ABSTRACT
This paper describes the design,
construction and commissioning testing program for a laboratory facility at the
School of Mining Engineering, UNSW, Australia, aimed at evaluating
reinforcement tendon performance when subjected to shear loading. The project
has been financially supported through the coal industry ACARP research
initiative. The single failure plane design adopted in the test rig has been
successful in allowing shear loading to be directly applied to fully installed
(and where relevant, pre-tensioned) rock bolts. Initial results are extremely
encouraging, and illustrate a significant axial component of load development
as shearing takes place – particular in lower strength, softer material types.
The paper also outlines future plans for use of this facility.
INTRODUCTION
The Australian coal industry produces in
excess of 350 Mtpa of black coal, resulting in over 270 Mtpa of saleable
product, of which 73% is exported. This dominance of export markets results in
a priority focus on consistent, reliable supply of product, together with tight
constraints on mine operating costs.
product, of
which 73% is exported. Approximately 27% (94 Mtpa) of the total national
production is mined from underground mines, predominantly (78%) from retreating
longwall operations. There are currently 24 producing longwall mines in Australia. Typically, each of these mines would develop in excess of 10 km of development
drivage per year. This longwall focus further drives the need for efficient,
safe and cost-effective rapid and stable roadway drivage. Inadequate or
inappropriate roof support systems can lead to unsafe roof conditions
threatening both personnel safety and production capacity.
ently 29
producing longwall mines in Australia. This longwall focus also drives the
need for both rapid and stabl
Throughout the Australian underground
industry, roof support in both primary and secondary development roadways is
achieved through the use of fully-encapsulated resin-anchored bolts installed
as primary support, together with varying combinations of plates, straps and
steel mesh, usually installed at or near the face during a cyclic in-line cut
and bolt development system. This primary support is generally made up of high
tensile strength X-bar 22mm diameter bolts, and is often pre-tensioned.
Primary support density can vary from as few as 2 or 3 bolts per metre of
roadway, to in excess of 15 bolts per metre (including rib bolts). In
addition,
where required (such as in major roadways
and intersections, or regions of poor ground conditions or additional
mining-induced stresses (such as tailgates)), secondary support is also
installed in the form of cable bolts and other longer bolts and tendons, plus
standing support, where necessary.
A current research project being
conducted by the School of Mining Engineering at The University of New South
Wales (UNSW) involves the development of laboratory facilities to evaluate the
installed performance of bolts acting in shear resistance across different
types of rock mass discontinuities. The effect of different levels of
pre-tensioning on bolt shear performance will be included in the project
testing program. This work is considered of great significance in improving
the knowledge of how rock bolts perform when subjected to shear loading – a
condition that is prevalent in most laminated coal mine roof strata. An
initial description of the project objectives and early design concepts,
together with a background review of previous similar work was provided by
Hartman & Hebblewhite [1].
Rock reinforcement tendons – be they
rock bolts or cables – provide a significant proportion of their ground control
capability through offering resistance to shear movement of adjacent rock
masses or blocks. This potential shear movement may take the form of sliding
on horizontal bedding planes leading to strata bending; or block displacement
along other geological structures such as joints or similar discontinuities.
The shear resistance offered can be far greater than simply the shear strength
of the tendon material (steel), since the tendons can play a major role in
clamping the adjacent blocks of rock together – leading not only to
mobilisation of the frictional properties of the adjacent rock surfaces, but
also applying a normal force across the discontinuity to further increase the
frictional resistance.
Much has been reported about this type
of behaviour of rock bolts and other tendons, in theoretical concepts.
However, there is a shortage of quality data available on the exact nature and
quantitative significance of this mechanism for shear resistance, and the role
played by parameters such as pre-tensioning. A clearer understanding of the
nature and significance of this type of behaviour has major implications for
rock reinforcement tendon materials and installation design.
LABORATORY TESTING OBJECTIVES
The objectives of this research project
are aligned within the initial objectives of the ACARP project C12010 ‘Mechanical
Behaviour of Reinforcement Elements Subjected to Shear’.
The objectives of this research project
were:
1.
To research the current
understanding of the performance of reinforcement elements in shear.
2.
To design and develop a
testing rig which meets the need of the required testing.
3.
To conduct a series of controlled
laboratory experiments using the facility, and to study the effect of the
following variables on the performance of reinforcement elements in both direct
shear resistance and indirect shear resistance through axial clamping:
o
geomechanical properties
of test block
o
element pre-tensioning
o
applied loading rate
SHEAR TESTING FACILITY DESIGN
During the initial stages of the testing
facility design the key focus was on replicating an actual underground mining
excavation and identifying the key features of reinforcing elements in
resisting shear deformation. This facility (which incorporates the actual
installation of reinforcing elements in a block or bedded rock mass) steers
away from the common guillotine testing facility which can determine the direct
shear strength of the tendon material, by more closely representing the
interaction of the rock mass and reinforcing elements, which have a major
influence on the stability of the underground environment.
An Avery-Denison compressive testing
machine, previously used to test rock mass cylinders and cores (Figure 1) and
capable of loads up to 3600 kN was converted into the required shear testing
facility.
Located within the School of Mining Engineering at the University of New South Wales, this new testing facility provides a
controlled environment to allow for future evaluations on installed reinforcing
products and/ or concepts under shear loading to be undertaken.
In order to replicate a typical
underground mining environment, objectives for the design of the shear testing
facility included:
§
Determining the shear
displacement along an anticipated plane of weakness and final deformation
§
Inducing loading at
right angles
§
Inducing loading at
acute angles
§
Determining the axial
load distribution along the reinforcing element as a result of the shear load
and displacement
§
Determining the stress
distribution around the reinforcing element (within the concrete/rock mass)
Figure 1 – Avery-Denison testing
facility
In order to achieve the above objectives,
two separate concrete specimens were cast and consequently coupled together through
the installation of a reinforcing element. A rock mass of known material
properties and strength was recreated and cast from concrete. A smooth single
dominant joint surface was created during the casting of the concrete, to which
the shear loads could be applied. The purpose of recreating a rock mass through
casting concrete was to control the rock mass material and accurately measure
the mechanical and material properties of the concrete. The joint plane that
was created with the use of concrete can be controlled and altered by changing
the location, size, roughness and other properties of the joint.
The concrete specimens were cast in an
8mm thick steel RHS casing. The main purpose of this cast was to fully enclose
the concrete specimens to prevent splitting of the concrete rock mass once a
shear load was applied.
FACILITY CONSTRUCTION AND COMMISSIONING
A data acquisition system (DAQ) was introduced within the
shear testing facility. This DAQ consists of a high frequency multi-channel
monitoring system capable of measuring multiple arrays of 64 channels (Figure
3).
Figure 3 – Data Acquisition
System (DAQ)
The system is based on a personal
computer and DASYLab software. Sampling rates can be taken at up to 1,000
readings per
second with a typical experiment containing
between 5,000 and 10,000 data sets. The DAQ was configured to record the
voltage output of the pressure transducers from two channels and up to three
channels for displacement readings from the LVDT transducers. The DAQ is able
to provide the data to create load versus displacement curves (shear and axial)
and allow accurate calculations of loading rates within each test.
Pressure transducers introduced within
the testing facility record the pressure applied by the piston onto the
reinforced sample (later in the testing program an additional pressure
transducer was required to measure the axial load that was induced in the bolt between
the face plate and concrete surface). This pressure transducer, based on strain
gauge technology consisting of a one piece stainless steel body, has a capacity
of 400 bar and provides an output of 0 to 5 volts. The hydraulic load cell
(with the installed pressure transducer positioned within the two steel plates
prior to testing) can be seen below in Figure 4.
The DAQ system was upgraded during the
stage 3 testing program to allow for the data acquisition of strain-gauged rock
bolts. Three custom strain-gauged rock bolts were manufactured specifically
for this and created from the same batch of rock bolts used in the stage 2
tests.
Stage 3 testing included nine pairs of
diametrically opposed strain gauges being mounted within the rock bolt to
provide both axial and bending strain profiles either side of the shear plane.
The locations of the stain gauges were determined to gain readings as close to
the shear plane as possible. Figure displays the dimensions of the installed rock
bolt in the rock mass and the locations of the strain gauges in relation to the
shear plane.
Figure 4 – Application of the hydraulic load cell and
pressure transducer system
Each strain-gauged rock bolt was
calibrated prior to installation and testing of the reinforced sample. Even
though the rock bolts had nine pairs of strain gauges along their length, the
DAQ system had a limitation of only reading eight pairs at a time through two
AMX interface cards. The furthest positioned pair of strain gauges was ignored
due to the current limitations of the DAQ system.
The final set-up of the shear testing
facility can be seen below in Figure 5. Figure 6 displays the dimensions of
the installed rock bolt in the rock mass and the strain gauge locations,
relative to the shear plane.
Figure 5 – Shear testing facility
PRELIMINARY TESTING PROGRAM
Test Program
Tests were undertaken in
three stages. Stage 1 included two test samples that were used primarily to
determine the functionality of the testing facility and provide initial results
and scope for further modifications/adjustments to the facility. Stage 2
introduced additional instrumentation and modifications to the testing facility
and testing was undertaken on six reinforced samples. Stage 3 saw shear testing
performed on six reinforced samples and included the installation of three
strain-gauged rock bolts.
Installation of the reinforcing elements
was undertaken by an ARO roof bolter, commonly used in underground coalmines
throughout the world (Figure 7 and 8). Using a roof bolter to install the
reinforcing elements raised minor issues in the repeatability of installation
practices with variations in the borehole size and applied pre-tension across
some of the reinforced samples.
Figure 7 – Drilling borehole in concrete sample 1
Figure 8 – Drilling borehole in concrete sample 2
Once the samples were drilled and
reinforced they were transported back to the laboratory ready for testing and
analysis (Figure 9).
Figure 9 – Reinforced sample prior to testing
Initial Test Results
With no face plate installed on the
reinforced sample during the testing of the stage 2 samples, each sample was
loaded to a maximum of four cycles due to the maximum allowable shear
displacement within the shear testing facility (40mm). The load-displacement
curves that were generated during each test exhibited similar characteristics -
a much stiffer initial loading phase and a distinct transition point where the
stiffness of the curve dramatically decreased and remained relatively constant
until the load was removed from the system. This characteristic curve was
repeated in subsequent loading cycles.
Even with a faceplate installed on the
bolt collars during the stage 3 testing program the load-displacement
characteristics were similar across all three stages of testing. Figure 10
shows the typical load-displacement characteristics that were recorded during
the shear testing process. The top line indicates the applied shear load over
the duration of the test until failure, typically achieving peak shear loads of
up to 400 kN. (This is more than double the shear load required to cause direct
shear failure of the actual bolt steel). There is evidence of yielding, just
prior to the failure of the rock bolt. The bottom line indicates the
load-displacement measured by the load cell at the collar of the borehole
between the faceplate and the rock mass surface (250mm from the shear plane).
This result also indicates the presence of an initial pre-tension in the bolt
(approx. 40kN) applied to the reinforced sample. The shearing process has
clearly resulted in an increase in the axial bolt tension. The axial bolt load
line trends in three phases, with the load finally remaining relatively
constant at 160 kN until failure.
The reinforced samples that were not
confined with a faceplate and nut did not fail within the maximum allowable
shear displacement. This was due to the rock bolt slipping within the
encapsulation and subsequently not having the additional confining pressure of
the faceplate and nut at the borehole collar.
The reinforced samples confined with a
faceplate and nut at the collar of the borehole all failed at or above the
ultimate tensile strength of the rock bolt itself. Figure 11 shows the side
profile of a failed rock bolt.
Figure 10 – Typical shear-displacement characteristics
of a reinforced sample
Figure 11 – Side profile of failed rock bolt subject to
an applied shear load
A preliminary series of tests were
conducted using strain-gauged bolts, to assess the load distribution away from
the shear plane. Unfortunately, initial problems were encountered with
installation and orientation of the bolts and gauge alignment. The strain-gauged bolts were connected to
the DAQ system as shown below in Figure 12.
The strain gauge readings
were very volatile and did not produce reliable quantitative results within
each test. Interpreting the strain gauge results in relation to shear
displacement produced strains in excess of +/- 15,000,000 microstrain. The
strain-gauged rock bolt calibrations only attained approximately 500
microstrain after 100 kN of applied axial load during the calibration of the rock
bolts. This discrepancy in magnitude between the calibration results and actual
testing results removed the validity of the strain gauge data. The problem was
believed to be due to the DAQ hardware recording system. An outside source of
‘noise’ within the laboratory was also evident during testing, contributing to
the inconclusive results.
Figure 12 – Connection of strain-gauged rock bolt to DAQ
Although the quantitative results are
invalid, the correlation between strain (within the rock bolt), applied shear
load and shear displacement can be visually interpreted to analyse the position
on the load-displacement curve that strain in the rock bolt changes and how
this strain develops within the rock bolt during the test. Figure 13 (below)
also depicts the testing phase.
Figure 13 – Strain and load-displacement graph for
sample S3-6
Failure of the reinforcing
element
All the reinforcing elements that failed
due to an applied shear load, did so in a ductile manner. This final ductile
failure of the rock bolt occurred between the two plastic hinges (bending
regions) that were formed due to the shear displacement of the two concrete
blocks (where bending moment is greatest). Between these two hinges, the
reinforcing element was subject to an axial load causing the element to fail
axially, in tension (hence the bolt failing at shear loads well in excess of
the steel shear strength), see Figure 14.
Figure 14 – Side profile of a typically failed
reinforcing element subjected to a shear load
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, 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 (Figure
15). The shear lip is negligible at the sides of the element where the applied
shear load is perpendicular to the element.
Figure 15 – Typical end profile of a failed reinforcing
element subjected to a shear load
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.
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).
ANALYSIS OF RESULTS – INITIAL FINDINGS
The design and construction of the shear
testing facility has enabled quantitative shear testing of installed
reinforcing elements under controlled conditions. Numerous modifications and
alterations were implemented throughout the commissioning project to achieve a
testing facility to isolate and analyse the parameters that can influence the
shear performance of a reinforced rock mass.
The maximum load that was attained in the
reinforced samples that were not confined with a face plate and nut was
approximately 160 kN. This applied load is the load at which the resin
encapsulation fails due to the increased shear stress along the resin and
concrete rock mass interface. The hydraulic load cell that was installed at the
collar of the borehole for sample S3-2 onwards recorded maximum axial load
values between 150 and 164 kN, which can also be correlated to the load that
the resin encapsulation starts to fail and this maximum load is maintained
linearly over the duration of the test until failure of the rock bolt.
The introduction of strain-gauged rock
bolts to analyse the strain throughout the rock bolt (either side of the shear
plane) changed the failure mode of the rock bolts. The slot that is created in
the rock bolts to accommodate the gauges creates a weakness point and
subsequently intersects the zone of maximum bending stress, initiating failure
at this point at relatively low applied loads compared to the standard rock
bolt. This failure point is located within one of the plastic hinges that are
created during the shearing process at the zone of maximum bending moment.
With a higher pre-tension applied to the
reinforcing element, the load-displacement curve of the bolt under applied
shear load developed a higher stiffness until reaching a transition point on
the curve. The transition point is where the curve gradient decreases and
maintains a consistent residual level of reduced shear stiffness until failure
of the rock bolt. Initial results indicate that this transition point was
attained at a much lower shear displacement when an increase pre-tension was
applied to the rock bolt, i.e. higher initial bolt shear stiffness due to
pre-tension. The stiffness post transition point is similar for reinforced
samples with or without pre-tension applied.
The results also confirmed that by
activating a finite length of bolt either side of the shear plane to which the
loading is applied, the bolt bends and ultimately fails in an axial tensile
failure. The shear load at failure is significantly higher than the load
required to exceed the direct shear strength of the steel in the bolt
(typically about 50% of the tensile strength). This type of behaviour is very
dependent on both the strength and stiffness
of the surrounding rock or concrete material.
CONCLUSIONS
The various conclusions are as follows:
·
Standard rock bolts
installed in the concrete rock mass can offer a shear resistance more than
double the shear strength of the steel, and in fact also higher than the
ultimate tensile strength (UTS) of the rock bolt steel itself due to the
friction between the two specimen surfaces. There are two distinct loading
characteristic stages generated from the applied shear load and shear
displacement curve. The initial loading regime is much stiffer until reaching a
transition point, beyond which the load-displacement stiffness is reduced and
maintained until failure of the rock bolt. Even upon removing the applied shear
load and reapplying a shear load, the subsequent load-displacement curves
follow this same loading regime.
·
At relatively higher
loading rates the stiffness of the shear load-displacement curves is much
higher compared to the tests with a relatively lower applied loading rate.
·
A relative stiffer and stronger
rock mass material caused the rock bolt to fail within a lower shear
displacement compared to a relatively softer and weaker material. This was due
to the reduced bolt length “activation zone” in the stronger rock mass with
minimal crushing around the extent of the borehole compared to the weaker rock
mass.
·
When there was no
confining pressure from a face plate and nut at the collar of the borehole, the
rock bolt was slowly pulled through the borehole at a constant applied load,
which indicates a failure between the resin-rock mass interface.
·
A pre-tensioned
reinforcing element was shown to prevent early shear displacement at higher
applied shear loads, which is beneficial in minimising initial shear movement
within the surrounding rock mass. Beyond this initial loading regime the
pre-tensioned element then reduces in stiffness to a consistent level until
failure of the element.
·
The strain-gauged rock
bolts indicate a dramatic increase of a positive axial load within the
reinforcing element, which is maintained consistently through the element until
failure. Due to minor issues with the installation and positioning of the
strain-gauged rock bolts a majority of the results were inconclusive, in regard
to load distribution along the bolt.
·
All the reinforcing
elements failed in a typical ductile manner. The failure initiated in the
centre of the necked region of the element with the crack, then progressed
laterally towards the edge of the element in the area known as the radial zone.
The fracture was completed via a shear lip on the outer extremities of the
element.
·
When the strain gauges
were orientated and aligned with the shear plane, failure of the rock bolt is
initiated from one of the gouges at a location where there is maximum bending
stress.
RECOMMENDATIONS
Further testing reporting and
verification is required to fulfill the objective of this ongoing project and
provide a stepping-stone for the greater objectives within the ACARP project.
More definitive testing is required to study the influential parameters, including:
·
Borehole and element
geometry
·
Element orientation
relative to discontinuity
·
Element and
encapsulation material geomechanical properties
·
Block geometry
·
Further element pre-tensioning
·
Characteristic of
discontinuity
·
Discontinuity aperture
Further theoretical, mechanistic and
computational studies are required to support the experimental program with the
final objective to prepare a set of industry guidelines for the application of
the reinforcement systems in discontinuous materials, with respect to their
performance in shear resistance.
Recommendations have been made to further
control the laboratory experiments and
minimise any variations in the
results, including manually installing the reinforcing element to prevent the
influence of variability in the drilling and installation process.
ACKNOWLEDGEMENTS
The authors would like to acknowledge the
following organisations and personnel for their support of this research
project.
·
ACARP for financial
support
·
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
REFERENCES
1. Hartman W & Understanding
the performance of rock Hebblewhite B reinforcement elements under shear
loading (2003) through Hebblewhite laboratory
testing: a 30-
year history. 1st Aust. Ground Control in Mining
Conf., Sydney, Nov. 2003, pp 151-160.
2. Crosky A (2005) Personal
Communication
Original version of manuscript published in Proceedings
of 24th International Conference on Ground Control in Mining, Morgantown, University of West Virginia, August 2005.
