Study on the Long-Distance Gas Pre-Drainage Technology in the Heading Face by Directional Long Borehole

Abstract

Gas control in the heading face of a coal roadway is an important and difficult point in coal mining in China. On the basis of analyzing the disadvantages of high gas control cost and long drainage period in the existing mine heading face, a long-distance pre-drainage method of long-distance drilling is proposed to control the gas in the heading face so as to improve the tunneling speed. Applied to the engineering geological conditions of Changcun coal mine, the technology is studied in detail. First, a gas migration model considering permeability changing with time is established, and the model is put into the numerical simulation software to study the variation law of permeability and gas pressure under the conditions of single borehole and multi-borehole drainage. The results show that with the increase of drainage time, the permeability around the borehole increases gradually, the gas pressure decreases gradually, and the permeability at the borehole boundary increases the most, reaching 1.2 times the initial permeability. In the process of multi-borehole drainage, there will be mutual influence between boreholes, but with the increase of borehole spacing, the degree of this influence gradually decreases. Second, according to the results of numerical simulation, a reasonable gas drainage scheme is designed and applied in the field. The field application shows that the technology has a good gas drainage effect, the gas drainage concentration and flow are at a high level for a long time, the drilling cuttings quantity is always lower than the critical value, and the excavation length of roadway increases by more than 50 m per month. These results indicate that this technology is a promising method to realize the safe and rapid excavation of a mine coal roadway.

1. Introduction

Gas disaster is one of the main disasters in coal mines in China. The occurrence of gas accidents not only affects the safety production of the mine, but also seriously affects the life safety of workers. Therefore, it is necessary to control the coal seam gas [1,2,3]. Coal seam gas drainage is an effective method to prevent the occurrence of gas accidents. At the same time, the use of the extracted gas not only increases the total amount of energy, but also is conducive to environmental protection [4,5].

According to the different location of drainage, coal seam gas drainage can be divided into working face drainage and heading face drainage. Among them, the heading face is difficulty of gas control, which often leads to abnormal gas emission accidents due to a poor drainage effect [6,7]. At present, gas drainage in the heading face mainly includes cross-measure borehole gas drainage technology in the floor rock roadway and in-seam borehole gas drainage technology in the heading face [8,9,10]. Cross-measure borehole gas drainage technology is to extract coal seam gas by drilling in the rock roadway under the coal seam (Figure 1a). This method has the advantages of high safety, good borehole structure stability and long drainage time [11,12,13]. However, due to the need to dig out a rock roadway, there are some shortcomings, such as high cost, large engineering quantity, long construction period and so on [14]. In-seam borehole gas drainage technology is to extract the coal seam gas by drilling along the coal seam in the heading face (Figure 1b). Compared with the cross-measure borehole technology, this method does not need to excavate a rock roadway, which reduces the construction cost. However, the length of one-time pre-drainage is short, and can only reach 80–100 m. After passing the inspection, the excavation can be carried out, and then the next drainage cycle is carried out, that is “Drilling-Drainage-Inspection-Excavation-Drilling” [15,16], which leads to the slow excavation speed of roadway.

With the development of technology, directional long borehole gas drainage technology has been gradually applied in coal mines. Wang et al. [17] summarized the application of directional long boreholes in China’s coal mines, including pre-mining and post-mining drainage, and found that it can effectively improve the gas drainage efficiency. Lu et al. [18] conducted the gas drainage testing of directional long boreholes in Daning Coal Mine, China, which effectively controlled the gas outburst accident risk of the mine. Compared with the ordinary borehole drainage method, the proportion of coal seam drilling is higher, and the drainage effect is better. Wang et al. [19], Li et al. [20] and Hao et al. [21] studied the gas drainage in goaf by directional long boreholes instead of a separate drainage roadway, analyzed and determined the reasonable horizon of directional long boreholes in the roof, and found that this method can effectively solve the problem of gas overrun in the upper corner of the working face, and can replace the roadway-based gas drainage, thereby saving much work.

Based on the advantages of the directional long borehole and the disadvantages of existing ordinary borehole gas drainage in the heading face, a new method of long-distance gas drainage in the heading face by using the directional long borehole is proposed. Through the construction of boreholds by drilling in the completed roadway (not floor rock roadway), the coal seam gas near the pre-excavation roadway is extracted. The construction diagram is shown in Figure 2. Compared with the cross-measure borehole gas drainage technology, this method does not need to excavate a rock roadway, and has the advantages of low cost. Compared with the in-seam borehole gas drainage technology in the heading face, the drainage distance is longer, which increases the length of what was once the excavation roadway, and can extract the coal seam gas for a long time. Taking the return air roadway of working face 2302 in Changcun coal mine as an example, this paper establishes a gas migration model considering the change of permeability with drainage time, studies the gas migration law and borehole layout of directional long borehole drainage, carries out industrial testing on site, and analyzes the drainage effect.

2. Project Summary and Basic Parameter Test

The Changcun coal mine is located in the Changzhi City, Shanxi Province, China. It is a large modern mine of Shanxi Lu’an environmental protection energy development Co., Ltd. (Changzhi, China). The main mining area is 3# coal seam. The geological structure is mainly fold. The strata strike nearly north-south and incline westward with an inclination of 3–6°. In the east, it is mainly a tilt structure with near east-west undulation. In the west, there are nearly north-south folds. The geological structure provides a good environment for coal seam gas storage. The gas content of the coal seam is generally 8–10 m³/t, and the gas content is large.

2.1. Project Summary

2302 working face is located in No. 23 mining area of the mine, mining 3# coal seam; the thickness of the coal seam is generally 4.84–7.32 m, and the average thickness is 6.09 m. The average depth of the coal seam is about 500 m. The roadway layout adopts the “two air intakes and one return” mode; the 2302 belt transportation roadway and auxiliary transportation roadway provide for entry of the air, and the 2302 return air roadway returns the air. A 2302 belt transportation roadway and auxiliary transportation roadway have been excavated, and 2302 return air roadway is the pre-excavation roadway. The mine location and roadway layout of 2302 working face are shown in Figure 3. According to the field measurement data, the gas content of 2302 working face is 8.5 m³/t, the gas pressure is 0.35 MPa, and the permeability is 8.09 × 10⁻⁸ m². The attenuation coefficient of borehole gas flow is 0.1726–0.3025 d⁻¹. According to the difficulty degree of coal seam drainage table (

Table 1. Difficulty degree of coal seam gas drainage.

Classification	Attenuation Coefficient of Borehole Gas Flow/d−1	Coal Seam Permeabilit/m2
Easy to be extracted	<0.003	>2.5 × 10−4
Can be extracted	0.003–0.05	2.5 × 10−6–2.5 × 10−4
Difficult to be extracted	>0.05	<2.5 × 10−6

), the gas drainage in 2302 working face is classified as difficult.

The gas pressure is produced by the thermal movement of free gas. Changcun coal mine has high gas content and low gas pressure, which indicates that the coal seam has strong adsorption capacity, and the adsorption content of coal seam gas accounts for more. Gas desorption is a slow process. Research shows that under natural conditions, when the coal particle size is 1 cm, the time required for desorption of 90% of the gas is 15 years [22]. Therefore, for the coal seam with strong adsorption capacity, gas drainage should be carried out for a long time even under negative pressure.

2.2. Basic Parameter Test

The coal samples were taken on site, and the coal samples were processed into standard samples and pulverized with different particle sizes, respectively. The coal samples are used to test the mechanical parameters, industrial analysis and gas basic parameters of coal. The test results are shown in

Table 2. Mechanical parameters of coal samples.

Elastic Modulus (E/GPa)	Poisson’s Ratio (υ)	Cohesion (c/MPa)	Internal Friction Angle (φ/°)
1.1	0.3	0.8	28

and

Table 3. Industrial analysis and gas basic parameters.

Water Content (W/%)	Ash Content (A/%)	Porosity (η/%)	Maximum Gas Adsorption Capacity (a/m3/kg)	Adsorption Constant (b/MPa−1)
1.14	7.76	4.29	37.08	0.82

.

3. Gas Migration Equation in Coal Seam

3.1. Control Equation of Coal Seam Permeability Considering Gas Pressure Variation and Gas Adsorption and Desorption

The research shows that coal seam permeability k can be expressed as an exponential function related to effective stress [23]:

$$k=k_0\exp \left(-3C_f\Delta {\sigma }_e\right)$$

(1)

where k₀ is the initial permeability, m²; C_f represents cleat volume compressibility, MPa⁻¹; Δσ_e is effective stress variation, MPa.

The cleat volume compressibility C_f can be expressed by the following formula [24]:

$$\left\{\begin{array}{c}{C}_f=\frac{1}{K_p}\\ K_p=\eta ·K=\eta ·\frac{E}{3\left(1-2\nu \right)}\end{array}\right.$$

(2)

where K_p is pore bulk modulus, GPa; η is porosity of coal seam, %; K is elastic modulus of coal matrix, GPa; E is elastic modulus of coal, GPa; υ is Poisson’s ratio of coal.

There are three changes in the process of drilling and gas drainage. Firstly, the coal seam is disturbed in the process of drilling, and the stress around the borehole redistributes, resulting in the change of effective stress. Secondly, the gas in a free state is extracted under negative pressure, and the decrease of gas pressure leads to the increase of effective stress, resulting in the compression of coal pores and the decrease of gas flow channels. Thirdly, the decrease of gas pressure promotes gas desorption and coal matrix shrinkage, resulting in the increase of coal fracture channels. Therefore, the variation of effective stress can be expressed as three parts:

$$\Delta {\sigma }_e=\Delta {\sigma }_s+\Delta {\sigma }_p+\Delta {\sigma }_x$$

(3)

where Δσ_s is effective stress variation caused by stress redistribution, MPa; Δσ_p is effective stress variation caused by gas pressure drop, MPa; Δσ_x is effective stress variation caused by gas desorption, MPa.

Coal is a kind of soft elastic-plastic body. After stress redistribution, the tangential stress around the borehole will be higher than the strength of coal, so that the coal will be destroyed and a plastic zone and elastic zone will appear. Assuming that the coal is in the limit equilibrium state in the plastic zone, the tangential stress around the borehole can be described by the following formula [25,26]:

$$\Delta {\sigma }_t=\left\{\begin{array}{c}c·cot\phi ·\left[\left(\frac{1+sin\phi }{1-sin\phi }\right)·{\left(\frac{2x}{R_0}+1\right)}^{\frac{2sin\phi }{1-sin\phi }}-1\right],x\le H\\ {\sigma }_0·\left[1+\frac{4H^2}{{\left(2x+R_0\right)}^2}\right]-\frac{4H^2}{{\left(2x+R_0\right)}^2}·c·cot\phi \left[{\left(\frac{2H}{R_0}\right)}^{\frac{2sin\phi }{1-sin\phi }}-1\right],x>H\end{array}\right.$$

(4)

where σ₀ is initial stress, MPa; c is cohesion of coal, MPa; φ is internal friction angle, °; x is distance from coal body to borehole wall, m; R₀ is borehole diameter, m; H is distance from plastic zone boundary to borehole wall, m. H can be expressed as follows:

$$H=\frac{R_0}{2}\left\{{\left[\frac{{\sigma }_0·\left(1-sin\phi \right)}{c·cot\phi }+1\right]}^{\frac{1-sin\phi }{2·sin\phi }}-1\right\}$$

(5)

According to the coal seam buried depth of 500 m, the initial stress σ₀ is 12.5 MPa, the diameter of long hole is 113 mm. The tangential stress as shown in Figure 4 can be calculated by substituting the cohesion c and internal friction angle φ measured in the laboratory and original stress and borehole diameter into Equations (3) and (4). If the position where the stress changes by 5% is taken as the influence boundary of the borehole, it can be seen from Figure 4 that the radius of the influence range of the directional long borehole is 0.313 m, which is smaller than the thickness of the coal seam. Therefore, the influence of the stress change around the borehole on the permeability of the coal seam can be ignored, and the change of the effective stress can be expressed as follows:

$$\Delta {\sigma }_e=\Delta {\sigma }_p+\Delta {\sigma }_x$$

(6)

The effective stress variation caused by gas pressure drop can be expressed by the following formula:

$$\Delta {\sigma }_p=P_0-P$$

(7)

where P₀ is initial gas pressure of coal seam, MPa; P is gas pressure of coal seam, MPa.

The effective stress variation caused by gas adsorption and desorption can be expressed by the following formula:

$$\Delta {\sigma }_x=K·\Delta {\epsilon }_x$$

(8)

where ε_x is the coal bulk strain caused by gas adsorption and desorption, which satisfies the Langmuir equation. It can be expressed by the following formula [24]:

$${\epsilon }_x=\frac{{\epsilon }_l·P}{P+P_l}$$

(9)

where ε_l is Langmuir volume strain constant; P_l is Langmuir pressure constant.

The control equation of coal seam permeability considering gas pressure variation and gas adsorption and desorption can be obtained by simultaneous Formulas (1), (6)–(9):

$$k=k_{0}exp\left\{-3C_f\left(P_0-P\right)·\left[1-\frac{E}{3·\left(1-2\nu \right)}·\frac{{\epsilon }_l{P}_l}{\left(P+P_l\right)\left(P_0+P_l\right)}\right]\right\}$$

(10)

3.2. Control Equation of Gas Flow

Coal seams are porous media, in which gas seepage conforms to the mass conservation equation [27]:

$$\frac{\partial X}{\partial t}+\nabla \left(\rho V\right)=0$$

(11)

where X is the gas content in unit volume coal, kg/m³; t is the time variable, s; ρ is the gas density in the coal seam, kg/m³; V is the gas seepage velocity, m/s.

There are two forms of gas in a coal seam, namely free state and adsorption state. Therefore, the gas content in a coal seam includes free gas content and adsorption gas content, which can be expressed by a gas state equation and Langmuir equation, respectively:

$$\left\{\begin{array}{c}{X}_1=\eta \rho \\ X_2=\frac{abp}{1+bp}·\frac{100-A-W}{100}·\frac{1}{1+0.31W}\\ X_3=X_1+X_2\end{array}\right.$$

(12)

where X₁ is the free gas content, kg/m³; X₂ is the content of adsorbed gas, kg/m³; a is the maximum gas adsorption constant per unit mass of coal, m³/kg; b is the adsorption constant of coal, MPa⁻¹; A is ash content of coal, %; W is moisture content of coal, %.

Assuming that the gas is an ideal gas, the gas density in the coal seam is as follows:

$$\rho =\frac{M_{g}P}{RT}$$

(13)

where M_g is molecular weight of gas, 16 g/mol; R is the ideal gas constant, 8.314 J/(mol·K); T is the absolute temperature, K.

It is assumed that the flow of gas in the coal seam conforms to Darcy’s law [8,27]:

$$V=-\frac{k}{\mu }\nabla P$$

(14)

where k is the permeability of coal seam, m²; μ is the dynamic viscosity of gas, Pa·s.

The control equation of gas flow can be obtained by simultaneous Formulas (11)–(14):

$$\left[\frac{M_g\eta }{RT}+\frac{ab}{{\left(1+bp\right)}^2}·\frac{100-A-W}{100}·\frac{1}{1+0.31W}\right]\frac{\partial P}{\partial t}-k·\frac{M_g}{\mu RT}\nabla \left(P\nabla P\right)=0$$

(15)

It can be seen from Formulas (10) and (15) that gas drainage and permeability are influenced by each other. With gas drainage, gas pressure decreases and gas desorption occurs, which affects the permeability. Accordingly, the change of permeability will affect the results of gas extraction. The relationship between gas drainage and permeability is shown in Figure 5.

4. Variation of Permeability and Gas Pressure around Borehole

4.1. Single Borehole Drainage

4.1.1. Geometric Model and Parameter Setting

Since the directional borehole is parallel to the pre-excavation roadway after construction in the design area, the geological conditions of the area through which the borehole passes have little change, so the numerical calculation model can be simplified into a local three-dimensional model. It is assumed that gas is the only flowing gas in the coal seam and the coal seam is isotropic. According to the coal seam thickness and gas parameters of Changcun coal mine, the COMSOL numerical simulation software is used to establish the coal seam gas migration model. In order to reduce the influence of boundary effect, the length of the model is 70 m, the height is 6 m, and the diameter of the borehole is 113 mm. The model diagram is shown in Figure 6. At the same time, a 15 m long monitoring line is arranged from the borehole center to the depth of coal seam, and four monitoring points are arranged at 1 m, 3 m, 5 m and 7 m away from the borehole center to monitor the change of gas parameters. There are no flow boundary conditions around the model, and constant pressure boundary conditions around the borehole simulate the drainage pressure, and the drainage pressure is 20 kPa. The PDE module of numerical simulation software is used to substitute Formulas (10) and (15) into the calculation, and the calculation parameters are listed in

Table 4. Numerical calculation parameters.

Parameter	Numerical Value
Elastic modulus of coal (E/GPa)	1.1
Poisson’s ratio of coal (υ)	0.3
Internal friction angle of coal (φ/°)	28
Porosity of coal (η/%)	1.21
Original gas pressure (P0/MPa)	0.35
Maximum gas adsorption capacity of coal (a/m3·kg−1)	37.08
Adsorption constant of coal (b/MPa−1)	0.82
Ash content of coal (W/%)	1.14
Moisture content of coal (A/%)	7.76
Langmuir volume strain constant (εl)	0.01266
Langmuir pressure constant (Pl/MPa)	4.31
Molecular weight of gas (Mg/g/mol)	16
Ideal gas constant (R/J·mol−1·K−1)	8.314
Dynamic viscosity of gas (μ/Pa·s)	1.84 × 10−5
Absolute temperature (T/K)	273
Initial permeability of coal seam (k0/m2)	8.09 × 10−8

.

4.1.2. Mesh Independence Test of Model

In order to study the influence of model mesh on simulation results, it is necessary to test the mesh independence. The mesh is divided into normal, fine and finer, as shown in Figure 7. According to the theoretical formula, the permeability and gas pressure distribution law of drainage time is 10 d are simulated, as shown in Figure 8.

It can be seen from the figures that the mesh division type is from normal to finer, the permeability and gas pressure distribution around the borehole are almost unchanged. In order to more clearly analyze the influence of mesh division type on simulation results, the data measured by the monitoring line are plotted into a curve, as shown in Figure 9.

It can be seen from the figures that the three different mesh types have obtained the same observation curve, so it can be determined that the mesh type has no effect on the simulation results of single borehole drainage. In order to save the time of numerical simulation, the mesh type is selected as normal mode in the simulation of single borehole drainage.

4.1.3. Permeability Variation Law

According to the theoretical formula, the numerical simulation software is used to calculate the distribution characteristics of permeability around the borehole when the drainage time is 50 d, 100 d, 150 d, 200 d, 250 d and 300 d respectively, so as to observe the variation law of permeability around the boreholes. The curves from the monitoring data of monitoring line and monitoring points are shown in the figures below.

Figure 10 and Figure 11 show the change of coal permeability around the borehole at different drainage times. It can be seen from the figure that the permeability around the borehole is greater than the initial permeability. The further away from the borehole boundary, the smaller the permeability, and the closer to the initial permeability. At the same time, it can also be seen that with the increase of drainage time, the range of permeability around the borehole gradually increases. The monitoring data of monitoring line and monitoring points form curves, as shown in Figure 12. It can be seen from Figure 12a that the permeability of the borehole boundary increases by 1.2 times. With the increase of drainage time, the coal permeability also increases, but the permeability at the borehole boundary remains unchanged. It can be seen from Figure 12b that no matter how far the distance from the borehole is, with the increase of drainage time, the permeability gradually increases from the initial permeability, but the increase in the amplitude of permeability decreases with the increase of time. The closer to the borehole, the faster the permeability increases with time in the early stages.

4.1.4. Variation Law of Gas Pressure

Based on the theoretical formula, the numerical simulation software is used to calculate the distribution characteristics of gas pressure around the borehole when the drainage time is 50 d, 100 d, 150 d, 200 d, 250 d and 300 d, respectively, so as to observe the variation law of gas pressure around the borehole and to plot the curves from the monitoring data of monitoring line and monitoring points, as shown in the figure below.

Figure 13 and Figure 14 show the cloud diagram of gas pressure distribution around the borehole at different drainage times. It can be seen from the figure that the gas pressure increases gradually from the borehole boundary to the depth of the coal seam. With the increase of drainage time, the range of gas pressure decrease gradually increases. The monitoring data of monitoring line and monitoring points are plotted into curves, as shown in Figure 15. It can be seen from the figure that with the increase of drainage time, the gas pressure decreases more. The closer to the borehole, the greater the gas pressure drop rate in the early stage of drainage.

4.2. Multi-Borehole Drainage

4.2.1. Geometric Model Setting

Based on the single borehole drainage model, a multi-borehole drainage model has been established, as shown in Figure 16. Except for increasing the number of boreholes, other calculation parameters of the model have not changed. Because it is estimated that the drainage time of boreholes is more than 200 d, the gas drainage effect of boreholes with different spacing is studied based on the drainage time of 200 d. The spacing of boreholes is 4 m, 6 m, 8 m, 10 m, 12 m and 14 m, respectively. In the process of borehole drainage, the whole coal seam in the area covered by the borehole should be drained to meet the standard. Therefore, a 60 m long monitoring line is arranged at the top of the coal seam, that is, in the middle of the upper boundary of the model, to monitor the simulation results. At the same time, a monitoring point is arranged at the midpoint of the two boreholes, to monitor the simulation results in the middle of the boreholes.

4.2.2. Mesh Independence Test of Model

In order to study the influence of model mesh on simulation results, it is necessary to test the mesh independence. The mesh is divided into normal, fine and finer, as shown in Figure 17. According to the theoretical formula, the permeability and gas pressure distribution law of the simulation borehole spacing of 6 m and the drainage of 20 d are shown in Figure 18.

It can be seen from the figures that the mesh type varies from normal to finer, and there is no obvious change in the permeability and gas pressure distribution around the borehole. In order to more clearly analyze the influence of mesh type on simulation results, the data measured by the monitoring line are plotted into a curve, as shown in Figure 19.

It can be seen from the figures that the simulation results of three different mesh types are different between boreholes under the multi-borehole drainage. The mesh type is from normal to finer, the permeability between boreholes decreases, and the gas pressure increases. The maximum error is 0.093% from normal to fine mesh types, and 0.049% from fine to finer mesh types. The error of simulation results between different mesh types is small. Since it is necessary to determine the borehole spacing when simulating the multi-borehole extraction, in order to ensure the extraction effect and consider the error, the finer mesh type is selected for simulation.

4.2.3. Permeability Variation Law

According to the theoretical formula, the numerical simulation software is used to calculate the distribution characteristics of permeability around the borehole under different borehole spacing when the drainage time is 200 d, so as to observe the variation law of permeability around the borehole., and to draw the curves from the monitoring data of monitoring line and monitoring point, as shown in the figures below.

Figure 20 and Figure 21 shows the distribution of permeability around differently spaced boreholes. It can be seen from the figure that with the increase of borehole spacing, the range of permeability gradually increases, but the permeability between boreholes decreases with the increase of borehole spacing. The data monitored by the monitoring line and the monitoring point are drawn into curves, as shown in Figure 22. It can be seen from Figure 22a that with the increase of borehole spacing, the permeability between boreholes gradually changes from greater than at the top of the boreholes to less than at the top of the boreholes, indicating that the drainage boreholes will interact with each other, resulting in the increase of permeability between boreholes; however, with the increase of borehole spacing, the degree of interaction between boreholes gradually decreases. It can be seen from Figure 22b that when the drainage time is 200 d, and the spacing between boreholes is 6 m, 10 m and 14 m, that is, the distance between measuring point and boreholes is 3 m, 5 m and 7 m, the permeability ratios are 1.112, 1.100 and 1.088, respectively, and the permeability gradually decreases. Compared with 1.067, 1.056 and 1.046 at the same time and location of single borehole drainage, the increase is 4.22%, 4.17% and 4.02%, respectively, which can be concluded as above.

4.2.4. Gas Pressure Variation Law

According to the theoretical formula, the distribution characteristics of gas pressure around the boreholes are calculated respectively by using numerical simulation software under different borehole spacing when drainage time is 200 d. The gas pressure variation law around the boreholes is observed, and the data monitored by the monitoring line and the monitoring point are plotted into curves, as shown in the following figure.

Figure 23 and Figure 24 show the distribution of gas pressure around different boreholes according to their spacing. It can be seen from the figure that with the increase of borehole spacing, the gas pressure reduction range gradually increases, but the gas pressure between boreholes increases with the increase of borehole spacing. The data monitored by the monitoring line and the monitoring point are drawn into curves, as shown in Figure 25. It can be seen from Figure 25a that with the increase of borehole spacing, the gas pressure between boreholes gradually changes from less than the top of boreholes to more than at the top of boreholes, indicating that the drainage boreholes will interact with each other, resulting in the decrease of gas pressure between boreholes; with the increase of borehole spacing, the degree of interaction between boreholes gradually decreases. It can be seen from Figure 25b that when the drainage time is 200 d, and the distance between boreholes is 6 m, 10 m and 14 m, that is, the distance between measuring point and boreholes is 3 m, 5 m and 7 m, the gas pressure is 0.133 MPa, 0.154 MPa and 0.175 MPa respectively, and the gas pressure increases gradually. Compared with 0.213 MPa, 0.233 MPa and 0.253 MPa at the same time and position of single borehole drainage, the values decreased by 36.49%, 35.32% and 31.10%, respectively, similar to the findings above.

For a high gas coal seam with gas pressure lower than 0.74 MPa, the effective range of borehole drainage is often defined by relative pressure, that is, the boundary of the effective drainage range is 51% decrease in gas pressure [28]. Therefore, the effective drainage boundary of gas pressure is 0.172 MPa. It can be seen from Figure 25, when the spacing between boreholes is 14 m, the gas pressure between boreholes will be greater than 0.172 MPa, that is, the drainage between boreholes will not meet the standard. In order to more clearly analyze the effective drainage range between boreholes, the range of gas pressure lower than 0.172 MPa at different borehole spacing is drawn, as shown in Figure 26. It can be seen from the figure that with the increase of borehole spacing, the range of gas pressure lower than 0.172 MPa also gradually increases (the red part in the figure). However, when the borehole spacing is 14 m, there is a non-standard drainage zone between boreholes.

5. Field Application

5.1. Method Statement

According to the above numerical simulation results, the final borehole spacing is determined to be 12 m. In order to control the coal seam within 15 m on both sides of the pre-excavation roadway, four boreholes need to be arranged. According to the length of roadway, four drilling fields with the size of 8 m × 5 m × 5 m are designed and constructed in the 2302 auxiliary transportation roadway. The distance between drilling fields is 400 m. There are eight boreholes designed and constructed in 1# drilling field and four boreholes in 2–4# drilling fields. The borehole layout plan is shown in Figure 27. The sealing method of “two plugging, one injection and one row” is adopted, and the sealing depth of the borehole is 20 m to ensure tight sealing without air leakage. And each borehole in the drilling fields of 1–4# is equipped with a concentration measuring port and orifice flowmeter, which is convenient for real-time monitoring of gas concentration and gas drainage flow.

The drilling equipment used in the construction is shown in Figure 28, mainly including operation console, guidance system console, water pump, directional drill, rotary unit, front gripper, crawler, etc. The operation console controls the hydraulic system and water system of the drilling rig. The guidance system console displays the collected data of the unit in the borehole on the screen. The water pump is used to provide water for the downhole motor. A directional drill is used to control drilling direction. The rotary unit is used to rotate the drill pipe. The front gripper is used to push and dismantle the drill pipe, and also to guide the drill pipe. The crawler can enable the drilling rig to walk freely in the underground roadway.

5.2. Effect Analysis

After the completion of drilling construction, record the gas drainage flow and gas concentration after different drainage times. After the completion of drainage, test the effects on the coal seam within the scope of the pre-excavation roadway. At the same time, record the excavation speed of the heading face during the excavation, and analyze the drainage effect.

Figure 29 shows the data record during borehole drainage and 2302 return air roadway excavation. Figure 29a,b show the gas concentration and gas drainage flow of a single borehole in four drilling fields in the first 50 d. It can be seen from the figure that the gas concentration and gas drainage flow change little in the first 35 d, with gas concentration higher than 70% and gas drainage flow higher than 0.38 m³/min. The gas concentration is still higher than 65% and the gas drainage flow is still higher than 0.33 m³/min after 50 d of drainage, with little decrease and good drainage effect. Figure 29c shows the measurement of the drilling cutting quantity in the coal seam before the roadway excavation (after each excavation distance of the roadway, it is necessary to predict the outburst in front of the heading face, and the drilling cutting quantity is one of the important indicators). It can be seen from the figure that the drilling cutting quantity is 3.1–4.7 kg/m, which is lower than the critical value of outburst of 6.0 kg/m, and there is no danger of outburst, i.e., there is no gas dynamic phenomenon in the process of roadway excavation. Figure 29d shows the excavation length per month. It can be seen from the figure that the excavation length of the long-distance gas pre-drainage method is more than 80 m per month; compared with the in-seam ordinary borehole gas drainage technology (Figure 1b), the excavation length can be more than 50 m per month, and the excavation speed is greatly increased.

6. Discussion

This paper introduces and studies the technology of long-distance advance pre-drainage of gas in the heading face, and has achieved good application results in the field. This technology can enable drilling of boreholes to control the gas in the area of the roadway to be excavated by use of the roadway that has been excavated. This technology is applicable to the following situations: (1) The mine is a mine with high gas or coal and gas outburst. (2) The mine needs long-term gas drainage to solve the gas problem. (3) There is a shortage of mining and replacement, which affects the efficient production of mine. This technology can save a large amount of cost compared with the technology of driving bottom drainage, and can save a large amount of time compared with the technology of digging while performing drainage, which has great technical advantages.

7. Conclusions

In view of the shortcomings of the existing commonly used gas drainage methods in the heading face, this paper proposes to adopt the long-distance gas pre-drainage technology in the heading face by directional long borehole. The main conclusions are as follows:

• Taking the return air roadway of 2302 working face in Changcun coal mine as the research object, a gas migration model considering the change of permeability with drainage time was established. The results show that the permeability decreases gradually from the borehole boundary to the coal depth in the same drainage time. Whether with single borehole or multi-borehole drainage, the range of permeability increases gradually with the increase of drainage time. In the process of multi-borehole drainage, the permeability between boreholes will increase due to the interaction between boreholes. The maximum permeability does not change with drainage time and is 1.2 times the initial permeability.
• With the increase of drainage time, the gas pressure around the boreholes gradually decreases. When multi-boreholes are used for drainage, due to the mutual influence between boreholes, the decrease in the range of gas pressure between boreholes increases. Compared with single borehole drainage, the effective drainage radius of boreholes increases, but with the increase of borehole spacing, the degree of mutual influence between boreholes decreases. In the field construction, a reasonable borehole spacing should be selected.
• According to the field conditions and research results, a reasonable gas drainage scheme was designed and implemented. The field application shows that the concentration and purity of gas drainage can maintain a high level for a long time. The drilling cutting quantity index is always lower than the critical value of outburst, and there is no gas dynamic phenomenon in the excavation process. Compared with the in-seam ordinary borehole gas drainage technology, the excavation length of the roadway is more than 50m per month, which greatly increases the excavation speed of the roadway. This technology provides a new way of thinking for solving the problem of mining replacement shortage in high gas or coal and gas outburst mines that need long-time extraction.