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Rock bolts are commonly used to stabilise slopes and underground excavations but their installation can sometimes be difficult in challenging grounds due to excessive leakage or washout of the grout. To manage these problems, tubular geotextile socks have been adopted in the industry but there has been a lack of research on their properties. To advance the understanding of sock behaviour, the water permittivity and filtration properties of two types of grout sock have been investigated in their unstretched state. From the permittivity tests, it is shown that water flow through these socks will be laminar and that using more than one layer in a setup can effectively reduce the flow velocity. Moreover, it has been shown that the effect of the tester should be considered in the analysis or else the permittivity of the geotextile could be underestimated. From the filtration tests, it is shown that both types of grout sock tested are effective in minimising grout loss and that the permittivity of the filter cake will be many orders of magnitude lower than that of the socks. The results also show that the water–cement ratio of the grout will be reduced as a result of the dewatering process.

The effects of reinforcement system on pillars were tested in laboratory, using three types of pillars with different strengths. The tests were performed using the UCS machine, to test pillar without reinforcement, pillar with rock bolt reinforcement, pillar with shotcrete reinforcement and pillar with the combination of both rock bolt and shotcrete reinforcement. Uniaxial compressive strength (UCS) testing aims to determine the effects of the reinforcement system on pillar strength. The results of this study indicate that the reinforcement system on high strength pillars causes a strength increase of 14.93% on pillar with rock bolt reinforcement, 21.45% on pillar with shotcrete reinforcement and 34.67% on pillar with combination of rock bolt and shotcrete reinforcement. On medium strength pillars, reinforcement installation shows a strength increase of 16.27% on pillar with reinforced rock bolt, 19.83% on pillar with reinforced shotcrete and 44.40% on pillar with combination of rock bolt and shotcrete reinforcement. Likewise, on low strength pillars, reinforcement installation causes a strength increase of 13.13% on pillar with reinforced rock bolt, 36.21% on pillar with reinforced shotcrete and 53.85% on pillar with combination of rock bolt and shotcrete reinforcement. The results of laboratory testing and numerical modeling indicate that the increase in strength occurs because the horizontal displacement on the surface of the pillar wall is detained by shotcrete and faceplate on rock bolt, so that the pillar seems to have confining pressure throughout the pillar wall surface, which is called as equivalent confining pressure.

In coal mining roadway support design, the working resistance of the rock bolt is the key factor affecting its maximum support load. Effective improvement of the working resistance is of great significance to roadway support. Based on the rock bolt’s tensile characteristics and the mining roadway surrounding rock deformation, a mechanical model for calculating the working resistance of the rock bolt was established and solved. Taking the mining roadway of the 17102 (3) working face at the Panji No. 3 Coal Mine of China as a research site, with a quadrilateral section roadway, the influence of pretension and anchorage length on the working resistance of high-strength and ordinary rock bolts in the middle and corner of the roadway is studied. The results show that when the bolt is in the elastic stage, increasing the pretension and anchorage length can effectively improve the working resistance. When the bolt is in the yield and strain-strengthening stages, increasing the pretension and anchorage length cannot effectively improve the working resistance. The influence of pretension and anchorage length on the ordinary and high-strength bolts is similar. The ordinary bolt’s working resistance is approximately 25 kN less than that of the high-strength bolt. When pretension and anchorage length are considered separately, the best pretensions of the high-strength bolt in the middle of the roadway side and the roadway corner are 41.55 and 104.26 kN, respectively, and the best anchorage lengths are 1.54 and 2.12 m, respectively. The best anchorage length of the ordinary bolt is the same as that of the high-strength bolt, and the best pretension for the ordinary bolt in the middle of the roadway side and at the roadway corner is 33.51 and 85.12 kN, respectively. The research results can provide a theoretical basis for supporting the design of quadrilateral mining roadways.

In order to study the mechanical mechanism of the new tension–compression composite anchor, the finite element software is used to establish the numerical calculation model of the new composite anchor, and the numerical drawing test is carried out. Through numerical experiments, the optimal position range of the bolt bearing plate is obtained, and the axial load and shear stress distribution of the bolt were analyzed. The results show that the ultimate pullout force of the new tension–compression composite bolt increases first and then decreases with the movement of the bearing plate position, and the optimal position is at the position where the length ratio of the tension–compression anchorage section is 3:4. At the same time, the study shows that the optimal position of the bearing plate has an optimal position interval. The ultimate uplift capacity of the new tension–compression composite anchor is 118.6% and 94% higher than that of the traditional tension anchor and pressure anchor, respectively. The shear stress distribution of the new tension–compression composite anchor is more uniform. The peak shear stress first appears at the bearing plate and begins to transfer to both sides of the bearing plate with the increase in the drawing force, which increases first and then decreases. Since the new tension–compression composite anchorage section is in an unbonded state, the axial load of the bolt is always equal to the drawing force. The axial force distribution of the tensile anchorage section is similar to that of the traditional tensile bolt and decreases exponentially along the axial direction of the bolt. Under the action of ultimate drawing force, the plastic zone distribution of the new tension–compression composite anchor is different from that of the traditional tension anchor and the traditional pressure anchor. The plastic zone first appears in the position of the bearing plate and begins to transfer to both sides of the bearing plate with the increase in the drawing force. When the plastic strain exceeds the allowable strain of the interface, the bolt pulls out from the anchor hole and the bolt support structure fails.

Rock bolts have been widely used to enhance the structural stability of underground infrastructures. Careful tracking of rock bolt positions is highly significant since it assists with operational success of ground support and has applications to predictive maintenance practices. This paper presents a practical algorithm, CFBolt, to detect rock bolts from a 3D laser scanned point cloud. Considering that rock bolts are relatively tiny objects, CFBolt follows a two-step coarse-to-fine strategy. It first computes a single-scale proportion of variance (POV) for each point as the local point descriptor and filters out near 95% not-bolt points with a simple but effective classifier, Linear Discriminant Analysis (LDA), which allows for the pruned point cloud to be then used as a compatible input to a deep neural network, designed and trained to precisely detect rock bolts from the pruned point cloud. CFBolt was tested for detecting rock bolts from LiDAR scan data collected from Sydney’s civil tunnelling project site. The entire dataset contains more than 160 million points. The obtained scores of Intersection over Union (IoU) and precision for individual bolt points were 89.33% and 92.04%, respectively. For rock bolt objects, the precision and recall were 98.34% and 98.73%, respectively. The detection quality of CFBolt is superior to the state-of-the-art 3D object detection algorithms and the newest rock bolt detection algorithm, demonstrating the robustness and effectiveness of CFBolt.

Rock bolt anchorage performance is crucial for tunnel support safety. We investigate the mechanical behaviour of reinforced rock and the bolts that reinforce it from the micro-scale to the macro-scale. Bolt pull-out tests were performed on soft rock using the distinct element method, in which a new contact model that considers bond size, is employed to constrain the main rock mechanical behaviour. The minimum sample width and height values for which the boundary effect can be neglected are first proposed through numerous tests on the influence of sample size on peak load and bond breakage. The influence of sample width is substantially greater than that of sample height. We then select an appropriate sample size to study the influence of bolt embedment length and confining pressure on the mechanical behaviours of the rock and bolt. The results show that increased rock bolt embedment length and confining pressure can increase the peak load; however, the bolt length effect is limited when exceeding the critical anchorage length. In cases without confining pressure, bond breakage occurs in the rock around the grout-rock interface and the breakage zone is rectangular, whereas in cases under confining pressure, the breakage zone presents an inverted cone shape. We use our results to discuss the influence of bond strength at the bolt–grout interface on the peak load and failure mode. The failure mode changes gradually from complex failure to single failure along the bolt–grout interface with decreasing interfacial bond strength.

Determining the rock bolt response under the typical in situ loading conditions of tension, shear and bending is an important part of the support design process. Calibrated numerical modelling can be an important tool for simulating the rock bolt response under in situ loads. However, currently available structural support models do not accurately model the rock bolt behaviour under shear loads. These models need to be modified to simulate the actual rock bolt response as determined under laboratory testing. Results from double shear laboratory tests by Kostecki (2019) have been used to calibrate the numerical models in ANSYS (ANSYS® Academic Research Mechanical 2019). The calibrated numerical models were used to study the state of stress along the rock bolt under different concrete strength. These numerical models are compared with the analytical model proposed by Pellet and Egger (1996). The results show that the rock reaction and consequently the shear load on rock bolt in the analytical model is underestimated (by up to 50% in case of high strength concrete). In high strength concrete, the yield and failure behaviour is determined by the shear stress rather than the bending and tensile stress as predicted by the analytical models. The analytical model is modified to better predict the shear forces in high strength concrete. The pile structural element available in FLAC3D (FLAC3D — Fast Lagrangian Analysis of Continua in Three-Dimensions 2017) finite difference method neglects the transverse shear strain in the element. In this paper, the pile structural element is modified to simulate the correct response of rock bolt under shear. The FLAC3D code is also modified to simulate the bolt yielding and failure under combined axial, bending and shear load. The modified model is validated with laboratory results from Mchugh and Signer (McHugh and Signer 1999) and is shown to be better at simulating the shear response and failure of rock bolt under shear compared to the original pile model.

Rock bolts are components of tunnel support structures and their efficacy in reinforcement requires the appropriate estimation of the bolt pre-tension. However, cases of ineffective rock bolt application due to loosening are common which leads to the potential failure of the support structures, thus negatively affecting the tunnel stability. The effective employment of pre-tensioned bolt within support structures requires an appropriate installation tension as a result of sufficient bolt head/nut pre-load. Henceforth, this paper presents a load indicator washer (LIW) capable of withstanding assigned static loading as well as provides an appropriate pre-tension. The different LIW breadths are numerically analysed in the finite element code (ABAQUS) to estimate deformation when a designated 60-kN pre-load is applied. The numerical simulation employs a bi-linear isotropic hardening constitutive law governed by the von Mises yield criterion. The simulation also considers the nominal temperature of the nut-LIW-rock bolt connection. Therefore, a long-term relaxation effect on bolt pre-tension as a result of stress relaxation is neglected. The compressive force required to achieve the LIW-predetermined deformation is the bolt pre-tension. It is drawn from the simulation results that the 1.5-, 2.0-, 3.0-, 3.5-, and 4.0-mm-thick LIWs are completely flattened when subjected to 28-, 42.5-, 65-, 76-, and 88-kN pre-load respectively.

This paper presents results of 71 in situ nondestructive rock bolt and tendon pull tests conducted in the Powerhouse Complex of the Uma Oya Multipurpose Development Project, located in Sri Lanka. The pull tests are performed on 6 and 8 m Double Corrosion Protected fully and partially grouted rock bolts (high strength threadbar) as well as 4 m normal fully grouted rock bolts and 26 m pre-stressed tendons, which are installed in different types of Gneiss. Based on the recorded load–displacement curve of the rock bolt head in each test, stiffness of fully and partially grouted rock bolt is measured in the bonded section. Effect of some parameters on the stiffness such as rock mass types and characteristics, bond length as well as rock bolt diameter and steel type are investigated. Primarily, the pull tests are done to approve the working load capacity of rock bolts and tendons and quality control of the installation process. Besides, the stiffness of bonded rock bolt is also calculated. The results show that the stiffness declines by decreasing the rock mass mechanical characteristics, rock bolt bond length, and rock bolt diameter as well as steel bar grade.

The research has proved that rock mass stability can be ensured using steel resin rock bolting including rods, fast-setting resin capsules, bearing plates, strapping and tension nuts. Interaction of these structural components with roof rocks and sidewalls involves obligatory displacement of rocks. The article presents the studies into applicability of steel resin rock bolting in difficult geological conditions in coal mines in Kuzbass. The steel resin rock bolting is considered in this case as the method of ‘active’ ground control. The ‘active’ ground control consists in the capacity of steel resin rock bolts to alter main strength characteristics of rocks, to increase their resistance and stability, and, thereby, to ensure efficiency of roof support. It is found that horizontal and vertical stresses in underground excavations depend on the mining depth. It is proved that rock bolts installed in the roof rocks can prevent or confine rock fracture.

Corrosion of rock bolts is a relatively slow process yet a potential problem for the underground mining industry. Mine groundwater in the strata through natural or mining-induced cracks together wi...

The corrosion of rock bolts buried in rock and soil layer is obviously different from that of general ground structure. The main corrosion sources of the latter come from the air and the corrosion sources of rock bolts are not only from the air, but also the rock-soil body and groundwater. At the same time, the corrosion of rock bolts is under the stress state in practical engineering. To study the stress-corrosion effect of grouted rock bolts, using a self-developed rock bolt stress corrosion device, corrosion tests were conducted under different working stresses (10 kN, 35 kN and 60 kN), corrosion environments (5% Na2SO4, 1.5% NaCl + 2.5% Na2SO4 and 3% NaCl) and corrosion times (3 d, 7 d and 14 d). The results show that the NaCl solution is more corrosive than the Na2SO4 solution for the grouted rock bolts with more rust productions. The maximum tension of rock bolts under the condition of no corrosion reaches 60.16 kN, and under the T-60-A-14d working condition, the maximum tension of rock bolts is only 31.32 kN. Compared with the condition of no corrosion, the maximum tension of grouted bolts is lowered by 25%–50% under the different working stresses and corrosion conditions. The maximum difference values of the relative bonding strength of rock bolts is 14.9% under corrosion time, 13.1% under working stress and 9.4% under corrosion environment. It shows that the corrosion effect of the rock bolt under the corrosion time is the most significant, and the working stress is second. The influence of the corrosion environment on the bonding strength of the grouted rock bolts is the least important. The slip curves of the rock bolts under the environment of no corrosion have an obvious rising stage, and uplift of tension accounts for about 62.9% of the maximum tension. The final residual tension is stable at approximately 25 kN. The bonding slip curves of corroded rock bolts, of which the “initial tension rising stage” almost disappears, demonstrate a trend of steady decline. The minimum final residual tension is less than 10 kN (T-35-B-14d). The larger the working stress is, the more significant the degradation effect of anchor tension is. It is shown that the current test and study on the corrosion of anchor bolts can not accurately reflect the corrosion law of rock bolts in practical engineering if only the material or the axial force with small value is applied. The influence of stress on the corrosion test of rock bolts can not be ignored.

Rock bolts have replaced the conventional cog/chock as roof support from underground coal mines working with continuous miner (CM) technology in Bord and Pillar mining method with caving system. Rock bolts installed in the gallery at the goaf edge by aligning the adjacent stable rib during the CM-based mechanised depillaring have proven its potential by restricting goaf encroachment towards the working area. However, conventional depillaring with stowing and caving system still uses skin-to-skincog/chock as goaf edge support in Indian coalfields which poses threat to the safety of miners during its erection and withdrawal. Further, bearing capacity of such goaf edge support also reduces with increase in height of working. Field study found that the adjacent in-bye rib of the working slice remains intact and stable due to less span of hanging roof inside the goaf during conventional depillaring with stowing compared to caving. These field outcomes are also validated during the numerical simulation and observed that the goaf edges are over-supported in a stowing panel as it adopted the similar goaf edge support pattern as that of caving. Further, a parametric study is conducted on numerical models by varying the nature of roof, gallery width, and depth of cover to estimate the rock load height (considering 1.5 safety factor contour) around the goaf edges in stowing panels. Multivariate regression analysis was carried out to establish a relationship among influencing parameters for estimation of rock load height at the goaf edges in stowing panels for a given geo-mining condition. This paper presents a novel design of rock bolts as goaf edge support on the basis of roof-rib interaction in conventional depillaring with stowing panel.

High stress in surrounding rock mass could cause serious stability problems such as larger squeezing deformation in soft rock and rock burst in hard rock. The support system applied in high in situ stress conditions should be able to carry high load and accommodate large deformation of rock mass. This paper presented a new yielding rock bolt, called tension and compression-coupled yielding rock bolt, which is promising to provide support for both squeezing and burst-prone rock mass encountered in mining or tunneling at depth. The new bolt mainly consists of a steel rod and two additional anchors. The steel rod is a round shape bar with varying surface conditions. The inner segment is processed into rough surface, while the middle segment of the rod has smooth surface. Two additional anchors were arranged on both ends of smooth segment. The bolt is fully encapsulated in a borehole with either cement or resin grout. The rough segment and the inner anchor are firmly fixed in the bottom of the borehole, while the smooth segment has no or very weak bonding to the grout, which can stretch to accommodate rock dilatation. First, direct quasi-static pull tests were performed to examine the load capacity of tension and compression-coupled anchor. The results showed that the coupling action of tension to the rough rod and compression on the inner additional anchor by grout in different positions can increase the ultimate bearing capacity of inner anchoring segment significantly. Second, the performance of the new bolt and the fully encapsulated rebar bolt was tested under fracture opening condition. Results showed that the load and strain concentration could result in premature failure of fully encapsulated rebar bolt. However, the smooth segment of TCC Yielding rock bolt can detach from the grout under pull loading and provide a larger deformation to accommodate rock dilations. Third, shear tests were performed to examine the deformation mechanism of the new rock bolt under fracture sliding condition. Results showed that the smooth section of the new bolt specimen can deform freely to accommodate the sliding of fracture. The maximum shear displacement of the new bolt specimen is much larger than the fully encapsulated rebar bolt specimen, which is promising a better ability to accommodate the large displacement sliding of fracture in engineering practice.

The shear performance of bolted rock joints under cyclic loading condition was studied through laboratory shear tests. The tests were performed on a cyclic shear testing apparatus. The fully encapsulated rock bolts were inserted in the rock joints. The upper block was fixed in the horizontal direction, and the cyclic shear load was applied laterally on the lower block. A total of 16 shear tests were conducted for two types of rough joints. The results showed that the failure mode of rock bolts in cyclic shear test was different with the case of direct shear test. When the cyclic displacement was small, the rock bolt played a very small role in the second-to-fifth cycles, the rock bolt remained unbroken during cyclic shearing. The shear resistance of the rock bolt would gradually recover after the shear displacement has exceeded the cyclic displacement in the termination cyclic after five cycles. When the cyclic displacement was large, the rock bolt broke during cyclic shearing, and the shear resistance could not be recovered in the termination cyclic. It is indicated that the rock bolt had completely lost its supporting role after cyclic shear loading. The shear strength reduction of bolted joints under the cyclic loading condition was much significant compared to the un-bolted joints. Therefore, the influence of cyclic loading on the shear resistance of the rock bolt was much larger than its counterpart of the rock joint itself. These results showed that the shear performance of a rock bolt inserted in a rock joint was strongly influenced by cyclic shear loading.

This paper aims to propose new numerical procedures for elasto-plastic analysis of the interactions between fully-grouted bolt and rock with slip and non-slip cases at a potential-failure interface. In the proposed procedure with the slip case, the iteration process accounts for the elongation of the bolt, and the calculation process is greatly simplified by considering only the initial and final states (i.e., no intermediate finite states as is the case for existing procedures) without compromising the accuracy of the predictions. In the proposed procedure with the non-slip case, the elasto-plastic solution of the plastic rock strain is rigorously derived; the two scenarios in which the length of the rock bolt is contained within the plastic zone or within the elastic zone are comprehensively considered. The obtained solutions of tensile force in the bolt, shear stress at the potential-failure interface, rock displacement and stress components in the rock from the proposed procedures are validated by analyses with purely analytical methods and a commercial finite-difference program. With influences of rock quality and plastic softening behaviour of rock mass, the differences of the solutions in the slip and non-slip cases at potential-failure surface are compared; the dependency of the relative movement on the shearing stiffness at the potential failure interface is also investigated.

The problem of proper assessment of the technical functionality of rock bolt support systems is still valid. Many research centers have undertaken efforts to diagnose and monitor the technical state of such a support system used in mines and tunneling. With that aim the method of quality assessment of grouted rock bolts was invented and a relevant apparatus was constructed. The method concerns non-destructive identification of discontinuity of a resin layer (grout) surrounding rock bolts. The method is based on an impact excitation of a rock bolt and uses modal analysis procedures. Assuming that the installed rock bolt acts as an oscillator, different lengths and positions of grouting discontinuity alter its modal parameters. The extraction of these modal parameters, of which a resonant frequency is seen as the most valued, enable the relevant identification of grout discontinuity. After constructing a prototype version and validating the results for known cases of resin discontinuity in an experimental coal mine, the apparatus fulfilling ATEX requirements was developed. Subsequently that version was also verified both in laboratory conditions and in an experimental coal mine. As necessary for proper identification of discontinuity length, the reference data base was developed and elaborated consisting of a very large number of finite element models (FE models), namely discontinuity cases. The models encountered different rock bolt lengths and diameters, different rock strata parameters and different positions and lengths of resin layers. Then the method was used in a working coal mine to monitor a technical state of rock bolt support system mounted to reinforce long underground openings. The data base was utilized as reference for investigated rock bolts.

In coal mining roadway support design, the working resistance of the rock bolt is the key factor affecting its maximum support load. Effective improvement of the working resistance is of great significance to roadway support. Based on the rock bolt’s tensile characteristics and the mining roadway surrounding rock deformation, a mechanical model for calculating the working resistance of the rock bolt is established and solved. Taking the mining roadway of the 17102 (3) working face at the Panji 3# Well Coal Mine of China as a research site, with a quadrilateral section roadway, the influence of pretension and anchorage length on the working resistance of high-strength and ordinary rock bolts in the middle and corner of the roadway is studied. The results show that when the bolt is in the elastic stage, increasing the pretension and anchorage length can effectively improve the working resistance. When the bolt is in the yield and strain-strengthening stages, increasing the pretension and anchorage length cannot effectively improve the working resistance. The influence of pretension and anchorage length on the ordinary and high-strength bolts is similar. The ordinary bolt’s working resistance is approximately 25 kN less than that of the high-strength bolt. When pretension and anchorage length are considered separately, the best pretensions of the high-strength bolt in the middle of the roadway side and the roadway corner are 41.55 and 104.26 kN, respectively, and the best anchorage lengths are 1.54 and 2.12 m, respectively. The best anchorage length of the ordinary bolt is the same as that of the high-strength bolt, and the best pretension for the ordinary bolt in the middle of the roadway side and at the roadway corner is 33.51 and 85.12 kN, respectively. The research results can provide a theoretical basis for supporting the design of quadrilateral mining roadways.

Concrete lining and rock bolts are usually provided to support excavated tunnels on different ground conditions. These two support systems are worthy and need in-depth consideration as their impact on tunnel cost and deformation are critical. The optimum use of these support systems is a challenge that demands the best design solution. In order to facilitate the determination of the optimal combination of the two, an optimization scheme has been devised in this study using Finite Element Analysis (FEA) and MATLAB software for a specific under-construction tunnel in Karakore, Ethiopia. The tunnel displacements found from the FEA and the cost of tunnel construction were modeled using polynomial functions with a multi-objective optimization since the problem has two objectives. Considering displacement and cost functions, the optimum combination of rock bolt and concrete lining has been found. The result shows that the optimum combination can be found by increasing the number of rock bolts. The finite element analysis gives a total displacement of 6.34 cm in the best scenario case, which is slightly greater than the observed total displacement of about 6 cm. The total cost of rock bolt and concrete lining for the optimum combination becomes $30,700 per meter length of tunnel rock bolt and concrete lining construction.

The work results are part of the “Theories and me-thods development for geotechnical systems state ma-nagement to ensure mines performance intensification” research under National Academy of Sciences of Ukraine funding program.

In this paper, a unified rock bolt model is proposed and incorporated into the two-dimensional discontinuous deformation analysis. In the model, the bolt shank is discretized into a finite number of (modified) Euler–Bernoulli beam elements with the degrees of freedom represented at the end nodes, while the face plate is treated as solid blocks. The rock mass and the bolt shank deform independently, but interact with each other through a few anchored points. The interactions between the rock mass and the face plate are handled via general contact algorithm. Different types of rock bolts (e.g., Expansion Shell, fully grouted rebar, Split Set, cone bolt, Roofex, Garford and D-bolt) can be realized by specifying the corresponding constitutive model for the tangential behavior of the anchored points. Four failure modes, namely tensile failure and shear failure of the bolt shank, debonding along the bolt/rock interface and loss of the face plate, are available in the analysis procedure. The performance of a typical conventional rock bolt (fully grouted rebar) and a typical energy-absorbing rock bolt (D-bolt) under the scenarios of suspending loosened blocks and rock dilation is investigated using the proposed model. The reliability of the proposed model is verified by comparing the simulation results with theoretical predictions and experimental observations. The proposed model could be used to reveal the mechanism of each type of rock bolt in realistic scenarios and to provide a numerical way for presenting the detailed profile about the behavior of bolts, in particular at intermediate loading stages.

Rock bolts are commonly used in conjunction with Fibre Reinforced Shotcrete (FRS) linings to stabilise ground around excavations. A steel cover plate is often placed over the protruding end of a rock bolt to dissipate load from the bolt into the FRS lining. The process of drilling and installing rock bolts takes place after the lining is sprayed, so a hole must be made through the lining to insert the bolt. To examine the effect of this on lining behaviour, an investigation has been carried out to determine the influence of a central hole on the punching capacity of a FRC panel. Grout or concrete is commonly placed between the cover plate and lining to distribute the load more evenly, so the effect of this aspect of rock bolting practice on punching behaviour has also been examined.

The existence of uncertainty in the design parameters of anchored rock slopes warrants reliability based study. An approach for reliability based design optimization of anchored rock slopes against plane failure is described under different hydraulic distributions. Planar slide failure along with water pressure has been considered in conjunction with Barton-Bandis failure criterion to compute the factor of safety. Nonlinear strength criterion is employed for the computation of strength of jointed rock mass. The rock bolt and rock mass interaction model is used to analyze the stabilizing force in anchors. The variability associated with Barton-Bandis parameters, namely joint compressive strength (JCS), joint roughness coefficient (JRC), and basic friction angle (ψ) and a joint factor is considered. The target reliability based design optimization (TRBDO) is used to obtain an optimum length of the rock bolt and optimum anchor force required to stabilize the rock slope in order that the reliability index must meet a desired level of safety.

Usually, in a fully grouted rock bolt pullout test the load-displacement curve of the rock bolt head is recorded. This paper presents an analytical method to use this curve for determining the bond (bolt-grout and grout-rock interface) shear strength parameters. For this purpose, the fully grouted rock bolt interaction with grout and surrounding rock in the pullout test is investigated and the load-displacement curve of the bolt head (beginning of the bonded section) is obtained analytically. For modeling the bolt-grout interface behavior a distribution of the shear stress along the fully grouted rock bolt by consideration of bolt shank failure is used. In this regard, different stages including complete bonding, partial decoupling, decoupling with the residual shear strength and complete decoupling are considered. With increasing the applied load, two possible cases involving the rock bolt complete pullout and bolt shank yielding are taken into account. Based on the presented analytical method, the obtained bolt head load-displacement curve can be compared with the one recorded in the pullout test. With this, the relevance of selected shear strength parameters compared to real parameters can be assessed. A flowchart for determining the bolt bond shear strength parameters is presented using the trial and error method (coded in Matlab). The proposed solution is used to determine two experimental pullout shear strength parameters. The results show good agreement between predicted and calculated load-displacement curves.

This paper illustrates how the Bieniawski rock mass rating geomechanics classification, within the support system used by medium-scale mining companies, allows for the development of a new anchor rock bolt selection method aimed at improving horizontal access stability in underground mines. However, this case study only seeks to select the most efficient anchor rock bolt for any given horizontal access. A proper support selection method is very important for mining companies because this decision will safeguard both miners and mining infrastructure. This selection process will ultimately prevent fatal accidents, which is critical for mining companies today, especially considering the constant operation standstills reported in Peru.

Roof falls are currently one of the most dangerous threats associated with underground mining at great depth. Every occurrence of such an event poses a significant risk to the mining crew and disturbs the continuity of the mining process, which clearly affects the economy of the exploitation process. The development of a reliable monitoring system may significantly reduce the impact of eventual roof failure and will have a positive effect on the sustainability of the extraction process. Within this research study, a prototype of an instrumented rock bolt developed for continuous stress measurement is presented. The procedure of a 4-groove multilevel instrumented rock bolt is described and the calibration process is shown. Then, preliminary results of long-term in situ monitoring are presented. Based on the continuous monitoring of stress distribution within immediate roof strata, it was concluded that the developed instrumented rock bolt provides reliable results and is a very useful device, ensuring the possibility of early warning for miners about increasing roof fall risk.

Optimizing or revising support systems are inevitable due to the existence of high probabilities and uncertainties in underground mines. For this aim, field data and measurements are the only scales available to truly measure the performance of these systems. This paper presents a site-dependent approach for optimizing/revising the applied rock bolt support systems and estimating the efficiency of installation operation based on the field data for underground coal mines. For this purpose, a comprehensive database containing field data from 128 different mining locations of an underground coal mine was taken into account. At first, the relationships between field data and measured displacement by installed tell-tale extensometers (TTD) in the roof to determine the effective parameters on the roof instability were analyzed. In the following, a method was introduced to calculate the weighting factor of the effective parameters in the system using coefficient of determination (R2) and sensitivity index (SI). The sensitivity index was developed for sensitivity analysis of all considered parameters on TTD under fair conditions. Then, a method entitled field data-based risk index (FDRI) model was presented to evaluate the risk of roof instability supported by rock bolt systems which was validated by the considered case study. Finally, a process was proposed to optimize/revise a rock bolt design using the results of FDRI model and introduced data-driven efficiency (DDE) theory. DDE theory was provided to relative estimation of the efficiency of the installation operation of rock bolts. This research presents a systematic method for selecting the effective parameters on an issue or hazard, determining their weight in the system and carrying out relevant ratings needed for risk assessment.

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.

Anchoring mechanism and failure characteristics of composite soft rock with weak interface usually exhibit remarkable difference from those in single rock mass. In order to fully understand the reinforcement mechanism of composite soft roof in western mining area of China, a mechanical model of composite soft rock with weak interface and rock bolt which considering the transverse shear sliding between different rock layers was established firstly. The anchoring effect was quantified by a factor defined as anchoring effect coefficient and its evolution equation was further deduced based on the deformation relationship and homogenized distribution assumption of stress acting on composite structure. Meanwhile, the numerical simulation model of composite soft rock with shear joint was prompted by finite element method. Then detailed analysis were carried out for the deformation features, stress distribution and failure behavior of rock mass and rock bolt near the joint under transverse load. The theoretical result indicates that the anchoring effect of rock-bolt through weak joint changes with the working status of rock mass and closely relates with the physical and geometric parameters of rock mass and rock bolt. From the numerical results, the bending deformation of rock bolt accurately characterized by Doseresp model is mainly concentrated between two plastic hinges near the shear joint. The maximum tensile and compression stresses distribute in the plastic hinge. However, the maximum shear stress appears at the positions of joint surface. The failure zones of composite rock are produced firstly at the joint surface due to the reaction of rock bolt. The above results laid a theoretical and computational foundation for further study of anchorage failure in composite soft rock.

Rock bolts have been widely applied with roadway excavation in underground coal mines to prevent roadway collapse and improve the stability of roadway surrounding rocks. Overloading and failures of rock bolts could result in accidents or casualties in coal mine roadways. Consequently, monitoring axial forces and work conditions of rock bolts plays an increasingly important role in ensuring safe operations of underground coal mines. Conventional mechanical or electronic rock bolt monitoring systems are typically affected by electromagnetic interference, corrosive groundwater, and dusty circumstance in underground working sites. This work proposed a FBG-based instrumented rock bolt. Quasi-distributed FBG sensors were installed on a slotted rock bolt and encapsulated by epoxy resin that was used to fix FBG sensors on the rock bolt and protect FBG sensors from damage. The FBG sensors were calibrated before the in situ application. Monitoring results indicated that the axial forces of rock bolts installed on the roof of the roadway were higher than that of others, and the maximum axial forces of each rock bolt were typically detected near the middle portion of rock bolts. A real-time and accurate rock bolt monitoring system was established by integrating instrumented rock bolts to the existing monitoring system of the coal mine.

A rock bolt is a reinforcing bar commonly used in geotechnical engineering, and it plays an important role for the safe operation of the reinforced structures, such as tunnels, underground stations, and stabilized slopes. The length of an installed rock bolt is an important quality index that ensures the performance of a rock bolt for the stability of the reinforcing structure. In this paper, based on the general principle of stress wave propagation, the method of Automatic Extraction Algorithm (AEA) is developed using self-correlation analysis to measure the length of an in-service rock bolt, whose length is mostly embedded. The propose method has advantages of simple computation and high accuracy. Through laboratory experiments and field tests, we compared the performances of AEA with the traditional Phase Analysis Method (PAM) and the results demonstrated that the proposed AEA showed improved performance in terms of both detection speed and accuracy in rock bolt length estimation. In addition, the AEA method was successfully implemented in several engineering scenarios with high accuracy.

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.

In this study, different plate fixing designs were tested in order to improve rock bolting performances. Instead of commonly used steel nuts, different plate fixing designs were investigated by both laboratory and site studies. In addition, the usability of nuts and energy absorbers made of the polyamide type engineering polymer was investigated to assess whether they are advantageous for the aim of supplying high-energy absorbing features. In this experimental study, deformation-controlled load tests were carried out to evaluate the bearing and energy absorption capacities of totally eight different groups of specimens. According to the findings from this study, use of the polyamide material with steel nuts as an energy absorber material, rather than using instead of steel nuts supplies good results. As another outcome, it has been determined that use of proper washers between steel nuts and polyamide absorbers can significantly improve the energy absorption performances. It has been evaluated that the energy absorption capacity of the bolt plates can be economically increased up to 10 times by using suitable polyamide energy absorbers placed between the plate and steel nut parts. It was concluded that the polyamides can be used to remarkably improve the rock bolting performances especially in the rock masses where the high energy absorption capacity property is needed, like those with rock bursting, squeezing and swelling problems.