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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]

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]

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]

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]

Shear behavior of the bolt–grout interface basically depends on site-specific boundary conditions, i.e., initial normal stress and normal stiffness of the borehole wall. Few studies have been conducted to investigate the shear behaviors between the bolt and grout material under different boundary conditions. To better understand the effect of boundary conditions on the shear behaviors of the bolt–grout interface, simplified two-dimensional (2D) bolt–grout interface specimens are prepared and tested using direct shear tests at various initial normal stresses and normal stiffnesses in this context. The testing results showed that both initial normal stress and normal stiffness significantly influence the bolt–grout interface shearing behaviors. Increasing the normal stiffness or initial normal stress would increase the peak and residual shear strength. However, the degree of the brittleness after the peak, normal displacement, and peak friction coefficient was reduced. It is noted that the peak and residual shear strength points under the constant normal stiffness conditions located near the peak and residual strength envelopes for the constant normal load tests, respectively. Besides, the shear failure process of the bolt–grout interface was captured by PAC acoustic emission (AE) monitoring and digital camera technology, a good correlation between the evolution of AE parameters and the shear stress curves was obtained. [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]

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]

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]

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]

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]

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]

Bolt installation quality is influenced by various factors, some are well known and others are less recognised. A programme of field and laboratory studies was undertaken to examine various factors of relevance to the load transfer mechanism between the bolt, resin and rock to ensure test methods truly represent field performance. Short encapsulation tests were undertaken as part of the Australian Coal Association Research Program (ACARP) funded project (C21011) with the ultimate aim of developing standard test methods for assessing bolt encapsulation with chemical resin anchor installations. The field study consisted of a series of Short Encapsulation Pull Tests (SEPT) carried out in three mines with different geological conditions to determine the most representative and practical method of SEPT. Additional field work included installation of bolts into threaded steel tubes for subsequent removal and laboratory evaluation. A series of pull tests was carried out by installing bolts in overhead rig mounted sandstone block, cast in concrete with controlled encapsulation length. Factors of importance considered included; borehole diameter, resin annulus thickness, installation time (including bolt spin to the back and 'spin at back'), the effect of gloving and hole over drill. It was found that the borehole diameter had a detrimental effect on the encapsulation bonding strength. Bolt installation time of approximately 10 s constituted an acceptable time for effective bolt installation and within the resin manufacturers recommended time of 14 s. Maintaining constant length of encapsulation was paramount for obtaining consistency and repeatability of the test results. Finally, a numerical simulation study was carried out to assess the capabilities of FLAC 2D software in simulating the pull testing of rock bolts. [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]

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]

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]

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]

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]

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]