Mechanism and Application of Roof Cutting by Directional Energy-Cumulative Blasting along Gob-Side Entry

Abstract

In this paper, roof-cutting technology of directional energy-cumulative presplitting blasting is taken as the research object. Through the numerical simulation software Ansys/Ls-dyna3D, the process of energy-cumulative blasting and non-cumulative blasting is simulated and analyzed by using the ALE algorithm. Moreover, the evolution processes of tensile strain energy, detonation stress wave of explosives, stress state of rock mass, and rock crack damage cumulative are compared in two conditions. In the energy cumulative state, the detonation wave acts centrally on the hole wall in the energy cumulative direction to form an initial crack, and then under the action of the jet of energy cumulative, the crack continues to propagate until it runs through. In the non-energy cumulative state, the crack propagates uniformly around the hole wall, forming irregular short cracks. The simulation is verified by the field test, and the law of crack propagation is the same with the simulation. Therefore, directional energy-concentrated presplitting blasting has good practicability in a roof presplitting operation.

1. Introduction

In today’s world energy system, coal is still the main supply of energy in the world, including in China. Roadway is the lifeblood of China’s coal mining industry; roadway roof structure stability plays an important role in the coal mining industry. The roadway support system has been strengthened and optimized to prevent roadway deformation, but relying solely on the support system is not a long-term strategy. The crack structure of hard roofs and high stress concentrations are the two main factors leading to roadway failure [1,2,3]. When the roadway is subjected to the coupling of dynamic and static loads, it may become unstable. The phenomenon of high stress concentration under static load and the movement of the fracture structure under dynamic load will have an impact on the surrounding rock of the roadway [4]. Based on my pressure effect and rock swelling characteristics, He [5] put forward the technology of bi-directional concentrated energy tensile blasting to control the structural stability of the roadway roof. With this technology, the roadway roof is presplit by using concentrated energy presplitting blasting, after which the roof collapses and expands along the presplitting surface and fills the goaf, effectively supporting the high roof to improve the surrounding rock stress environment and reduce the occurrence probability of rock burst accidents. Blasting technology is widely used in coal mines, tunnels, and other industries in China because of its simple process and high economic benefits [6]. Aiming at the roof cutting cumulative presplitting blasting of roadway along goaf, experts and scholars around the globe have conducted a significant amount of research. Based on the cracking mechanism of coal and gangue in multi-directional concentrated blasting, Pan [7] and others analyzed the propagation law of the blasting stress wave. Through similar simulation experiments, it is proved that multi-directional concentrated blasting is scientific and feasible. Ren [8] studied the roof-cutting parameters, roadway support parameters, and stress and deformation law of roadway-surrounding rock under complex thin coal seam by means of numerical simulation, theory, and field experiments. Guo [9] and others carried out the study of roof cutting by concentrated blasting according to the characteristics of shallow and deep coal seam, and determined the key parameters, such as cutting angle, height, and hole spacing. Zong and Meng [10] analyzed the propagation law of the explosive detonation wave and the energy transfer characteristics under different charge structures. Liang [11], according to the propagation mechanism and fracture mechanism of the detonation stress wave in a bi-directional concentrated tension blasting, established a numerical model to study the damage degree of blasthole surrounding rock, the law of detonation stress wave energy propagation, and crack propagation under energy cumulative and non-energy cumulative conditions. Guo [12], based on the dynamic action theory of concentrated blasting, the energy cumulative mechanism, and the forming mechanism of shaped jet, analyzed the characteristics of crack expansion in coal seam blasting, established the numerical model, and analyzed the formation and migration process of shaped jet. Yang [13] et al. used the dynamic caustics test system to analyze the dynamic process and characteristics of crack propagation under the action of a single blasthole, and explored the law of crack propagation in the direction of connection between the empty hole and the blasthole under the action of blasting. Esen et al. [14] proposed a new engineering model to predict the degree of fragmentation around the blasting hole. Based on the crack-forming mechanism of bi-directional energy-cumulative blasting, Shi et al. [15] calculated the crack length of the stress wave, established the numerical model of energy-cumulative blasting, and analyzed the crack propagation length of rock under the action of blasting. Wu and others [16] established the mechanical model of concentrated blasting based on the mechanism of concentrated blasting, and analyzed the mechanism of initial crack formation in the state of concentrated energy.

Therefore, in order to improve the stability of the roadway roof and reduce the probability of accidents in the roadway, research on the roof cutting technology of concentrated blasting is of great significance. Through the methods of theoretical analysis, numerical simulation, and field test, this paper takes the 6302 working face of the Baoshan Coal Mine of Yitai Group in Inner Mongolia as the background. The north side of the Baoshan Coal Mine is the 6303 working face, while the southeast of the Baoshan Coal Mine is the 6301 working face. The length of the working face is 200 m, the thickness of the coal seam is 1.5~1.6 m, the buried depth is 53.5 m~73.7 m, the dip angle of coal seam is 1~3°, and the roof lithology includes mudstone, sandy mudstone, and medium-grained sandstone. The direct bottom and the old bottom are sandy mudstone and medium-grained sandstone, respectively.

2. Fracture Principle of Rock Mass by Concentrated Blasting

2.1. Mechanism of Bi-Directional Energy-Concentrated Tensile Blasting

Bi-directional concentrated tensile blasting mainly makes use of the characteristics of high compressive strength and low tensile strength of rock. This technology can change the dynamic process of interaction between detonation wave and surrounding rock after explosive explosion. Detonation wave acts on the surrounding rock of blasthole, uniformly compressed in the non-energy cumulative direction and concentrated tension in energy cumulative direction [5]. The explosive gas pressure is converted to the tension in the energy cumulative direction to the maximum extent, so as to produce an effective slit surface and achieve the purpose of cutting the top. The mechanism is shown in Figure 1.

After the explosion, under the action of the energy cumulative device, the detonation wave is first unloaded from the set direction and acts directly on the surrounding rock of the blasthole along the set direction, resulting in an initial crack on the surrounding rock of the blasthole. Under the guidance of energy, the high-pressure, high-temperature, and high-speed impact gas [17,18] produced by detonation still flows into the initial crack along the set direction and forms a “gas wedge”, which causes the crack to continue to expand until the explosive in the blasthole is exhausted. In the non-set direction, because the energy cumulative device itself has a certain strength and thickness, it will reduce the propagation of the stress wave in the non-set direction to a certain extent and restrain the detonation wave, and because of the energy cumulative device and the annular space with the blast hole wall, it will also reduce the direct action and damage of detonation gas on the hole wall, resulting in rock mass tensions that are cracked along the set direction, resulting in the formation of slits. The surrounding rock with a non-set direction still maintains good integrity relative to the jet direction.

2.2. Mechanism of Crack Initiation and Propagation

In the non-energy cumulative state, the detonation products will disperse irregularly inside the blasthole, so the rock mass around the blasthole is seriously damaged, and the cracks are irregularly divergent [19]. After the detonation of the explosive, the detonation air flow generated first gathers in the set direction and acts directly on the hole wall rock along the energy concentrator. In the early stage of blasting, the borehole wall is mainly affected by the explosion stress wave, and in a later stage, it is mainly affected by the static pressure of explosive gas. The shaped charge jet on the surrounding rock of the blasthole produced a certain range of failure zone and crack development zone under the action of the explosion stress wave, that is, the initial oriented slit, and the explosive gas also entered the initial crack and promoted the further expansion of the crack.

According to the theory of fracture mechanics, $\partial {\sigma }_{\theta }/\partial \theta =0,{\partial }_{}^2{\sigma }_{\theta }/\partial {\theta }_{}^2<0$, and $K_H$ = 0, then by the formula:

$$\left(\begin{array}{l}{\sigma }_{\gamma }=\frac{1}{2{(2\pi \gamma )}_{}^{1/2}}[K_1(3-{\cos }^{}\theta )\cos \frac{\theta }{2}+K_H(3{\cos }^{}\theta -1){\sin }^{}\frac{\theta }{2}]\\ {\sigma }_{\theta }=\frac{1}{{(2\pi \gamma )}_{}^{1/2}}\cos \frac{\theta }{2}[K_1{\cos }^2\frac{\theta }{2}-\frac{3}{2}{K}_H\sin \theta ]\\ {\sigma }_{\gamma \theta }=\frac{1}{2{(2\pi \gamma )}_{}^{1/2}}\cos \frac{\theta }{2}[K_1\sin \theta +K_H(3{\cos }^{}\theta -1)]\end{array}\right)$$

(1)

The propagation angle of the crack can be obtained:

$${\theta }_0=0$$

Therefore, it can be obtained that under the action of static pressure, the new cracks in the surrounding rock will continue to expand in accordance with the original cracks, resulting in a large range of extended radial cracks in the set direction of the energy concentrator, forming top-cutting cracks.

The stress intensity factor at the tip of crack propagation [20] is:

$$K=PF\sqrt{\pi \left(r+\alpha \right)}+\sigma \sqrt{\pi \alpha }$$

(2)

where P is the pressure of explosive gas; F is the correction coefficient of the stress intensity factor at the crack tip; $r$ is the radius of the hole; $\alpha$ is the instantaneous length of crack propagation; $\sigma$ is tangential stress.

At the end of the concentrated energy flow penetration, the stress intensity factor at the crack tip can be expressed as:

$$K_1=P_{0}F\sqrt{\pi \left(r+{\alpha }_0\right)}+\sigma \sqrt{\pi a_0}$$

(3)

where P₀ the pressure when the explosive gas fills the hole at the end of the penetration; F is the correction coefficient of the stress intensity factor at the crack tip; r is the radius of the hole; ${\alpha }_0$ is the length of the crack; $\sigma$ is tangential stress.

According to the theory of rock mechanics, when the fracture toughness of rock is less than the stress intensity factor at the crack tip, the rock is destroyed, which leads to crack propagation [21]. Therefore, initiation conditions of cracks are obtained as follows:

$$P_0>\frac{K_{IC}-\sigma \sqrt{\pi {\alpha }_0}}{F\sqrt{\pi (r_{}+{\alpha }_0)}}$$

(4)

where $K_{IC}$ is the fracture toughness of rock; $P_0$ the pressure when the explosive gas fills the hole at the end of the penetration; $\sigma$ is tangential stress; $F$ is the correction coefficient of the stress intensity factor at the crack tip; ${\alpha }_0$ is the length of the crack; r is the radius of the hole.

Under the action of the static pressure of the later explosive gas, the crack expands further, and with the expansion of the crack, the pressure of the explosive gas decreases continuously. In order to ensure the continuous expansion of the crack, the instantaneous pressure of the explosive gas needs to meet the following formula:

$$P_{}>\frac{K_{IC}-\sigma \sqrt{\pi {\alpha }_0}}{F\sqrt{\pi (r_{}+{\alpha }_{})}}$$

(5)

where $K_{IC}$ is the fracture toughness of rock; $P$ is the pressure of explosive gas; $\sigma$ is tangential stress; $F$ is the correction coefficient of the stress intensity factor at the crack tip; ${\alpha }_0$ is the length of the crack; $r$ is the radius of the hole; $\alpha$ is the instantaneous length of crack propagation.

In the process of crack expansion in the non-energy-concentrated state, because there is no concentrated energy flow penetration, the stress intensity factor at the crack end is:

$$K=PF\sqrt{\pi \left(r+\alpha \right)}$$

(6)

where P is the instantaneous pressure of explosive gas; F is the correction coefficient of the stress intensity factor; r is the radius of blasthole; $\alpha$ is the instantaneous length of crack propagation.

Similarly, the initiation conditions of cracks in the non-energy-accumulating direction are as follows:

$$P_0>\frac{K_{IC}}{F\sqrt{\pi (r_{}+{\alpha }_0)}}$$

(7)

where $K_{IC}$ is the fracture toughness of rock; $P_0$ the pressure when the explosive gas fills the hole at the end of the penetration; $F$ is the correction coefficient of the stress intensity factor; $r$ is the radius of blasthole; ${\alpha }_0$ is the length of the crack.

The conditions for continuous expansion of cracks in the non-energy-gathering direction are as follows:

$$P>\frac{K_{IC}}{F\sqrt{\pi (r_{}+\alpha )}}$$

(8)

where $K_{IC}$ is the fracture toughness of rock; $P$ is the instantaneous pressure of explosive gas; $F$ is the correction coefficient of the stress intensity factor; $r$ is the radius of blasthole; $\alpha$ is the instantaneous length of crack propagation.

From the above formula, it can be seen that the penetration of concentrated energy flow in the energy cumulative direction reduces the pressure needed for crack initiation and expansion, and promotes the crack propagation distance to be larger. The results show that compared with non-concentrated blasting, concentrated blasting has the advantage of cracking and can improve the roof cutting effect.

3. Arbitrary Lagrangian-Eulerian (ALE) Algorithm

The arbitrary Lagrangian-Eulerian algorithm was first proposed by Noh (1964) as the coupled Euler-Lagrange algorithm. In this algorithm, the grid can move in any form in space, so that the moving interface of the object can be accurately described and the reasonable shape of the unit can be maintained by specifying a reasonable form of grid motion. The pure Lagrangian algorithm and the Eulerian algorithm are in fact two special cases of the ALE algorithm, that is, it is reduced to the Lagrangian algorithm when the velocity of the grid point is equal to the velocity of the material point, and to the Eulerian algorithm when the grid is fixed in space. The characteristic of the ALE algorithm is that the grid it uses is neither the fixed grid of the Eulerian algorithm nor the body grid of the Lagrangian algorithm, but constructs a set of appropriate grids according to the flow boundary of the material region.

The body derivative described by ALE can be written as follows:

$${\frac{\partial F}{\partial t}|}_X={\frac{\partial F(\xi ,t)}{\partial t}|}_{\xi }+{\mathrm{c}}_{\mathrm{i}}\frac{\partial F}{\partial {\chi }_{\mathrm{i}}}$$

(9)

where ${\mathrm{c}}_{\mathrm{i}}$ = u_i − ω_i is the convective velocity described by ALE; u_i is the physical velocity of particle X; ω_i is the grid speed.

Through Formula (9), the governing equation in the current configuration of ALE algorithm can be obtained [22]:

The mass conservation equation:

$${\frac{\partial \rho }{\partial \mathrm{t}}|}_{\epsilon }+{\mathrm{c}}_{\mathrm{i}}\frac{\partial \rho }{\partial {\chi }_{\mathrm{i}}}+\rho \frac{\partial {\upsilon }_{\mathrm{i}}}{\partial {\chi }_{\mathrm{i}}}=0$$

(10)

The momentum conservation equation:

$${\rho \frac{\partial {\upsilon }_{\mathrm{i}}}{\partial \mathrm{t}}|}_{\epsilon }+\rho {\mathrm{c}}_{\mathrm{i}}\frac{\partial {\upsilon }_{\mathrm{i}}}{\partial {\chi }_{\mathrm{j}}}=\frac{\partial {\sigma }_{\mathrm{ij}}}{\partial {\chi }_{\mathrm{j}}}+\rho {\mathrm{f}}_{\mathrm{i}}$$

(11)

The energy conservation equation:

$${\rho \frac{\partial \mathrm{e}}{\partial \mathrm{t}}|}_{\epsilon }+\rho {\mathrm{c}}_{\mathrm{i}}\frac{\partial \mathrm{e}}{\partial {\chi }_{\mathrm{i}}}={\sigma }_{\mathrm{ij}}\frac{\partial {\upsilon }_{\mathrm{i}}}{\partial {\chi }_{\mathrm{j}}}-\frac{\partial {\mathrm{q}}_{\mathrm{i}}}{\partial {\chi }_{\mathrm{i}}}$$

(12)

where $\rho$ represents density; f_i represents physical strength per unit mass; $\chi$ represents the coordinates of matter in the grid; $\upsilon$ represents the speed of motion; t represents time; c represents the convection velocity; ε represents the position vector in the reference coordinates; ${\sigma }_{\mathrm{ij}}$ represents the Cauchy stress tensor; e represents internal energy per unit mass; q_i represents the heat flux.

4. Numerical Simulation Research

4.1. Numerical Modeling and Meshing

In the field construction, the roof cutting by concentrated blasting usually adopts the structure of continuous charge. In order to facilitate the calculation, the blasthole and rock mass profiles are obtained for analysis. The nonlinear dynamic software is used to simulate the two-way concentrated blasting, the Ansys software is used for modeling, and the SOLID164 element in the software is used to establish the model. The model is divided into four parts: rock, air, explosive, and energy concentrator. In the model, the area of air and rock mass is 100 cm × 100 cm square, the inner diameter of the concentrating tube is 3.6 cm, the outer diameter is 4.2 cm, the slit width is 0.4 cm, and the diameter of the explosive is 3.2 cm. The calculation model is simplified to the plane stress state, and 1 cm is taken as the thickness direction. After the establishment of the model, the meshing is divided by mapping and sweeping, while the local mesh refinement is beneficial to the transmission pressure, so the air and rock grids near the explosive area are refined, and the energy concentrator and explosive are divided into uniform grids, as shown in Figure 2. Because the rock and air elements will produce a large deformation in the blasting process, and the mesh distortion caused by large deformation may affect the calculation results, the ALE algorithm is used here. The Lagrange element is used for the rock and energy tube, the ALE element is used for explosive and air, and the coupling of structure and air is realized by the keyword * CONSTRAINED_LAGRANGE_IN_SOLID. In the calculation process, the explosive is initiated by the central point, while the boundary conditions of air and rock are non-reflective. In order to obtain the energy cumulative effect of energy cumulative blasting, a model with the same size and the same parameters except the energy cumulative tube is established, while the energy cumulative and non-energy cumulative stress evolution laws are analyzed and compared, and the energy cumulative effect of energy cumulative blasting is obtained.

4.2. Determination of Model Calculation Parameters

4.2.1. Explosive Parameters

The explosive adopts two-stage emulsion explosive, the explosive material model adopts the * MAT_HIGH_EXPLOSIVE_BURN model, and the detonation pressure is expressed by the JWL equation of state [23]:

$$\mathrm{p}=A(1-\frac{\omega }{A{R}_1}){\mathrm{e}}^{{-\mathrm{R}}_1\mathrm{V}}+B(1-\frac{\omega }{A{R}_2}){\mathrm{e}}^{{-\mathrm{R}}_2\mathrm{V}}+\frac{\omega E}{V}$$

(13)

where p is the pressure of detonation products; $A$, $B$, $R_1$, and $R_2$ are material constants; $V$ is the relative volume of detonation products; E is the initial internal energy density of the detonation products.

The explosive parameters are shown in

Table 1. Mechanical parameters of explosives.

RO/ (g·m−3)	D/ (m·μs−1)	PCJ/ Mbr	JWL Equation of State
			A/ Mbr	B/ Mbr	R1	R2	ω
1.18	0.5122	0.0953	2.762	0.0844	5.2	2.1	0.5

.

4.2.2. Material Parameters of Energy Concentrator

PVC material is used in the energy concentrating tube in the project site, this kind of material is a heat-related material, which has a certain strength in the initial stage of blasting reaction, but under the impact of detonation products and high temperature, it will be destroyed [24,25], so it is simulated by the * MAT_PLASTIC_KINEMATIC plastic constitutive model. The * MAT_ADD_EROSION keyword is used to set a small restart, and a part of the energy concentrator is deleted after the failure of the energy concentrator to prevent the failure of the energy concentrator from affecting the simulation results.

The material parameters of the concentrator are shown in

Table 2. Mechanical parameters of concentrating tube materials.

RO/(g·cm−3)	PR	E/ Mbr	SIGY/ Mbr	ETAN/ Mbr
8.93	0.35	1.17	0.004	0.001

.

4.2.3. Air Material Parameters

Because the energy concentrating tube does not fully fit the hole wall, the air domain should be set. Air is regarded as an ideal gas. The equation of the air material selection* MAT_NULL model is described by * EOS_LINEAR_POLYNOMIAL:

$$P=C_0+C_1\mu +C_2{\mu }^2+C_3{\mu }^3+(C_4+C_5\mu +C_6{\mu }^2)E$$

(14)

where C₀, C₁, C₂, C₃, C₄, C₅, C₆ are constant; μ is the specific volume; $E$ is the specific internal energy.

The air material parameters are shown in

Table 3. Mechanical parameters of air materials.

RO/(g·cm−3)	* EOS_LINEAR_POLYNOMIAL
	C0	C1	C2	C3	C4	C5	C6	E0
1.2	0	0	0	0	0.4	0.4	0	2.5 × 10−6

.

4.2.4. Rock Material Parameters

In the process of explosion, the rock mass is damaged in two forms: one is tensile failure and the other is compression failure. Furthermore, the rock around the blasthole is in a state of large strain, high strain rate, and high stress during blasting; according to this characteristic, the rock adopts * MAT_JOHNSON_HOLMQUIST_CONCRETE, that is, the HJC model. Since rock cracking and crack propagation are realized by rock failure, the keyword * MAT_ADD_EROSION is used to add tensile stress and shear strain damage failure criteria to simulate the dynamic failure effect of rock [26].

The rock material parameters are shown in

Table 4. Mechanical parameters of rock materials.

RO/ (g·cm−3)	G/ Mbr	A	B	C	N	FC/ Mbr	T/ Mbr	EPS0/ µs−1	Efmin	Sfmax
2.6	0.287	0.28	2.5	0.00186	0.79	0.00154	1.708 × 10−4	1.0 × 10−11	0.01	5.0
Pc/ Mbr	μc	Pl/ Mbr	μl	D1	D2	K1/ Mbr	K2/ Mbr	K3/ Mbr	Fs
1.0 × 10−4	0.00162	0.012	0.012	0.04	1.0	0.12	0.25	0.42	0.035

.

4.3. Comparison and Analysis

The crack propagation process is shown in Figure 3 and Figure 4.

In order to obtain the change law of mechanical properties of concentrated blasting and non-concentrated blasting during blasting, measuring points A, B, C, D, and E were arranged at 5 cm, 15 cm, 25 cm, 35 cm, and 45 cm, respectively, to monitor the changes of mechanical properties of each point during blasting. The location of the measuring point is shown in Figure 5.

4.3.1. Energy Evolution Process

It can be seen from Figure 6 that under the action of the energy accumulator, the energy produced by the explosive explosion in the state of 11.9 μs produces a jet in the energy cumulative direction, and the energy in the energy cumulative direction is enhanced and weakened in other directions. The energy wave propagates outward in a fan shape along the energy cumulative direction. At 33.9 μs, the energy first acts on the hole wall from the set direction to form a guiding initial crack, then the energy propagates outward along the direction of the initial guiding crack, and a high stress zone is formed at the crack tip. Because the stress in the high stress area is greater than the fracture toughness of the rock, the crack deepens continuously in the range of 33.9 μs to 399.9 μs. As can be seen from Figure 7, under the non-energy cumulative condition, the explosion stress wave propagates gradually and uniformly along the blasting hole, and the blasting energy acts uniformly on the wall of the blast hole. At 11.9 μs, the detonation wave acts on the surrounding rock of the blast hole, and the failure area is approximately circular. The crack expands continuously between 33.9 μs and 121.9 μs. At 189.9 μs, because the stress in the high stress area at the crack tip is lower than the fracture toughness of the rock, the crack stops expanding and the surrounding stress tends to be stable. Under the action of the energy cumulative device, the length of the rock crack in the energy cumulative state is much longer than that in the non-energy cumulative state. It can be seen from Figure 8 that the energy of the detonation wave in two states decreases with the increase of propagation distance, and the energy of each measuring point in the energy cumulative direction is much larger than that of each measuring point in the non-energy cumulative state, i.e., 2.3 times, increasing the action time. The utilization rate of explosive energy is improved.

4.3.2. Detonation Stress and Velocity Propagation Process

From the combination curve of the detonation wave stress and velocity cloud diagram (Figure 9, Figure 10, Figure 11, Figure 12, Figure 13 and Figure 14), it can be seen that in the energy cumulative direction, the explosive propagates preferentially to the energy cumulative direction, forms a high stress zone and a high velocity zone in the energy cumulative direction, and becomes stronger with time. After the detonation wave propagates to the wall of the blasthole along the energy cumulative direction, the stress strength of the detonation wave is greater than the dynamic tensile strength of the rock mass, and a guide crack is formed under the action of the jet. Since then, under the influence of the propagation characteristics of the detonation wave, the detonation wave continues to propagate along the energy cumulative direction. At 61.9 μs, the energy cumulative tube is greatly damaged by high stress deformation, and the detonation wave also has an effect on the hole wall in the non-energy cumulative direction. At this time, the hole wall is destroyed. Part of the detonation wave will continue to reflect in the process of propagation, and the new stress wave will continue to act on the tip of the notch, making the fracture continue to develop. In the process from 61.9 μs to 399.9 μs, with the penetration of a detonation wave, the crack deepens until it penetrates. In the non-energy cumulative direction, under the influence of the energy cumulative device, the detonation energy transferred to the borehole wall is weakened and the failure area is reduced. In the non-energy-gathering state, the detonation wave propagates outward in a ring, acting uniformly on the hole wall, and the hole wall is uniformly forced to form a near-circular failure area. The detonation wave stress and velocity around the blasthole propagate uniformly and gradually decrease with the propagation distance until the end of the blasting process. There are differences in the arrival time and magnitude of the peak at each measuring point under energy cumulative and non-energy cumulative conditions. The velocity transfer of detonation wave in the non-concentrated state lags about 20 μs compared with that in the concentrated state, and the peak value of each measuring point in the concentrated state is about 2.2 times of that in the non-concentrated state, and the propagation speed is fast. The peak value of stress and velocity of detonation wave at measuring point An is the largest, and the peak value decreases gradually with the distance from the hole wall, but under the action of concentrated jet, the attenuation velocity of peak stress and velocity of detonation wave under the condition of concentrated energy is smaller than that under the action of non-concentrated energy. The peak stress of detonation wave at each measuring point in the energy cumulative state is larger than that in the non-energy cumulative state, so the peak stress of rock mass in the energy cumulative direction is larger under the action of energy cumulative device during blasting, which promotes the development and expansion of cracks. In addition, due to the action of a jet, the attenuation velocity of detonation wave stress in rock mass in the energy cumulative direction is the smallest, which expands the range of detonation wave stress and effectively improves the utilization rate of explosive energy.

4.3.3. Rock Tension-Compression Stress and Rock Damage Evolution Process

It can be seen from rock tension and compression stress, rock damage cloud map, and history curve(Figure 15, Figure 16, Figure 17, Figure 18, Figure 19 and Figure 20), in the energy cumulative state, at the beginning of blasting, that the detonation gas first ejects from the set direction, the stress concentration occurs in the set direction at 11.9 μs, and the stress wave propagates outward in a fan shape along the energy cumulative direction. Then, the detonation wave acts on the wall of the blasthole to form an initial crack, that is, the guiding crack, guides the fracture to continue to spread outward in this direction, and then, with the continued propagation of the stress wave, the damage area of the rock continues to expand. However, with the increase of time, the stress wave will continue to decrease, which will reduce the degree of rock damage, and at this time, the stress is mainly concentrated in the crack tip. Under the action of the “gas wedge”, the new crack at the tip is constantly under tension, and because the stress intensity factor at the tip is greater than the rock fracture toughness, the crack continues to expand. After 61.9 μs, the damage area is basically unchanged, and the stress of the surrounding rock of the blasthole decreases. Under quasi-static action of explosive gas, the crack expands under the tensile stress, forming a through crack. In ordinary blasting, due to the use of uncoupled charge. The detonation wave cannot act on the blasthole wall the first time. At 11.9 μs, the detonation wave is transmitted to the blasthole wall, and the tension and compression stress of the rock is mainly distributed around the blasthole. A high stress zone is formed on the rock around the blasthole, and the detonation wave stress strength greatly exceeds the tensile strength of the rock. At 33.9 μs, the rock around the blasthole is completely destroyed, the stress wave expands outward, and the damage range of the rock increases continuously. Moreover, under the action of the detonation wave, the rock around the blasthole continues to destroy, and because of the randomness of the detonation wave, irregular cracks are formed in the rock outside the blasthole. Between 33.9 μs and 121.9 μs, under the action of explosive stress wave, the failure range of rock increases continuously, and the maximum effective stress also increases. At 121.9 μs, the crack stops developing, the tension and compression stress of rock begins to attenuate, the effect on rock decreases until the explosive is exhausted, and the stress distribution of rock outside the blasthole becomes gradually uniform. Under the non-energy cumulative condition, the stress wave always propagates uniformly outward in a ring, and the effective stress at the same distance from the center of the hole is basically the same. Compared with the non-concentrated blasting, the cracks of the concentrated blasting are mainly distributed in the set direction of the concentrated tube, but not the cracks caused by the concentrated blasting spreading irregularly. Due to the existence of the energy cumulative tube, the stress factors are concentrated in the energy cumulative direction, which not only protects the integrity of the surrounding rock around the non-energy cumulative direction, but also can effectively precrack the roof and form an effective crack. Under the energy cumulative condition, the tension and compression stress of rock at each measuring point is about 2.3 times higher than that in the non-energy cumulative state. In both states, the tension-compression stress and rock damage decrease with the increase of the distance between the measuring point and the blasthole, but the interaction time between detonation wave and rock is longer in the energy cumulative state.

4.3.4. Blasting Crack Displacement

It can be seen from Figure 4 that the cracks produced by non-concentrated blasting spread out uniformly with the blasthole at the center, resulting in irregular divergent cracks and the development of cracks are more uniform. It can be seen from Figure 3 and Figure 21 that under the energy cumulative condition, the stress concentration occurs in the set direction of the energy cumulative tube, and the detonation stress wave preferentially acts on the hole wall along the energy cumulative direction, resulting in an obvious guiding crack at 33.9 μs. Under the guidance of the guide crack, the subsequent cracks continue to extend in this direction until they run through, and there are fewer cracks in the non-energy cumulative direction. It can be seen from Figure 22 that the curve changes of each measuring point have an obvious dynamic load stage, static load stage, and stable stage. Under the energy cumulative condition, there are two kinds of action: the compressive strain and tensile strain, and the tensile strain is mainly affected by the compressive strain, while in the non-energy concentrated state, there is only the compressive strain. Under the energy cumulative condition, with the increase of distance, the peak value of the compressive stress at the five measuring points of Arecoire, C and D is gradually weakened, while the tensile stress is gradually increased, which is due to the concentrated tension in the set direction due to the action of the energy gathering device. The detonation pressure can be converted to the tensile effect on the surrounding rock in the set direction to the maximum extent, resulting in an effective slit surface. Compressive stress in the non-energy-gathering state decreases with the increase in the distance until it is 0. By comparison, it can be found that the detonation wave is enhanced in the energy cumulative direction by the energy cumulative device. With the formation of the guide crack, a large amount of detonation gas is poured in, which strengthens the load and lasts longer, resulting in effective slit cracks.

5. Field Test

5.1. Test Scheme

Experiments are carried out in the working face on the basis of numerical simulation, and the presplitting slit blasting materials is shown in Figure 23, including emulsion explosives, gun mud, detonators, wires, energy accumulators, and connectors. In the field test, the diameter of the borehole is 50 mm, the optimum charge quantity in the stable area of the roof is 3:2, there are two concentrating tubes in each hole, the specification of the charge cartridge is φ32 × 200 mm, and the quality is 300 g/roll. The optimum charge quantity in the compound broken roof area is 2:1, there are two energy collecting tubes in each hole, the specification of the charge cartridge is φ32 × 200 mm, and the mass is 300 g/roll. The method of segmenting charge is adopted, and when the crack rate of this test reaches 80%, the charge structure is adopted in the follow-up test. When the crack rate is less than 80%, the charge needs to be increased. The slit hole is arranged at a certain angle between the side of the roadway and the roof, the angle between the crack hole and the plumb line is 15°, the hole depth is 5 m, and the sealing mud length is 2 m.

5.2. Test Process

First of all, DCA- 45 roofs slotting hole drilling rig is used for drilling. The detonators and lead are installed in the roadway according to the blasting charge design parameters. Starting from the bottom of the hole, after the charge is finished, the lead is passed through the second energy concentrator, and the first energy concentrator is connected with the second one with connectors. Then, the charge is started and the leads are installed in the second tube, and so on, after which the charge of the concentrator is completed in turn. After all placements, the directional rod is used to adjust the slit direction of the energy concentrator, so that the slit direction of the energy concentrator is consistent with the cutting top line. Then, the hole is sealed with blasting mud, and finally, the blasting test is realized. The blasting process is shown in Figure 24.

5.3. Analysis of Test Results

At the end of the test, a CXK6 snooping device was used to measure the length of the crack in the slit hole. The blasting results are shown in Figure 25. From the blasting results, it can be seen that under the action of energy cumulative blasting, there are two cracks on the inner hole wall and the outer rock wall of the blast hole, and the rock wall in the non-set direction still maintains good integrity. Therefore, the crack in the slit hole is a single crack surface, not multiple crack surfaces, and this kind of one-sided crack occurs in the shallow, middle, and deep parts of the hole. After the statistics, the average crack rate of the shaped blasting slit hole is 82%, which meets the requirement of the cracking rate. After presplitting, the roof collapse is shown in Figure 25c. Through the comparison of the results, it can be seen that the results of the field test and the numerical simulation are consistent, and the penetration cracks are produced in the energy cumulative direction. Therefore, directional energy-cumulative presplitting blasting can realize directional roof cutting.

6. Conclusions

(1)

In the state of energy accumulation, the detonation wave produced by blasting gathered in the set direction, resulting in the concentration of tensile strain energy. Under the action of the energy accumulation device, the surrounding rock of the blasthole in the set direction was destroyed and an initial crack was formed. Under the action of the energy-accumulating jet, the crack continued to propagate until penetration. In the non-energy-gathering state, the detonation wave and energy produced by the explosion acted uniformly on the whole wall around the blast hole, resulting in divergent short cracks.

(2)

Under the action of detonation wave and detonation energy, two penetrating cracks in the set direction were formed in the surrounding rock under the condition of concentrated blasting, and the cracks developed irregularly around the blasthole under the non-concentrated blasting condition.

(3)

Through the field test of the 6302 working face in the Baoshan Coal Mine, two single cracks in the set direction were produced on the inner wall of the blast hole and the rock wall outside the hole, which met the cracking requirements. The field test results were consistent with the simulation results. Therefore, directional energy-concentrated presplitting blasting can cut off the connection between roadway roof and stope, effectively control roadway roof structure, and realize the effect of directional roof cutting.