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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.

Proper design of rock bolt support in underground mines is critical to avoid incidents, accidents and loss of production. The traditional design approach only considers the axial (tensile) capacity and this is clearly not the situation in situ, where a rock bolt is subjected to both axial and shear/bending loads which determines its overall performance and failure behaviour. To demonstrate and analyse the shear displacement in bedded roof, scaled physical models of underground excavation were created. From the models it was found that the shear displacement between the layers depends on the vertical roof deformation and thickness of beds. To analyse the effect of combined loading on rock bolt design for suspension and beam building models, analytical methods were used to calculate the required spacing of rock bolt for a given safety factor. Numerical models were then created using Rocscience RS2 software to establish the stresses on the rock bolt. The results show a significant reduction in safety factor for suspension as demonstrated in an example (reduced from 3.5 to 2.0) and beam building (2.0 to 1.36) when the rock bolt capacities are calculated considering the effect of combined loading as opposed to just the axial or shear loads.

Rock bolts are subjected to different loading conditions along their lengths such as axial, bending, and/or shear forces, which can cause failure at lower loads than those considered for design purposes. The common existing methodologies do not consider the actual loading of the rock bolts and assume it is only pure axial or pure shear. This study was conducted to investigate the un-grouted rock bolt performance under combined load conditions. Two loading regimes were evaluated: the effect of initial shear displacement on axial load capacity and displacement, and the effect of axial displacement on the shear load capacity. The first regime was also conducted for shear with a gap, when there is a spacing between the shear interfaces. The results of this study showed that the rock bolt can resist higher axial loads than shear under pure or combined load conditions. Under combined load conditions, the rock bolt capacity decreased significantly for both regimes. However, when applying the shear load with a gap, the rock bolt load capacity was not affected significantly. Also, the total bar deformation was improved for shear and axial. The findings of this study show the need to improve the rock bolt design considering the complex loading conditions in situ with/without a gap. Keywords: Rock bolt, Combined load, Gap test, Failure mechanism, Axial load, Shear load

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.

Underground construction has become a necessity considering social, technological, economical, and sustainable advancements in society. Tunnel boring machines (TBMs) can be considered as an efficient method of tunneling because of its high productivity and applicability in a variety of subsurface conditions. The performance of a rock TBM is assessed by its advance rate that is a function of penetration rate and utilization factor. While there are various existing models that can reasonably predict the penetration rate, only few models have the capability to estimate the utilization, and the accuracy of models predicting the utilization is highly dependent on the incorporation of complexities involved in different tunneling activities. The geology-related delays include those incurred for ground stabilization, including installation of rock bolt, wiremesh, strap, or channel, placing of ribs and lagging, gripping process, control of water inflow, cutter change and inspection, probe drilling, scaling, muck removal, and foam or grout injection for ground stabilization. This study aims at determining the most critical ground support activities affecting the TBM performance using discrete event simulation. For this purpose, six projects that utilized open TBMs were analyzed. Comparing the individual ground support measure installations, steel ribs showed the maximum downtime of 7% compared to water inflow at 4.8% and cutter change at 2%. The results show that geology-related delays indeed contribute significantly to the variations in utilization factor.

Coal auger mining is a promising technology used for excavating thin coal seams. The efficiency of auger mining is largely related to the stability of mine roadways in the influence zone of the coal-face. Roof bolting systems are promising in such conditions. An adequate choice of roof bolting parameters is only possible if one understands the features of the stratification of the rocks and stages of deformation of the array in auger mining. Modern monitoring methods of the condition of rocks are based on the use of mechanical benchmarks, sounding of the mine array and the use of optical devices. There are few studies concerning roadways with auger mining. The innovations presented in this manuscript are a determination of the research results of the in-situ processes of rock deformation around a roadway in auger mining which will help to better understand the features of deformation processes in the technological method and design an adequate support system. Some field studies were undertaken in order to investigate the geo-mechanical processes that can be observed while auger mining a roadway with fully grouted bolts of 2.4 m in length. The research included monitoring rock stratification with the help of mechanical telltales, the convergence in the roadway using contour benchmarks, measurements of altitude and rock falls, and visual observations. The presented results show that roof-bolting can be used to support the roadways for auger mining., Rudarenje ugljena svrdlima (augerima), tj. strojevima za iskopavanje ugljena bušenjem, tehnologija je primjerena za vađenje tankih ugljenih slojeva, a učinkovitost uvelike ovisi o stabilnosti rudničkih okana na ugljenome čelu, što se rješava podgradama. Prikladan odabir varijabli podgrade moguć je samo kada se razumiju svojstva slojeva u toj krovini te njihova podložnost deformacijama kod bušenja svrdlima. Suvremene metode praćenja svojstava stijena temelje se na mehaničkim provjerama, sondiranju i optičkome ispitivanju. Postoji nekoliko studija takvih uvjeta, a ovdje su dane inovacije kod opažanja deformacija in situ, tj. u oknima u kojima se buši. Tako se bolje opisao proces deformiranja te tehnologija izgradnje adekvatnih podgrada. Odabrano je nekoliko terenskih studija kojima su izučena geomehanička svojstva stijena tijekom bušenja uz uporabu injektiranja u dužini od 2,4 m. Praćenjem uslojenosti i pojava mehaničkih nedostataka u oknima kartiranjem, mjerenjem visine i vizualnim opažanjem uočavani su nedostatci. Rezultati su pokazali kako je podgrađivanje sidrenjem poželjno.

The instability of rock slopes and underground engineering is usually caused by the shear failure of non-persistent joints. Rock bolt, as the main support method, has an important influence on the shear behaviour of joints. However, the shear behaviour of bolted non-persistent joints has rarely been studied. In this research study, direct shear tests were carried out to explore the effect of different bolting schemes on shear behaviour of non-persistent joints. The test results showed that fully encapsulated bolts can greatly improve the shear strength of non-persistent joints, while end encapsulated bolts behave better at a larger shear displacement. The bolt breakage is closely related to the shear failure of non-persistent joints. The bolt passing through the coplanar coalescence areas usually has a larger shear deformation and breaks first. The shear displacement of bolt breakage tends to be smaller as the joint persistency decreases. Furthermore, acoustic emission tests showed that the fully encapsulated bolt noticeably reduced the shear damage, while it increased the shear failure intensity. In addition, the bolting effect tends to be smaller under a larger joint persistency.

In roadways with high ground stress or burial depths, the joints distributed within rock formations are subject to complex stresses and interlayer misalignments frequently. Rock bolts and cable bolts anchored in the rock formations are subject to tensile and shear forces. Most of the bolts used in roadway engineering are local anchored, resulting in insufficient shear strength at the bolt free end close to roadway surface and increasing bolts breaking. The anchor cable and C-shaped tube (ACC) is a highly prestressed cable bolt that can withstand high shear force in its free end. This paper examines the effect of the relationship between C-shaped tube length and joint location on the shear resistance of ACC by double shear tests. To fully exploit the ACC’s shear resistance, the C-shaped tube ends should be at least 30 cm beyond the joint. The effect of preload and concrete spray thicknesses on roadway deformation and plastic zone is investigated by numerical simulation. Results show that ACC and concrete spraying layer can form a stable extruded arch structure, so that the broken and soft rock within the loosen zone is in three-dimensional-stress state, effectively improving surrounding rock properties and controlling its deformation size. Based on these results, the ACC support design method is proposed.

Since the failure modes of fully grouted bolts subject to displacements of rock joints are difficult to predict, the joint displacements in this study are categorised into two types: opening displacement and shear displacement. At first, a closed-form solution incorporating the nonlinear constitutive model of interfaces between bolts and grout is developed, which is able to formulate the full-range behaviour of bolts subject to the opening displacements of rock joints. Then, a novel numerical approach considering the nonlinear rock reaction force is proposed, which can analyse stress and displacement conditions on each point along a bolt. The results of the proposed numerical approach are validated by existing test results. Combining the proposed numerical approach with the improved closed-form solution, a comprehensive model for predicting yield and failure patterns of rock bolts is proposed. In this way, decoupling between the bolt and grout, the yield and failure patterns of two potential points along the bolt are also highlighted and discussed. Further discussions about the influences of bolt inclined angles to the discontinuity and the dilatancy effect of the joint face are carried out at last. This work provides a useful tool to evaluate the behaviours and estimate the load-bearing capacity of fully grouted bolts in jointed rock masses.

In this study, a new spring plate was designed and investigated for its usability with the split set type friction bolts. According to the results of this study, remarkable active support pressures can be practically supplied by compression of the new spring plate during the insertion of the split sets. In addition to easy supply of the active support pressure owing to the insertion of the bolts, the energy absorption capacity can be highly improved by use of the new spring plate. Furthermore, the new plate design is remarkably advantageous against the nonaxial loading condition which is very wide in the bolted rock masses.

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

Rock bolt is a support system extensively utilized in Civil and Mining Engineering applications, especially for underground excavations projects. The main function of a rock bolt is to stabilize the rock mass around the opening of an excavation by fastening to further stable formations. Previous technical work developed pull-out testing in order to improve understanding regarding the mechanical response within this system. However, due to limitations with conventional laboratory methods of capturing strain, there still exists a lack of understanding in this mechanical response at the micro-scale; more specifically, in the detailed strain profile distribution along fully grouted rock bolts. Aiming to address this knowledge gap, two research programmes were conducted and will be addressed in this paper: numerical modelling of pull-out tests through finite element method (FEM) and hybrid finite-discrete element method (FDEM), and laboratory pull-out tests on concrete samples utilizing a newly developed distributed optical strain sensing technology that provides a spatial resolution of 0.65 mm in order to capture the strain along the rock bolt. Experimental and modelling results were compared to evaluate the numerical approaches feasibility and their correspondence with the experiments. FDEM modelling demonstrated ability in capturing general mechanical behavior, while the FEM software seems to be a good option to be utilized as a design tool, yet improvements to the interaction model are necessary to achieve a better calibration with experimental results.

Rock bolts are widely used in the main structure of the DC transit system and play the role of the main body reinforcement and support. Because of the extensive stray current in the subway system, the corrosion risk of rock bolts is greatly increased. Therefore, the rock bolt was experimentally studied by simulating the input stray current and monitoring the polarization potential in this paper. After the stray current corrosion (SCC) process was finished, a tensile and breaking experiment of the corroded specimens was conducted for the mechanical properties. As an important parameter to indirectly characterize the corrosion strength of the stray current, the polarization potential was analyzed. Two types of rock bolts were compared for their mechanical properties and corrosion parameters. The result shows that the rock bolt with stranded steel wire as the rod has better corrosion resistance than the HRB500 rebar. Hence, the bonded stranded steel wire is more resistant to the stray current corrosion than the HRB500 rebar, which is adaptive for rock bolts in the direct current (DC) transit system.

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.

For coal mines, rock, coal, and rock bolt are the critical constituent materials for surrounding rock in the underground engineering. The stability of the “rock-coal-bolt” (RCB) composite system is affected by the structure and fracture of the coal-rock mass. More rock bolts installed on the rock, more complex condition of the engineering stress environment will be (tensile-shear composite stress is principal). In this paper, experimental analysis and theoretical verification were performed on the RCB composite system with different angles. The results revealed that the failure of the rock-coal (RC) composite specimen was caused by tensile and shear cracks. After anchoring, the reinforcement body formed inside the composite system limits the area where the crack could occur in the specimen. Specifically, shearing damage occurred only around the bolt, and the stress-strain curve presented a better post-peak mechanical property. The mechanical mechanism of the bolt under the combined action of tension and shear stress was analyzed. Additionally, a rock-coal-bolt tensile-shear mechanical (RCBTSM) model was established. The relationship (similar to the exponential function) between the bolt tensile-shear stress and the angle was obtained. Moreover, the influences of the dilatancy angle and bolt diameter of the RCB composite system were also considered and analyzed. Most of the bolts are subjected to the tensile-shearing action in the post-peak stage. The implications of these results for engineering practice indicated that the bolts of the RCB composite system should be prevented from entering the limit shearing state early.

The seepage force is of importance on tunnel face stability. The formulas to determine the seepage force are introduced based on Darcy's Law. The formula of the stability safety factor containing the tunnel face of the seepage force is derived and applied to the actual tunnel engineering using the upper bound limit analysis combining with the strength reduction method. The results show that: as the water level increases, the average seepage force increases, and the safety factor decreases. If the horizontal seepage force is considered only and the vertical seepage force is ignored, the result is unsafe. The stability of the tunnel face can be improved by reducing the excavation height and adopting some reinforcement, such as grouting or rock bolt to increase the cohesion and friction angle of the rock.

This paper proposes a particulate interface model to simulate the interface by using a particulate discrete element method. Six modifications are adopted in the numerical calculations to improve the stability and accuracy. The interface model provides two failure criteria: one is the Mohr–Coulomb failure criterion and the other is the Barton–Bandis shear strength criterion. To verify the proposed model, the failure envelope, the shear-displacement curve, the closure curve, and the dilation curve were analyzed and simulated. The simulations demonstrated good agreement with the theoretical curves. The influence of the particle size was also investigated; the results proved that the quantity or size of the particle did not affect the performance of the proposed model. Finally, a rock bolt pull-out problem was simulated. The results proved that the proposed interface model can reasonably predict the shear stress acting on the interface between the rock bolt and rock.

Underground surge pool cavern for the storage and lifting of water is being constructed in the part of Telangana State of India. For better rock mass characterization and support design, 3D engineering geological mapping was carried for the heading portion of the surge pool. 3D geologic mapping of heading portion is very important for large underground cavern to know about the cavern behavior during benching down. The direction of the longest axis of the cavern was finalized based on the in-situ hydrofracturing testing inside the deep borehole. The assessment of tunnel quality index ‘Q’ for the exposed granitic rock mass was done based on the information available of the rock joints and their nature, 3D geological mapping and in-situ stress measurement. Roof support and wall support pressure was estimatedbased on Q-values, and in this case applicable for the non-squeezing ground condition. For structural stability of crown and walls, rock support arrangement was recommended based on NMT technology and geological conditions of site, which includes steel fibre reinforced shotcrete, grouting, rock bolt and drainage hole provisions. The capacity of support system is determined to evaluate the efficacy of the proposed support system.

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.

This article presents the methodology and results of single shear tests of bolt rods under dynamic impact loading generated by means of a drop hammer. Comparative analysis was also performed for bolt rod load capacity, stress and shear work under static and dynamic (impact) loading. The developed method of single shear testing of bolt rods under impact loading makes it possible to obtain repeatable test results concerning maximum bolt rod shearing force, shear stress and shear work values. Comparative shear tests of four types of bolt rods under static and impact loading showed that the APB-type bolt rods made of AP770 steel, which was characterised by having the highest strength, exhibited the greatest shear work. AM22-type bolt rods exhibited a very similar work value. Though the AM22-type bolt rods made of A500sh steel demonstrated lower strength than the APB-type bolts, as well as a smaller diameter and cross section, they dissipated the impact energy better thanks to their higher plasticity. This could indicate the direction of optimisation for bolt rods in order to increase their impact strength. Mathematical relationships were also formulated for selected tests, describing the real single shear courses F d =f(t) of bolts under impact loading. The obtained relationships could be applied in the load assessment process of bolt rods intended for use under roof caving, tremor and rock burst conditions.

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.

Developing an improved understanding of the interaction between the immediate roof and the rockbolts is fundamental in obtaining an effective roof support system. Effective instrumentation along the full length of fully grouted rock bolts would provide the required understanding, which in turn could improve the rock related safety in the underground mines. The rock bolt sensing technology has rapidly developed in the last decade, where the instrumentation has changed from short electrical strain gauges to optical strain and shape sensing using more than 2 slots (which used to be the standard). In addition, the instrumentation has been increased from two orthogonal slots to three or four. With these recent advances of the technologies, there lies a challenge in analysing and understanding the new type of data and in effectively adopting the new technology into the underground mines. This paper deals with the challenges of instrumenting rock bolts with fibre optics and presents techniques for improving the existing technology. Furthermore, it includes a prototype design for an effective implementation of this technology to better suit the harsh underground environment.

Rock bolts are widely used in rock engineering projects to improve the shear capacity of the jointed rock mass. The bolt inclination angle with respect to the shear plane has a remarkable influence on the bolting performance. In this study, a new artificial molding method based on 3D scanning and printing technology was first proposed to prepare bolted joints with an inclined bolt. Then, the effects of the bolt inclination angle and boundary conditions on the shear behavior and failure characteristic of bolted joints were addressed by conducting direct shear tests under both CNL and CNS conditions. Results indicated that rock bolt could significantly improve the shear behavior of rock joints, especially in the post-yield deformation region. With the increase of bolt inclination angle, both the maximum shear stress and the maximum friction coefficient increased first and then decreased, while the maximum normal displacement decreased monotonously. Compared with CNL conditions, the maximum shear stress was larger, whereas the maximum normal displacement and friction coefficient were smaller under the CNS conditions. Furthermore, more asperity damage was observed under the CNS conditions due to the increased normal stress on the shear plane.

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.

The shear behaviours of rock joints with and without rock bolt support are numerically studied using the discrete element code PFC2D. A cohesive contact model was employed to reproduce the damage response of the synthetic intact rock (i.e. asperity degradation). We used the smooth-joint model to simulate the micro-scale roughness of the joint surface. We calibrated and validated the proposed numerical framework against the experimental results. A series of numerical constant normal stiffness (CNS) direct shear tests were conducted on idealised and natural rock joints to investigate the influence of the boundary condition on the shear behaviour of rock joints. In particular, the importance of CNS condition, surface roughness, asperity angle, and the initial normal stress were studied. Additionally, the shear and damage mechanism of bolted rock joints under constant normal load (CNL) and CNS condition was numerically investigated. The results presented show that the shear resistance of the joint increases under both CNL and CNS conditions, but at a high degree of roughness, no significant enhancement was observed on the value of peak shear stress.

The support of rock bolt plays a significant role in maintaining the stability of a roadway in steeply inclined geological formations, yet choosing a reasonable support scheme of rock bolt in this situation is still not well understood. To reveal the force or stress distribution mechanism of anchor bolts and design an effective support scheme, an in situ anchorage experiment was conducted in a deep-buried roadway in steeply inclined rock strata. This results of this study demonstrate that in the steeply inclined rock formations, the intersection angle of a bolt length direction and the interface of rock layers have a great effect on the distribution and variation of axial force in bolts. The axial forces at the positions near the intersection of the interface and bolt increase a lot with time. In addition, the existence of no small or non-zero axial force at the end of anchor bolts indicates that much longer bolts should be used to prevent further deformation in steeply inclined rock.

The stability of longwall coal mines has been a challenging concern for geotechnical and mining engineers. Rock-bolting is the commonly used support system preventing roof and wall convergence and floor heave. This study attempts to suggest an optimized new rock-bolting pattern to effectively prevent the convergence of transportation and ventilation tunnels in Tabas coal longwall mine at panels 2 and 3. Three-dimensional numerical analysis was performed on the support system using FLAC3d code. Investigating the rock bolts axial force applied to the ventilation tunnels roofs shows that, the safety factor of the new support system compared to the old one can be reduced by the ratio of 1.2 to 2, while the amount of the roof displacements are still within the allowable limit. Moreover, rock bolts installed in the transportation tunnels walls have yielded about 10 m from the working face, which became stable in the redesigned rock-bolting pattern. In order to decrease the floor heave, a support system with inclined bolts was proposed, in which the amount of floor heave in the transportation and ventilation tunnels decreased about 40% and 33%, respectively.

Development of durable anchorage systems for slope support in high-cold and high-altitude areas has been becoming a global research hotspot. This paper is addressed to the deterioration mechanism of bond and anchorage performance of grouted rock bolts under the action of freeze–thaw cycles. A laboratory-based physical simulation was employed to study the anchoring force, displacement, anchor stress, and bond strength between bolts and cement mortar. It was shown that the anchoring force decreased by 27% after 30 freeze–thaw cycles by means of the pull-out tests and microstructural characterization. It was found that the maximum anchor stress existed in the middle part of the anchorage, where the shearing failure was most likely to happen. Using bond-slip curves, four stages of bolt slippage were established. The analysis on the bolt deformation indicated that shearing failure was caused by a decrease of strength and stiffness of the cement mortar, which led to a loss of bonding ability. A degradation model for anchoring limit was established and the reliability of this model was ascertained. This may serve as a reference for future application of grouted rock bolts designed to be used in low-temperature conditions.

Load-bearing structural elements such as rock bolts and reinforcing bars (rebar) are ubiquitously present in both mining industries and in infrastructure projects, for securing tunnel walls and ceilings. Being able to detect and localize defects in previously installed rock bolts, or even quantitatively estimate the load they are taking, would be very valuable when planning maintenance and service. Unfortunately, there are no such techniques available today. Previous work shows that ultrasound, being a mechanical wave, is sensitive to mechanical changes in rock bolts. The ultrasound signature is, however, rather complex, and traditional modeling of the wave propagation rapidly becomes challenging. In this paper, we show that the ultrasound signature measured from a 3.2 m long dynamic rock bolt can be used to accurately determine the load applied to the bolt and the resulting elongation of the bolt. The method is based on training a Partial Least Squares Regression (PLSR) model to estimate the force an elongation from power spectra of backscattered ultrasound signatures.

Sandstone-roofed roadways are susceptible to deformation and failure caused by reservoir-water-induced disturbances, thereby compromising human safety. Using rock-mechanics testing techniques, numerical simulations, and engineering principles, this study investigates the strength, deformation, and pore-structure characteristics of sandstone roofs as well as means to support the surrounding rock structure. The results obtained in this study reveal that the residual strain is proportional to the pore-water pressure, which, in turn, causes a significant reduction in the elastic modulus during the unloading phase. Furthermore, an increase in the pore-water pressure causes the shear failure of specimens in compression. The delay between crack initiation and specimen-volume expansion decreases. Moreover, the specimen demonstrates increased deformation and failure responses to changes in the confining pressure, thereby resulting in accelerated conversion. Changes in water inflow can be correlated to crack initiation, propagation, and fracture. This water inflow gradually increases with an increase in the osmotic pressure. Correspondingly, the volumetric strain required for maximum water inflow undergoes a gradual decrease. The increased water inflow can be considered a precursor to specimen failure. In addition, fractures in the surrounding rock structures are mainly caused by joint dislocations. The increase in pore pressure promotes the development of dislocation fractures in the deep surrounding rocks. Subsequently, these fractures overlap with their open counterparts to form large fractures; this increases the roadway-roof subsidence and layer separation of the shallow surrounding rocks, thereby further increasing the fracture count. Lastly, the use of high-performance rock bolts, cable-bolt reinforcements, and W-shaped steel bands is expected to ensure the stability of rocks surrounding sandstone-roofed roadways subject to water-pressure disturbances.

The use of unlined/shotcrete lined pressure tunnels and shafts are cost-efective solutions for a hydropower project and are being implemented worldwide. To implement this concept, the ground conditions at the area of concern should be favorable regarding minimum principal stress magnitude, which should be higher than hydrostatic water head acting on the tunnel periphery. In addition, the rock mass should be relatively unjointed or joints in the rock mass should be relatively tight. Among the most important issues in the design of unlined/shotcrete lined pressure tunnels is the extent of hydraulic jacking and water leakage out of the tunnel during operation. This manuscript first presents fluid flow and potential hydraulic jacking assessment of two selected locations of the headrace tunnel of Upper Tamakoshi Hydroelectric Project (UTHP) in Nepal using the UDEC. It is noted here that the 7960 m long headrace tunnel will experience a hydrostatic water head that will vary from 2.9 to 11.5 bars (0.29–1.15 MPa). The headrace tunnel is supported by sprayed concrete (shotcrete) in combination with systematic rock bolts in the tunnel walls and crown. The invert of the tunnel and few hundred meters downstream end (at surge shaft area) of the headrace tunnel is being concrete lined after the completion of all other works. The qualitative fluid flow assessment carried out using UDEC indicated considerable pressure built-up in the joint systems suggesting potential hydraulic jacking. This was especially the case at the downstream segment (downstream from chainage 7100 m) of the headrace tunnel. The manuscript further presents the quantitative results of water leakage estimation from the headrace tunnel carried out using Panthi (Panthi KK (2006) Analysis of engineering geological uncertainties related to tunneling in Himalayan rock mass conditions. PhD Thesis, NTNU, Trondheim, Norway;Panthi, Note on estimating specific leakage using Panthi’s approach, NTNU, Trondheim, 2010;) approach. The leakage assessment carried out indicated an average specific leakage of about 2.5 l/min/m tunnel, which may result in over 210 l/s leakage from the headrace tunnel. The evaluation also indicated that the outer reach (860 m downstream segment) of the headrace tunnel after chainage 7100 m seems extremely vulnerable and over 80 l/s water leakage may occur only from this headrace tunnel segment during operation of the hydropower plant. This article is licensed under a Creative Commons Attribution 4.0 International License, which permits use, sharing, adaptation, distribution and reproduction in any medium or format, as long as you give appropriate credit to the original author(s) and the source, provide a link to the Creative Commons licence, and indicate if changes were made. The images or other third party material in this article are included in the article’s Creative Commons licence, unless indicated otherwise in a credit line to the material. If material is not included in the article’s Creative Commons licence and your intended use is not permitted by statutory regulation or exceeds the permitted use, you will need to obtain permission directly from the copyright holder. To view a copy of this licence, visit http://creativecommons.org/licenses/by/4.0

Rock bolts are one of the main measures used to reinforce unstable blocks in a rock mass. The embedment length of fully grouted bolts in the stable and competent rock stratum behind the unstable rock blocks is an important parameter in determining overall bolt length. It is required that the bolt section in the stable stratum must be longer than the critical embedment length to ensure the bolt will not slip when loaded. Several series of pull tests were carried out on fully grouted rebar bolts to evaluate the pull-out mechanics of the bolts. Bolt specimens with different embedment lengths and water/cement ratios were installed in either a concrete block of one cubic meter or in steel cylinders. Load displacement was recorded during testing. For some of the bolts loaded beyond the yield load, permanent plastic steel deformation was also recorded. Based on the test results, three types of failure mechanisms were identified, corresponding to three loading conditions: (1) pull-out below the yield strength of the bolt steel; (2) pull-out between the yield and ultimate loads, that is, during strain hardening of the steel; and (3) steel failure at the ultimate load. For failure mechanisms 2 and 3, it was found that the critical embedment length of the bolt included three components: an elastic deformation length, a plastic deformation length and a completely debonded length due to the formation of a failure cone at the borehole collar.

Coal and gas outburst is an important risk faced by coal seam mining in the Huainan region of China. In order to control gas outburst, the gas is predrained by digging a floor gas drainage roadway. To study deformation due to dynamic pressure, the failure characteristics of the floor, and their effect on the stability of the floor gas drainage roadway, a comprehensive monitoring method combining Brillouin optical time-domain reflectometry- (BOTDR-) distributed fiber optics and self-potential exploration was adopted. Dynamic data monitoring of the rock strata between the 11123 working face floor and the floor gas drainage roadway of a mine in Huainan was carried out. The field data obtained showed that, when stabilized by rock bolts and other fixed components in the surrounding rock mass of the floor gas drainage roadway, under the influence of mining, the area of concentrated stress appeared at a depth of 20.7 m, when cracks eventually formed, but the overall structural stability of the surrounding rock mass remained good. The stress distribution and crack evolution of the bottom plate under the influence of dynamic pressure showed spatiotemporal characteristics. Of these, the effect of the lead support stress was 107.48 m, and the range of effect of the hysteresis stress was 34.42 m. When the working face mining position arrives and is far from the monitoring station, the failure depth of floor rock stratum shows the following rule: unchanged in the early stage, deepened continuously in the middle stage, and finally remained stable. It takes about eight days for the dynamic adjustment of this process to finally stabilize. The results of this study can provide guidance for devising suitable procedures for carrying out intelligent green safety mining and for warning about the hazards of roadway damage.

Currently, the application of coal pillars to protect an adjacent roadway is a common method in Vietnam when exploiting according to the longwall system. Therefore, the width of a coal pillar is an important issue for the stability of a roadway. In order to reduce coal loss in these coal pillars, they tend to be designed in a narrow coal pillar style but still have to ensure that the adjacent roadway can meet safe coal production conditions. The stability of roadways and coal pillars is related to many factors such as technical mechanical characteristics, physical and mechanical properties of coal, stress environment and support methods. The bearing structure of the coal pillar and the around rock a roadway is analyzed and it has been shown that enhancing roadway support and improving the carrying capacity of coal pillars can control the deformation of the surrounding rock. A study related to the stability and safety of roadways and small coal pillars in the longwall mining system has been carried out. Stabilization factors have been considered, especially the state of stress in the coal pillars and the deformation of the roadway. By applying the numerical simulation method, the stress of the coal pillar and the deformation of the adjacent roadway under different supporting solutions were analyzed and evaluated. By using this method, the rock bolt roadway support solution combined with the long cable bolt in the roadway roof and the coal pillar was selected in the safe condition of the mining process. Because cable bolt can improve the flexibility of the coal pillar such as: reducing the size of the plastic area on both sides of the pillar; enhancing coal pillar stability in the core area by providing great drag and tensile for coal pillars; contributing to improving the anchor point fixation of rock bolt. The conclusions obtained may provide a certain reference parameters to improve mining efficiency and labor safety in underground coal mines.

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.

A distributed optical strain-sensing technique is presented as a solution for measuring the strain distribution along ground support members used in tunnelling and mining works. The technique employs a Rayleigh optical frequency domain reflectometry technology, which measures strain at a spatial resolution of 0.65 mm along the length of a standard optical fiber. A rationale for selecting this technology as a potential monitoring technique for ground support elements over alternative commercially available technologies is discussed. The development of a technique to couple optical fiber sensors with rock bolt, umbrella arch, and cable bolt support members is also demonstrated. A robust laboratory investigation of such optically instrumented support members demonstrated the capability of the technique to capture the expected in situ support behaviour in the form of coaxial, lateral, and shear loading arrangements as would be anticipated in the field. Moreover, the micro-scale data obtained by this optical sensing technique are shown to provide unprecedented insight into the local/micro-scale geomechanistic complexities associated with the bearing capacity of ground support members, especially when compared with data obtained by discrete strain-sensing technologies.

In mining operations, the stability of excavations is critical to ensure the safety of production. Rock bolt support is the most effective and economical method commonly used to improve ground stability. The glass-fiber-reinforced plastic (GFRP) bolt has been developed and is becoming widely used because of its corrosion resistant character and its high tensile strength. However GFRP bolts are more susceptible to damage by blast loads and therefore it is important to understand the mechanical behaviour of GFRP bolts under dynamic loads so the support system can be better designed. In this paper we propose a new approach to describe the mechanical behaviour of fully grouted GFRP rock bolts under the combination of pre-stressed static load and blast dynamic load using the Kelvin’s solution. The method was applied to a case study and it was found that large axial and shear stresses of fully grouted GFRP rock bolts are mainly distributed in a small region near the bolt collar. It was further demonstrated that the strength of blasting, the distance between the rock bolt and the blasting point and the Young’s modulus of the rock mass and the bolt material are the most important factors having significant influences on the mechanical behaviour of GFRP bolts. For the cases examined where the bolt is within the range of 15 m from the blasting point, the influence of blast load on the axial and shear stresses of the rock bolt is significant and the dynamic load is the main factor causing the support damage. The proposed model provides a method to quantitatively evaluate the effect of blasting on the mechanical behavior of GFRP rock bolts so that a better ground support system can be designed in practical mining applications.

Tectonic shear belts, with their low mechanical strengths, large spatial extents and obvious shearing damage, present challenges for the excavation of high slopes. In this study, we described the basic unloading performances of shear belts induced by excavation through observational data and numerical back analysis. The unloading deformation and failure processes of deep outward-dipping shear belts exposed on a hydraulic slope nearly 300 m tall were investigated first. The field observation and monitoring data indicated that the excavation-induced deformation pattern of the outward-dipping shear belt can be divided into two stages: primarily opening deformation normal to the shear belt during the first stage and primarily shear deformation along the shear belt during the second stage, manifested as opening and shear cracks along the active shear belt. Then, the numerical and analytical analysis were performed to expose the displacement evolution of the shear belt during the multistep excavation of the slope from the top down. The back analysis showed that the grading deformation mechanism of the shear belt was primarily related to the reduction of normal compressive stress on the shear belt's plane due to slope excavation. The reinforcement schemes, including a prestressed anchor cable to limit deep unloading deformation and a prestressed rock bolt to restrain the surface relaxation of the rock mass of the slope, were also discussed. In practice, actual excavation and corresponding monitoring data indicated that the unloading mechanism and supporting measures for the shear belt presented in this study were reasonable and effective. These presented experiences can also enrich the stability analysis and support design for similar high slopes.

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

Geotechnical properties of red shale encountered in deep underground mining were characterized on both laboratory and field scale to reveal its unfavorably in geoenvironment. Its constituents, microstructure, strength properties and water-weakening properties were investigated. In situ stress environment and mining-induced fractured damage zone after excavation were studied to reveal the instability mechanism. The results show that red shale contains swelling and loose clayey minerals as interstitial filling material, producing low shear strength of microstructure and making it vulnerable to water. Macroscopically, a U-shaped curve of uniaxial compressive strength (UCS) exists with the increase of the angle between macro weakness plane and the horizon. However, its tensile strength reduced monotonically with this angle. While immersed in water for 72 h, its UCS reduced by 91.9% comparing to the natural state. Field sonic tests reveal that an asymmetrical geometrical profile of fractured damage zone of gateroad was identified due to geological bedding plane and detailed gateroad layout with regards to the direction of major principle stress. Therefore, red shale is a kind of engineering soft rock. For ground control in underground mining or similar applications, water inflow within several hours of excavation must strictly be prevented and energy adsorbing rock bolt is recommended, especially in large deformation part of gateroad.

Traditionally steel has been used for rock bolts in tunnelling. Faced with the challenge of reducing mankind's environmental impact, it is necessary to look at using better alternatives. Glass-fibre-reinforced polymer (GFRP) has been used for many years in tunnelling applications, primarily for temporary applications. GFRP has many advantages such as its excellent durability characteristics, its light weight, which reduces manual handling impacts on workers in the tunnel, and its lower environmental impact. One obstacle to the more widespread adoption of GFRP for permanent applications has been a concern over its performance in shear. It is well understood that during shearing the rock bolts actually deform under bending and the tensile strength of the bolt governs the behaviour. This paper will describe a new design method to account for the benefit of the bolts in resisting shearing on discontinuity planes. This method can be used in numerical models to demonstrate that GFRP bolts function as well as steel bolts in terms of resisting both tensile and shear loading from the rock mass. This opens the path for the wider use of GFRP rock bolts, which offer superior durability and a lower environmental impact.

The influence of jointing on the stability of adjacent rock mass is analyzed using the coefficient of structural weakening and the joint spacing. The gravity–tectonic stress field of ore bodies in Gornaya Shoria, including the azimuth of the principal horizontal stresses is characterized. The laboratory tension tests of combinations of support systems as well as the small-scale extension tests of rock bolts made of different materials are discussed. The framework for a multi-factor digital model of safe and efficient support design for iron ore mines in Gornaya Shoria is based on the stability assessment of adjacent rock mass in terms of jointing and stress state.

Bolt reinforcement is a standard reinforcement method for jointed rock masses. However, regarding rock anchoring, the mechanical characteristics of the joint surface, as well as the angle between the bolt and the joint sliding surface, are important factors that affect rock support. Therefore, to understand the influence of a set angle, length, normal load, and other conditions that affect the shear strength of bolt joints, this study uses numerical software to establish the shear sliding model of bolt rock masses and analyzes the influence of the setting conditions of the bolt on the shear strength of a bolt rock mass, which can be done by changing the setting method of the bolt and normal mechanical conditions of the sliding surface. The results show that the shear strength of the anchor joint is not affected after the anchor reaches a certain length. The angle of the anchor strongly influences the shear strength of the anchor joint, and the shear strength curve is V-shaped, where the anchor angle is less than 90°. Moreover, the shear strength curve indicates a downward trend when the anchor angle is greater than 90°, and the shear strength of the anchorage joint increases with the increase of the normal load. Under the same anchor length (4 cm) in the anchor angle and shear strength coordinate system, the shear strength curve of the single anchor is above the shear strength curve of the double anchor, which is exclusively in the local anchor angle section under the condition of a large normal load and a large anchor angle. The shear strength curve of the double anchor is above the shear strength curve of the single anchor.

To study the mechanical responses of large cross-section tunnel reinforced by pretensioned rock bolts and anchor cables, an analytical model is proposed. Considering the interaction between rock mass and bolt-cable support, the strain softening characteristic of rock mass, the elastic-plastic characteristic of bolt-cable support, and the delay effect of installation are considered in the model. To solve the different mechanical cases of tunneling reinforced by bolt-cable support, an analytical approach has been put forward to get the solutions of stress and displacement associated with tunneling. The proposed analytical model is verified by numerical simulation. Moreover, parametric analysis is performed to study the effects of pretension force, cross-section area, length, and supporting density of bolt-cable support on tunnel reinforcement, which can provide references for determining these parameters in tunnel design. Based on the analytical model, a new Ground Response Curve (GRC) considering the reinforcement of bolt-cable support is obtained, which shows the pretension forces and the timely installation are important in bolt-cable support. In addition, the proposed model is applied to the analysis of the Great Wall Station Tunnel, a high-speed railway tunnel with a super large cross-section, which shows that the analytical model of bolt-cable support was a useful tool for preliminary design of large cross-section tunnel.

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.

Numerical simulation tests were performed on the arch-bolt combined supported mining tunnel through an improved numerical simulation approach. The typical soft rock roadway was took as the background, and the influencing factors such as ground stress level, lateral pressure coefficient, and support type and parameters were considered. The failure mechanism of a semicircular roadway with two straight walls was analyzed; results showed that the arch legs’ inward bending deformation and the arch-rock separation are the breakthrough of the global failure of the supporting system, and rock bolts breakage promoted the failure process. The effects of different controlling measures were analyzed including enlarging the bolt diameter, replacing the conventional bolts with energy-absorbing bolts, and setting arch locking bolts on the arch legs. The field test of the concrete-filled steel tube (CFST) arch-bolt composite support scheme was carried out in a high-stress soft rock roadway, and the results indicate the reliability of the main conclusions.

The reaction of earth to pull-out process of frictional rock bolts was here modeled by the distinct element method (DEM). Ten frictional bolts were prepared; the expanding shells of five bolts included convex edges and the others had the shells with concave bits. The strength of bolts was measured by applying a standard pull-out test; the results confirmed that the strength of shells with convex edges was remarkably more than the strength of other shells. Furthermore, a two-dimensional DEM model of the test was developed by a particle flow code; the obtained results showed that the reaction of rock particles to the contacts occurring between the convex edges and earth was considerably more than those of the concave bits. In the other words, the convex edges transferred the pull-out force into a large area of the surrounded rock, causing these bolts to have the highest resistance against earth movements.