Abstract
Irregular coal pillars inevitably appear in the layout of the current long-wall mining method, which easily forms stress concentrations and becomes a heavy disaster area of rock burst. In order to solve the impact risk of irregular coal pillar working face, it is necessary to study the instability mechanism of the coal pillar and put forward effective prevention and control measures. Based on the research background of 14320 working face of the Dongtan Coal Mine in the Yanzhou mining area of China, this paper studies the prediction and prevention of rock bursts in this kind of coal pillar by means of theoretical calculation, numerical simulation, engineering analogy, and field monitoring. The results show that (1) the absolute stability of coal pillar is that the width of coal pillar B reaches twice the support pressure of 2L, and the possibility of instability from large to small is coal pillars 2, 5, 3, 1, and 4. (2) The ratio of coal pillar strength to its average load determines the stability coefficient of the coal pillar, and it is judged that coal pillars 1 and 4 are in a stable state, coal pillars 3 and 5 are in a limit equilibrium state, and coal pillar 2 is in an unstable state. The numerical simulation shows that the maximum stress value inside the coal pillar during the mining process is basically consistent with the theoretical calculation of the bearing strength of the coal pillar. (3) The new evaluation method is used to evaluate the rock burst risk degree of the working face roadway: 156.75 m is a strong rock burst risk zone, 728.18 m is a medium rock burst risk zone, and 176.88 m is a weak rock burst risk zone. (4) Regional prevention and local prevention measures are proposed for the risk of rock burst in the roadway, which reduces the stress concentration of the coal pillar. It is verified that the pressure relief effect is remarkable, and the safe mining of such an irregular coal pillar working face is completed, which provides a solution for studying and solving such rock burst risk.
1. Introduction
China is an important country in terms of coal production and coal consumption. Most of China’s coal production bases are distributed in the central, northeastern, northern, and northwestern parts of the country. More than 95% of the raw coal is mined via underground mining methods.1,2 The production mode of underground operations is very different from that of surface mining. There are many technical problems in underground mining, and many coal mine production accidents will be faced at the same time.3−5 Rock burst is a dynamic phenomenon in which the elastic potential energy of the surrounding coal and rock mass is suddenly and violently released under certain conditions during coal mining production, resulting in severe vibration and severe damage to the coal and rock mass, making this phenomenon very destructive.6−11
In coal mining, the long-wall mechanized mining method can greatly improve the output efficiency. Under the influence of geological conditions, most adjacent working faces will inevitably form either a single irregular coal pillar or a continuous irregular coal pillar, continuing to be affected by factors such as the goaf.12 As the coal mining face advances, the irregular coal pillar of the mine will bear the weight of the overlying roof rock on the working face and the multidirectional lateral abutment pressure near the goaf. The superposition of stress leads to the accumulation of a large amount of elastic energy, creating the main area of stress concentration. The mining disturbance of the working face can easily lead to the instability of the coal pillar, resulting in a rock burst accident.13,14 The complexity of stress changes and the immediacy of disasters pose a serious threat to the safety of the coal mining face.15 At present, more and more working faces are affected by the stress of coal pillars in the adjacent goaf, resulting in rock burst accidents. Therefore, many experts have carried out abundant theoretical and practical research on the mechanism and prevention of rock burst in working faces affected by regular and irregular coal pillars. Based on the upper irregular coal pillars and faults encountered in working face 7301 in the Zhaolou Coal Mine, He et al. studied the stress and energy evolution when the working face passed through the influence area of the coal pillars.16 Fang et al. established a mechanical model of a triangular structure at the end of the goaf based on the roof fracture characteristics of a working face and analyzed the instability principle of a triangular coal pillar via theory and numerical simulation.17 Wang et al. studied triangular coal mining technology and roof control technology for the problem of coal resource waste caused by coal pillar areas with irregular boundaries and triangular coal pillars.18 Liu et al. analyzed the mechanism of rock burst induced by a thick and hard roof and residual coal pillars through field investigations, theoretical analysis, and numerical simulations based on the engineering background of the Shendong Buertai Coal Mine.19 Regarding research on rock burst prevention and control methods, Gao et al. studied the seismic isolation technology and implementation method of a deep rock burst roadway and effectively protected the stability of the structure of the rock surrounding the roadway through dual protection means of isolation and shock absorption.20 Xu et al. studied the dynamic response characteristics of the surrounding rock and support system of a roadway under different vibration intensities and proposed the full anchor cable yield support technology, which effectively controls the stability of the rock surrounding the roadway under dynamic rock burst.21 At present, most researchers have studied and analyzed the reasonable retention of the coal pillar width and the combination of faults and roofs. Research on prevention and control technologies has also been conducted from the perspective of support. Few studies have been conducted on the rock burst of continuous irregular triangular coal pillars on working face rock bursts and prevention and control measures.
In summary, this paper studies the stress concentration problem of five continuous irregular triangular coal pillars in the 14320 mining face of Dongtan Coal Mine and provides methods and ideas for solving the prediction of rock burst risk and effective prevention and control of irregular coal pillars. The relationship between the influence distance of the abutment pressure and the safety size of the coal pillar is determined through theoretical calculation and analysis, and the correlation between the stability index of the coal pillar and its stability state is established. The internal stress change of the irregular coal pillar is analyzed by the numerical simulation method. Combined with the safety factor of the coal pillar, the safety size of the coal pillar, and the stress difference between theory and simulation, the prediction criterion of rock burst risk of irregular coal pillar is obtained. Finally, through engineering analogy and on-site monitoring methods, targeted measures to prevent rock burst are proposed, and the scour prevention effect is fully certified.
2. Conditions of Test Working Face and Continuous Irregular Coal Pillar
2.1. Basic Situation of 14320 Working Face
Fully mechanized caving face 14320 is located in the western part of mining area 14 in the Dongtan Coal Mine. The Xinglongzhuang Coal Mine is located in the north, and a large number of goafs are located in the south. With continuous mining, five continuous irregular triangular coal pillars will be formed within the southern goaf. The Baodian Coal Mine boundary coal pillars are located in the west, with a width of 50.2 m (Figure 1).
Figure 1.
Plane layout of working face 14320.
2.2. Continuous Irregular Triangular Coal Pillar Area Conditions
Continuous irregular triangular coal pillars are left after the completion of mining in working faces 14317, 14318, and 14319. Because the mining block section in the western region is irregular, working face 14320 is planned here after several optimization designs. For the irregular coal pillar outside the tail entrance, the size is divided into five sections. The sizes of the first, second, third, fourth, and fifth sections are 0–145, 11–105, 12–133, 15–146, and 0–126 m, respectively (Figure 2).
Figure 2.
Diagram of the sizes of the irregular triangle coal pillars.
3. Stability Analysis of Continuous Irregular Triangular Coal Pillar Area
3.1. Characteristics of Elastic–Plastic Deformation Zone of Coal Pillar
For the coal mining face, the coal pillar is stable enough to ensure the safety of its mining and its width and stress distribution determine its own stability. When one side of the working face remains after the end of mining, and the other side is being mined, its stability is mainly affected by the abutment pressure caused by the mining distance L and the width of the coal pillar B, mainly in the following situations22 (Figure 3).
-
(1)
When B ≥ 2L, the stress is evenly distributed in the coal pillar and is equal to the stress of the original rock. The edge is the stress increase area, and the stress is more concentrated. x3 is the influence distance of the abutment pressure, x2 is the length of the pressure peak from the coal wall, and 0–x3 corresponds to the three states of the coal pillar for rupture, plasticity, and elasticity. The stress distribution at this time results in the coal pillar being in a relatively stable state. When the coal pillar is affected by external conditions such as faults, disturbances, and other factors, it has a strong ability to maintain stability. In this case, it is the absolute width of its stability.
-
(2)
When 2L > B > L, due to the continuous increase in the bearing pressure, the stress in the middle of the coal pillar is superimposed, and thus, the stress in the middle is greater than that in the original rock. However, it does not reach the ultimate bearing strength of the coal pillar, and there is still elastic coal. Therefore, the coal pillar has a certain resistance in this state, and the width is a general width that can maintain stability.
-
(3)
When B = L, the peak stress just reaches the limit of the bearing strength of the coal pillar, and the elastic body in the middle disappears and no plastic failure occurs. In this case, the width is the limit width, and the coal pillar is in dynamic equilibrium, but there is no resistance to external influences in this state. If any unfavorable factors occur, instability and failure of the coal pillar will occur.
-
(4)
When B < L, the peak stress of the coal pillar exceeds its bearing limit, resulting in its destruction and the release of energy, and the bearing capacity is reduced.
Figure 3.
Characteristics of elastic–plastic deformation zone of the coal pillar: (a) stress and elastic–plastic zone distribution under absolute width of the coal pillar; (b) stress and elastic–plastic general width of the coal pillar; (c) elastic–plastic distribution of stress under distribution limit width of the coal pillar; (d) stress and elastic–plastic under instability of the coal pillar.
3.2. Coal Pillar Stability Calculation and Analysis
As the coal body breaks, external stress is transferred inward in multiple directions. As the coal body undergoes destruction, its strength allows it to withstand multidirectional stress loading at a specific internal position, achieving equilibrium. This area represents the limit equilibrium zone of the coal body destruction. A small unit with a width of dx is extracted from the plastic zone in Figure 3, and its mass is disregarded. The equation, based on the limit equilibrium method, is derived from Figure 4:
 |
1 |
where C is the cohesion of the coal body (MPa); f is the friction coefficient of the contact surface between the coal seam and the roof and floor; σy is the vertical stress in the plastic zone (MPa); σx is the horizontal stress in the plastic zone (MPa); M is the thickness of the coal seam (m).
Figure 4.
Stress analysis of the plastic zone element.
After simplifying formula 1, the following equation is obtained:
 |
2 |
The limit equilibrium condition is
 |
3 |
where φ is the internal friction angle of the coal (deg).
By letting 
and substituting formula 3 into formula 2, the
following equation is obtained:
 |
4 |
Solving formula 4 and letting x = 0 and σx = P1, the following equation is obtained:
 |
5 |
where P1 is the resistance of the support to the coal side (MPa).
According to the stress distribution of the coal pillar in Figure 3, the maximum elastic stress is
 |
6 |
where K is the stress increase coefficient; γ is the average bulk density of the overlying strata (kN/m3); H is the burial depth of the coal seam (m).
Within a certain width from the edge of the coal pillar, the ultimate stress is equal to the abutment stress, and this width is the width of the plastic zone of the coal pillar. Formulas 5 and 6 are used to calculate the width of the plastic zone, denoted as X0:
 |
7 |
The element in Figure 5 is taken from the elastic zone in Figure 3, and from the horizontal force balance, it can be seen that
 |
8 |
Figure 5.
Stress analysis of the elastic zone unit body.
In the elastic zone, the following conditions hold:
 |
9 |
where λ is the lateral pressure coefficient.
Substituting formula 9 into formula 8 and considering the boundary conditions, the following equation can be obtained:
 |
10 |
Substituting σy = γH into formula 10, the influence distance L of the abutment pressure is obtained:
 |
11 |
According to the geological data for working face 14320, M is 9.2 m, H is 530 m, φ is 26°, C is 1.0 MPa, P1 is 0.1 MPa, γ is 25 KN/m3, f is 0.20, and K is 1.5. When these values are substituted into formula 7, the width of the plastic zone X0 is calculated to be 12.80 m.
When λ = 0.39 is substituted into formula 11, L is calculated to be 36.71 m. Therefore, the width of the coal pillar required to maintain absolute stability is 2L, that is, 73.42 m. The limit width of the coal pillar in the limit stable state is L; that is, 36.71 m. When the width is in the range of 36.71–73.42 m, it has a certain stability and can withstand certain stress changes.
For working face 14320 to maintain absolute stability to achieve safe mining, the coal pillar width must be greater than or equal to 73.42 m. From the width size of the irregular triangular coal pillars in Figure 2, it is concluded that the proportions of the widths that maintain absolute stability are 48.2, 31.2, 43.8, 51.5, and 40.5%. The percentages of coal pillars 1, 2, 3, and 5 are all less than 50%. Among them, the proportion of coal pillar 2 is the lowest, accounting for only 31.2%. Although coal pillar 4 accounts for more than 50% of the total area, it does not go beyond the range. Therefore, the possibility of coal pillar instability from greatest to least is as follows: coal pillars 2, 5, 3, 1, and 4.
The above formula derivation involves hypothetical reasoning based on the characteristics of stress transfer to the deep part of the coal body following the yielding of the coal wall under the influence of dynamic abutment pressure. It can be seen that the distribution characteristics of dynamic abutment pressure are related to factors such as the coal seam thickness, buried depth, and pressure measurement coefficient. However, in addition to the influence of dynamic abutment pressure, the stability of the coal pillar is also affected by static abutment pressure. When the dynamic and static abutment pressures are superimposed and analyzed, the stress state at this time is in line with the actual situation.
3.3. Coal Pillar Stability Safety Factor
3.3.1. Coal Pillar Strength Calculation
The width, height, and overall structural characteristics of the coal pillars determine their strength. The calculation method is mainly based on minefield investigation and laboratory tests. There are three classical formulas:23−25
-
(1)Bieniawski formula
where Sp is the strength of the coal pillar (MPa); σ1 is the uniaxial compressive strength (MPa), taking 9.709 MPa; b1 is the effective width of the coal pillar (m); h1 is the height of the coal pillar (m).
12 -
(2)UNSW formula
13 -
(3)Obert–Duvall formula
where σm is the critical uniaxial compressive strength of laboratory cylindrical specimens (MPa).
14
Bieniawski proposed the concept of “critical size” through the study of cubic specimens, and the critical strength of coal pillars obtained in the laboratory is as follows:
 |
15 |
where b2 is the diameter of the laboratory cylindrical specimen (m); h2 is the height of the cylindrical specimen in the laboratory (m).
In the laboratory, the cylindrical specimen has a diameter of 5 cm and a height of 10 cm. The uniaxial compressive strength of the specimen is 18.5 MPa, and by substituting the parameters into formula 15, the critical compressive strength is calculated as 13.081 MPa.
Due to the irregular shape of the coal pillar, the bearing capacity at the sharp corner is low and the bearing capacity at the wide and thick parts is high. To reasonably determine the strength and load of an irregular coal pillar, its effective width is used in the calculation, which is the maximum inscribed circle diameter.26Figure 6 presents a schematic diagram of the effective width.
Figure 6.
Diagram of the effective width of the irregular coal pillar.
The effective widths of continuous irregular triangular coal pillars are 118.72, 69.77, 86.40, 98.90, and 82.68 m, respectively. All parameters are then substituted into the above formula. The bearing strength of five irregular triangular coal pillars is calculated by formula 12: 51.32, 32.72, 39.04, 43.79, and 37.63; 28.33, 18.64, 21.55, 24.17, and 20.77 MPa are calculated by formula 13. The values of 47.65, 32.20, 37.45, 41.39, and 36.27 MPa are calculated using formula 14.
The applicable conditions of formula 12 are relatively wide. Formulas 13 and 14 are applicable to the range of 1–8 and 1–10, respectively. According to the effective width and height of the coal pillars, the width–height ratios are 12.9, 7.6, 9.4, 10.8, and 8.9, respectively. From the perspective of safety, the bearing capacity is selected to be small; therefore, the calculation results of formula 13 are relatively appropriate, but formula 13 obviously does not meet the requirements of the aspect ratio of irregular coal pillars here. Formula 14 yields a smaller result than formula 12, but formula 12 is chosen due to two width–height ratio ranges of coal pillars not meeting the requirements.
3.3.2. Average Load Calculation of Coal Pillar
According to Wilson’s two-zone constraint theory,27 when the width of the goaf on both sides of the coal pillar is greater than 0.6 times the average mining depth, the average load of the coal pillar is
 |
16 |
The average loads of irregular triangular coal pillars 1–5 were calculated to be 31.00, 43.45, 37.63, 34.55, and 38.73 MPa by substituting the parameters into eq 4.
3.3.3. Stability Coefficient of Coal Pillar
In the mining face, the internal stress of the coal pillar gradually changes, so its failure is a gradual process. The stability coefficient of the coal pillar can distinguish the stable state of the coal pillar (Table 1), and the applicable range includes regular and irregular coal pillars.28
Table 1. Coal Pillar State Discrimination Table.
| serial number | coal pillar stability coefficient | coal pillar state |
|---|---|---|
| 1 | 0 < k < 1 | unstable state |
| 2 | k = 1 | state of limit equilibrium |
| 3 | 1 < k < 1.5 | steady-state |
| 4 | k ≥ 1.5 | absolute stable state |
According to the determined irregular triangular coal pillar bearing strength Sp of working face 14320 and its average load p, the coal pillar stability coefficient k can be obtained as follows:29
 |
17 |
The stability coefficients k of coal pillars 1–5 are 1.66, 0.75, 1.04, 1.27, and 0.97. According to the coal pillar state discrimination table, coal pillar 2 is in an unstable state. Although coal pillars 3 and 5 are in a stable state, their stability coefficients are close to the limit equilibrium state.
4. Numerical Simulation of Stress Distribution in Continuous Irregular Triangular Coal Pillar Area
4.1. Constructing the Numerical Model
According to the geological conditions and mining conditions of working face 14320, RHINO6 software was used to establish a three-dimensional model, and FLAC3D software was used to calculate and analyze the stress changes in the stress concentration area of the irregular coal pillars during the mining process of the working face. The grid division of the model is shown in Figure 7, the length × width × height is 1300 m × 693.2 m × 130 m, and the Mohr–Coulomb yield criterion is adopted. Table 2 presents the distribution and mechanical parameters of each rock layer in the model. Vertical constraints are applied at the model point, horizontal constraints are applied around it, and the top is set as a free boundary. Finally, it is necessary to apply an equivalent load in the vertical direction at the top of the model to simulate the gravity of the overlying strata. Through calculations, it is determined that the equivalent load that should be applied to the top is 13.75 MPa, and the load applied to the top is 20.63 MPa.
Figure 7.
Three-dimensional mesh division of the model.
Table 2. Mechanical Properties of the Rock Strata.
| lithology category | density (kg·m–3) | bulk modulus (GPa) | shear modulus (GPa) | force of cohesion (MPa) | angle of internal friction (deg) |
|---|---|---|---|---|---|
| interbedded silty fine sandstone | 2700 | 3.72 | 3.41 | 5.79 | 35 |
| post office box stone | 2750 | 4.08 | 3.63 | 6.13 | 38 |
| medium-grained sandstone | 2810 | 4.21 | 3.82 | 6.97 | 39 |
| coal | 1900 | 3.93 | 1.83 | 2.52 | 29 |
| siltstone | 2600 | 3.26 | 2.85 | 5.27 | 34 |
| mudstone | 2382 | 3.07 | 2.54 | 6.23 | 37 |
4.2. Analysis of Simulation Results
According to the actual mining sequence, after the excavation of adjacent working faces 14317, 14318, 14319, and 7302, in the Baodian Coal Mine, the stress distribution cloud diagram before the working face is mined is shown in Figure 8. Afterward, working face 14320 is mined in stages I–V, and each 50 m of the adjacent irregular coal pillar is a stage. The corresponding mining distances of each stage are 220, 398, 644, 912, and 1043 m, and a total of eight stress monitoring points are set up at 25 m around the sharp corner of the coal pillar.
Figure 8.
Stress distribution cloud diagram of the working face before mining.
After conducting stagewise mining on the 14320 working face, we obtained the stress cloud map for stages I to V mining (see Figure 9). Figure 10 illustrates the stress changes at each monitoring point during the working face mining process derived from numerical simulation data. The local stress changes at monitoring points reflect the overall behavior of the irregular coal pillar. Initially, when the working face is not mined, the internal stress values for the five irregular coal pillars are 24.7, 27.5, 27.1, 24.5, 23.1, 28.3, 23.5, and 23.9 MPa, respectively, all below the theoretical values. This indicates stability, with stress changes occurring first in the irregular coal pillar closest to the mining face. As the mining face advances, the internal stress of the five irregular coal pillars in the 14320 working face gradually increases due to the influence of the goaf. At the completion of each mining stage, we determine the maximum stress when the 14320 mining face completely passes through the irregular coal pillar. The maximum stress values for the eight stress monitoring points during the mining process are 49.8, 32.1, 33.2, 39.4, 41.3, 42.1, 44.9, and 39.6 MPa. The theoretical bearing strengths of the irregular coal pillars are calculated as 51.32, 32.72, 39.04, 43.79, and 37.63 MPa, respectively. The simulated stress values of the irregular coal pillars align closely with the theoretically calculated data. This stress value primarily represents the peak abutment pressure on coal pillars. The stress concentration range of the irregular coal pillars gradually increases with the mining distance. Additionally, the stress concentration area predominantly exists in the edge and sharp-angle areas of the irregular coal pillar, confirming that the side of the coal pillar is influenced by the goaf, resulting in abutment pressure and subsequent stress concentration.
Figure 9.
Distribution cloud diagram of maximum principal stress in the different mining stages: (a) completion of stage I; (b) completion of stage II; (c) completion of stage III; (d) completion of stage IV; (e) completion of stage V.
Figure 10.
Stress change curves for the eight monitoring points.
5. Rock Burst Risk Assessment of Continuous Irregular Coal Pillar Area
5.1. Criterion for Judging the Degree of Rock Burst Risk
The coal pillar area produces stress concentration, especially the irregular shape of the coal pillar, which may be affected by the superposition of stresses in several directions. In addition, these coal pillars are prone to stress concentration when a sudden change in size occurs, so a rock burst is prone to occur near the coal body.30−32 In the past 20 years, about 60% of rock burst risks were caused by coal pillars, resulting in the destruction of working faces and roadways and a large number of casualties. Therefore, coal pillars are a key factor affecting rock burst risks. In the previous section, the actual stable state and stress changes of the coal pillar were analyzed. In this section, the possibility of rock burst is judged. The probability criteria for rock burst risks caused by each factor are as follows:
-
(1)
Safety size of coal pillar stability: B < L indicates a certain occurrence of a rock burst; B = L indicates the probable occurrence of a rock burst; 2L > B > L indicates the possible occurrence of a rock burst; B ≥ 2L indicates that the occurrence of a rock burst is impossible.
-
(2)
Coal pillar state: The instability state is certain; the limit equilibrium state is very likely; the steady state is possible; absolute stability is impossible.
-
(3)
The deviation between the simulated maximum principal stress and the average load stress of the coal pillar: If the deviation is greater than 20 MPa, it is certain that a rock burst will occur; if the deviation is 10–20 MPa, it is very likely that a rock burst will occur; if the deviation is 0–10 MPa, it is possible that a rock burst will occur; if the deviation is negative, it is impossible for a rock burst to occur.
5.2. Division of Roadway Rock Burst Risk Zone Affected by Coal Pillar Area
The probability of risk associated with the five continuous irregular triangular coal pillars is assessed using the criteria for rock burst risk, and the classification of rock burst risk in the roadways within the area affected by the coal pillars is determined based on the principles outlined in Table 3.
Table 3. Principles of Rock Burst Risk Classification.
| rock burst risk level | probability factor |
|---|---|
| strong rock burst risk | one certain probability and other probability superposition |
| three probable probability superpositions | |
| medium rock burst risk | one most likely probability superimposes with other probabilities |
| superposition of three possible probabilities | |
| weak rock burst risk | two possible probabilities are superposed with other probabilities |
During the mining period, the tailentry is affected by the coupling of the continuous irregular triangular coal pillars, and it is expected that 156.75 m of the roadway is a strong rock burst risk zone; 728.18 m of the roadway is a medium rock burst risk zone; and 176.88 m of the roadway is a weak rock burst risk zone (Figure 11).
Figure 11.
Working face tailentry rock burst risk zone evaluation results.
6. Prevention and Control of Rock Burst in Irregular Coal Pillar Area of Working Face
6.1. Prevention Measures of Rock Burst in Irregular Coal Pillar Area
Due to the large degree of stress concentration and uneven stress in the irregular coal pillar area, the disturbed area is prone to risks. Based on the risk level classification of rock burst in working face 14320, it is essential to implement prerelief measures in the irregular coal pillar area and prioritize regional prevention, followed by local prevention.33−36 Region rock burst prevention measures in the mining area include the use of a mining protective layer, optimizing the roadway layout, and judiciously determining the mining speed. As the rock burst prevention measures implemented here are part of the mining process for the working face, the roadway layout and mining methods have already been determined. Hence, scour prevention measures for controlling the mining speed of the working face are selected. Local scour prevention measures include coal seam drilling pressure relief, coal seam water injection, and more. In view of the complexity of roadway stress in the working face of the irregular coal pillar area and the economy of implementing scour prevention measures, large-diameter drilling pressure relief and roof hydraulic fracturing pressure relief are selected.
6.1.1. Reasonable Determination of Working Face Mining Speed
Due to the physical characteristics of the creep of the coal and rock mass, its own strength will change over time. Acceleration of the mining speed causes the stress to be insufficiently transmitted forward, and thus, the peak pressure increases and moves closer to the coal wall (Figure 12). Increasing the mining speed leads to a cumulative increase in the energy value in the coal and rock strata, and the elastic zone moves closer to the side of the roadway. Moreover, the faster the mining speed of the working face is, the longer the roof of the overlying rock is and the larger the pressure step is. In addition, the intensity and range of the roof and the breaking movement of the overlying rock are stronger than those for slow mining. The risk of a rock burst induced by rapid mining is greater. Therefore, for a working face with a rock burst risk, the mining speed should be reasonably controlled and slow uniform mining should be conducted.
Figure 12.
Distribution characteristics of the abutment pressure of coal under different mining speeds.
Figures 13 and 14 are the frequency statistics of mine earthquakes corresponding to different mining speeds in the 14319 working face of Dongtan Mine. Throughout the mining period of the 14319 working face, there were 131 mine earthquakes with a magnitude of 1.5 and above and 34 mine earthquakes with a magnitude of 2.0. The mining speed of the working face has a great correlation with the frequency of mine earthquakes. When the mining speed of the working face is 0.75–1.5 m/day, there is no mine earthquake above 2.0; when the mining speed of the working face does not exceed 2.25 m/day, the proportion of the frequency of mine earthquakes above 1.5 to the total frequency is 10.7%, and most of them are mine earthquakes within 1.5–2.0, and the proportion of the frequency of mine earthquakes above 2.0 to the total frequency is 2.9%. When the mining speed of the working face exceeds 3 m/day, the frequency of large energy mine earthquakes increases rapidly. Therefore, it is proved that the frequency of large energy mine earthquakes can be greatly reduced by controlling the mining speed, so as to reduce the possibility of rock burst in working face.
Figure 13.
Frequency of mine earthquakes above 1.5 corresponding to different mining speeds.
Figure 14.
Frequency of mine earthquakes above 2.0 corresponding to different mining speeds.
According to the experience of scour prevention in the Dongtan Coal Mine and in nearby mines under similar mining conditions in the past as well as statistical analysis of the mining speed and the number of large-energy microseismic events in working face 14318, it is concluded that when the mining speed is less than 3.0 m/day, the number of microseismic events accounts for only 7%. When the speed is 3.0–5.0 m/day, the number of microseismic events accounts for 22%. When the speed exceeds 5.0 m/day, the number of microseismic events accounts for 71% (Figure 15). Therefore, the reasonable scour prevention recovery speed is determined as follows:
-
(1)
It is predicted that when mining in a zone with weak rock burst, the daily advance distance should not exceed 5 m.
-
(2)
It is predicted that when mining in a zone with medium rock burst, the daily advance should not exceed 4 m.
-
(3)
It is predicted that when mining in a zone with strong rock burst, the daily advance should not exceed 3 m.
Figure 15.
Relationship between mining speed and large energy microseismic events in working face 14318 in the Dongtan Mine.
6.1.2. Prerelief of Large-Diameter Boreholes
6.1.2.1. Scour Prevention Mechanism of Large-Diameter Boreholes
The primary purpose of large-diameter borehole pressure relief is to utilize the elastic energy stored in the coal seam under high-stress conditions. This helps to disrupt the coal body around the borehole, alleviate stress in the coal seam, and mitigate the risk of rock bursts.37,38 The implementation of large-diameter drilling induces structural damage to the surrounding rock at a specific depth of the roadway, creating a weakening zone (Figure 16). This weakening zone alters the stress peak value and its distance from the roadway surface. Consequently, the stress condition in the rock burst’s limit equilibrium zone is significantly reduced, disrupting the conditions for rock burst occurrence.
Figure 16.
Drilling pressure relief principle diagram.
6.1.2.2. Design of Pressure Relief Parameters for Large-Diameter Borehole
At present, the commonly used layout methods are single-row layout and double-row layout.39 Considering the specific conditions of the roadway pressure relief area and the stress concentration area, the single-row layout is finally selected. The distance between the borehole and the roadway floor is about 1.2 m, and construction ahead of the working face should be at a distance of at least 350 m. Because the diameter of the drilling hole directly affects the pressure relief effect of the drilling hole, the diameter of the drilling hole is 150 mm. The drilling depth should be tailored to the stress concentration range and peak position of the coal body to be relieved. At the same time, the drilling depth must also be adapted to the thickness of the coal seam. Because the current coal seam thickness is 9.2 m greater than 8 m, the drilling depth is 25 m, and it is vertically constructed with the roadway side. Because one side of the mining roadway that needs to be relieved by drilling is the coal seam of the working face and the other side is a continuous irregular triangular coal pillar, the width of the triangular coal pillar will continue to change with the advancement of the mining work. When the width of the coal pillar is less than 30 m, the drilling depth is based on the 5 m coal body left in the adjacent goaf. When the coal pillar size is less than 10 m, no pressure relief drilling is constructed. The principle of determining the spacing of pressure relief boreholes is to ensure that the pressure relief areas around each borehole are interconnected to form a weakening zone. The spacing of pressure relief boreholes is generally 1–3 m. Spacing can be calculated using formula 18 and finalized after adjustment through the empirical analogy method.
 |
18 |
where D0 is the distance between pressure relief boreholes (m); n0 is the correction coefficient of the distance between pressure relief boreholes. For the weak rock burst risk area, it is 18.84, for the medium rock burst risk area, it is 12.56, and for the strong rock burst risk area, it is 6.28. d0 is the diameter of the pressure relief borehole (m); KE is the coal rock burst energy index.
According to the test results of coal samples taken from the field, KE is 3.678, which is substituted into the formula 18. According to the previous experience, the drilling spacing is finally determined as follows: the strong rock burst risk zone is 1 m drilling spacing, the medium rock burst risk zone is 2 m drilling spacing, and the weak rock burst risk zone is 3 m drilling spacing.
6.1.3. Roof Hydraulic Fracturing Prerelief
Roof hydraulic fracturing is a scour prevention technology that injects high-pressure liquid into the roof rock layer to generate new or expand existing cracks, control roof fractures, and release energy. Due to the presence of medium-grained sandstone with a thickness of 20.2 m in the main roof of the coal seam at the 14320 working face, the hanging roof phenomenon is likely to occur in the goaf behind the support of the working face and the coal pillar side of the goaf. This leads to a stress concentration in the coal and rock mass at the roadway, resulting in rock bursts. To prevent this, hydraulic fracturing is implemented in the roof to weaken it and provide pressure relief. Roof hydraulic fracturing has the advantages of simplicity, high safety, and strong adaptability (not limited by gas). The prerelief method of hydraulic fracturing in the gob-side entry is primarily used to sever the hanging roof of the irregular coal pillar goaf and presplit the hard roof of the working face. Presplitting the roof of the working face helps reduce advanced abutment pressure, lowering the dynamic load of the tailentry. Cutting off the hanging roof extending above the irregular coal pillar to the goaf reduces the lateral abutment pressure and the static load of the tailentry. The medium-grained sandstone with a thickness of 20.2 m is selected as the roof-breaking layer of the 14320 working face, with the crack initiation position in the middle and upper parts. The depth of the hydraulic fracturing borehole on the side of the working face of the tailentry is 18.9 m, with a dip angle of 90°. The drilling depth in the irregular coal pillar side is 20 m, and the dip angle is 72° (Figure 17). Because the irregular triangular coal pillar not only bears the pressure imposed by the goaf of working faces 14317 and 14318 and other working faces but also bears the superposition of the secondary lateral abutment pressure when working face 14320 becomes the goaf, under the action of high stress, the coal pillar will continue to deform rapidly. To effectively control the mine pressure appearance of the roadway, the spacing of the hydraulic fracturing holes in the continuous irregular side is designed to be 10 m. The spacing of the hydraulic fracturing holes on the side of the solid coal is designed to be 15 m so that the roof can break in time and not concentrate energy.
Figure 17.
Schematic diagram of hydraulic fracturing parameters.
6.2. Scour Prevention Effect Test
The drilling cuttings method, the mining stress method, and the microseismic method are used to monitor the rock burst risk in mining.40 After prerelief prevention and control, the prevention and control effect are analyzed according to the data obtained in actual mining.
6.2.1. Monitoring Using Drilling Cuttings Method
The drilling parameters are as follows: the drilling depth is 14 m, the drill diameter is 42 mm, and the hole spacing is 20 m. The data for eight boreholes near the stress monitoring points in the numerical model are analyzed. The critical coal powder amounts corresponding to the drilling depths of 1–5 and 6–14 m are 3.5 and 6.8 kg/m, respectively. The actual coal powder amounts of the eight boreholes are listed in Figure 18.
Figure 18.
Actual amount of pulverized coal in each borehole.
According to the data in Figure 13, the maximum amount of pulverized coal in the actual borehole corresponding to the two borehole depths of 1–5 and 6–14 m reaches 3.38 and 6.68 kg/m, which do not exceed the corresponding critical amount of pulverized coal. In addition, there is no obvious dynamic phenomenon such as sticking and suction during the borehole detection, and all of the indicators are normal. It is proven that the energy accumulated in the coal rock of the roadway with rock burst risk is released after the prerelief, and the risk of rock burst is reduced.
6.2.2. Mining Stress Method Monitoring
According to the mining conditions of 14320 working faces, the condition of high stress in the early stage of tailentry is affected by irregular triangular coal pillars. Therefore, after the implementation of prerelief and scour prevention measures, the stress online monitoring system is adopted in the inner side of tailentry to monitor the stress change in the area affected by abutment pressure in front of the working face, so as to monitor the occurrence of rock burst risk.
Throughout the mining period of the working face, the predicted rock burst risk area and a set of monitoring stations are arranged according to the distance between the strong rock burst risk zone not more than 10 m, the medium rock burst risk zone not more than 20 m, and the weak rock burst risk zone not more than 30 m. Stress sensors at each monitoring station group are drilled at depths of 8 and 14 m (14 m on the open-off cutting side and 8 m on the other side), with a 1 m interval. The drilling is 1.5 m away from the floor. The initial pressure of the sensor is not less than 5 MPa. Drilling is perpendicular to the coal wall and follows the slope of the coal seam. With the mining of the working face, it is continuously installed and extended outward and is installed and arranged 300 m ahead of the working face. Simultaneously, monitoring indicators and critical warning values are established (see Table 4).
Table 4. Early Warning Index Table of Stress Online Monitoring System.
| stressometer | initialization value | early warning index |
|---|---|---|
| shallow hole (8 m) | ≥5 MPa | ≥14 MPa |
| deep hole (14 m) | ≥5 MPa | ≥16 MPa |
| early warning of stress increase (stress increase within 24 h) | ≥4 MPa | |
By sorting out the online monitoring data of stress after mining in 14320 working face, the stress values of deep and shallow holes in working face are drawn into stress change curves (Figure 19). It can be seen that with the advance of mining in the working face, the stress value increases and decreases periodically, but most of the time it is below the stress warning value. It can be seen that the rock burst disaster of the working face is effectively controlled.
Figure 19.
Online variation of stress in deep and shallow holes during the mining process of 14320 working face.
6.2.3. Microseismic Monitoring
Microseismic monitoring is based on the principle of sound and energy produced during the failure process of the coal and rock masses. It records the microseismic waveform resulting from the fracture of coal and rock masses, determining essential parameters such as occurrence time, source location, and magnitude of microseismic events. This monitoring method aids in predicting the risk of rock bursts and evaluating the effectiveness of rock burst prevention and control measures.
The arrangement of microseismic geophones follows Figure 1, primarily monitoring vibration events during the mining of the working face roadway in the irregular triangular coal pillar area. Key monitoring indicators include the microseismic frequency and total energy. Referring to other working faces in Dongtan Mine, the critical early warning value is set at 1 × 104 J, indicating a risk of rock burst if exceeded.
Daily microseismic activity frequency and daily maximum energy values from the working face advancing length of 50–940 m during the mining process after the pressure relief of the 14320 working face are selected for statistical analysis, as depicted in Figure 20. Results show that after implementing prerelief and scour prevention measures the maximum daily microseismic energy remains low. There is a significant increase when influenced by the first and second squares of the working face, but it does not reach the critical early warning value. Most microseismic energy is less than 1 × 103 J, with frequencies above 10 times primarily associated with small energy vibrations. Additionally, Figure 21 illustrates that large energy events above 1 × 104 J, which can induce shocks, do not occur. The distribution of microseismic events is relatively scattered, indicating that the implementation of pressure relief measures not only reduces the energy accumulation degree of the surrounding rock of the roadway in the 14320 working face but also significantly diminishes the energy of the continuous irregular coal pillar area. Consequently, scour prevention measures effectively reduce the rock burst risk in the irregular coal pillar stress concentration area.
Figure 20.
Microseismic activity in the mining process of 14320 working face.
Figure 21.
Distribution of microseismic events after implementation of pressure relief measures.
7. Conclusions
-
(1)
The influence of abutment pressure influence distance L and coal pillar width B on the stability of the coal pillar is analyzed. The absolute stable width of the coal pillar is 2L. It is concluded that the possibility of instability of five irregular triangular coal pillars from large to small is coal pillars 2, 5, 3, 1, and 4.
-
(2)
The safety factor of coal pillar stability is determined by the ratio of the actual bearing strength of the coal pillar to the average load of the coal pillar. It is judged that coal pillars 1 and 4 are in a stable state, coal pillars 3 and 5 are in a limit equilibrium state, and coal pillar 2 is in an unstable state. At the same time, the numerical simulation analysis of the mining process of the irregular coal pillar working face is carried out. The stress of the coal pillar is less than the theoretical value before the mining of the working face. When the mining distance of the working face increases, the internal stress gradually increases, and it is verified that the maximum stress of the five coal pillars is basically consistent with the theoretical value.
-
(3)
The prediction criterion of rock burst in irregular coal pillar areas is established with the safety size of the coal pillar, the state of the coal pillar, the average load of the coal pillar, and the simulated stress deviation value as the evaluation index. The evaluation of the rock burst risk degree of roadway in working face is as follows: 156.75 m is a strong rock burst risk zone, 728.18 m is a medium rock burst risk zone, and 176.88 m is a weak rock burst risk zone. This prediction criterion only takes into account the factors of the coal pillar itself, does not involve the analysis of the external environmental impact, and has certain limitations. Therefore, in order to improve the accuracy of the prediction, it is necessary to increase the analysis of the geological conditions and even the mining technical conditions of the coal pillar in the subsequent research.
-
(4)
The scour prevention measures combining regional prevention and local prevention are put forward. Considering the specific conditions of the working face, the measures of reducing the mining speed, large-diameter borehole pressure relief, and roof hydraulic fracturing pressure relief are adopted, and the scour prevention effect is comprehensively tested by the drilling cuttings method, mining stress monitoring method, and microseismic monitoring method. According to the actual data, it is verified that the pressure relief effect is good, which can effectively solve the problem of rock bursts in the irregular coal pillar area of the working face. The above prevention measures are to reduce the internal stress of the coal body. In addition, it can also be considered to increase the scour prevention capacity of the working face roadway and improve the roadway support capacity to achieve the prevention and control of rock burst risk.
Acknowledgments
We thank LetPub (www.letpub.com) for its linguistic assistance during the preparation of this manuscript.
The authors declare no competing financial interest.
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