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Rock bolts play a critical role in ground support in mining, tunneling and construction. Being able to obtain timely information on rock bolt load condition and deformation could help immensely in safeguarding workers' safety, optimization of ground support system, and ensuring stability and longevity of underground structures. The objective of this work was to develop a practical ultrasonic rock bolt sensing technology for monitoring axial load and deformation of a full-bodied rock bolt in both elastic and plastic deformation regimes. To this end, empirical mathematical models involving simultaneous use of times of flight of longitudinal and shear ultrasonic waves propagating along the axial direction of the rock bolt were developed, sensors were designed and fabricated, and laboratory studies were conducted. The results showed that the technology was able to measure load change, provide early detection of yield, and measure both plastic and total elongations of the rock bolt undergoing a pull test. Additionally, sectional load and elongation information could be obtained by applying the technology to rock bolts divided into sections by drilled holes along their shanks, therefore, providing the capability to assess the condition of grouted bolts at different depths within a rock mass. This work has laid the groundwork for the deployment of the technology in the field. Highlights: Empirical models based on simultaneous use of times of flight of longitudinal and shear ultrasonic waves were proposed for condition monitoring of full-bodied rock bolts Methods for instrumenting full-bodied rock bolts were developed Calibration methods for determining values of parameters used in the empirical models were established Pull testing showed that the technology could measure load changes within and beyond the elastic limit, detect yield, and measure plastic and total elongations of the rock bolt Sectional load and elongation information could be obtained by applying the technology to rock bolts divided into sections by drilled holes along their shanks [ABSTRACT FROM AUTHOR]

Estimating the axial force of rock bolts is vital for the stability and safety of underground excavations. Current methods suffer from being labor-intensive, time-consuming, and prone to inaccuracies. This paper presents a novel method for estimating the axial force of a rock bolt, a critical element in underground engineering, by analyzing the deformation field of the bearing plate. This innovative approach combines 3D laser scanning for data acquisition with a developed convolutional neural network (CNN) model for data processing, resulting in an efficient and non-destructive external monitoring technique. The proposed method was verified with 27 laboratory experiments. The results showed an accurate estimation of the axial forces of rock bolts with an average error of ± 5 kN, underlining the method's potential for in-situ application in underground engineering. This research not only contributes to the development of intelligent systems in rock mechanics and engineering but also has significant implications for industries, such as mining, tunneling, and underground engineering. Highlights: The study introduces a novel method that utilizes 3D laser scanning and a developed CNN model to accurately estimate the axial force of rock bolts. The proposed method achieves accurate estimations of rock bolts' axial forces with an average error of ± 5 kN. Future work includes the integration of object detection algorithms for large-scale 3D laser scanning and the creation of a database of various bearing plates to enhance the method's robustness. [ABSTRACT FROM AUTHOR]

The mechanical and deformation properties of bolted joints are important research issues in rock mechanics and rock engineering. Quite a few tests have been performed to study the performance of rock bolts in resisting shear loading. The effect of cyclic shear loadings on the bolts in the joints under different joint roughness was studied through a series of shear tests. The results illustrate that the peak shear strength was substantially equal to the residual strength with the increase in the number of cycles, and the slope of the normal displacement curve gradually decreased. When the number of shears is greater than three, the shear stress curve and the slope of the normal displacement curve tend to be stable gradually, which means it has completed the asperity degradation. The cyclic shear loading affects the rock bolt and reduces its bearing capacity. For specimens with larger JRC, the peak shear stress in both the direct shear and the first cycle was significantly elevated compared with small JRC, and so was the normal displacement. An evaluation index of shear energy loss ratio (SELR) was introduced to evaluate the shear performance of the rock bolts and rock joints under a cyclic shear loading condition. When the JRC is large enough, the asperities are difficult to degrade, and the jointed rock still has a more remarkable ability to resist shear deformation. This shows that JRC significantly affects the performance of joint plane and rock bolts under cyclic shear loading. Highlights: The effect of cyclic shear loading on the bolted rock joints under different JRC conditions were investigated. The cyclic shear load affects the rock bolt and reduce its supporting capacity. An evaluation index of shear energy loss ratio (SELR) was introduced to evaluate the shear performance of the rock bolts and rock joints under a cyclic shear loading condition. [ABSTRACT FROM AUTHOR]

This Suggested Method proposes impact tests to determine selected properties of full-scale rock bolt assemblies subject to an axis-aligned impact force. These properties include the stiffness of the bolt assembly, the rise time, the tensile strength, the ultimate displacement, and the energy absorption of the rock bolt assembly. These properties may be used to design or select ground support for burst-prone ground conditions. This Suggested Method is limited to the mass freefall impact test method. It is performed using a test frame in the laboratory. The Method specifies the requirements for the testing apparatus and measurement sensors, the test procedures, the data processing methods and the reporting items. [ABSTRACT FROM AUTHOR]

Excavation of shallow-buried, hard rock masses usually leads to the instability of rock blocks delimited by discontinuity planes. Classic block theory only deals with the stability of rock blocks on the exposed surfaces. However, in rock engineering practice, it is quite common that when surface unstable blocks collapse along with their constraints to adjacent blocks released, more inner blocks will be exposed to the surface and become unstable or fall down. This process is termed the progressive failure of blocky rock mass. Based on three-dimensional (3D) block-cutting analysis, the removable blocks in progressive failure process can be determined through geometrical and kinematics analysis. This study proposes to further identify unstable key blocks from removable blocks using mechanical analysis. On this basis, the anchoring force required for all of randomly distributed key blocks is calculated, and the spatial distribution of the required anchoring force per unit area of the free surface are analyzed, with suitable probability distribution obtained. Afterwards, adopting reliability-based design concept, the design scheme for systematic rock bolting to prevent progressive failure of blocky rock mass is discussed, and the method is presented for estimating the anchoring force of individual rock bolt, bolt spacing, and anchorage length. Highlights: Progressive failure of blocky rock is analyzed through the geometrical-mechanical identification of key blocks. A new approach to assess the required anchoring force per unit area of rock excavation surface is proposed. Systematic rock-bolting design scheme for coping with progressive failure is presented and discussed. [ABSTRACT FROM AUTHOR]

Conventional tunnelling methods (i.e., the drill and blast tunnelling method or the new Austrian tunnelling method) have been frequently applied for the construction of tunnels/caverns under a high overburden and with a large cross section. Under such adverse conditions, tunnel support materials may yield because of ground pressure and tunnel deformation. As a result, a tunnel may lose its stability due to a reduced effectiveness of the tunnel supports. Rock bolts have been widely used as an essential component of tunnel support, and many studies have been conducted on the performance of rock bolts in strengthening the jointed rock mass. However, there is a lack of understanding on the interface behaviour between the rock bolt and the bond material, especially the crack initiation and propagation inside the bond material. This study aims to investigate the interface behaviour between rock bolts and bond materials using laboratory tests (shear tests) and numerical simulations (Discontinuous deformation analysis). By assessing the impact of the key parameters in the rock bolting system, such as the ribs, rib angle, strength of the bond material and the confining pressure, we are able to better understand the supporting mechanism and the effects of rock bolting. [ABSTRACT FROM AUTHOR]

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. [ABSTRACT FROM AUTHOR]

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. [ABSTRACT FROM AUTHOR]

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. [ABSTRACT FROM AUTHOR]

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 (DDA) 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. [ABSTRACT FROM AUTHOR]

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 FLAC 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. [ABSTRACT FROM AUTHOR]

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. [ABSTRACT FROM AUTHOR]

In this paper, after determining an appropriate experimental design and testing scheme to test the shear performance suitably, the impacts of pretension and shear interfaces on fibreglass rock bolts were investigated. For this purpose, at the first stage, the double shear testing apparatus was modified to address the issues and shortcomings of the established testing methodology. Then, 20-tonne and 30-tonne rock bolts were tested with pretension loads of 0 kN, 10 kN, 15 kN, and 20 kN for each test scheme. For this, a 40 MPa small aggregate concrete was used for all samples to simulate the host rock. The grout used to anchor the rock bolts was mixed to the product's recommended guidelines to ensure close resemblance to field mixtures. To maintain consistency, all samples were manufactured with the same concrete and grout recipes. The results of double shear tests revealed that increasing the pretension also increased the confining pressures at the shear interfaces for clean joints. This, in turn, reduced the damage propagating from the bolt at the shear interface and decreased the hinge point bending. The results can be applied for further analytical and numerical studies, which could utilise for a better understanding of the shear behaviour of fibreglass rock bolts. Highlights: Two types of fibreglass rock bolts, the 20-tonne and 30-tonne rock bolts, were tested utilising the modified double-shear apparatus. The double-shear testing apparatus was modified to address the issues and shortcomings of the established testing methodology. The double shear performance profiles of the bolts could be described by three regions categorised as the elastic, strain-softening, and failure regions. Increasing the pretension of each rock bolt type resulted in a reduction in the severity of the recorded shear force dip. The increased pretension yielded less physical damage to the rock bolt propagating away from the shear plane. [ABSTRACT FROM AUTHOR]

In underground mining practice, the rock bolt support system is the major support pattern to control the deformation and stability of openings. A rock bolt is generally subjected to complex loads including tension, torsion, bending and shear, which result from the deformation of excavations and exposure to dynamic loads that are generated by rockbursts. An understanding of the response of rock bolt under complex conditions is of great importance for rock bolt support design and practice. New sophisticated equipment has been developed for this purpose. This work involved a comprehensive experimental study on the mechanical behavior of rock bolts under complex loads. The results show that rock bolt pre-tensioning by torque application to the nut can result in decreases in tensile strength and elongation because the rock bolt is subjected to a combination of tension and distortion. When a pre-tensioned rock bolt is subjected to a shear load, the maximum shear force can reach up to 80% of the tensile capacity of the rock bolt. Higher impact energy results in a longer period of dynamic loading and a larger irreversible plastic deformation on the rock bolt, in contrast to a rock bolt that is subjected to low impact energy. The capacity and especially the deformation capacity of a rock bolt may decrease significantly after successive containment of the deformation of the surrounding rock mass from rockbursts. [ABSTRACT FROM AUTHOR]

In this article, the rock bolt reinforcement mechanism is discussed, and the gravity method as well as the finite element method for the bolted rock is presented. These methods are applied to study the gravity dam foundation of the Xiaoxi Hydropower Project, which is influenced by the presence of a large fault with a cracked zone over 180 m wide. Rock bolt reinforcement was used to improve the stability of the dam foundation, and the reinforcement parameters were determined from a semi-empirical equation controlled by in situ test. The stability analysis was conducted using the above methods, and the improvement in terms of deformation and stress as well as stability of the dam foundation due to the reinforcement is assessed. Rock bolt reinforcement was completed successfully, and the dam started normal operations in January 2008. [ABSTRACT FROM AUTHOR]

Wide-span openings in horizontally layered jointed rock were analysed using the distinct element method (DEM) to provide additional insight into rock bolt support mechanisms and forces. The rock bolt forces in the DEM calculations exhibited distinct mechanisms, identified here as reinforced rock unit, beam building or suspension mechanisms. Each of these mechanisms was found to be associated with certain characteristics in the jointed rock structure. In jointed rock with thick horizontal beddings, the rock bolt forces at the roof tapered off with distance leading to a reinforced rock unit where each layer exhibited a self-support mechanism akin to a voussoir beam. In jointed rock with thinner horizontal beddings not capable of self-supporting, a beam building mechanism took place. When the rock bolts were extended into a rigid layer above the thinner layers, a suspension mechanism took place. Potential inaccuracies of thrust line predictions and the benefits of rock bolt tensioning were investigated, and their effects were found to be sensitive to the bedding thickness. A case history of the cavern roof of the underground car park for the Sydney Opera House was modelled. The calculated deflections of the cavern roof compared well with field measurements reported in the literature. The findings in this paper highlight the importance to ensure that the reinforcement support for wide-span underground excavations in horizontally layered jointed rock are tailored to the geological conditions being encountered. [ABSTRACT FROM AUTHOR]

Fully grouted rock bolts are one of the main forms of rock reinforcement used in underground mines, especially in coal mines. Analytical models describing the deformation behaviour of rock bolts under various loading conditions are used in numerical models for designing appropriate rock reinforcement. A reliable analytical model capable of accurately predicting rock bolts' responses under different loading conditions is, therefore, important for more realistic numerical models and effective reinforcement design. In this paper, a new analytical model for shear behaviour of rock bolts is proposed which can better predict rock bolt behaviour under large shear deformations. A method to consider the effect of post-elastic strain hardening on bending stiffness of rock bolt is presented. The analytical model is validated with results from laboratory shear tests and found to be consistent with the laboratory results. To analyse the effect of post-elastic strain hardening on shear behaviour of a rock bolt, results from the proposed analytical model is compared with analytical model without strain-hardening effect. Without strain hardening, the analytical models overestimate the ultimate shear displacement by up to 35% in the examples shown in the paper. [ABSTRACT FROM AUTHOR]

Intelligent rock bolts can be used in mining operations to acquire information on the bolt's status and movements of the rock mass. The future value of this technology resides in optimizing the support layout and increasing safety. The primary goal of this research is to investigate the sensor spatial distribution necessary for efficiently identifying roof wedges in mines. Roof wedges occur when three joints intersect over the roof, forming a triangular shape. They present a hazard by falling and causing damage. Sensor density refers to the number of sensors on a single rock bolt and the number of intelligent rock bolts in the mine or one section. Too many sensors drive costs up, and too few sensors compromise measurement quality. This research visualizes and investigates the results of the simulated models of intelligent rock bolts depending on the number of installed sensors and intelligent bolts to get a better understanding of the problem and describe the technique used, including numerical simulations and the considered parameters. Furthermore, it offers insight into the findings' possible implications, providing knowledge about the viability of using intelligent rock bolts. [ABSTRACT FROM AUTHOR]

An inflatable bolt is integrated in the rock mass through the friction and mechanical interlock at the bolt-rock interface. The pullout resistance of the inflatable bolt is determined by the contact stress at the interface. The contact stress is composed of two parts, termed the primary and secondary contact stresses. The former refers to the stress established during bolt installation and the latter is mobilized when the bolt tends to slip in the borehole owing to the roughness of the borehole surface. The existing analysis of the inflatable rock bolt does not appropriately describe the interaction between the bolt and the rock since the influence of the folded tongue of the bolt on the stiffness of the bolt and the elastic rebound of the bolt tube in the end of bolt installation are ignored. The interaction of the inflatable bolt with the rock is thoroughly analysed by taking into account the elastic displacements of the rock mass and the bolt tube during and after bolt installation in this article. The study aims to reveal the influence of the bolt tongue on the contact stress and the different anchoring mechanisms of the bolt in hard and soft rocks. A new solution to the primary contact stress is derived, which is more realistic than the existing one in describing the interaction between the bolt and the rock. The mechanism of the secondary contact stress is also discussed from the point of view of the mechanical behaviour of the asperities on the borehole surface. The analytical solutions are in agreement with both the laboratory and field pullout test results. The analysis reveals that the primary contact stress decreases with the Young's modulus of the rock mass and increases with the borehole diameter and installation pump pressure. The primary contact stress can be easily established in soft and weak rock but is low or zero in hard and strong rock. In soft and weak rock, the primary contact stress is crucially important for the anchorage of the bolt, while in hard and strong rock it is the secondary contact stress that plays a vital role. [ABSTRACT FROM AUTHOR]

As coal mining progresses to greater depths, the rate and severity of coal burst hazards have greatly increased, resulting in significant operational and safety challenges. Appropriate rock support design for coal burst-prone roadways requires the understanding of the mechanical characteristics of the support system under impact loads and the underlying damage mechanisms. In this study, the response of rock bolts under coal burst conditions is assessed, aiming to advance the knowledge concerning the application of rock bolt systems in coal burst-prone mines. This is achieved by a review of typical failure patterns of rock bolts caused by coal bursts in a typical coal burst coalfield in China, a field study on the response of rock bolts subjected to a simulated coal burst via blasting and a laboratory study on the mechanical properties of rock bolts that have suffered coal bursts. It is believed that the outcome of this study has added to the improved understanding of the performance and design of rock bolt support systems in coal burst-prone mines. [ABSTRACT FROM AUTHOR]

Most accessible mineral deposits are nearly exhausted. Consequently, some resources that were previously abandoned are now being considered for remining. This paper describes a case study that aims at revealing the instability mechanism and selecting the optimal control countermeasures for a cataclastic roadway regenerated roof (CRRR) during the extraction of the remaining mineral resources. Based on the structural model of a cataclastic regenerated roof (CRR) of a remaining lower slice, a modified natural arch mechanical model for a CRRR was established by considering the lateral pressure and stress concentration coefficient of the bearing arch, through which the modified bearing arch curve equation and arch rise expression were deduced. The relationship between the modified arch rise and the lateral stress was obtained through a practical case. Next, a numerical model of a CRRR, which accounted for the cataclastic features, was built in the Universal Distinct Element Code (UDEC). The development of subsidence and fractures in the roadway roof was determined through modeling under various roof support scenarios, namely no support, rock bolt (RB) support, rock bolt and cable (RBC) support and a combined support of the rock bolt, cable and steel beam (RBCS). The theoretical analysis, numerical simulation and monitoring results of the RB and RBC stress all indicated that a bearing arch existed in the CRRR and that the arch rise underwent changes during the roof hanging time and with different support scenarios. The support structure should exceed the bearing arch rise in height and should be leak proof to avoid roof leakage. After a comparative analysis of the numerical and field support effects provided by various supporting scenarios, the RBCS support scenario was shown to be the optimal roof support choice, providing a good controlling effect for a CRRR and meeting the production requirements. [ABSTRACT FROM AUTHOR]

The selection of ground support for underground excavations typically focuses on matching a system to the anticipated ground conditions. In this context, the emphasis on materials selection involves picking a rock bolt type that meets a set of predefined capacity and deformation properties to manage the applied loads. Exposure of ground support to corrosive environments in underground hard rock mines can result in significant material degradation and loss of mechanical performance. This can have significant repercussions on the structural integrity of excavations in rock and potential falls of ground that may endanger the safety of workers and equipment. Consequently, there is a need to ensure that the selected rock bolt provides sufficient resistance to corrosion for the working life of the excavation. Current design practice is limited in selecting rock bolt types perceived as more resistant to corrosion, e.g., grouted versus friction rock stabilizer bolts. This investigation suggests, however, that there are significant variations in corrosion susceptibility of rock bolts that may otherwise display similar mechanical behavior. This has significant implications for mines operating in very corrosive environments. [ABSTRACT FROM AUTHOR]

The article offers information on improving performance of rock bolt using polyurea coating. Topics discussed include use of rock bolts by rock engineers for mining and tunnelling process, polyurea as an isocyanate- based copolymer used for application of surface treatment and used as spraying membrane in tunneling, and testing the capacity of polymer-coated rock bolts (PCRBs) and corrosion prevention capacity of the polyurea coating.

Shear failure is the main failure mode of anchorage bolts in coal mines. To improve ground control, it is crucial to improve the shear strength of roadway anchorage systems. In this study, theoretical analysis, numerical simulation and laboratory experiments were carried out to study the shear characteristics of the cuneiform reaming anchorage. The results show that the cuneiform reaming anchorage can reduce the maximum shear stress on the cross section of the anchorage to improve the shear strength. The stress concentration near the joint planes can be transferred to middle parts of the bolts. Moreover, the cuneiform structure can also prolong the period in which the shear load of the anchorage decreases and shorten the period in which the shear load increases, thus reducing the shear load of the anchorage. This paper provides a new economical and practical method for raising the shear strength of an anchorage system and enhancing roadway stability. [ABSTRACT FROM AUTHOR]

This paper experimentally compares the shear behavior of fiber glass (FG) bolt, rock bolt (steel rebar bolt) and cable bolt for the bolt contribution to bolted concrete surface shear strength, and bolt failure mode. Two double shear apparatuses of different size were used for the study. The tensile strength, the shear strength and the deformation modulus of bolt control the shear behavior of a sheared bolted joint. Since the strength and deformation modulus of FG bolt, rock bolt and cable bolt obtained from uniaxial tensile tests are different, their shear behavior in reinforcing joints is accordingly different. Test results showed that the shear stiffness of FG bolted joints decreased gradually from the beginning to end, while the shear stiffness of joints reinforced by rock bolt and cable bolt decreased bi-linearly, which is clearly consistent with their tensile deformation modulus. The bolted joint shear stiffness was highly influenced by bolt pretension in the high stiffness stage for both rock bolt and cable bolt, but not in the low stiffness stage. The rock bolt contribution to joint shear strength standardised by the bolt tensile strength was the largest, followed by cable bolts, then FG bolts. Both the rock bolts and cable bolts tended to fail in tension, while FG bolts in shear due to their low shear strength and constant deformation modulus. [ABSTRACT FROM AUTHOR]

Summary The three dimensional elasto-viscoplastic composite element method is formulated in this paper for rock masses reinforced by a fully-grouted bolt. If a bolt segment penetrates a finite element representing the rock mass, then a composite element is formed including five sub-elements corresponding to the rock material, the grout material, the bolt material, the rock-grout interface and the bolt-grout interface. The displacements in each sub-element are interpolated from the corresponding nodal displacements of the composite element. By the virtual work principle the governing equation for the solution of the nodal displacements can be formulated. The elasto-viscoplastic characteristics of the materials are considered in the formulation. The new model can be incorporated into the conventional finite element analysis grid, in which several composite elements have fully grouted bolts embedded. In this way the mesh generation of large scale bolted rock structures becomes convenient and feasible. The model has been implemented in a FEM program, and a comparative study between the numerical analysis and a pull out field test has been carried out, from which the validity and the robustness of the new model are justified. [ABSTRACT FROM AUTHOR]

Highlights: The fully encapsulated bolt is prone to stress concentration, which leads to the bolt breaking suddenly. Energy-absorbing bolt adapted to the large deformation of surrounding rock through slight yield but long-term support. When filling thickness to asperity height ratio (t/a) is small, the joint shear failure behavior is controlled by the asperities and the filling. When t/a is large, the joint shear failure mainly exists within the filling. [ABSTRACT FROM AUTHOR]

Understanding the performance and failure mechanisms of surface support in mine roadways is paramount to the success of mine productivity. A sophisticated facility has been developed to evaluate the mechanical behavior of surface support under quasi-static loading. Three different types of mesh, welded wire mesh, chain-link mesh, and geogrid mesh are tested in conjunction with different surface components including rock bolt plate, steel strap and W-strap. It is found that the welded wire mesh has the higher stiffness and load-bearing capacity than the diamond mesh and the geogrid mesh. For each mesh, the incorporation of steel strap or W-strap helps the load to distribute more uniformly to more wires than just adopting bearing plate alone, leading to a substantial improvement in both the stiffness and the load-bearing capacity. The welded wire mesh can still withstand large deformation after peak load. In contrast, the diamond mesh and the geogrid mesh lose its capacity quickly after peak load. [ABSTRACT FROM AUTHOR]

For road tunnels in most Nordic countries, temporary rock support installed during tunnel excavation usually becomes a part of the permanent rock support. Rock bolts are the most common rock support measure in road tunnels excavated in hard rock, and their conditions over the period of tunnel operation play a significant role in the safety of these tunnels. The rock bolt types and aggressive environmental conditions considered in this research are focused on those used and observed in Norwegian and Swedish road tunnels. Findings elsewhere in similar environments are included to highlight the different kinetics of degradation. Based on the collected data, the corrosion levels in these road tunnels are comparable to those in the most polluted industrial areas. Aggressive groundwater conditions for rock bolts include a groundwater pH below four (which can be caused by sulfuric acid formation due to the oxidation of the mineral pyrite in the rock), marine groundwater, and flowing groundwater with a high concentration of dissolved oxygen. Furthermore, chloride-bearing deicing salts commonly used on roads located in cold climate regions during winter can promote corrosion in rock bolts from the tunnel room. For these environments, this research proposes a lognormal probability function to quantify the expected steel corrosion level for 25, 50, and 100 years of exposure time. The corrosion protection given by cement grouting, hot-dip galvanizing and epoxy coating are also addressed to explore their contributions to the lifespan extension of rock bolts in acidic and chloride-rich environments. [ABSTRACT FROM AUTHOR]

The 3-m-long rock bolt, steel pipe and accessories are hot-dip galvanized and epoxy coated (Espedal and Nærum 1992). c CT-Bolt used in the Frøya subsea road tunnel constructed between 1998 and 2000. 3 Some rock bolts used as permanent rock supports in Norwegian road tunnels. a Uncoated rebar cement grouted in the Haukeli road tunnel constructed between 1968 and 1969. [Extracted from the article]

A systematic approach of using the distinct element method (DEM) to provide useful insights for tunnel support in moderately jointed rock masses is illustrated. This is preceded by a systematic study of common failure patterns for unsupported openings in a rock mass intersected by three independent sets of joints. The results of our simulations show that a qualitative description of the failure patterns using specific descriptors is unattainable. Then, it is shown that DEM analyses can be employed in the preliminary design phase of tunnel supports to determine the main parameters of a support consisting of rock bolts or one lining or a combination of both. A comprehensive parametric analysis investigating the effect of bolt bonded length, bolt spacing, bolt length, bolt pretension, bolt stiffness and lining thickness on the tunnel convergence is illustrated. The highlight of the proposed approach of preliminary support design is the use of a rock bolt and lining interaction diagram to evaluate the relative effectiveness of rock bolts and lining thickness in the design of the tunnel support. The concept of interaction diagram can be used to assist the engineer in making preliminary design decisions given a target maximum allowable convergence. In addition, DEM simulations were validated against available elastic solutions. To the authors' knowledge, this is the first verification of DEM calculations for supported openings against elastic solutions. The methodologies presented in this article are illustrated through 2-D plane strain analyses for the preliminary design stage. More rigorous analyses incorporating 3-D effects have not been attempted in this article because the longitudinal displacement profile is highly sensitive to the joint orientations with respect to the tunnel axis, and cannot be established accurately in 2-D. The methodologies and concepts discussed in this article, however, have the potential to be extended to 3-D analyses. [ABSTRACT FROM AUTHOR]

A new analytical approach able to predict the mechanical behavior of fully grouted rock bolts subjected to pull-and-shear load is been proposed. The elastic theory of a semi-infinite length beam and approximate differential equation of the deflection curve (elastic and elastoplastic) are employed to analyze the lateral behavior of a bolt under surrounding rock pressures. Bolt axial behavior is based on a realistic tri-linear bond-slip model with residual bond strength at the grout–bolt interface. According to the Tresca yield criterion, important failure parameters such as shear displacement, tangent direction, tensile displacement, and tensile force can be determined via the analytical approach while the shear force and other basic parameters are provided. Results of the proposed analytical approach for 0°, 20°, 40°, 60°, and 90° displacing angles correlate well with experimental test results. Parametric analysis of both the rock and bolt mechanical behavior are also carried out using this new analytical approach. [ABSTRACT FROM AUTHOR]

This paper presents a case study on the impact of jointed rock mass deformation on liner stability for an inclined excavation at the Bulianta Coal Mine. The tunnel was excavated by a single shield tunnel boring machine (TBM), with part of the excavation passing through a heavily jointed rock mass at great depth. Serious local instabilities occurred at one side of the liner due to highly asymmetric loading caused by large ground stresses and joint deformation. A support system with anchored cables and compressible mortar was combined with segmental lining to overcome these instabilities. With this case as a baseline, some numerical sensitivity analyses were performed to study how the ground might have responded to different combinations of support. The results show that an appropriately anchored cable system was able to limit the propagation of the asymmetric deformation of the jointed rock mass and significantly reduce the degree of anisotropy and the magnitude of the loads on the liner. The use of compressible mortar between the liner and the jointed rock mass further lowered the anisotropy of the load on the liner. There are optimum values for the rock bolt length and the compressible mortar thickness, beyond which these parameters have limited effects on the behaviour of the system. The combined effect of compressible mortar and anchor cables in promoting ground and liner stability is greater than the sum of the two individual support effects. These results clearly demonstrate that the combination of anchored cables and compressible mortar had a significant positive effect on the overall support system stability. [ABSTRACT FROM AUTHOR]

The mechanical performance of fully grouted rock bolts is essential in the stability of underground excavations in jointed rock masses. This research implements a new cohesive contact model in distinct element codes (PFC2D) to investigate the fracturing response of rock-like and grout material, as well as the bolt–grout interface. The results are compared in detail with experimental observations. The proposed modelling approach, used in conjunction with the distinct element method (DEM), successfully predicted the behaviour of grout failure and the bolt–grout interface's shear response. We then developed a novel numerical, stepwise pull-and-shear test (SPST) scheme to further analyse the mechanical behaviour of bolted rock joints subjected to simultaneous pull–shear loading. The cohesive DEM framework proposed in this paper was used to carry out the SPST scheme numerically. The mechanism involved in enhancing the shear strength of bolted rock joint was determined by monitoring the σ n i and its corresponding contact chain force network during the pull-out test. The influence of pretension stress, the rig angle of the bolt profile, and the constant normal stiffness (CNS) condition are assessed systematically. In particular, the pretension stress magnitude at which the synthetic rock bolting system exhibits the highest shear resistance is identified. The findings from this research highlight the sensitivity of bolted rock joints to the simultaneous pull-and-shear loading, boundary conditions, and bolt–grout interface configurations. [ABSTRACT FROM AUTHOR]

Anchoring is a highly effective means to reinforce jointed rock masses. In addition to the axial force from the anchor bolt, the effect of its transverse force in constraining the displacement of the structural plane cannot be ignored. Thus, this paper studies the combined action of the axial and shear forces of an anchor bolt on enhancing the shear strength of the anchored joint. In this work, first, laboratory direct shear tests are conducted to study the evolution of the peak shear strength of the anchored joint with increasing angle between the shear direction and bolt inclination. Second, to further verify this trend, numerical simulation with the finite difference method is used to simulate the same direct shear tests to reveal the lateral and axial stresses and deformations of the bolt. Finally, according to classical beam theory, the mechanical analysis model of a bolt subjected to the combined action of axial tension and lateral shear is established, which takes the realistic stress and deformation patterns of the bolt from the numerical simulation into account. The results show that the lateral restraint of the bolt plays a significant role in the shear strength of the anchored joint, its enhancing effect first increases and then decreases with increasing installation angle, and the optimal installation angle providing the maximum shear strength is approximately 60°. In addition, it is concluded that the shear resistance of the bolt is mainly due to its axial force when the installation angle is relatively small; conversely, the shear resistance of the bolt is mainly due to its shear force when the installation angle is relatively large. Highlights: A new approach for calculating the anchoring force considering the combined action of the transverse restraint and axial restraint of the bolt is presented. A calculation method for the optimal anchoring angle is presented, and its value is suggested to be approximately 60° in this study. The contribution rates of the axial force and tangential force to the anchoring force are not equal and vary with the change in the anchoring angle. [ABSTRACT FROM AUTHOR]

In this work, dynamic test results of D-bolts are presented. The rock bolt specimens studied are 20 and 22 mm in diameter and 0.8-1.5 m in section length. The bolts were tested at an impact velocity of 5.4-6.2 m/s and with impact energy varying from 10 to 60 kJ. In total, over 50 drop tests were conducted during a period of three years. The dynamic tests show that a D-bolt section of 22 mm × 1.5 m can sustain an impact of 56 kJ of a dynamic impact and absorb 60 kJ of energy prior to failure. The maximum impact energy of the 22 mm bolt is thus 37 kJ/m of bolt and the maximum energy absorption is 40 kJ/m. The displacement of the D-bolt increases linearly with the impact energy. A theoretical solution has been obtained for the relationship between the impact energy and the displacement of the bolt. It states that the energy absorption of a D-bolt section is proportionally related to the volume of the bolt section and the tensile strength and ultimate strain of the bolt material. It was also found that the magnitude of the impact energy determines whether or not the bolt fails, while the impact momentum determines how long the impact lasts. The impact duration increases linearly with momentum as long as the bolt shank does not fail. [ABSTRACT FROM AUTHOR]

Pre-driven recovery rooms allow for the safe and rapid extraction of longwall panel face equipment. Optimum support design requires an understanding of the loading mechanisms of pre-driven longwall recovery rooms subjected to large abutment pressures. This paper presents a case study evaluating the ground response of a pre-driven recovery room. The recovery room was supported by a rock bolt and cable support system in conjunction with two rows of concrete cribs. A numerical analysis of the pre-driven recovery room was conducted using the distinct element code UDEC. The numerical results were found to be in good agreement with field observations in terms of the patterns and magnitude of stress changes, roof-to-floor convergence and failure patterns. The present results suggest that the stresses carried by the outby pillar and inby fender began to significantly increase when the longwall face was approximately 20 m away. When the longwall face entered the recovery room, the stress concentration coefficient ranged from 3.0 to 3.5 in the inby fender and from 2.0 to 2.5 in the oubty pillar, resulting in spalling failure of the room ribs. The set of longwall face equipment was safely and successfully recovered. The concrete cribs, in conjunction with the rock bolts and cables, were considered effective, but conservative. It was also found that the stiffness of the concrete crib is critical to the ground response and must be considered when determining the required capacity. From the study results, design guidelines for determining the optimal support requirement of a pre-driven recovery room are proposed. [ABSTRACT FROM AUTHOR]

Electro-hydraulic and jumbo drills are commonly used for underground coal mines and tunnel drives for the purpose of blasthole drilling and rock bolt installations. Not only machine parameters but also environmental conditions have significant effects on drilling. This study characterizes the performance of button bits during blasthole drilling in coal measure rocks by using multiple regression analyses. The penetration rate of jumbo and electro-hydraulic drills was measured in the field by employing bits in different diameters and the specific energy of the drilling was calculated at various locations, including highway tunnels and underground roadways of coal mines. Large block samples were collected from each location at which in situ drilling measurements were performed. Then, the effects of rock properties and machine parameters on the drilling performance were examined. Multiple regression models were developed for the prediction of the specific energy of the drilling and the penetration rate. The results revealed that hole area, impact (blow) energy, blows per minute of the piston within the drill, and some rock properties, such as the uniaxial compressive strength (UCS) and the drilling rate index (DRI), influence the drill performance. [ABSTRACT FROM AUTHOR]

This paper analytically investigates the performance of fully grouted rock bolts in typical scenarios, including pullout test, suspending loosened block, and increasing joint aperture, respectively. The interface shear stress distribution follows the model proposed by Li and Stillborg (Int J Rock Mech Mining Sci 36:1013-1029, ), while the axial behavior of the bolt shank obeys the elasto-plastic (yielding-hardening) constitutive model of steel. Three different failure modes are taken into account: tensile failure of bolt shank, bolt shank being pulled out along the bolt/rock interface, and loss of face plate. The evolution of the interface shear stress and the axial tensile stress are examined for both long and short bolts under displacement and load boundary conditions. The derived charts are able to predict the load capacity of fully grouted bolts in pullout test, the minimum length requirement of the bolt to suspend a loosened block, and the maximum allowed opening displacement of a rock joint for a fully grouted bolt. In addition, different potential failure modes are specified. Full range load-displacement curves are produced and compared for various failure modes. The derived charts could be directly used in rock-bolting design. [ABSTRACT FROM AUTHOR]

A combined support scheme was proposed to address the issue of large deformation in the deep soft rock roadway at the Tangkou coal mine in Shanxi Province, China. This support scheme includes high constant resistance, elongation, and prestressed negative Poisson's ratio (NPR) bolts as the core element, along with metal mesh, shotcrete, cable, and bottom angle grouting bolt. A numerical simulation test using Flac3D was conducted to verify the feasibility of the NPR bolt-anchored soft rock roadway. In addition, a constitutive model for a rock mass supported by NPR bolts was established. The model is based on a parallel model of a generalised Kelvin body and elastic–viscous sliding body, which was used to reveal the mechanical behaviour of an NPR bolt-anchored rock mass. The results showed that the stress–strain relationship of the NPR bolt-anchored rock mass was obtained as a piecewise function consisting of the initial, slipping, and sticking stages. The stress and strain of the NPR bolt-anchored rock mass exhibit distinct patterns under different stress conditions. Specifically, when the stress remains constant, the strain of the NPR bolt-anchored rock mass increases continuously while the stress decreases in a step-like manner. On the other hand, when the stress increases linearly with time, the strain of the anchored rock mass increases linearly, while the stress of the anchorage rock mass increases slowly in a sawtooth pattern. Furthermore, a field support test fully confirmed the effective control of the NPR bolt support on the large deformations of soft rock roadways. These findings can guide further theoretical research, engineering design, and field applications of NPR bolt-anchored rock masses. Highlights: A new combined support scheme is proposed using the NPR bolt as the core. The mechanical behaviour of an NPR bolt-anchored rock mass is revealed. Both numerical simulations and field tests demonstrated that the combined support scheme effectively controls the large deformations of soft rock roadways. [ABSTRACT FROM AUTHOR]

With the rapid rise of infrastructure construction in southwestern mountainous areas of China, tunnel engineering has become increasingly frequent in crossing faults with strong activity. However, there has been a long-term lack of an effective theoretical method for identifying the stability and controlling the deformation of extremely weak surrounding rocks inside active faults. Therefore, this paper researched stability analysis and deformation control of a tunnel with extremely weak surrounding rocks caused by strike-slip fault activity. Firstly, uniaxial and triaxial compression tests were conducted on completely decomposed granite within active faults under drained and undrained conditions, and the mechanical properties of granite after desertification and argillization were analyzed. Then, based on the upper-bound limit theory and according to the relationship between the safety factor, instability trace, and geo-stress, a theoretical method to evaluate the stability of the unsupported surrounding rock under different geo-stress levels was proposed. Furthermore, a comprehensive deformation control method, including advanced support, excavation face reinforcement, and radial grouting, was proposed, and the applicability of this method was analyzed and verified through field mechanics and deformation measurements. The results showed that under the triaxial stress state, the post-peak stress drop characteristics of completely decomposed granite basically disappear, and the plastic deformation characteristics of rock become prominent. The comprehensive strengthening scheme has a significant deformation control effect, and the surrounding rock deformation can be reduced by over 65%. In addition, the pressure on the supporting structure will be more than 3.2 times higher with no reinforcement scheme. Highlights A comprehensive deformation control scheme has been proposed to prevent and control deformation disasters in extremely weak surrounding rocks caused by strike-slip fault activity. An upper-bound limit theory considering the influence of geo-stress is established to judge the stability of the unsupported surrounding rock. The mechanical properties of completely decomposed granite caused by strike-slip fault activity under drained and undrained conditions are analyzed. [ABSTRACT FROM AUTHOR]

D-bolt is a type of energy-absorbing rock bolt. It is made of a smooth steel bar with anchors spaced along the bolt length. A typical section between adjacent anchors is approximately 1-m long, but it can be adjusted to adapt to the rock conditions. The bolt is fully encapsulated with either cement or resin grout in a borehole. The anchors are firmly fixed into the grout, while the smooth bolt sections can freely deform to absorb deformation energy. Full-scale static pull tests were carried out at different testing facilities in two laboratories. The tests show that a smooth bolt section between anchors may elongate by 110-167 mm depending on the section length. Field trials of the D-bolt were conducted in deep metal mines. The measurements showed that the D-bolts were equally loaded within every anchor-between section, avoiding load peaks and premature bolt failure due to stress concentrations caused by fracture/joint opening. The field trials of rebar and D-bolts in a largely deformed mine tunnel showed that the D-bolts behaved satisfactorily, with only a few failed bolts, while a number of the rebar bolts failed at the thread. [ABSTRACT FROM AUTHOR]

The article presents field observations on the loading conditions and the failure modes of rock bolts. It shows the shortcomings of the conventional rock bolts and determines the features that rock bolts should possess for rock support in high stress rock masses. It was found that in situ rock bolts are most loaded in the part close to the collar of the borehole. The shortcoming of the conventional fully bonded rebar bolt is cited.

An inflatable rockbolt is a steel tube that has been rolled into an omega shape. During installation, water is injected into the tube, which expands the profile of the rockbolt, generating a radial force against the borehole wall. This reinforces the rock mass through the friction at the bolt–rock interface. This study aims to quantify the effects of borehole size, borehole roughness, and installation water pressure on the pull load capacity of the bolt. This was achieved through a series of pull tests under laboratory conditions. The test results showed that the load capacity was higher in boreholes that were either close to the initial profile diameter of the bolt or the fully unfolded diameter of the bolt, that were 26.8 mm and 38.7 mm, respectively, in the study. The load capacity and the radial stiffness of the bolt were lowest in the medium-sized boreholes (33 and 33.5 mm). In small boreholes, the shoulders of the bolt tongue are tightly compressed such that the outward elastic deformation of the bolt tube is locked in after installation. This deformation locking enhances the load capacity of the bolt. In addition, the load capacity was found to be higher in percussively drilled boreholes than in diamond-drilled boreholes. The additional friction angle of the percussive boreholes was back-calculated to be approximately 5.83°. The load capacity was also found to increase as the installation water pressure increases in the range of pressures tested. It was observed that the inflatable bolt was clamped against the borehole in three zones: on the two tongue shoulders and the side of the bolt opposite the tongue. Highlights: The pull load of an inflatable bolt is minimum as the borehole diameter is in the middle of the initial and fully unfolded diameters of the bolt. In small boreholes, the clamping of the shoulders of the tube tongue results in higher contact stress and consequently higher pull load capacities. Borehole roughness is vital in enhancing the pull load capacity of inflatable bolt. The roughness angle of a percussive borehole is around 5.83°. The pull load capacity of the inflatable bolts increased as the installation water pressure increased from 24 to 30 MPa. An installed inflatable bolt contacts the borehole at the two tongue shoulders and in a zone around the opposite of the tongue. [ABSTRACT FROM AUTHOR]

In this paper, a simplified method for predicting rock mass resistance against tensile forces from rock anchors (anchor) is presented. The interaction of anchors and rock mass was investigated using three-dimensional discontinuous numerical modelling. Several patterns of rock discontinuities were assumed in the numerical modelling while a single anchor is embedded in it. The numerical results show that the existence of a discontinuity set sub-parallel to the anchor significantly improves the rock mass resistance against the tensile force from the anchors. This phenomenon is due to the rock block interlocking at the sub-parallel discontinuity set. Rock block interlocking generates a zone of stress concentration inside the rock mass which has an arch shape (i.e., a pressure arch), resisting against the anchor's tensile force. The load-bearing capacity of the pressure arch plays a significant role in the resistance of the rock mass against the forces from the rock anchor. A voussoir beam analogy was utilised to study the load-bearing capacity of the pressure arch. A simplified analytical approach was developed to assess the load-bearing capacity of the voussoir beams. Then, it was used in combination with the weight of the mobilised rock mass by the anchor to assess the maximum anchoring resistance of the rock mass (anchoring capacity). The suggested method was calibrated by numerical modelling and relevant published pull-out test results. The technique developed in this paper shows the significance of rock block size, shear behaviour of rock discontinuities, Young's modulus of the rock mass, and uniaxial compressive and tensile strength of the intact rock in anchoring capacity of rock masses. Highlights: A theoretical framework is presented on how rock mass and rock anchors interact based on numerical modelling. An analytical method is developed to calculate the anchoring capacity of a rock mass which has at least one joint set sub-parallel to the anchor. The importance of the rock block size and the shear behaviour of rock discontinuities in determining the anchoring capacity of the rock mass is discussed. [ABSTRACT FROM AUTHOR]

The untimely collapse of the lateral cantilever structure is one of the main factors in the deformation of the surrounding rock in hard-roof mines. In this study, a method for weakening the hard roof using dense drilling is proposed to promote roof collapse. Based on the geological engineering conditions in the Tashan coal mine, an FLAC3D numerical model was developed by considering the plastic-softening characteristics of the hard roof to investigate the dynamic evolution process of the weakening roof surface. As the load increased, failure around drillings occurred gradually, and the failure zones expanded in an x-shape from the drillings toward each other, forming a weakened macro-surface. The influence of the drilling arrangement on weakening the hard roof was investigated using numerical models of various drilling diameters and spacings. The results indicated that a large diameter or small spacing was helpful for the formation of a weakened macro-surface. However, differences were observed in the response of key parameters to the weakening effect. The peak stress, fracture stress, and energy absorbed during loading and dissipated during the rock failure process displayed a downward trend for all models with an increasing drilling diameter. In contrast, these parameters, except for the fracture stress, tended to increase with a decreasing spacing. Field monitoring revealed that the average rock strength around dense drillings decreased from 69.0 to 60.5 MPa during the advancing of the work face after roof cutting by implementing dense drillings. Furthermore, the maximum axial tension increments of the cable and bolt reached 80 kN and 10 kN, respectively. Highlights: A roof-cutting method involving dense drillings to weaken hard roofs was proposed. The progressive failure mechanism of the rock around dense drillings was investigated. The influence of the drilling arrangement on weakening hard roofs was investigated. The effectiveness of the roof-cutting method was revealed by conducting a case study. [ABSTRACT FROM AUTHOR]

A new rockbolt for rock support in deep underground engineering is developed. The bolt, which is called 2S-bolt, is composed of a threadbar, a sliding piston, a sliding-resistance cone pipe, an anchorage section, a plate, and a plate-nut. The deformation of the bolt is composed of frictional sliding of the sliding piston in the sliding-resistance cone pipe in the first deformation stage and steel stretching of the threadbar in the second deformation stage. The structure and working principle of the new bolt is introduced first, followed by a laboratory study of the sliding resistances of each component of the bolt. Next, laboratory static pull-out tests of 2S-bolts with different sliding piston lengths are conducted and the results analyzed. The test results show that the bolt with a 30 cm long sliding piston can increase the load from 120 kN in the first deformation stage to 210 kN in the second deformation stage and it has a deformation capacity and an energy absorption capacity of 500 mm and 83.5 kJ, respectively. The average anchor force of the bolt with a 45 cm long sliding piston is 181.7 kN with an energy absorption capacity of 77.8 kJ. When utilized for rock support in rockbursting and squeezing grounds, the newly developed 2S-bolt can absorb a large amount of energy at a higher load capacity. Highlights: A new rockbolt called 2S-bolt is developed for rock support in deep underground engineering. The deformation of the 2S-bolt is composed of two stages: frictional sliding and steel stretching. The 2S-bolt uses a sliding piston and a conical tube to maintain high load resistance during sliding. Laboratory tests show that the 2S-bolt can increase load capacity, have large deformation and energy absorption capacities, and is suitable for use in squeezing and rockbursting grounds. [ABSTRACT FROM AUTHOR]

The shear mechanical properties are the important factors affecting the failure of rock mass. At present, tests measuring the shear mechanical behavior characteristics of rocks under three-dimensional stress are rare. In particular, the influence of lateral stress on the shear failure mechanism of rock is less often considered. The true triaxial test device was used to conduct direct shear, true triaxial stress shear, and confining shear tests under three-dimensional states of stress. Through the three different types of shear tests, the deformation, failure modes, strength, and surface roughness of sandstone were investigated. Focusing on the acoustic emission (AE) characteristics during the failure of sandstone specimens, the influences of lateral stress on the shear failure mechanism of sandstone under true triaxial stress were studied. The test results show that lateral stress can increase the peak strength and residual strength of sandstone, significantly affecting its lateral deformation. Under high normal stress, the weaker the lateral restraint on the rock, the more severe the lateral deformation failure. The analysis found that lateral stress increased the compressive compactness of the mineral particles inside the sandstone and the occlusal hinge strength of the particles. It also restricts the lateral deformation of the sandstone and the germination of the number of internal cracks. Crack propagation is also accelerated under the combined action of high normal stress and lateral stress. Highlights: Direct shear, true triaxial shear and confining shear tests on sandstone specimens were carried out using true triaxial test device. Through the different types of shear tests, the strength, deformation, JRC and failure modes of sandstone were investigated. Focusing on the AE characteristics, the influences of lateral stress on the shear failure mechanism of sandstone were studied. [ABSTRACT FROM AUTHOR]

Impact-induced rockburst hazards pose a severe threat to the construction safety of deep-buried, high-stress tunnels. Negative Poisson's ratio (NPR) cable has the characteristics of strong energy absorption and impact resistance. In order to explore the control effect of NPR cable on impact-induced rockburst, this paper designed and developed a model NPR cable based on the similarity principle. With the help of the self-developed new true triaxial rockburst experimental apparatus, indoor simulation experiments of rockburst under dynamic disturbance at varying strain rates were conducted. The rockburst characteristics under the conditions of NPR cable support were compared to those without support. On rockburst, the influence mechanism of strain rate and NPR support was discussed. The results indicate that the rock strength and rockburst intensity increase as the strain rate of dynamic disturbance increases. The support of NPR cables enhances energy dissipation and rock fragmentation, as well as increases the strength of the rock. Importantly, with NPR support, the energy release rate during rock failure slows, which significantly minimizes the rockburst intensity (kinetic energy release). The research results provide reference and guidance for the engineering practice of rockburst control. Highlights: A cable model with negative Poisson's ratio (NPR) structure effect was developed and its mechanical properties were verified. The experiments of rockburst caused by dynamic disturbance with different strain rates were carried out, and the situation with or without NPR support was compared. With the increase of strain rate of dynamic disturbance, rock strength and rockburst intensity increase. NPR cable improves rock-bearing capacity and effectively reduces rockburst intensity by enhancing fracture and reducing energy release. [ABSTRACT FROM AUTHOR]