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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.

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.

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...

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.

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.

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.

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.

Rock bolts are commonly used on transportation projects to stabilize cut slopes, reinforce undermined or unsupported blocks, and reduce potential for large rock slides. Installing bolts can be challenging. Locations where bolts are useful may not be easily accessible for personnel or large equipment, and targeted blocks generally are marginally stable. In addition, jointed rock with openings and voids, which are typical conditions, complicate hole advancement and full grout encapsulation. These issues can lead to construction delays, claims, reduced corrosion protection, or shorter design life. This paper describes research conducted to address two rock bolting challenges in open-jointed rock: i) advancement of drill holes to design depth, and ii) successful completion of grouting through the application of injected polyurethane foam (PUF). Testing utilized double-stacked concrete blocks aligned in rows to simulate a rock mass with vertical fractures. Steel spacers were placed between the stacks to simulate 2-, 4-, 6-, and 8-inch apertures. A standard rock drill advanced holes using a 3-inch bit. Research showed PUF could: bridge simulated aperture sizes up to 8 inches, survive drilling action during secondary reaming, resist drilling air and grouting pressures, and reduce grout consumption.

The position determination for rock bolt support design based on rock stress distribution is important in ensuring full mobilization of rock bolt tensions during excavation process. The monitoring of the progressive rock stress variation in magnitude and orientation can play an important role in optimal installation of rock bolt. The aim of this paper is to determine the optimal positions for rock bolt installation by computing the progressive directional variation of rock stresses and plot the stress distribution in a two-dimensional stereographic projection to capture the stress transformation during excavation process. Finally, an algorithm based on stress transformation determination for optimal rock bolt installation is proposed.

Rock bolting is a method used for rock reinforcement in underground mining and tunnelling. There is a large variety of different types of rock bolts with different support functions. The behaviour of a rock bolt in a rock mass depends on the function and material of the bolt itself, combined with the mechanical properties of the rock mass, deformation capacity, strength and level of stress. Testing of rock bolts in full-scale laboratory-controlled conditions is therefore of great importance. At the rock mechanics laboratory of SINTEF and Norwegian University of Science and Technology (NTNU) in Trondheim, a rock bolt test rig has been developed for full-scale testing for pull, shear and combination pull-shear tests. This paper describes the principles behind this quasi-static full-scale testing and includes the results and analyses of recent tests performed on different types of rock bolts. It also discusses the applicability of the test rig for rock bolt selection and rock support design.

The article describes designs of injection rock bolts that combine the functions of bolts and injectors in reinforcement of jointed rocks during mining. The data of search for low-toxic compositions are presented. The injection rock bolting installation technology is developed.

The article presents the impact of geological and mining factors on the stability of room excavations in the Legnica-Głog&oacute, w Copper District (LGOM) in Poland. In underground mining, the primary task of bolting of mining excavations is to ensure their stability as an essential condition of work safety. Appreciating the role and importance of the rock bolting in Polish ore mining, rock bolt load sensors were designed, manufactured and tested under laboratory conditions. The purpose of the research was to characterize the sensors and determine the elastic range of the bearing plate, which are an integral part of the sensor. The sensors have been verified in industrial conditions. The tests were carried out in the underground copper ore mine in Poland. Three rooms in the exploitation field were selected for testing, where exploitation was carried out at a depth of 809-820 m below the ground surface with the application of room and pillar with roof deflection and maintaining the central part of the field. The exploitation field included 60 rooms and pillars. The effectiveness of the mechanical load sensor of the expansion rock bolt support has been experimentally confirmed. Based on mine research, it was found that the largest increases in the load of the rock bolting, vertical stress and convergence occur in the middle of the mining field.

The article presents methods of making the rock bolt support more yieldable, especially for a stratified roof. Alongside the increasing depth of exploitation of raw material deposits, rock bolt support units are more often designed taking into account more intensive deformations and displacements of underground excavations. In the article, a room and pillar method with mined roof bending and roof reinforcement with bolt patterns of 1 m &times, 1 m, 1.5 m &times, 1.5 m and 2 m &times, 2 m is presented. Moreover, the laboratory tests included 1.8 m long bolts, which were embedded segmentally on the lengths of 100 mm, 150 mm and 200 mm were tested. Based on the load&ndash, displacement characteristics, the deformation energy for flat and profiled dome bearing plates was calculated. Making the segmentally embedded resin rock bolt support yieldable enabled it to perform additional work. Furthermore, it was found that rock bolt support with a dome bearing plate took over 2.5 times more energy compared to a rock bolt support equipped with a flat bearing plate.

Rock bolting is one of the most effective and economical means of rock mass reinforcement. Existing studies of rock bolt reinforcement are mostly focused on rock masses without flaw, with a single flaw, or with parallel flaws. However in rock masses, cracks or flaws usually exist in the form of cross-flaws. In order to understand the impact of cross-flaws on rock bolt reinforcement and to further explore the differences of bolt reinforcement between rock mass with cross-flaws and rock mass with a single flaw, reinforced analog specimens with cross-flaws and with a single flaw were tested under uniaxial compressive condition. The experimental results show that the uniaxial compressive strength of the reinforced rock mass with cross-flaws in this research is higher than that of reinforced rock mass with a single flaw. This observation can be explained by the difference in the failure modes of reinforced specimens: the reinforced rock masses with a single flaw fail due to the formation of a shear crack while reinforced rock masses with cross-flaws fail as a result of a tensile fracture or interaction between tensile fracture and shear fracture.

To prevent large tunnel deformations caused by the rock bursts or the squeezing ground conditions, the rock bolts used should satisfy both strength capacity and the required deformability. Currently, the energy-absorbing rock bolts, such as the cone bolt and D bolt, have been successfully utilised in the deep mining to avoid sudden tunnel collapses. However, a rigid type rock bolt (e.g. a fully grouted rock bolt) is still commonly used in civil engineering tunnels even when they are excavated under a high overburden pressure with poor geological conditions. In such cases, the rock bolt might fail in tension due to the large deformation. This paper proposed a new energy-absorbing rock bolt, which is referred to as a deformation-controlled rock bolt (DC-bolt). The performance of the proposed DC-bolt was verified by the numerical simulations using the discontinuous deformation analysis (DDA) and by the prototype laboratory tests. This study concluded that the DC-bolt possesses both the high loading capacity and the deformable capacity. Additionally, the DC-bolt can limit the rock surface movement when it reaches a certain displacement, thus, it can be a useful tunnel support for civil engineering tunnels excavated under the squeezing ground condition.

The key factor in the operation safety for the shaft mining is stabilization of the working surface [1]. The time required to install support element is crucial [2, 3]. Surface stabilization can be done three ways: using frame support or rock bolt or their combination [4]. The rock stabilizing method is realized using significant number of design options of rock stabilizer systems [5, 6]. Friction rock bolts with tubular profile have become widespread recently [7 -9]. Its fixation in borehole is based on generation of friction force at contact surface. There are two methods realized: Swellex and Split Set. According to the first method the closed tubular profile (after being freely inserted into borehole) that has working liquid inside forced under the high pressure is used. Under the pressure outside dimension of profile is increased and its forcing against the borehole surface is provided [10]. The second method is based on use of spindle with unclosed tubular profile and outside diameter that exceeds the spindle diameter [11]. When spindle is inserted into borehole it elastically deforms at cross section. By means of that the distributed pressure is provided on contact surface. This operating principle of spindle and borehole interaction determined name of the support as self-installing rock bolt (SARB). Usage of SARB is preferable for building temporary support because of smaller time expenditure on installation.

In underground coal mining, high abutment loads caused by the extraction of coal can be a major contributor to many rock mechanic issues. In this paper, a large-scale physical modeling of a 2.6 × 2.0 × 1.0 m entry roof has been conducted to investigate the fundamentals of the fracture mechanics of entry roof strata subjected to high abutment loads. Two different types of roof, massive roof and laminated roof, are considered. Rock bolt system has been taken into consideration. A distinct element analyses based on the physical modeling conditions have been performed, and the results are compared with the physical results. The physical and numerical models suggest that under the condition of high abutment loads, the massive roof and the laminated roof fail in a similar pattern which is characterized as vertical tensile fracturing in the middle of the roof and inclined shear fracturing initiated at the roof and rib intersections and propagated deeper into the roof. Both the massive roof and the laminated roof collapse in a shear sliding mode shortly after shear fractures are observed from the roof surface. It is found that shear sliding is a combination of tensile cracking of intact rock and sliding on bedding planes and cross joints. Shear sliding occurs when the abutment load is much less than the compressive strength of roof.

Rock bolts are widely used in coal mines throughout China. Approximately 8,000 km of roadways are excavated in coal mines every year in China, 80 % of which are supported by rock bolts. At present, the design of rock bolt support schemes is mainly based on analogies and experience from previous projects. In the present study, in order to evaluate the design of rock bolt support in roadways, several cross sections of a roadway were monitored for rock bolt stress during the roadway excavation and mining. The study results show that the stress in the rock bolts varied in the areas 20 m behind the excavating face and 30 m ahead of the mining face. For the rock bolts observed in this study, the max axial force was within the design limit of the bolts, thus the support design was shown to be acceptable. Then, numerical simulation was performed using FLAC3D to investigate the stress evolution in the rock bolts during the mining of the fully mechanized longwall face. The simulation results show an overall agreement with the in situ measurements. Finally, parametric study pertaining to length, anchorage length, and rock bolt spacing was carried out with the numerical model, and several suggestions for the support design were proposed.

Computer modeling can be used to explore and gain new insights into the impacts of rock bolt intersecting joints in rock masses, and to estimate the effectiveness of the rock reinforcement system. In order to achieve this goal, we couple a rock bolt element into the two-dimensional discontinuous deformation analysis (DDA2D) program. The coupling algorithm is based on the analytically-derived interface behavior between a rock bolt and the rock material for grouted rock bolts. The shear force generated by slippage along the interface is assumed to have a linear relationship with respect to the relative slipping distance between the rock bolt and the rock. The linear elastic criterion is applied to determine the material behavior of rock bolts before the axial stress reaches the yield value. The pullout tests are simulated to verify the coupling algorithm and the effects of the proposed rock bolt elements. Parametrical studies are also carried out to analyze the effectiveness of the rock bolts under various end conditions, joint locations and bond stiffness. In addition, the performance of the rock bolt during the interface debonding is analyzed using two types of constitutive laws, i.e., the friction law and the reduction law. The simulation results show that the proposed rock bolt models can predict the shear forces and axial loading along the rock bolts.

Hollow grouted rock bolts are widely used to reinforce the surrounding rock mass in underground engineering, such as deep roadway of coal mines, deep tunneling. The anchorage capacity of hollow grouted rock bolts greatly influences the support quality. However, the anchorage capacity of hollow grouted rock bolt is usually destructively measured. In this paper, the grout ratio, defined as ratio of the grouted volume to the over-coring volume of anchorage, is introduced to measure the anchorage capacity. The quantitative relationship between the grout ratio and anchorage capacity of hollow grouted rock bolt by pull-out tests is established. Then, the non-destructive inspection on grout ratio of anchorage for hollow grouted rock bolts by stress wave reflection method is theoretically and experimentally investigated. By combining the two steps, a new non-destructive inspection method on anchorage quality of hollow grouted rock bolts is proposed. The results indicate that: (i) The anchorage capacity increases with the grout ratio increases up to 75% beyond which the benefits are lessened. (ii) The stress wave propagating along the bolt rod reflects and transmits at each interface impedance mismatch. The reflected signals can be used to identify the start and end of anchorage, and void part in anchorage. (iii) The results of wavelet transform analysis indicate the low-pass filters employed to reflected signals can be used to identify the start of anchorage, while the high-pass filters can identify the void part and the end of anchorage. (iv) A new supporting capacity monitoring method combined with pull-out tests and non-destructive method on hollow grouted rock bolts has been proposed, which can be used to inspect the on-site hollow grouted rock bolts in roadway on a large scale.

This paper used LS-DYNA simulation software to simulate rock bolt breaking position, displacements at different locations and the plastic deformation under dynamic load. Major drew the following conclusions: plastic deformation of the bolt away from breaks 600mm is only 1/20 compared with breaks; deformation of the rock bolt increased as the distance away from dynamic load applied position reduced, rock bolt deformation under dynamic load has a local effect; rock bolt deformation under dynamic load and rock bolt deformation under static load had essentially different.

One of the most serious hazard in the underground copper mining in Poland is the roof fall hazard. The scale and intensity of this type of events depends on several factors. Due to seismic activity, weaker roof strata can loosen and deform, generating additional load on the rock bolt support. The type of load depends strictly on the course of stratum deformation as well as on the intensity of the dynamic effect in the form of a seismic wave. Continuous monitoring is then necessary to determine the risk level of stability loss within existing excavations in order to minimize the probability of serious accidents. The following study analyses the impact of tremors recorded within the near-wave field upon the stress change in the instrumented rock bolts.

The purpose of this study was to investigate the effect of bolt profile on load transfer mechanism of fully grouted bolts in jointed rocks using analytical and numerical methods. Based on the analytical method with development of methods, a new model is presented. To validate the analytical model, five different profiles modeled by ANSYS software. The profile of rock bolts T3 and T4 with load transfer capacity, respectively 180 and 195 kN in the jointed rocks was selected as the optimum profiles. Finally, the selected profiles were examined in Tabas Coal Mine. FLAC analysis indicates that patterns 6+7 with 2NO flexi bolt 4 m better than other patterns within the faulted zone. Keywords: Fully grouted bolt, Load transfer mechanism, Jointed rocks, Analytical and numerical methods