Research on Disaster Prevention and Control Technology for Directional Hydraulic Fracturing and Roof Plate Unloading

Abstract

In coal seam groups where the spacing between the upper and lower seams is small, the lower seam working face is significantly influenced by residual coal pillars from the upper seam and the void spaces created during mining. This presents considerable challenges for underground mining safety. Through field investigations, the layout of the coal seam quarry above the working face of the 3⁻¹ coal seam in Yanghuopan Mine was examined, along with the distribution of the residual coal pillars. This allowed for the identification of the interlayer rock strata characteristics. Subsequently, we analyzed the mechanism of directional hydraulic fracturing and decompression to determine the key parameters of the 3⁻¹ coal seam. Using the Rock Fracture Process Analysis 3D (RFPA 3D) numerical simulation, we evaluated the effects of various factors on the initiation and propagation of hydraulic fracturing-induced cracks, formulated the evolution law of these fractures, and incorporated the damage variables into the analysis. Additionally, we assessed the influence of different parameters on crack initiation and extension during hydraulic fracturing, using RFPA 3D simulations to derive the evolution law governing directional hydraulic fractures. This allowed us to define the hydraulic fracturing parameters for the 3⁻¹ interbedded rock layers by integrating the process parameter calculations with the damage variables. Based on these findings, an on-site implementation plan was developed and executed, followed by a comprehensive evaluation of the construction results. The study concludes that directional hydraulic fracturing and decompression effectively contribute to the prevention and control of roof-related disasters in the mining of lower coal seams where seam spacing is minimal. This research offers valuable theoretical insights and practical reference for disaster prevention and control in similar geological conditions.

1. Introduction

The incidence of mine roof accidents in China remains significantly high compared to other types of mining accidents, presenting a substantial risk to the safety of coal mining operations [1]. In recent years, the depletion of shallow coal reserves in certain mines in northern Shaanxi has necessitated the extraction of deeper coal seams within the same coal seam clusters [2]. Due to the small spacing between the upper and lower coal seams, the impact of mining between these coal seams causes superimposition, resulting in peripheral rock movements in adjacent coal seams; this makes the stress relief zone and the concentration of the area of the superimposed impact on each other. The influence of coal pillars and the airspace areas left behind by mining in the upper seam means that the lower seam mining overburden breakage law will be different from the geological conditions of a single seam mine. The lower seam working face is prone to all kinds of pressure frame accidents during the mining process. One of the major issues that must be resolved quickly in order to produce mines in this area is the problem of mining close to the coal seam group [3,4,5,6].

Qingxiang Huang et al. [7,8] carried out a study on the overburden collapse law and damage reduction in shallow buried coal seam group mining and concluded that the pressure of the working face is determined by the key layer of the spacing rock stratum and the disturbed roof of the upper coal seam. They also found that it has the phenomenon of large and small cycles of pressure. Chuangye Wang et al. [9] combined on-site measurements, simulation tests, and other means of studying the mine pressure characteristics of the working face of 31,201 in an overburdened mine and the leftover coal pillars of the Ohtai Mine. They also studied the abnormal pressure that occurs during the advancing process of the working face under the mine void zone, which may trigger the abnormal coming pressure when passing through the upper coal pillar. It was concluded that the pressure was abnormal during the advancement of the working face under the mining hollow area, and that when passing through the overlying coal pillar, the pressure was more intense and accompanied by the phenomena of coal wall flake ganging and increased subsidence of the support column, which might trigger a disaster in the pressure frame. The work of Jie Zhang [10,11,12] and other similar simulation experiments on the repeated mining of the shallow buried close coal seam group in the Hanjia Mine showed that under repeated mining conditions, the overburden rock appeared to have off-seam fissures and longitudinal fissures, which were accompanied by the generation of super fissures. Overburden fissures undergo six dynamic cycles of generation, expansion, closure, regeneration, penetration, and re-closure in the formation of “mine-area-face” and “mine-area-area-face” structures. Keshavarz et al. [13] used hydraulic fracturing to improve inter-fracture connectivity and proposed enhancing the electrical conductivity of the fracture system to maintain its effectiveness during the production period after the fracturing. Scholars at home and abroad have carried out a lot of research on the theory of overburden rock activity and the law of mineral pressure manifestation in coal seam groups. This further explains various special mineral pressure phenomena occurring in the mining process of close coal seam groups and lays a theoretical foundation for research developing control technologies for the surrounding rock in close coal seam mining. The method of directional hydraulic fracturing to unload pressure is proposed to prevent and control roof disasters.

Based on a theory from previous research, this paper examines the Yanghuopan Mine as a case study and proposes the method of directional hydraulic fracturing and decompression to prevent and deal with roof disaster accidents. The hydraulic fracturing decompression treatment for roof disaster prevention in the Yanghuopan Coal Mine mainly focuses on the 30,116 working faces, 30,119 working faces, and the 301-disk area of the 3⁻¹ coal seam of the mine. Through on-site research, we clarified the characteristics of the interbedded rock layers above the 3⁻¹ coal seam and analyzed the main causes of the pressurized roof accident in the lower coal seam based on the structural morphology of the overburden rock from the proximal coal seam group. Next, we determined the key parameters of the 3⁻¹ coal seam through an analysis of the mechanism of directional hydraulic fracturing and unloading from the pressurized roof. We then conducted an analysis of the factors influencing fracture initiation and propagation using RFPA 3D (Rock Fracture Process Analysis) numerical simulations, which allowed us to derive the evolution law of directional hydraulic fracturing. By integrating the hydraulic fracturing process parameter calculation method with the damage variable, we determined the optimal hydraulic fracturing parameters for the 3⁻¹ coal seam. Based on these findings, we developed and implemented a field plan for the 30,116 and 30,119 working faces at Yanghuopan Mine, followed by on-site testing. A subsequent evaluation of the construction outcomes was conducted. This study provides a robust theoretical foundation for the prevention and control of roof-related disasters during mining operations in similar geological settings and offers a practical reference for mining the lower seams of coal seam groups under comparable conditions.

2. Overview of the Mine and Workings

Yanghuopan Coal Mine is located in Dianta Town, Shenmu City, Shaanxi Province, and belongs to the shallow southwestern part of the Xinmin Mining Area of the Shenfu Mining Area in the Jurassic Coal Field of North Shaanxi Province. Most of the coal mine surface is covered by Quaternary System and Neoproterozoic System sediments, and the coal seams are shallowly buried. The mineable coal seams are divided into upper and lower coal groups, with two disk areas for each coal seam. The 2⁻², 3⁻¹, and 3⁻² coal seams are upper coal groups, while the 4⁻³ coal seam and 5⁻¹ coal seam are lower coal groups. The 30,116 face is located in the central-western part of the well field, west of the mine’s 3⁻¹ coal 301 disk area, and the interbedded rock layers are fine-grained sandstone and siltstone, with the RQD (Rock Quality Designation) of the core of the rock being above 70% and with a good degree of completeness. The 30,119 working face is located in the east-central part of the well field, next to the 3⁻¹ coal 301 disk area of the mine; the interlayer rock layer is siltstone and fine-grained sandstone, and the core integrity is good. The working face layout is shown in Figure 1. The surface topography of the work face is complex, with rolling beams and mounts, gullies, and ravines, and the highest elevation of the beams and mounts is +1256 m. Overlying strata on the work face range in thickness from 0 to 161.5 m, overlying strata on the cutting eye range in thickness from 10 to 66.2 m, overlying strata on the retreat tunnel range in thickness from 117 to 161 m, the bedrock ranges in thickness from 42.87 to 161.58 m, and loose layers range in thickness from 0 to 15.19 m.

The 3⁻¹ seam studied here is located in the upper part of a seam (which is split into two consecutive seams) and acts as the main seam. This coal seam has the characteristic of medium thickness, with a recoverable area of 22.66 km², burial depth of 0–173.26 m, and thickness of 0–4.81 m (average thickness of 2.58 m). The 2⁻² coal seam and the 3⁻¹ coal seam of the Yanghuopan Mine are adjacent seams, and the thickness of the rock strata between them is about 23 m. The rock strata between the seams are hard, and part of the area above the working face of the 3⁻¹ seam, 30,116, is part of the working face of the 2⁻² coal seam. Above the working face, there are coal pillars left after the mining of several workings in the 2⁻² coal pan area. The 30,116 working face in the 3⁻¹ coal seam is under the loading of coal pillars in the 2⁻² coal seam and the overlying rock layer, and there is a stress concentration phenomenon on the roof plate of the lower coal seam in the area of the stopping line, which is very prone to the occurrence of dynamic pressure in the mining field. In the mining of the upper seam, there is a permanent coal pillar in the working face, and as the air space on both sides of the coal pillar continuously falls and compacts, part of the weight of the rock body above and the over-supporting pressure formed during mining will be concentrated on the coal pillar, resulting in concentrated stress. This concentrated stress is transmitted downward, which will lead to an increase in rock fissures on the top plate of the lower coal seam and intensify the broken top plate of the coal seam. When the lower working face advances to the area affected by the concentrated stress of the coal pillar, severe flaking of the coal wall and the top of the end face will occur. If the masonry beam structure is in an unstable structural condition after breaking the overburden key layer after mining the upper coal seam, the masonry beam structure after breaking the lower key layer structure during mining of the lower coal seam will be prone to slipping and destabilization and the mine pressure at the working face will be abnormally strong. Furthermore, the working face of the lower coal seam will be prone to livid loaded mine pressure, and even to pressure frame accidents similar to those under the conditions of the overburden single key layer and composite single key layer structure during mining of the upper coal seam [14,15]. Therefore, the main threat to the safety of the comprehensive mining working face is the stress concentration problem when the close coal seam is mined under the coal pillar.

3. Mechanism Analysis of Directional Hydraulic Fracturing Unloading and Determination of Key Parameters

3.1. Mechanism Analysis of Unloading Mechanism in Directional Hydraulic Fracturing

The primary technology for managing high stress involves transforming the coal and rock mass into weak structures, thus facilitating stress redistribution [16]. On this basis, the stress transfer theory of the coal rock cracking in weak structures is put forward: artificial fracturing of the coal and rock mass reduces its mechanical strength, forming a weak structure that lowers its load-bearing capacity and redistributes the stress and energy fields. RFPA 3D (Rock Failure Process Analysis 3D) was employed to simulate the structure and stress variation rule of the weak structure after loading [17], thereby demonstrating the stress variation effect of the weak structure, as shown in Figure 2. Under further loading, the weak structure will fracture before the surrounding coal and rock, weakening the area's mechanical properties and creating a low-stress zone that shifts high stress to adjacent areas.

After the weak structural body is loaded, the tips of the artificial cracks will gradually crack and expand, producing a certain degree of microscopic damage. When the cracks penetrate into each other and reach the critical instability conditions of the coal rock body, the coal rock body will undergo shear slippage [16]; when the load is further increased, the macroscopic cracks will also be further increased and eventually form a network of cracks (as shown in Figure 3). Stress homogenization of weak structural bodies is of great significance for the prevention and control of coal rock dynamic disasters.

Based on the effect of stress change in weak structures, the concept of stress transfer in coal cracking of weak structures is proposed: generally, the approach employed is to drill holes into the roof, coal seam, or floor on both sides of the mining roadway or at the head of the excavation roadway and subsequently implement measures such as hydraulic fracturing in the area outside the protection of the coal and rock mass, thereby creating the requisite artificial cracks in the designated area and forming a reasonable weak structure (as shown in Figure 4).

3.2. Directional Hydraulic Fracturing Process Design Process and Steps

The design flowchart for the parameters of the directed hydraulic fracturing process is shown in Figure 5. The calculation steps are as follows:

• Determine the range of fracturing L_h according to the desired damage variable D;
• Given an initial fracturing displacement Q and fracturing time T, calculate the radius of fracture propagation R;
• Based on the radius of fracture propagation R, determine the spacing of the drilled holes d;
• Based on the spacing of the drilled holes d and the range of fracturing L_h, the number of drilled holes m can be determined, and the arrangement of the boreholes can be determined by m;
• Knowing the damage variable D and the number of drilled holes m, the number of fracturing segments n for each borehole can be calculated. Through the fracturing range L_h and the number of fracturing segments n, the fracturing spacing d₀ can be calculated, and since the number of fracturing segments n must be an integer, accordingly, the fracturing time T can be adjusted by feedback until n is an integer ≥1.

Figure 5. Flowchart of the calculation method of hydraulic fracturing process parameters based on damage variables.

Figure 5. Flowchart of the calculation method of hydraulic fracturing process parameters based on damage variables.

3.3. Determination of Key Parameters for Directional Hydraulic Fracturing

• Determination of the extent of the fracturing section

Hydraulic fracturing changes the range of influence of the support stress from x_t to x_td, so it is necessary to fracture all the coal rock body in the range of x_td; thus, the range of fracturing L_h is determined as x_td. Equation (1) is as follows:

$$x_{\mathrm{td}}={\displaystyle \frac{M}{2f}}\left({\displaystyle \frac{\left(\frac{1+\sin \phi }{1 - \sin \phi }{\sigma }_3+\frac{2C\cos \phi }{1 \sin \phi }\right)\left(D_a-D_b\right)}{{\sigma }_{3\mathrm{a}}{ - \mathsf{\sigma }}_{3\mathrm{b}}}}+1\right)\ln \left[{\displaystyle \frac{K\gamma \gamma H}{N_0}}\right]+{\displaystyle \frac{M\beta }{2f}}\ln K$$

(1)

2.

Determination of drilling angle

The currently adopted prefabricated fracturing technique mainly involves cutting transverse slots in the radial direction of the borehole (the direction of the fracture is perpendicular to the axial direction of the borehole). Therefore, the angle of the drill hole can be determined as (90° − θ), where θ is the starting angle of the target hydraulic fracturing crack.

3.

Determination of drilling spacing

On the basis of the highly elliptical morphology of the hydraulic fracture seam, the displacement field relation (2) for the crack in the plane stress state according to the linear elastic fracture mechanics crack is as follows:

$$V\left(x\right)={\displaystyle \frac{2\sigma }{E}}\sqrt{a^2-{\mathrm{x}}^2}$$

(2)

where σ is the fluid pressure in the seam, E is the modulus of elasticity of the coal rock, and a is the length of the fracture. The established crack geometry is shown in Figure 6.

Consider the fracturing crack as a synthesis of many I-shaped penetrating cracks arranged continuously in the direction of the seam length. Applying the formula to any crack leads to the seam width in Equation (3):

$$W(\mathrm{x},\mathrm{z})={\displaystyle \frac{4{\sigma }_x}{E}}\sqrt{H^2-{\mathrm{z}}^2}$$

(3)

Equation (4) is then obtained from the principle of conservation of volume:

$$Q\cdot t=2{{\displaystyle \int }}_0^L{{\displaystyle \int }}_0^{H}2W(\mathrm{x},\mathrm{z})\mathit{dzdx}$$

(4)

In the formula, H is the seam height in the Z direction, which should be a variable, considering that the model will not be solved if it is analyzed as a variable, so it is considered that the seam heights are equal, and the simplified analysis can be obtained as Equation (5):

$$Q\cdot t=4{{\displaystyle \int }}_0^L\left[{{\displaystyle \int }}_0^H{\displaystyle \frac{4{\sigma }_x}{E}}\sqrt{H^2-{\mathrm{z}}^2}\mathit{dz}\right]\mathit{dx}$$

(5)

Replacing σ_(x) with the mean value σ and organizing it gives the crack radius as Equation (6):

$$a={\displaystyle \frac{QTE}{2\sigma \pi H^2}}$$

(6)

The spacing of the drilled holes is given in Equation (7):

$$d=2L={\displaystyle \frac{QTE}{2\sigma \pi H^2}}$$

(7)

Taking Q = 7.2 m³/h, t = 0.5 h, E = 5 GPa, σ = 35 MPa, and H = 5 m, the parameters were substituted into the above equation. The effect of fracturing time T on the extension range of hydraulic fracturing cracks and damage variables was obtained as shown in Figure 7. It can be observed that as the fracturing time increases, the length of the hydraulic fractures exhibits linear growth, and the damage variable of the coal rock mass similarly increases in a linear manner.

4.

Determination of the number of fracturing segments per borehole

Assume that the volume of a fracturing area is V, the width of the area is W, the height is H, and the depth of fracturing is L. Therefore, V = W × H × L. The number of drilled holes is m, the number of fracturing segments per hole is n, and the total number of directional fractures made is N. Then, N = m × n. Here, N can be calculated from the damage variable to find the number of fracturing segments n. The equation is as follows (8):

$$D=\left\{\begin{array}{c}0\\ 1-{\displaystyle \frac{1}{1-\frac{7.56BN{(\frac{QTE}{2\sigma \pi H^2})}^2\cdot {\tau }^2{\mathrm{e}}_{\mathrm{ff}}\ln \left(\cos \frac{\pi \phi }{2}\right)}{V{\phi }^2\cdot {\sigma }^{2}eq\left\{\frac{2}{3}(1+v)+3(1-2v){\left(\frac{{\sigma }_{\mathrm{m}}}{{\sigma }_{\mathrm{eq}}}\right)}^2\right\}}}}\end{array}\right.$$

(8)

The total number of cracks can be obtained as Equation (9):

$$N={\displaystyle \frac{MV{\phi }^2\cdot {\sigma }_{\mathrm{eq}}^2\left\{\frac{2}{3}(1+v)+3(1-2v){\left(\frac{{\sigma }_{\mathrm{m}}}{{\sigma }_{\mathrm{eq}}}\right)}^2\right\}}{7.56(D-1)B{(\frac{QTE}{2\sigma \pi H^2})}^2\cdot {\tau }_{\mathrm{eff}}^2\ln (\cos \frac{\pi \phi }{2})}}$$

(9)

The number of fracturing segments per borehole is Equation (10):

$$n={\displaystyle \frac{n}{m}}={\displaystyle \frac{MV{\phi }^2\cdot {\sigma }_{\mathrm{eq}}^2\left\{\frac{2}{3}(1+v)+3(1-2v){\left(\frac{{\sigma }_{\mathrm{m}}}{{\sigma }_{\mathrm{eq}}}\right)}^2\right\}}{7.56(D-1)B{(\frac{QTE}{2\sigma \pi H^2})}^2\cdot {\tau }_{\mathrm{eff}}^2\ln (\cos \frac{\pi \phi }{2})m}}$$

(10)

The meanings of the variables involved in the above 10 equations are shown in

Table A1. The meanings of the variables involved in the equations.

Variable	Meaning
D	Damage variable
Lh	Range of fracturing
Q	Initial fracturing displacement
T	Fracturing time
R	Radius of fracture propagation
d	Spacing of the drilled holes
d0	Fracturing spacing
m	Number of drilled holes
n	Number of fracturing segments
xt, xtd	Range of influence of the support stress
σ3	Minimum principal stress in the initial state
σ3a, σ3b	Minimum principal stress in a given state
φ	Angle of internal friction
M	Material parameters
C	Cohesion of the soil
Da, Db	Location-specific damage variables
β	Inclination of the crack
K, Kγγ	Soil permeability coefficient
V(x)	Displacement function, indicating the displacement at position (x)
E	Modulus of elasticity of the coal rock
L	Fracture depth
H	Length of the crack
σ	In-seam hydraulic pressure
σx	Liquid pressure in the seam in the x-axis direction
x	Horizontal distance
z	Distance in the vertical direction
W(x,z)	Slit width of the fractured fissure at a specific location (x,z)
t	Fracturing time
a	Radius of the crack
B	Constants of the material
τeff	Effective shear stress
ϕ	Porosity of the substrate
V	Volume of fractured area
σeq	Constant force
σm	Average stress
ν	Velocity of fluid percolation in the matrix

.

The effects of the density of drilling holes and the number of fracturing segments on the fracture density and damage variables are shown in Figure 8. Drill hole density refers to the number of drill holes per unit volume of coal rock, while fracture density represents the number of fractures per unit volume. As the density of drill holes and the number of fracturing segments increase, the fracture density and damage variables rise steadily. However, the rate of increase slows as the process continues.

According to the characteristics of the lithology of the top plate of the Yanghuopan 3⁻¹ coal seam, the parameters of the directional hydraulic fracturing of the top plate were derived based on the above simulation and the calculation method of the hydraulic fracturing process parameters of the damage variables:

• Pre-fracture pressure: Controlled at 10~25 MPa.
• Pre-fracture penetration radius: 15 m.
• Fracturing time of single hole: 30 min.
• Number of fracturing segments per borehole: 5~6.
• Fracturing method: Transverse fracture.
• The ratio of coal rock active water was determined according to the actual water quality coefficient of the coal seam at the site.
• Pre-cracking volume: 0.7 m³ for single hole, 1~2 m³ cumulative.
• Pre-cracking flow rate: according to experience, it should be 0.7~2.0 m³/h for dynamic pressure pre-cracking.

4. Directional Hydraulic Fracturing Fracture Evolution Law

4.1. Establishment of RFPA 3D Numerical Simulation

Directional hydraulic fracturing is an important technical way to preform artificial fractures in the coal rock body, and then carry out high-pressure water injection fracturing to realize the structural modification and stress transfer of the coal rock body. The directional hydraulic fracturing with preformed cracks is one of the commonly used hydraulic fracturing methods in coal mines [18,19]. The principle is shown in Figure 9.

The fracture is first prefabricated in the direction specified by the borehole, and a stress concentration is generated at the end of the prefabricated fracture. When water is injected into the borehole, the hydraulic fracture crack preferentially expands along the direction of the wedge groove and generates pore water pressure P₁ at the front end of the crack A. Combined effects of P₁ and the stress field cause the hydraulic fracture crack to initiate and expand. The RFPA 3D numerical model of coal rock body is established to analyze the crack initiation and expansion law of the fracture cracks as shown in Figure 10. The drill hole is first set in the center of the model, and then the drill hole is fractured. The geometry of the numerical model is 20 m × 20 m, the model is divided into 240 × 240 cells, totaling 57,600. The diameter of the drill holes is 150 mm. The model type is a Moore Coulomb model. This information has been added to the manuscript. Displacement convergence criterion, stress convergence criterion, residual convergence criterion and maximum number of iterations convergence criterion were used during numerical simulation. The finite element method (FEM) and finite difference method (FDM) were used for the numerical simulation of the rock rupture process. During the rupture process, RFPA 3D may encounter numerical instability, especially when the rupture propagates rapidly, and the local stress concentration may lead to computational dispersion. In order to reduce the numerical instability, we adopt a cell softening model to try to ensure numerical stability and improve the computational accuracy: a damage model is used in the rupture simulation to deal with the nonlinear behavior of the material and to prevent instability caused by sudden stress release. According to the burial depth of 100 m and the lateral pressure coefficient of 0.75 (The Lateral Pressure Coefficient (LPC) is commonly used in soil mechanics and geotechnical engineering, especially when analyzing the lateral pressure of soil or rock. This coefficient represents the ratio between the pressure on a soil or rock in the vertical direction and the pressure on it in the horizontal direction. When the lateral pressure coefficient (K = 0.75) is used, this means that the stress in the horizontal direction is 75% of the stress in the vertical direction. In other words, if the stress in the vertical direction is 100 KPa, the stress in the horizontal direction will be 75 KPa), the stress applied to the upper boundary of the model is σ₁ = 8.9 MPa, and that applied to the left and right boundaries is σ₃ = 6.3 MPa. because the quantitative pumps are usually used for fracturing in the field, the water injection with constant displacement is used in the simulation. Constant displacement water injection. The hydraulic fracturing process was simulated by increasing the water pressure in the borehole step by step. The initial value of water pressure in the borehole is 2 MPa, and 0.5 MPa is added in each step, and the specific parameters are shown in

Table 1. Numerical simulation of mechanical parameters.

Parameters	Mean Value of Modulus of Elasticity/MPa	Degree of Homogeneity/m	Friction Angle/(°)	Average Compressive Strength/MPa	Poisson’s Ratio
Parameter value	9600	3	37	30	0.25
Parameters	Permeability Coefficient/m·d−1	Pore Water Pressure/MPa	Permeability Coefficient After Rupture	Pore Water Pressure Coefficient	Loading Method
Parameter value	0.001	1.0	10	0.1	Stress loading

.

4.2. Factors Affecting Crack Initiation and Expansion

4.2.1. Effect of Different Angles on Crack Initiation and Expansion

• 0° prefabricated crack

When the angle of the prefabricated crack was 0°, high stress concentration occurred at the crack end during the stress accumulation stage (1~6.5 MPa). This is analogous to mining a thin protective layer in the rock, which disrupts the stress concentration zone surrounding the borehole and transfers the high stress to the crack tip. This mechanism effectively fractures the intact rock mass, leading to a “corkscrew effect” that enhances the rock's permeability, increases the unloading range of a single borehole, and facilitates complete unloading of the surrounding coal body. During the stage of local crack expansion (6.5~8.5 MPa), with the increase in drilling pressure, at step 5 (the water pressure in the drilling hole was 6.5 MPa), fine cracks formed at the end of the cracks. By step 10, a macroscopic microcrack appeared at the crack tip, and the crack progressively extended and expanded as the water pressure in the borehole increased. At this moment, the cracks were distributed on both sides of the borehole and penetrated with the borehole. During the crack extension and penetration phase (8.5~27 MPa), once the borehole stress exceeded 8.5 MPa, the cracks connected with the borehole and expanded outward, forming additional crack groups, as depicted in Figure 11a.

2.

30° prefabricated crack

When the angle of the prefabricated crack was 30°, in the stress accumulation stage (1~5 MPa), the water pressure in the borehole was 3 MPa, and a high stress concentration phenomenon appeared at the end of the 30° crack in the borehole. With the increase in applied pressure in the borehole, at step 5 (water pressure in the borehole was 5 MPa), fine cracks formed at the crack end. At the stage of localized crack extension (5~8.5 MPa), with the increase in the applied pressure in the borehole, by step 10 (the water pressure in the borehole was 8.5 MPa), a macroscopic tiny crack formed at the crack end, and the crack gradually extended and expanded. At this point, the cracks were distributed on both sides of the borehole, and the direction of crack expansion was upward and downward, centered on the end of the crack. At the stage of crack expansion and penetration (8.5~27 MPa), when the stress in the borehole increased to 8.5 MPa (after 12 steps), the formed cracks were connected with the borehole, penetrated with the cracks, and further expanded outward to form a larger number of crack clusters, as shown in Figure 11b.

3.

60° prefabricated crack

When the angle of the prefabricated crack was 60°, high stress concentration was observed at the crack end during the stress accumulation stage (1~5 MPa). At the stage of local crack extension (5~8 MPa), with the increase in pressure applied to the borehole, at step 10 (the water pressure in the borehole was 8 MPa), a macroscopic tiny crack was formed at the end of the crack, and the crack gradually extended and expanded. At this time, the cracks were distributed on both sides of the borehole to expand outward, and the whole crack extension direction was 60°. At the crack extension and penetration stage (8~27 MPa), when the borehole stress increased to 8 MPa, i.e., after 12 steps, the cracks formed were connected with the borehole, penetrated with the cracks, and further extended outward, forming more crack groups, as shown in Figure 11c.

4.

90° prefabricated crack

When the prefabricated crack angle was 90°, high stress concentration occurred at the crack end during the stress accumulation stage (1~4 MPa), as in the case of the prefabricated crack angle of 60°. At the stage of local crack extension (4~7 MPa), with the increase in applied pressure in the borehole, at step 5 (the water pressure in the borehole was 4 MPa), a macroscopic tiny crack was formed at the crack end, and the crack gradually extended and expanded. At step 10, the water pressure in the borehole was 7 MPa, the cracks were distributed on both sides of the borehole and extended outward, and the whole crack extended to the center of the borehole at 90°. At the stage of crack extension and penetration (7~27 MPa), when the stress in the borehole increased to 7 MPa, after 10 steps, the cracks formed were connected with the borehole, penetrated with the cracks, and further extended outward to form more crack clusters; the crack clusters involved a large radius, as shown in Figure 11d.

In conclusion, fracture drilling plays a crucial role in directing the expansion of hydraulic fractures. Fracture initiation and expansion follow the direction of drilling, and as the angle with the minimum stress increases, the range of fracture expansion widens.

4.2.2. Effect of Fracturing Displacement on Fracture Initiation and Expansion

The effect of fracturing displacement on the initiation and propagation patterns of hydraulic fracturing cracks is illustrated in Figure 12. Initially, the cracks developed and extended along the direction of the prefabricated fracture, gradually deflecting toward the vertical orientation and eventually expanding along the vertical axis. As the fracturing displacement increased, the likelihood of deflection decreased over longer fracture lengths, resulting in a larger angle between the fracture direction and the vertical axis. This occurs because increased fracturing displacement accelerates the growth of the pore water pressure gradient at the crack tip, altering the pore water pressure distribution and reducing its sensitivity to the principal stress, thereby stabilizing the fracture extension direction. Additionally, there was a linear increase in the damage variable of the weak structural body as the fracturing displacement increased. This is attributed to the enhanced directional extension of hydraulic fractures, which causes the crack angle within the coal rock to approach 45°.

4.2.3. Effect of Fracturing Segment Spacing on Fracture Initiation and Expansion

The effect of fracturing segment spacing d₀ on the fracture initiation and expansion pattern of hydraulic fracturing cracks is shown in Figure 13. When d₀ = 2 m, some of the lower segmented hydraulic fracturing cracks converged with the upper segmented hydraulic fracturing cracks, while the other cracks expanded independently. Unaffected by the upper segmented hydraulic fracturing cracks, 2~3 hydraulic fracturing cracks were produced eventually. When d₀ = 3 m and 4 m, the hydraulic fracturing cracks all expanded independently along the direction of σ₁, with no influence on each other, and finally produced three hydraulic fracturing cracks. When d₀ = 5 m and 6 m, the fracturing section could only carry out two fracturing segments, and finally, two hydraulic fracturing cracks were formed. When d₀ = 7 m, the fracturing section could only carry out one fracturing section and eventually formed one hydraulic fracturing fracture. Overall, with the increase in segment spacing, the number of hydraulic fracturing cracks produced first increased and then decreased. The number of hydraulic fracturing cracks was highest when the segmentation distance was 3 m. With the increase in segment spacing, the damage variable of the weak structural body increased and then decreased, and the damage variable was the largest when the segment distance was 3 m. The damage variable of the weak structural body increased and then decreased with the increase in segment spacing. The reason for this is that overly small and overly large segment spacing will result in a decrease in the number of hydraulic fracturing cracks eventually formed, which will reduce the crack density in the coal rock body.

4.2.4. Effect of Fracture Number on Fracture Initiation and Propagation

The effect of the number of cracks n_a on the expansion pattern of hydraulic fracturing cracks is shown in Figure 14. When the number of cracks n_a = 1, only one hydraulic fracturing crack was produced, with hydraulic fracturing cracks initiating and expanding along the 45° crack direction. When the number of cracks n_a = 2, two hydraulic fracturing cracks were produced, and the hydraulic fracturing cracks were also initiated and extended along the 45° crack direction. When the number of cracks n_a = 3, only one hydraulic fracturing crack was produced, and the hydraulic fracturing crack was initiated and expanded along the 90° crack direction. From the above, it can be seen that with the increase in the number of cracks, the number of hydraulic fracturing cracks first increased and then decreased, and there is a maximum number of hydraulic fracturing cracks for two sets of crack numbers. As can be seen, increasing the number of cracks does not increase the number of hydraulic fracturing cracks.

5. Field Application of Directional Hydraulic Fracturing Unloading Technology

Based on the mechanism of directional hydraulic fracturing and the key parameters of the 3⁻¹ coal seam in Yanghuopan Coal Mine, combined with the theoretical knowledge and experimental data of the fracture evolution law, the feasibility of the directional hydraulic fracturing decompression roof disaster prevention and control technology program was verified by on-site implementation in the 30,116 and 30,119 working faces of Yanghuopan Coal Mine.

5.1. Implementation Program Design

• 30,116 working face

After collecting relevant information and combining the data from the on-site research, we carried out targeted roof management construction for the six key areas where the stresses in the 30,116 working face are concentrated. The construction area map is shown in Figure 15.

The layout of the fracturing drill holes is depicted in Figure 16. A total of 35 drill holes were constructed, categorized into two types: A and B. Drill hole A was 80 m deep with an inclination angle of 17°, forming a 90° angle with the working face and coal pillar. Drill hole B was 50 m deep with an inclination angle of 27°, forming a 10° angle with the coal wall of the working face on the return side and a 45° angle with the coal pillar on the auxiliary transportation side. In the 30,116 auxiliary transportation channel, 12 holes of type A and 6 holes of type B were drilled, while 12 holes of type A and 5 holes of type B were drilled in the 30,116 return wind channel.

2.

30,119 working face

A total of 44 drill holes were constructed, categorized into four types: C (40 m depth, 20° inclination), D (30 m depth, 30° inclination), E (40 m depth, 20° inclination), and F (20 m depth, 40° inclination). The arrangement included 14 holes of type C, 14 holes of type D, and 6 holes of type F on the cut-off mining side. Additionally, 3 holes of type D and 2 holes of type E were placed in the transportation chute, while 3 holes of type D and 2 holes of type E were installed in the return chute. The layout for cut-off hydraulic fracturing and initial roof placement in the 30,119 working face is illustrated in Figure 17.

5.2. Analysis of Fracturing and Pressure Relief Effects

• Fracturing effect analysis

When most of the drill holes in the 30,116 return wind trench were pressed to the position of 11~17 m from the mouth of the borehole (most of the drill holes in the auxiliary transportation trench were pressed to the position of 20~29 m from the mouth of the borehole), the anchor ropes started to produce water. The closer the drill holes were to the borehole mouth, the greater the fracturing out of water, as shown in Figure 18a,b. This reflects that the surrounding rock above 17 m (29 m) of the borehole is relatively hard and complete and the fractures did not develop; under the action of high-pressure water, the fractures were also extended and diffused and new fractures were generated. The fractures in the surrounding rock below 17 m (29 m) were more developed, the penetration under the action of high-pressure water was better, and a fracture network was formed, as shown in Figure 19. Most of the fracturing values are distributed between 10 and 25 MPa, without an obvious pressure drop, and the phenomenon of neighboring holes drenching during the fracturing indicates that the cracks were extended to the neighboring holes, and the design of the fracturing scheme is reasonable.

The primary difference between the 30,119 and 30,116 working faces is that in the 30,119 working face, most of the drill holes were pressurized to a depth of 10–19 m from the hole opening, at which point the anchor cables began to release water (as shown in Figure 18c,d). The majority of the fracturing pressures ranged between 10 and 22 MPa. The remaining principles governing the process are consistent with those applied in the 30,116 working face.

2.

Analysis after pressure relief

To further assess the impact of roof management in the 3⁻¹ coal seam, pressure monitoring was performed on the brace resistance of the 30,116 working face at the Yanghuopan Mine following roof depressurization. Two monitoring channels were selected in each section of the upper, middle, and lower parts of the working face to analyze the mine pressure data (as shown in Figure 20). For data comparability, six channels—4-4, 6-6, 10-10, 12-12, 17-17, and 19-19—were used for data acquisition.

Analysis of the measured stent resistance data from the 30,116 working face, as shown in Figure 20, revealed distinct cyclic variations in stent resistance. The working face exhibited a cycle pressure step of 10–12 m, a cycle pressure range of 37–41 MPa, and a cycle duration of 3–4 days. This indicates that the directional hydraulic fracturing technology effectively reduced concentrated stress in the roof plate of the working face gradually, alleviated dynamic pressure in the quarry, and allowed the stent to function normally.

6. Conclusions

Analyzing the characteristics of the 30,116 face at Yanghuopan and the overburden pressure in the quarry, it is evident that the mining of the 30,116 face is influenced by the concentrated stress from the coal pillar in the 2⁻² coal seam void. During the extraction process above the coal pillar, dynamic pressure occurs in the quarry, posing a potential risk to the safe operation of the working face. As a result, it is proposed to implement directional hydraulic fracturing and pressure relief technology to prevent and mitigate roof-related hazards.

• The key parameters of the 3⁻¹ coal seam were determined based on the analysis of the mechanism of unloading pressure of directional hydraulic fracturing.
• Various angles, fracture displacements, segment spacings, and fracture numbers were modeled using RFPA 3D numerical simulation software. The influence of these factors on hydraulic fracture initiation and propagation was comparatively analyzed, leading to the formulation of the directional hydraulic fracture evolution law. The results indicated that the extent of damage to the weak structural body of the coal rock is dependent on the hydraulic fracturing crack parameters. Variations in hydraulic fracturing process parameters lead to differing fracture morphologies, damage variables, and corresponding stress transfer effects in the weak structural body. Additionally, the hydraulic fracturing process parameters combined with damage variables determined the use of transverse fractures for directional fracturing in the 30,116 and 30,119 working faces of the 3⁻¹ coal seam at the Yanghuopan Mine. The drill holes were spaced 15 m apart, with 5 to 6 fracturing segments per hole, a single-hole fracturing duration of 30 minutes, and pre-fracturing pressure controlled between 10 and 25 MPa.
• The on-site monitoring results show that under the effect of directional hydraulic pressure on the rock body of the roof slab of the 30,116 general mining face (under the effect of directional hydraulic pressure on the roof slab of the 30,119 general mining face with open cut eyes), the fracture network expansion reached the expected effect. The fracture network developed well, effectively releasing concentrated stress in the rock body. During mining operations, the bracket pressure remained stable and exhibited clear periodicity, achieving the desired effect of pressure relief to prevent and control roof-related hazards.