*Article* **New Technology of Pressure Relief Control in Soft Coal Roadways with Deep, Violent Mining and Large Deformation: A Key Study**

**Shengrong Xie 1,2, Hui Li 1, Dongdong Chen 1,\*, Shaohua Feng 1, Xiang Ma 1, Zaisheng Jiang <sup>1</sup> and Junqi Cui <sup>1</sup>**


**Abstract:** Previous studies have shown that the influence of deep dynamic pressure on the surrounding rock control of a coal roadway is one of the difficulties in mine roadway support. Based on the investigation of the headgate 11231 in a coal mine, this study analyzes the damage characteristics of coal roadway surrounding rock affected by deep dynamic pressure, expounds on the difficulties of controlling the roadway surrounding rock, and creatively proposes a cooperative control technology of external anchor–internal unloading for regulating large deformation of roadways. The vertical stress distribution and transfer law of surrounding rock with different hole-making depths, spacing, and lengths after roadway excavation were simulated and studied, and an appropriate parameter range of hole-making space in the stage without dynamic pressure influence was obtained. Considering the influence of mining dynamic pressure, the surrounding rock pressure relief effect of each optimized hole-making parameter was analyzed. In addition, the optimal hole-making parameters (the hole-making depth, spacing, and length were 8 m, 3.2 m, and 3 m, respectively) that can effectively reduce the high stress of roadway shallow surrounding rock in two stages (without and with dynamic pressure) and ensure integrity of the shallow surrounding rock were obtained. The actual field application shows that the new technology can reduce the higher rib deformation by approximately 850 mm and achieve a good surrounding rock control effect. The research and practice show that the pressure relief control for soft coal roadways with deep, violent mining and large deformation has achieved success, providing technical support for the maintenance of the same type of roadway.

**Keywords:** deep mine; coal roadway; numerical simulation; pressure relief; external anchor–internal unloading; surrounding rock control technology

#### **1. Introduction**

With the deep mining of coal resources gradually becoming the norm [1–3], the engineering response problems of discontinuous, uncoordinated large deformation and large-scale instability of roadway surrounding rock caused by the typical deep "three high" occurrence environment and coal mining dynamic pressure have become urgent engineering problems that need to be solved [4,5]. Research shows that the complex stress field [6], deformation brittle–ductile transition [7], continuous deterioration [4,8], and strong rheology of deep coal and rock mass are the major reasons for the continuous large deformation of deep roadway surrounding rock. In China, coal roadways excavated by underground coal mines account for approximately 80% of the total roadway excavation volume [9]. Coal science and technology workers have conducted a lot of research on the

**Citation:** Xie, S.; Li, H.; Chen, D.; Feng, S.; Ma, X.; Jiang, Z.; Cui, J. New Technology of Pressure Relief Control in Soft Coal Roadways with Deep, Violent Mining and Large Deformation: A Key Study. *Energies* **2022**, *15*, 9208. https://doi.org/ 10.3390/en15239208

Academic Editor: Adam Smoli ´nski

Received: 29 October 2022 Accepted: 21 November 2022 Published: 5 December 2022

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**Copyright:** © 2022 by the authors. Licensee MDPI, Basel, Switzerland. This article is an open access article distributed under the terms and conditions of the Creative Commons Attribution (CC BY) license (https:// creativecommons.org/licenses/by/ 4.0/).

surrounding rock properties [10,11], stress environment [12,13], and roadway deformation characteristics [14,15] of deep coal roadways through field investigation, theoretical analysis, experimental research, and numerical simulation.

In recent years, a lot of work has been carried out to study the impact of mining on the rock mass and to recommend measures for reducing strain changes of rock mass and minimizing the stress on the surface. Since mining production has a significant effect on the stress-strain behavior of rock mass, the issues of reducing this influence are very relevant and scientists around the world are trying to minimize it. Adigamov, A.E., et al. [16] established the stress-strain behavior model of disturbed rock mass with regard to anisotropy and discontinuities, which can be used to calculate the strength of underground rock mass; Khayrutdinov, A.M., etc., [17] studied the change of stress-strain characteristics of rock mass after using different-strength backfill and carried out research on the stress-strain relationship of disturbed rock mass under different conditions. On the basis of studying the influence of mining activities on the stress-strain relationship of rock mass [18,19], scholars at home and abroad have developed many targeted coal roadway surrounding rock support technologies. High prestressed strong bolt support [20], high strength, high stiffness and high prestressed bolt support [21], and butted long bolt support [22] technologies can effectively control the generation and development of separation, sliding, and cracks in shallow surrounding rock as well as improve the post-peak strength of deformed surrounding rock and its mechanical parameters. Support technologies, such as the constant resistance and large deformation bolt support technology [23] and high-strength pressure relief anchor box beam support system [24] can release the uncontrollable deformation energy in the surrounding rock after the roadway support is completed, and then, give full play to the support capacity of the bolt, so that the bolt and the pressure relief surrounding rock can form a stable support bearing body. The anchor cable truss support technology [25] can simultaneously provide extrusion stresses in the horizontal and vertical directions of the roadway roof and rib, thus effectively reducing the maximum tensile stress and maximum shear stress of the surrounding rock in the anchorage zone. The anchor-grouting combined support technology [26] can re-condense the broken surrounding rock of the coal roadway, effectively improve the mechanical properties of the surrounding rock, and ensure the integrity and safety of the roadway surrounding rock. In addition, the active–passive coupling support technology [27] involves combining bolts (cables) with the passive support components to maintain stability of the coal roadway surrounding rock under special geological conditions and fully utilize the initial support resistance from the active support and the high deformation resistance of the passive support. To control the deformation of deep coal roadway surrounding rock, while seeking innovation of the support technology, it has become an important research direction for the surrounding rock control of deep coal roadways to realize the release or transfer of high concentrated stress in the coal roadway surrounding rock through stress control (surrounding rock pressure relief) to achieve stability of the surrounding rock. Currently, many roadway pressure relief control methods are available, such as arranging the roadway in the stress reduction zone [28], physical pressure relief of surrounding rock (e.g., by drilling [29] and blasting [30,31]), excavating the pressure relief roadway [32], and roof cutting pressure relief [33]. The pressure relief control technology can solve the problem of controlling the surrounding rock under various working conditions to a certain extent.

Analyzing the above, it can be noted that the surrounding rock control in soft coal roadways with deep, violent mining and large deformation is a very topical issue, but the existing research results have not made a big breakthrough in the surrounding rock control of this type of roadway. Given the above problems, this study creatively proposes a cooperative control technology of external anchor–internal unloading of surrounding rock in deep coal roadways. First, the roadway shallow surrounding rock is strengthened through an anchor-grouting combined support. Then, a hydraulic hole-making machine is used to create pressure relief space with appropriate spacing in a certain range of the deep coal roadway to transfer the high concentrated stress from the shallow to the deep rock and effectively improve the stress environment of the roadway surrounding rock. The drawings in Figure 1 show how by combining the characteristics of new and old pressure relief technologies, both conventional pressure relief drilling [29] and hole-making pressure relief can effectively realize the transfer of high concentrated stress in the shallow surrounding rock to the deep rock. However, the conventional pressure relief dense drilling makes the shallow surrounding rock more broken, reducing its strength. In contrast, the hole-making pressure relief technology uses steel pipes to effectively solidify the borehole surrounding rock and ensures the integrity and effectiveness of the shallow surrounding rock and support structure. The new cooperative control technology was first applied to the deep dynamic pressure mining roadway and achieved success, ensuring the roadway rib did not expand during the service period. This has important research value for the surrounding rock pressure relief support of a high stress and large deformation coal roadway.

**Figure 1.** Comparison between the internal hole-making pressure relief technology and conventional pressure relief technology of surrounding rock.

#### **2. Project Overview**

#### *2.1. Engineering Geology and Problems*

The headgate 11231 of a coal mine is located in the 1100 southern mining area. In the south of the headgate is solid coal panel 11231, in the northern part is fault SF86, in the eastern part is the −760 m horizontal main roadway, and in the western part is the concentrated roadway of the 1100 southern mining area. The average buried depth of the headgate is 740 m, and the roadway size is 5.0 × 3.5 m (width × height), with an anchor mesh cable combined support. The average thickness of the No. 2 coal seam is 4.5 m, and the average dip angle is 14◦. Panel 11231 is a gangue backfilling panel, and the compressed ratio of the gangue after compaction is approximately 80%. A drawing of the location and columnar of panel 11231 is shown in Figure 2.

In the process of advancing each panel in the 1100 southern mining area, although the goaf is filled with gangue, the surrounding rock of the section roadway within 90–120 m in front of the panel still has large deformation owing to the mining dynamic pressure. As the mining progresses, the surrounding rock is continuously renovated by a team to ensure normal mining, as illustrated in Figure 3. According to statistics, during the service period of the 1100 southern mining area, the section roadways need to be renovated at least once, and some sections even require two to three renovations. The high frequency of renovation work has greatly increased the cost of roadway support and the labor intensity of workers. Moreover, it has seriously restricted the safe and efficient production of the mine. To solve the engineering problem caused by the repeated roadway renovation and improve the production efficiency of the mine, this study selects a reasonable position of headgate 11231 in a coal mine as a test roadway for applying the new support technology. The position of the test roadway is displayed in Figure 2.

**Figure 2.** Location of panel 11231 and the columnar section of the coal and rock strata. (**a**) Location of panel 11231; (**b**) Columnar section of the coal and rock strata.

**Figure 3.** Renovation site of headgate 11231. (**a**) Deformation and crack diagram of roadway surrounding rock; (**b**) Roadway rib renovation site; (**c**) Repair depth of roadway rib; (**d**) The roadway rib has been renovated.

As indicated in the curve chart in Figure 4, the accumulated approach of the two ribs is approximately 610 mm (no mining influence stage) between the completion of headgate 11231 and 9 months before the mining dynamic pressure. The influence range of the increased dynamic pressure of panel 11231 is approximately 90 m. Because of the influence of mining dynamic pressure on the panel, the deformation of the surrounding rock increases sharply. The data show that within the period (approximately 21 days) from the influence of dynamic pressure on the roadway surrounding rock to 30 m from the panel, the accumulated approach of the two ribs of the roadway surged to 1240 mm (mining influence stage). Then, the roadway was renovated, and the displacement monitoring was performed again. The accumulated approach of the two ribs of the renovated roadway to the completion of mining of the panel was approximately 520 mm (mining influence stage). Therefore, the total approach of the two ribs of the roadway is approximately 1760 mm. The continuous large-scale extrusion of the two ribs of headgate 11231 (Figure 4(II)-a–c) led to damage of the support structure (Figure 4(I)-a–c). The reduction in support strength of the roadway rib coal aggravated the continuous large deformation of the coal body, forming a vicious cycle of large deformation of the surrounding rock and damage of the support body, which posed a great threat to the safety of underground personnel and equipment.

**Figure 4.** Deformation overview of the two ribs of headgate 11231 surrounding rock. (**I-a**) Tearing of metal mesh; (**I-b**) Large deformation of roadway rib; (**I-c**) Anchor bolt failure; (**II-a**) Large deformation of the panel rib; (**II-b**) Large deformation of non-panel rib; (**II-c**) Roadway rib bulges out.

#### *2.2. Analysis of Control Difficulties of Roadway Surrounding Rock*

Through investigation and analysis of the mining and production geological conditions of panel 11231, the reasons for the large deformation of headgate 11231 are as follows:


(3) The coal body is soft and broken. The strength of the coal body in a coal mine is low, owing to the loose and soft properties of the coal body. The borehole peep (Figure 5c-I) shows that the coal body is broken when the roadway rib depth is 3 m. It can be seen from Figure 5c-II that when the roadway rib has no large deformation, the soft coal body in the roadway rib peels off under an effective support. The low strength and weak self-supporting capacity of the coal body is the main reason for the large deformation of headgate 11231.

**Figure 5.** Schematic for the analysis of surrounding rock failure difficulties of headgate 11231. (**a**) Roadway deformation caused by high field stress; (**b**) Roadway deformation caused by complex tectonic stress; (**c-I**) Borehole peeping; (**c-II**) Crushed coal; (**d**) Large deformation of roadway rib; (**e**) Large mining height.


#### **3. Principle and Key Technical Parameters of the Cooperative Control Technology of External Anchor–Internal Unloading of Coal Roadway Surrounding Rock**

According to the above analysis, to solve the problem of repeated renovation of mining roadways caused by their large deformation, technical research should be conducted from the following two perspectives: (1) complex stress environment in coal roadways (difficulties 1, 2, and 5) and (2) soft and broken characteristics of surrounding rock (difficulty 3). Based on this, the cooperative control technology of external anchor–internal unloading of coal roadway surrounding rock is proposed in this study.

#### *3.1. Principle of Cooperative Control Technology of External Anchor–Internal Unloading of Coal Roadway Surrounding Rock*

The cooperative control technology of external anchor–internal unloading refers to strengthening the shallow surrounding rock of the roadway through the combined strategy of anchor-grouting and then using physical means to make pressure relief holes with reasonable spacing in a certain range of the deep part of the roadway to improve the stress environment of the roadway surrounding rock. The pressure relief holes can also provide a large compensation space for the deep coal body to transfer to the roadway space and effectively block the intermediate source of large deformation of roadway surrounding rock. In addition, the strengthening of the surrounding rock in the shallow part of the roadway can restrict the shallow coal body from moving to the pressure relief space. Generally, the internal pressure relief method involves making large-diameter holes with reasonable spacing in a certain range in the deep part of the roadway by using hydraulic hole-making equipment. The method mainly includes determining reasonable technical parameters using geological steel pipes with an appropriate diameter to effectively support the ordinary boreholes in the shallow part of the roadway and creating large-diameter holes in the deep part of the roadway.

Figure 6 illustrates the principle of cooperative control technology of external anchor– internal unloading. This technology mainly includes two aspects, namely "external anchor" and "internal unloading". As mentioned above, the external anchor is a high-efficiency and strong pre-tightening support structure of anchor cable truss beams formed by strong anchor cables and a channel steel or steel belt beam. The shallow coal roadway is grouted to form an anchor-grouting reinforced bearing body of shallow surrounding rock of the roadway (Figure 6A), which creates a good surrounding rock environment for internal hole-making and pressure relief. Internal unloading mainly includes two objectives. (1) The pressure relief space in the deep part of the roadway makes the peak area of the abutment stress caused by the roadway rib substantially transfer to the deep part (Figure 6B,C), which reduces continuous damage (due to high concentrated stress) to the shallow coal and rock mass and improves its stress environment. (2) Continuous large-diameter holes provide a large compensation space for the transfer of deep coal body to the roadway space and effectively block the intermediate source of large deformation of roadway surrounding rock (Figure 6D). It should be pointed out that the structural integrity of the surrounding rock in the shallow anchorage zone should not be damaged during internal hole-making and pressure relief, and the pipe-fixing method can be used to effectively support the shallow ordinary drilling area to ensure that the strength of the coal and rock mass in this area is not reduced by the hole-making.

**Figure 6.** Principle of cooperative control technology of external anchor–internal unloading of roadway surrounding rock.

#### *3.2. Key Technical Parameters of Internal Hole-Making and Pressure Relief*

Effective implementation of the internal hole-making and pressure relief technology mainly involves the reasonable selection of three technical parameters, as shown in Figure 7.

**Figure 7.** Schematic for the layout of hole-making in the two ribs of the roadway.


rock loose areas. Appropriate spacing of hole-makings can connect the loose areas of two adjacent holes-making along the axial direction of the roadway, and the loose areas of continuous hole-making connect with each other in the axial direction of the roadway to form a pressure relief continuous zone (Figure 6E). When the spacing is extremely large, the pressure relief continuous zone cannot be formed. In addition, the coal body between the hole-making spaces is still in a state of high concentrated stress, and the surrounding rock pressure relief effect is poor. In contrast, when the spacing is extremely small, dense ordinary boreholes cause great damage to the shallow surrounding rock, which affects its integrity, and at the same time increases the labor of workers and reduces construction efficiency.

(3) Hole-making length *L*3. The hole-making length has a significant influence in terms of two aspects: (1) the greater the hole quantity and length, the greater the coal output and the more obvious the compensation effect of the hole-making space on deep coal; (2) pressure relief amplitude, namely, the transfer distance of highly concentrated stress. In general, when the hole-making depth is appropriate, the greater the holemaking length, the greater the transfer distance of the peak value of the roadway rib abutment stress to the deep rock. When the hole-making length is extremely small, the pressure relief range is small, the pressure relief effect is not sufficient, and the pressure relief hole-making is closed in a short time, preventing its blocking effect on the transfer of deep coal body to the roadway space. On the other hand, when the hole-making length is extremely large, the deeper hole-making space has little effect on the pressure relief of the shallow surrounding rock of the roadway.

Owing to the large dip angle of the coal seam and limited by the technical capability of the hydraulic hole-making equipment, the coal output effect of the internal hole-making space in the lower rib of the headgate 11231 test section is poor and cannot meet the technical requirements. Meanwhile, comparing the displacement monitoring data of headgate 11231, the higher rib deformation of the headgate accounts for more than 65% of the total displacement of the two ribs of the roadway. Therefore, effectively limiting the deformation of the higher rib surrounding rock of the headgate can greatly relieve the renovation pressure of headgate 11231 and meet the engineering requirements of the panel.

#### **4. Study on Key Technical Parameters of Pressure Relief by Internal Hole-Making**

*4.1. Establishment of Numerical Model and Research Ideas*

To determine the key technical parameters of pressure relief by internal hole-making in deep coal roadways affected by dynamic pressure, a numerical simulation study was carried out according to the actual situation on-site and the technical principle of pressure relief by internal hole-making combined with the FLAC3D finite element software. The numerical model is depicted in Figure 8. The numerical model size is 130 × 70 × 90 m. The length of panel 11231 is 59 m. The mining height is 4.5 m, and the dip angle of the coal and rock strata is 14◦. Considering that panel 11231 is a gangue backfilling panel, the final compressed ratio is approximately 80%. It is considered that the equivalent mining height of the panel is 0.9 m. The mining roadways are arranged along the coal seam roof. The roadway width is 5 m, the center height of the roadway is 3.5 m, the distance from the borehole to the roadway floor is 1.5 m, and the upward angle of the hole-making in different schemes ranges from 3.0 to 8.5◦. Each coal and rock stratum in the numerical model adopts the Mohr-Coulomb constitutive model. The left and right boundaries of the model are fixed with horizontal displacement along the x-direction, the front and rear boundaries of the model are fixed with horizontal displacement along the y-direction, and the bottom boundary of the model is fixed with z-displacement along the vertical direction. Moreover, a load of 17.115 MPa is applied to the top boundary of the model to simulate the overburden weight. The coefficient of lateral pressure of the model is 1.2. Based on the basic mechanical parameters of coal rock mass measured in the laboratory and data from the literature, the mechanical parameters were calculated [34–36]. Table 1 presents the physical and mechanical parameters of each coal and rock stratum.

**Figure 8.** Numerical simulation model.

**Table 1.** Mechanical parameters of coal and rock strata.


Given the key technical parameters (*L*1, *L*2, and *L*3) of pressure relief by internal holemaking of headgate 11231 in a coal mine, a variety of numerical simulation schemes were set up by using a control variable method. The goals were to study the vertical stress distribution law of the surrounding rock under different hole-making parameters and determine the final hole-making parameters of the roadway. As presented in Table 2, there are 15 numerical simulation schemes in total (the technical parameters of Schemes 3, 8, and 13 are the same).


**Table 2.** Numerical simulation schemes.

According to the field investigation, the surrounding rock deformation of headgate 11231 can be divided into two stages: the no mining influence stage and the mining influence stage. Based on the displacement monitoring data in Figure 4, the displacement of the two ribs and deformation of the higher rib, respectively, account for 34.1 and 33.9% of the total deformation in the no mining influence stage. That is, the deformation of the two ribs of the coal roadway in the no mining influence stage accounts for approximately 1/3 of the total deformation, and the roadway first goes through the no mining influence stage after the support is completed. In this stage, the surrounding rock of the deep coal roadway is mainly affected by the high abutment stress of the roadway rib formed after roadway excavation. Therefore, it is necessary to study the pressure relief law of the coal roadway surrounding rock under different hole-making parameters in the no mining influence stage after roadway excavation to determine the appropriate hole-making parameters and reduce the deformation of the roadway rib at this stage. The abutment pressure caused by coal mining has changed the original stress distribution state of the roadway surrounding rock. The superposition of the advanced abutment pressure and the original high abutment stress of the roadway rib makes the stress of the roadway rib (panel rib) complex and changeable. Based on the key technical parameters of pressure relief by hole-making determined in the stage without mining influence, the parameters of pressure relief by hole-making are continuously optimized so that it can effectively relieve the surrounding rock pressure and reduce the influence of the superimposed high stress on the roadway rib on the surrounding rock in the mining influence stage.

#### *4.2. Key Technical Parameters of Pressure Relief by Internal Hole-Making in the No Mining Influence Stage*

(1) Hole-making depth

The vertical stress distributions in the higher rib of the roadway with different holemaking depths are illustrated in Figure 9b–f. When the hole-making depths are 4, 6, and 8 m, the peak stress of the higher rib of the roadway is located in the deeper part of the hole-making space. When the hole-making depth increases from 8 m to 10 m and 12 m, a new high stress peak zone is formed between the roadway space and the hole-making space, which is not conducive to the stability of the roadway surrounding rock. The stress monitoring lines are arranged at the higher rib of the roadway along the direction of holemaking, as displayed in Figure 9(A1-A1) to (F1-F1). According to the stress distribution value of the higher rib of the roadway under the conditions of no hole-making and different hole-making depths, a comparison diagram of the stress curve of the roadway rib is drawn, as indicated in Figure 10.

**Figure 9.** Layout scheme under different hole-making depths and distribution nephogram of vertical stress.

To effectively evaluate the pressure relief effect of different hole-making depths on the surrounding rock, two evaluation indices were selected according to the stress transfer law. First is the transfer amplitude of the high concentrated stress, which is the transfer distance from the original stress peak position to the deep part of the roadway. Second is the reduction in the high concentrated stress in the shallow surrounding rock of the roadway, that is, the pressure relief effect on the surrounding rock in the high abutment stress zone of the roadway rib after roadway excavation. The pressure relief effect from high to low is categorized as excellent, good, fair, no, and poor. As shown in Figure 10a, when the hole-making depth is 4 m, the transfer distance from the original stress peak position to the deep part is 0.5 m. The pressure relief space only reduces the stress in the low stress zone of the shallow roadway, which cannot achieve the purpose of inward movement of the original high concentrated stress of the roadway rib, and there is almost no pressure relief effect. As depicted in Figure 10b, when the hole-making depth is 6 m, the transfer distance is 1.5 m. The pressure relief space makes the stress in the original stress peak zone of the roadway rib decrease significantly: the stress at the peak position decreases by 7.6 MPa, and the high stress zone transfers to the deep part. Thus, the pressure relief effect is evident. As illustrated in Figure 10c, when the hole-making depth is 8 m, the transfer distance is 3.0 m. The pressure relief space makes the stress in the original stress peak zone of the roadway rib decrease overall: the stress at the peak position decreases by 17.3 MPa, and the high stress zone transfers to the deep part considerably. Therefore, the pressure relief effect is remarkable. As displayed in Figure 10d,e, when the hole-making depths are 10 and 12 m, the stress distributions on the original stress peak zone and shallower surrounding rock are not changed significantly after pressure relief. Furthermore, the original stress peak position is not moved inward effectively, and even the shallow abutment stress increases locally, aggravating the damage on the surrounding rock.

**Figure 10.** Comparison diagram of the roadway rib stress curve at different hole-making depths. (**a**) Hole−making depth = 4 m; (**b**) Hole−making depth = 6 m; (**c**) Hole−making depth = 8 m; (**d**) Hole−making depth = 10 m; (**e**) Hole−making depth = 12 m.

#### (2) Hole-making spacing

The vertical stress distributions on the higher rib of the roadway with different values of hole-making spacing are indicated in Figure 11a–e. When the hole-making spacing is 2.4 and 3.2 m, a good pressure relief zone can be formed between the hole-making spaces, so that multiple holes form a continuous pressure relief zone along the axis of the roadway. Moreover, the high concentrated stress in the shallow part of the surrounding rock can be uniformly transferred to the deep surrounding rock. Thus, the pressure relief effect is good. When the hole-making spacing is increased to 4.0 m, the transfer effect of the high concentrated stress of the coal mass between the two holes to the deep surrounding rock becomes worse, and a stress concentration zone (Zones A and B) near the inner and outer ends of the hole-making area appears. When the hole-making spacing is 4.8 m, Zones A and B are connected, and the stress of the coal mass between the two holes is restored to the original stress state, resulting in no pressure relief effect. When the hole-making spacing continues to increase to 5.6 m, the high stress moving inward to the deep part accumulates

in the coal mass between the two holes, forming a stress-increasing area, which not only has no pressure relief effect, but also causes great damage to the coal mass.

**Figure 11.** Layout scheme under different hole-making spacing and distribution nephogram of vertical stress.

To accurately evaluate the pressure relief effect of different values of hole-making spacing on the surrounding rock, the two evaluation indices proposed above were used. It can be seen from Figure 11 that the coal mass between the two holes (the position with the worst pressure relief effect) has the greatest impact on the stress change of the surrounding rock with different values of hole-making spacing. Therefore, the vertical stress monitoring lines ((F2-F2) in Figure 9 and (A2-A2) to (E2-E2) in Figure 11) were arranged at the center of the two holes and parallel to the holes in the different hole-making spacing schemes. Figure 12 presents the stress curve comparison diagram of the roadway rib under different values of hole-making spacing.

As shown in Figure 12a, when the hole-making spacing is 2.4 m, the transfer distance from the original stress peak position to the deep part is 3.0 m. After pressure relief, the stress in the original stress peak zone decreases significantly, and the stress at the peak position decreases by 13.5 MPa. Thus, the pressure relief effect is excellent. As depicted in Figure 12b, when the hole-making spacing is 3.2 m, the transfer distance is 2.5 m. After pressure relief, the stress at the original stress peak position decreases by 6.0 MPa, and the pressure relief effect is good. When the hole-making spacing is increased to 4.0 m, the stress in the peak zone of the original stress does not decrease significantly after pressure relief but increases slightly in local areas, and the pressure relief effect is generally fair. As illustrated in Figure 12d,e, when the hole-making spacing values are 4.8 m and 5.6 m, the stress in the original stress peak zone increases significantly after pressure relief. The stress at the peak position increases by 3.8 MPa and 4.5 MPa, respectively, and the transfer distances are 2.0 m and 1.0 m, respectively. Hence, the pressure relief effect is fair.

**Figure 12.** Comparison diagram of the roadway rib stress curve at different vales of hole-making spacing. (**a**) Hole−making spacing = 2.4 m; (**b**) Hole−making spacing = 3.2 m; (**c**) Hole−making spacing = 4.0 m; (**d**) Hole−making spacing = 4.8 m; (**e**) Hole−making spacing = 5.6 m.

(3) Hole-making length

The vertical stress distributions of the higher rib of the roadway with different holemaking lengths are indicated in Figure 13b–f. With an increase in the hole-making length, the transfer distance from the peak stress zone of the roadway rib to the deep part gradually increases, and the range of the stress peak zone of the roadway rib gradually decreases. When the hole-making length increases to 3 m, the range of the stress peak zone of the increased hole-making length does not change significantly.

**Figure 13.** Layout scheme under different hole-making lengths and distribution nephogram of vertical stress.

The layout of stress monitoring lines is shown in Figure 13(A3-A3) to (F3-F3), and Figure 14 presents a comparison of the roadway rib stress curves with different hole-making lengths. The two indices proposed above were used to evaluate the pressure relief effect of the surrounding rock with different hole-making lengths. In addition, to analyze the influence of the different hole-making lengths on the range of the peak stress zone, the stress of 25 MPa was defined as the stress boundary value to measure the range of the stress peak zone. Thus, the reduction coefficient *K* of the surrounding rock stress peak zone under different hole-making lengths is

$$k = \frac{s\_i}{s\_0} \times 100\% \tag{1}$$

where *S*<sup>i</sup> is the area of the stress peak zone (greater than 25 MPa) when the hole-making length is 1–5 m, *i* = 1–5; and *S*<sup>0</sup> is the area of the stress peak zone (greater than 25 MPa) without hole-making. By defining the area of the original stress peak zone (greater than 25 MPa) as 1, measuring Figure 13, and combining with Equation (1), we obtain *k*<sup>i</sup> = 1.90, 0.90, 0.56, 0.51, and 0.54 (*i* = 1–5). Analysis shows that the smaller the value of *k*, the greater the pressure relief effect.

As depicted in Figure 14a–e, when the hole-making lengths are 1, 2, 3, 4, and 5 m, the transfer distances from the original stress peak position to the deep part are 1.5, 2.0, 3.0, 3.5, and 5.0 m, respectively, and the stress reduction values at the peak position are 4.8, 14.8, 17.3, 19.31, and 20.3 MPa, respectively. From the stress curve, it can be observed that the different hole-making lengths have good pressure relief effects. To effectively evaluate the cost performance *K* of the different hole-making lengths in transferring the original stress peak point to the deep part, the ratio of the transfer distance from the original stress peak position to the deep part to the hole-making length (which can represent the construction cost, that is, the greater the hole-making length, the greater the construction cost) is calculated as follows:

$$K = \frac{I\_i}{L\_{2-i}} \times 100\% \tag{2}$$

where *l*<sup>i</sup> is the transfer distance (in m) from the original stress peak position to the deep part, *i* = 1–5; and *L*2−<sup>i</sup> is the hole-making length (in m), *i* = 1–5. Substituting the known data into Equation (2), we obtain *K*<sup>i</sup> = 1.5, 1.0, 1.0, 0.875, and 1.0, respectively. It can be

observed from the analysis that the greater the *K* value, the higher the cost performance of hole-making and pressure relief.

**Figure 14.** Comparison of the roadway rib stress curves with different hole-making lengths. (**a**) Holemaking length = 1 m; (**b**) Hole-making length = 2 m; (**c**) Hole-making length = 3 m; (**d**) Hole-making length = 4 m; (**e**) Hole-making length = 5 m.

#### (4) Research and analysis of key technical parameters

According to the above analysis, the pressure relief effect of the different hole-making schemes was evaluated with the vertical stress as the main index. The comprehensive evaluation results of each key technical parameter evaluation index on the pressure relief effect of the surrounding rock of different hole-making schemes are listed in Table 3. According to the analysis, Schemes 2 and 3 can be selected with respect to the hole-making depth, that is, the hole-making depths are 6 and 8 m, respectively. Schemes 6, 7, and 8 can be selected in terms of the hole-making spacing; thus, the hole-making spacing values are 2.4, 3.2, and 4.0 m, respectively. When the hole-making length is 1 m, the reduction coefficient of the range of the stress peak zone *k*<sup>1</sup> > 1. In addition, the transfer distance from the original stress peak position to the deep part is small, so this scheme should be excluded. For the other hole-making length schemes (Schemes 12–15), it is not appropriate

to analyze the pressure relief effect simply from the perspective of stress, as it should be analyzed in combination with other factors.


**Table 3.** Evaluation results of surrounding rock pressure relief for different hole-making schemes.

*4.3. Optimization Analysis of Key Technical Parameters of Pressure Relief by Internal Hole-Making in the Mining Influence Stage*

Panel 11231 adopts step-by-step mining with an interval of 5 m. Figure 15 shows the nephogram of abutment stress distribution and the stress monitoring curve in front and behind the panel near the rib of the headgate. By analyzing Figure 15, the following can be inferred: (1) The stress reduction zone is from the coal rib of the panel to 4 m in front of the panel. Moreover, the stress rise zone is 4 m in front of the panel and further away, and the distance between the stress peak position and the coal rib of the panel is 12 m. (2) Affected by the dynamic pressure of coal mining, the stress value in the higher rib of the headgate increases sharply. The vertical stress values of the coal body at depths ≥ 6 m are greater than the primary rock stress, forming a high concentrated stress zone. The stress peak zone of the roadway rib shifts deeper under the influence of mining, with a transfer distance of 4–5 m. (3) The advanced abutment pressure caused by coal mining and the high abutment stress of the roadway rib are superimposed on each other, which jeopardizes the stability of the roadway rib coal in the area affected by dynamic pressure. To summarize, transferring the high concentrated stress in the shallow coal body of the roadway using hole-making and reducing the continuous damage caused by the high stress to the shallow coal and rock are essential for maintaining the stability of the surrounding rock of the coal roadways affected by mining.

**Figure 15.** Distribution law of abutment stress in front of panel 11231. (**a**) 3D stress distribution of roadway rib; (**b**) Line layout; (**c-1**) 22 m in front of the panel (**c-2**) 12 m in front of the panel; (**c-3**)2m in front of the panel.

Based on the key technical parameters of hole-making determined at the stage without dynamic pressure influence and in combination with the dynamic pressure disturbance law of panel 11231, we continued to optimize the hole-making parameters and arranged stress monitoring lines at reasonable positions in the stress rise zone in front of the panel to analyze the stress distribution law of surrounding rock with different hole-making schemes under the influence of superimposed stress.

#### (1) Hole-making depth

As illustrated in Figure 16, when the hole-making depth is 6 m or 8 m, the stress decreases in the area where the stress on the higher rib of the panel increases during the dynamic pressure influence stage. Comparing Figures 15a and 16b, the following can be observed: (1) When the hole-making depth is 8 m, the range of the stress reduction zone and the pressure relief degree are larger than that when the hole-making depth is 6 m. (2) When the hole-making depth is 8 m, the pressure relief space transfers the peak zone of the superimposed stress of the roadway rib to the deeper part, and the pressure relief effect is good. (3) When the hole-making depth is 6 m, the hole-making space causes great damage to the coal body at the roadway depth of 5 m, which affects the stability of the rock mass of bolt-grouting coal in the shallow part of the roadway. Based on the above analysis, Scheme 3 is preferred with respect to the hole-making depth of headgate 11231; that is, the hole-making depth is 8 m.

**Figure 16.** Comparison of stress curves of the roadway rib at different hole-making depths in the mining influence stage. (**a**) Hole-making depth = 6 m; (**b**) Hole-making depth = 8 m.

#### (2) Hole-making spacing

Figure 17(a-1), (b-1), and (c-1) show the vertical stress nephogram of the higher rib of the roadway with different values of hole-making spacing in the advanced dynamic pressure zone. The center positions of the two holes are respectively selected to monitor their stress values. The corresponding stress curves are displayed in Figure 17(a-2), (b-2), and (c-2). Figure 17(a-2), (b-2) indicate that the pressure relief zone is still formed at the center of the two holes. When the hole-making spacing increases to 4.0 m, stress concentration between the two holes occurs, and the hole-making space threatens the stability of the roadway rib coal body, which is not conducive to the stability of the roadway rib surrounding rock in the advanced dynamic pressure zone. The hole-making spacing is extremely small, and the boreholes are dense. The dense boreholes cause great damage to the coal body in the shallow anchorage zone of the roadway, which is not conducive to the stability of the roadway surrounding rock. Through comprehensive comparison and analysis, it is determined that the optimal hole-making spacing is 3.2 m.

**Figure 17.** Stress distribution on roadway rib at different values of hole-making spacing in the mining influence stage. (**a-1**) Stress nephogram with spacing of 2.4 m; (**b-1**) Stress nephogram with spacing of 3.2 m; (**c-1**) Stress nephogram with spacing of 4.0 m; (**a-2**) Stress curve with spacing of 2.4 m; (**b-2**) Stress curve with spacing of 3.2 m; (**c-2**) Stress curve with spacing of 4.0 m.

#### (3) Hole-making length

As shown in Figure 18, with an increase in the hole-making length, the pressure relief range increases, the peak position of the superimposed stress shifts deeper into the surrounding rock, and the pressure relief effect becomes more evident. However, considering the construction capacity of hole-making machines and the use efficiency of the hole-making space, the hole-making length cannot be increased without limit. Comparing Figure 18a–d, the hole-making lengths of 3 and 4 m can meet the pressure relief requirements of the surrounding rock in the mining influence stage. As depicted in Figure 18b,c, when the hole-making lengths are 3 and 4 m, the effective pressure relief ranges of the surrounding rock are 4.5 m and 5.0 m, respectively. As the hole-making length increases by

1 m, the effective pressure relief range does not increase significantly. In combination with equation (2), *K*<sup>3</sup> = 1.0 > *K*<sup>4</sup> = 0.875. Comparing the pressure relief effect and construction quantities of the two schemes, the hole-making length of 3 m can better meet the pressure relief requirements, reduce the construction quantities, and obtain the optimal solution of hole-making pressure relief and efficient construction.

**Figure 18.** Comparison of stress curves of roadway rib at different hole-making lengths in the mining influence stage. (**a**) Hole-making length = 2 m; (**b**) Hole-making length = 3 m; (**c**) Hole-making length = 4 m; (**d**) Hole-making length = 5 m.

#### (4) Determination of key technical parameters

According to the stress distribution law of the surrounding rock with different holemaking schemes in the advanced dynamic pressure zone of panel 11231 and the comprehensive analysis of factors such as construction benefit, cost, and construction period, the key parameters for pressure relief of higher rib by hole-making in headgate 11231 were determined as follows: the hole-making depth, spacing, and length are 8 m, 3.2 m, and 3 m, respectively.

#### **5. Cooperative Control Technology of External Anchor–Internal Unloading of Surrounding Rock in Deep Coal Roadways**

*5.1. Technical Parameters of External Anchor–Internal Unloading of Surrounding Rock in the Test Section of Headgate 11231*

The cooperative control technology of external anchor–internal unloading of surrounding rock in the test section of headgate 11231 was implemented in two stages, namely, strengthening of the shallow surrounding rock and pressure relief by hole-making in the deep part. As illustrated in Figure 19I-a–c, reinforcement of the shallow surrounding rock is first carried out for the test section of the headgate, and three steel strand anchor cables with dimensions of ϕ21.8 × 10,500 mm and a row spacing of 2.0 × 1.6 m are added to the roof. Moreover, three steel strand anchor cables with dimensions of ϕ21.8 × 4500 mm are added to the two ribs. The row spacing between the higher ribs is 1.3 × 1.6 m, whereas that between the lower ribs is 1.1 × 1.6 m. Each anchor cable uses three resin anchor agents

of model Z2360 for each hole. The roof and rib anchor cables are connected by H-type double steel belt beams supporting large square pallets (400 × 400 mm) and small pallets (200 × 200 mm) to form the roof and rib anchor cable truss beam structure, which can resist the overall outward heave of the two ribs of soft coal in the roadway. The pre-tightening force of the anchor cables is not less than 250 kN. After completing the anchor cables, shotcrete was applied on the higher rib surface of the roadway in the test section, and C20 concrete was sprayed with a thickness of 30 mm. After completion of the surrounding rock shotcrete, coal grouting was conducted on the roadway rib. Three grouting holes were arranged on the higher rib of the roadway with a spacing of 1.2 m, and two grouting holes were arranged on the lower rib with a spacing of 1.4 m. The grouting material was cement slurry and soluble silicate, in which the water–cement ratio was 1.3:1–1.4:1 (mass ratio), and the ratio of the cement slurry to soluble silicate was 1:0.5–1:1.0 (volume ratio). The cement used was P.O42.5 normal Portland cement, and the Baume density of soluble silicate (liquid sodium silicate) was 35◦ Bé. The depth of the grouting hole was 3 m, the grouting pressure was not less than 2.5 MPa, and the row distance of the grouting hole was 3.0 m.

**Figure 19.** Schematic of cooperative control technology of external anchor–internal unloading of coal roadway surrounding rock. (**I-a**) 3D schematic diagram of roadway support; (**I-b**) Front view; (**I-c**) Top view; (**II-a**) Drilling rig; (**II-b**) Steel pipe; (**II-c**) Grouting materials.

After completing the bolt-grouting support of the shallow surrounding rock of the roadway, deep coal hole-making was performed to relieve the pressure, as indicated in Figure 19. The distance from the borehole to the roadway floor is 1.5 m, and the upward angle is 6.5◦. The diameter of the shallow drilling hole is 100 mm. To protect the integrity of the shallow surrounding rock and reduce the damage caused by hole-making to the shallow coal, steel pipes were imbedded in shallow boreholes (Figure 19(II)-a), and grouting was conducted for pipe fixing. The deep part is the hole-making space, and its key parameters are as described above: the hole-making depth, spacing, and length are 8 m, 3.2 m, and 3 m, respectively.

#### *5.2. Analysis on the Effect of Cooperative Control Technology of External Anchor–Internal Unloading in Roadway Surrounding Rock*

(1) Analysis of the prestressed field of the surrounding rock reinforced anchor cables in the coal roadway test section

According to the support parameters of the original bolts and reinforced anchor cables in the test section of headgate 11231, the numerical simulation software was used to simulate the prestressed field formed by the roadway shallow surrounding rock and hole-making space, and the effective control range and control principle of the cooperative control technology of external anchor–internal unloading were comprehensively analyzed. As shown in Figure 20, two supporting bodies (anchor coal mass and hole-making space, as depicted in Figure 20a) and three zones (external anchor zone, buffer protection zone for anchorage zone, and internal unloading zone, as illustrated in Figure 20b) are formed in the roadway surrounding rock after adopting the cooperative control technology of external anchor–internal unloading. After implementing the anchor-grouting support in the roadway shallow surrounding rock, anchored coal mass is formed in the range of 0–5 m depth of the surrounding rock. The strength of the surrounding rock is improved effectively, which provides a basic protection environment for resisting large deformation of the surrounding rock caused by deep high ground pressure and strong mining. After the hole-making space is formed, a pressure relief deterioration zone is created within 7–12 m depth of the surrounding rock, which transfers the high stress in the shallow surrounding rock to the deep part, alleviating the continuous damage due to the high stress in the shallow surrounding rock. At the same time, the large-diameter hole-making space can effectively absorb the continuous deformation caused by the transfer of the deeper coal body to the roadway space due to the high horizontal stress. The depth of the surrounding rock (5–7 m) is the buffer protection space, whose function is to avoid damage to the shallow anchored coal mass caused by the hole space.

**Figure 20.** Schematic of the external anchor prestressed field and the hole−making space of coal roadway surrounding rock. (**a**) Three-dimensional prestress field; (**b**) Three zones of roadway rib; (**c**) Anchor bolt (cable) simulation scheme.

(2) Observation results and analysis of roadway surrounding rock displacement

The roadway surrounding rock displacements in the test section (using the new technology) and in the non-test section were monitored on-site. The results are displayed in Figure 21. The values of the deformation of the higher rib of the roadway in the non-test section are 400 mm and 780 mm in the no mining and mining influence stages, respectively. The deformation of the higher rib of the roadway in the test section is 190 mm after

hole-making, and the accumulated reduction in the higher rib deformation of the roadway is approximately 850 mm. As indicated in Figure 22, after the new support technology is adopted, the available width of the roadway is not less than 4.2 m, which can always meet the ventilation and transportation requirements of the roadway. Furthermore, there is no need to renovate the roadway rib. This technology effectively limits the continuous deformation of the rib surrounding rock in the test section of headgate 11231 and ensures the stability of the surrounding rock of the mining roadway affected by the deep dynamic pressure.

**Figure 21.** Deformation curve of coal roadway surrounding rock.

**Figure 22.** Advanced support section of the panel after adopting the external anchor–internal unloading technology.

#### **6. Conclusions**


results, the major technical parameters for pressure relief during hole-making were finally determined as follows: the hole-making depth, spacing, and length were 8 m, 3.2 m, and 3 m, respectively.

(4) The field practice was carried out in the test section of headgate 11231. The monitoring results showed that the displacement of the hole-making rib was reduced by 850 mm, making the available width of the roadway no less than 4.2 m, which meets the ventilation and transportation requirements of the roadway, and eliminates roadway rib renovation, which ensures the stability of the surrounding rock. Thus, the new technology is of great significance for the further development of strategies for controlling the surrounding rock of deep coal roadways affected by dynamic pressure.

**Author Contributions:** Conceptualization, S.X.; Methodology, H.L. and D.C.; Software, S.F. and X.M.; Validation, S.F., X.M. and Z.J.; Writing—original draft, H.L.; Writing—review & editing, J.C.; Visualization, Z.J.; Supervision, D.C.; Project administration, H.L.; Funding acquisition, S.X. and D.C. All authors have read and agreed to the published version of the manuscript.

**Funding:** This work was supported by the National Natural Science Foundation of China (No. 52074296).

**Conflicts of Interest:** The authors declare no conflict of interest.

#### **References**


**Gang Xu \*, Yaping Hou, Hongwei Jin and Zhongwei Wang**

College of Safety Science and Engineering, Xi'an University of Science and Technology, Xi'an 710054, China **\*** Correspondence: xugang25193@xust.edu.cn

**Abstract:** It is of great significance to obtain the source of mixed gas emission from the working face and the law of gas emission from each coal seam for the targeted implementation of gas control measures. Based on the principle that the hydrocarbon isotope values of gas in different coal seams have significant variability, a hydrocarbon isotope method for identifying the source of gas emission is proposed. Taking Pingmei No. 6 Coal Mine as the study area, the distribution characteristics of each value were obtained by testing the values of carbon and hydrogen isotopes in the gas of mined coal seams and adjacent coal seams; by testing the hydrocarbon isotope value of CH4 in the mixed gas of coal seam, the proportion of gas emission in each coal seam is determined and the law of gas emission in each coal seam is studied. The results show that the variation law of the proportion of gas emission in each coal seam can be divided into three stages: the dominant stage of gas emission in the mining layer (stage I), the stage of gas emission in the long-distance adjacent coal seam (stage II), and the dynamic equilibrium stage of gas emission in each coal seam (stage III). In the process of working face mining, the amount of gas emission in the mining layer remains in a small fluctuation state, and the proportion of gas emission decreases rapidly in stage I and stage II, and remains stable in stage III; the amount of gas emission and the proportion of gas emission in adjacent coal seams increase rapidly in stage I and stage II, and remain stable in stage III; the mixed gas emission of the working face increases rapidly in stage I and stage II, and remains stable in stage III. The calculation formula of the gas emission rate of the adjacent coal seam is established; during the development of the height of the mining fractured zone, the gas emission rate of the adjacent coal seam increases exponentially, and the gas emission ratio and gas emission amount of the adjacent coal seam increase; after the height of mining fracture zone tends to be stable, the gas emission rate, the proportion of gas emission, and the amount of gas emission remain of adjacent coal seams remain in a small fluctuation state.

**Keywords:** multi-seam mining; gas emission of working face; hydrocarbon isotope; gas emission rate of adjacent coal seam; mining fracture zone height

#### **1. Introduction**

Gas emission in the working face has the characteristics of high prediction difficulty, large damage range, and serious consequences, so it has always been the focus of mine disaster prevention and control. In particular, under the condition of multi-coal seam mining, when the layer spacing of each coal seam is small, the gas of the adjacent coal seam will pour into the mining coal seam, resulting in a relatively high gas emission of the mining coal seam, which increases the risk of the working face and the difficulty of gas control [1,2]. Gas extraction is the fundamental measure of gas control in order to reasonably and efficiently formulate the gas extraction measures of the working face, identify the source of gas emission from the working face, and analyze the law of gas emission from the working face [3,4].

The traditional prediction methods of gas emission mainly include the source prediction method [5] and the statistical analysis method [6]. The gas geology research group

**Citation:** Xu, G.; Hou, Y.; Jin, H.; Wang, Z. Study on Source Identification of Mixed Gas Emission and Law of Gas Emission Based on Isotope Method. *Energies* **2023**, *16*, 1225. https://doi.org/10.3390/ en16031225

Academic Editor: Sergey Zhironkin

Received: 9 January 2023 Revised: 19 January 2023 Accepted: 20 January 2023 Published: 23 January 2023

**Copyright:** © 2023 by the authors. Licensee MDPI, Basel, Switzerland. This article is an open access article distributed under the terms and conditions of the Creative Commons Attribution (CC BY) license (https:// creativecommons.org/licenses/by/ 4.0/).

of Jiaozuo Mining Institute uses the statistical analysis method to analyze the geological factors affecting gas emission and predict the amount of gas emission, the influence of geological factors on gas emission is deeply studied from qualitative analysis to quantitative research [7]. However, the mathematical model established by this method fails to consider the influence of mine pressure and adjacent coal seam on gas emission during mining. Lunarzewski [8] used the geomechanics of "Floor gas" and "Roof gas" and the gas emission model to calculate the contribution ratio of different gas sources to the total gas content, but this method cannot realize the dynamic prediction of gas emission in the working face. Zhang et al. [9] used statistical analysis and source prediction methods to predict the gas emission of mining faces and analyzed the influencing factors and sources of gas emission in mining faces; however, this study failed to obtain the dynamic change characteristics of gas emission during mining. Whittles et al. [10] studied the influence of geological factors on the gas flow in the goaf of the longwall coal mining face in the United Kingdom by numerical simulation; however, this method failed to obtain the influence of working face mining on gas emission and the source of gas emission in the working face.

With the development of artificial intelligence technology, some mathematical models and methods have also been introduced into the prediction of gas emission. Gu and Zhang et al. [11,12] established a new prediction model of gas emission by combining gray theory and a wavelet neural network. According to the measured data of gas emission and related geological factors in the mining area of the mine, Zhang et al. [13] established a multifactor mathematical geological model for predicting gas emission by using quantitative theory [14], considering various influencing factors including mining depth. Xiao and Zhu et al. [15,16] established a BP neural network source prediction model by combining the source prediction method with the neural network prediction technology. The above gas emission prediction method can obtain higher accuracy for short-term prediction in the case of existing gas emission data, but there is a large error for long-term prediction, and it is also difficult to obtain the source of gas emission in the working face.

The research on the law of gas emission in the process of working face mining has also achieved more results. Xu and Li et al. [17,18] studied the gas emission characteristics and gas distribution law of fully mechanized mining face through theoretical analysis and field measurement. Gao et al. [19] believed that when the mine pressure appeared, the absolute gas emission of the working face increased obviously, and the mining intensity was consistent with the change trend of gas concentration in the upper corner of the working face. Zhang et al. [20] believed that in the early stage of mining, the absolute amount of gas emission increased continuously, reached the peak when the initial pressure came, and then showed a wave-like downward trend. Cui et al. [21] considered that the absolute gas emission of the working face was positively correlated with daily output, and the relative gas emission was negatively correlated with output. Yuan and Dai et al. [22,23] mainly used the source prediction method to study the gas emission law of the protective layer working face and then took the corresponding gas control measures.

From the above, researchers have carried out a lot of research on gas emission prediction methods and emission laws, but mainly based on the constructed model to predict gas emission. The disadvantage of this method is that the applicability of the model is poor, and it is difficult to reflect the change in geological conditions over time. In addition, the research on the law of gas emission under the condition of multi-coal seam mining mainly adopts the source prediction method, which cannot realize the dynamic prediction of the proportion of gas emission and the amount of gas emission in each coal seam at the initial stage of mining. In view of the above problems, the hydrocarbon isotope method is used to dynamically predict the mixed gas in the working face, and the proportion of gas emission in each coal seam is quantitatively obtained. Then, the gas emission rate of adjacent coal seams and the variation law of gas emission in each coal seam during the mining process of the working face are studied, and the calculation formula of gas emission rate in adjacent layers is proposed, which provides a theoretical basis for formulating gas control measures in the working face under multi-coal seam mining conditions.

#### **2. Materials and Methods**

#### *2.1. Determination Method of Mixed Gas Emission Source in the Working Face*

#### 2.1.1. Theoretical Foundation

Coal seam gas is generated with the formation of coal, the main gas component of coal seam gas is CH4. During the peatification phase, i.e., the biochemical gas generation period, anaerobic microorganisms decompose organic matter to produce a large amount of CH4; in the period of coalification metamorphism, under the action of high temperature and pressure, the volatile of organic matter decreases and the fixed carbon increases, and a large amount of CH4 will be generated. Due to the different strengths of biodegradation or pyrolysis during the generation of CH4, the effect of isotope fractionation is also different, resulting in obvious differences in the carbon and hydrogen isotope values of CH4 in each coal seam gas, the difference in carbon and hydrogen isotope values is the theoretical basis for studying the source of mixed gas emission in the working face [24,25].

#### 2.1.2. Calculation Model of Mixed Gas Emission Source in the Working Face

The mixed gas emission from the working face during multi-coal seam mining comes from different coal seams. The gas is only a simple physical mixture, and its chemical properties have not changed, but the carbon and hydrogen isotopes of the gas from different coal seams in the mixed gas are different, which provides the possibility for calculating the composition of gas emission from the working face. Based on this, the calculation formula of hydrocarbon isotope value in the mixed gas is derived according to the principle of mass conservation [26,27]:

$$\delta\_{\rm mix} = \frac{V\_A \delta\_A + V\_B \delta\_B}{V\_A + V\_B} \tag{1}$$

Among them, *δmix* is the measured value of CH4 hydrocarbon isotope in mixed gas, ‰; *δ<sup>A</sup>* is the hydrocarbon isotope value of CH4 in the first coal seam, ‰; *VA* is the volume of CH4 in the first coal seam, m3; *δ<sup>B</sup>* is the hydrocarbon isotope value of CH4 in the second coal seam, ‰; *VB* is the volume of CH4 in the second coal seam, m3; *δmix*, *δA*, and *δ<sup>B</sup>* can be directly measured by an isotope mass spectrometer.

For the unit volume of mixed gas, *VA* = *aZA*, *VB* = *bZB*, where *a* is the gas emission proportion of the first coal seam in the mixed gas, %; *b* is the gas emission proportion of the second coal seam in the mixed gas, %, *a* + *b* = 1; *ZA* is the gas component of CH4 in the first coal seam, %; *ZB* is the gas component of CH4 in the second coal seam, %; *ZA* and *ZB* can be tested by a gas chromatograph for the gas composition of each coal seam.

Similarly, when the composition source of mixed gas has *n* endmembers, the calculation formula of hydrocarbon isotope value in mixed gas can be expressed as:

$$\begin{cases} \delta\_{\text{mix}} = aZ\_A \delta\_A + bZ\_B \delta\_B + \dots + nZ\_N \delta\_N\\ a + b + \dots + n = 1 \end{cases} \tag{2}$$

The gas emission ratio *a* and *b* ··· *n* of each coal seam in Equation (2) can be solved by the software MATLAB.

#### *2.2. Test Method for Hydrocarbon Isotope Value*

#### 2.2.1. Field Test Background and Conditions

Pingdingshan Tian'an Coal Co., Ltd.'s sixth mine is affiliated with China Pingdingshan Shenma Group, the administrative division is under the jurisdiction of Pingdingshan City and Baofeng County, Henan Province. The approved production capacity of the mine is 3.2 million t/a, the mine adopts the multi-level development mode of vertical shaft and inclined shaft, adopts the long wall retreating mining technology, and manages the roof by all caving methods. It is a coal and gas outburst mine [28].

The main mining method of the Ding5−6 coal seam, the Wu8 coal seam, and the Wu9−10 coal seam in the minefield is multi-coal seam mining. The Ding5−6 coal seam belongs to the most unstable minable coal seam, the Ding5−6 coal seam is divided into two layers: Ding5 and Ding6, the direct roof of the coal seam is sandstone and sandy mudstone, the main roof is sandy mudstone, and the floor is sandy mudstone or mudstone. The coefficient of variation of coal thickness of the Wu8 coal seam is 39.88%, the mining index is 0.91, the elevation of the coal seam is 60 to −1000 m, the burial depth is 150 to 1100 m, and the coal-bearing area is 36.3 km2. In terms of the characteristics of the roof and floor, the direct roof is sandy mudstone and mudstone, and the main roof is fine-grained sandstone; the floor is dark gray mudstone and sandy mudstone, and the old bottom is medium-grained sandstone, which belongs to the more stable roof and floor. The coefficient of variation of coal thickness of the Wu9−10 coal seam is 38.37%, the recoverability index is 0.95, the elevation of the coal seam is 60 to −1000 m, the burial depth is 150 to 1100 m, and the coal-bearing area is 31.3 km2. It is a more stable area-wide recoverable coal seam, the characteristics of the top and bottom plate are: the direct top plate is mainly mediumgrained sandstone or sandy mudstone, the old top is sandy mudstone, and the sandy mudstone pseudo-top can be seen locally. The direct bottom plate is mudstone, and the old bottom is sandy mudstone, which is a more stable top and bottom plate. The Wu8 −32010 working face is located in the third-level second mining area, and the working face elevation is −570 to −660 m. The strike length of the working face is 2300 m, the dip length is 220 m, and the mining height of the working face is 3.7 m. The long wall retreating mining technology is adopted, and the roof is managed by all caving methods, the working face strike section is shown in Figure 1. In order to reduce the impact of the risk of coal and gas outbursts, they first mine the Wu8 coal seam as a protective layer and then mine the Ding5−6 coal seam and the Wu9−10 coal seam. However, the average distance between the Wu8 coal seam and the upper Ding5−6 coal seam is 71 m, and the average distance between the Wu8 coal seam and the lower Wu9−10 coal seam is 13 m. During the mining process of the Wu8 coal seam, a large amount of gas from adjacent coal seams will enter the working face, which will bring potential safety hazards to the mining of the Wu8 coal seam. Therefore, it is of great practical significance to obtain the proportion of gas emission and the law of gas emission in each coal seam during the mining process of the Wu8 coal seam to ensure the safe and efficient mining of the working face.

**Figure 1.** Wu8 −32010 working face strike section.

2.2.2. Testing of Hydrocarbon Isotope Value of Gas in Each Coal Seam

According to the standard AQ1018−2006 "mine gas emission prediction method" middle spacing and adjacent coal seam emission rate relationship curve [29], the Ding5−6 coal seam and Wu9−10 coal seam gas will pour into the Wu8 coal seam, other coal seam gas will not pour into the Wu8 coal seam. Therefore, desorption gas is collected in the Ding5−6 coal seam, the Wu8 coal seam, and the Wu9−10 coal seam, respectively, to determine the gas composition and hydrocarbon isotope values, including 13C (CH4), 13C (CO2), 13C (C2H6), and 2H (CH4). Four samples are collected from each coal seam, and a total of twelve samples are collected.

The specific test method is as follows: first, six sealed tanks with a capacity of 1 L should be prepared, the sealed tank should be washed and dried before use, and the air tightness of the sealed tank should be ensured to be intact, so there is no air leakage at 300 to 400 kPa; then the coal sample containing gas is drilled by the special drilling rig for coal core in the selected place of the coal mine, and the coal sample is put into the prepared sealed tank; after leaving the well, the sealed tank is connected with the experimental equipment for gas component test and hydrocarbon isotope value test. The instrument used for the gas component test is the GC−2000 TCD gas chromatograph, and the instrument used for the hydrocarbon isotope value test is the Delta V stable isotope mass spectrometer. The experimental instruments are shown in Figure 2.

**Figure 2.** Experimental instrument diagram. (**a**) GC−2000TCD gas chromatograph. (**b**) Delta V stable isotope mass spectrometer.


The two areas of the upper corner and goaf of the Wu8−32010 working face are selected as the sampling sites of mixed gas samples, in which the sampling point of the goaf is 20 m deep into the goaf. The specific sampling site is shown in Figure 3. The sampling time is calculated from the beginning of the working face, and the sampling time is one month, that is, from 7 June 2018 to 6 July 2018. The samples are collected once a day after the samples are collected, and the samples are sent to the laboratory for mixed gas component test and CH4 hydrocarbon isotope value test. During this period, a total of 30 groups of mixed gas samples are collected.

**Figure 3.** Mixed gas sample collection location diagram.

(2) Sample collection method of mixed gas

The collection method of mixed gas samples in the upper corner of the working face is as follows: the sampling personnel stands on the footplate of the end support of the return air side and extends the expansion rod to a distance of 1.5 to 2 m from the support and a distance of 300 mm from the top side; the other end of the telescopic rod is connected with a high negative pressure suction tube, the gas in the sampling tube and the airbag is discharged, and the sampling bag is connected to start sampling. It is strictly prohibited to extend the head into the windshield during sampling.

The collection method of mixed gas samples in goaf is as follows: the sampling tube with a diameter of 20 mm is arranged in the return airway; one end of the sampling tube is 20 m deep into the goaf and a sampling device is arranged; the other end is connected to the extraction system in the return air trough, and the valve is set when connected to facilitate sampling.

#### **3. Results**

#### *3.1. Test Results of Hydrocarbon Isotope Value of Gas in Each Coal Seam*

The test results of gas composition and hydrocarbon isotope value of each coal seam are shown in Table 1. From Table 1, it can be seen that CH4 is the main gas component of each coal seam, among which the average proportion of CH4 gas component in the Ding5−6 coal seam is 71.563%, the average proportion of CH4 gas component in Wu8 coal seam is 87.340%, and the average proportion of CH4 gas component in the Wu9−10 coal seam is 88.642%, which indicates that there are some differences in CH4 gas components in each coal seam. In addition to CH4, the gas composition of each coal seam also includes N2, CO2, and C2H6.

**Table 1.** Test results of gas components and hydrocarbon isotope values of each coal seam.


In order to more intuitively show the distribution characteristics of hydrocarbon isotope values of gas in each coal seam, the tested hydrocarbon isotope values are presented in a box diagram (Figure 4), the upper and lower bounds of the box represent 75% and 25% quantiles of the data, respectively, and the hollow point in the middle of the box represents the average value of the data, while the upper and lower bounds of the vertical lines represent the maximum and minimum values of the data, respectively.

**Figure 4.** Box diagram of hydrocarbon isotope value distribution of gas in each coal seam.

It can be seen from Figure 4a that the distribution range (25% to 75%) of the carbon isotope values of CH4 gas in each coal seam does not overlap, so the carbon isotope value of CH4 gas has the condition to identify the source of mixed gas emission. It can be seen from Figure 4b that the distribution range (25% to 75%) of the carbon isotope values of CO2 gas in the Wu8 coal seam and the Wu9−10 coal seam overlaps, so the carbon isotope value of CO2 gas does not have the condition to identify the source of mixed gas emission. It can be seen from Figure 4c that the distribution range (25% to 75%) of carbon isotope values of C2H6 gas in the Ding5−6 coal seam, the Wu8 coal seam, and the Wu9−10 coal seam overlaps, so the carbon isotope values of C2H6 gas does not have the condition to identify the source of mixed gas emission. It can be seen from Figure 4d that the distribution range (25% to 75%) of hydrogen isotope values of CH4 gas in each coal seam does not overlap, so the hydrogen isotope value of CH4 gas has the condition to identify the source of mixed gas emission. Based on the above analysis, the hydrocarbon isotope value of CH4 gas can be selected to identify the source of mixed gas emission.

#### *3.2. Test Results of Mixed Gas Emission Source in the Working Face*

The working face mixed gas CH4 hydrocarbon isotope value test results are shown in Table 2.


**Table 2.** Test results of CH4 hydrocarbon isotope value in the mixed gas of the working face.

We substitute the test results of CH4 hydrocarbon isotope values in Tables 1 and 2 into Equation (2), and use MATLAB software to obtain the proportion of gas emission from each coal seam in the upper corner and goaf during the mining process of the working face (Figures 5 and 6). From Figures 5 and 6, it can be seen that the change trend of gas emission proportion of each coal seam in the upper corner and goaf during the mining process of the working face is basically the same. The gas emission proportion of the Wu8 coal seam shows a change rule of decreasing first and then stabilizing, and the gas emission proportion of the Wu9−10 coal seam and the Ding5−6 coal seam shows a change rule of increasing first and then stabilizing.

**Figure 5.** Gas emission proportion of each coal seam in the upper corner.

Further analysis of Figures 5 and 6 shows that the change law of the proportion of gas emission in each coal seam can be divided into three stages: the dominant stage of gas emission in the mining layer (stage I), the stage of gas emission in long-distance adjacent

coal seams (stage II), and the dynamic equilibrium stage of gas emission in each coal seam (stage III).

**Figure 6.** Gas emission proportion of each coal seam in goaf.

The mining time of the stage I working face is from 7 June to 15 June, and the advancing distance of the working face is 54 m. The proportion of gas emission in the Wu8 coal seam shows a trend of rapid decline, and the proportion of gas emission decreases from 95% at the beginning of mining to 60%; the proportion of gas emission in the Wu9−10 coal seam shows a trend of rapid increase, and the proportion of gas emission increases from 3% at the beginning of mining to 30%; the proportion of gas emission in the Ding5−6 coal seam shows a trend of slow increase, and the proportion of gas emission increases from 1% at the beginning of mining to 5%.

The mining time of the stage II working face is from 16 June to 22 June, and the advancing distance of the working face is 42 m. The proportion of gas emission in the Wu8 coal seam shows a trend of slow decline, and the proportion of gas emission decreases from 60% to 53%; the proportion of gas emission in the Wu9−10 coal seam shows a slow downward trend, and the proportion of gas emission decreases from 30% to 27%; the proportion of gas emission in the Ding5−6 coal seam increases rapidly from 5% to 18%.

The mining time of the stage III working face is from 23 June to 6 July, and the advance distance of the working face is 84 m. The proportion of gas emission in each coal seam enters the dynamic equilibrium stage. The proportion of gas emission in the Wu8 coal seam is stable at about 54%, the proportion of gas emission in the Wu9−10 coal seam is stable at about 28%, and the proportion of gas emission in the Ding5−6 coal seam is stable at about 18%.

#### **4. Discussion**

#### *4.1. Gas Emission Law of the Working Face and Each Coal Seam*

While testing the CH4 hydrocarbon isotope value of the working face, the gas concentration value in the return airflow of the working face is recorded. Combining this with the test results of the air volume of the working face and the gas emission ratio of each coal seam in the upper corner, the gas emission of each coal seam can be obtained (Table 3).


**Table 3.** Gas emission test results of each coal seam.

Based on the basic parameters of the working face, the "mine gas emission prediction method" (AQ1018−2006) is used to calculate the gas emission of the working face. Combining this with the gas emission test results of each coal seam in Table 3, the change trend of the measured value and the calculated value of the gas emission can be obtained (Figure 7).

The measured value in Figure 7 is the test result using the method in this paper, and the calculated value is the calculated result using the "mine gas emission prediction method" (AQ1018−2006). It can be seen from Figure 7 that the gas emission of each coal seam and working face has different variation rules. During the mining process of the working face, the gas emission of the Wu8 coal seam is maintained in a small fluctuation state, and the difference between the measured value and the calculated value is small. From the beginning of mining to 15 June, the gas emission of the Wu9-10 coal seam increases rapidly, after 15 June, the gas emission of the Wu9-10 coal seam remains in a small fluctuation state when it reaches the calculated value. From the beginning of mining to 15 June, the gas emission of the Ding5-6 coal seam increases slowly, from 15 June to 22 June, the gas emission of the Ding5-6 coal seam increases rapidly, and remains in a slight fluctuation state after reaching the calculated value. The mixed gas emission of the working face increases rapidly before 22 June and remains in a small fluctuation state after reaching the calculated value.

**Figure 7.** Variation trend of measured and calculated values of gas emission.

In summary, it can be seen that the gas emission amount of the working face and the measured value of the gas emission amount of each coal seam show an increasing trend in the early stage of the working face mining. The change range and change time of the gas emission amount are related to the geological conditions of the working face mining. The existing gas emission prediction method [30,31] makes it difficult to dynamically display and describe this change law; when the working face advances for a certain distance, the measured values of the gas emission amount of the working face and the gas emission amount of each coal seam will remain in a small fluctuation state. This small fluctuation state actually reflects the dynamic changes in coal seam mining conditions and geological conditions. The existing gas emission prediction method [32,33] usually cannot show this small fluctuation state. From the test results, it can be seen that the measured gas emission in this paper shows a small fluctuation state, which indicates that the mining conditions and geological conditions of the working face in this paper change little; however, in some cases, the changes of coal seam mining conditions and geological conditions are more intense, at this time, the gas emission will also fluctuate greatly, which will bring serious hidden dangers to the safe production of coal mines. Through the test method in this paper, the dynamic change law of gas emission in each coal seam can be obtained, so as to formulate and implement gas control measures in a targeted manner and avoid the occurrence of coal mine safety production accidents.

#### *4.2. Analysis of Influencing Factors of Gas Emission*

#### 4.2.1. Analysis of Influencing Factors of Gas Emission in the Mining Layer

According to the standard "mine gas emission prediction method" (AQ1018 –2006), the calculation method of the mining layer gas emission is as follows:

$$Q\_1 = K\_1 \cdot K\_2 \cdot K\_3 \cdot \frac{m}{M} \cdot (W\_0 - W\_c) \tag{3}$$

where *Q*<sup>1</sup> is the relative gas emission of the mining coal seam (including surrounding rock), m3/t; *K*<sup>1</sup> is the gas emission coefficient of the surrounding rock, and the value range is 1.1 to 1.3. When the roof is managed by all caving methods, 1.3 is taken; *K*<sup>2</sup> is the coal loss coefficient of the working face, which is the reciprocal of the recovery rate of the working face; *K*<sup>3</sup> is the influence coefficient of roadway pre-drainage gas on gas emission in mining layer; *m* is the thickness of mining layer, m; *M* is the mining height of the working face, m; *W*<sup>0</sup> is the original gas content of coal seam, m3/t; *Wc* is the residual gas content of coal seam, m3/t.

The calculation method of *K*<sup>3</sup> is as follows:

$$K\_3 = \frac{L - 2\text{h}}{L} \tag{4}$$

where *L* is the length of the mining face, m; *h* is the width of the coal seam gas emission zone in the roadway, m, which is considered according to the average exposure time of 200 days in the roadway.

It can be seen from Equation (3) that the gas emission of the mining layer is related to the gas emission coefficient of the surrounding rock, the coal loss coefficient of the working face, the influence coefficient of the pre −drainage gas of the preparation roadway, and the original gas content of the coal seam. In the test of this paper, the basic parameters such as the gas emission coefficient of surrounding rock, the coal loss coefficient of the working face, the influence coefficient of pre-discharge gas in preparation roadway, and the original gas content of coal seam do not change, so the gas emission of mining layer remains in a small fluctuation state and the difference between the measured value and the calculated value is small (Figure 7).

#### 4.2.2. Analysis of Influencing Factors of Gas Emission in Adjacent Coal Seam

The standard "mine gas emission prediction method" (AQ1018 –2006) also gives the adjacent coal seam gas emission calculation method as follows:

$$Q\_2 = \sum\_{i=1}^{n} \left(\mathcal{W}\_{0i} - \mathcal{W}\_{ci}\right) \cdot \frac{m\_i}{M} \cdot \eta\_i \tag{5}$$

where *Q*<sup>2</sup> is the relative gas emission of the adjacent coal seam, m3/t; *mi* is the thickness of the *i*th adjacent coal seam, m; *M* is the mining thickness of the mining layer, m; *W0i* is the original gas content of the *i*th adjacent coal seam, m3/t; *Wci* is the residual gas content of the *i*th adjacent coal seam, m3/t; *η<sup>i</sup>* is the *i*th adjacent coal seam gas emission rate, %, which can be seen in Figure 8.

1: Upper adjacent coal seam. 2: Adjacent coal seam under gently inclined coal seam. 3: Adjacent coal seam under inclined and steeply inclined coal seams

**Figure 8.** Relationship curve between gas emission rate of adjacent coal seam and layer spacing.

Equation (5) is a common method for calculating the gas emission of adjacent coal seams [34]. According to Equation (5), the gas emission of adjacent coal seams is positively correlated with the original gas content of adjacent coal seams, the thickness of adjacent coal seams, and the gas emission rate of adjacent coal seams. In the process of working face mining, the original gas content of the adjacent coal seam, the thickness of the adjacent coal seam, and other parameters remain unchanged, they have no effect on the change of gas emission of the adjacent coal seam, so the gas emission rate of the adjacent coal seam is the main parameter affecting the change of gas emission of the adjacent coal seam. When using Equation (5) to calculate the gas emission of adjacent coal seams, the gas emission rate of adjacent coal seams is usually selected from Figure 8; however, the gas emission rate of adjacent coal seams in Figure 8 is the gas emission rate when the dynamic equilibrium stage is reached after the full mining of the working face. Before the dynamic equilibrium stage is reached, the calculation of the gas emission rate of adjacent coal seams in Figure 8 will inevitably produce large errors, which is the problem with using Equation (5) to calculate the gas emission of adjacent coal seams.

The gas emission rate of adjacent coal seams is related to the gas flow characteristics of adjacent coal seams. The principle of adjacent coal seam gas flowing to the working face is (Figure 9): after the mining of the mining layer, a certain mining space will be formed, the coal rock layer around the coal seam will move to the mining space, and the original equilibrium relationship of the coal rock mass will be disturbed and destroyed, so that the original stress–strain state of the coal rock mass will change. The results of the change will lead to the release of pressure and elastic potential energy in the coal rock around the coal seam, thus forming mining fractures in the coal rock around the coal seam. As an associated gas of the coal seam, the gas itself has fluidity, and the generation of mining cracks will provide a channel for the seepage movement of gas in the coal seam. During the mining process of the mining layer, the pressure-relief gas of the adjacent coal seam flows into the mining face through the mining-induced cracks, which increases the gas emission of the mining face [35].

**Figure 9.** Gas flow characteristics of adjacent coal seams during advancing of the working face. (**a**) Unmined coal seams. (**b**) Stage I. (**c**) Stage II.

In order to further analyze the influence of the gas emission rate of the adjacent coal seam on the gas emission of the adjacent coal seam, based on the geological conditions of this paper, FLAC3D software is used to simulate the development characteristics of the fracture zone in the process of the working face advancing. According to the numerical simulation results, the change trend of the development height of the mining fracture zone with the advancing distance of the working face can be obtained. Then the adjacent coal seam gas emission rate is calculated according to Equation (5) by using the measured data of the adjacent coal seam gas emission (Figure 10).

**Figure 10.** Variation of mining fracture zone height and gas emission rate of adjacent coal seams with the advancing distance of the working face.

It can be seen from Figure 10 that the height of the mining fracture zone increases with the advance of the working face. The height of the fracture zone increases rapidly in the early stage of mining, and the growth rate slows down after full mining; when the working face advances to a certain distance, the height of fracture zone does not increase anymore. It can also be seen in Figure 10 that the adjacent coal seam gas emission rate is a dynamic change parameter in the working face mining process.

Based on the test results of this paper, combined with Figures 9 and 10, the influence of the gas emission rate of adjacent coal seams on gas emission of adjacent coal seams can be analyzed as follows:


roof, resulting in the rapid increase of the gas emission rate of the Ding5−6 coal seam in stage II.

(3) With the advance of the working face, in stage III, the mining-induced fracture zone height of the coal seam roof and floor tends to be stable. The Ding5−6 coal seam and the Wu9−10 coal seam also enter into the mining-induced fracture zone of the coal seam roof and floor, respectively, the gas emission rate of adjacent coal seams and the gas emission of adjacent coal seams will remain in a small fluctuation state until the end of working face mining.

According to the above analysis, the gas emission rate of adjacent coal seams in the process of the working face shows a phased change characteristic, which is the result of the combined effect of the working face mining method and geological conditions. Therefore, the gas emission of adjacent coal seams is the result of a combination of multiple factors, it is difficult to predict the gas emission of adjacent coal seams only by establishing a mathematical model or calculation formula [36–38], and there will be obvious defects in accuracy and timeliness.

#### *4.3. Gas Emission Rate of Adjacent Coal Seams*

In Section 4.2., the influencing factors and changing trends of the gas emission rate of adjacent coal seams have been analyzed. Because the gas emission rate of adjacent coal seams shows a phased change during the mining process of the working face, the calculation formula for the gas emission rate of adjacent coal seams will also be established in stages.

Firstly, according to Figure 11, the calculation formula of the height of the fracture zone of the coal seam roof with the advancing distance of the working face is obtained by fitting:

$$h\_1 = -0.004l^2 + 1.337l - 5.884 \quad \text{( $R^2 = 0.994$ )} \quad l \le l\_1 \tag{6}$$

where *l* is the advancing distance of the working face, m; *h*<sup>1</sup> is the height of the coal seam roof fracture zone, m; *l*<sup>1</sup> is the advancing distance of the working face when the height of the fracture zone of the coal seam roof reaches the upper adjacent coal seam, m.

**Figure 11.** Variation of the gas emission rate of lower adjacent coal seams with the advancing distance of the working face.

Similarly, according to Figure 11, the calculation formula of the height of the fracture zone in the coal seam floor with the advancing distance of the working face is obtained by fitting:

$$h\_2 = -0.002l^2 + 0.572l - 1.055 \quad \text{( $R^2 = 0.994$ )} \quad l \le l\_2 \tag{7}$$

where *l* is the advancing distance of the working face, m; *h*<sup>2</sup> is the height of the coal seam floor fracture zone, m; *l*<sup>2</sup> is the advancing distance of the working face when the height of the fracture zone of the coal seam floor reaches the lower adjacent coal seam, m.

Then, according to Figure 11, the calculation formula of the gas emission rate of the upper adjacent coal seam with the height of the coal seam roof fracture zone is obtained by fitting:

$$P\_1 = 1.236e^{0.042h\_1} \quad \left(R^2 = 0.98\right) \quad l \le l\_1 \tag{8}$$

where *P*<sup>1</sup> is the upper adjacent coal seam gas emission rate, %.

Similarly, according to Figure 11, the calculation formula of the gas emission rate of the lower adjacent coal seam with the height of the fracture zone of the coal seam floor is obtained by fitting:

$$P\_2 = 2.609e^{0.168h\_2} \quad \left(R^2 = 0.976\right) \quad l \le l\_2\tag{9}$$

where *P*<sup>2</sup> is the gas emission rate of the lower adjacent coal seam, %.

According to the gas flow characteristics of the adjacent coal seam during the advancing process of the working face, the higher the height of the mining fracture zone, the higher the gas emission rate of the adjacent coal seam, and the gas emission rate of the adjacent coal seam and the height of the mining fracture zone are exponentially related. From the layer spacing of adjacent coal seams, the smaller the layer spacing of adjacent coal seams, the higher the gas emission rate of adjacent coal seams, and the layer spacing of adjacent coal seams is inversely proportional to the gas emission rate of adjacent coal seams. In addition, it can be seen from Equations (8) and (9) that the gas emission rates of upper adjacent coal seams and lower adjacent coal seams have similar variation trends with the height of the fracture zone, and the gas emission rates of upper adjacent coal seams and lower adjacent coal seams can be combined for study.

Based on the above analysis, combined with Equation (6)–(9) and Figure 8, the calculation formula of the gas emission rate of the upper adjacent coal seam with the advancing distance of the working face can be obtained by introducing the interlayer spacing *H*<sup>1</sup> of the upper adjacent coal seam and the interlayer spacing *H*<sup>2</sup> of the lower adjacent coal seam:

$$\begin{cases} \begin{aligned} h\_1 &= -0.004l^2 + 1.337l - 5.884 \quad (R^2 = 0.994) & l \le l\_1\\ P\_1 &= (-0.024)H\_1 - 16(+2.609)e^{(-0.00218(H\_1 - 16) + 0.168)h\_1} & l \le l\_1 \end{aligned} & \text{(10)} \end{cases} \tag{10}$$

where *H*<sup>1</sup> is the distance between the upper adjacent coal seam and the mining layer, m.

Similarly, the calculation formula of the gas emission rate of the lower adjacent coal seam with the advancing distance of the working face can be obtained:

$$\begin{cases} \begin{aligned} h\_2 &= -0.002l^2 + 0.572l - 1.055 & \left( R^2 = 0.994 \right) & l \le l\_2\\ P\_2 &= \left( -0.024 \right)H\_2 - 16 \left( +2.609 \right)e^{\left( -0.00218 \left( H\_2 - 16 \right) + 0.168 \right)h\_2} & l \le l\_2\\ P\_2 &= \left( 50 - H\_2 \right)/50 & l > l\_2 \end{aligned} \tag{11} $$

where *H*<sup>2</sup> is the distance between the lower adjacent coal seam and the mining layer, m.

According to Equations (10) and (11), the variation trend of the gas emission rate of adjacent coal seams with the advancing distance of the working face can be drawn under different interlayer spacings (Figures 11 and 12). From Figures 11 and 12 it can be seen that as the working face advances, the gas emission rate of the upper adjacent coal seam and the lower adjacent coal seam has the same change trend. In the early stage of working face mining, the gas emission rate of adjacent coal seams increases exponentially with the advancing distance of the working face. After the working face is fully mined, the gas emission rate of adjacent coal seams remains stable; the larger the layer spacing with the mining coal seam is, the smaller the gas emission rate of the adjacent coal seam is and the longer the distance for the working face to reach the dynamic equilibrium state. In Figures 11 and 12, the curve of seam gas emission rate of adjacent coal seam with an interlayer spacing of 74 m and the curve of seam gas emission rate of adjacent coal seam

of the Ding5 –6 coal seam basically coincide, and the curve of seam gas emission rate of adjacent coal seam with an interlayer spacing of 16 m and the curve of seam gas emission rate of adjacent coal seam of Wu9-10 coal seam basically coincide. This shows that the error between the calculated value and the test result is small, and the accuracy of the gas emission rate of adjacent coal seam calculated by Equations (10) and (11) is high, which can meet the needs of gas emission prediction of the adjacent coal seam in production practice.

**Figure 12.** Variation of the gas emission rate of upper adjacent coal seams with the advancing distance of the working face.

The test method of gas emission in adjacent coal seams and the calculation formula of the gas emission rate in adjacent coal seams proposed in this paper can make up for the shortcomings of the existing methods, and have good applicability under complex geological conditions. Its popularization and application under the complex geological conditions of multiple coal seams can obtain considerable safety and economic benefits.

#### **5. Conclusions**


**Author Contributions:** Conceptualization, G.X. and Y.H.; Data curation, G.X., Y.H. and Z.W.; Formal analysis, G.X., Y.H. and H.J.; Funding acquisition, G.X., Z.W. and H.J.; Methodology, G.X. and Y.H.; Project administration, H.J.; Software, G.X., Y.H. and Z.W.; Supervision, H.J.; Writing—original draft, G.X. and Y.H.; Writing—review and editing, G.X., H.J., Y.H. and Z.W. All authors have read and agreed to the published version of the manuscript.

**Funding:** All of these are gratefully acknowledged. This work was financially supported by the Natural Science Foundation of China (51904231), the National Key Research and Development Program (2018YFC0807805), and the Natural Science Basic Research Program of Shaanxi (2019JM-072).

**Institutional Review Board Statement:** Not applicable.

**Informed Consent Statement:** Not applicable.

**Data Availability Statement:** Not applicable.

**Acknowledgments:** Many thanks to all the contributions and support given by the authors in preparing the writing of this article. Moreover, thanks to the Sixth Mine of Pingdingshan Tian'an Coal Co., Ltd. for providing the field test for this study.

**Conflicts of Interest:** The authors declare no conflict of interest.

#### **References**


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### *Review* **Application of Gob-Side Entry Driving in Fully Mechanized Caving Mining: A Review of Theory and Technology**

**Dongdong Chen, Jingkun Zhu, Qiucheng Ye, Xiang Ma, Shengrong Xie \*, Wenke Guo, Zijian Li, Zhiqiang Wang, Shaohua Feng and Xiangxiang Yan**

> School of Energy and Mining Engineering, China University of Mining and Technology-Beijing, Beijing 100083, China; chendongbcg@163.com (D.C.); zjk17860715825@163.com (J.Z.); yeqiucheng0314@163.com (Q.Y.); max\_cumtb@163.com (X.M.); gwk0129@163.com (W.G.); lzj\_te\_amo\_j@163.com (Z.L.); wang18336811606@163.com (Z.W.); fshahsx@163.com (S.F.); yxxcumtb@163.com (X.Y.)

**\*** Correspondence: xsrxcq@163.com

**Abstract:** China has abundant coal resources, and the distribution of coal seams is complex. Thick coal seams account for more than 45% of all coal seams. Fully mechanized top coal caving mining has the advantages of large production, high efficiency, and low cost. In fully mechanized caving mining, especially in fully mechanized caving mining of extra-thick coal seams, the mining space is ample, the mine pressure is severe, and the roadway maintenance is complex. As a result, it is necessary to summarize and discuss the gob-side entry driving of fully mechanized caving in theory and technology, which will help to promote the further development of fully mechanized caving gob-side entry driving technology. First, in recent years, the research hotspots of gob-side entry driving have focused on the deformation mechanism and the control method of the roadway surrounding rock. Secondly, this paper discusses the theoretical models of the "triangle-block" and "beam" for the activity law of the overlying strata in gob-side entry driving, including the lateral breaking "large structure" model, compound key triangle block structure model in the middle and low position, the high and low right angle key block stability mechanics model, elastic foundation beam model, low-level combined cantilever beam + high-level multilayer masonry beam structure model, and the vertical triangular slip zone structure model. It introduces the "internal and external stress field theory" and the "stress limit equilibrium zone model". Thirdly, it summarizes several numerical simulation analysis methods in different conditions or research focuses and selects appropriate constitutive models and simulation software. Finally, it introduces surrounding rock control technology, including two ribs, the roof, and under challenging conditions. It provides a method reference for support in similar projects.

**Keywords:** fully mechanized top coal caving mining; gob-side entry driving; triangle-block structure; beam structure; cable truss; support

#### **1. Introduction**

The recoverable reserves of thick coal seams in China account for about 45% of the total reserves of production mines [1]. The thick coal seam mining method has always been an important research topic in the coal industry [2–4]. The most common method of mining thick coal seams in China is fully mechanized top coal caving. In mining, most of the working face roadways will leave a certain width of a section coal pillar to protect the roadway, which is responsible for supporting the overlying strata and isolating the goaf water and harmful gas [5–8]. The width of the coal pillar increases as the mining intensity and geological conditions become more complex, resulting in a significant loss of coal resources [9,10]. Nonpillar or narrow coal pillar mining methods have been proposed to improve the recovery rate of coal resources. Gob-side entry driving is an important method of nonpillar mining in mining roadways [11]. Maintaining narrow coal pillars not only puts the roadway in the stress reduction zone [12,13] but also reduces coal pillar loss

**Citation:** Chen, D.; Zhu, J.; Ye, Q.; Ma, X.; Xie, S.; Guo, W.; Li, Z.; Wang, Z.; Feng, S.; Yan, X. Application of Gob-Side Entry Driving in Fully Mechanized Caving Mining: A Review of Theory and Technology. *Energies* **2023**, *16*, 2691. https:// doi.org/10.3390/en16062691

Academic Editor: Krzysztof Skrzypkowski

Received: 6 February 2023 Revised: 9 March 2023 Accepted: 10 March 2023 Published: 13 March 2023

**Copyright:** © 2023 by the authors. Licensee MDPI, Basel, Switzerland. This article is an open access article distributed under the terms and conditions of the Creative Commons Attribution (CC BY) license (https:// creativecommons.org/licenses/by/ 4.0/).

and improves the resource recovery rate. As a result, gob-side entry driving mining has been extensively used in thick coal seam mining.

Scholars from both home and abroad have researched the related problems of gobside entry driving [14–16]. In terms of the activity law of the overlying strata [17] in gob-side entry driving, they put forward various theories based on the mechanical model of the "triangle-block" [18–21] and "beam". In fully mechanized caving section roadway surrounding rock control theory, they put forward the "internal and external stress field theory" and the "stress limit equilibrium zone model". The proposal of these theoretical models has extensively promoted the development and application of gob-side entry driving in fully mechanized caving mining. Based on these theories, we can make correct guidance for production practices. However, the current theory has limitations, and theoretical research must be strengthened. The numerical simulation [22] of gob-side entry driving in fully mechanized caving found that the strain-softening constitutive model is mainly used to study the reasonable width of the coal pillar [23–26]. Considering the compaction effect of gangue in goaf, the double-yield model is often used to study its influence on stress redistribution [27]. Numerical simulation is an indispensable technical means in mining engineering. Although the numerical simulation results sometimes do not reflect the actual situation well, they can also provide some reference. To control the surrounding rock of gob-side entry driving in fully mechanized caving, with the development of coal mine support technology [28–30], this method gradually changed from an initial shotcrete wall [31,32] and shed support [33] to a high prestressed anchor bolt and anchor cable support. In addition, for the asymmetric deformation problem of gob-side entry, they proposed an asymmetric cable truss [34–37] and step-bundled anchor cable [31,32] support way and achieved an excellent supporting effect. In order to ensure efficient and safe production, roadway surrounding rock control has always been an important research topic in the mining field. The mine pressure is severe in a fully mechanized caving roadway, and roadway maintenance is complex. Therefore, a safe and efficient support method is needed.

Gob-side entry driving is usually arranged at the edge of the goaf of the previous working face. After the overlying rock in the goaf has collapsed and become stable, retaining the smaller width coal pillar (generally 5–8 m) plays an isolation role. It drives the roadway along the edge of the goaf [38]. The popularization and application of gob-side entry driving mining are conducive to promoting the development of mining roadway support theory, effectively improving the recovery rate of coal resources and having obvious social and economic benefits. As shown in Figure 1, according to the mining geological conditions, gob-side entry driving in fully mechanized caving mining is mainly applied to incline extra-thick coal seam caving mining, three soft coal seam caving mining, large mining height caving mining, thick and hard basic roof caving mining, island working face caving mining, deep well caving mining, and large section caving mining [39–41].

With theoretical research and mine equipment development, gob-side entry driving has gradually become an essential means in fully mechanized caving mining [42]. Therefore, the coal recovery rate increases, the surrounding rock of the roadway is effectively controlled, and the incidence of accident disasters is reduced, promoting high-quality and efficient coal mining. The research hotspot map was obtained by analyzing the research status of gob-side entry driving in fully mechanized caving mining in recent years, as shown in Figure 2.

It can be seen from Figure 2 that the research hotspots of gob-side entry driving in fully mechanized caving mainly focus on the following five aspects: <sup>1</sup> stability of the surrounding rock; <sup>2</sup> deformation mechanism of the surrounding rock; <sup>3</sup> reasonable roadway positions and the coal pillar width; <sup>4</sup> numerical simulation analysis of gob-side entry driving; and <sup>5</sup> roadway surrounding rock control technology.

By searching the keywords "gob-side entry driving" in CNKI, more than 170 related pieces of literature in recent years were obtained. Figure 3 depicts the subject words of these research papers: gob-side entry driving, extra-thick coal seam mining, fully mechanized top coal caving mining, reasonable width of coal pillar, surrounding rock control [43], large mining height mining, research and application, large section, numerical simulation, and island working face. Among them, research on the appropriate section coal pillar width [44,45] of gob-side entry driving in fully mechanized caving is a hot topic, accounting for 25%.

**Figure 1.** Some research directions of gob-side entry driving in fully mechanized caving mining.

**Figure 2.** Research hotspot map of gob-side entry driving in fully mechanized caving.

**Figure 3.** Research on gob-side entry driving technology in fully mechanized caving mining in China: (**a**) 10 research hotspots and (**b**) research trends in recent 5 years.

In this paper, the technology of gob-side entry driving with fully mechanized caving in recent years will be summarized in an all-around way. The related theories, numerical simulation constitutive models, research methods, and surrounding rock control technology will be systematically expounded. With proposals such as "carbon neutralization" and "carbon peak", coal mining is increasingly advocating for the development of safe mining technology with a high recovery rate. Fully mechanized caving along the goaf is widely used to realize a coal pillar reduction or no coal pillar. Therefore, this paper takes this as the research object. We analyze the application status and development prospects of gob-side entry driving in fully mechanized caving. When some experts or scholars need to preliminarily understand the relevant knowledge on gob-side entry driving, this paper can provide some guidance.

#### **2. Research on the Theoretical Model of Gob-Side Entry Driving in Fully Mechanized Caving Mining**

Fully mechanized top coal caving mining is a necessary technical means for high yield and high efficiency of coal mines in China. In terms of the movement law of the overlying strata in gob-side entry driving, scholars have constructed a mechanical model of the "triangle-block" and "beam" of the overlying rock, including the lateral breaking "large structure" model, compound key triangle block structure model in the middle and low position, the "high and low right angle key block stability mechanics model", elastic foundation beam model, low-level combined cantilever beam + high-level multilayer masonry beam structure model, and the vertical triangular slip zone structure model. In fully mechanized caving section roadway surrounding rock control theory, they put forward the "internal and external stress field theory" and the "stress limit equilibrium zone model".

#### *2.1. Activity Law of Overlying Strata in Gob-Side Entry Driving* 2.1.1. Mechanical Model of the "Triangle-Block"

(1) The lateral breaking "large structure" model

When the working face is left with narrow coal pillars, mining will significantly affect the surrounding rock, and the movement of the overlying strata will be violent, which will cause potential safety hazards. By analyzing the characteristics of the surrounding rock, scholars put forward the stability control principle of the "large structure" of the lateral breaking of the overlying strata in the adjacent goaf and the "small structure" of the overlying roof of the roadway surrounding rock. They established the large structure model of the overlying rock mass of the roadway driving along the goaf, as shown in Figure 4 [46]. By analyzing the stability of the triangular arc block of a basic roof, as shown in Figure 5, scholars proposed a surrounding rock control mechanism. Among them, *F*<sup>Z</sup> is the resultant force of the self-weight of key block B, *F*<sup>R</sup> is the resultant force of the self-weight of the upper weak overburden rock, the vertical force and horizontal force of rock block A to key block B are *R*<sup>A</sup> and *F*A, respectively, and the vertical shear force and horizontal thrust of structural block C to key block B are *R*<sup>B</sup> and *F*B, respectively, the supporting force of the gangue, end coal, and lateral coal in the goaf to key block B are *F*G, *F*SM, and *F*M.

**Figure 4.** The lateral breaking structure model [46].

**Figure 5.** Mechanical analysis of the lateral breaking "large structure".

(2) Compound key triangle block structure model in the middle and low position

The construction of a mechanical model of a middle and low composite key triangular plate structure with synchronous and asynchronous migration of a basic roof (low position) at the end of the adjacent goaf of a fully mechanized caving face and the adjacent key hard rock layer (middle position) is shown in Figures 6 and 7. The stability characteristics and engineering disaster conditions of the key triangular plate structure in three-time states (before the formation of the section coal lane, after the formation of the section coal lane, and at the time of mining) were explored. The fracture position of the main roof of fully mechanized caving and its influence were obtained [47]. The same symbols in Figures 5 and 7 also have the same meaning. In addition, *F*' <sup>Z</sup> is the resultant force of the self-weight of key block B (mesoposition), and *F*' <sup>R</sup> is the resultant force of the self-weight of the upper weak overburden (mesoposition). The movement of weak rock strata above key block B separates from the hard rock strata above it and loses the transmission of force, namely, *F*<sup>X</sup> = 0.

**Figure 6.** Compound triangle block structure model [47].

**Figure 7.** Mechanical analysis of the key triangular plate structure [47].

(3) The high and low right angle key block stability mechanics model

In fully mechanized caving mining of extra-thick coal seam with a hard and thick main roof, the mining space is large, and the strength is high, which causes the overlying strata to fracture, move, and collapse, and the influence is vast. Therefore, the low-key and high-key strata play a key role in the stability of the surrounding rock of the gob-side entry. Based on the theory of internal and external stress field and limit equilibrium theory, as shown in Figures 8 and 9, scholars have deduced a mechanical model for the stability of high and low right-angle key blocks with periodic breaking [48] and analyzed the joint stability of high and low key blocks.

**Figure 8.** Fracture morphology of key blocks [48].

**Figure 9.** Mechanical analysis of high and low right angle key blocks [48].

2.1.2. Mechanical Model of the "Beam"

(1) Elastic foundation beam model

This model was based on the theory of masonry beams with lateral roof fracture, considering the deformation characteristics of coal seam and immediate roof. Scholars have established the basic roof elastic foundation beam model, as shown in Figure 10. They deduced the expressions of the lateral roof bending moment and displacement and then obtained the basic roof breaking position; the influence of the main roof, immediate roof, and coal seam thickness and elastic modulus on the lateral fracture position of the roof was explored [49]. Figure 11 shows the overburdened structure after the basic roof is broken.

**Figure 10.** Structural model of elastic foundation beam [49].

(2) Low-level combined cantilever beam + high-level multilayer masonry beam structure model

The mining space formed by the mining of the fully mechanized caving face in the extra thick coal seam is large, the roof breaking and migration range are wide, and the development height of the overburdened caving zone and the fracture zone is significantly increased. When there are multiple layers of hard rock strata in the roof, the overlying hard key stratum breaks not only one layer but multiple layers. Therefore, for the mining of extra-thick coal seam with a hard roof, scholars have proposed a lateral overburden

structure model of "low-level combined cantilever beam + high-level multi-layer masonry

**Figure 12.** Low-level combined cantilever beam + high-level multilayer masonry beam structure model [50].

#### (3) The vertical triangular slip zone structure model

Scholars took a fully mechanized caving face with a large mining height in an extrathick coal seam as their research object and analyzed the activity range, fracture field distribution, motion characteristics, and structural characteristics of the overlying strata at the end of the working face with a large mining height in the extra-thick coal seam. It was proposed that there is a stable stress reduction zone with a triangular slip zone structure at the end of the goaf. This is conducive to the layout of gob-side entry driving with a small coal pillar and the maintenance of a small coal pillar roadway. According to the movement characteristics of the overlying strata in the goaf and the time–space relationship, the reasonable position and time of the small coal pillar driving along the goaf were determined, as shown in Figure 13 [51].

**Figure 13.** The vertical triangular slip zone structure model [51].

*2.2. Surrounding Rock Control of Fully Mechanized Caving Section Roadway* 2.2.1. The "Internal and External Stress Field Theory"

Scholars have established the structural mechanics model of gob-side entry driving by the theoretical analysis method. They established the expression of "internal stress field" width and determined the reasonable position of gob-side entry driving and the reasonable width of the coal pillar. They predicted the deformation of the surrounding rock of gob-side entry driving, as shown in Figure 14 [52]. The movement law and deformation failure characteristics of the surrounding rock of the roadway and the curved triangular block of the end basic roof in different stages were studied. The overall mechanical environment of the fully mechanized caving along the goaf was analyzed, and stability control theory was preliminarily formed. In Figures 15–17 [53], scholars also established a mechanical model of the roadway surrounding rock structure. The relationship between the key block's rotation angle and the coal pillar's overlying load was obtained, and the coal pillar's width was calculated to determine the fracture position of the basic roof. In Figure 14, the interval calibrated by *S*<sup>1</sup> is called the "internal stress field", and the interval calibrated by *S*<sup>2</sup> is called the "external stress field". *σ*<sup>y</sup> is the lateral support pressure; *K* is the stress concentration factor; and *γ* is the average bulk density of the overlying strata. Moreover, *H* is the buried depth of the roadway.

**Figure 14.** Structural mechanics model of gob-side entry driving [52].

**Figure 15.** Structure model with main roof fracture line above the solid coal [53].

**Figure 16.** Structure model with main roof fracture line above the roadway [53].

**Figure 17.** Structure model with main roof fracture line above the coal pillar [53].

2.2.2. The "Stress Limit Equilibrium Zone Model"

In Figures 18 and 19, based on the stress characteristics of the surrounding rock of gob-side entry driving in deep-well fully mechanized caving, considering the strengthsoftening characteristics of coal and rock mass at the interface between the roadway side and roof and floor, scholars have established the mechanical analysis model of two ribs and deduced the theoretical calculation of the limit equilibrium zone width and the coal stress displacement of two ribs [54]. Based on the distribution characteristics of the inclined abutment pressure of the coal body on the goaf side and the limit equilibrium theory of a coal pillar in roadway protection, the analytical expressions of the upper and lower limits of a reasonable width of a narrow coal pillar in roadway protection were determined [55]. Based on analyzing the stress environment of gob-side entry driving, the principle of damage mechanics was used to analyze the abutment pressure distribution of the solid coal side of gob-side entry driving under the given deformation, and the relationship between the abutment pressure distribution and parameters such as the coal rock thickness and elastic modulus was discussed. It is of great significance to the upkeep of gob-side entry

as well as the study of the floor heave mechanism and control [56]. In Figure 18, the shear stress at the interface between the coal seam and the roof and floor is *τ*xy, the vertical pressure is *σ*y, the horizontal stress inside the coal rock mass is *σ*x, and the roadway rib support resistance is *f* i. In Figure 19, *f* <sup>s</sup> is the support resistance of the general goaf side, *f* <sup>z</sup> is the support resistance of the roadway side, and the peak stress in the limit equilibrium zone of the coal pillar side should be *σ*ym.

**Figure 18.** The mechanical model of solid coal rib [54].

**Figure 19.** The mechanical model of coal pillar rib [54].

Among the above many theoretical models, the rock beam model belongs to a relatively simplified model, and the triangular block structure is a theoretical model that is more recognized by scholars at present. It is also more often used in the related research of gob-side entry driving.

#### **3. Numerical Analysis Method of Gob-Side Entry Driving in Fully Mechanized Caving Mining**

Numerical simulation is an indispensable technical means in mining engineering. At present, the most commonly used numerical analysis methods are the finite element method, boundary element method, finite difference method, weighted residual method, discrete element method, rigid body element method, discontinuous deformation analysis

(DDA) method, manifold element method [57], etc. In the study of fully mechanized caving gob-side entry driving engineering problems, scholars often use FLAC3D, UDEC, PFC, and other numerical software according to different working conditions and select the appropriate software analysis [58–61]. As shown in Figure 20, it shows the constitutive model used in FLAC3D.

**Figure 20.** Constitutive model of FLAC3D used gob-side entry driving in fully mechanized caving.

#### *3.1. Constitutive Model of FLAC3D*

#### 3.1.1. Strain-Softening Model of Yielding Coal Pillar

In studying the surrounding rock control of gob-side entry driving in fully mechanized caving, many scholars mostly choose the strain-softening constitutive model when studying reasonable coal pillar width. Compared with the Mohr–Coulomb constitutive model, it can more truly reflect the yield of the small coal pillar, especially when retaining a small coal pillar. It can also provide an accurate and reliable basis for formulating a sensible coal pillar width and support [23–26]. The strain-softening model reflects the real failure properties of coal pillars as follows: the elastic stage is consistent with the Mohr–Coulomb model. After entering the plasticity, the cohesion and friction will gradually decrease with the plastic strain [30,62,63].

FLAC3D finite difference software has a strain-softening constitutive model [60]. Scholars have established a standard specimen model for a uniaxial compression 1:1 simulation and carried out the corresponding parameter iterative inversion through indoor test and numerical simulation then fitted the standard specimen parameters for numerical simulation [64–66].

#### 3.1.2. Double-Yield Model of Goaf

Due to the compaction of gangue in the goaf, the stress state of the surrounding rock in the gob-side entry will be affected [67]. Therefore, the double-yield model can well-simulate the influence of gangue on stress redistribution [27]. In the gob-side entry driving process, the coal pillar's bearing capacity needs to be considered, and the influence on the goaf cannot be ignored [68–73]. In the numerical simulation, the "cap pressure" is the parameter that mainly determines the compaction characteristics of the goaf material in the simulation, which is controlled by Table [74,75]. According to the classical theoretical formula of Salamon, the corresponding parameter inversion is carried out by establishing a 1 m × 1 m × 1 m model in the numerical simulation so that the parameters

can precisely reflect the actual condition [76,77]. The specific parameter inversion process in the numerical model is shown in Figure 21.

**Figure 21.** Inversion fitting process of constitutive model parameters in numerical simulation: (**a**) strain-softening model and (**b**) double-yield model.

#### *3.2. UDEC Simulation of Coal Pillar Fracture*

Through the UDEC (a discrete element software), many scholars can intuitively see the fracture development and damage degree in the coal pillar of gob-side entry driving [78–81]. By analyzing the crack propagation morphology and plastic state in the coal pillar, they can guide the reasonable setting position of the gob-side entry and coal pillar width to minimize the influence of the coal pillar on the stability of the roadway [82–84]. As shown in Figure 22, it indicates the development state of coal pillar cracks in gob-side entry driving.

**Figure 22.** Development state of coal pillar fracture of gob-side entry driving in fully mechanized caving: (**a**) fracture pattern, (**b**) cracks, and (**c**) state.

#### *3.3. Other Numerical Analysis Methods*

Many scholars have carried out uniaxial and triaxial compression simulation tests by sampling coal on-site and combining discrete element software PFC2D/PFC3D to explore the crack propagation law of coal specimens under the condition of prefabricated cracks [85–87] to guide the setting of coal pillars in gob-side entry driving to ensure the optimal stress field environment of roadway surrounding rock. Some scholars have studied the dynamic, progressive failure process of coal rock samples through CDEM to analyze the influence of cracks on the stability of coal rock columns [88,89].

#### **4. Surrounding Rock Control Technology of Gob-Side Entry Driving in Fully Mechanized Caving Mining**

The structural characteristics of overburdened roadway rock differ in different mining stages of gob-side entry driving, which has a significant impact on the surrounding rock support [90]. It is essential for the stability of the surrounding rock to optimize the support parameters according to the structural characteristics of the overlying rock [91]. Scholars have put forward various supporting technologies for roadway driving along the goaf. This paper will introduce the surrounding rock control technology from three aspects: two ribs, the roof, and other complex conditions.

#### *4.1. Two-Rib Support of the Roadway*

#### 4.1.1. Support of Coal Pillar Rib

In gob-side entry of fully mechanized caving, the coal pillar rib usually adopts general support forms, such as a bolt + ladder beam of steel (W, JW steel strip) + mesh and anchor cable support. However, with the progress of fully mechanized caving mining, the fracture development of the coal pillar is obvious, and the stress concentration at the end of fracture is obvious, which leads to the weakening of the bearing capacity and an antideformation and failure ability of the coal pillar [92,93]. The linkage between the coal pillar and top coal is large. The failure of the coal pillar reduces the stability of the top coal, increasing the deformation and pressure of the roadway and increasing the difficulty of support. Using an ordinary bolt and cable support on the coal pillar's rib to maintain stability is challenging. Roadside support can assist the coal pillar in bearing roof pressure and improve the bearing capacity of the coal pillar.

The roadside support of gob-side entry driving is mainly divided into concrete wall support on the side of the coal pillar roadway and filling support on the side of the coal pillar goaf, as shown in Figure 23. The pouring concrete wall support on the side of the coal pillar roadway refers to establishing a certain width of the reinforced concrete wall in the roadway. The supporting wall is connected with the roof, floor, and the coal pillar rib through the preset high-strength bolt. Therefore, the reinforced concrete wall and the surrounding rock are coordinated [31,32]. The filling support of the goaf side of the coal pillar refers to the injection of foam, fly ash material, high water material, paste material, or cement slurry near the goaf side of the coal pillar. It can replace part of the falling gangue or directly fill it, thereby reducing the roof activity space [94,95].

**Figure 23.** Support diagram of coal pillar rib: (**a**) filling support on the side of the goaf, (**b**) concrete wall support on the side of the roadway, and (**c**) general bolt and anchor cable support. (**a**,**b**) [31,32] and (**c**) [94,95].

#### 4.1.2. Support of Virgin Coal Rib

The degree of mine pressure on the virgin coal rib is small, and the damage to the surrounding rock of the roadway is also tiny. As a result, the support method is simple. Bolt + ladder beam of steel (W, JW steel strip) + mesh is often used, and sometimes a single anchor cable is also used for reinforcement support.

#### *4.2. Roof Support of Gob-Side Entry Driving*

For a long time, scholars have conducted much research on the problem of roof control in gob-side entry driving. The roof is usually controlled with a combination of various support methods, among which, the most commonly used is bolt support [44]. The arrangement of bolt support in various fully mechanized caving roadways is the same, so this paper focuses on roof control technology with an anchor cable as the core.

#### 4.2.1. Single Anchor Cable Support or Anchor Cable + Steel Strip Support

When the coal seam thickness is less than six m, two or three independent single anchor cables are arranged on the roof for support, as shown in Figure 24a,b, and the anchor cable can be anchored to the stable rock stratum. W steel strips are also commonly used to connect the anchor cables, as shown in Figure 24c,d.

#### 4.2.2. Anchor Cable Truss Support

The anchor cable truss comprises a long anchor cable and a special connecting lock device [96]. Figure 24e,f represents symmetric and asymmetric layouts, respectively. The special connecting lock device connecting the long anchor cable is shown in Figure 24g.

Figure 24h indicates its control principle [97,98]: The cable truss system gradually locks during roof rock deformation, increasing the compression value of shallow surrounding rock and preventing excessive deformation of roadway surrounding rock; the anchor cable truss has a long length and solid shear resistance. It crosses the greatest shear stress area at the coal pillar–roof junction obliquely, enhancing the surrounding rock's shear resistance and maintaining its stability in the coal pillar's corner area.

#### 4.2.3. Cable Beam Truss Support

The cable beam truss structure comprises a long anchor cable, channel steel support beam, steel support beam, and lock. The single anchor cable is first connected with a high-strength steel support beam, and the anchor cable near the side of the coal pillar is connected with a channel steel support beam for a secondary connection. The support structure is arranged near the side of the coal pillar rib, as shown in Figure 24i,j. Figure 24k is the on-site support diagram, and Figure 24l is the supporting principle diagram.

The control principle is as follows: After applying a high pretightening force, the anchor cable, steel (channel steel) support beam, and coal–rock mass form an inverted trapezoidal bearing structure. When subjected to unbalanced abutment pressure, the inverted trapezoidal structure jams the two corners. The greater the load, the greater the force of the anchor cable and the formation of a stress arch with a base point at the two corners. The formation of the stress arch weakens the transfer of external pressure to the interior, reducing the asymmetric subsidence of the roof and horizontal extrusion deformation [99]. The anchor cables are connected by the high-strength steel support beam, which is more flexible to adapt to extrusion deformation and can prevent the connection structure from failing due to horizontal dislocation of surrounding rock. Increasing the internal hole size of the channel steel can reserve the deformation space for horizontal movement and avoid the stress concentration between the channel steel and the anchor cable due to the horizontal movement of the rock stratum [100].

#### 4.2.4. Anchor Cable + Channel Steel Support

In mining hard and thick main roof coal seams, there are some control problems, such as severe overburden activity and asymmetric deformation in the roof. Scholars have proposed the asymmetric combined control technology of the roof with anchor cable– channel steel combination [101]. Figure 24m,n shows that each row contains four or five anchor cables. The anchor cables near the two ribs are deflected outward by 15◦, and the anchor cables in the middle position are arranged perpendicular to the roof. Figure 24p is the supporting principle diagram.

#### 4.2.5. Step Bundled Anchor Cable Support

For the roof support of ultra-thick coal seams (up to 15 m), some scholars have proposed the supporting technology of step-bundled anchor cables, as shown in Figure 24q,r. Moreover, the step bundle anchor cable comprises 5 anchor cables and a bundle anchor cable tray arranged in a "2-1-2" manner. The center is a 22 mm × 10,300 mm anchor cable surrounded by two 22 mm × 6300 mm and two 22 mm × 8300 mm anchor cables. The anchor cables are arranged diagonally and connected by a porous tray [32], as shown in Figure 24s [32].

#### *4.3. Support under Difficult Conditions such as a Broken Roof*

Figure 25 depicts an early support form of roadway driving along the goaf, primarily shed support, including I-steel and U-steel support. Secondary or multiple mining may influence the roads during layered mining and the mining of coal seam groups, and the deformation and failure of the surrounding rock are severe. The shed support and steel mesh combination can effectively limit the roadway's severe deformation. At the same time, it can improve the stress environment and the mechanical properties of the surrounding rock with a high-strength grouting anchor cable, indirectly improve the majestic residual strength and self-bearing capacity, significantly slow down the large deformation of the surrounding rock, and ensure the stability of the surrounding rock [33].

**Figure 24.** *Cont*.

**Figure 24.** The roof support of gob-side entry driving: (**a**) single anchor cable support, (**b**) on-site support diagram, (**c**) anchor cable + W steel strip support, (**d**) on-site support diagram, (**e**) symmetric cable truss support, (**f**) asymmetric cable truss support, (**g**) connecting lock device, (**h**) supporting principle diagram, (**i**) asymmetric double anchor cable support, (**j**) cable truss + single anchor cable support, (**k**) on-site support diagram, (**l**) supporting principle diagram, (**m**) channel steel + four anchor cables support, (**n**) channel steel + five anchor cables support, (**o**) on-site support diagram, (**p**) supporting principle diagram, (**q**) single-step bundled anchor cable support, (**r**) step bundled anchor cable support, (**s**) stepped anchorage beam cable, and (**t**) supporting principle diagram. (**f**) [96], (**h**) [97,98], and (**m**,**n**) [101].

A single support method in the gob-side entry of fully mechanized caving often cannot meet the support requirements. Therefore, the combined support form with an anchor cable support as the core and other support methods (bolt, metal mesh, etc.) is often used to achieve adequate control of the surrounding rock, as shown in Table 1. The '√' in Table 1 represents the support form used in a certain spatial orientation of the roadway.


**Table 1.** Roof combined support table of gob-side entry driving in fully mechanized caving.

**Figure 25.** Shed support-grouting anchor cable cooperative support diagram [33].

#### **5. Engineering Monitoring**

In addition to the above research, other scholars also analyzed the stress and deformation laws of surrounding rock in gob-side entry driving using field engineering monitoring.

The lateral roof structure type and lateral abutment pressure distribution characteristics were determined using microseismic monitoring and stress dynamic monitoring [44,102]. As a result, the deformation, failure mechanism, and control of fully mechanized caving roadway along the goaf were studied. The borehole peeping method was used to measure the main roof's fracture position and the roof's two-way movement characteristics. Combined with the CT identification of the asymmetric evolution process of the microcracks in the roof coal, comprehensive support technology was proposed [37,103]. In addition, to evaluate the feasibility of the support scheme and understand the working state of the support scheme in detail, the surface displacement of the roadway was monitored by arranging the measuring station; using the steel ruler and the measuring line, the separation sensor monitored the roof separation; the stress of the coal pillar was monitored with the borehole stress meter; and the bolt cable dynamometer monitored the working resistance of the bolt cable [104,105]. As shown in Figure 26, represents a variety of monitoring instruments.

**Figure 26.** *Cont*.

**Figure 26.** Engineering monitoring instruments: (**a**) detection equipment for internal fracture of the coal mass, (**b**) industrial CT scanning system, (**c**) roadway surface displacement monitoring equipment, (**d**) roof abscission layer instrument, (**e**) borehole stress gauge, and (**f**) anchor cable dynamometer.

#### **6. Discussion**

Gob-side entry driving is usually arranged at the edge of the goaf of the prior working face, retaining a certain width coal pillar (generally 5–8 m).

This paper summarized seven theoretical models of the overlying strata activity law and surrounding rock control in fully mechanized caving gob-side entry driving (the lateral breaking "large structure" model, compound key triangle block structure model in the middle and low position, the high and low right angle key block stability mechanics model, elastic foundation beam model, low-level combined cantilever beam + high-level multilayer masonry beam structure model, the vertical triangular slip zone structure model, the "internal and external stress field theory" and the "stress limit equilibrium zone model"). Three kinds of constitutive models (strain-softening, Mohr–Coulomb, double-yield) and numerical simulation methods were discussed. The support methods of coal pillar ribs, virgin coal ribs, and roofs of gob-side entry driving in a fully mechanized caving face are summarized. The following conclusions and prospects have been reached:


is essential in optimizing the support design scheme using microseismic monitoring, dynamic stress monitoring, and borehole peeping for on-site engineering monitoring.

(5) Gob-side entry driving in fully mechanized caving is an important method of thick coal seam mining. In the future, we still need to strengthen the research on the basic theory to help us gain a more in-depth understanding of the various problems in gobside entry driving. Further, we need to explore the nonpillar mining technology and develop more effective surrounding rock control technology to improve the recovery rate of coal resources in fully mechanized caving mining. In addition, the future method of gob-side entry driving with fully mechanized caving will also take into account precision, automation, and greening to realize safe and efficient mining of coal mines.

In this paper, we have summarized the relevant theories and technologies of gob-side entry driving in fully mechanized caving face and put forward that its future development direction should focus on theoretical research, nonpillar mining, and efficient surrounding rock control, which is helpful to promote the further development of gob-side entry driving technology in fully mechanized caving face.

**Author Contributions:** Conceptualization, D.C. and S.X.; data curation, J.Z.; formal analysis, Q.Y. and X.M.; funding acquisition, D.C.; investigation, W.G., Z.L. and Z.W.; methodology, D.C., J.Z. and Q.Y.; project administration, D.C. and S.X.; software, D.C., X.M., S.F. and X.Y.; supervision, S.X.; validation, X.M. and W.G.; writing–original draft, D.C., J.Z. and Q.Y. writing–review and editing, D.C., J.Z. and Q.Y. All authors have read and agreed to the published version of the manuscript.

**Funding:** This work was financially supported by the National Natural Science Foundation of China (Grant No. 52004286), the Fundamental Research Funds for the Central Universities (Grant No. 2022XJNY02), the China Postdoctoral Science Foundation (Grant No. 2020T130701, 2019M650895), all of which were gratefully acknowledged.

**Conflicts of Interest:** The authors declare no conflict of interest.

#### **References**


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