Mine Stability and Water Inrush in Coal Mine

Abstract

This chapter describes three major parts associated with mining stability and mine water inrush: The first part discusses the stability issues related to rock burst in deep underground coal mines, the second part studies the height variations of water-conducting fractured zone in shallow underground coal mines, and the last part presents a numerical study on the mechanisms of a major water inrush incident occurred in China in 2010 in which 32 people died.

1 Introduction

With the worldwide economic and social development , the shallow mineral resources are gradually depleted , and the depth of mining or tunnelling is increasing (Tan et al. 2015; Zhao et al. 2016). Under the influence of high stress and high water pressure, the deformation and failure modes of deep rock mass are different from that of shallow one, and large scale fracturing becomes one of the dominant characteristics (Cai 2013; Tan et al. 2017). Phenomena of rock burst, large deformation, sudden water inrush etc., increase in deep exploitation. Case studies of rock burst and water inrush can build a foundation for the mitigation and control of these risks. In addition, the height variations of water-conducting fractured zones are of significant importance for the prevention of water inrush induced disasters in longwall mining. In this chapter, some typical cases in coal mines are chosen to demonstrate the application of knowledge and predicting tools of rock fracturing to rock burst and water inrush disaster.

For rock bursts, we choose Muchengjian Coal Mine of Jingxi Coalfield, which has witnessed more than 30 rock bursts during the past 10 years. The mine has complex geological structures, with notable faults, folds, and tectonic movement affected areas. One typical mining area is selected for this study. Firstly, we present the rock burst distributions, the related spatial and temporal characteristics, and the geological characteristics. Secondly, we propose a rock burst mechanism through the analysis of these representative rock bursts. Thirdly, on the basis of this mechanism, we propose a progressive mitigation method of rock bursts (PMMRB). And finally, we validate the PMMRB method by field application in a similar mining area.

Regarding the height variations of water-conducting fractured zones, it begins with the establishment of numerical models of No. 1101 coal face at Luxin Mine, China, using FLAC3D. We then calculate the damage zone using numerical models and obtain the heights of water-conducting fractured zones for different face lengths and mining thicknesses. At last, the optimal mining method, face length and mining thickness of No. 11 coal seam are determined.

The chapter also discusses a detailed water inrush case study. A coupled model is designed using FRACOD to investigate the mechanisms of water inrush disaster (known as “Luotuoshan accident”) occurred in China in 2010 in which 32 people died. This model applies the fracture-hydraulic flow coupling function, and focuses on the rock fracturing processes in a karst collapse column, which is a geologically altered zone linking several rock strata vertically formed by the long-term dissolution of the flowing groundwater. The study is based on the actual geological conditions of Luotuoshan mining area, and various materials with actual geological characteristics are used to simulate the rock fracturing processes surrounding the coal seam. The influences of several key factors for this water inrush, such as in-situ stresses, fractures in the formation and development of water inrush channels, are investigated after roadway excavation.

2 Progressive Mitigation Method of Rock Bursts under Complicated Geological Conditions

2.1 Engineering Geological Characteristics and in Situ Stress Field of Jingxi Coalfield

2.1.1 Overview of Jingxi Coalfield

Jingxi Coalfield is located in Mengtougou and Fangshan Districts in Beijing with a length of 45 km from east to west and a width of 35 km from south to north. The mineable zone is about 1019 km². This area has three major coal-producing underground mines, including Muchengjian Coal Mine, Datai Mine, and Da’anshan Mine. The rock formation in this area is under compression from the southeast to northwest (Han et al. 2014). The area also has five major syncline folds and one large-scale fault. From west to east, these structures are the Baihuashan, Tiaojishan-Miaoanling, Jiulongshan-Xiangyu, Shijingshan and Beiling synclines, and Babaoshan Fault, as shown in Fig. 15.1 (Wang et al. 2017). Because of these coal-bearing synclines influenced by tectonic movement and sedimentary environment, the dip angle of strata is steep in the southeast and is relatively flat in the northwest. Many subsidiary folds exist, and the direction and dip angle of strata varies significantly. In addition, different kinds of fault tectonics exist, which were caused by a pushing force from the southeast. The formation of Jingxi Coalfiled is influenced not only by the late secondary structures, but also by connate deposit. The tectonic areas with geological changes, such as the change of coal seam thickness, coal seam dip angle, and coal quality (as shown in Fig. 15.2) are formed easily by different deposition velocities, various oscillation frequencies of the crust, and diverse sedimentary environments. All of these factors lead to the complicated geological conditions in Jingxi Coalfield. The Muchengjian Coal Mine is a typical representation of this complexity.

Fig. 15.1

Geological structure sketch map of Jingxi Coalfield (Wang et al. 2017)

Fig. 15.2

Tectonic areas with facies change in Jingxi Coalfield. (a) Change of coal seam thickness (b) Change of coal seam dip angle and coal quality

2.1.2 Engineering Geological Characteristics of Muchengjian Coal Mine

Figure 15.3 illustrates the main part of the geological structures of Muchengjian Coal Mine, and the locations of the two research areas: West No. 1 Ming Area of + 250 Level (W1 Area) and No. 1 Mining Area of + 150 Level (M1 Area). This mine has about 11 coal seams, and the main mining coal seam is the No.3 coal seam, as shown in Fig. 15.3b. The folds and faults seriously influence these two research areas, causing sharp changes in coal seam thickness, coal seam dip angle and coal quality. The north of W1 Area is located at the bottom of a synclinal fold, and two large faults exist near the south boundary. The average mining depth of W1 Area reaches more than 700 m. M1 Area also is located across that synclinal fold, and its average mining depth reaches more than 900 m. Longwall technology was adopted in these two mining areas.

Fig. 15.3

Schematic map of geological structures of Muchengjian Coal Mine. (a) Part of the geological occurrence of No.3 coal seam (b) Tectonic section of No.8 prospecting line

The exploration boreholes reveal that the coal seam thickness in W1 Area varied from 1.0 to 2.5 m with an average of 1.8 m. The dip angle varied from 3° to 20°, and the average value was 15°. Table 15.1 lists the lithological characteristics of the roof and floor (site details of M1 Area are presented in Sect. 13.2.4).

Table 15.1 Lithological characteristics of roof and floor in W1 Area

2.1.3 Stress Field

The in situ stress field is a fundamental factor causing dynamic disasters in underground mining, especially when mining depth increases or a geological structure exists. Thus, we established three measurement stations to measure the in situ stress, one near the synclinal axis, one near W1 Area, and one in W1 Area, as shown in Fig. 15.3a. We used the overcoring method of hollow inclusion in this test. Table 15.2 lists the test results. According to the field test, the primary characteristics of the geo-stress field are summarized as follows:

• The maximum principal stress increases with the increase of depth, except for the No.3 measurement station where the stress value is the highest (28.7 MPa) although the buried depth is only 581 m. Obviously, this stress is caused by the fold syncline, which is an area of high in situ stress concentration.

• The dip angle of the maximum principal stress varies from −19.3° to 0.4°. This result means that the direction of in situ stress is near horizontal. Thus, the tectonic stress is the main stress field in the Muchengjian Coal Mine.

• The azimuth angle of the maximum principal stress is primarily within the range of 182.6°–267.9°, indicating that the stress direction changes slightly. The direction of the maximum principal stress is basically perpendicular to the synclinal axial, especially for the No.1 and No.2 measurement stations.

• In W1 Area, the median principal stress σ₂ is nearly 19.6 MPa, which is almost equal to the gravity stress. The minimum principal stress σ₃ is 9.8 MPa in the horizontal direction with a dip angle of −13.2°. The dip angle of the maximum principal stress σ₁ is −19.3°. Therefore, the W1 Area is in the plane of σ₁ and σ₃.

Table 15.2 Test results of in situ stress

2.1.4 Coal Burst Liability

Coal burst liability (CBL) is an inherent property of coal, which arises with an increase of stored strain energy in the coal seam. It is the intrinsic factor causing rock bursts. Researchers have proposed various CBL indexes based on energy, stiffness, strength, and failure duration. The standard CBL indexes widely used in China include the uniaxial compressive strength (R_C), elastic strain energy index (W_ET), bursting energy index (K_E), and dynamic failure duration (D_T). Table 15.3 lists CBL classifications according to the Chinese standard (GB/T 25217.2-2010, 2010) and their calculations.

Table 15.3 Index values of CBL classification in the Chinese standard

The CBL indexes were measured through coal samples obtained from W1 Area and M1 Area, and Table 15.4 lists the test results. These results indicate that the No.3 coal seam has a strong CBL, meaning that it has the capacity to store energy and is prone to dynamic breakage.

Table 15.4 Test results of CBL

2.2 Description and Analysis of Rock Bursts

These rock bursts are graded according to the weight of ejected coal, as shown in Table 15.5. Three weak bursts (“2.28”, “3.6”, and “8.26” events) and two strong bursts (“6.28” and “8.29” events) occurred in W-No. 1 working face, one median burst (“5.24” event) and one strong burst (“11.10” event) occurred in E-No. 1 working face, one weak burst (“12.18” event) and one strong burst (“12.4” event) occurred in W-No. 4 working face, and two strong bursts (“4.16” and “7.8” events) occurred in W-No. 5 working face.

Table 15.5 Grade statistics of rock bursts

The data in Table 15.5 shows that both the occurrence probability and damage intensity of rock bursts in the stoping process are greater than that in the tunneling process. Moreover, the six strong rock bursts, accounting for 55% of the total bursts, all occurred in the stoping process. Figure 15.4 shows that every working face has one or two strong bursts. All of the strong bursts occurred just after destress blasting except the “6.28” event in W-No. 1 working face. This phenomenon indicates that blasting-induced disturbance has a significant influence on the occurrence of rock bursts. Thus, we select these six strong rock bursts for further analysis and discussion in this section.

Fig. 15.4

Rock burst distributions, geological characteristics and other information of the working faces. (a) W-No. 1 working face (b) E-No. 1 working face (c) W-No. 4 working face (d) W-No. 5 working face

2.2.1 “6.28” and “8.29” Events in the W-No. 1 Working Face

During early recovery on June 28, 2009, a rock burst event occurred in the headgate and the longwall face when stoping the position where the longwall face was about 140 m away from the uphill roadway. From the north end of the longwall face to the east about 17 m, the headgate had a south rib spalling 0.8–1.2 m wide; from the north end to south about 16.5 m, the longwall face had a rib spalling 0.4–1.0 m wide. The “6.28” event occurred when stoping the area where the coal seam thickness fluctuated from 3.5 to 2.2 m, as shown in Fig. 15.4a. Stoping from the thick to the thin area was not advantageous for rock burst control, because a high stress concentration area could form easily and relatively large elastic energy was stored (Zhao et al. 2016). Another factor causing stress concentration was the fault just across this section. Without any blasting disturbance, normal mining disturbance activated this event. Three obvious causes, including the change of coal seam thickness, fault, and mining activities, led to the “6.28” event.

During the final exploitation process, another strong burst event occurred on August 29, 2009. The distance between the longwall face and the uphill roadway was only 28 m, as shown in Fig. 15.4a. After the “8.29” event, the height of the 48 m length uphill roadway was reduced to only 0.8–1.2 m because the entire west rib had been pulled out. In addition to the inevitable mining-induced stress, the “8.29” event was apparently related to the coal pillar formed during the end-stoping process. This coal pillar was located just between two faults. Therefore, in the “8.29” event, the high stress concentration was caused by the coal pillar and the faults, and the blasting disturbance from the south side of the longwall face was the inducing factor.

2.2.2 “11.10” Event in the E-No. 1 Working Face

On November 10, 2009, an intensive rock burst event was activated by a blasting disturbance from the south side of the longwall face during the final exploitation process. The longwall face, the tailgate, and the uphill roadway were all damaged, as shown in Fig. 15.5. The coal ribs of the uphill roadway between the middle tailgate and tailgate had pulled out about 1.2–1.5 m in width, and the roadway’s height was reduced to less than 0.8 m, as shown in Fig. 15.5a. The height of the tailgate from the tailgate throat to the longwall face became only about 1.0–1.2 m, as shown in Fig. 15.5b. The longwall face suffered rib spalling of about 1.0 m in width, as shown in Fig. 15.5c.

Fig. 15.5

Photos of roadway’s damage after the “11.10” event in the E-No. 1 working face. (a) Uphill roadway (b) Tailgate (c) Longwall face

When “11.10” event occurred, the distance between the tailgate and the uphill roadway was only 41 m. From Fig. 15.4b, it is clear that the “11.10” event occurred when stoping the isolated coal pillar. In addition, two other coal pillars existed next to this area, one in the west (10 m width) and the other in the south (4–10 m width). This area did not have any surrounding geological structures. Thus, the isolated coal pillar was the main cause for the “11.10” event.

2.2.3 “12.4” Event in the W-No. 4 Working Face

During the stoping process, a strong burst occurred in the tailgate of the W-No. 4 working face on December 4, 2008, when the north side of the longwall face was destressed by a blasting method, as shown in Fig. 15.4c. This event caused a rib spalling of 13 m long on the south sidewall of the tailgate, beginning 5 m away in front of the longwall face. Figure 15.4c shows that a fault and a coal pillar were near the damaged section, both of which could have led to high stress concentration. Therefore, a remote blasting disturbance is believed to have induced the “12.4” event.

2.2.4 “4.16” and “7.8” Events in the W-No .5 Working Face

While stoping the W-No. 5 working face, two strong rock burst events occurred on April 16 and July 8, 2013. The distances between the headgate and the uphill roadway were 171 m and 130 m, respectively, as shown in Fig. 15.4d. Both events were activated by the remote destress blasting. In the “4.16” event, the headgate experienced a rib spalling of 0.8–1.2 m wide, starting from 13 m ahead of the longwall face; in the “7.8” event, the north sidewall of the headgate of 79 m long had a rib spalling of 0.3–1.2 m wide.

The W-No. 5 working face did not have any coal pillars or obvious geological structures. The serious damaged sections of these two events were located in or near the peak region of the abutment pressure, which could not be avoided. The two significant external causes in these two events are that (a) the W-No. 5 working face was just located in the core of a synclinal fold (a high in situ stress concentration area) and that (b) the stoping direction was subparallel to the synclinal axis.

2.2.5 Rock Burst Mechanism

This analysis indicates that the high stress concentration is primarily controlled by the mining engineering condition (i.e., coal pillar, island coal face, and final exploitation process, which cause high mining-induced stress) and the geological condition (i.e., mining depth, fault, fold, coal seam thickness change, and coal seam dip angle change, which cause high in situ stress). The rapid release of stress concentration is induced by the disturbance stress, such as blasting and mining activities. Thus, from the stress point of view, the occurrence of rock bursts is generally controlled by three major factors: the in situ stress and the mining-induced stress leading to high stress concentration, and the disturbance stress. The stress criterion for rock bursts occurrence can be expressed as Eq. (15.1):

$$ {\sigma}_i+{\sigma}_m+{\sigma}_d>\left[\sigma \right] $$

(15.1)

where σ_i is the in situ stress, σ_m is the mining-induced stress, σ_d is the disturbance stress, and [σ] is the critical stress required for rock burst formation. Equation (15.1) indicates that it is possible to control rock bursts by reducing σ_i, σ_m, and σ_d.

The rock burst mechanism can be classified based on the main causes for high stress concentration, as shown in Fig. 15.6. The three types are I (“high in situ stress” rock burst), II (“high mining-induced stress” rock burst), and III (“high in situ stress plus high mining-induced stress” rock burst). The explanations are as follows:

• For I (“high in situ stress” rock burst), when tunneling or stoping approaches the geological structure (high in situ stress concentration areas) and small mining induced stress is added, the critical stress for rock burst formation can be reached, such as the “6.28” event, “4.16” event, and “7.8” event.

• For II (“high mining-induced stress” rock burst) in the absence of high in situ stress concentration, the critical stress can also be reached easily because of the influence of coal pillar, island working face, and end stoping process, such as the “11.10” event.

• For III (“high in situ stress plus high mining-induced stress” rock burst), both the geological structure and the coal pillar contribute to reaching the critical stress, such as the “8.29” event and “12.4” event.

Fig. 15.6

Schematic drawing of rock burst mechanism in Jingxi Coalfield

According to this analysis, the rock bursts occurrence in Jingxi Coalfield can be described as follows. When a coal seam with burst proneness experiences high stress concentration caused by the superposition of in situ stress and mining-induced stress, an unsteady dynamic balance condition will be reached. If this balance is influenced by external disturbance, a rock burst probably occurs. The damage of the rock burst is directly related to the released energy.

2.3 The Progressive Mitigation Method of Rock Bursts – PMMRB

The characteristics and occurrence mechanisms of rock bursts in W1 Area during the stoping process indicate that the disturbance stress, especially the blasting induced stress, greatly influences the occurrence of rock bursts. Yan et al. (2015) found that rock bursts may be a result of instantaneous dynamic extension of rock cracks induced by multiple superimposed disturbances of the blast-induced loading stress wave, unloading stress wave and stress wave induced by the rapid in situ stress redistribution. Destress blasting was adopted to mitigate the rock burst in W1 Area, which could be the cause of some rock burst events here. Therefore, before using the blasting method, new destress techniques, which have weak disturbance, need to be applied to mitigate the in situ stress concentration and the mining-induced stress concentration. The two widely used destress techniques, including the large-diameter drilling and water infusion, are good choices. Water infusion of the coal specimens can decrease their compression strength, elastic modulus, bursting energy index and pre-peak accumulated energy in different degrees (Su et al. 2014), which can mitigate the CBL. For example, Díaz Aguado and Nicieza (2007) found that water infusion constitutes an improvement in the mining sequence and in safety in two different coal mines influenced by coal and gas outbursts. Large-diameter drilling also can mitigate the rock burst because an X-shape area of plastic deformation will be formed around the drilling hole (Zhu et al. 2009). The main functions of large-diameter drilling are to reduce elastic modulus, increase the deformability, and absorb the energy of rock mass. This method was applied successfully in the No. 1302 Mining Face of Xinjulong Coal Mine to prevent rock burst (Liu et al. 2014). Therefore, large-diameter drilling and water infusion are good choices for rock burst prevention with weaker dynamic load than destress blasting.

According to the rock burst mechanism in Jingxi Coalfield, two kinds of large diameter drilling schemes are designed and water infusion is added to the stoping process, as shown in Fig. 15.7a. The borehole length is 10–15 m, the borehole diameter is 110 mm, and one row of boreholes is located in the middle of each coal sidewall of the gateroad. One obvious difference exists between the two schemes: Scheme I has one borehole in every drilling site to mitigate “high in situ stress” concentration or “high mining-induced stress” concentration, whereas Scheme II has three boreholes in every drilling site to mitigate “high in situ stress plus high mining-induced stress” concentration. The second difference is the drilling spacing, which is 5 m for Scheme I, and 9 m for Scheme II. Additionally, boreholes vertical to the coal wall are also used for water infusion whose length is about 50–70 m in front of the longwall face, and water infusion will be stopped when water flows out.

Fig. 15.7

Destress schemes for large-diameter drilling and water infusion, and PMMRB. (a) Schemes of large-diameter drilling and water infusion to mitigate rock bursts during the stoping process; (b) PMMRB

We introduce the progressive mitigation method of rock bursts (PMMRB), as shown in Fig. 15.7b. This method include three steps: Firstly, carefully survey the geological and engineering characteristics of the working face and then estimate the kind of stress concentration each part of the working face belongs. Secondly, according to the stress concentration classification, the scheme of the large-diameter drilling and water infusion is chosen. Secondly, estimate the destress effect through the electromagnetic radiation (EMR ) method and drilling cutting method. Thirdly, if the coal wall still has stress concentration, apply destress blasting until the stress concentration is released. The parameters for destress blasting are the same as the original parameters and are given in Fig. 15.8. The critical advantage of PMMRB is that the strength and the burst liability of coal are weakened by water infusion and crack zones are formed by large diameter drilling before applying destress blasting. Not only will the stress concentration improve but also the weak and crack zones will absorb the blasting impact. Therefore, the risk of rock burst can be mitigated more effectively, and the rapid release of stress concentration induced by disturbance stress can be avoided. Notably, PMMRB is established to add weak disturbance destress techniques without optimizing blasting parameters.

Fig. 15.8

Technique parameters of destress blasting

2.4 Practical Experience

2.4.1 Site Description of M1 Area

The selected mining area (M1 Area) adjacent to the east of W1 Area is located at the bottom of a fold, as shown in Fig. 15.3a. The area also contains faults and regions of coal seam thickness change. The average mining depth reaches more than 900 m (the mining depth of No. 1 and No. 2 working faces is about 1000 m), which is deeper than that of W1 Area, indicating that the in situ stress in M1 Area are higher. Figure 15.3b shows the geological and mining engineering characteristics of M1 Area in detail. Longwall technology is adopted for the underground mining. The coal seam thickness varies from 1.6 to 4.0 m, and the average thickness is 3.4 m. The dip angle varies from 18° to 40°, and the average value is 20°. Table 15.6 lists the lithological characteristics of roof and floor.

Table 15.6 Lithological characteristics of roof and floor in M1 Area

2.4.2 Stress Concentration Estimation and Destress Scheme Selection

Before exploiting M1 Area, the stress concentration was estimated according to the geological and mining engineering characteristics. The estimated results are as follows:

• The No. 1 working face is located near the core of the synclinal fold and is an area of coal seam thickness change. Meanwhile, the working face can be regarded as a large coal pillar because of its size, which could easily lead to high mining induced stress. Thus, the No. 1 working face is classified as “high in situ stress plus high mining-induced stress” concentration.

• The north side of the No. 2 working face is an area of coal seam thickness change, and there are also several faults. Additionally, this is influenced by the coal pillar and abutment pressure all formed by the No.1 working face. Thus, the north side is “high in situ stress plus high mining-induced stress” concentration, and the other two sides are “high in situ stress” concentration or “high mining-induced stress” concentration.

• The north side and the south side of the No. 3 working face are mainly influenced by “high mining-induced stress”, which is generated by stoping the No. 2 working face. The west side, however, is influenced by “high in situ stress” because of two faults (Note: The coal seam pinch-out area was not found until the stoping passed this station).

• The upper-dip part of the north side of the No. 4 working face is “high in situ stress plus high mining-induced stress” concentration because of the influence of two faults and the exploitation of the No. 3 working face. The down-dip part of the north side and the other two sides are “high in situ stress” or “high mining-induced stress” concentration.

As shown in Fig. 15.9, different destress schemes are applied according to their stress concentration classification. Scheme I destressed is marked with a black oval, and Scheme II destressed with a blue oval.

Fig. 15.9

Geological and mining engineering characteristics of M1Area, and application of PMMRB

2.4.3 Prevention Effects from Applying PMMRB

The application of PMMRB has achieved good results. The most noteworthy is the excellent results of the No. 1 and No. 2 working faces where no bursts occurred, especially when the 1000 m mining depth, the fold, faults, areas of coal seam thickness change, and coal pillars all played a role in the stress concentration of the two working faces. The electromagnetic radiation method (EMR) monitoring results of these two working faces are shown in Fig. 15.10 (No. 1 working face) and Fig. 15.11 (No. 2 working face). These results show that the EMR intensity was kept low in the whole stoping process, which varied from 7.8 to 34.9 mV.

Fig. 15.10

EMR monitoring results of the No. 1 working face in 2015. (a) January (b) February (c) March

Fig. 15.11

EMR monitoring results of the No. 2 working face in 2015. (a) April (b) May (c) June (d) July (e) August

In addition, Figs. 15.10 and 15.11 also revealed that the EMR intensity value was high during the final exploitation process. For example, Fig. 15.10 illustrates that the EMR intensity varied from 7.8 to 20.6 mV in January and February, but in March, during the late recovery period, the variation range was 12.1–24.5 mV; Fig. 15.11 illustrates that the EMR intensity varied from 9.6 to 25.9 mV in April, May, June, and July, but in August, the variation range was 15.1–34.9 mV. Obviously the mining-induced stress was high in the late recovery period. From Fig. 15.11, it can also be seen that the EMR intensity was larger during the early and final exploitation processes than during the middle exploitation process. For example, the EMR intensity varied from 13.1 to 25.9 mV in April and from 15.1 to 34.9 mV in August, but in May and June, the variation range was only 9.6–21.7 mV. Moreover, destress blasting method was also applied when the EMR intensity showed an increasing trend, and no bursts were induced. That is to say, the application of destress methods with weak disturbance might be a good way to absorb the blasting load wave.

Although two rock bursts occurred during the final exploitation process of the No. 3 working face, it does not mean that PMMRB failed. Figure 15.9 shows that an area of coal seam pinched out near the events’ site had high in situ concentration. Superimposed with the high mining-induced stress caused by the end exploitation, this area was “high in situ stress plus high mining-induced stress” concentration, but stress Scheme I was applied, which led to an insufficient release of stress concentration. If Scheme II had been applied, the two bursts might have been controlled or mitigated.

2.5 Discussions

Although PMMRB has a good practical use, it still has a few shortcomings. For instance, both the stress estimation and the corresponding destress schemes could be classified more carefully because various factors can cause high stress concentration. We introduce PMMRB roughly without placing too much consideration on parameter optimization. Thus, although PMMRB is initially established, more work needs to be done for improvement. Firstly, the rock burst mechanism can be classified in more detail based on the three types of rock burst mechanism presented in this section. For example, Scheme I (“high in situ stress” rock burst) induced by geological structures could be classified into fault slip burst, fold burst, and coal thickness change burst because of variations in the main affecting factors (Han et al. 2014; Zhao et al. 2016). According to this classification, we can investigate and understand the rock burst mechanism further. Secondly, the mitigation effect of the same destress scheme can be applied to different kinds of rock bursts. According to a more detailed classification of rock burst mechanisms, we can study the most suitable destress method as well as the related parameter optimization for each type of rock burst mechanism. At last, precursory information can be applied to different kinds of rock burst and different kinds of destress mitigation methods. Thus, we can take more targeted actions by investigating the relationships among the various types of rock bursts.

2.6 Conclusions

In this section, we studied the rock burst mechanism based on rock burst cases in Muchengjian Coal Mine. We introduced PMMRB based on the rock burst mechanism, and we discussed and analyzed the practical experience of PMMRB. Primary conclusions are as follows:

• 1. In addition to the property of the strong coal burst liability (CBL) of coal seam, the occurrence of rock bursts is caused by the superposition of three kinds of stresses: in situ stress, mining-induced stress, and disturbance stress. Thus, the analysis of rock bursts accompanied by careful study of the geological and mining engineering characteristics indicates that mining depth, fault, fold, and coal seam thickness change can lead to high in situ stress concentration; and e coal pillar, island mining face, and final exploitation process can lead to high mining induced stress concentration. In addition, the rapid release of stress concentration is induced by the disturbance stress, such as blasting and mining activities.

• 2. The rock burst mechanism in Jingxi Coalfield can be classified into three types based on the main causes for the high stress concentration: I (“high in situ stress” rock burst), II (“high mining-induced stress” rock burst), and III (“high in situ stress plus high mining-induced stress” rock burst).

• 3. On the basis of this classification and without considering the optimization of blasting parameters, we recommended the PMMRB method of weakening the strength and burst liability of coal seam by water infusion and of expanding cracked zones by applying large diameter drilling before using destress blasting. The two advantages of PMMRB are (1) reducing the stress concentration in the mine structures and (2) absorbing dynamic stress concentrations from blasting. Practical experience from applying PMMRB in Jingxi Coalfield showed that the method can effectively mitigate rock bursts, and that large diameter drilling and water infusion are effective techniques to reduce rock burst under complicated geological and mining conditions. The monitoring results from electromagnetic radiation (EMR ) measurements also revealed that the EMR intensity is strong during the final exploitation process of a working face.

3 The Height of Water-Conducting Fractured Zones in Longwall Mining of Shallow Coal Seams

3.1 Geological and Mining Conditions

Numerical models were built based on No. 11 coal seam at Luxin Mine, China. No. 1101 coal face is the first face of No. 11 coal seam in Luxin Mine. The depth of No. 1101 coal face ranges from 168 to 256 m, with an average of 218 m. The thickness of the coal seam ranges from 0.6 to 20.9 m, with an average of 10.76 m. The dip angle ranges from 3° to 5°, with an average of 4.1°. The coal seam is simple in geological formation and rich in joint development. The main roof is sandy mudstone with thickness ranging from 29.8 m to 48.2 m, and the immediate roof is also sandy mudstone with thickness ranging from 18.9 to 31.7 m. Characteristics of roof and floor are shown in Table 15.7. There is a water-rich aquifer in the strata overlying the coal seam, and its distance from the coal seam ranges from 70 to 93.6 m.

Table 15.7 Characteristics of roof and floor of No. 1101 coal face

3.2 Model Building

3.2.1 Model Size and Grid

Based on “Pillar design and mining regulations under buildings, water, rails and major roadways” guidelines in China, the height of the water-conducting fractured zone can be obtained from the following empirical formula (National Bureau of Coal Industry of China 2000).

$$ {H}_{\mathrm{li}}=\frac{100\sum M}{1.6\sum M+3.6}\pm 5.6 $$

(15.2)

where H_li is the height of the water-conducting fractured zone, m; M is the mining thickness, m.

The calculated heights of the water-conducting fractured zone of No. 1101 coal face range from 46.1 to 57.3 m. According to the mining practices, the height of the water-conducting fractured zone does not increase further when the working face advances more than 300 m. To ensure that there are no significant boundary effects, two groups of numerical models are built, group A and group B. The size of the numerical model in group A (length × width × height) is set to 2 l (l = face length) × 300 × 200 m and the vertical distance below the seam is set to 30 m. The depths of No. 11 coal seam in both groups are set to 218 m. The size of the numerical model in group B is set to 3 l × 300 × 220 m and the vertical distance below the seam is set to 50 m. Both the haulage way and tail roadway are excavated in the coal seam along the floor of the coal face, with 4 m in width and 3 m in height.

Cuboid mesh is used to create the FLAC3D grid. To improve the accuracy, the elements of coal, immediate roof, main roof and floor should be small, the side length of which is less than 2.0 m. The side length of other elements ranges from 3.0 to 5.0 m. The model grid of group A is shown in Fig. 15.12.

Fig. 15.12

Grid for No. 1101 coal face model in group A

3.2.2 Boundary Conditions

Displacement boundary condition is applied on the boundaries of the model (Itasca Consulting Group, Inc., 1997). X-velocity is set to be zero at both left and right boundaries. Y-velocity is set to zero at both front and back boundaries. X-velocity, y-velocity and z-velocity are all set to be zero at bottom boundary. The top boundary is free.

3.2.3 Initial Conditions

To reach the equilibrium state rapidly, the vertical and horizontal stress are initialized to σ_v and σ_h, respectively, where σ_v is obtained from Eq. (15.3) and σ_h is obtained from Eq. (15.4) (Itasca Consulting Group, Inc. 1997).

$$ {\sigma}_v=\gamma h $$

(15.3)

$$ {\sigma}_h=\frac{\mu }{1-\mu }{\sigma}_v $$

(15.4)

where σ_v is the vertical stress based on gravity loading of overlying strata, Pa; σ_h is the horizontal stress due to self-loading following Poisson’s effect, Pa; γ is bulk density of rock masses, N/m³; h is the depth of coal seam, m; μ is the Poisson’s ratio.

3.2.4 Constitutive Model

Coal and rock in the model are represented using Mohr-Coulomb plasticity model. The roadway and goaf excavated are represented using null model. The shear yield complies with the Mohr-Coulomb criterion, as shown in Eq. (15.5). The tensile yield is estimated by tensile strength, as shown in Eq. (15.6).

$$ {f}_s={\sigma}_1-{\sigma}_3\frac{1+\sin \varphi }{1-\sin \varphi }-2c\sqrt{\frac{1+\sin \varphi }{1-\sin \varphi }}>0 $$

(15.5)

$$ {\sigma}_3={\sigma}_t $$

(15.6)

where σ₁ is the maximum principal stress , MPa; σ₃ is the minimum principal stress, MPa; φ is the internal friction angle, °; c is the cohesion, MPa; σ_t is the tensile strength, MPa.

3.2.5 Material Properties

There are ten rock strata in the model. The physical and mechanical parameters of coal and rocks are listed in Table 15.8.

Table 15.8 Physical and mechanical parameters of coal and rocks in model

3.2.6 Calculating Procedures

The calculating procedures are as follows:

• 1. Calculating the model to its initial stress equilibrium state without any excavation. Results show that the vertical stress, the maximum horizontal stress and the minimum horizontal stress are 4.58 MPa, 1.93 MPa and 1.12 MPa, respectively, as shown in Fig. 15.13. Comparing with the measured results (4.6 MPa, 1.9 MPa and 1.1 MPa) by the hydrofracturing method, we can reach the conclusion that the numerical model can describe the real stress state of No. 11 coal seam in Luxin Mine.

• 2. For each group, excavating No. 1101 coal face with different face lengths and mining thicknesses as shown in Table 15.9 and iterating to the stress equilibrium state.

• 3. Comparing vertical stress, horizontal stress, vertical displacement and damage zone distribution of model in group A with those of corresponding model in group B to find out whether there are significant boundary effects, as shown in Fig. 15.14. It can be easily concluded from the comparisons that any boundary effects are indeed negligible.

• 4. By analyzing the damage zone distribution in models of group A, the area consisting of damage zones that directly connect the goaf can be seen as water-conducting fractured zone, as shown in Fig. 15.15. Then the heights of water-conducting fractured zones of different models can be obtained.

Fig. 15.13

Horizontal stress distribution without excavation (a) vertical stress (b) the maximum horizontal stress

Table 15.9 Face lengths and mining thicknesses of different models

Fig. 15.14

Some numerical results of models (face length 150 m, mining thickness 3.0 m) in group A and group B (a) vertical stress (b) damage zone distribution (c) horizontal stress (d) vertical displacement (Paralleling the face, at y = 150 m)

Fig. 15.15

Water-conducting fractured zone obtained from the damage zone distribution in model (paralleling the face, at y = 150 m)

3.3 Effects of Face Length on the Height of the Water-Conducting Fractured Zone

The results of Models 1–1 to 1–6 are shown in Fig. 15.16. Results of other models are similar to Fig. 15.16 and not listed. By analyzing the damage zone distribution in models, the heights of the water-conducting fractured zone were obtained. The variations of the height of the water-conducting fractured zone with different face lengths are shown in Fig. 15.17.

Fig. 15.16

Damage zone distribution in models (a) Model 1–1 (b) Model 1–2 (c) Model 1–3 (d) Model 1–4 (e) Model 1–5 (f) Model 1–6 (Paralleling the face, at y = 150 m)

Fig. 15.17

Heights of water-conducting fractured zones for different face lengths (Note: the different curves are for different mining thicknesses, 3.0 m, 4.5 m, 6.0 m, 7.5 m and 9.0 m, respectively)

As shown in Fig. 15.17, for the same mining thickness, the heights of water-conducting fractured zones increase with the face length. Several curves all show an inflection points at the face length of 250 m. The heights increase slowly before the inflection points and rapidly thereafter. Then the heights reach peak values at the face length of 300 m and remain nearly constant. The largest height of the water-conducting fractured zone ranges from 140 to 180 m, about 3 times larger than that at the face length of 100 m. It can be concluded that in shallow coal seams the height of the water-conducting fractured zone generally increases with the face length. And there is a critical value of face length where the height of the water-conducting fractured zone remains nearly constant after the face length increases to the critical value. The critical value of face length of No. 11 coal seam is about 300 m.

3.4 Effects of Mining Thickness on the Height of the Water-Conducting Fractured Zone

The variations of the height of the water-conducting fractured zone with mining thickness are shown in Fig. 15.18. The height of the water-conducting fractured zone generally increases steadily with the mining thickness.

Fig. 15.18

Heights of water-conducting fractured zones with different mining thicknesses (Note: the different curves are for different face lengths, 100 m, 150 m, 200 m, 250 m, 300 m and 350 m, respectively)

3.5 Discussions

To ensure the safety of the shallow coal mine, the mining induced water-conducting fractured zone cannot reach the water-rich aquifer in the overburden. And there must be a certain thickness of the protective layer between the water-rich aquifer and the water-conducting fractured zone. According to “Pillar design and mining regulations under buildings, water, rails and major roadways” guidelines in China, the minimum thickness of the protective layer needed can be obtained from Eq. (15.7). The thicknesses of the protective layer in Models 1–1 to 5–6 can be obtained from Eq. (15.8). The minimum thicknesses needed and the thicknesses of protective layers in Models 1–1 to 5–6 are shown in Table 15.10.

$$ {H}_{\mathrm{min}}=3M $$

(15.7)

where H_min is the minimum thickness of the protective layer needed, m; M is the mining thickness, m.

$$ {H}_N=\left\{\begin{array}{l}H-{H}_{li}\kern1.5em H>{H}_{li}\\ {}0\kern4em H\le {H}_{li}\end{array}\right. $$

(15.8)

where H_N is the thickness of the protective layer in the numerical model, m; H is the minimum distance between coal seam and the water-rich aquifer, m; H_li is the height of the water-conducting fractured zone obtained from numerical calculations, m.

Table 15.10 The minimum thicknesses needed and the thicknesses of protective layers in numerical models

As shown in Table 15.10, only Models 1–1, 1–2, 1–3, 2–1 and 2–2 can ensure enough thickness of the protective layer during mining. Therefore, No.11 coal seam had to adopt slicing mining method. Comparing Models 1–1, 1–2, 1–3, 2–1 and 2–2, it was easy to exclude Models 1–1, 1–2 and 2–1 for their relatively inefficiency in mining. Considering that the abutment pressure of coal face with larger mining thickness is much higher, we selected Model 1–3 as the mining model in the first slice of No. 11 coal seam. Of course, to mine the remaining resources of No. 11 coal seam, some measures such as borehole dewatering and draining should be adopted to dewater the water-rich aquifers in the overburden. Or we can use some special mining methods such as backfill mining to greatly reduce the height of the water-conducting fractured zone.

3.6 Conclusions

• 1. Numerical simulations show that the height of the water-conducting fractured zone increases with both the face length and mining thickness of coal face in a shallow coal seam. And there is a critical value of face length, which is about 300 m in No. 11 coal seam at Luxin Mine, China.

• 2. The slicing mining method is suggested to be adopted in No. 11 coal seam, and the face length and mining thickness of the first slice are recommended as 200 and 3.0 m, respectively. The thickness of the protective layer is expected to be 11.3 m.

• 3. To mine of next slice of No. 11 coal seam, measures such as borehole pumping should be adopted to dewater the water-rich aquifer in overburden. Or some special mining methods such as backfill mining should be used.

4 Water Inrush After Excavation in Coal Mine

4.1 The Luotuoshan Water Inrush Accident

The Luotuoshan coal mine is located in Wuhai City, in the Inner Mongolia Autonomous Region of China, and the original design production capacity was 1.5 Mt annually. On March 1, 2010, the main return roadway of the No.16 longwall panel at the +870 m level experienced a sudden increase in water inflow. The average rate of inrushing water reached 65,000 m³/h, and the total volume reached 75,000 m³. Because the quantity of maximum inrushing water was far greater than the drainage capacity of the mine, the mine was flooded and 77 people were trapped underground. The underground water level rose to +1093 m after stability, as shown in Fig. 15.19. This water inrush accident ultimately led to a loss of 32 lives and a direct economic loss up to RMB 48.53 million (US$7.5 million).

Fig. 15.19

Location of the water inrush accident in the Luotuoshan coal mine

4.2 Characteristics of Collapse Column and Stratum

Formation of karst collapse column is a dynamic geological phenomenon, in which the rock mass or deposits overlying a karstified zone subside along a karst cavity, resulting in a collapse pit or sinkhole (Chen 1988). Based on the detailed geological prospecting work conducted after the water inrush accident (Wang W and Hu B 2011, Zhang W 2015), the spatial distribution of the collapse column is shown in Fig. 15.20. Z₁, Z₂ and Z₃ are geological drilling, Z_1-1, Z_2-1 and Z_3-1 are supplementary drilling, T₁ is exploratory drilling of collapse column after the accident.

Fig. 15.20

Plane view (a) and A-A section view (b) of the collapse column. (After Zhang W 2015)

Based on pumping-out test data of limestone karst aquifer of Ordovician system in Luotuoshan, an average value of the aquifer thickness ranges is 23.33 m, and the average level of groundwater ranges is + 1264.35 m; The water inflow is estimated at 0.0291–0.0311 L/s, the units-inflow ranges from 0.480 × 10⁻³ to 0.512 × 10⁻³ L/s·m, with the permeability coefficient of 1.62 ~ 2.05 mm/d, and water pressure of around 4.1 MPa. The location of water inrush accident , main return roadway of No.16 longwall panel, is 33 m from Ordovician.

From the plane view and section view of collapse column (Fig. 15.20), the small collapse column consists of two parts, the main cave and secondary cave. The main cave is less than 10 m in front of the working face of the No.16 main return roadway, and the secondary cave is buried at the bottom of the coal seam, which has well developed fractures that allow water flow.

4.3 FRACOD Models to Simulate the Water Inrush

There are three factors affecting mine water inrush: water source, water inrush channels and water inrush condition. Among them, the presence of a water inrush channel is a key factor; yet research results on this topic are very vague. Previous studies mostly concentrated on the floor damage characteristics and did not provide a detailed description of the formation of water inrush channels under the joint action of the mining process, geological structure and high water pressure. The aim of this study is to elucidate the formation of a water inrush channel during the process of tunnel excavation.

Based on this, a roadway excavation model was established along the direction of face advance, including the collapse column and limestone karts aquifer of Ordovician, see Fig. 15.21. The simulated main return roadway in the No.16 Seam is 36 m long and 3.66 m high, excavated at a depth of about 414 m below the ground surface. The roadway was located in the upper part of the No.16 coal seam, and the roadway floor is 38 m above the limestone karts aquifer.

Fig. 15.21

FRACOD models used for investigating the effect of the collapse column in Luotuoshan mine. The monitoring lines Lv1~4 are vertical and are located between the collapse column and the working face of the roadway and monitoring lines Lh1~4 are horizontal and are beneath the roadway, respectively

Note that the modelling was done in two-dimensional (2D). The 2D representation of a roadway along its axis direction is problematic because it assumes the infinite width of the roadway, which is clearly an unrealistic assumption. Nevertheless, the 2D model is considered to be more conservative than reality because it can lead to a higher stress concentration in front of the roadway face. The mechanisms of rock fracturing in the 2D model should still be reasonably representative of reality.

The mechanical properties of the simulated rocks and coal are given in Table 15.11. During this numerical study, the model includes two parts: surrounding rock and collapse column, as shown in Fig. 15.21. In the first part, the surrounding rock mass is treated as strata of different homogeneous, isotropic and elastic media, and geostatic stress is considered. In the second part, we have introduced cracks with random orientations inside collapse column. These cracks are assumed to have a hydraulic aperture that enables them to hold water pressure. However, due to the limitation in modelling capacity, it is impossible to define the boundary of the collapse column and use the actual sizes of the micro-cracks. Hence, as a simplification, we used cracks with a length of 1.4 m and a crack density of 0.5 cracks per square metre. In the model, the inner boundary of the roadway is stress free. The bottom boundary of the model is a displacement boundary, and both sides are roller boundaries.

Table 15.11 Mechanical properties of simulated rocks and coal

4.4 Modelling Results and Discussions

4.4.1 A Description of the Water Inrush Event That Occurred

Because the natural stress field changes due to roadway excavation, an area of stress concentration forms in the rock surrounding the roadway, causing deformation or destruction of the roadway. Meanwhile, the fracture and damage of the rock mass significantly change the permeability of the surrounding rock, thus causing water inrush accidents in coal mines. In this model, the Factor of Safety (FoS) is used to analysis failure pattern of surrounding rock at main return roadway of the No.16 Seam. FoS, is also known as safety factor (SF), is a term describing the load carrying capacity of a system beyond the expected or actual loads. Essentially, the factor of safety measures how much stronger the system is than an intended load.

As is shown in Fig. 15.22, the failure zone of surrounding rock is gradually increasing with time near the roadway face, especially between the main cave of collapse column and the roadway face. After cycle 9, the failure zone between the collapse column and the roadway face reaches the main cave, and the failure zone in the roadway floor started to expand towards the secondary cave. After cycle 12, the diameter of the failure zone at the top of the main cave is predicted to be approximately 5 m. Fluid flow in the rock mass in the failure zone of the main cave may cause localized tensile stress, which further drives the pre-existing fractures to propagate. After cycle 14, water pressure drove the short factures in the main cave to propagate towards the roadway face due to expansion of the failure zone and coalescence with the main fracture. At cycle 24, the major fractures coalesce, water inrush channels form and water rushes into the roadway. After cycle 24, the surrounding rock mass of main return roadway of the No.16 seam was generally stable, and the range of the failure zone does not seem to change after coalescence of water inrush channel .

Fig. 15.22

Distribution of modeled total FoS in the surround rock mass at (a) Cycle 9, (b) Cycle 12, (c) Cycle 14, (d) Cycle 24

4.4.2 Stress of Surrounding Rock at Roadway

As shown in Fig. 15.23a, the simulated normal stresses in the x direction along monitoring lines Lv1-4 (Fig. 15.21) for cycle 14 follow a similar trend and have very close values at the lower part of the lines (the left part of the curves in Fig. 15.23a) but large differences at the upper part of the lines in the region in front of the roadway face (the right part of the curves in Fig. 15.23a). This is because, at the cycle 14, the rock in the region at the lower part of the monitoring lines is relatively intact, while fractures are well developed in the region of the coal seam at front of the roadway face, see Fig. 15.22c. The maximum normal stress is observed at the top end of monitoring line Lv1, near the corner of the roof. At the upper part of the monitoring lines, the normal stress decreases with increasing distance from the roadway face. The variation in stress far from the roadway area influenced by mining always exhibited a hysteresis, unlike that of the rock mass near the same area. Hence, the distribution within the high-stress zone was not uniform in the pillar. The zone was first produced in front of the roadway face, and then fractures constantly initiated over time. As a result, more water could flow into this area due to the greater number of fractures.

Fig. 15.23

Comparison of the differential stresses in the coal pillar and floor for cycle 14. (a) The vertical axis is the normal stress in the x direction on monitoring lines Lv1~4 in the coal pillar, the horizontal axis is the distance in the y-coordinate direction. (b) The vertical axis is the normal stress in the y direction on the monitoring lines Lh1~4 in the floor, and the horizontal axis is the distance in the x-coordinate direction

A similar trend is observed for the simulated normal stress in the y-direction at the left end part of the monitoring lines in the roadway floor. Geometrically, the abovementioned part is similar to the top part of a vertical monitoring line: both are near the corners. The maximum normal stress is at the end near the floor corner of the monitoring line Lh1 (Fig. 15.23b). The variations in the normal stress in the region at the right end of the monitoring lines are more complex. This may be due to the effect of the secondary cave of the collapse column (Fig. 15.21). Some complex variations can also be seen in this region in Figs. 15.22c and 15.24b. What’s more, the fractures will change the stress state in the rock mass and has important implications for the permeability of the rock mass.

Fig. 15.24

Modelled fracture initiation, propagation and coalescence in the coal pillar and rock mass surrounding the roadway cavern at (a) cycle 9, (b) cycle 24

4.4.3 Fracture Initiation and Propagation

The FRACOD model predicted that a major fracture initiated at the top corner of the roadway face due to a stress concentration and propagated towards the collapse column (Fig. 15.24a). Meanwhile, the initial short fractures at the collapse column started to propagate towards the roadway face. Although the number of those fractures was very small, they play an important role in the formation of water inrush channel due to the fracture coalescence, eventually forming several long-inclined fractures and then the water inrush channels (Fig. 15.24b).

4.4.4 Formation and Development of Water Inrush Channels

A new numerical simulation of a roadway is conducted, in this cause, the model consists of the cross section of the roadway and some pre-existing fractures at one side representing the collapse column, see Fig. 15.25.

Fig. 15.25

Modelled formation and development of water inrush channels before (a) and after (b) fracture coalescence

Fractures filled with high-pressure water were observed to propagate in the collapse column towards the roadway face (Fig. 15.25a), and a major fracture without water at the top corner of the roadway face propagated towards the direction of the collapse column. In the figure, the blue arrows on the pre-existing fractures represent the relative magnitudes of the fluid pressure. After the fractures coalesced (Fig. 15.25b), water inrush channels formed, and high-pressure water gushed into the roadway face. The high pressure from the water released, and numerically, the water pressure decreased to zero. The newly initiated fractures are not predicted to propagate under the stress conditions, and no water inrush channel has been predicted to for. At this stage and after, no stress-induced fracture initiations are predicted because the stresses are released rather than elevated due to excavation, and the surrounding rock mass of the main return roadway of No.16 generally stablilizes.

4.5 Conclusions

The case study conducted for the Luotuoshan accident using a FRACOD model indicates that, the water pressure drove the short factures in the main cave to propagate towards the roadway face and eventually coalesce with the factures in the coal pillar. These fractures propagated and coalesced with other newly initiated short fractures and those initiated from the roadway corner, eventually forming several long fractures and hence a water inrush channel. Based on the simulation results, it is determined that the water source of the accident was the Ordovician limestone aquifer beneath the mining seam, which is connected by the collapse column to the No.16 coal seam. Water inrush occurred when the roadway face was at some distance from the main cave in the collapse column, where fracture initiation and propagation finally linked the roadway to the main cave and formed water inrush channels .