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AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014 1
INTRODUCTION
Rock burst frequently occurs in a sudden or violent manner in
the excavation face or on a working panel of an underground
excavation at great depth. Although the evolution of a rock
burst is generally recognised as a process of crack initiation,
propagation and coalescence of internal fractures in the rock
masses, a comprehensive understanding of the mechanisms
of rock bursts have so far been illusive due to its sophisticated
and non-linear nature.
In recent years, rock burst phenomena have been extensively
investigated by many researchers (CAMIRO, 2005; He, 2009;
Kaiser, 2009) through a variety of experimental devices
and theoretical approaches. It is critical to understand the
phenomenon of rock burst, focusing on the patterns of
occurrence of these events so as to save costs and possibly
lives.
There are several mechanisms that originate rock burst. The
main source mechanisms are usually associated with local
underground geometry and the existing geology (Ortlepp
and Stacey, 1994; CAMIRO, 1995; Castro, Bewick and Carter,
2012; He et al, 2013). They normally occur in large-scale
mining operations, but also in civil works. The most common
phenomenon is strain bursting, although buckling and face
crashing may also occur. In addition, impact-induced rock
burst can occur in less stressed and deformed rock formations
due to blasting and excavation of adjacent cavities.
Some researchers have conducted experimental studies
on rock burst using uniaxial compression tests, combined
uniaxial and biaxial static-dynamic tests, true triaxial loading
tests and conventional triaxial unloading tests. In recent
years, some researchers have promoted the use of acoustic
emission (AE) technology in rock mechanics, and considerable
achievements have been made in the characterisation of rock
failure and rock burst mechanisms (Yang and Wang, 2005;
Zhao, 2006).
A new method for rock burst studies involves the use of a
modied true triaxial apparatus that can unload the samples
being tested on one surface (He et al, 2011). It includes the
development of a test system for simulating strain burst in
a laboratory under deep ground excavation conditions, rock
masses and an output in response to the external perturbed
forces.
In addition to understanding rock burst mechanisms,
controlling rock burst is the most important issue for the
safety of mining operations and other deep underground
Experiments on Rock Burst and its
Control
M He1 and L R Sousa2,3
ABSTRACT
Rock burst is common in deep underground excavations and is characterised by a violent ejection
of block rocks from excavation walls. It is critical to understand the phenomenon of rock burst and
particularly the failure mechanism. Laboratory experiments are one of the ways. For that purpose,
a rock burst laboratory test system was developed at the State Key Laboratory for Geomechanics
and Deep Underground Engineering at the China University of Mining and Technology, Beijing.
A true triaxial test system, designated the Deep Underground Rock Burst Analogue Test Machine,
was developed, which can transfer a stress state from three-direction-compression on six surfaces
to the compression on ve surfaces with one set free. The rock burst experimental system
comprises an acquisition of acoustic emission signals, a high-speed recording in order to accurately
record the kinetic characteristics of rock fragments ejected during a rock burst event and the
possibility of infrared thermography showing the surface temperature of the samples. In this
paper, analyses of typical results from rock burst tests are presented with a description of the
obtained results. The mechanisms of rock burst are illustrated, and a rock burst classication is
proposed based on these laboratory tests. For the control of rock burst, the constant resistance,
large deformation (CRLD) bolt or anchor is presented for accommodating the large displacement
of surrounding rock masses and absorbing the impact of the sudden release of rock burst while
outputting a constant resistant force in response to the external load. The CRLD bolt or anchor
has had practical applications in underground mines. Finally, in situ tests performed in a tunnel at
Qingshui coalmine, China, are analysed in detail.
1. Director, State Key Laboratory for Geomechanics and Deep Underground Engineering, China University of Mining and Technology, Beijing, 16 Tsinghua East Road, Haidian District, Beijing
100083, China. Email: hemanchao@263.net
2. Researcher, State Key Laboratory for Geomechanics and Deep Underground Engineering, China University of Mining and Technology, Beijing, 16 Tsinghua East Road, Haidian District, Beijing
100083, China. Email: sousa-scu@hotmail.com
3. Professor, University of Porto
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
2
structures. Bolts and anchors are efcient measures for the
control of rock burst in underground excavations, performing
different roles such as the reinforcement of the rock and to
hold the retained reinforced rock mass.
These supports were widely investigated by many
researchers worldwide (Kaiser, 2009; He et al, 2011). Among
them is the cone bolt, developed by a Canadian company.
This bolt does not have the ability to adjust itself to the
external loading to output a constant resistance. Its maximum
deformation was only 120 mm and it is unable to adapt to
the support requirements of larger deformation in mine
roadways (greater than 200 mm). Another compressible rock
bolt (Roofex) was developed by Atlas Copco in Australia,
which was suitable for supporting a soft rock tunnel and
capable of maintaining a constant resistance while the
supported rock mass underwent large deformations. This
type of bolt has a maximum extension value of up to 300 mm
and a constant resistance force of 80 kN. However, as mining
depths increase, the demands for anchor bolts with larger
extension and higher loading capacity are growing. This paper
introduces the state-of-the-art anchor and bolt technology, the
constant resistance, large deformation (CRLD) bolt or anchor,
developed by the State Key Laboratory for Geomechanics and
Deep Underground Engineering (SKLGDUE) at the China
University of Mining and Technology (CUMTB), Beijing. The
CRLD bolt or anchor has the ability to accommodate larger
deformations of the surrounding rock masses of tunnels at
great depth in response to external forces. Field tests were
performed and analysed in the last section of a coalmine in
China.
A LABORATORY SYSTEM FOR ROCK BURST
The mechanism of rock burst
The evolution of rock burst can be considered in a simplied
way by two stages (He et al, 2007): the stress evolution and
the plate structure evolution. Stress evolution refers to the in
situ stress state of the rock mass, in which the stress state is
transformed from a 3D six-surface state to a 3D ve-surface
state of stress due to an underground excavation (Figure 1).
The so-called plate structure evolution refers to the structural
response of the rock mass, which can be divided into the
following three phases (Figure 2):
1. cracking through the surface parallel to the excavation
2. buckling of the surface near the excavation
3. the rock burst ejection process.
The essence of the mechanical behaviour transformation
process dened by the rock burst is an alteration of the
mechanical behaviour of the rock under external conditions,
which has been proved by laboratory triaxial loading and
unloading experiments.
Experimental set-up
Laboratory modelling of rock burst under the conditions in
deep underground engineering is one of the major goals of
SKLGDUE at the CUMTB. To test this, the Deep Underground
Rock Burst Analogue Test Machine (DURAT M) was
developed in 2006 (Figure 3).
The testing system is a true triaxial testing scheme consisting
of three main parts: a loading/unloading device, a high-speed
data acquisition system and an AE detecting system. The
loading/unloading device of the DURAT M is comprised of a
main stand, hydraulic control apparatus and force-measuring
transducers, which can provide dynamic loading and
unloading independently in three principal stress directions.
During the test, one surface of the specimen can be unloaded
immediately from the true triaxial compression condition,
simulating the stress condition for rock mass at the free
excavation boundary in underground excavations. A series of
physical modelling tests on the rock burst phenomena was
conducted in the laboratory to calibrate the machine with real-
world conditions in deep mining and to make adjustments
to the physical and mechanical parameters of the DURAT M
system (He et al, 2007).
The DURAT M system is unique, with three main features
(He et al, 2012b):
1. Transition of the mechanical behaviour of the rock mass at
great depth can be simplied using this testing equipment.
2. Transformation of the stress state can be achieved with
the DURAT M system. Rock masses located in deep
ground are under compressive stress conditions in all
FIG 1 – Stress model for excavation-induced rock burst.
FIG 2 – Structural changes after rock burst.
FIG 3 – Experimental system for simulation of rock burst process at great depth.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
EXPERIMENTS ON ROCK BURST AND ITS CONTROL
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three principal directions. When an underground cavity
is excavated, a parallelepiped volume element of the rock
mass will transfer its compressive stress state in three
directions into a compression state with one free surface.
3. One surface of the specimen can be unloaded immediately
from the true triaxial compression condition. Ensuring
that one surface was free when unloading the existing
stress was crucial for the simulation of the in situ rock burst
phenomenon. The design of the Single Face Unloading
Device in the DURAT M was one of the difculties during
the initial period of experimentation (Figure 4).
In conclusion, the information measuring system for the
laboratory rock burst tests include data acquisition, AE
monitoring, high-speed image recording and, optionally,
infrared image recording. The data acquisition consists of
sensors, ampliers, data acquisition instruments, computers
and appropriate software that can automatically collect, edit
and process the test data. The AE released during the tests can
be monitored using an acquisition card, a continuous current
source, a preamplier and an AE sensor. The system is also
equipped with a high-speed digital camera that accurately
records the kinetic characteristics of the rock fragments that
are ejected during the rock burst tests. Finally, there is the
possibility of incorporating infrared thermography, which
shows changes in the surface temperature of the samples.
These changes in temperature allow us to better understand
the mechanism of rock burst.
CASES OF LABORATORY ROCK BURST TESTS
General
Since the rst rock burst laboratory tests were performed in
2006, more than 200 tests have been carried out on samples
from China, Italy, Canada and Iran by using the system
developed at SKLGDUE. About 11 rock lithologies have been
investigated (He et al, 2011).
A database of 139 cases was created from the rock burst
tests where complete information was obtained. A form was
elaborated and a great number of elds were considered,
including: location of the test, dimensions of the sample, rock
material, main minerals and cracks, stresses before loading,
stresses during tests, characteristics of rock burst test and
critical depth. The results were analysed and several data
mining techniques were applied in order to obtain models for
the maximum principal stress obtained in the rock burst tests
and for a rock burst risk index (He et al, 2014a).
Two examples are described in the following sections. One
relates to a test in Laizhou mine in the Shandong province,
China, while the other relates to a rock burst test performed
in a sample from Creighton mine, Canada.
Rock burst test at Laizhou mine, China
This section describes the results of a granite rock burst test
from Laizou mine in Shandong province, China (He et al,
2012a). The dimensions of the sample were 150 × 60 × 30 mm3.
X-ray diffraction analysis showed that the sample was
27 per cent quartz, 68 per cent feldspar and ve per cent clay
minerals.
The loading/unloading path is presented in Figure 5. The
in situ stresses were σ1 = 101.1 MPa, σ2 = 59.8 MPa and σ3 =
29.7 MPa. The stresses were kept constant for 15 minutes,
before the load on the surface of σ3 direction was removed
suddenly. The sample failed with stresses of σ1 = 129.3 MPa,
σ2 = 58.5 MPa and σ3 = 0.0 MPa, with the ejection of particles.
The accumulated AE energy release is shown in Figure 6.
Initially, AE energy releases at a low rate when the stresses are
kept constant. Later, the AE accumulated was characterised
by a sudden increase in the rst unloading followed by a
sharp increase with the occurrence of instantaneous burst.
The movement process of the rock fragments ejected from the
unloaded surface was recorded using the high-speed camera,
as shown in Figure 7. It shows the surface of the sample before
rock burst. The rst falling down is also illustrated, and when
the rock bursts occurred, a lot of fragments were ejected from
the upper region of the sample.
Granite sample from Creighton mine, Canada
Rock burst tests were also performed at Creighton mine,
which is located in southern part of Sudbury Basin, Canada
(Camiro, 1995; He et al, 2014b). Creighton’s sulde orebodies
are present in the lower sublayer of the hanging wall norites.
FIG 4 – Sketch of dowel steel and pressure head.
FIG 5 – Loading/unloading path for a granite sample.
FIG 6 – Accumulate acoustic emission energy release of the burst.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
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The footwall rocks are mainly granite. The overall geotechnical
parameters, estimated to a depth of about 2100 m, are the
following: Young’s modulus – 30 GPa; Poisson ratio – 0.25;
strength parameters – σt = 0 MPa, c = 22 MPa and φ = 35°.
The samples for the rock burst tests were obtained to a depth
of 2420 m. Figure 8 illustrates one of the tests for the sample
CG4# from the mine. The dimensions of the prismatic sample
were 97.9 × 42.0 × 20.6 mm3. In the rock burst test, taking into
consideration concentration factors, the in situ state of stress
was supposed to be equal to σ1 = 240.7 MPa, σ2 = 97.2 MPa and
σ3 = 90.6 MPa.
Figure 9 indicates the loading path of the granite sample
used for the rock burst test. The sample was rst loaded with
three principal directions loading step-by-step to simulate the
original stress state referred. Next, the horizontal minimum
principal stress on one surface of the sample was unloaded,
and the rock burst process was simulated following the stress
path indicated in Figure 9. The critical rock burst stresses
were then obtained with the following values: σ1 = 282.1 MPa,
σ2 = 93.3 MPa and σ3 = 0.0 MPa.
MECHANISMS OF ROCK BURST AND ITS
CLASSIFICATION
Indoor rock burst tests play an important role in understanding
the mechanisms of rock burst, the calibration of numerical
models, the evaluation of mechanical parameters and the
identication of the stress state when a dynamic event may
be initiated. Laboratory tests have been used by several
researchers and include uniaxial compression tests (Badge
and Petrosav, 2005; Zuo et al, 2006), combined uniaxial and
biaxial static-dynamic tests, true triaxial loading tests (Chen
and Feng, 2006; Cheon et al, 2006) and conventional triaxial
unloading tests (Xu, 2003).
However, the rock burst simulation tests could not provide
correct in situ stresses on the near-face region during
underground excavation. With respect to the triggering
mechanisms, rock bursts may occur under high in situ stress
conditions. For the surrounding rocks in lower stress states
due to external disturbances (such as blasting, caving and
adjacent tunnelling), rock bursts can also be triggered. In this
paper, rock bursts are classied into two major types: strain
bursts and impact-induced burst, as shown in Figure 10.
Strain bursts are frequently encountered during tunnel
excavation, and they are also associated with pillar and room
mining cavities. According to different stress paths and failure
locations, strain bursts can be divided into three subtypes:
1. instantaneous burst
2. delayed burst
3. pillar burst.
After excavation, the surfaces of the cavities and pillars may
also suffer rock bursts due to the impact waves generated
by mining disturbances. According to their formation
FIG 7 – Analysis of the burst process.
FIG 8 – Rock burst test for sample CG4# from Creighton mine.
FIG 9 – Rock burst loading path for the sample from Creighton mine.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
EXPERIMENTS ON ROCK BURST AND ITS CONTROL
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mechanisms, impact-induced burst can be divided into three
subtypes:
1. rock bursts induced by blasting or excavation
2. rock bursts induced by roof collapse
3. rock bursts induced by fault slip.
Rock burst experimental equipment was briey presented
in the previous section. A new experimental system was
developed in order to simulate impact-induced burst
(Figure 11) (He et al, 2012a). It consists of a main stand,
servo-controller and hydraulic power. As indicated in He et
al (2012a), it can generate various types of disturbance wave
signals.
The following types of strain bursts can be simulated by this
testing system:
•Instantaneous burst – one surface of the sample is
unloaded suddenly from a true triaxial stress state to
simulate the strain bursts immediately after excavation.
A schematic diagram of the loading/unloading path is
shown in Figure 12a.
•Delayed burst – one surface of the specimen is unloaded
suddenly from a true triaxial stress state and then the
vertically-imposed stress (σ1) is increased based on the
stress concentration to simulate the strain bursts that occur
sometime after excavation due to stress redistribution.
A schematic diagram of the loading/unloading path is
presented in Figure 12b.
•Pillar burst – due to excavation, the pillar size decreases,
thus increasing the vertical stress. The horizontal stresses
(σ2 and σ3) can be gradually decreased to simulate the
formation of a pillar until the burst occurs. A schematic
diagram of the loading/unloading path is illustrated in
Figure 12c.
Rock burst induced by blasting or excavation can also be
considered. Firstly, the static load stresses are applied on
the sample to simulate the in situ stress state. Secondly, the
disturbance wave is loaded in one, two or three directions and
the burst phenomena is observed, in which the disturbance
load is used to simulate site excavation, blasting, earthquakes
or a mechanical vibration waveform. The schematic diagram
of the loading/unloading path is presented in Figure 13.
Different rock burst criteria were analysed in detail by He
et al (2012a).
Figure 14 shows three different stress paths for strain
bursts using Hoek–Brown criterion for the strength of the
rock mass (Sonmez and Gokceoglu, 2006). Figure 14a shows
the instantaneous burst path. The area Z1 is the potential
zone for the occurrence of rock burst of this type. Point A
represents the initial stress state before excavation. σc and
σr are the uniaxial compressive strength and the long-term
FIG 10 – Laboratory experimental methods based on
rock burst classication (He et al, 2012a).
FIG 11 – Illustrations for the new deep rock nonlinear mechanical system.
A
B
FIG 12 – Loading paths for strainburst.
A B C
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
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peak strength respectively. Instantaneous burst occurs with
the release of σ3 if the maximum principal stress σ1c is greater
than σc. Figure 14b shows the delayed burst path. Area Z2 is
the potential burst-prone zone of this type. Point B represents
the initial stress state before excavation. As σ1 is lower than σc,
the delayed burst will not occur when σ3 is suddenly released
unless there is enough energy that can be released in the form
of kinetic energy or other. The increase of σ1 may be attributed
to the tangential stress concentration due to excavation and
to the damage of the surrounding rocks by eld engineering
disturbances, such as excavation and blasting. Figure 14c
shows the stress path for pillar burst. The areas Z1 and Z2
can be the potential burst-prone zones of the pillar. Points C1
and C2 represent the initial stress states in the areas Z1 and Z2
before excavation respectively. Increasing σ1 and decreasing
σ3 will result in the occurrence of pillar burst.
Figure 15 shows schematic diagrams of impact-induced
burst criteria. Theoretical considerations about the criteria
used for the situations of impact by blasting, excavation, roof
collapse or fault slip are analysed in detail in He et al (2012a).
CONSTANT RESISTANCE LARGE
DEFORMATION BOLTS
Bolts are a major method used in rock support. However, in
high-stressed rock masses, bolts frequently break because
they cannot adapt to large deformations. A bolt with constant
working resistance and steady large deformation, under large
deformations and impact loads, was developed at SKLGDUE
(He et al, 2011). The CRLD bolt or anchor device consists of
two parts, the constant resistance element and the bolt rod
(Figure 16). The constant resistance element is composed of a
slide track sleeve and a constant-resistant body.
Figure 17 illustrates the supporting principle of the CRLD
bolt over the following three different stages (A is the
anchored segment, B is the wall of the drilling hole, C is the
constant resistance element and D is the bafe plate):
1. Elastic deformation (Figure 17a) – the deformation energy
of the surrounding rocks could be converted to the bolt
rod in the bolt assembly through the bafe plate and inner
anchorage segment. In the case that a relative deformation
for the surrounding rocks and the axial force loaded
by the rock deformation is less than the rated constant
resistance for the CRLD bolt, the bolt will not elongate by
the displacement of the constant-resistance element, but
will resist to the deformation and failure of the rock, solely
relying on the elastic deformation of the bolt rod itself.
2. Structural deformation stage (Figure 17b) – with the
deformation building up, the axial force on the build rod
will be increasing and may be equal to or larger than the
rated constant resistance for the CRLD bolt, leading to the
frictional-sliding displacement of the constant-resistant
body along the sleeve track (ie the CRLD bolt elongates).
While elongating, the bolt will be keeping the constant-
FIG 13 – Loading/unloading path for rock burst
induced by blasting or excavation.
FIG 14 – Schematic diagram of strain bursts criteria.
A B C
FIG 15 – Schematic diagrams of impact-induced burst criteria.
A B C
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
EXPERIMENTS ON ROCK BURST AND ITS CONTROL
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resistant characteristics and resist the deformation and
failure of the rock mass by its elongation (ie the structural
deformation of the constant-resistant element).
3. Ultimate deformation stage (Figure 17c) – after undergoing
the elastic deformation of the bolt rod itself at the rst stage
and large deformation that equals the elongation of the
constant-resistant element, the deformation energy for the
rock mass in abutment to the excavations has been fully
released. In this case, the external load will be smaller than
the rate of constant resistance, and the constant-resistant
body will stop sliding due to fractional drags. Therefore,
the surrounding rock mass for the excavation has been
stabilised.
Consequently, under conditions of deformation of the
surrounding rock masses, the bolt can absorb the deformation
energy of the rock mass, which will release the energy stored
in the surrounding rocks. Over the structural deformation
stage, the bolt is still able to elongate steadily while keeping its
working resistance constant in response to the external forces.
Thus, the stabilisation could be realised for the CRLD bolt-
supported rock masses adjacent to excavations, mitigating
potential disasters such as rockfall, collapse, slabbing and
splitting and oor heaving.
The development of the CRLD bolt or anchor has been tested
in situ, in the laboratory and in some practical applications.
The maximum extension of bolt rock for CRLD is about
1000 mm, which can fully accommodate the displacement
extent of the rock mass adjacent to the deep underground
excavations. In comparison with currently existing large-
deformation anchors from abroad, this novel bolt has much
longer extension length under the same external pulling
force. At the same time, its maximum load-carrying capacity
is much larger, as illustrated in Figure 18.
Static and dynamic tests were developed at SKLGDUE for
this type of bolt.
The aim of the static tests was to evaluate the static
characteristics of the CRLD bolts, including the constant
resistance, magnitudes of the resistance and maximum
elongation extent. The equipment is illustrated in Figure 19.
The nominal length of the test CRLD bolts is 1 m, the diameter
of the bolt rod is 22 mm and the inner diameter of the slide-
track sleeve tube is 34 mm. The bolt was fastened on the two
ends of the testing machine by the holding device (Figure 20).
Proles of the resistance against displacement for a 20 t
CRLD bolt and two anchors (35 and 85 t) are indicated in
Figure 18. These are compared with the Canadian cone bolt
and the Australian Roofex bolt, with maximum displacements
of about 120 mm and 300 mm respectively.
Dynamic impact tests are also performed at SKLGDUE.
The testing system can be used to evaluate the performance
of resistance and absorbing impacting energy for CRLD bolts
or anchors by measuring the extension length of the bolt
(anchor) rod body itself and the radial deformation of the rod
FIG 16 – Layout of the constant resistance, large deformation bolt.
FIG 17 – Supporting principle for a constant resistance, large
deformation bolt. (A) elastic deformation stage; (B) structural
deformation stage; (C) ultimate deformation stage.
A
B
C
FIG 18 – Comparison among dierent bolts or anchors.
FIG 19 – Static tension test set-up for the constant
resistance, large deformation bolt.
FIG 20 – Detail of the experimental set-up of the static tension testing system.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
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body. The testing system is shown in Figure 21, and has the
following technical specications:
•maximum impact energy – 15 000 J
•effective range of the impacting height – 0–1.5 m
•hammer height with ve grade – 840–1000 kg
•allowed diameter of the bolt – 34 mm and 22 mm
•length – 1500 m and 2500 m.
The purpose of the dynamic impact test is to evaluate
the energy absorbing capacity for the bolts under dynamic
loading. Before the impact, the testing bolt was fastened to
the test machine with the holding device. The bafe plate was
then installed and the initial displacement was set up. After
setting the maximum dropping height and conrming that
all the procedures were ready, the test was conducted with
a cyclic-loading scheme until the completion of all the cycles.
Experimental results are illustrated in Figure 22, which
shows the time domain curves of impacting load versus
deformation of bolt under dynamic loading with an impacting
height of 10 mm. Figure 22a shows the whole time domain
waveform gure with an impacting height of 10 mm for a bolt
and Figure 22b shows the time domain waveform gure after
rst impact.
Advancements in Hopkinson dynamic tests were also
developed for the CRLD bolts. Experimental studies with one
and two bolts are still being analysed, complemented by the
use of numerical models using LS-Dyne software.
IN SITU TESTING RESULTS
General
In situ tests with CRLD anchors were performed at Hongyang
coalmine in Liaoning province, China, using a blasting method
in order to simulate rock burst that had occurred frequently
in this mine. An abandoned roadway tunnel was used for
the location of the tests at a depth of about 780 m (Figure 23).
The geological stratigraphy near a section of the roadway
is illustrated in Figure 24, showing that there is a mudstone
layer above the tunnel and medium sandstone layer below.
The average thickness of the coal seam is about 2.6 m.
The roadway tunnel was used for ventilation purposes and
has a rectangular cross-section (Figure 25). The supporting
conditions were generally good, with a stable roof, and the
tunnel was over 100 m from other tunnels. In situ program
tests were performed in the roadway tunnel and included
two kinds of tests using different support methods. The rst
type of support contained normal bolts of 2.2 m in length and
20 mm in diameter, anchors of 6.5 m in length and 21.7 mm in
diameter and wire mesh. The second type added 35 t CRLD
anchors with the same length of 6.5 m instead of normal
anchors. These supports were submitted to impact tests. The
total length of the testing area was 40 m, which was divided
into four sections of 10 m each (Figure 26).
Sections I and IV used the rst type of supports, while
sections II and III used supports with CRLD anchors.
Chambers were excavated between sections I and II and
between sections III and IV, with dimensions of 4.5 m in
length, 3.0 m in width and 2.0 m in height. Figure 27 presents
an overview of the testing method. Figure 27a presents a
side view of the testing area, while Figure 27b provides a
view from up to down as well as the chambers (caves). In
the chambers, two blastholes of about 7–8 m in length were
drilled in both sides in the coal seam for the placement
of the explosives, as illustrated in Figures 27 and 28. The
explosive loads were 4.6 kg and 6.0 kg, which approximately
represented the seismic events occurring in this mining area.
In the chamber between sections III and IV, the explosives
were applied again, therefore the explosives were doubled
with two times blasting.
Monitoring and results
A monitoring plan was established with the purpose of
manually measuring surface displacements at the roadway
tunnel in the four sections. Later, a real-time monitoring
system, photographs, forces and elongations for normal and
CRLD anchors were used. The real-time monitoring system
used BeiDou satellites, and the information was transmitted
in real-time to SKLGDUE in Beijing (Figure 29).
FIG 21 – Dynamic impact test set-up.
FIG 22 – Time domain waveform gure with an impacting
height of 10 mm: (A) the whole time domain waveform; (B)
time domain waveform gure after rst impact.
A
B
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EXPERIMENTS ON ROCK BURST AND ITS CONTROL
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Figures 30, 31 and 32 are before and after photographs of
normal section I and the CRLD-supported sections II and
III respectively. In the normal section, it is clear that the
upper part of the roadway tunnel collapsed after blasting,
with failure of the existing supports. Following the blast,
this section of the tunnel was no longer available for safety
reasons. The other two sections supported by CRLD anchors
remained stable after the large deformations, with only the
steel mesh being broken.
Some monitored results were obtained. Figure 33 illustrates
measured displacements at sections II and III from the heads
of the CRLD anchors for the rst explosion. The minimum
values were about 25 mm and the larger values were almost
60 mm, but the average values were between 25 mm and
30 mm. Other measurements included forces for the types of
anchors.
FIG 23 – Hongyang Mine; location of the testing area.
FIG 24 – Hongyang Mine; geological formations near the roadway.
FIG 25 – Section of the roadway and location of constant-
resistance and large deformations anchors.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
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CONCLUSION
SKLGDUE has developed a true triaxial apparatus that is able
to reproduce the rock burst phenomenon in a laboratory. The
initial feature lies in a single-face unloading that simulates the
stress paths existing in the face of underground excavations.
The experimental set-up comprised several systems,
including devices for loading/unloading, high-speed
acquisition, AE detecting and infrared thermography. The
distribution law of the dominant frequency bandwidth for
different rock samples under varied stress paths during rock
burst simulation tests can also be obtained based on time-
frequency analysis and the discrete Fourier transformation
from the AE waveform data. The investigations performed
have laid a solid foundation for further development of the
state-of-the-art theories and experimental devices developed
by SKLGDUE. Also, a database with a large number of rock
burst tests was created. The information was analysed with
several data mining techniques in order to obtain models for
the maximum stresses obtained in the tests and for a rock
burst risk index.
The rock burst triaxial laboratory was adapted in order
to simulate impact-induced bursts by generating different
types of disturbance wave signals analogous to the impact
produced by the drill-and-blast method. Based on these rock
burst testing devices, a new classication for rock burst was
proposed, including the experimental procedures, stress paths
imposed and the released energy-based rock bursts criteria.
A
B
FIG 27 – Overview of the testing method.
FIG 26 – In situ testing area.
FIG 28 – Location of the blastholes and explosives.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
EXPERIMENTS ON ROCK BURST AND ITS CONTROL
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FIG 29 – Automatic monitoring transmission system.
FIG 30 – Photographs of a normal support section before and after blasting.
FIG 31 – Photographs of constant resistance, large deformation-supported section II before and after blasting.
AUSROCK 2014: THIRD AUSTRALASIAN GROUND CONTROL IN MINING CONFERENCE / SYDNEY, NSW, 5–6 NOVEMBER 2014
M HE AND L R SOUSA
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Control of rock burst is a very important issue. Therefore,
the CRLD bolt or anchor, which permits constant resistance
and large deformations, was developed at SKLGDUE. The
performance of these bolts was experimentally veried with
the development of static and dynamic equipment. In situ tests
were performed at a coalmine in Liaoning province, China.
The feasibility of the new bolt or anchor was successfully
veried, which is expected to have a signicant role in control
and the prevention of the occurrence of rock burst in deep
underground engineering.
ACKNOWLEDGEMENTS
This work was supported by the Key Project of Natural
Science Foundation of China (No 51134005) and the General
Program of National Natural Science Foundation of China
(Nos 40972196 and 41172263).
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