Search

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

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

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

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

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.

Understanding the interaction between rock bolts and underground rock movement is critical for safe and cost-effective underground excavation design. Although early research on this subject involved a balance of theoretical analysis and field measurement, the current work focused on analytical and numerical studies. The present study aimed to study fully grouted rock bolt in Tabas Coal Mine using numerical and instrumentation methods. Tabas Coal Mine is the first fully mechanized coal mine in Iran, producing 1.5 million tons of coal per year. The mine extracts coal by both longwall and room and pillar methods. The roadways have a rectangular profile of 4.5 m width and 3.5 m height. The field investigations and geotechnical characteristics of rocks showed that the rock masses are weak, requiring a suitable support system. The roadway is intersected by a major fault zone. For the designed roadway, four patterns (two types of bolt with transfer load capacity of 180 and 195 kN) were considered. FLAC analysis showed that 13 bolt 2.4 m with 2 NO flexi bolt 4-m patterns were better than other patterns within the faulted zone. Assessing the optimum pattern through instrumentation, there was compromise between the results of numerical modeling with instrumentation reading such as telltales, sonic extensometer, and strain gauge rock bolt. Thus, there was an inspiring agreement between numerical and field methods.

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

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

Combined with the 3D FEM, end-anchored anchorage bolts were simulated by implicit anchorage bolt element. Implicit anchorage bolt elements hide in the elements of rock mass and extremely simplify the element subdivision. The calculated value of anchorage bolt stress is larger than the measured one for the most time. we further analyzed the reciprocity of anchorage bolt and rock mass, and then deduced the analytical equations of anchorage bolt stress and rock mass deformation under elasto-plastic state. The results indicate that it is essential to revise the anchorage bolts stress by using the formulas deduced when rock mass is softened or significantly deformed. Finally, a case study indicates that the calculated results agree with those measured. Some helpful methods are offerd for more accurate simulation of the support effect and anchorage bolts real stress state.

A ground convergence-support reaction analysis is extended to the case of an opening excavated in horizontally layered strata. Simplifying assumptions include linear elastic behaviour and classical beam theory. The method is used to demonstrate the influence on bolt loads of a number of factors including the timing of support installation, failure of the roof, number, thicknesses and flexibilities of the rock layers, type of rock bolt, bolt spacings, bolt lengths, roof span, and roof beam thickness.

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

Numerical simulation is a useful tool in investigating the loading performance of rock bolts. The cable structural elements (cableSELs) in FLAC3D are commonly adopted to simulate rock bolts to solve geotechnical issues. In this study, the bonding performance of the interface between the rock bolt and the grout material was simulated with a two-stage shearing coupling model. Furthermore, the FISH language was used to incorporate this two-stage shear coupling model into FLAC3D to modify the current cableSELs. Comparison was performed between numerical and experimental results to confirm that the numerical approach can properly simulate the loading performance of rock bolts. Based on the modified cableSELs, the influence of the bolt diameter on the performance of rock bolts and the shear stress propagation along the interface between the bolt and the grout were studied. The simulation results indicated that the load transfer capacity of rock bolts rose with the rock bolt diameter apparently. With the bolt diameter increasing, the performance of the rock bolting system was likely to change from the ductile behaviour to the brittle behaviour. Moreover, after the rock bolt was loaded, the position where the maximum shear stress occurred was variable. Specifically, with the continuous loading, it shifted from the rock bolt loaded end to the other end.

Rock bolts are highly important underground to prevent falls of ground and thus, ensure safety. In situ, the ground reinforcement is subjected to a combination of loads, such as shear, tensile, and/or bending loads. Studies have been conducted to evaluate the effect of either shear, tensile, or bending loads on the ground support. However, only a few studies have been done to evaluate the effect of combined load on the rock bolt performance. In situ, the order of loads acting on the bar cannot be predicted, and more than one load can act on the bar at the same time. Thus, a hybrid failure criterion was developed that covered all combined loading conditions. Fifty-eight un-grouted laboratory rock bolt tests were conducted in total under two load conditions: applying shear displacement to the rock bolt and then tensile load to failure, and applying tensile displacement to the rock bolt and then shear load to failure. From the test results, the failure criterion was developed based on regression analysis. A nonlinear relationship was determined for the loading condition variations. The failure criterion presented in this paper can be used as a guide in the rock bolt capacity design under all loading conditions.

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.

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.

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.

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.

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.

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.

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.

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.

Apparent discontinuities can be easily found in natural rock medium owing to the constant motion and change of the Earth’s crust, which contains large amount of discontinuous surfaces such as faults, joints, cracks and so forth. Rock bolt is one of the most effective and economical reinforcing tools in practical geotechnical engineering for a long time. This paper investigates the mechanical properties, anchorage effect, cracking, and coalescence process of intact rock-like specimens, rock-like specimens containing flaws, and bolted rock-like specimens containing flaws. A series of uniaxial compression tests, splitting tests, and biaxial compression tests are performed on these specimens. Some findings can be observed from this study. (1) The number of rock bolt(s) and the anchoring angle have a great influence on the anchorage effect of rock bolt(s). With the increasing number of rock bolt(s), the uniaxial compressive strength (UCS), the splitting peak strength (which can be converted to the tensile strength), and biaxial compressive strength (BCS) can all be improved, whose variation tendencies do not follow a linear relationship. (2) The contributions to the tensile strength for rock bolt is greater than that to the UCS for the same-type specimen. (3) In the biaxial compression tests, with the increasing of the lateral pressures, the anchorage effect of rock bolt gradually declines and the lateral pressure plays a dominant role in improving the strength of the specimen. The failure characteristics of three types of laboratory tests have also been systematically analyzed in this paper.

Although an appropriate tunnel construction method can be conducive to safeguarding the deformation of surrounding rock mass, too much emphasis on controlling deformation leads to ignoring the importance of accelerating tunnelling speed and saving construction costs. Therefore, this paper presents a historical case of large cross-section tunnel in loose ground, which was excavated by optimal multi-step excavation method. First, to investigate the influence on rock mass deformation by the centre diaphragm (CD) method and optimal tunnelling method, comparative studies were performed by numerical simulation of the hypothetical construction schemes. Then, field monitoring was carried out to further reveal rock mass deformation and structural load characteristics with optimal excavation method. Numerical simulation results show that, compared with CD method, the optimal construction method is more helpful in controlling tunnel invert uplift but has minor disadvantages of increasing horizontal displacement of surrounding rock mass. However, tunnel crown settlement excavated by CD method is approximately similar to that of optimal construction method. Field test results confirm the effectiveness of optimal tunnelling method on controlling excavation-induced tunnel deformation; the maximum vault settlement and horizontal convergence of drift are approximately 57.5 mm and 8.5 mm, respectively. Moreover, the maximum inside and outside forces of steel set are 172.5 MPa and 151.5 MPa. The rock bolt is subjected mainly to a tensile force with a maximum value of up to 29.3 MPa. Therefore, as the crucial part of initial support structure, steel sets and rock bolts are under appropriate conditions and do not exceed their bearing capacity.

Construction of underground tunnel for highway in the urban area is a challenge to overcome many problems such as subsidence of residential houses and commercial buildings, broken of utilities lines, and vibrating and noise due to rock blasting. This paper presents the analysis of stability of 5.6 km long underground tunnel both partially completed tunnel and planed tunnel section, and the cause of tunnel collapse due to sinkhole. Several methods applied to analyze the stability of tunnel by the monitoring the instrumented data; settlement of ground, ground water level, and lateral and vertical displacement of internal face of tunnel. Soil investigation was also conducted to identify the exact soil layer characteristics and strength of soil and rock, and stratification of soil ground. The rock core was obtained and evaluated for the specimens from the 50 m long horizontal pre-boring process. The numerical analysis was carried out for the stability of tunnel in terms of internal displacement of tunnel, shotcrete bending compressive stress, and rock bolt axial force. GPR survey was also conducted for detection of sinkhole along the route of underground highways.

The main access road of Mae Mao dam hydropower project, Chiang Mai province, Thailand, has failed during a sunny day. The question has been raised of what might have triggered the failure. At the day of failure, the precipitation amount was very small and the accumulated rainfall amount in the past days was also not significant to trigger the failure. Detail geotechnical investigations were carried out including the geological survey, seismic refraction survey, bore holes drilling and various laboratory testing. The rock found at site was Phyllitic and Quartzitic sandstone. It appears that the engineering properties of Phyllitic sandstone are more susceptible for slope failure. Rock and soil slope stability analysis have been done, however none of them was found to be the cause of failure. After carefully reexamining the geologic structure, it was found that there is a weak plane layer that resulted in the failure. A suitable stabilization method has been selected as the economical remedy of the slope failure. In this case, flattening the back slope is the economic method along with the drainage system specially designed to reduce the pore water pressure of the weak plane layer. Rock bolts and rock anchors were found to be uneconomical and unsuitable. Probabilistic approach was used to optimize the engineering safety and construction cost.

For better rock mass classification and support design of supporting system, geotechnical investigations were carried out for the proposed 196.80 m long, 50.30 m high and 25.24 m wide underground pump house cavern of a lift irrigation scheme. Geotechnical site investigations, rock support design, excavation and related works and modification of supports design according to the observational construction method are the principal activities for the construction of underground rock cavern. The investigation includes engineering geological mapping, geological logging of drill holes, in situ permeability test and laboratory testing on core samples. Detailed geotechnical investigations carried out supporting the feasibility of underground excavations of cavern within the hillock. Support system recommended is the rock bolt and steel fibre reinforced shotcrete based on empirical approaches and their capacity is determined. Capacity of support system was conducted to evaluate the efficacy of the proposed support system.

Roof failure related to laminated and jointed rock masses has been recognized as a major cause of fatalities in underground coal mines. This paper presents numerical simulations of the failure of a laminated and jointed roof using the discrete element method, while focusing on the initiation and propagation of (and the interactions between) microcracks and macroscopic fractures and the development of the associated damage, stress and deformation in the laminated and jointed roof. Numerical modeling shows that the behaviors of bedding planes and vertical joints governed the progressive failure of the laminated and jointed roof. Microseismic activity corresponding to the development of microcracks in the roof intensified before major roof displacements and subsided in the later stages; such activity could serve as a precursor of roof fall events. The effectiveness of various support elements (rock bolts, metal meshes, and anchor cables) on the roof stability was numerically evaluated. Rock bolts and/or anchor cables produced a banded compressive stress in the laminated roof, restricting the separation of bedding planes. Rock bridges in the bolted roof suppressed the formation of macroscopic fractures and significantly reduced the interactions between fractures. The roof skin support with a metal mesh provided confinement to the fractured roof, thereby preventing fragmented rocks from falling and restricting progressive spalling between bolts. The numerical simulations presented in this paper can be helpful for understanding the likely failure mechanisms of laminated and jointed roofs and for finding potential solutions.

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.

Columnar jointed basalt (CJB), characterized by a hexagonal joint network, poses a great challenge for geotechnical designs due to its time-dependent propensity to deformation and damage. In this paper, the geometrical structure and failure modes of CJB exposed on a dam foundation is presented. The study was conducted during the unloading excavation of the Baihetan dam, China, aiming to provide a more in-depth understanding of the CJB damage mechanism. A systematic field test for the columnar jointed rock mass was also carried out on the dam foundation using ultrasonic P wave measurements. It was found that the excavation damage zone of the CJB not only showed time-dependent damage both in depth and degree but also showed spatially inhomogeneous distribution of the damage degree. Sequential regression analysis further demonstrated the time-dependent behavior of the CJB. Successive observations using digital borehole cameras in the field showed that the unloading damage mechanism of the Baihetan CJB was mainly related to the time-dependent opening of the initially closed joints. Several reinforcing methods for improving the stability of the damaged CJB, including the injection of liquid cement and the installation of the prestressed rock bolts, are discussed.

Shear testing of large size rock mass in the laboratory is a tedious and challenging task. Difficulties further arise when testing is done with rock bolts. Complex interaction between intact blocks of mass, joint and bolts creates problems during shear testing. Hence, for testing of large rock mass with bolts, special technique and facilities are required. The present paper summarized the details of the development of a large sized direct shear testing facility for passive bolt—reinforced rock masses. A specially designed servo-controlled direct shear test apparatus having shear box of size 750 mm × 750 mm × 1000 mm has been fabricated in the Geotechnical engineering laboratory of IIT Roorkee. The normal loading capacity of the apparatus is 1500 kN while shear loading capacity is 2000 kN. To evaluate the performance of the apparatus, specimens of blocky mass were prepared and tested without and with bolts. The present paper discusses the details about the large direct shear apparatus and testing methodology.

Because of the chemical corrosion of underground water on rock bolts that persistently present in complicated geological conditions, the mechanical properties and integrity of bolts may be affected, even bolt failure may occur. In the present article, first, the effect of the chemical reaction and the corrosion expansion on bond strength has been analyzed. Second, a coupled creep model of anchor structures and soil has been proposed based on the stress condition and deformation characteristics of the anchor segment. In this model, the Mohr–Coulomb (M–C) plastic element was used to simulate the complete creep deformation, and the constitutive equation of the coupled creep and the loss of the anchor force were obtained. A comparison between previous studies and the present model indicates that the results of this model agree well with previous results. The coupled creep of the anchor segment and soil is properly reflected in this model. Thus, this model can accurately predict the variation of the anchorage force.

Roof collapses during the constructions of underground excavations in weak rocks is a serious problem. In particular, excavations in shallow depths and in incompetent rocks may initiate the caving of the overburden materials and sinkhole formation. Sinkholes have significant environmental impacts and more importantly, they threaten the stability of surface and subsurface infrastructures above the excavations. This study investigates the formation of sinkhole subsidences in shallow excavations in poor and problematic rocks. The progressive collapse and sinkhole subsidence in the Dolaei road tunnel is considered as a case history. Understanding the geological and geotechnical characteristics of rocks is the fundamental step for analysing the mechanisms of instability. Rock mass characteristics are reviewed, and the most effective factors impacting on the tunnel’s stability are identified and discussed. The role of the method of excavation and support in controlling ground movements are assessed through numerical modellings. The effect of pre-support as a practical technique for controlling ground movements and preventing sinkhole formation in weak rocks is also discussed. Outcomes of this study indicate that the rock mass surrounding the Dolaei tunnel consists of highly tectonised and foliated metamorphic rocks. Schistosity and foliation considerably impact on the strength and deformability of the rock mass. This study shows that the geological characteristics of the rocks in the Dolaei tunnel had substantial effect on the collapse and sinkhole formation. The numerical findings also reveal that employing pre-supporting techniques, particularly forepoling, using a staged excavation and applying composite support systems (consisting of rock bolts and reinforced shotcrete) are practical remedies to prevent the progressive collapse, and to avoid the formation of sinkholes during excavation in weak rocks in shallow depths.

This paper presents the results of the geological and geotechnical studies of the rock masses along the Meram railway tunnel. The study area consists of limestone which is the base rock and covers a large area between the entrance and exit portals and sandstone which is only observed at the slope of the hill that overlies the tunnel. Results obtained from rock mechanics testing indicate that the limestone along the tunnel route is a moderately strong rock with an average uniaxial compressive strength of 61.88 MPa. Rock mass qualities of the rock units along the tunnel were determined by means of rock mass rating (RMR), Geomechanics Classification (Q) system and New Australian Tunneling Method (NATM). The study shows a very poor-to-fair rock mass conditions, with a rock mass rating (RMR) ranging from 40 to 49 and rock mass quality (Q-system) ranges between 0.69 and 2.52. The limestone was classified according to NATM into two ground categories, which are B2 (friable) and B3 (very friable). Correlations with the RMR, Q-systems and NATM suggested rock bolts and shotcrete as a suitable preliminary support system for the rock mass during excavation.

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

Electro-hydraulic and jumbo drills are commonly used for underground coal mines and tunnel drives for the purpose of blasthole drilling and rock bolt installations. Not only machine parameters but also environmental conditions have significant effects on drilling. This study characterizes the performance of button bits during blasthole drilling in coal measure rocks by using multiple regression analyses. The penetration rate of jumbo and electro-hydraulic drills was measured in the field by employing bits in different diameters and the specific energy of the drilling was calculated at various locations, including highway tunnels and underground roadways of coal mines. Large block samples were collected from each location at which in situ drilling measurements were performed. Then, the effects of rock properties and machine parameters on the drilling performance were examined. Multiple regression models were developed for the prediction of the specific energy of the drilling and the penetration rate. The results revealed that hole area, impact (blow) energy, blows per minute of the piston within the drill, and some rock properties, such as the uniaxial compressive strength (UCS) and the drilling rate index (DRI), influence the drill performance. © 2015, Springer-Verlag Wien.

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.

The rock bolt support system is one of the major elements of safe maintenance of roof sections in coal mines almost all over the world. Leading coal producers such as the USA, South Africa, and Australia use the rock bolt system as practically the only one in roadways, room, and pillar sections. It has also been increasingly popular in other coal-producing countries including Poland. The function of the rock bolt support system is to anchor and reinforce the rock zone in the near field of an underground opening to deeper rock strata. In contrast to the conventional steel framework, the crucial part of the rock bolt support system is concealed from the observer. The developed method uses modal analysis procedures and is based on impact excitation. It consists of two main phases: experimental, which entails the measurement of modal characteristics of an investigated object; and theoretical involving reference to the Finite Element (FE) model databases. The main assumption of the method is that boundary conditions have to be such that the in situ bolt behaves like a cantilever. Since an installed rock bolt acts as an oscillator, different lengths of grouting discontinuity form different boundary conditions and change its modal parameters.

For many decades, a tunnelling method has been in use which effectively minimizes the use of concrete, which should be one of the goals in our CO2-producing planet. We call the method NMT (Norwegian Method of Tunnelling) and emphasize its ‘single-shell’ characteristics, to distinguish it clearly from double-shell NATM (the so-called New Austrian Tunnelling Method), which is recommended to have (ASG, NATM: the Austrian practice of conventional tunnelling 2010): shotcrete, mesh, lattice girders, rock bolts (if in-rock), drainage fleece, membrane, and the final load bearing and often steel-reinforced concrete lining, including the invert when in poor rock conditions. This tunnelling method is inevitably several times more expensive, uses many times the volume of concrete, takes longer to build, and requires at least a ten times larger labour force than single-shell NMT. The single-shell tunnels for road or rail or hydropower or water transfer, or for large caverns for storage of oil or food, or for hydropower machine and transformer halls, can be made stable by judicious application of a well-used (>2000 case record based) the so-called Q-system of rock mass quality estimation. The latter encompasses a rock mass quality scale from 0.001 (equivalent to a serious fault zone, where we also may need a local concrete lining) to 1000 (equivalent to massive unjointed rock) where careful blasting will remove the need even for shotcrete. In general, rock masses where we need tunnels or caverns will lie closer to ‘mid-range’ (i.e. closer to Q = 1 which is described as ‘poor quality’). Here we would need combinations of corrosion-protected rock bolts and high quality fibre-reinforced shotcrete, with stainless steel or polypropylene fibres. We may also need systematic high-pressure pre-injection of micro-cement and micro-silica, which may add 20% to the (low) starting cost of the NMT excavation. Written as B + S(fr) in short-hand, NMT has rock bolt c/c spacing in metres and shotcrete thickness in centimetres, as specified by the range of Q values and excavation dimensions. The details are also affected by the planned use. For instance, at our record-breaking Olympic cavern of 60 m span (for housing 5400 spectators or later concert goers), B = 2.5 m c/c + S(fr) 10 cm were (and remain 25 years later) the stabilizing and permanent measures of support and reinforcement. Deformation monitoring and distinct element (jointed rock) numerical verification showed 7–8 mm of maximum deformation in the arch. The moderate Q value range of quality of 2–30 (poor/fair/good) and RQD = 60–90 indicated a well-jointed gneiss, which had only moderate UCS = 90 MPa compressive strength.

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.

An increasing number of tunnels have been constructed worldwide for a diverse range of infrastructure projects, including highways, subways, railways, and various utilities. As a result, tunnel engineers are increasingly aware of the importance of safe and economic tunnel construction. Peck (1969) addresses the following three issues in terms of tunnel construction: maintaining stability and safety during construction, minimizing unfavorable impacts on third parties, and performing the intended function over the life of a project. Among these issues, the first is directly related to the behavior and stability of the tunnel support system. In rock tunneling, the New Austrian Tunneling Method (NATM) is a widely used tunneling method in which the shotcrete liner acts as the main support. Shotcrete is sprayed on the tunnel excavation surface very shortly after being excavated and supports the ground load in conjunction with the rock bolt and the steel rib. Barrett and McCreath (1995) explained the failure modes for the shotcrete liner (Fig. 1). It has been shown that the behavior of the shotcrete liner is highly dependent on the adhesion strength at the shotcrete–rock contact. For example, if this strength is sufficient enough, then the mode of failure may be a direct shear failure. Otherwise, the shotcrete liner may fail under a flexural condition. In addition to the behavior of structural characteristics, sufficient adhesion strength can merge the surrounding ground and the shotcrete liner into a single body, increasing tunnel stability. Therefore, the adhesion strength at the shotcrete– rock contact can influence the failure mode for the shotcrete liner, which, in turn, can impact the overall tunnel stability. This indicates a need for a better understanding of this adhesion strength. An adhesion strength test method (Swedish Standard 1987) from Sweden drills two circular slots with a doublediamond drill. Here, the inner circular slot is drilled through the substrate and shotcrete, and the test equipment is attached to the shotcrete drill core for the pullout test. The test method can be used only for relatively stiff shotcrete. Brennan (2005) explains a shotcrete bond test method from the International Concrete Repair Institute (ICRI). This method uses a core bit to drill the applied shotcrete and substrate, and then attaches a rigid steel disk to the top of the core by using epoxy. Then, the test equipment applies some tensile load to the disk until it fails. This test method is similar to the direct test of tensile bond strength by the American Concrete Institute (ACI 2004) or the American Society for Testing and Manuals (ASTM 2004). A thin spray-on liner (TSL) has been developed and used in the mining industry since the 1990s. Ozturk and Tannant (2010) compare various adhesion test methods based on TSL and suggest a pulloff test of an elevator bolt (33 mm in diameter) attached to the liner with strong epoxy. This test uses a thin-walled diamond core bit to isolate the test area by overcoring. Ozturk (2012) provides a method for calculating the liner–substrate adhesion M. Son (&) Department of Civil Engineering, Daegu University, Jillyang, Gyeongsan, Gyeongbuk 712-714, South Korea e-mail: mson@daegu.ac.kr

A simplified method was proposed for the design of concrete lining in underground rock cavern/tunnel against shock loading. The loading may result from the detonation of explosives on ground surface or ground penetration projectiles exploding adjacent to the cavern/tunnel. The resulting problem necessitates the solution of the dynamics of a beam loaded by a transient pressure uniformly distributed over the span. According to mechanical characteristics of the system with rock bolt and shotcrete, a dynamic support design method based on equivalent single degree of freedom (SDOF) was put forward. The SDOF method was applied to obtaining the maximum displacement at the mid-span of the beam, which is often the controlling factor in the blast-resistant design. In the formulation of the problem, the proposed method combines the phenomena of spalling and structural dynamics theory. An example is provided to demonstrate the applicability of this simplified method.

This paper deals with the modeling of jointed rock masses reinforced by rock bolts. It is well known that rock bolts are extremely effective in reinforcing jointed rocks. However, if a continuum approach is adopted for modeling jointed rock masses, it is often misleading to evaluate the effectiveness of the rock bolts by numerical analyses such as the finite element method. This may be due to the fact that since no more joints exist in the continuum, the effectiveness of the rock bolts in constraining the relative displacements along the joints cannot be evaluated properly. In order to investigate the reinforcement effect of rock bolts, physical model tests were performed in the laboratory. The test results revealed that jointed rock masses should be modeled as an equivalent continuum after the installation of rock bolts and that the mechanical parameters of the equivalent continuum should be evaluated by considering the reinforcement effect of the rock bolts. Therefore, the values of the mechanical parameters differ from place to place in accordance with the relation between joint orientation and rock bolt direction, even though joint systems are homogeneous. In conclusion, in the continuum approach for modeling jointed rock masses, it should be emphasized that rock bolts and jointed rock masses should not be modeled separately but should be modeled simultaneously by considering the reinforcement effect of the rock bolts in constraining joint movement. The modeling of shotcrete reinforced by steel ribs is also discussed in comparison to the modeling by rock bolts.

1 IntroductionRock bolting has been widely used for rock reinforcementin civil and mining engineering for a long period. Initiallysporadic bolting was used to secure loose blocks in shallowunderground openings. Bolts function as hanging elementsto fasten loose blocks to stable formations behind. A simpledemand, in this case, is that bolts have to be strong enoughto sustain the dead weight of the unstable blocks. There-fore, the strength of the bolt is important for rock support inshallow locations. In rock masses at depth in which the insitu stresses are usually high, loose blocks seldom existfollowing excavations, but rather stress-induced rock fail-ure occurs in many cases. In general, the rock bolts used inunderground openings at depth are still the same types asthose used in shallow openings. The main difference maybe in the manner of bolting. Systematic, rather sporadic,bolting is used more than often today for rock support indeeply located openings. The loading condition of a rockbolt at depth is more complicated than a bolt for securing aloose block. It has been often experienced that conven-tional rock bolts do not provide satisfactory rock rein-forcement because of their premature failure in high stressrock conditions. New types of rock bolts are desired inorder to deal with the harsh rock conditions at depth. Inorder to figure out the desired performance of such rockbolts, it is necessary to have a good knowledge of how rockbolts behave in high stress rock conditions. In this articlethe loading conditions and the failure modes of rock boltsare presented through field observations.The aim of the presentation of the field observations isto show the shortcomings of the conventional rock boltsand to figure out the features rock bolts should possess forrock support in high stress rock masses.2 Rock Bolts Observed in the FieldIt is an obvious fact that the in situ rock stresses becomehigh at depth. The consequence of the high in situ stressesis that rock fails either in the form of time-dependentdeformation (creeping) in weak and soft rocks, or in theform of dynamic (rock burst) or quasi-dynamic (slabbing)ejection in hard rocks. The failure mode of rock has a directimpact on the design of ground support systems.The field observations presented in this paper wereconducted in cut-and-fill mines. With the cut-and-fillmining method, stopes are mined in 3–5 m high slicesalong the strike of the ore body by breasting using drilljumbos. After a slice has been mined out it is backfilledwith waste rock and tailing. Because of the nature of themining method, the roof rock of a previous slice cut will beexposed on the advance face of the current slice cut.Therefore, the cut-and-fill mining method provides aunique opportunity to observe the interaction between boltsand the rock in situ. The observations on the advance faceof mine stopes will reveal the loading conditions as well asthe failure modes of rock bolts in high stress rockconditions.Figure 1 shows a part of an advance face in a cut-and-fillmine stope. A number of sub-horizontal stress fracturestogether with a rock bolt are exposed on the face. The boltwas one of the roof bolts that had been installed in theprevious slice of the stope. The fractures have a domedform with the dome peak at the position of the bolt. This

Squeezing failure is a common failure mechanism experienced in underground coal mine roadways due mainly to mining-induced stresses, which are much higher than the strength of rock mass surrounding an entry. In this study, numerical simulation was carried out to investigate the mechanisms of roadway squeezing using a novel UDEC Trigon approach. A numerical roadway model was created based on a case study at the Zhangcun coal mine in China. Coal extraction using the longwall mining method was simulated in the model with calculation of the mining-induced stresses. The process of roadway squeezing under severe mining-induced stresses was realistically captured in the model. Deformation phenomena observed in field, including roof sag, wall convexity and failed rock bolts are realistically produced in the UDEC Trigon model.

Numerical simulations have been performed to assess the stability of a large underground powerhouse in the Himalayas, using both finite element and distinct element approaches. Large convergences (around 250–300 mm) along the 45 m high wall of the Machine Hall have been measured by total station measurements during construction and up to 60 mm in the post construction and operational phase. Displacements are continuing at a reduced rate of about 3 to 6 mm per year. A large number of rock bolts have failed in the powerhouse and the shotcrete liner presents cracks and failure marks, thus questioning the long-term integrity and stability of the existing rock support system. The purpose of the numerical analysis was to back-calculate the prevailing rock mass conditions surrounding the machine hall, thereby gaining a better understanding of the mechanisms responsible for the instability in the cavern. The results from the analysis indicate that there was possibly an underestimation of the rock support requirements needed for the cavern, coupled with a not so good installation (incomplete grouting of the bolts). Both the length and the capacity of the rock bolts were underestimated presumably due to the existence of a high stress regime in the area, which may not have been taken into consideration in the design of rock support. The effect of a possible earthquake, frequent in this area, has also been studied and proves to induce a significant increase of the displacements and the support failure.

The aim of this study is to examine cracking and instability of the high and steep left bank slope with weak rock mass structures, and effectiveness of the reinforcements designed for the slope at the Jinping hydroelectricity power station, southwestern China. A new geomechanical model testing method is first proposed for evaluating slope safety factors. In the proposed geomechanical model test, the slope is constructed on a flat testing bed, which can be rotated by hydraulic lifts. By increasing the rotation angle of the testing bed, the forces tending to induce the sliding of the slope are increased, which may cause crack initiation and propagation in the slope and result in rock mass slippage. Thus, through the proposed geomechanical model test, the slope failure mechanism, progressive failure process and final failure pattern can be studied. Moreover, the stability safety factors can be evaluated according to displacement data monitored by sensors installed in the geomechanical model. The geomechanical model test of the Jinping left-bank slope reveals that the bending and toppling cracks occur simultaneously in the unreinforced zone of the slope together with strong relaxation and tension cracks. It is found that the factor controlling the rock mass failure and instability is the structurally weak rock mass and the dominant failure mode is the integral catastrophic instability mode, in which the slope energy is totally dissipated and the slope destabilizes at a limit state. The reinforcement installed in the large area above the elevation of the dam-slope abutment and the unloading action due to the excavation of rocks lying above the dam platform have effectively improved the anti-slide safety factor of the slope, and thus increased its inherent safety factor. On the basis of the geomechanical model testing results and their comparisons with field monitoring results, it is concluded that the installed rock bolts and long anchor reinforcement measures are very effective in keeping the cracks closed in the rock mass and maintaining the slope stability.

This paper analytically investigates the performance of fully grouted rock bolts in typical scenarios, including pullout test, suspending loosened block, and increasing joint aperture, respectively. The interface shear stress distribution follows the model proposed by Li and Stillborg (Int J Rock Mech Mining Sci 36:1013–1029, 1999), while the axial behavior of the bolt shank obeys the elasto-plastic (yielding-hardening) constitutive model of steel. Three different failure modes are taken into account: tensile failure of bolt shank, bolt shank being pulled out along the bolt/rock interface, and loss of face plate. The evolution of the interface shear stress and the axial tensile stress are examined for both long and short bolts under displacement and load boundary conditions. The derived charts are able to predict the load capacity of fully grouted bolts in pullout test, the minimum length requirement of the bolt to suspend a loosened block, and the maximum allowed opening displacement of a rock joint for a fully grouted bolt. In addition, different potential failure modes are specified. Full range load–displacement curves are produced and compared for various failure modes. The derived charts could be directly used in rock-bolting design.