A study of the behaviour of
reinforced elements subjected to shear

L Mahony and PC Hagan

The University of New South Wales (UNSW), Sydney

This paper outlines the design, construction, commissioning and subsequent investigation of a shear test laboratory facility at the University of New South Wales (UNSW). The facility was developed with the financial support of ACARP to understand the impact of shear loading conditions on the performance of reinforcement elements such as rockbolts and cablebolts. While it is acknowledged that shear loading commonly occurs, progress towards a better understanding of its impact has been limited due to issues in modelling behaviour. Preliminary testwork indicates that many of these issues have been overcome in the design of the current facility.

The facility is based on a single shear failure plane design which has overcome many of the issues associated with double shear designs such as dealing with high resultant loads and deformation to the support structure. A program of preliminary tests indicates that behaviour of the rock reinforcement system is less a function of the behaviour of the individual components than the interactions that occurs between the components that make up the system.

 

Introduction

Strata control is one of the core risk areas in underground coal mining. The use of appropriate technology for ground support whether it be roof or ribs, primary or secondary and the effective management of this technology can be pivotal to achieving a safe and economically viable mine.

Rock reinforcement elements provide a significant proportion of their ground control capacity through their ability to resist shear movement of the surrounding rock mass. This potential shear movement may take the form of sliding on horizontal bedding planes leading to strata bending; or block displacement along geological structures such as joints and other discontinuities. The shear resistance offered is far greater than simply the shear strength of the reinforcement element since the shearing action results in a normal or axial load in the element which is greatest at the shear plane and dissipates with distance away from the shear plane. The normal force clamps together the rock surfaces, mobilising the frictional properties acting over the large surface area of rock thereby contributing to frictional resistance.

While much research in the past has focused on the effects of axial loading of reinforcement elements in rock, less attention has been directed at understanding how these elements behave when subjected to shear loading. It is widely acknowledged that shear loads can have an important bearing on the stability of underground excavations; however, progress in this area has been hampered by the complexity of building a physical model that can reliably simulate shear loading conditions and the interplay that occurs between the different components.

The research conducted at the School of Mining Engineering at The University of New South Wales (UNSW) initially involved a review of the current understanding followed by design and building of a shear test laboratory facility to monitor performance of reinforcement elements and undertaking a preliminary round of experiments to begin to understand the behaviour of reinforcement elements on a rockmass.

The review found there is only a limited understanding. The review also highlighted the poor understanding of the effect of installation method of reinforcement elements and the influence of varying the loading conditions such as loading rate, pre-tension, torque and normal tension/loading on performance.

Design of Shear Testing Facility

The review as reported by Hartman and Hebblewhite (2003) commented that poorly designed and/or constructed test facilities had often had a detrimental impact on the quality of subsequent research activities in terms of integrity of results and/or validity of the findings. Hence a much greater effort in the design phase was required than had originally been envisaged with computer modelling of the load distribution being important in optimising the design.

The design objective was to be able to simulate conditions that are commonly encountered in underground mining environments. The initial design incorporated a double shearing action but it was subsequently found that a design with a single shear failure plane could be constructed with sufficient rigidity that minimal block rotation would occur about the reinforcement element. This design had an important advantage in that it effectively halved the load that needed to be applied to the system during testing and hence lowered the amount of reinforcement necessary to maintain stiffness of the facility.

In anticipation of the later experimental work, design of the laboratory facility had to cater for:

The as-built laboratory facility is based on a hydraulically actuated Avery-Denison compression test machine with an axial load capacity of 3600 kN. The facility has a rated shear loading capacity of 600 kN and is capable of isolating many of the operational variables necessary for experimentation. A schematic of the facility is shown in Figure 1.

Figure 1. Schematic of shear test laboratory facility.

The data acquisition system used in the test program incorporated several pressure transducers, displacement transducers (LVDT) and a load cell.

Each “rock” sample used in the test program was comprised of two concrete blocks cast separately in specially fabricated 10 mm thick steel casings. The blocks remained in their casing during a test to ensure an even load distribution as well as to provide confinement. The lengths of the two concrete blocks were 250 mm and 1000 mm with cross-sectional dimensions of 280 mm x 280 mm. After curing, the ends of the two blocks were bolted together in such a manner that there was a 50 mm offset between the centre-lines of the two blocks. A hole was then drilled through the two blocks and a rockbolt installed as shown in Figure 2. The blocks of concrete were cast with the casings standing upright. This was to ensure the two shear surfaces would have the same level of surface roughness.

Figure 2. Layout and dimensions of a test sample with rockbolt in place.

The shear test laboratory facility comprising test sample and monitoring equipment with structural modifications to the compression test machine is shown in Figure 3.

Figure 3. Shear test laboratory facility showing test sample in place.

Test PROGRAM

Three series of experiments were undertaken in the test program. While the Series 1 tests provided some initial results, its purpose was primarily to confirm functionality of the laboratory facility. Some modifications were made to the facility following the tests.

Six tests were undertaken in each of Series 2 and Series 3 with three strain-gauged rockbolts used in Series 3. The concrete mix was altered between the three test series in order to assess any effect of rock strength on shear behaviour. Properties of the concrete in the three test series are summarised in Table 1.

Table 1

Material Property Characteristics of Concrete

 

Series 1

Series 2

Series 3

UCS (MPa)

65.9

46.9

68.7

Young’s Modulus (GPa)

38.4

34.5

32.2

Poisson’s Ratio

0.12

0.13

0.13

Cure time

52 days

31 days

34 days

 

A 23 mm BX rockbolt was used in the test program with typical UTS and yield strength of 335 kN and 240 kN respectively. For practical purposes, rockbolts were installed using an ARO roofbolter at Hydramatic Engineering in Newcastle using an industry-scale rig found in common use in underground coalmines. While use of the roofbolter was necessary because of the length of borehole and strength of rock, it also introduced some minor issues especially regarding repeatability of installation. The drilling arrangement can be seen in Figure 4.

Figure 4. Drilling arrangement for installation of rockbolt in a test sample.

During the Series 2 tests, each test sample was loaded and unloaded up to four times. This process was followed as a consequence of three factors. First, the shear displacement of the laboratory facility was limited to 40 mm. Second, the strength of the rock mass /concrete was 47 MPa, the lowest in the three test series. Third, no nut and face plate was installed on the end of the most of the rockbolts used in Series 2. The combination of these factors limited the maximum shear resistance that could be developed.

The load-displacement curves for each test in Series 2 exhibited a similar trend. As shown in Figure 5, there was an initial stiff loading phase up to some transition point followed by some yielding with continued displacement between the shear surfaces.

Figure 5. Load-Displacement graph showing four loading and unloading cycles.

In one test, resistance to shearing increased with each subsequent loading cycle up to a maximum load of 350 kN whereas in other tests, the peak load reduced in subsequent cycles.

In all the tests where the rockbolt was not constrained by a faceplate and nut, failure or rupture of the rockbolt was never achieved. It was thought that this was due to failure at the rockbolt/resin interface allowing slippage within the relatively short length of encapsulation. Hence the rockbolt was free to deform, limiting the amount of shear resistance and axial load developed in the rockbolt.

The relatively low strength of the concrete allowed deformation of the rockbolt to occur with failure over a significant area of the concrete around the rockbolt. The lack of fixed constraints on the rockbolt due to a combination of slippage in the rock mass and mobilisation of the pivot points meant the system was able to sustain a reasonable level of resistance to shear loading over a large range of shear displacement.

The extent of yielding that occurred and failure of concrete around a rockbolt is illustrated in Figures 6 and 7.

Figure 6. Extent of yielding in rockbolt and sustained
failure around the rockbolt.

Figure 7. Schematic indicating the amount of yielding and damage to concrete.

With a faceplate and nut attached during the Series 3 tests, a somewhat different load-displacement characteristic was observed. Figure 8 shows a typical load-displacement result for a pre-tensioned rockbolt. The top (dark grey) line indicates the variation in shear load with shear displacement. The load initially increased sharply to the pretension load indicating the system is quite stiff. Thereafter stiffness reduced as there was a gradual increase in load up to some peak value with failure occurring shortly after. The peak load sustained in the test was approximately 400 kN at a corresponding shear displacement of 45 mm. This is much higher the ultimate tensile strength and nearly double the shear strength of a rockbolt.

Figure 8. Variation in shear and axial loads with shear displacement
when using a pretensioned rockbolt.

The lower (light grey) line in the graph indicates the change in level of normal or axial load as measured at the load cell at the collar of the borehole some 250 mm from the shear plane. The line begins from a threshold axial load of 40 kN, equivalent to the pretension applied to the rockbolt. It would appear that there was initially little measurable change in axial loading at the collar until a displacement of approximately 6 mm corresponding to a shear load of 150 kN. Above this level, axial load increased at a rate of 6 kN/mm up to a maximum of 150 kN compared to a slightly higher rate of 7 kN/mm for shear load.

Figure 9. Profile of failed rockbolt from shear loading.

Figure 9 shows the profile of a failed rockbolt. It was found that there was appreciably less failure of the concrete surrounding the rockbolt as compared to that observed in the Series 2 tests. This is in accord with the higher strength concrete used in the Series 3 test samples. The combined constraints of stronger concrete and fixed end point of the rockbolt meant sufficient stresses could be developed to cause failure of the rockbolt.

This is in accord with previous findings that the level of induced normal force, sn, varies directly with applied shear force, t, such that

                                 1)

where is m the friction in the system principally the friction between the rock surfaces. Re-arranging and taking account of the clamping stress, Co, we get the more usual equation

                           2)

A set of tests with three strain-gauged bolts was undertaken to examine the variation in load with distance along the rockbolt from the shear plane. Unfortunately, due to a combination of issues associated with installation, orientation of the bolts, gauge alignment and instrumentation little quantifiable results were obtained. The strain gauge readings during a test fluctuated widely with strains in excess of ±15,000,000 microstrain being measured. During earlier calibrations tests, approximately 500 microstrain was measured with an applied axial load of 100 kN. Further work will be needed to determine the cause for this discrepancy.

Figure 10. Variation in axial strain along a strain-gauged rockbolt
subjected to shear loading.

Although the magnitude of values are questionable, Figure 10 shows there was some correlation especially between the level of strain along the rockbolt and the level of applied shear load and shear displacement.

Failure of the Reinforcing Element

An examination of all failed rockbolts indicated they did so in a ductile manner with necking evident as can be seen in Figure 11. Ductile failure occurred between the two plastic hinges (bending regions) on either side of the shear plane associated with deformation caused by the shear displacement; this is where the bending moment was greatest. Between the two hinge points, the loading regime was altered such that given sufficient shear loading, the rockbolt failed axially in tension. This would account for the higher failure loads observed that were well in excess of the shear strength of the rockbolt.

Figure 11. Profile of a rockbolt that failed in shear.

Inspection of the failed surface of the reinforcing element confirmed the failure mechanism as being a typical ductile bending, necking and then tensile failure. The failure initiated in the centre of the necked region with the crack, then progressed laterally towards the edge of the element in the area known as the radial zone. The fracture is then completed via a shear lip on the outer extremities of the element. A reinforcing element that is subject to a pure axial load creates a symmetrical shear lip around the outer edge of the failed rock bolt section as indicated in Figure 12, whereas the shear lip in the failed element subject to a shear load creates a more ellipsoid shape, engaging at the upper and lower section of the element. The shear lip is negligible at the sides of the element where the applied shear load is perpendicular to the element.

The development of this unique shear lip can be due to the final rupture of the rock bolt occurring at the ends where maximum stress is located in this section of the reinforcing element. When a shear load is applied to the element, the greatest stress within the rock bolt is located in the same plane as the applied load where the element is subjected to a tensile and/or compressive stress at either extremity. This final rupture of the element due to the shear lip occurs predominately in the same plane where the shear load is applied, compared to the uniform smooth annular area formed adjacent to the free-surface of the element when subjected to a pure axial load.

Figure 12. View of the cupped failure surface.

To further analyse the failure mechanisms within the reinforcing element, scanning electron microscope (SEM) analysis was undertaken of the fracture surface of the failed element by the School of Materials Science and Engineering (UNSW). The two SEM results indicated the phenomenon of a dimpled rupture, which occurs via the process of microvoid coalescence. The two fractures started in the centre of the section of the reinforcing element and then radiated outward. Once the crack was near the surface the stress state changed from triaxial to plane strain and this was responsible for the change from flat face fracture that is perpendicular to the tensile axis, to slant fracture (45 degrees to the tensile axis) that produces the shear lip (Crosky, 2005).

conclusion

In summary, the main findings of the test program were as follows.

1.    A standard BX rockbolt exhibited a greater resistance to shear loading than had been anticipated; greater than both the ultimate tensile strength (UTS) and shear strength of the individual rockbolt element. The amount of shear displacement and deformation of the rockbolt was much greater than had been expected; nearly double that which had been allowed for in the initial design of the laboratory facility.

Failure loads of up to 400 kN were observed compared to typical UTS values of 250-300 kN.   

This result emphasises that behaviour of the complete system is not solely a function of the individual elements that make up the system such as the reinforcement elements. Rather behaviour is significantly influenced by the interaction that occurs between the system’s various components such as the rock reinforcement elements and the rockmass.

One potential ramification of this finding is that the extent of resistance to shear and the degree of deformation which is allowed for in the design of underground support systems may be well underestimated.

2.    Strength of the rockmass was shown to affect the performance of the system. In tests using higher strength concrete, the amount of shear displacement was less than that observed in comparable tests with weaker strength concrete samples. Stiffness of the system increased with strength of the rockmass. Conversely, maximum load resistance decreased with rockmass strength.

The stronger concrete is thought to have limited the extent of the “activation zone” along the rockbolt. Less crushing of the concrete about the rockbolt was observed in the stronger concrete samples indicating the material was less compliant.   

Hence in design of an underground support system, cognisance must be given to the strength parameters not only of the rockbolt but also of the rockmass. The result indicates that in endeavouring to design for a level of performance account has to be made of the rockmass, for example:

·      in strong rock, the support system is likely to be less compliant and stiff. The system is better able to maintain integrity of the laminated beds and hence contribute to overall stability.     

·      in weak rock, the system is likely to be more compliant and allow for more differential movement between bedding plains. Conversely the strength of the system would be enhanced as the rockbolt is capable of sustaining a higher resistance to shear load than can be achieved in a stronger rockmass.

3.    The performance of a rockbolt subjected to shear loading as characterised in a plot of applied shear load versus shear displacement demonstrated two distinct zones of behaviour. Initially, the system was relatively stiff with resistance increasing dramatically with very little shear displacement up to some level of load beyond which yielding was observed until the rockbolt eventually failed.

4.    Stiffness of a system is unaffected by cyclic loading. In earlier tests, with the limited shear displacement capacity, load on the sample had to be temporarily withdrawn to allow packers to be installed and the load was re-applied. The load-displacement curve was found to follow a similar path as in continuously loaded tests.     

Hence cyclic loading and unloading to less than the yield point is unlikely to impact performance of the rockbolt support system.

5.    Over the range of loading rates examined, stiffness of the system varies with the rate of load application; higher loading rates result in greater stiffness.

6.    Examination of the fracture surfaces of the failed rockbolts showed the rockbolts failed in a typical ductile manner and not in a manner usually associated with shear failure. Failure was initiated in the centre of the necked region of the element with cracks radiating outwards towards the surface. The fracture was completed via a shear lip on the outer extremities of the element.

7.    Although a rockbolt may have failed axially at the rockbolt/resins interface, it may still be capable of offering appreciable resistance to shear loading and hence provide some support to the rockmass.     

In tests where no face plate and nut were used at the collar of the borehole, the rockbolt generally could not be made to fail. At some point during a test, the limited length of encapsulation in the shorter block was insufficient to react against the axial load generated in the rockbolt resulting in failure of the rockbolt/resin interface. With continued shear displacement between the two test blocks, the level of resistance remained constant as the rockbolt was extruded through the borehole as it was not fixed or constrained by any face plate.

8.    Pre-tensioning of a rockbolt increased its initial resistance to shear displacement. When pre-tensioned, a rockbolt initially exhibited a high level of stiffness. With continual loading, a point was reached when shear displacement increased with load at a rate similar to that observed in untensioned elements.

The magnitude of load necessary to initiate shear displacement increased with the level of pre-tension.

Hence pre-tensioning is beneficial to increasing the stiffness of a rock support system dependent on the level of pre-tensioning.

9.    Use of strain-gauge rockbolts confirmed that shear loading generated an axial tensile load in the rockbolt, effectively clamping the shear surfaces together. The level of axial or normal load increases with shear load. This resultant normal force activates the frictional forces between the two rock surfaces that enhance resistance to shear loading.

When the orientation of the strain gauges was aligned with the shear plane, failure of the rockbolt was initiated in the corner of the one of the longitudinal slots of the strain-gauged rockbolt where the bending stress is at a maximum. Here plastic hinges are created that fractured the rockbolt.

Acknowledgements

The authors acknowledge the support of the Australian Coal Association Research Program (ACARP) for funding the research project. The authors also wish to thank the contributions made by Peter Craig, Jennmar Australia; Adam Raine and Troy Robertson, Hydramatic Engineering; Andrew Sykes, Minova Australia; Robin Genero, Sandvik; A/Prof. Alan Crosky, School of Materials Science and Engineering, UNSW; Bill Terry, Paul Gwynne, Tony Macken and Frank Sharpe, School of Civil and Environmental Engineering, UNSW; and, Bruce Hebblewhite and Wouter Hartman., School of Mining Engineering, UNSW.

References

Hartman, W and Hebblewhite, B, 2003. Understanding the performance of rock reinforcement elements under shear loading through laboratory testing: a 30- year history, in Proc.s 1st Aust. Ground Control in Mining Conference, Sydney, November, pp 151-160 (UNSW).

Crosky, A, 2005. University of NSW. Personal Communication