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3D Physical Experimental Study of
Shield–Strata Interaction Under
Dynamic and Static Disturbance
Shengli Yang
1
,
2
, Hao Yue
1
,
2
*, Ruihao Zhai
2
, Zhiwei Cui
2
and Xia Wei
3
1
State Key Laboratory of Water Resource Protection and Utilization in Coal Mining, Beijing, China,
2
School of Energy and Mining
Engineering, China University of Mining and Technology-Beijing, Beijing, China,
3
Department of Numerical Control Engineering,
Shanxi Institute of Mechanical and Electrical Engineering, Changzhi, China
With the increasing depth and intensity of coal mining, there is an increasing risk to the
working face due to high static load and periodic breakage of the roof. The relation
between the support and the surrounding rock under static–dynamic coupling loading
disturbance is an important factor affecting the stability of the working face. In this study, a
3D physical modeling platform is developed to study the interaction between the shield and
strata under dynamic and static disturbance. In the experiment, the static load is set to 0,
0.5, 1, and 2 MPa, respectively. The different dynamic load is realized by changing the fall
height of the iron plate. The change in hydraulic support resistance is recorded by the
pressure monitoring system. The displacement of the coal wall is monitored by using an
infrared rangefinder. The results show that the change in static load and dynamic load will
affect the support resistance and coal wall displacement. With the increase in dynamic
load, coal wall displacement, and bracket resistance increase, the increase is not linear.
The larger the dynamic load, the greater the increase. Static load change has little effect on
bracket resistance, and the impact on coal wall displacement is large. With the increase in
static load, coal wall displacement is reduced and then increased. In static load, for the
stability of coal wall, there is a threshold; below the threshold, the static load can improve
the stability of coal wall, exceed the threshold, but accelerate the destruction of coal wall. At
the same time, the stability coefficient of the quarry bracket and surrounding rock is
defined. The sensitivity analysis of the main parameters is carried out. The method of
controlling the stability of the quarry bracket and surrounding rock is proposed.
Keywords: static–dynamic coupling loading, support and surrounding rock relation, similar simulations, coefficient
of stability, quarry
INTRODUCTION AND BACKGROUND
The relationship between coal mine quarry supports and surrounding rock is an important part of
mine pressure theory. With the rapid development of all areas of coal mining (Deng et al., 2020;Deng
et al., 2022;Chen Jianhang et al., 2021;Chen Yulong et al., 2021), the understanding of the
relationship between quarry supports and surrounding rock is also in-depth. The quarry support is
an important part of the support system, and scholars at home and abroad have conducted research
on different types of support systems. Shiroma et al. (2004) studied the standardization of tunnel
support reduction patterns on the basis of the test construction results (Shiroma et al. (2004)). A
strong support system was established suitable for a squeezing tunnel for the purpose of addressing
Edited by:
Yulong Chen,
China University of Mining and
Technology, Beijing, China
Reviewed by:
Haiqiang Jiang,
Northeastern University, China
Yin Wei,
Huaiyin Institute of Technology, China
Weiqing Zhang,
China University of Mining and
Technology, China
*Correspondence:
Hao Yue
yuehaocumtb@163.com
Specialty section:
This article was submitted to
Structural Geology and Tectonics,
a section of the journal
Frontiers in Earth Science
Received: 06 April 2022
Accepted: 14 April 2022
Published: 18 May 2022
Citation:
Yang S, Yue H, Zhai R, Cui Z and Wei X
(2022) 3D Physical Experimental Study
of Shield–Strata Interaction Under
Dynamic and Static Disturbance.
Front. Earth Sci. 10:913903.
doi: 10.3389/feart.2022.913903
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139031
ORIGINAL RESEARCH
published: 18 May 2022
doi: 10.3389/feart.2022.913903
problems exhibited in the extreme deformation of rock mass,
structure crack, or even failure during the excavation phase
(Wang et al., 2019). Zhang C analyzed and discussed stress
characteristics and deformation patterns of the primary
support and secondary lining structures, and reveals the
influencing mechanisms of different tunneling modes on the
stress characteristics of the support and lining structures
(Zhang and Jiang, 2020). New methods and theories have
been emerging. However, in the case of coal mines (Li, et al.,
2019), the support system has also changed due to the influence of
mining. It is well known that the force on the support of the
quarry comes from the activity of the basic top and the direct top
rock layer (Qian et al., 1994). It is known that the incoming
pressure of the roof plays a key role in the mine roof accident. The
key to the incoming pressure of the roof lies in the mutual
coupling relationship between the breakage of the basic roof,
the direct roof, the support, the coal body, and the bottom plate.
Therefore, it is of great significance to study the relationship
between the bracket and the surrounding rock in the mining site
for the safe and efficient production of coal mines.
For this topic, scholars have carried out a lot of research,
supplementing and improving the traditional brace envelope
relationship (Cao et al., 1998). The relationship between the
working resistance of the stent and the amount of roof sinking
has been studied from an energy perspective (Gao et al., 1999). An
algorithm was proposed to consider the actual conditions of rock
mass and support interaction and the algorithm implementation
method to ensure efficient calculation of stresses in rock and
support (Seryakov and Kurlenya, 2018). Two-dimensional (2D)
continuum finite-element models of longwall panels were
developed and analyzed to understand the interaction of
hydraulic-powered support with surrounding rock strata
(Verma and Deb, 2013). The influence of the three-
dimensional stress on the surrounding rock supported with
different stiffness was studied (Wu et al., 2016). Stress
distribution laws of surrounding rock plastic zone of the
tunnel, the mechanism of load-bearing and acting relation
between surrounding rock and support were studied (Zhang
et al., 2008). The study was conducted on the basis of
traditional mining pressure theory. For different mining
conditions, the correlation between the bubbling and flaking
gang and the support parameters of large inclination isolated
working face was studied (Xie et al., 2013). The relationship
between the pressure pattern and the surrounding rock of the coal
seam above 10 m and below 10 m was studied (Yan, 2013). The
characteristics of the basic top rock layer breakage and the
working resistance of the support when crossing the old void
area were studied (Bi, 2015). A model of the relationship between
the brace envelope of the upper thin and lower thick coal seam
group and the brace envelope of the same mining was established
(Lu et al., 2017). The support system when the working face was
retrieved over the trap column was studied (Zhao, 2019). The
mechanism of mine pressure and the mechanical relationship
between the support and the surrounding rock at the 8.8 m super-
high comprehensive mining face was analyzed (Yang, 2020;Xu,
2021). The relationship between the support and the surrounding
rock under comprehensive mining in two directions was
analyzed: direction and tendency. This provides experience
and methods for the study of the bracket-rock relationship
under different mining conditions, based on which scholars
have conducted in-depth studies in different directions (Zhu,
2020). The relationship between the initial bracing force and the
stability of the roof slab was studied (Wan et al., 2011). The
“three-factor method”to consider the actual use effect of the
bracket was proposed (Zhang et al., 2014). A binary criterion for
determining the working resistance of the brace by balancing the
roof load and maintaining the stability of the coal wall was
established (Wang et al., 2014). The “single rock beam”and
“double rock beam”structures were proposed (Wu et al., 2016).
The method of moving the roof chaser with pressure rubbing was
proposed (Yang et al., 2018). The preliminary method of
determining the impact force was obtained (Yang et al., 2019).
A technical idea of using the mining pressure of the working face
to break the hard coal wall was proposed (Li et al., 2019). A
“macro-large-small”structural mechanics model of the quarry
overburden was established (Li et al., 2020). The relationship
between the support rock and the surrounding rock in two
qualitative and quantitative ways was demonstrated (Zhang
et al., 2020).
However, most of the scholars have studied the relationship
between the support and the surrounding rock at the working
face based on the traditional static load conditions, the
consideration of dynamic load is missing. In recent years some
scholars have taken the dynamic load effect caused by the roof
breakage into consideration in the study of the relationship
between the support and the surrounding rock (Yang, 2019).
However, most of the studies are only based on dynamic load
conditions and do not meet the reality that dynamic and static
loads exist simultaneously which are inseparable in the quarry.
Therefore, in order to better fit the actual engineering situation,
both must be considered when studying the bracket-rock
relationship. On this basis, a similar simulation experiment
platform that can simulate the relationship between the
bracket and surrounding rock in the quarry under dynamic
and static load conditions is developed. The relationship
between the bracket and surrounding rock under dynamic and
static load disturbance is further studied.
MODEL DEVELOPMENT
Experimental Platform Development
The independently developed dynamic and static load similarity
simulation platform can study the relationship between the
quarry support and the surrounding rock under different static
and dynamic load conditions. In this system, the platform mainly
consists of five parts, namely static load application device,
dynamic load application device, hydraulic support,
experimental main frame body, and pressure monitoring
system, as shown in Figure 1. The experimental platform’s
length is 130 cm, width is 65 cm, and height is 90 cm. The
mainframe of the platform is fitted with similar simulated
materials as required. From the bottom to the top are the
simulated coal wall and the direct roof, while the hydraulic
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139032
Yang et al. 3D Physical Experimental Study
support is installed in front of the coal wall. The coal body is
generally made up of sand, lime, gypsum, and water in a certain
ratio. The thickness of the coal body is 60 cm, the length is 65 cm
and the width is 65 cm. The direct roof is made up of sand, lime,
gypsum, water, and cement in a certain ratio. The thickness of the
direct top is 45 cm, the length is 130 cm and the width is 65 cm.
The hydraulic support consists of two hydraulic cylinders with a
load capacity of 8 t. The static load application device consists of a
hydraulic cylinder and a counterforce device, where the hydraulic
cylinder has an ultimate load capacity of 8 t. The magnitude of the
applied static load pressure is monitored and controlled by a
pressure gauge with a range of 10 MPa. The dynamic load
application device is a frame beam fixed with four
electromagnets with extremely strong adsorption force on the
iron plate, where the electromagnets can be adjusted by adjusting
the space position to simulate different degrees of dynamic load
impact effect. At the same time, the up and down movement of
the electromagnets can better simulate the different amounts of
delamination between the direct top and the basic top with an
accuracy of up to 1 mm and a maximum movement distance of
20 cm. The left and right movement of the electromagnets on the
track can better restore the location of the top plate breakage by
changing the iron plates under different conditions. It is possible
to simulate the different dynamic load effects of the overlying
rock layer on the roof plate. The specific operation is: after the
electromagnet is disconnected, the iron plate falls under the
action of gravity to simulate the basic roof breakage. The
pressure monitoring system consists of a load cell, a data relay
base station, and a computerized acquisition unit with a data
acquisition frequency of 50–100 Hz.
Experimental Process
In this experiment, the relationship between the quarry
support and the surrounding rock is simulated by the
orthogonal design of multiple comparison experiments with
different combinations of static and dynamic loads. Different
static loads represent the original rock stresses of different
coal mines, and different heights represent the strength of the
impact on the working surface when the roof is broken.
According to the similar scale and model size, the static
load of 0.5–2 MPa can represent the static load of the field
degree.Thestaticloadsaresetat0,0.5,1,and2MPa.
Considering the feasibility and simplification of the
experiment, the weight of the iron plate is 100 kg and the
height range is set to 1–10 cm to represent the impact of the
fundamental top fracture on the direct top. Table 1 shows the
specific experimental program.
The iron plate is dropped from a height of 1–10 cm above the
direct top under different static load conditions as follows. The
material proportions have been determined by several tests and
can be found in the literature (Yang et al., 2021).
1) Lay the simulated mock-up material, and mix well according
to the ratio of sand, lime, and gypsum of 20:1:1. Add 7% of the
weight of the well-mixed material with water, continue mixing
well, lay it into the box, tamp the coal body simulated material
using a tamping tool, and air-dry it for 3–5 days until the
model is moderately dry.
2) Place hydraulic support and pressure monitoring system in
front of the coal simulation material after it has air-dried.
3) Lay direct roof simulation material over the coal body
simulation material and hydraulic supports in a ratio of 5:
10:1:1 of sand, lime, gypsum, and cement, tamping with a
tamping tool and air-drying for 3–5 days until the direct roof
simulation material is moderately dry.
TABLE 1 | Experimental program.
Static loads (MPa) Hight/cm
0 1,2,3,4,5,6,7,8,9,10
0.5 1,2,3,4,5,6,7,8,9,10
1 1,2,3,4,5,6,7,8,9,10
2 1,2,3,4,5,6,7,8,9,10
FIGURE 1 | 3D physical modelling platform for the dynamic and static test.
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139033
Yang et al. 3D Physical Experimental Study
4) Application of the corresponding static load to the simulated
material by means of a static load application device.
5) Attach the iron plate to the electromagnet and drop it from a
height of 1–10 cm above the direct top.
6) The changes in hydraulic support resistance are recorded by a
pressure monitoring system and the coal wall displacement is
monitored by an infrared rangefinder.
RESULTS AND ANALYSIS
Analysis of Experimental Phenomena
Figure 2 shows the damage of similar simulated materials during
the experiments. These figures are all phenomena produced at
2 MPa and a fall height of 10 cm. (a) and (c) are phenomena after
the 3rd shock. (b) and (d) are the phenomena of the 10th impact.
(a)showsthecoalwalljuststarttoflake. The area of flaking is
mainly concentrated in the upper part of the coal wall,
accompanied by a small amount of damage in the lower part.
The damage is mainly in the form of tensile cracking damage. (b) is
a large area of the coal wall, mainly concentrated in the middle and
upper part of the coal wall, with the same tensile cracking damage.
Under the action of static load and impact load, the coal wall
produced tensile stress greater than its tensile strength and finally
produced tensile cracking damage. When the direct roof was
subjected to a smaller impact, a fracture was produced as shown
in Figure 2C, the length and width of the fracture were smaller.
After the direct roof has been subjected to a larger impact, a fracture
as shown in Figure 2D is produced and the entire direct roof is cut
down along the coal wall. In the coal mine site, when the amount of
inter-top delamination is small, the dynamic load is small, and the
extent of damage to the coal wall and roof is small. As shown in
Figures 2A,C, where the coal wall only produces a small area of
sheeting and the roof fissures begin to develop slowly. When the
amount of inter-roof departure is larger, the dynamic load is larger,
and the damage to the coal wall and roof is greater. As shown in
Figures 2B,D, where the coal wall starts to have large flake helpers
and the roof plate produces large cracks or complete fractures.
Development of Shield Loads
It was found that as the static load and the drop height of the iron
plate changed, the brace resistance and coal wall displacement
also changed. As shown in Figures 3,4, with the increase in the
height of the iron plate drop, the amount of offset between the top
slab in the coal mine site, the coal wall displacement, and the
resistance of the bracket keeps increasing. They are not linear.
The larger the amount of offset, the larger the increase. Therefore,
FIGURE 2 | Diagram of the experimental process.
FIGURE 3 | Impact forces at different heights. FIGURE 4 | Coal wall displacement at different heights.
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139034
Yang et al. 3D Physical Experimental Study
the core of the stability of the relationship between the bracket
and the surrounding rock in the mining site is to control the
amount of delamination between the basic top and the direct
top. So that the amount of delamination between the top plates
is as small as possible. The mine pressure will be relatively
moderate when the pressure comes in the mining cycle. Figure 3
shows the working resistance of the hydraulic support under a
static load of 0, 0.5, 1, and 2 MPa respectively. Through analysis,
it is found that with the increase of static load, the working
resistance of the hydraulic support is also increasing. However,
the increase is not large, which means that the change of static
load has little effect on the working resistance of the hydraulic
support. This is mainly because the weight of the direct top and
the impact force of the basic top on the direct top are borne by
the hydraulic support and the coal body. The coal body bears
most of the weight. The hydraulic support only bears a small
part of the weight. So, when the static load increases, most of the
energy is borne by the coal body, resulting in a significant
increase in the displacement of the coal wall, and the
phenomenon of the piece gang appears. The role of the
hydraulic support is mainly to control the amount of
departure between the direct top and the basic top, rather
than to resist the weight of the overburden. For working
faces with unstable coal walls, the stability of the coal wall
should be controlled as much as possible so that the coal body in
front of the working face can better bear the weight of the
overburden.
Development of Horizontal Displacement of
the Coal Wall
Figure 4 shows the displacement of the coal wall under different static
load conditions. According to the data in the figure, it can be found
that, also at 10 cm impact height, the maximum displacement of the
coal wall is about 7 cm without static load restraint. The maximum
displacement of the coal wall is 3 cm with 0.5 MPa static load
restraint. The maximum displacement is reduced by about 57%,
which can be considered that a certain degree of static load can
improve the stability of the coal wall. As the static load continues to
increase, the maximum displacement of the coal wall rapidly
increases to more than 8 cm, indicating that there is a threshold
value for the promotion of the static load on the stability of the coal
wall, below which the static load is beneficial to the stability of the coal
wall, beyond which it will instead accelerate the sheeting of the coal
wall, thus affecting the stability of the quarry support and the
surrounding rock system. With the increase of static load, the
change of coal wall displacement is more obvious than the
hydraulic bracket working resistance. With the increase in the
amount of offset, the displacement of the coal wall keeps
increasing. The increase is not linear and greater under the
conditionofalargestaticload.Itcanbeconsideredthatthe
increase in static load does not have a significant impact on the
working resistance of the hydraulic support, but has a greater impact
on the displacement of the coal wall. So, the stability of the quarry
support and the surrounding rock system can be controlled by
increasing the resistance of the hydraulic support which is limited.
Thestabilityofthequarrysupportandthesurroundingrocksystem
should be studied from both the stability of the coal wall and the
amount of delamination between the roof plates. The support force of
the coal body should also be considered in the design and selection of
the on-site hydraulic support.
Development of Stability Coefficient
From the previous analysis, the load on the quarry is borne by
the coal wall and the bracket together. The source of the
dynamicloadismainlyduetotheimpacteffectonthedirect
topcausedbythebreakageoftheroofplateduringthe
periodic incoming pressure. So, the stability of the bracket
and the surrounding rock system of the quarry can be
achieved by enhancing the ultimate bearing capacity of the
coal wall or the maximum working resistance of the hydraulic
bracket. The role of hydraulic support is to control the amount
of separation between the basic roof and the direct roof. In the
coal mine site, the initial support force of the hydraulic
support can be increased to reduce the amount of
separation between the roof slabs, thus reducing the
amount of dynamic load impact caused by roof breakage.
FIGURE 5 | Shear failure model of coal wall (Kong et al., 2017).
TABLE 2 | Values for each parameter.
G
(MN)
α
(°)
φ
(°)
C
(MN)
H
2
(m)
Fmax (MN) m
(kg)
Q
(MN)
L
(m)
H
1
(m)
t
(s)
p
(g/
cm
3
)
Q
Z
(MN)
E
(MPa)
0.3 20 20 1 0.5 15 3,000 0.5 20 5 0.5 2 3 3.0 × 10
4
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139035
Yang et al. 3D Physical Experimental Study
The stability of the support and the surrounding rock depends on
the ultimate bearing capacity of the coal wall and the maximum
working resistance of the hydraulic support. The quarry support and
surrounding rock system will become unstable when the sum of the
ultimate bearing capacity of the coal wall and the maximum working
resistance of the hydraulic support is less than the load given to the
direct top by the weight of the basic top and the direct top. Therefore,
the stability factor for the quarry bracket and the surrounding rock
system is defined Kastheratioofthesumoftheultimatebearing
force of the coal wall and the maximum working resistance of the
hydraulic support to the load applied to the quarry.
KQmax +Fmax
Fc+QZ
,(1)
where Q
max
is the ultimate bearing force of the coal body, F
max
is
the maximum working resistance of the hydraulic support, and F
c
is the impact force on the direct roof after the basic roof
destabilization (Yang et al., 2019) and Q
z
is the weight of the
direct roof.The ultimate bearing capacity of the coal body Q
max
is
(Wang et al., 2014)
Qmax Gsin αtan φ+CH2sec α+Q0cos αtan φ+sin α
cos α−sin αtan φ,
(2)
where Gis the gravity of coal sliding body, αis the included
Angle between shear surface and coal wall, φis the Angle of
internal friction of coal body, Q
0
is the force of protective plate
on coal wall, Cis the cohesion of coal body, H
2
is the failure
height of coal wall. The specific parameters and position
relationship can be referred to in Figure 5.
Simplifying Fq. 2, as the force of the coal wall guard plate in
the field has little effect on the coal wall, the calculation of Q
0
will
be ignored, and the following formula is obtained:
Qmax Gsin αtan φ+CH2sec α
cos α−sin αtan φCH2sec α
cos α−psin α−G.(3)
Because the impact force caused by slip instability is greater than
the impact force caused by rotary deformation instability, F
c
calculates the impact force according to the slip instability (Yang
et al., 2019), and it is simplified to obtain Eq. 4.Letp=tanφ,and let n
=EH
1
Lρ.
Fcm
t
2Q2
EH1Lρtan φsin β+cos βtan φ2+2gΔ
+Q
m
t
2Q2
npsin β+pcos β2+2gΔ
+Q.(4)
In Eqs. 4,tis the impact time after the basic roof is broken,
mis the mass of the basic roof, Qis the overburden load of the
basic roof, Eis the elastic modulus of the basic roof, H
1
is the
thickness of the basic roof, Lis the length of the basic roof
FIGURE 6 | Effect of stability coefficients for each parameter.
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Yang et al. 3D Physical Experimental Study
block, ρis the density of the basic roof, Δis the off-layer
volume, and gis the gravitational acceleration.
Substituting Eqs. 3,4into Eq. 1 for simplification, we obtain
the following equation:
K
CH2sec α
cos α−psin α−G+Fmax
m
t
2Q2
np(sin β+pcos β)2+2gΔ
+Q+QZ
.(5)
A sensitivity analysis of the stability coefficient Kof the
quarry bracket and the surrounding rock system was carried
out according to Eq. 5. The parameters are assigned according
to the actual situation on-site, as shown in Table 2.
Figure 6 shows the effects of off-layer volume, maximum
working resistance of the support, modulus of elasticity, and
direct top weight on the stability of the quarry bracket and the
surrounding rock system. As the off-layer volume increases
(Figure 6A), the stability of the quarry stands and the
surrounding rock system becomes worse. It is consistent with
the derivation from the previous study, from which (Yang et al.,
2019) it is also evident in this study that as the amount increases,
F
冲
increases significantly. This leads to greater loading of the
quarry, and in turn, makes the quarry support and the surrounding
rock system less stable. The increase in the maximum working
resistance of the hydraulic support (Figure 6B)enhancesthe
stability of the quarry bracket and the surrounding rock system.
The increase is linear, but the maximum working resistance of the
hydraulic support in the field is constrained by equipment and
technology. It cannot be increased indefinitely, so the use of
support to control the stability of the quarry bracket and the
surrounding rock system is limited.The modulusof elasticity of the
basic roof (Figure 6C) is positively correlated with the stability of
the quarry bracket and the surrounding rock system. The increase
in the modulus of elasticity makes the quarry more stable, but the
increase is not linear. As the modulus of elasticity increases, the
growth rate of the stability coefficient of the quarry bracket and the
surrounding rock system gradually tends to level off. As the direct
top weight increases (Figure 6D), the static load required on the
support and coal body increases, making the quarry less stable and
more prone to instability.
The stability coefficient Kindicates the stability degree of
the quarry bracket and the surrounding rock system. When K
is greater than 1, it indicates that the quarry bracket and the
surrounding rock system are stable. The greater the value of K
indicates that the quarry can bear the load the stronger the
ability. When Kis less than 1, it indicates that the quarry
bracket and the surrounding rock system will be destabilized.
The specific destabilization forms macroscopic performance,
that is, the coal wall sheet gang, hydraulic bracket pressure
frame, and inverted frame, the smaller the value of Kindicates
that the quarry can bear the load the weaker the ability. As can
be seen from the above equation, in addition to improving the
stability of the quarry bracket and the surrounding rock
system can improve the ultimate bearing capacity of the
coal body. It can also increase the maximum working
resistance of the hydraulic bracket and reduce the load on
the working face. For the coal wall, the ultimate bearing
capacity of the coal body can be increased by means of
“flexible reinforcement.”In working faces with unstable
coal walls, increasing the stability of the coal wall is more
effective in controlling the stability of the quarry bracket and
the surrounding rock system than increasing the maximum
working resistance of the hydraulic support. For the
hydraulic bracket, raising the maximum working resistance
as much as possible will play a certain role in promoting the
quarry bracket and the surrounding rock system, but not
fundamentally. The amount of separation between the roof
plate should be reduced by raising the initial bracing force of
the hydraulic bracket, to achieve the reduction of the dynamic
load impact of the roof plate, which is the most effective
measure to reduce the dynamic load effect of the quarry.
Comprehensively, it can be considered that the core of
controlling the stability of the quarry bracket and
surrounding rock system is to improve the stability of the
coal wall and reduce the amount of off-layer volume.
CONCLUSION
1) A similar simulation platform for dynamic and static loads
has been developed, allowing the study of the relationship
between the quarry support and the surrounding rock
under simultaneous consideration of dynamic and static
load disturbances, with different static load constraints
and different drop heights set up for the experiments.
2) The core of the stability of the relationship between the
bracket and the surrounding rock in the quarry is to
control the amount of delamination between the basic top
and the direct top so that the amount of delamination between
the top plate is as small as possible or no delamination is
produced.
3) The stability of the quarry bracket and the surrounding rock
system is limited by increasing the resistance of the hydraulic
bracket in the quarry, and it needs to be considered more from
the stability of the coal wall and the amount of separation
between the roof plates.
4) Define the stability coefficient Kofthequarrybracketand
the surrounding rock system. The amount of separation,
the direct jacking weight, and the stope support are
nonlinearly inversely related to the stability of the
surrounding rock system. The maximum working
resistance of the bracket is linearly positively correlated
with the stability coefficient. The elastic mold is
nonlinearly positively correlated with the stability
coefficient.
DATA AVAILABILITY STATEMENT
The original contributions presented in the study are included in
the article/Supplementary Material, further inquiries can be
directed to the corresponding author.
Frontiers in Earth Science | www.frontiersin.org May 2022 | Volume 10 | Article 9139037
Yang et al. 3D Physical Experimental Study
AUTHOR CONTRIBUTIONS
SY provided the overall framework of the thesis. HY
completed the physical experiment section. RZ and ZC
analyzed experimental data. XW changed the formatting
and images of the article. All authors contributed to
manuscript revision and read and approved the submitted
version.
FUNDING
This study was supported by the National Natural Science
Foundation of China (51974320), Natural Science Foundation
of Hebei (E2020402041), “State Key Laboratory of Coal Mining
Water Conservation and Utilization”2017 Open Fund Project
Grant (SHJT-17-42.4), and the Science and Technology
Innovation Project of China Energy Investment Corporation
(SHJT-17-38).
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