The effects of wave frequency and grouted length on the group velocity and attenuation of the guided ultrasonic waves were then evaluated. The test results indicated clear but different trends for the group velocity in the free and the grouted bolts. The attenuation in free bolts was not affected by bolt length and frequency. However, in grouted bolts it increased with frequency and grouted length. It was also found that the two main sources of attenuation are the setup energy loss, which has a fixed quantity for a specific type of test setup, and the dispersive and spreading energy loss which varies with frequency and bolt length.

In order to characterize the behavior of a main gate during retreating of longwall faces and to evaluate the performances of five different support types, including rock bolts and steel sets, a total of 21 convergence and 18 floor-heave stations were installed. Consequently, about 11,500 convergence and 7200 floor-heave data points were obtained over a period of 14 months. In order to simulate the behavior of a supported main gate roadway, a convergence model, represented by an exponential function, was developed on the basis of statistical analyses. In addition to the convergence model, convergence–velocity and convergence–acceleration models were also derived and interpreted in terms of gate road behavior. It was determined that the convergence depends on the loading region, the time, the location of the upper and lower longwall faces, the support type and pattern, and the static and dynamic support loads. Convergence–velocity and acceleration curves provide better means to interpret the performance of supports and gate roadway behavior.

Three analytical models have been developed for rock bolts: one for bolts subjected to a concentrated pull load in pullout tests, one for bolts installed in uniformly deformed rock masses, and one for bolts subjected to the opening of individual rock joints. The development of the models has been based on the description of the mechanical coupling at the interface between the bolt and the grout medium for grouted bolts, or between the bolt and the rock for frictionally coupled bolts. For rock bolts in pullout tests, the shear stress of the interface attenuates exponentially with increasing distance from the point of loading when the deformation is compatible across the interface. Decoupling may start first at the loading point when the applied load is large enough and then propagate towards the far end of the bolt with a further increase in the applied load. The magnitude of the shear stress on the decoupled bolt section depends on the coupling mechanism at the interface. For fully grouted bolts, the shear stress on the decoupled section is lower than the peak shear strength of the interface, while for fully frictionally coupled bolts it is approximately the same as the peak shear strength. For rock bolts installed in uniformly deformed rock, the loading process of the bolts due to rock deformation has been taken into account in developing the model. Model simulations confirm the previous findings that a bolt in situ has a pick-up length, an anchor length and a neutral point. It is also revealed that the face plate plays a significant role in enhancing the reinforcement effect. In jointed rock masses, several axial stress peaks may occur along the bolt because of the opening of rock joints intersecting the bolt.

Fiber reinforced shotcrete has been widely used as part of permanent tunnel support during the last 15 years especially in connection with the application of the Norwegian Method of Tunnelling (NMT). The interaction of the fiber reinforced shotcrete and the rock bolt reinforcement can now be numerically modelled with the Distinct element method (DEM). The discontinuous code UDEC (Universal Distinct Element Code) is used to investigate the overall stability of an excavation, to predict the expected stresses and deformations caused by the excavation and to investigate the optimal excavation sequence to be followed. The jointed rock geometry of a triple lane traffic tunnel in jointed volcanic tuff has been considered for demonstating the fiber reinforced shotcrete subroutine. The results have shown that by using S(fr) and subsequently rock bolts as primary support in the tunnel, the load attained by some of the rock bolts is reduced by approximately half compared to the case were only rock bolts were used.

The Yucca Mountain Project is conducting a Single Heater Test (SHT) in the Exploratory Studies Facility at Yucca Mountain, NV to evaluate coupled thermal-mechanical-hydrologic-chemical processes in situ . This paper describes the thermomechanical aspects of the SHT and presents preliminary results from the test. The 5-m long heater deployed in the SHT was energized on 26 August 1996 with the heater power set at 4 kW. Thermomechanical instrumentation was installed within and on the rock mass surrounding the SHT. The thermomechanical instrumentation includes temperature measurements using thermocouples, RTDs, and thermistors; displacement measurements using multiple-point borehole extensometers (MPBXs), tape extensometers, and surface-mounted wire extensometers; load measurements from load cells on rock bolts installed both within thermally perturbed and ambient regions; and rock mass modulus measurements using the NX borehole jack. Pretest analyses of the thermomechanical response were conducted using the thermohydrologic code TOUGH2 and the thermomechanical structural code JAC3D. The test design, preliminary data, and comparisons between pretest predictions and early data from selected thermal and mechanical sensors are presented.

The holding mechanism of under-reamed rockbolts differs from that of conventional rockbolts, in which the bonding or the friction along the element–rock interface provides the holding capacity. The under-reamed end is kinematically blocked by the surrounding rock mass and can provide a greater holding capacity, especially in soft rock, whereas the strength of a soft rock frequently controls the bonding strength of the element–rock interface. Both an experimental study and numerical analyses were performed to examine the holding mechanisms of model under-reamed rockbolts subjected to direct pull out loading and pre-tensioning. When subjected to direct pull out loading, the holding capacity originates from the capacity of the rock resisting tensile fracture. Failure is characterized by the formation of a smooth, conical region bounded by a tensile crack, which subsequently separates from the surrounding rock. Correspondingly, the holding capacity is related to the tensile strength of the rock, bolt length and size of the under-reamed end. When subjected to pre-tensioning, the holding mechanism is provided by the ability of the rock to form two conical zones between the faceplate and the under-reamed end, and to prevent subsequent indentation of the two cones. Major factors influencing the holding capacity of under-reamed bolts include the size of the under-reamed end, bolt length and properties of the rock.