Mechanical Properties and Control Rockburst Mechanism of Coal and Rock Mass with Bursting Liability in Deep Mining

In order to study the mechanism of the dynamic disaster of rockburst in a deep coal mine and the prevention and control measures of weakening shock, the MTS815.03 servocontrolled rock mechanics test system is used to test the coal, rock, and combined specimens with the buried depth of nearly 1200m in Xinwen Mining Area. And their mechanical properties, energy evolution, and bursting properties are studied and analyzed. The rationality of the test results is also veriﬁed by the in-situ engineering practice. The key conclusions are as follows: (1) There is a relation between the ratio of elastic modulus E e before peak strength to descending modulus E d after peak strength and the bursting properties. For the fractured coal, the descending modulus E d is relatively small, and the E e / E d is relatively large and presents progressive ductile failure with low probability and risk of rockburst. For the less fractured rock, the descending modulus E d is relatively large, and the E e / E d is relatively small and presents brittle failure, which is very similar to the characteristics of rockburst. (2) For the same type of rock, with the increase of conﬁning pressure, the E e / E d gradually increases, indicating the reduction of rockburst strength. Therefore, the greater the support strength provided to the surrounding rock surface of the roadway, the smaller the failure degree of rockburst. (3) With the increase of conﬁning pressure, after peak strength, the elastic energy of coal specimens decreases slowly, and the dissipated energy increases slowly, indicating that the increase of conﬁning pressure can eﬀectively limit the energy dissipation and release after coal specimen failure. So, in the in-situ engineering practice, it is an important measure to improve the surface restraint and support strength of the coal roadway for reducing the occurrence intensity and probability of rockburst. (4) The combined measures of “the mining of double liberating seam+the implementation of large-diameter pressure relief borehole in advance of working face” is the very eﬀective way to eliminate the rockburst accidents of working face in a protected coal seam and has an important guiding signiﬁcance for the safe mining of rockburst mine.


Introduction
China is rich in coal, but it has little gas and oil. In the past, coal has provided the main driving force for the rapid development of China's national economy. Now, although China is promoting the development of the new energy industry, the status of coal as the main energy in China will not change for a long time. With the continuous increase of mining intensity and depth, the number of rockburst mines in China is increasing rapidly, and the distribution range is also expanding significantly. According to incomplete statistics, there occur several rockburst accidents in China every year, causing a large number of casualties and equipment damage [1][2][3][4][5]. erefore, how to prevent the rockburst accident is a serious problem faced by China's mining industry. In the past decades, many experts and scholars have done a great deal of research and exploration of the theory, the monitoring and the early warning, prevention, and control technology about rockburst.
In terms of the theory of rockburst, the three-factor theory put forward by Qi et al. [6], the theory of rockburst start-up during coal mining by Pan et al. [7], and the instability theory of deformation system by Zhang [8] are very representative. In addition, some other scholars have systematically summarized the mechanism of rockburst according to different production and geology conditions. For example, Zhu et al. [9] studied the mechanism of rockburst caused by fault sliding instability during mining in isolated island working face and summarized the relationship between fault sliding instability and rockburst risk. Shi et al. [10] studied the classification of rockburst from the viewpoint of energy storage and release and analyzed the occurrence mechanism of various types of rockburst. He et al. [11] studied the coupled mechanism of compression and prying-induced rockburst in steeply inclined coal seams and principles for its prevention. Jiang et al. [12] studied the influence of fault parameters on rockburst. Zhang and Jiang [13] proposed the stress criterion and energy criterion for rockburst occurrence, according to different positions of roadways in the coal seam.
For the monitoring and early warning of rockburst, the online monitoring system of mining stress, the microseismic and underground sound monitoring systems are widely used [14][15][16][17][18]. ese systems effectively reduce the rockburst accidents. In addition, many scholars have studied the prediction method of rockburst. For example, Cai et al. [19] put up a new coal burst forecasting method in underground coal mines based on the seismic strain energy. Wen et al. [20] studied the rockburst hazard evaluation method. Zhu et al. [21], under the analysis of load transfer mechanisms of key stratum and extremely thick alluvium stratum (ETAS), proposed a theoretical computational model of abutment pressure for ETAS longwall panels. And then, the model was applied to the determination of the abutment pressure distribution in the working face.
In terms of the prevention and control of rockburst, the methods of mining the liberating seam, large-diameter borehole pressure relief of coal seam, water injection of the coal seam, and deep hole blasting of the roof are often used to reduce or eliminate the risk of rockburst [22][23][24][25][26]. For the effect of prevention measures, the drilling cutting method is generally used for verification [27,28].
Many theories, monitoring and early warning, prevention, and control technology about rockburst, have been achieved, but rockburst accidents still occur from time to time, which shows that there are still many deficiencies in the current understanding and research of rockburst problems. Especially in the deep mining over 1 km, the high ground stress and strong mining stress are very prominent. When the coal seam, the roof, and the floor strata are all in danger of rockburst, how to effectively prevent the dynamic disaster accident is a serious problem. Based on this, taking the deep mining area of Huafeng coal mine in Shandong as the engineering background, this paper studies the above rockburst problems by a lab test, theoretical analysis, and in-situ engineering practice. e research results have some guiding significance for the prevention and control of rockburst in deep mining.

Specimen Preparation.
e coal and rock samples were from the deep mining area of the Huafeng coal mine in Shandong, China. ese samples were sealed on-site and transported to the laboratory. Each sample was processed into the standard cylindrical specimens with a diameter of 50 mm and a height of 100 mm; the flatness error of the specimen section does not exceed 0.02 mm, and the axial perpendicularity is less than 0.001 radian, which is suggested by the International Society for Rock Mechanics (ISRM). In addition, the combined specimen is composed of the roof, coal, and floor specimens, and their heights are 30 mm, 40 mm, and 30 mm, respectively. Some processed specimens are shown in Figure 1.

Test Procedures.
e tests were conducted using an MTS815.03 servocontrolled rock mechanics test system (MTS Systems Corporation, Eden Prairie, MN, USA) of Shandong University of Science and Technology (see Figure 2). e conventional compression tests were carried out on the coal, roof, floor, and combined specimens. Prior to the test, according to the different types of specimens, a corresponding initial force of 1∼2 kN was applied to the specimen to ensure close contact between the test machine pressure head and the specimen. e displacement control mode was used in the loading process of the specimen, and the displacement loading rate is 0.003 mm/s. In the conventional triaxial compression test, the confining pressure loading rate of 0.05 MPa/s was chosen.

Stress-Strain Characteristics and Constitutive Relation
3.1.1. Stress-Strain Curve. Figure 3 shows the stress-strain curves of the roof, floor, and coal specimens under different confining pressures. From Figure 3, the prepeak stress-strain curves of deep coal and rock specimens are not smooth, and there are many stress jumping points. Showing that in the deep high ground stress environment, there are abundant microcracks in coal and rock masses, and with the prominent heterogeneity in them. In addition, under the conventional compression loading, the deformation and failure of coal, rock specimens show obvious phased characteristics; that is, their whole stress-strain curve can be roughly divided into prepeak and postpeak parts according to the strength changing trend. e prepeak stress-strain curve can be roughly divided into three phases, namely: the microcrack compaction, elastic deformation, and elastic-plastic deformation phases. In the microcrack compaction phase, with the increase of stress, dσ/dε increases gradually, indicating that the microcracks in the specimen are gradually compacted. In the whole elastic deformation phase, dσ/dε is constant, and the specimen shows linear deformation. After falling into the elastic-plastic deformation phase, with the increase of stress, dσ/dε gradually decreases, and the nonlinear deformation characteristics become gradually obvious. Showing that the microfracture events in the specimen gradually develop from a disordered state to a concentrated state, and the number of failure elements also increases. When the axial stress exceeds the peak strength of the specimen, the specimen falls into the postpeak phase. However, the failure characteristics of different types of specimens are diverse under varying confining pressures. Under uniaxial stress state, the failure of roof and floor specimens is brittle, and the coal specimen is progressive. But, under the confining pressures of 3 and 5 MPa, the failure of the three specimens is slowly progressive.

Constitutive Relation.
e dynamic failure of coal and rock masses is the basic problem in deep mining. In the research, the dynamic failure process of coal and rock masses should be focused on, especially the dynamic evolution process of crack initiation, propagation, and penetration. erefore, it is necessary to establish the elastic-plastic constitutive relation and the energy criterion for judging the failure of coal and rock masses.
According to elastic-plastic mechanics, the mechanical model can be simplified as a double line stress-strain relation model (see Figure 4). In the prepeak stress-strain curve, assume the linear relation between stress and strain, and the slope of the curve is called elastic modulus E e . In the postpeak stress-strain curve, also assuming the linear relation between stress and strain, the absolute value of the slope of the curve is called descending modulus E d .
e E e /E d values indicate different bursting liability indexes: In pre − peak phase, σ � E e ε, ε < ε c , where E e is the linear elastic modulus in the prepeak curves, and E d is the linear descending modulus in the postpeak curves. e comparison of the prepeak linear elastic modulus E e and postpeak linear descending modulus E d of the roof, the floor, and the coal specimens under different confining pressures is shown in Table 1. Under the confining pressure of 0, 3, and 5 MPa, the elastic modulus E e and descending modulus E d of the roof and floor specimens are significantly higher than that of coal specimen, indicating that the high strength roof and floor will represent a more violent failure process than coal. However, the E e /E d of roof and floor specimens is obviously smaller than that of coal. So, the E e / E d can be used as an index to evaluate the rockburst strength of coal and rock masses. at is, the smaller the E e /E d , the greater the rockburst strength. In addition, for the same type of rock, with the increase of confining pressure, the E e /E d increases, that is, the rockburst strength decreases. is conclusion shows that the greater the support strength provided to the surrounding rock surface of the roadway, the smaller the failure degree of rockburst.      [29]. When the rock element is deformed by compression loading, assuming that all the work done by the external force is absorbed by the rock, the energy "U" input to the rock can be expressed as follows [30]: where U e is elastic strain energy and U d is dissipated energy. e elastic strain energy "U e " is expressed as follows [31]: where E u is the elastic modulus and μ u is Poisson's ratio.
In the conventional triaxial compression test, the energy "U" is expressed as follows [32]: where U 0 is the energy absorbed in the hydrostatic pressure phase; U 1 is the energy absorbed in the specimen under the axial force; U 3 is the consumed energy under the confining pressure (a negative value). When both U 0 and U 3 are 0, it is the energy absorbed in the specimen under uniaxial compression. e absorbed energy by the specimen "U 0 " during the hydrostatic pressure phase is expressed as follows [32]: e energy "U 1 "and "U 3 " can be calculated by the integral accumulation method according to the stress-strain curve in the lab test. Figure 5 shows the calculation principle of the energy "U 1 ." e calculation equations of energy "U 1 " and "U 3 " are as follows: where σ 1,i is the axial stress at the ith sampling point on the axial stress-strain curve, MPa; σ 1,i+1 is the axial stress at the (i+1)-th sampling point on the axial stress-strain curve, MPa; ε 1,i is the axial strain corresponding to the sampling point σ 1,i ; ε 1,i+1 is the axial strain corresponding to the sampling point σ 1,i+1 ; σ 3,i is the confining pressure corresponding to the sampling point σ 1,i ; σ 3,i+1 is the confining pressure corresponding to the sampling point σ 1,i+1 ; ε 3,i is the circumferential strain corresponding to the sampling point σ 1,i ; ε 3,i+1 is the circumferential strain corresponding to the sampling point σ 1,i+1 Under the uniaxial stress state, the energy equations are simplified as follows: where σ 1i is the axial stress of the ith sampling point on the stress-strain curve, MPa; σ 1i+1 is the axial stress of the (i + 1)th sampling point on the stress-strain curve, MPa; ε 1i is the axial strain of the ith sampling point on the stress-strain curve; ε 1i+1 is the axial strain of the (i + 1)-th sampling point on the stress-strain curve. "E u " is the unloading elastic modulus, but, so far, there is no consensus on the value of "E u ." In order to facilitate the calculation, the elastic modulus "E 0 " at the late elastic phase is chosen instead of E u [30].

Energy Evolution Characteristics
(1) Influence of Confining Pressure on Energy Evolution of Specimen. Taking the coal specimen as an example, the influence of confining pressure on the specimen energy evolution is illustrated. Under the confining pressure of 0, 3, and 5 MPa, the energy evolution curves are shown in Figure 6. From Figure 6, under the confining pressure of 0, 3, and 5 MPa, the evolution trend of input energy, elastic energy, and dissipated energy are the same, but there is an obvious difference at the increasing and releasing rates of energy. Especially, in the postpeak phase, compared with the uniaxial stress state, the   Shock and Vibration 5 elastic energy decreases slowly, and the dissipated energy increases slowly under the confining pressures of 3 and 5 MPa, which shows that the confining pressure can effectively limit the energy dissipation and release after the coal specimen failure. So, in the in-situ engineering, providing enough support strength for the roadway surrounding rock surface can effectively slow down the energy release rate and reduce the rockburst strength.
(2) Energy Evolution Characteristics of Different Types of Specimens under Uniaxial Stress State. Because the most rockburst accidents occur in the coal pillar or roadway with a free surface, this paper focuses on the energy evolution characteristics of roof, floor, coal, and combined specimens under the uniaxial compression. Figure 7 shows the energy evolution curves of the four types of specimens. e energy evolution of roof and floor specimens is different from that of coal and combined specimens. As for the high strength roof and floor specimens, the input energy, elastic energy, and dissipated energy are very low in the compaction phase. In the prepeak phase, the input energy and the elastic energy increase significantly, but the dissipated energy basically remains the same. After the peak strength is reached, the elastic energy is violently released and the dissipated energy increased dramatically. As to the low strength coal and combined specimens, the energy evolution process is complex. In the early compaction phase, the input energy, elastic energy, and dissipated energy are very low, but in the late compaction phase, the input energy and dissipated energy increase significantly, while the elastic energy only increases slightly. In the elastic phase, the input energy and the elastic energy increase rapidly, while the dissipated energy basically remains the same. In the elastic-plastic phase, the input energy continues to increase because there are many microcracks in the coal. Several local fractures takes place and result in a significant increase of the dissipated energy, while the elastic energy increases in the fluctuation.

Characteristics of the Bursting Energy Index.
Under uniaxial compression loading, the prepeak energy, postpeak energy, and bursting energy index of four types of specimens are shown in Table 2. e bursting energy index is the ratio of prepeak energy to postpeak energy. From Figure 7 and Table 2, in the four types of specimens, the strength of roof and floor specimens is highest, the prepeak input energy and elastic energy are the largest, but the postpeak energy release rate is the fastest and the bursting energy index is the largest, showing that the rockburst destruction force of brittle rock is greater than that of coal. erefore, in the in-situ engineering, the prevention and control of the roof and floor rockburst should be highly valued, because once the rockburst occurs, the destructive force is much greater than that of the coal seam. For the roof and floor to easy accumulate energy, the water injection softening and blasting can be used to increase their internal cracks and reduce the strength so that the failure mode is changed from brittleness to ductility and finally reduce the accumulated energy in the strata and slow down the energy release rate. Compared with the coal specimen, the bursting energy index of the combined specimen is slightly larger, which indicates that the rockburst destruction force of the thin or medium-thick coal seam clamped by the rock strata is greater than that of the single thick coal seam. e main reason is that the elastic energy accumulated in the roof and floor is released rapidly and converted into the kinetic  Shock and Vibration 7 energy of the coal mass movement when the coal seam is collapsed by the rockburst.

Discussion
(1) From Figure 7, the released energy of the specimen at the peak strength is not necessarily the largest. e maximum released energy usually occurs at some point on the postpeak stress-strain curve. Because a large amount of crack penetration needs to take some time to develop after peak strength.
is well explains the reason that the rockburst does not occur immediately when the stress concentration exceeds the peak strength of coal and rock masses in the stope. So, it is the key to prevention and control of dynamic disasters to study the postpeak stress-strain curve of coal and rock. According to the stress-strain curves of coal and rock specimens, when the brittle failure is more serious, the larger the E d , the smaller the E e /E d , and the more likely the rockburst occur. On the contrary, when progressive ductile failure is more obvious, the smaller the E d , the larger the E e /E d , and the possibility of rockburst is reduced.
(2) e dissipation and release of energy are the essential properties during the damage and failure process of loaded coal and rock. So, the dissipated energy "U d " at peak strength can be used as an energy indicator to predict the occurrence of rockburst disasters. U d � σ 2 c /2E c ; σ c is the peak strength, E s is the elastic modulus. In the in-situ engineering, rockburst will occur when the deformation energy stored in the whole coal and rock mass exceeds the index "U d ." (3) e test results show that the increase of confining pressure can effectively limit the energy dissipation and release after coal specimen failure. So, in the insitu engineering, it is an important measure to improve the surface restraint and support strength of coal roadway for reducing the occurrence intensity and probability of rockburst. (4) According to the above findings, when mining the deep rockburst coal seam, both the regional and local control measures should be adopted. As for the regional control of rockburst, liberating seam mining can be the most important method. After the mining of the liberating seam, the cracks in the roof and floor strata of the protected coal seam increase and reduce their stress concentration and the bending energy index. As for the local control of rockburst, the large-diameter borehole pressure relief, water injection cracking, and blasting can be carried out in advance of the working face. After the implementation of local control measure, the internal cracks of coal and rock masses increase, the stress concentration of working face reduces, and the energy dissipation increases, and finally, the probability and intensity of rockburst will be significantly reduced or eliminated. After synthesis evaluation, the eight areas of No. 1412 working face are classified as risk zones of rockburst. In order to facilitate the management, the adjacent risk zones are combined, and finally, there are four risk zones. Among them, there are two low-risk zones, one medium risk zone, and one high-risk zone. e upper roadway range affected by the first roof pressure of No. 1412 working face is determined as the low risk zone (see green zone of No. 1412 upper roadway in Figure 8). e range away from the cutting hole from 0 to 157 m is determined as the medium risk zone (see yellow zone in Figure 8). e lower roadway range away from the cutting hole from 157 to 530 m is determined as the low risk zone (see green zone of No. 1412 lower roadway in Figure 8). e upper roadway range away from the cutting hole from 157 to 530 m is determined as the high-risk zone (see red zone of in Figure 8).

Local Prevention and Control of Large-Diameter Borehole Pressure Relief in Advance of Working Face.
According to the theories of mine rock mechanics and mine pressure, although in mining pressure relief condition of double liberating seams, the stress concentration still occurs in the key parts of the working face, such as the first roof pressure, the working face square, and the geological structure area. In order to ensure that the working face is always mined under the low-stress state, the local control measure must be carried out, such as the large-diameter pressure relief borehole, water injection cracking, blasting, and so on. Especially, the implementation of a large-diameter pressure relief borehole can effectively reduce the stress concentration of the coal mass, increase the energy dissipation, and finally reduce the probability and strength of rockburst. e pressure relief principle of the large-diameter borehole is shown in Figure 9. e parameters of deep large-diameter pressure relief borehole in coal seam mainly include the depth, spacing, and diameter of the borehole. e drilling depth should ensure that the coal seam is in a nearly three-dimensional stress   state after pressure relief, and the coal mass not easy to rush out (the resistance is greater than the rush out force). e borehole spacing should be determined according to the burial depth, hardness, and rockburst risk of the coal seam. Based on the elastic-plastic mechanics analysis of the breaking, plastic, elastic, and original stress zones around the circular borehole and the stress in these zones, and combined with the in-situ practice experience, the diameter of the pressure relief borehole is chosen as 182 mm, and the depth is not less than 20 m.
(1) Prepressure Relief of Large-Diameter Borehole in Advance of Working Face. Before mining, the large-diameter borehole is carried out for the coal mass pressure relief within 300 m in advance of the upper and lower roadways of No. 1412 working face. e drilling equipment is the CMS1-6000/55 hydraulic drill, as shown in Figure 10. Within the range of 300 m in advance of working face, the low risk, medium risk, and high-risk zones, the large-diameter pressure relief borehole with a depth of 30 m is constructed at the spacing of 5 m, 3 m, and 1 m, respectively. Because both sides of No. 1412 lower roadway are the solid coal mass, the large-diameter pressure relief boreholes are arranged on both sides of the roadway. e dip angle of the boreholes is the same as that of the coal seam. According to the distribution characteristics of the abutment pressure in the stope, and after considering the safety factor, the borehole depth is determined to be 20 m. e lower side of No. 1412 upper roadway is the solid coal mass, and the upper side is near the No. 1411 goaf, so the borehole is arranged at the lower side of the roadway. Similarly, according to the distribution characteristics of the abutment pressure in the stope, and after considering the safety factor, the borehole depth is determined to be 30 m. In the in-situ engineering practice, it is necessary to ensure that the pressure relief boreholes are constructed within 300 m in front of the working face.
For the special zones requiring enhanced pressure relief, after the implementation of the large-diameter borehole, it is necessary to use the drilling cutting method to test the effect of pressure relief. If the amount of coal scraps exceeds the standard or there are dynamic phenomena (such as borehole sticking, roof shaking, coal blasting, etc.) in the process of drilling, the effect of pressure relief is not up to the standard. At this time, after the synthesis evaluation of these risk zones and determination of the risk level, the large-diameter boreholes are constructed again according to the required pressure relief parameters. And then test the effect of pressure relief until up to the standard.

(2) Safety Coordination Control of Reinforcement-Pressure
Relief in Rockburst Roadway Surrounding Rock. According to the conclusion of Section 3, the increase of confining pressure can effectively limit the energy dissipation and release after coal specimen failure. So, in the in-situ engineering, it is an important measure to improve the surface restraint and support strength of coal roadway for reducing the occurrence intensity and probability of rockburst.
rough in-situ measurement and engineering practice, the supporting scheme of the roadway surrounding rock is determined. e support parameters of the roof are as follows: the Φ6 mm metal net with a grid size of 100 mm × 100 mm is laid on the roof and two sides of the roadway. en, the one trapezoid-typed steel band with a length of 4 m and the six Φ22 mm × L4300 mm anchor cables are combined to install in the roadway roof, and each anchor cable installed a 150 mm × 150 mm × 30 mm load-bearing plate. Each anchor cable is anchored using two MSZ2850 resin anchorage agents with a pretension of 80 kN, and the anchorage force is not less than 150 kN. Besides, along the roadway strike, the anchor cable beam is installed in the roadway roof and consists of one beam with a length of 4.8 m and three Φ22 mm × L10300 mm anchor cables with the interval of 1800 mm. Each anchor cable is anchored using three MSM2350 resin anchorage agents with a pretension of 80 kN, and the anchorage force is not less than 200 kN. e spacing of the anchor cable beam is set to 1.0 m. e support parameters of the solid coal side are as follows: e one trapezoid-typed steel band with a length of 2.2 or 1.3 m and the five Φ22 mm × L2500 mm metal screw equal-strength bolts are combined to install in the solid coal side of the roadway. e bolts in the top corner of the roadway are inclined to the roadway roof, with an angle of 15°∼20°to the horizontal direction. And the bolts in the bottom corner of the roadway are inclined to the roadway floor, with an angle of 10°∼30°to the horizontal direction. Besides, the channel steel with a length of 3.2 m is vertically installed on the two sides of the roadway.
ere are three anchor cable holes in the channel steel, and the spacing of holes is 1.2 m. e Φ22 mm × L10300 mm anchor cables are installed in the top hole, the Φ22 mm × L2500 mm metal screw equal-strength bolts installed in the middle hole, and the Φ22 mm × L6300 mm anchor cables installed the bottom hole. e spacing of channel steel is set to 1.0 m. e support parameters of the roadway side near goaf are as follows: e anchor cable beam is supported by the shed leg made of the U29 steel. e shed leg is composed of two parts, the upper part is 2.0 m long, the lower part is 3.2 m long, and the connecting part of the two is overlapped by 0.5 m. e pretension moment of each bolt is not less than 400 N·m, and the anchorage force not less than 120 kN.  maximum is up to 35. However, in No. 1412 working face, after implementation of large-diameter pressure relief borehole, the maximum frequency of microseismic events is only 11. Figure 12 shows the energy release and event frequency monitored by the geoacoustic monitoring system during the mining of No. 1412 working face. e event frequency is very high, but the energy very low, indicating that there is a small-sized fracture in coal mass. e smallsized fracture releases the energy accumulated in the coal mass and reduces the stress concentration. rough the data of microseism and geoacoustic monitoring systems, it can be concluded that the large-diameter pressure relief borehole can effectively reduce the bursting risk of the working face.

Real Time Stress Monitoring and Analysis.
In order to effectively evaluate the pressure relief effect of the largediameter borehole, in the upper and lower roadways of No. 1412 working face, the borehole stress sensors are arranged to monitor the stress evolution of coal mass in real time. e first measuring station is arranged at a spot of 30 m from No. 1412 cutting hole, and the subsequent stations are arranged according to the interval of 30 m in turn. Each station is equipped with two measuring points with the spacing of 2 m, and the installation depth of borehole stress sensors is 14 m and 18 m, respectively. e early warning of borehole stress is divided into two levels, namely: yellow and red warnings; the former warning value is set as 10 MPa, and the latter warning value is 12 MPa. With the advance of the working face, timely dismantle the borehole stress sensors and move them forward to arrange to ensure that the monitoring range is not less than 300 m. Figure 13 shows the monitoring results of the advancing abutment stress in No. 1411 working face with a section coal pillar. e maximum stress and the maximum change rate appear in the right borehole of No. 51 measuring station, the depth of this borehole is 18 m, and 169.8 m away from the   e combined measures of "the mining of liberating seam + the implementation of large-diameter pressure relief borehole in advance of working face" provide the theoretical guidance and engineering reference for the safe mining of rockburst mine.

Conclusions
rough the lab test and in-situ engineering practice, the main conclusions of this paper are as follows: (1) e maximum released energy of the specimen usually occurs at some point on the postpeak stress-strain curve. Because a large amount of crack penetration needs to take some time to develop after peak strength. So, it is the key of Shock and Vibration 13 prevention and control of dynamic disasters to study the postpeak stress-strain curve of coal and rock. (2) e E e /E d can be used as an index to evaluate the coal and rockburst strength, and the smaller the E e /E d , the greater the rockburst strength. In addition, for the same type of rock, with the increase of confining pressure, the E e /E d increases, that is, the rockburst strength decreases. So, the greater the support strength provided to the roadway surrounding rock surface, the smaller the failure degree of rockburst. (3) With the increase of confining pressure, after peak strength, the elastic energy of coal specimen decreases slowly and the dissipated energy increases slowly, indicating that the increase of confining pressure can effectively limit the energy dissipation and release after coal specimen failure. So, in the insitu engineering practice, it is an important measure to improve the surface restraint and support strength of coal roadway for reducing the occurrence intensity and probability of rockburst. (4) e combined measures of "the mining of double liberating seam + the implementation of large-diameter pressure relief borehole in advance of working face" is the very effective way to eliminate rockburst accidents of working face in protected coal seam and provide the theoretical guidance and engineering reference for the safe mining of rockburst mine.

Data Availability
e data used to support the findings of this study are available from the corresponding author upon request.

Conflicts of Interest
e authors declare that they have no conflicts of interest.