# Research on Optimal Design of Pressure Relief Blasting Parameters of Middle Rock Pillar in Near Vertical Coal Seam

Authors: Yadi Wang; Shiqi Yu; Hai Rong; Liting Pan
Journal: Shock and Vibration
 (2022)

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## Abstract

In order to prevent the occurrence of rockburst accident caused by the stress concentration of the middle rock pillar in the near-vertical coal seam and affect the mining of the lower coal seam in the known goaf, taking Wudong coal mine in Xinjiang as an example, based on the special geological and mining conditions of the mine, through the simulation experiment and numerical analysis of similar materials, the variation law of the stress distribution of the middle rock pillar with the increase in mining depth is analyzed. A presplitting blasting pressure relief technology of drilling a construction chamber in the middle of the rock column is proposed to form a buffer layer in the rock wall and effectively prevent the continuous downward movement of stress. In order to achieve the optimal blasting effect, the finite element numerical simulation software GTS NX is used to carry out the chamber rock mass hole blasting simulation experiment, the three influencing factors of blasting effect, such as charge quantity, blast hole diameter, and blast hole row spacing, are comprehensively considered, and the optimal pressure relief blasting parameters with the radius of borehole blasting crack circle and the width of plastic zone as the index are determined. The experimental results show that under the condition of determining the physical and mechanical parameters of coal and rock, the blast hole diameter has become the main control factor affecting the blasting pressure relief effect, followed by the charge amount. Through 9 groups of orthogonal tests, the optimal combination of the three indexes is 6 kg charge, 150 mm hole diameter and 10 m hole row spacing. The practical engineering application results of the combination are tested by downhole transient electromagnetic detection (TEM), and it is found that the pressure relief effect is obvious. The preliminary conclusions of the study have guiding significance for the pressure relief and erosion prevention of the middle rock pillar in Wudong mine and provide experimental methods and theoretical basis for the determination of blasting pressure relief parameters in other mines.


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## Body

## 1. Introduction

Rockburst is a complex mine dynamic disaster, which often releases elastic properties in the form of sudden, sharp, and violent, resulting in instantaneous damage of coal and rock strata, accompanied by throwing out of coal and rock mass, loud noise, and gas wave, resulting in serious casualties and property losses [1–9]. According to statistics, by the end of 2020, the number of rockburst mines in China has exceeded 144, which has a serious impact on the smooth output of China’s coal resources [10–14]. Many scholars at home and abroad have conducted in-depth research on the prevention and control of rockburst and put forward a variety of feasible methods. At present, blasting pressure relief, as a more effective antiscour measure, is widely used in rockburst mines to solve the problem of stress concentration in roadway surrounding rock [15]. For the mechanism of blasting pressure relief and antiscour, Zhao [16] and others established a multiphysical field numerical analysis model for the whole process of coal and gas outburst. In the process of coal and gas outburst under high stress, the response laws of stress, gas pressure, temperature, and seepage in different areas and time nodes around coal and rock mass are discussed. The results show that firstly, the stress response law of coal and rock mass around the impact hole is initial vibration-sudden attenuation-later stability. Wang [9] and others studied the principle of crack development in the process of underwater blasting based on explosion and impact dynamics, fluid dynamics, fracture dynamics, and field tests. The research shows that in the process of underwater blasting, the fracture of surrounding rock is due to the joint action of shock wave and stress wave, which leads to the initial rock fracture and subsequent water expansion. The results of the deep hole underwater blasting test of large rock confirm the effective utilization of explosives in the hole, so as to improve the safety conditions; Dou et al. [17] put forward the theory of strength weakening and impact reduction, and expounded the mechanism of blasting pressure relief and impact prevention according to this theory. By using LS Dynan software for simulation, Guo et al. [18] analyzed the mechanism of upper wall pressure-related drilling and blasting and studied the effectiveness of drilling blasting. Si [19] and others first proposed the application of energy release blasting (ERB) technology in underwater drilling and blasting, and developed a new nondestructive damage assessment method of underwater blasting. The experimental results also show that the active sensing method based on a piezoelectric transducer combined with the signal processing method based on transmission energy can monitor and quantify the rock mass damage caused by underwater drilling and blasting. Wei [20] and others used electromagnetic radiation monitoring technology to analyze the pressure relief blasting parameters and pressure relief effect under different conditions; Luo [21] et ., based on the excavation of Sandaoshan Mine and using ANSYS/LS-DYNA dynamic finite element software, simulated the damage and fracture area of roadway surrounding rock under three pressure relief blasting schemes and reduced the strength of rock mass in the failure area, and then, calculated the in situ stress balance. The results show that the non-through fracture zone formed by pressure relief blasting will increase the stress concentration of roadway surrounding rock, which is not conducive to engineering safety. When a reasonable blasting scheme is designed to form a through fracture zone in the surrounding rock of the roadway, the concentrated high ground stress in the surrounding rock can be transmitted to the deep part of the rock mass, so as to reduce the elastic strain energy and reduce the risk of rockburst. Guo [22] and others focused on the criterion of coal seam fracture caused by cumulative blasting. On this basis, they summarized the fracture zone of coal, the crack propagation process, and the key technology of shaped charge blasting. In addition, the latest research progress of drilling and blasting parameter optimization in the cumulative blasting field test is also introduced. The results show that the mechanism of cumulative blasting to improve permeability can be further strengthened, and its process and technical equipment need to be improved. Liu [23] and others used a numerical simulation method to carry out the deep hole blasting test research on roadway side. The test results show that the physical and mechanical properties of coal play a decisive role in the blasting effect. Lai [24] and others revealed the evolution law of the energy field of the rock pillar through field exploration and physical similarity simulation experiments. Wu Zhenhua [25] and others analyzed the structure between the roof and the coal seam of the near-vertical coal seam and established a mechanical model. The results show that the joint action of the roof and rock pillar on the coal seam is the main reason for the occurrence of rockburst in the coal seam. Chang [26] et al. implemented three pressure relief blasting schemes in the fully mechanized top coal caving face of +450 level B3 + 6 coal seam in Wudong coal mine, namely, “deep hole is the main, shallow hole is the auxiliary,” deep and shallow holes are arranged alternately, and “shallow hole is the main, deep hole is the auxiliary.” The test results show that after the shallow rock blasting holes are densified, the daily average energy and frequency of micro-earthquakes are significantly reduced, the stress concentration near the working face is weakened, and the pressure relief effect is improved; Compared with deep rock holes, the stage pressure relief project with the alternating arrangement of deep and shallow holes and shallow rock holes is more effective to reduce the stress of the working face, and the pressure reduction in the area with alternating arrangement of deep and shallow holes is more obvious. Gao [27] and others analyzed the mechanism of rockburst induced by rock slab between coal seams by studying the relationship between microseismic events and stress concentration in coal and rock. Based on the elastic theory, the mechanical model of rock plate breaking between coal seams is established, and the energy calculation formula of rock plate breaking is obtained. Sun et al. [28] analyzed the stability of hard rock column (HRP) sandwiched between extremely dense coal seams (ESTCS). The research shows that the thermal effect of the acoustic emission (AE) test and infrared radiation thermal imaging (IRT) results is an index of the physical characteristics of rock mass fracture under high stress. Lei [29] and others verified the mechanism of coal pillar-induced rockburst by analyzing the distribution law of high-energy microseismic events and the source location of impact behavior. In view of this inducing mechanism, the measures of rock injection and coal rock deep hole blasting are applied to the shear zone, which can effectively reduce the stress concentration in the shear zone of coal pillar, make the coal rock release elastic performance smoothly, and reduce the impact risk. Xue [30]and others conducted triaxial compression tests of coal under different gas pressure conditions. Based on the experimental data, the mechanical characteristics, acoustic emission (AE) energy characteristics, and nonlinear characteristics of energy evolution of gas-bearing coal are obtained. It is found that the energy rate can be used as a new effective mechanical parameter to analyze and predict the damage and failure characteristics of coal. The energy dissipation characteristics before the peak can be divided into high dissipation rate type (hdert) and low dissipation rate type (ldert), which represent different failure modes (plastic failure and brittle failure). On this basis, the ratio of the dissipative energy rate to the input energy rate is further defined to effectively distinguish the two types of dissipative energy rates of coal. The research results are helpful to explore the fracturing evolution and energy-driving mechanism of the coal body, and then play a guiding role in the prediction of rock instability and the selection of support and reinforcement measures.It can be seen that scholars at home and abroad have performed a lot of research on the action mechanism and parameter design of blasting pressure relief, but there are many factors affecting the blasting pressure relief effect, and the geological conditions and stress field conditions of different mines are also different, and the research mostly focuses on the blasting loose pressure relief of soft rock roadway and hard roof slab. There is a lack of research on the blasting pressure relief of high-stress rock pillar in the middle of near-vertical coal seam and the detection of pressure relief effect after blasting. This article will combine the special stress field conditions that the maximum principal stress direction and the included angle of mining roadway are nearly vertical in Shenhua Xinjiang Wudong coal mine. Based on the numerical simulation method, the optimization design test of blasting pressure relief parameters is carried out to obtain the main control factors affecting the blasting effect and the optimal blasting parameter combination. The parameter combination is applied to the actual production process of Wudong coal mine, and the underground transient electromagnetic detection technology is used to test and verify the actual blasting effect, in order to provide reliable experimental data and theoretical basis for the prevention and control of rockburst in Wudong coal mine.


## 2. Influence Mechanism of Hard Rock Pillar on Rockburst

### 2.1. Occurrence Conditions of Rock Pillar and Coal Seam

The surface elevation of Wudong coal mine is +800 m, which is located in the south wing of Badaowan syncline and contains 32 coal layers. At present, the main coal seams areB1+2 and B3+6. The maximum thickness of B1+2 coal seam is 39.45 m, the minimum thickness is 31.83 m, and the average thickness is 37.45 m. The dip angle of the coal seam is 83°, belonging to a steeply inclined coal seam. The two groups of coal are separated by dikes, which gradually become thinner from west to east, ranging from 53 m to 110 m, with an average thickness of nearly 80 m. The lithology is mainly siltstone. The mining layout of working face in Wudong coal mine is shown in Figure 1.Figure 1
Mining layout of working face in Wudong coal mine.


### 2.2. Similar Material Simulation Experiment of Hard Rock Column

#### 2.2.1. Design and Fabrication of Model

The similar material simulation of fully mechanized top coal caving mining in Wudong coal mine is based on the measurement results of physical and mechanical parameters of coal and rock, based on the similar material theory, using similar materials such as sand, lime, and gypsum to make a model similar to the actual situation in the field according to the corresponding proportion, and then simulate the actual situation in the model for “mining,” and observe the movement of rock stratum on the model due to “mining” deformation and failure, and analyze and infer the situation of the field rock stratum, so as to further analyze and explain the causes of special ground pressure behavior in the process of coal seam mining.The simulation contents of similar materials are as follows:(1)
During the mining process of +627 m mining level to + 400 m mining level, the movement, deformation, failure, and stress change law of coal and rock strata on both sides of roadway inB3 + 6 working face;(2)
During the mining process of +627 m mining level to + 400 m mining level, the overburden structure collapse, displacement change law, and mining influence spatial relationship after mining inB1 + 2 working face.According to the approximate criteria of a similar material simulation experiment, it is determined that the length ratio of the engineering prototype to the model is 200, the density ratio is  1.5, the strength ratio is 300, and the time ratio is 14. Model filling dimension length × wide × height is 2000 mm × 240 mm × 2200 mm。 After calculation, the maximum principal stress at the bottom of the model is 21.4 MPa and the lateral pressure value is 10 MPa. Taking the fully mechanized top coal caving mining of Wudong coal mine as the engineering background, the mine field is divided into north-south mining areas with Badaowan syncline as the boundary. The coal seam dip angle in the south mining area is 80° ∼ 89° and 83° in most areas. The two groups of coal are separated by rock pillars, and the rock wall gradually thins from west to east, ranging from 53 m to 110 m. The overall effect of a similar material experiment is shown in Figure2.Figure 2
Overall effect of a similar material simulation experiment.Seven stress measuring points are arranged in the coal seam, and seven stress measuring points are arranged in the roof and floor. A total of 42 stress measurement points are arranged in the two coal seams. During mining, the BW-5 micropressure box embedded in the model is connected with YJZ-32A intelligent digital strain gauge and TST-3822A intelligent digital strain gauge through the lead, and then, the digital strain gauge is connected with the computer to collect the data signals of stress and strain changes in real time for processing and analysis. The layout of stress measuring points is shown in Figure3.Figure 3
Layout of stress measuring points.


#### 2.2.2. Stress Monitoring and Analysis

Combined with yjz-32a intelligent digital strain gauge and tst-3822a intelligent digital strain gauge, real-time stress monitoring is realized. Through the processing and analysis of monitoring data, the stress change in rock stratum can be obtained. After the mining of the two working faces with the level of + 522 m to + 450 m is completed, the stress change monitoring of the respective top and bottom plates of the working face is emphatically analyzed.Along the coal seam dip direction, the stress changes of rock pillar and B1 + 2 coal seam at the same level after mining inB3 + 6 working face at different levels are analyzed. Generally speaking, after the B3 + 6 working face of each mining level is mined, the stress changes of rock pillar and B1 + 2 coal seam at the same level are basically the same. Take the +500 m mining level as an example for detailed analysis, as shown in Figure 4.Figure 4
Profile of stress distribution law of stress monitoring points of +500 m horizontal working face.The original rock stress value of 500 m mining level is 7.51 MPa. In the mining process ofB3 + 6 working face, under the influence of mining activities, the stress value within 60 m of floor rock of B3 + 6 coal seam increases, with an increase range of 0.17 MPa∼0.50 MPa. The stress value in the range of 40 m from the roof rock of B1 + 2 coal seam to 10m from the floor of B1 + 2 coal seam decreases, with a decrease range of 0.12 MPa ∼0.34 MPa.It can be seen that the mining of +500 m horizontalB3 + 6 working face has a certain protective effect on B1+2 coal seam, and the maximum reduction range of stress is 4.20% of the original rock stress, but it also increases the stress value in the local area of the rock pillar, and the maximum increase range of stress is 6.17% of the original rock stress. The phenomenon of stress concentration in the rock column increases the risk of rockburst.


### 2.3. Analysis of Mechanical Properties of Rock Column

#### 2.3.1. Establishment of Structural Model of Rock Column Cantilever Beam

Taking the geological conditions of the south mining area of Wudong coal mine as the research background, with the continuous mining of B1 coal seam and B3 coal seam, the rock column will bend and deform under the action of self-weight stress. The rock column is regarded as a cantilever beam structure model, which can be calculated by the theoretical method of material mechanics. The structural mechanical model of rock column cantilever beam temporarily does not consider the influence of the combined action of goaf and coal seam on both sides of the rock column, but only considers the influence of the gravity factor of the rock column and analyzes the stress characteristics on rock column. The calculation model is shown in Figure5.Figure 5
Structural mechanical model of rock column cantilever beam.The equilibrium equation of the rock column cantilever beam structure model is as follows:(1)
Basic equation:(1)∑FX=0⇒FX=0,∑FY=0⇒Fy−γ⋅x⋅H⋅b=0,∑M=0⇒Mo−γ⋅x⋅H⋅b⋅X2⋅tanθ=0,Mo=γxHb⋅X2⋅tanθ.(2)
Bending moment equation:(2)∑MA=γ⋅X⋅H⋅bH2+x2,=γXHbH2+x2tanθ∑MB=γ⋅X⋅H⋅b⋅H2−x2tanθ=γXHbH2−x2tanθ,where γ is the average unit weight of overlying strata; σC is the mining depth of coal seam; H is the horizontal length of the rock column; b is the distance in the direction of rock pillar strike; θ is the included angle between inclination direction and vertical direction of the rock column; MA is the bending moment of gravity at point a; and MB is the bending moment of gravity at point B:(3)σ=M⋅yIZ,IZ=bh312,where M is the bending moment at any point on the cross section; y is the ordinate of any point on the cross section; IZ is the moment of inertia of the cross section to the neutral axis; b is the width of cross section; and h is the height of cross section.According to the above equation, the stress at the roadway floor of rock pillar B3 is expressed as(4)σA=3γXH2H+xtanθcosθbh3−γx.The stress at the roof of rock pillar B2 roadway is expressed as(5)σB=3γXH2H−xtanθcosθbh3=+γx.


#### 2.3.2. Calculation of Rock Column Stress Results

According to the actual mining situation of Wudong mine field, the width of the rock column is 100 m and the dip angle of the rock column is 83°. Calculate the stress distribution of the rock column. The rock column is brought into the known quantity by the above formula, and the different stress variation curves with the mining value are generated by Mathcad software. The results are shown in Figure6. In the figure, the negative sign only indicates that point a is subjected to tensile stress, and the positive sign indicates that point B is subjected to compressive stress. As can be seen from Figure 6, the value increases with the increase in the x value. The calculation shows that the stress increases linearly with the increase in the mining depth.Figure 6
Comparison with different stress variation curves with thex value.


#### 2.3.3. Simulation Results of Static Distribution of Rock Column

The finite element numerical simulation software GTS NX is used to simulate the stress distribution state of the rock column under the condition of only self-weight stress. GTS NX is a general finite element analysis software developed for the geotechnical field, which is suitable for accurate modeling and analysis of various practical projects such as subway, tunnel, slope, foundation pit, pile foundation, hydraulic engineering, and mine. According to the actual mining situation of Wudong mine field, the width of the rock column is taken as 100 m and the dip angle of the rock column is taken as 83°. The analysis results are shown in Figures7 and 8.Figure 7
Stress distribution nephogram of the rock column under self-weight stress.Figure 8
Displacement nephogram of the rock column under self-weight stress.It can be seen that under the condition of only considering the self-weight stress, there is a high-stress area in the rock pillar, which will cause extrusion deformation to the roadway layout of coal seam mining. In order to accurately grasp the stress change of rock pillar in the mining process and carry out risk relief work in time, the forces of coal seam and goaf are applied to the side of rock pillar, and the mining depths are 300 m, 350 m, and 400 m. Horizontal displacement of the rock column is at 450 m.Figures9–12 show the deformation of the rock column when the mining depth is 300 m, 350 m, 400 m, and 450 m. As can be seen from the figure, when the mining depth is 300 m, the horizontal displacement in the rock column is 0.24 m. At a mining depth of 350 m, the horizontal displacement in the rock column is 0.44 m. When the mining depth is 400 m, the horizontal stress in the rock column is 0.75 m. At a mining depth of 450 m, the horizontal displacement in the rock column is 1.21 m. Theoretically, the greater the mining depth, the greater the deformation of the rock column. Under the coupling action of the maximum horizontal principal stress and the tensile stress on both sides of rock column elastic beam, the roadway stability is poor, the compressive stress of roadway top and floor is large, the floor heave is serious, and the deformation of two sides of roadway is serious under the action of tensile stress.Figure 9
Horizontal displacement of the rock column at a mining depth of 300 m.Figure 10
Horizontal displacement of the rock column at a mining depth of 350 m.Figure 11
Horizontal displacement of the rock column at a mining depth of 400 m.Figure 12
Horizontal displacement of the rock column at a mining depth of 450 m.Using the methods of a similar material simulation experiment, theoretical analysis, and numerical simulation, this article analyzes the change of stress concentration degree of hard rock pillar with coal seam mining depth in Wudong coal mine and preliminarily obtains that the stress concentration degree of hard rock pillar is positively correlated with coal seam mining depth. At the same time, the occurrence of Badaowan syncline will cause a further increase in coal stress and further accumulation of energy in Wudong coalfield, which will further increase the risk of rockburst in Wudong coal mine [12].According to the above analysis results, presplitting blasting and other measures can be adopted for hard rock columns to reduce the degree of stress concentration and provide guarantee for safe mining of the working face. In this article, Midas numerical simulation software is used to analyze the pressure relief effect and blasting parameters of a hard rock column in detail.


## 2.1. Occurrence Conditions of Rock Pillar and Coal Seam

The surface elevation of Wudong coal mine is +800 m, which is located in the south wing of Badaowan syncline and contains 32 coal layers. At present, the main coal seams areB1+2 and B3+6. The maximum thickness of B1+2 coal seam is 39.45 m, the minimum thickness is 31.83 m, and the average thickness is 37.45 m. The dip angle of the coal seam is 83°, belonging to a steeply inclined coal seam. The two groups of coal are separated by dikes, which gradually become thinner from west to east, ranging from 53 m to 110 m, with an average thickness of nearly 80 m. The lithology is mainly siltstone. The mining layout of working face in Wudong coal mine is shown in Figure 1.Figure 1
Mining layout of working face in Wudong coal mine.


## 2.2. Similar Material Simulation Experiment of Hard Rock Column

### 2.2.1. Design and Fabrication of Model

The similar material simulation of fully mechanized top coal caving mining in Wudong coal mine is based on the measurement results of physical and mechanical parameters of coal and rock, based on the similar material theory, using similar materials such as sand, lime, and gypsum to make a model similar to the actual situation in the field according to the corresponding proportion, and then simulate the actual situation in the model for “mining,” and observe the movement of rock stratum on the model due to “mining” deformation and failure, and analyze and infer the situation of the field rock stratum, so as to further analyze and explain the causes of special ground pressure behavior in the process of coal seam mining.The simulation contents of similar materials are as follows:(1)
During the mining process of +627 m mining level to + 400 m mining level, the movement, deformation, failure, and stress change law of coal and rock strata on both sides of roadway inB3 + 6 working face;(2)
During the mining process of +627 m mining level to + 400 m mining level, the overburden structure collapse, displacement change law, and mining influence spatial relationship after mining inB1 + 2 working face.According to the approximate criteria of a similar material simulation experiment, it is determined that the length ratio of the engineering prototype to the model is 200, the density ratio is  1.5, the strength ratio is 300, and the time ratio is 14. Model filling dimension length × wide × height is 2000 mm × 240 mm × 2200 mm。 After calculation, the maximum principal stress at the bottom of the model is 21.4 MPa and the lateral pressure value is 10 MPa. Taking the fully mechanized top coal caving mining of Wudong coal mine as the engineering background, the mine field is divided into north-south mining areas with Badaowan syncline as the boundary. The coal seam dip angle in the south mining area is 80° ∼ 89° and 83° in most areas. The two groups of coal are separated by rock pillars, and the rock wall gradually thins from west to east, ranging from 53 m to 110 m. The overall effect of a similar material experiment is shown in Figure2.Figure 2
Overall effect of a similar material simulation experiment.Seven stress measuring points are arranged in the coal seam, and seven stress measuring points are arranged in the roof and floor. A total of 42 stress measurement points are arranged in the two coal seams. During mining, the BW-5 micropressure box embedded in the model is connected with YJZ-32A intelligent digital strain gauge and TST-3822A intelligent digital strain gauge through the lead, and then, the digital strain gauge is connected with the computer to collect the data signals of stress and strain changes in real time for processing and analysis. The layout of stress measuring points is shown in Figure3.Figure 3
Layout of stress measuring points.


### 2.2.2. Stress Monitoring and Analysis

Combined with yjz-32a intelligent digital strain gauge and tst-3822a intelligent digital strain gauge, real-time stress monitoring is realized. Through the processing and analysis of monitoring data, the stress change in rock stratum can be obtained. After the mining of the two working faces with the level of + 522 m to + 450 m is completed, the stress change monitoring of the respective top and bottom plates of the working face is emphatically analyzed.Along the coal seam dip direction, the stress changes of rock pillar and B1 + 2 coal seam at the same level after mining inB3 + 6 working face at different levels are analyzed. Generally speaking, after the B3 + 6 working face of each mining level is mined, the stress changes of rock pillar and B1 + 2 coal seam at the same level are basically the same. Take the +500 m mining level as an example for detailed analysis, as shown in Figure 4.Figure 4
Profile of stress distribution law of stress monitoring points of +500 m horizontal working face.The original rock stress value of 500 m mining level is 7.51 MPa. In the mining process ofB3 + 6 working face, under the influence of mining activities, the stress value within 60 m of floor rock of B3 + 6 coal seam increases, with an increase range of 0.17 MPa∼0.50 MPa. The stress value in the range of 40 m from the roof rock of B1 + 2 coal seam to 10m from the floor of B1 + 2 coal seam decreases, with a decrease range of 0.12 MPa ∼0.34 MPa.It can be seen that the mining of +500 m horizontalB3 + 6 working face has a certain protective effect on B1+2 coal seam, and the maximum reduction range of stress is 4.20% of the original rock stress, but it also increases the stress value in the local area of the rock pillar, and the maximum increase range of stress is 6.17% of the original rock stress. The phenomenon of stress concentration in the rock column increases the risk of rockburst.


## 2.2.1. Design and Fabrication of Model

The similar material simulation of fully mechanized top coal caving mining in Wudong coal mine is based on the measurement results of physical and mechanical parameters of coal and rock, based on the similar material theory, using similar materials such as sand, lime, and gypsum to make a model similar to the actual situation in the field according to the corresponding proportion, and then simulate the actual situation in the model for “mining,” and observe the movement of rock stratum on the model due to “mining” deformation and failure, and analyze and infer the situation of the field rock stratum, so as to further analyze and explain the causes of special ground pressure behavior in the process of coal seam mining.The simulation contents of similar materials are as follows:(1)
During the mining process of +627 m mining level to + 400 m mining level, the movement, deformation, failure, and stress change law of coal and rock strata on both sides of roadway inB3 + 6 working face;(2)
During the mining process of +627 m mining level to + 400 m mining level, the overburden structure collapse, displacement change law, and mining influence spatial relationship after mining inB1 + 2 working face.According to the approximate criteria of a similar material simulation experiment, it is determined that the length ratio of the engineering prototype to the model is 200, the density ratio is  1.5, the strength ratio is 300, and the time ratio is 14. Model filling dimension length × wide × height is 2000 mm × 240 mm × 2200 mm。 After calculation, the maximum principal stress at the bottom of the model is 21.4 MPa and the lateral pressure value is 10 MPa. Taking the fully mechanized top coal caving mining of Wudong coal mine as the engineering background, the mine field is divided into north-south mining areas with Badaowan syncline as the boundary. The coal seam dip angle in the south mining area is 80° ∼ 89° and 83° in most areas. The two groups of coal are separated by rock pillars, and the rock wall gradually thins from west to east, ranging from 53 m to 110 m. The overall effect of a similar material experiment is shown in Figure2.Figure 2
Overall effect of a similar material simulation experiment.Seven stress measuring points are arranged in the coal seam, and seven stress measuring points are arranged in the roof and floor. A total of 42 stress measurement points are arranged in the two coal seams. During mining, the BW-5 micropressure box embedded in the model is connected with YJZ-32A intelligent digital strain gauge and TST-3822A intelligent digital strain gauge through the lead, and then, the digital strain gauge is connected with the computer to collect the data signals of stress and strain changes in real time for processing and analysis. The layout of stress measuring points is shown in Figure3.Figure 3
Layout of stress measuring points.


## 2.2.2. Stress Monitoring and Analysis

Combined with yjz-32a intelligent digital strain gauge and tst-3822a intelligent digital strain gauge, real-time stress monitoring is realized. Through the processing and analysis of monitoring data, the stress change in rock stratum can be obtained. After the mining of the two working faces with the level of + 522 m to + 450 m is completed, the stress change monitoring of the respective top and bottom plates of the working face is emphatically analyzed.Along the coal seam dip direction, the stress changes of rock pillar and B1 + 2 coal seam at the same level after mining inB3 + 6 working face at different levels are analyzed. Generally speaking, after the B3 + 6 working face of each mining level is mined, the stress changes of rock pillar and B1 + 2 coal seam at the same level are basically the same. Take the +500 m mining level as an example for detailed analysis, as shown in Figure 4.Figure 4
Profile of stress distribution law of stress monitoring points of +500 m horizontal working face.The original rock stress value of 500 m mining level is 7.51 MPa. In the mining process ofB3 + 6 working face, under the influence of mining activities, the stress value within 60 m of floor rock of B3 + 6 coal seam increases, with an increase range of 0.17 MPa∼0.50 MPa. The stress value in the range of 40 m from the roof rock of B1 + 2 coal seam to 10m from the floor of B1 + 2 coal seam decreases, with a decrease range of 0.12 MPa ∼0.34 MPa.It can be seen that the mining of +500 m horizontalB3 + 6 working face has a certain protective effect on B1+2 coal seam, and the maximum reduction range of stress is 4.20% of the original rock stress, but it also increases the stress value in the local area of the rock pillar, and the maximum increase range of stress is 6.17% of the original rock stress. The phenomenon of stress concentration in the rock column increases the risk of rockburst.


## 2.3. Analysis of Mechanical Properties of Rock Column

### 2.3.1. Establishment of Structural Model of Rock Column Cantilever Beam

Taking the geological conditions of the south mining area of Wudong coal mine as the research background, with the continuous mining of B1 coal seam and B3 coal seam, the rock column will bend and deform under the action of self-weight stress. The rock column is regarded as a cantilever beam structure model, which can be calculated by the theoretical method of material mechanics. The structural mechanical model of rock column cantilever beam temporarily does not consider the influence of the combined action of goaf and coal seam on both sides of the rock column, but only considers the influence of the gravity factor of the rock column and analyzes the stress characteristics on rock column. The calculation model is shown in Figure5.Figure 5
Structural mechanical model of rock column cantilever beam.The equilibrium equation of the rock column cantilever beam structure model is as follows:(1)
Basic equation:(1)∑FX=0⇒FX=0,∑FY=0⇒Fy−γ⋅x⋅H⋅b=0,∑M=0⇒Mo−γ⋅x⋅H⋅b⋅X2⋅tanθ=0,Mo=γxHb⋅X2⋅tanθ.(2)
Bending moment equation:(2)∑MA=γ⋅X⋅H⋅bH2+x2,=γXHbH2+x2tanθ∑MB=γ⋅X⋅H⋅b⋅H2−x2tanθ=γXHbH2−x2tanθ,where γ is the average unit weight of overlying strata; σC is the mining depth of coal seam; H is the horizontal length of the rock column; b is the distance in the direction of rock pillar strike; θ is the included angle between inclination direction and vertical direction of the rock column; MA is the bending moment of gravity at point a; and MB is the bending moment of gravity at point B:(3)σ=M⋅yIZ,IZ=bh312,where M is the bending moment at any point on the cross section; y is the ordinate of any point on the cross section; IZ is the moment of inertia of the cross section to the neutral axis; b is the width of cross section; and h is the height of cross section.According to the above equation, the stress at the roadway floor of rock pillar B3 is expressed as(4)σA=3γXH2H+xtanθcosθbh3−γx.The stress at the roof of rock pillar B2 roadway is expressed as(5)σB=3γXH2H−xtanθcosθbh3=+γx.


### 2.3.2. Calculation of Rock Column Stress Results

According to the actual mining situation of Wudong mine field, the width of the rock column is 100 m and the dip angle of the rock column is 83°. Calculate the stress distribution of the rock column. The rock column is brought into the known quantity by the above formula, and the different stress variation curves with the mining value are generated by Mathcad software. The results are shown in Figure6. In the figure, the negative sign only indicates that point a is subjected to tensile stress, and the positive sign indicates that point B is subjected to compressive stress. As can be seen from Figure 6, the value increases with the increase in the x value. The calculation shows that the stress increases linearly with the increase in the mining depth.Figure 6
Comparison with different stress variation curves with thex value.


### 2.3.3. Simulation Results of Static Distribution of Rock Column

The finite element numerical simulation software GTS NX is used to simulate the stress distribution state of the rock column under the condition of only self-weight stress. GTS NX is a general finite element analysis software developed for the geotechnical field, which is suitable for accurate modeling and analysis of various practical projects such as subway, tunnel, slope, foundation pit, pile foundation, hydraulic engineering, and mine. According to the actual mining situation of Wudong mine field, the width of the rock column is taken as 100 m and the dip angle of the rock column is taken as 83°. The analysis results are shown in Figures7 and 8.Figure 7
Stress distribution nephogram of the rock column under self-weight stress.Figure 8
Displacement nephogram of the rock column under self-weight stress.It can be seen that under the condition of only considering the self-weight stress, there is a high-stress area in the rock pillar, which will cause extrusion deformation to the roadway layout of coal seam mining. In order to accurately grasp the stress change of rock pillar in the mining process and carry out risk relief work in time, the forces of coal seam and goaf are applied to the side of rock pillar, and the mining depths are 300 m, 350 m, and 400 m. Horizontal displacement of the rock column is at 450 m.Figures9–12 show the deformation of the rock column when the mining depth is 300 m, 350 m, 400 m, and 450 m. As can be seen from the figure, when the mining depth is 300 m, the horizontal displacement in the rock column is 0.24 m. At a mining depth of 350 m, the horizontal displacement in the rock column is 0.44 m. When the mining depth is 400 m, the horizontal stress in the rock column is 0.75 m. At a mining depth of 450 m, the horizontal displacement in the rock column is 1.21 m. Theoretically, the greater the mining depth, the greater the deformation of the rock column. Under the coupling action of the maximum horizontal principal stress and the tensile stress on both sides of rock column elastic beam, the roadway stability is poor, the compressive stress of roadway top and floor is large, the floor heave is serious, and the deformation of two sides of roadway is serious under the action of tensile stress.Figure 9
Horizontal displacement of the rock column at a mining depth of 300 m.Figure 10
Horizontal displacement of the rock column at a mining depth of 350 m.Figure 11
Horizontal displacement of the rock column at a mining depth of 400 m.Figure 12
Horizontal displacement of the rock column at a mining depth of 450 m.Using the methods of a similar material simulation experiment, theoretical analysis, and numerical simulation, this article analyzes the change of stress concentration degree of hard rock pillar with coal seam mining depth in Wudong coal mine and preliminarily obtains that the stress concentration degree of hard rock pillar is positively correlated with coal seam mining depth. At the same time, the occurrence of Badaowan syncline will cause a further increase in coal stress and further accumulation of energy in Wudong coalfield, which will further increase the risk of rockburst in Wudong coal mine [12].According to the above analysis results, presplitting blasting and other measures can be adopted for hard rock columns to reduce the degree of stress concentration and provide guarantee for safe mining of the working face. In this article, Midas numerical simulation software is used to analyze the pressure relief effect and blasting parameters of a hard rock column in detail.


## 2.3.1. Establishment of Structural Model of Rock Column Cantilever Beam

Taking the geological conditions of the south mining area of Wudong coal mine as the research background, with the continuous mining of B1 coal seam and B3 coal seam, the rock column will bend and deform under the action of self-weight stress. The rock column is regarded as a cantilever beam structure model, which can be calculated by the theoretical method of material mechanics. The structural mechanical model of rock column cantilever beam temporarily does not consider the influence of the combined action of goaf and coal seam on both sides of the rock column, but only considers the influence of the gravity factor of the rock column and analyzes the stress characteristics on rock column. The calculation model is shown in Figure5.Figure 5
Structural mechanical model of rock column cantilever beam.The equilibrium equation of the rock column cantilever beam structure model is as follows:(1)
Basic equation:(1)∑FX=0⇒FX=0,∑FY=0⇒Fy−γ⋅x⋅H⋅b=0,∑M=0⇒Mo−γ⋅x⋅H⋅b⋅X2⋅tanθ=0,Mo=γxHb⋅X2⋅tanθ.(2)
Bending moment equation:(2)∑MA=γ⋅X⋅H⋅bH2+x2,=γXHbH2+x2tanθ∑MB=γ⋅X⋅H⋅b⋅H2−x2tanθ=γXHbH2−x2tanθ,where γ is the average unit weight of overlying strata; σC is the mining depth of coal seam; H is the horizontal length of the rock column; b is the distance in the direction of rock pillar strike; θ is the included angle between inclination direction and vertical direction of the rock column; MA is the bending moment of gravity at point a; and MB is the bending moment of gravity at point B:(3)σ=M⋅yIZ,IZ=bh312,where M is the bending moment at any point on the cross section; y is the ordinate of any point on the cross section; IZ is the moment of inertia of the cross section to the neutral axis; b is the width of cross section; and h is the height of cross section.According to the above equation, the stress at the roadway floor of rock pillar B3 is expressed as(4)σA=3γXH2H+xtanθcosθbh3−γx.The stress at the roof of rock pillar B2 roadway is expressed as(5)σB=3γXH2H−xtanθcosθbh3=+γx.


## 2.3.2. Calculation of Rock Column Stress Results

According to the actual mining situation of Wudong mine field, the width of the rock column is 100 m and the dip angle of the rock column is 83°. Calculate the stress distribution of the rock column. The rock column is brought into the known quantity by the above formula, and the different stress variation curves with the mining value are generated by Mathcad software. The results are shown in Figure6. In the figure, the negative sign only indicates that point a is subjected to tensile stress, and the positive sign indicates that point B is subjected to compressive stress. As can be seen from Figure 6, the value increases with the increase in the x value. The calculation shows that the stress increases linearly with the increase in the mining depth.Figure 6
Comparison with different stress variation curves with thex value.


## 2.3.3. Simulation Results of Static Distribution of Rock Column

The finite element numerical simulation software GTS NX is used to simulate the stress distribution state of the rock column under the condition of only self-weight stress. GTS NX is a general finite element analysis software developed for the geotechnical field, which is suitable for accurate modeling and analysis of various practical projects such as subway, tunnel, slope, foundation pit, pile foundation, hydraulic engineering, and mine. According to the actual mining situation of Wudong mine field, the width of the rock column is taken as 100 m and the dip angle of the rock column is taken as 83°. The analysis results are shown in Figures7 and 8.Figure 7
Stress distribution nephogram of the rock column under self-weight stress.Figure 8
Displacement nephogram of the rock column under self-weight stress.It can be seen that under the condition of only considering the self-weight stress, there is a high-stress area in the rock pillar, which will cause extrusion deformation to the roadway layout of coal seam mining. In order to accurately grasp the stress change of rock pillar in the mining process and carry out risk relief work in time, the forces of coal seam and goaf are applied to the side of rock pillar, and the mining depths are 300 m, 350 m, and 400 m. Horizontal displacement of the rock column is at 450 m.Figures9–12 show the deformation of the rock column when the mining depth is 300 m, 350 m, 400 m, and 450 m. As can be seen from the figure, when the mining depth is 300 m, the horizontal displacement in the rock column is 0.24 m. At a mining depth of 350 m, the horizontal displacement in the rock column is 0.44 m. When the mining depth is 400 m, the horizontal stress in the rock column is 0.75 m. At a mining depth of 450 m, the horizontal displacement in the rock column is 1.21 m. Theoretically, the greater the mining depth, the greater the deformation of the rock column. Under the coupling action of the maximum horizontal principal stress and the tensile stress on both sides of rock column elastic beam, the roadway stability is poor, the compressive stress of roadway top and floor is large, the floor heave is serious, and the deformation of two sides of roadway is serious under the action of tensile stress.Figure 9
Horizontal displacement of the rock column at a mining depth of 300 m.Figure 10
Horizontal displacement of the rock column at a mining depth of 350 m.Figure 11
Horizontal displacement of the rock column at a mining depth of 400 m.Figure 12
Horizontal displacement of the rock column at a mining depth of 450 m.Using the methods of a similar material simulation experiment, theoretical analysis, and numerical simulation, this article analyzes the change of stress concentration degree of hard rock pillar with coal seam mining depth in Wudong coal mine and preliminarily obtains that the stress concentration degree of hard rock pillar is positively correlated with coal seam mining depth. At the same time, the occurrence of Badaowan syncline will cause a further increase in coal stress and further accumulation of energy in Wudong coalfield, which will further increase the risk of rockburst in Wudong coal mine [12].According to the above analysis results, presplitting blasting and other measures can be adopted for hard rock columns to reduce the degree of stress concentration and provide guarantee for safe mining of the working face. In this article, Midas numerical simulation software is used to analyze the pressure relief effect and blasting parameters of a hard rock column in detail.


## 3. Design of Numerical Simulation Experiment for Pressure Relief of Rock Pillar Blasting

### 3.1. Analysis of Blasting Pressure Relief and Antiscour Mechanism

Pressure relief blasting belongs to internal blasting. After blasting, the shock wave first destroys the coal and rock mass, and then, the gas generated by blasting further ruptures the coal and rock mass [31]. Due to the action of air pressure, tangential tensile stress is formed, resulting in radial tensile fracture. After the pressure relief blasting, the coal and rock mass around the blasting hole wall forms a fracture zone and nondestructive disturbance zone (Figure 13). After pressure relief blasting, the strength and energy conditions of rockburst in coal and rock mass are destroyed due to the existence of crushing area, forming a certain unloading area, weakening or eliminating the risk of rockburst in coal [32].Figure 13
Borehole division of pressure relief blasting in coal seam.


### 3.2. Orthogonal Experimental Design for Optimization of Blasting Pressure Relief Parameters

#### 3.2.1. Establishment of Model

In order to prevent the occurrence of rockburst caused by the stress concentration of vertical rock column, combined with the experience of +500 m horizontal rockburst prevention and control, it shows that the presplitting blasting of the rock column in the construction chamber in the middle of the rock column can effectively release the stress, form a buffer layer in the rock wall, and effectively prevent the stress from moving down.As shown in Figure14, the numerical simulation software GTS NX is used to establish the chamber wall blasting pressure relief model with a model size of 50 m × 50 m×30 m. The number of nodes of the model is 51791, and the number of elements is 68001. The arrangement direction of blast holes is parallel to the direction of roadway strike, arranged in a single row, and the blast holes adopt centralized and continuous charging. As the blasting is powerful and destructive, so as to prevent damage to the mining roadway, the charging section of the blast hole shall be left with a certain length of resistance line from the coal wall of the working face. At the same time, millimeter millisecond blasting shall be adopted to reduce blasting impact and vibration [33].Figure 14
Blasting hole model of chamber rock mass.When defining the boundary conditions of the model, the establishment of general boundary conditions for dynamic analysis will produce great errors due to the reflection of waves. Therefore, in order to solve this problem, the model uses the viscous boundary proposed by Lysmer and Wass in 1972 to calculate the damping ratio of the corresponding soil inx, y, and z directions to define the viscous boundary, and uses the vibration mode damping calculation. The mechanical parameters of coal and rock required by the model are listed in Table 1.Table 1
Summary of measurement results of physical and mechanical property parameters of +500 m horizontal coal and rock samples.
Coal and rock parametersB3 coal seamB2 coal seamB3 coal seam floor rockB2 coal seam roof rockApparent density (kg/m³)1284125324782692Internal friction angle (°)24.0827.5466.4438.42Cohesion (MPa)2.41.75.46.9Modulus of elasticity (GPa)0.2040.2091.6742.343Poisson’s ratio0.210.190.230.22For the general blasting elastic analysis, the blasting pressure acts on the vertical direction of the hole wall, and the load used at this time uses the formula mentioned in the National Highway Institute of the United States. The blasting load per 1 kg is as follows:(6)Pdet=4.18×10−7×Sge×Ve21+0.8Sge,pB=Pdet×dcdh3,where Pdet is the bursting pressure; pB is the pressure on hole wall; Ve is the blasting speed; dc is the powder diameter; dh is the hole diameter; and Sge is the specific gravity.The above formula determines the aerodynamic pressure during blasting. B = 16338 is the load constant and the dynamic pressure per 1 kg charge.


#### 3.2.2. Orthogonal Experimental Scheme Design

Because there are many factors affecting the blasting pressure relief effect, it is impossible to carry out a single control variable test. Therefore, an efficient, fast, and economic orthogonal experimental method is introduced, which selects some representative points from the comprehensive test according to the orthogonality [34]. Nine groups of simulation experiments are designed with the charge quantity, The experimental level table of orthogonal factors is shown in Table 2. The design of the 3-factor 3-level test with the blast hole diameter and blast hole row spacing as the influencing factors, as shown in Figure 15, is shown in Table 3,Table 2
Factor level of the orthogonal test.
LevelFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010Figure 15
Effect diagram of the simulation experiment.Table 3
Design of factor 3 level test.
Experimental schemeFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010421201054150666908721508849010961206


### 3.3. Analysis of Simulation Test Results

The advantages and disadvantages of blasting effect are mainly determined according to two aspects: first, the range of the plastic zone formed after blasting; and second, the diameter of the plastic zone after blasting shall be slightly larger than the blast hole spacing to ensure that the range of plastic zone is connected with each other; at the same time, the radius of plastic zone shall be larger than the blast hole spacing to avoid excessive overlap of fracture circle and affect the blasting effect. However, considering the impact analysis of underground blasting safety, it will not damage the stability of mining roadway, and the charge must be within the controllable range. At the same time, a reasonable blast hole row spacing will also make the cracks between blast holes better connected together. The relationship between the variables and the radius of the plastic zone is shown in Figure16.Figure 16
Relationship between various factors and the radius of the plastic zone.The following conclusions can be drawn from the information reflected in Figure16:(1)
Under the condition of certain charge and hole spacing, the range of the plastic zone increases obviously with the increase in the hole diameter.(2)
Under the condition of constant blast hole diameter, with the increase in charge quantity and blast hole spacing, it can also be seen that the radius of the plastic zone tends to increase, but the increasing trend is not obvious due to the influence of each other’s blast hole blasting. It can be seen that charge quantity and blast hole spacing become the control factors in turn.(3)
According to the comprehensive analysis of the above nine groups of simulation experiment effect drawings, it can be seen that No. 3 experiment with a blast hole diameter of 150 mm, a blast hole spacing of 10 m, and a charge of 6 kg has the best blasting effect, and the plastic area produced by blasting is large. Taking the blast hole as the center of the circle, it is measured that the radius of the plastic area produced by a single blast hole is 5.33 m, and the crack area between the two blast holes overlaps, and the mold is not smooth. The proposed effect is good, and the pressure relief effect is obvious.According to the simulation results of the software, a reasonable blasting pressure relief parameter scheme is designed. The blast hole diameter, as the main control factor, can effectively adjust the plastic area range after blasting and control it between 80 and 120 mm as far as possible. In order to avoid the interaction between adjacent blast holes, the blasting test of a single blast hole can be carried out to determine the plastic area range, so as to determine the reasonable blast hole spacing and charge quantity.


### 3.4. Test of Simulation Results Based on Coal Rock Power System

The existence of geological structures such as activated faults and synclines and the rapid change of coal and rock mass thickness will cause the uneven distribution of stress and energy in the stratum, resulting in the emergence of high-stress and high-energy areas in local coal and rock mass, forming a “coal rock dynamic system.” [35] Coal rock dynamic system is an open and dissipative system with complex structure and behavior characteristics, which is characterized by high complexity and nonlinearity. Rockburst is a dynamic process of energy accumulation and release in the coal rock mass system, and is the result of the combined action of geological dynamic conditions and mining disturbance [36-37].In this article, explosive explosion is regarded as the energy source of coal and rock power system, and the accuracy of simulation results is tested through the coal mine coal and rock power system analysis software independently developed by the team.The unit explosive explosion can produce 400 kJ of energy. Input the rock mass mechanical parameters and energy into the software to calculate the corresponding radius of the damaged area. The radius of the damage area caused by 2 kg, 4 kg, and 6 kg explosives is shown in Table4.Table 4
Calculation results of the coal rock power system.
Explosive quantity (kg)Energy generated (J)Radius of failure zone (m)Radius of plastic zone in simulation experiment (m)Error value (％)24000004.264.44448000004.664.591.5612000005.034.961.4In order to reduce the influence of other factors on the radius of the failure area, three groups of experiments of 120 mm and 8 m with blast hole diameter and blast hole spacing as the middle value are selected for verification. The error between the results and the calculation results of the coal rock dynamic system is shown in Table4. It can be seen from the calculation results that the error between the plastic zone radius corresponding to different charge quantities in the simulation experiment and the verification results of the coal rock dynamic system is not more than 5%, which shows that the reliability of the simulation experiment results is high.


## 3.1. Analysis of Blasting Pressure Relief and Antiscour Mechanism

Pressure relief blasting belongs to internal blasting. After blasting, the shock wave first destroys the coal and rock mass, and then, the gas generated by blasting further ruptures the coal and rock mass [31]. Due to the action of air pressure, tangential tensile stress is formed, resulting in radial tensile fracture. After the pressure relief blasting, the coal and rock mass around the blasting hole wall forms a fracture zone and nondestructive disturbance zone (Figure 13). After pressure relief blasting, the strength and energy conditions of rockburst in coal and rock mass are destroyed due to the existence of crushing area, forming a certain unloading area, weakening or eliminating the risk of rockburst in coal [32].Figure 13
Borehole division of pressure relief blasting in coal seam.


## 3.2. Orthogonal Experimental Design for Optimization of Blasting Pressure Relief Parameters

### 3.2.1. Establishment of Model

In order to prevent the occurrence of rockburst caused by the stress concentration of vertical rock column, combined with the experience of +500 m horizontal rockburst prevention and control, it shows that the presplitting blasting of the rock column in the construction chamber in the middle of the rock column can effectively release the stress, form a buffer layer in the rock wall, and effectively prevent the stress from moving down.As shown in Figure14, the numerical simulation software GTS NX is used to establish the chamber wall blasting pressure relief model with a model size of 50 m × 50 m×30 m. The number of nodes of the model is 51791, and the number of elements is 68001. The arrangement direction of blast holes is parallel to the direction of roadway strike, arranged in a single row, and the blast holes adopt centralized and continuous charging. As the blasting is powerful and destructive, so as to prevent damage to the mining roadway, the charging section of the blast hole shall be left with a certain length of resistance line from the coal wall of the working face. At the same time, millimeter millisecond blasting shall be adopted to reduce blasting impact and vibration [33].Figure 14
Blasting hole model of chamber rock mass.When defining the boundary conditions of the model, the establishment of general boundary conditions for dynamic analysis will produce great errors due to the reflection of waves. Therefore, in order to solve this problem, the model uses the viscous boundary proposed by Lysmer and Wass in 1972 to calculate the damping ratio of the corresponding soil inx, y, and z directions to define the viscous boundary, and uses the vibration mode damping calculation. The mechanical parameters of coal and rock required by the model are listed in Table 1.Table 1
Summary of measurement results of physical and mechanical property parameters of +500 m horizontal coal and rock samples.
Coal and rock parametersB3 coal seamB2 coal seamB3 coal seam floor rockB2 coal seam roof rockApparent density (kg/m³)1284125324782692Internal friction angle (°)24.0827.5466.4438.42Cohesion (MPa)2.41.75.46.9Modulus of elasticity (GPa)0.2040.2091.6742.343Poisson’s ratio0.210.190.230.22For the general blasting elastic analysis, the blasting pressure acts on the vertical direction of the hole wall, and the load used at this time uses the formula mentioned in the National Highway Institute of the United States. The blasting load per 1 kg is as follows:(6)Pdet=4.18×10−7×Sge×Ve21+0.8Sge,pB=Pdet×dcdh3,where Pdet is the bursting pressure; pB is the pressure on hole wall; Ve is the blasting speed; dc is the powder diameter; dh is the hole diameter; and Sge is the specific gravity.The above formula determines the aerodynamic pressure during blasting. B = 16338 is the load constant and the dynamic pressure per 1 kg charge.


### 3.2.2. Orthogonal Experimental Scheme Design

Because there are many factors affecting the blasting pressure relief effect, it is impossible to carry out a single control variable test. Therefore, an efficient, fast, and economic orthogonal experimental method is introduced, which selects some representative points from the comprehensive test according to the orthogonality [34]. Nine groups of simulation experiments are designed with the charge quantity, The experimental level table of orthogonal factors is shown in Table 2. The design of the 3-factor 3-level test with the blast hole diameter and blast hole row spacing as the influencing factors, as shown in Figure 15, is shown in Table 3,Table 2
Factor level of the orthogonal test.
LevelFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010Figure 15
Effect diagram of the simulation experiment.Table 3
Design of factor 3 level test.
Experimental schemeFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010421201054150666908721508849010961206


## 3.2.1. Establishment of Model

In order to prevent the occurrence of rockburst caused by the stress concentration of vertical rock column, combined with the experience of +500 m horizontal rockburst prevention and control, it shows that the presplitting blasting of the rock column in the construction chamber in the middle of the rock column can effectively release the stress, form a buffer layer in the rock wall, and effectively prevent the stress from moving down.As shown in Figure14, the numerical simulation software GTS NX is used to establish the chamber wall blasting pressure relief model with a model size of 50 m × 50 m×30 m. The number of nodes of the model is 51791, and the number of elements is 68001. The arrangement direction of blast holes is parallel to the direction of roadway strike, arranged in a single row, and the blast holes adopt centralized and continuous charging. As the blasting is powerful and destructive, so as to prevent damage to the mining roadway, the charging section of the blast hole shall be left with a certain length of resistance line from the coal wall of the working face. At the same time, millimeter millisecond blasting shall be adopted to reduce blasting impact and vibration [33].Figure 14
Blasting hole model of chamber rock mass.When defining the boundary conditions of the model, the establishment of general boundary conditions for dynamic analysis will produce great errors due to the reflection of waves. Therefore, in order to solve this problem, the model uses the viscous boundary proposed by Lysmer and Wass in 1972 to calculate the damping ratio of the corresponding soil inx, y, and z directions to define the viscous boundary, and uses the vibration mode damping calculation. The mechanical parameters of coal and rock required by the model are listed in Table 1.Table 1
Summary of measurement results of physical and mechanical property parameters of +500 m horizontal coal and rock samples.
Coal and rock parametersB3 coal seamB2 coal seamB3 coal seam floor rockB2 coal seam roof rockApparent density (kg/m³)1284125324782692Internal friction angle (°)24.0827.5466.4438.42Cohesion (MPa)2.41.75.46.9Modulus of elasticity (GPa)0.2040.2091.6742.343Poisson’s ratio0.210.190.230.22For the general blasting elastic analysis, the blasting pressure acts on the vertical direction of the hole wall, and the load used at this time uses the formula mentioned in the National Highway Institute of the United States. The blasting load per 1 kg is as follows:(6)Pdet=4.18×10−7×Sge×Ve21+0.8Sge,pB=Pdet×dcdh3,where Pdet is the bursting pressure; pB is the pressure on hole wall; Ve is the blasting speed; dc is the powder diameter; dh is the hole diameter; and Sge is the specific gravity.The above formula determines the aerodynamic pressure during blasting. B = 16338 is the load constant and the dynamic pressure per 1 kg charge.


## 3.2.2. Orthogonal Experimental Scheme Design

Because there are many factors affecting the blasting pressure relief effect, it is impossible to carry out a single control variable test. Therefore, an efficient, fast, and economic orthogonal experimental method is introduced, which selects some representative points from the comprehensive test according to the orthogonality [34]. Nine groups of simulation experiments are designed with the charge quantity, The experimental level table of orthogonal factors is shown in Table 2. The design of the 3-factor 3-level test with the blast hole diameter and blast hole row spacing as the influencing factors, as shown in Figure 15, is shown in Table 3,Table 2
Factor level of the orthogonal test.
LevelFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010Figure 15
Effect diagram of the simulation experiment.Table 3
Design of factor 3 level test.
Experimental schemeFactorABCCharge quantity (kg)Blast hole diameter (mm)Blast hole spacing (m)129062412083615010421201054150666908721508849010961206


## 3.3. Analysis of Simulation Test Results

The advantages and disadvantages of blasting effect are mainly determined according to two aspects: first, the range of the plastic zone formed after blasting; and second, the diameter of the plastic zone after blasting shall be slightly larger than the blast hole spacing to ensure that the range of plastic zone is connected with each other; at the same time, the radius of plastic zone shall be larger than the blast hole spacing to avoid excessive overlap of fracture circle and affect the blasting effect. However, considering the impact analysis of underground blasting safety, it will not damage the stability of mining roadway, and the charge must be within the controllable range. At the same time, a reasonable blast hole row spacing will also make the cracks between blast holes better connected together. The relationship between the variables and the radius of the plastic zone is shown in Figure16.Figure 16
Relationship between various factors and the radius of the plastic zone.The following conclusions can be drawn from the information reflected in Figure16:(1)
Under the condition of certain charge and hole spacing, the range of the plastic zone increases obviously with the increase in the hole diameter.(2)
Under the condition of constant blast hole diameter, with the increase in charge quantity and blast hole spacing, it can also be seen that the radius of the plastic zone tends to increase, but the increasing trend is not obvious due to the influence of each other’s blast hole blasting. It can be seen that charge quantity and blast hole spacing become the control factors in turn.(3)
According to the comprehensive analysis of the above nine groups of simulation experiment effect drawings, it can be seen that No. 3 experiment with a blast hole diameter of 150 mm, a blast hole spacing of 10 m, and a charge of 6 kg has the best blasting effect, and the plastic area produced by blasting is large. Taking the blast hole as the center of the circle, it is measured that the radius of the plastic area produced by a single blast hole is 5.33 m, and the crack area between the two blast holes overlaps, and the mold is not smooth. The proposed effect is good, and the pressure relief effect is obvious.According to the simulation results of the software, a reasonable blasting pressure relief parameter scheme is designed. The blast hole diameter, as the main control factor, can effectively adjust the plastic area range after blasting and control it between 80 and 120 mm as far as possible. In order to avoid the interaction between adjacent blast holes, the blasting test of a single blast hole can be carried out to determine the plastic area range, so as to determine the reasonable blast hole spacing and charge quantity.


## 3.4. Test of Simulation Results Based on Coal Rock Power System

The existence of geological structures such as activated faults and synclines and the rapid change of coal and rock mass thickness will cause the uneven distribution of stress and energy in the stratum, resulting in the emergence of high-stress and high-energy areas in local coal and rock mass, forming a “coal rock dynamic system.” [35] Coal rock dynamic system is an open and dissipative system with complex structure and behavior characteristics, which is characterized by high complexity and nonlinearity. Rockburst is a dynamic process of energy accumulation and release in the coal rock mass system, and is the result of the combined action of geological dynamic conditions and mining disturbance [36-37].In this article, explosive explosion is regarded as the energy source of coal and rock power system, and the accuracy of simulation results is tested through the coal mine coal and rock power system analysis software independently developed by the team.The unit explosive explosion can produce 400 kJ of energy. Input the rock mass mechanical parameters and energy into the software to calculate the corresponding radius of the damaged area. The radius of the damage area caused by 2 kg, 4 kg, and 6 kg explosives is shown in Table4.Table 4
Calculation results of the coal rock power system.
Explosive quantity (kg)Energy generated (J)Radius of failure zone (m)Radius of plastic zone in simulation experiment (m)Error value (％)24000004.264.44448000004.664.591.5612000005.034.961.4In order to reduce the influence of other factors on the radius of the failure area, three groups of experiments of 120 mm and 8 m with blast hole diameter and blast hole spacing as the middle value are selected for verification. The error between the results and the calculation results of the coal rock dynamic system is shown in Table4. It can be seen from the calculation results that the error between the plastic zone radius corresponding to different charge quantities in the simulation experiment and the verification results of the coal rock dynamic system is not more than 5%, which shows that the reliability of the simulation experiment results is high.


## 4. Downhole Transient Electromagnetic Detection of Blasting Pressure Relief Effect

### 4.1. Principle and Method of Underground Transient Electromagnetic Detection

According to the above simulation results of blasting pressure relief, the optimal parameter combination is applied to + 500 m horizontal B3 roadway. The row spacing of blasting holes is 10 m, two blasting holes are constructed in each row, 1# hole angle is 15°, and the construction length is 25 m. The layout of the detection section of B3 coal seam track roadway is shown in Figure17, and the layout of the detection section is shown in Figure 18.Figure 17
Layout of the detection section of track roadway in B3 coal seam.Figure 18
Layout of the detection section.Transient electromagnetic method (TEM) is a time-domain electromagnetic induction method. Its detection principle is to supply a current pulse square wave on the transmission loop. Generally, a primary magnetic field propagating underground is generated at the moment when the square wave current is closed. Under the excitation of the primary field, the geological body will generate eddy current, and its size depends on the conductivity of the geological body. After the primary field disappears, the eddy current cannot disappear immediately, and it will disappear. There is a transition (attenuation) process. The transition generates an attenuated secondary magnetic field to propagate to the surface, and the secondary magnetic field is received by the receiving loop on the ground. The change of the secondary magnetic field reflects the electrical distribution of the underground geological body.Ycs150 mining intrinsically safe transient electromagnetic instrument is an electrical exploration instrument to find underground useful mineral resources and solve geological engineering problems by observing and studying the temporal and spatial distribution law of electromagnetic field. The instrument is composed of host, transmitting wireframe and receiving coil (Figure19–21). It collects the voltage of the detection location underground（μV） - time（μs）. Based on the attenuation change data, the fan-shaped contour distribution map of apparent resistivity (Ω·m) - depth (m) of surrounding rock is obtained by using the late calculation formula of resistivity (calculation of the whole region), and the abnormal body is interpreted in combination with the actual geological and engineering data on site.Figure 19
Host of mine intrinsically safe transient electromagnetic instrument.Figure 20
Launch wireframe.Figure 21
Receiving coil.


### 4.2. Detection Scheme

According to the above simulation results of blasting pressure relief, the optimal parameter combination is applied to +500 m horizontal B3 roadway. The row spacing of blasting holes is 10 m, three blasting holes are constructed in each row, 1# hole angle is 15°, and the construction length is 25 m.The underground transient electromagnetic detection was carried out for the track roadway of +500 m horizontal B3 coal seam +. A detection section is arranged in the track roadway of 500 m horizontal B3 coal seam. The detection section is located at 800 m of the track roadway of B3 coal seam, and the detection direction is the direction of rock pillar. The launching wireframe of the section is 80° to the horizontal plane, forming a fan-shaped section at 10° to the normal of the roadway wall. The detection section has been detected twice before and after the blasting hole of the bottom plate is detonated to analyze the blasting effect and test the rationality of blasting parameters. The underground test takes 60° to the normal of the coal wall as the starting angle and 15° as the angle interval, and detects in a counterclockwise direction, with a total of 5 angles tested, 60°, 75°, 90°, 105°, and 120°.


### 4.3. Pressure Relief Blasting Effect Test

Figures22 and 23 show the contour distribution of rock wall side apparent resistivity before and after floor hole blasting at 800 m of B3 track roadway. There is one high resistance anomaly in Figure 22. Through field comparative analysis, it is inferred that this anomaly is caused by the loosening and fragmentation of surrounding rock caused by floor hole blasting at 810 m of B3 track roadway. The isolines in other areas are evenly distributed in layers, indicating that this area is not affected by loosening blasting. After deep hole blasting, the contour distribution of apparent resistivity of surrounding rock on the side of the rock wall is shown in Figure 23. It can be seen from the figure that there is a high resistivity anomaly in the rock stratum. The apparent resistivity is between 10 Ω m and 70 Ω m. The abnormal areas are connected with each other to form a circular high resistivity zone.Figure 22
Contour of apparent resistivity of rock wall before blasting at 800m of B3 track roadway.Figure 23
Apparent resistivity contour of rock pillar side after blasting at 800m of B3 track roadway.The above detection results show that(1)
The apparent resistivity of rock mass in the same detection area changed significantly before and after underground borehole pressure relief blasting, indicating that the blasting caused the loosening and fracture of rock mass, and the effect is obvious.(2)
After the pressure relief blasting of underground drilling, the spherical failure areas of surrounding rock formed by each pressure relief hole are connected and overlapped with each other, which shows that the drilling spacing of 10 m is reasonable at present, which can produce continuous damage to surrounding rock and realize the pressure relief of rock pillar blasting.


## 4.1. Principle and Method of Underground Transient Electromagnetic Detection

According to the above simulation results of blasting pressure relief, the optimal parameter combination is applied to + 500 m horizontal B3 roadway. The row spacing of blasting holes is 10 m, two blasting holes are constructed in each row, 1# hole angle is 15°, and the construction length is 25 m. The layout of the detection section of B3 coal seam track roadway is shown in Figure17, and the layout of the detection section is shown in Figure 18.Figure 17
Layout of the detection section of track roadway in B3 coal seam.Figure 18
Layout of the detection section.Transient electromagnetic method (TEM) is a time-domain electromagnetic induction method. Its detection principle is to supply a current pulse square wave on the transmission loop. Generally, a primary magnetic field propagating underground is generated at the moment when the square wave current is closed. Under the excitation of the primary field, the geological body will generate eddy current, and its size depends on the conductivity of the geological body. After the primary field disappears, the eddy current cannot disappear immediately, and it will disappear. There is a transition (attenuation) process. The transition generates an attenuated secondary magnetic field to propagate to the surface, and the secondary magnetic field is received by the receiving loop on the ground. The change of the secondary magnetic field reflects the electrical distribution of the underground geological body.Ycs150 mining intrinsically safe transient electromagnetic instrument is an electrical exploration instrument to find underground useful mineral resources and solve geological engineering problems by observing and studying the temporal and spatial distribution law of electromagnetic field. The instrument is composed of host, transmitting wireframe and receiving coil (Figure19–21). It collects the voltage of the detection location underground（μV） - time（μs）. Based on the attenuation change data, the fan-shaped contour distribution map of apparent resistivity (Ω·m) - depth (m) of surrounding rock is obtained by using the late calculation formula of resistivity (calculation of the whole region), and the abnormal body is interpreted in combination with the actual geological and engineering data on site.Figure 19
Host of mine intrinsically safe transient electromagnetic instrument.Figure 20
Launch wireframe.Figure 21
Receiving coil.


## 4.2. Detection Scheme

According to the above simulation results of blasting pressure relief, the optimal parameter combination is applied to +500 m horizontal B3 roadway. The row spacing of blasting holes is 10 m, three blasting holes are constructed in each row, 1# hole angle is 15°, and the construction length is 25 m.The underground transient electromagnetic detection was carried out for the track roadway of +500 m horizontal B3 coal seam +. A detection section is arranged in the track roadway of 500 m horizontal B3 coal seam. The detection section is located at 800 m of the track roadway of B3 coal seam, and the detection direction is the direction of rock pillar. The launching wireframe of the section is 80° to the horizontal plane, forming a fan-shaped section at 10° to the normal of the roadway wall. The detection section has been detected twice before and after the blasting hole of the bottom plate is detonated to analyze the blasting effect and test the rationality of blasting parameters. The underground test takes 60° to the normal of the coal wall as the starting angle and 15° as the angle interval, and detects in a counterclockwise direction, with a total of 5 angles tested, 60°, 75°, 90°, 105°, and 120°.


## 4.3. Pressure Relief Blasting Effect Test

Figures22 and 23 show the contour distribution of rock wall side apparent resistivity before and after floor hole blasting at 800 m of B3 track roadway. There is one high resistance anomaly in Figure 22. Through field comparative analysis, it is inferred that this anomaly is caused by the loosening and fragmentation of surrounding rock caused by floor hole blasting at 810 m of B3 track roadway. The isolines in other areas are evenly distributed in layers, indicating that this area is not affected by loosening blasting. After deep hole blasting, the contour distribution of apparent resistivity of surrounding rock on the side of the rock wall is shown in Figure 23. It can be seen from the figure that there is a high resistivity anomaly in the rock stratum. The apparent resistivity is between 10 Ω m and 70 Ω m. The abnormal areas are connected with each other to form a circular high resistivity zone.Figure 22
Contour of apparent resistivity of rock wall before blasting at 800m of B3 track roadway.Figure 23
Apparent resistivity contour of rock pillar side after blasting at 800m of B3 track roadway.The above detection results show that(1)
The apparent resistivity of rock mass in the same detection area changed significantly before and after underground borehole pressure relief blasting, indicating that the blasting caused the loosening and fracture of rock mass, and the effect is obvious.(2)
After the pressure relief blasting of underground drilling, the spherical failure areas of surrounding rock formed by each pressure relief hole are connected and overlapped with each other, which shows that the drilling spacing of 10 m is reasonable at present, which can produce continuous damage to surrounding rock and realize the pressure relief of rock pillar blasting.


## 5. Conclusion

(1)
The stability analysis and stress state of the intermediate rock column are analyzed by the methods of similar material simulation and numerical analysis. The results show that the coal and rock mass is in a high-stress state under the joint action of horizontal tectonic stress and mining disturbance, and the stress moves downward with the increase in the mining depth, accumulating a large amount of elastic energy in the rock column, which is easy to impact instability, Finally, it leads to the occurrence of rockburst accident.(2)
Under the condition of numerical simulation test, the optimal combination of influencing factors of the deep hole blasting effect of the middle vertical rock column is obtained; that is, the charge is 6 kg, the hole diameter is 150 mm, and the spacing is 10 m. The numerical calculation model under the optimal combination shows that the pressure relief effect is obvious and the range of plastic zone is significantly larger than that of the other 8 groups of experiments.(3)
Through the analysis of the test results carried out by the numerical simulation method, it is found that the blast hole diameter is the main control factor for the blasting effect, and the charge quantity and the distance between holes become the main control factors in turn. The simulation results are tested by coal rock power system software, and it is found that the errors are less than 5%.(4)
The optimal parameter combination obtained from the numerical simulation test is applied to the actual project, and the underground transient electromagnetic is used to test the pressure relief effect of the rock column. The test results show that the apparent resistivity of the rock mass has changed significantly, indicating that the blasting has caused the loosening and fracture of the rock mass, and the effect is obvious.



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*Source: 1947283-2022-05-29.xml*