Rock failure process analysis method (RFPA) for modeling coal strata movement

Rock failure process analysis method (RFPA) for modeling coal strata movement

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Chun-an Tang1 and Tao Xu2 1 Dalian University of Technology, Dalian, China, 2Northeastern University, Shenyang, China

Chapter Outline 11.1 Introduction 345 11.2 Rock Failure Process Analysis Method 11.2.1 11.2.2 11.2.3 11.2.4

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Gas flow 347 Stress analysis 347 Damage induced permeability evolution 348 Characterization of heterogeneity 352

11.3 Case Studies

353

11.3.1 Overburden movement 353 11.3.2 Gas flow 363

11.4 Conclusions 374 Acknowledgments 376 References 376

11.1

Introduction

Nonlinear and discontinuous numerical methods have become an important tool in modeling mining engineering processes for the analysis of stresses and deformations in mining or rock engineering structures such as underground openings. Although many numerical methods, such as finite element method, boundary element method, finite-difference method, and discrete element method, can do well in simulating nonlinear behavior in rock deformation, most of them are not a physical modeling of the nonlinear behavior of brittle rock. In the nonlinear FEM, for example, the nonlinear behavior has been assigned by introducing a nonlinear constitutive relation to elements that are considered to be homogeneous material. This undoubtedly results in a nonlinear outcome. Though this method has gained a sufficient degree of functionality, and may also incorporate compaction features, it may Advances in Coal Mine Ground Control. DOI: http://dx.doi.org/10.1016/B978-0-08-101225-3.00015-3 Copyright © 2017 Elsevier Ltd. All rights reserved.

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not always help in our understanding of why rock material demonstrates nonlinear behavior. It also prevents us from approaching seismicity problems since the seismicity-prone material is by no means homogeneous. One of the most important factors affecting the progressive failure is heterogeneity. When rock is subjected to a stress field, cracks may nucleate, propagate, interact, and coalesce. During fracturing, the heterogeneity plays a marked influence in determining the fracture paths and the resulting fracture patterns. The influence of heterogeneity is pronounced on the progressive failure process. Even core specimens obtained from a seemingly homogeneous block of rock show variability both in deformation and strength properties. Thus, the distributive character of the heterogeneity plays a crucial role in determining the evolution of fractures. Therefore, a more reasonable numerical model for the rock or rock mass should be able to demonstrate the progressive failure due to heterogeneity, which results in nonlinear behavior. This may only succeed via a statistical approach. Thus, the numerical method, Rock Failure Process Analysis Code (RFPA), to modeling progressive failure of rock is developed (Tang, 1997) and further extended to model the observed evolution of damage and induced seismicity due to the progressive failure leading to collapse in brittle rock or rock mass (Tang et al., 2003, 2007, 2008). In this chapter, the RFPA method is described and the application of RFPA method to overburden movement and associated gas flow is reported.

11.2

Rock Failure Process Analysis Method

Numerical model is currently the most commonly used method in the solution of complex problems in rock mechanics and engineering, such as reservoir simulation and coal bed methane recovery (Connell and Detournay, 2009; Karacan et al., 2007; Lunarzewski, 1998; Pan and Connell, 2009; Xu et al., 2006). RFPA method for strata movement (Xu et al., 2006; Yang et al., 2007) was proposed to describe the rock failure process and associated with gas flow during coal mining. Here RFPA method was applied to investigate the mechanism of the complex gas migration during coal mining intending to gain an insight into the coupling mechanism between gas flow and coal deformation. When the model is formulated mathematically, various levels of complexity can be incorporated into each component, with the accuracy and versatility of the model depending on the refinement of the components description. For coal seams, the coupled effect of the medium deformation and fluid flow may be important to be formulated. The formulation of the coupled model may help understand the mechanism of gas migration in gas drainage, coal and gas outbursts in mining or drilling and gas disposal in engineering practice. For a method that can be used to describe the interaction between gas flow and coal or rock deformation, three components must be accounted for: (1) a gas flow description, (2) a stress description, and (3) a failure description. Hereby, the gas flow, the stress analysis, and the failure criteria that are implemented in the RFPA method for gas flow model are presented in the following sections.

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11.2.1 Gas flow The fundamental assumption in the model is that gas is saturated in the coal and thus the equations for two phase flow in media can be used. Based on the Darcy’s law, Zhou and Lin (1998) further developed the gas filtration equation followed by the linear relationship in Eq. (11.1). qi 5 2 λij

dP dn

(11.1)

where, qi denotes gas filtration rate (i 5 1, 2, 3) in m/s; λij is coefficient of gas filtration (i, j 5 1, 2, 3) in m2/(MPa2
S); and P is square of gas pressure in MPa2. Normally, the coefficient of gas filtration λij is about 2.5 3 10217 m2 times that of the intrinsic permeability k. Note that the parameters λij and P is different from those in Darcy’s law though they are similar. Generally, gas occurs in coal in two forms, free gas and adsorbed gas. The adsorbed gas typically accounts for over 95% of the total gas, depending on the pressure at which the gas is adsorbed. The free gas, only a little fraction of the total gas, is stored in the pores or cleats either free or in solution. The total gas content in coal can be approximated by empirical relationship in Eq. (11.2) (Zhou and Lin, 1998) pffiffiffi X5A p (11.2) where, X is gas content in gassy coal in m3/m3; A is the empirical coefficient of gas content in m3/(m3
MPa1/2); and p is gas pressure in MPa. According to the fundamental seepage theory of gas flow in porous media, the isothermal filtration gas flow in gassy coal and rock can be described as Eq. (11.3) αp
r2 P 5

@P @t

(11.3) 3

where, αp 5 4λA21 P4 .

11.2.2 Stress analysis The stress can be formulated in a number of ways. For a stress analysis in terms of effective stress, the stress equilibrium equations take the form σij; j 1 fi 5 0

(11.4)

where, σij is stress tensor, ði; j 5 1; 2; 3Þ in MPa; and fi is stress caused by the body forces per unit volume in MPa. Based on the Terzaghi’s effective stress principle, the stress equilibrium equation can be expressed as Eq. (11.5) for one- to two-phase materials: σij 5 σ0ij 1 α
p
δij

(11.5)

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where, σij is the total stress tensor; σ0ij is the effective stress tensor of the solid phase; p is the gas pressure; α is a positive constant, which is equal to 1 when individual grains are much more incompressible than the grain skeleton; and δij is the Kronecker delta function. Substitution of Eq. (11.5) into Eq. (11.4) leads to Eq. (11.6) as follows: σ0ij; j 1 fi 1 ðα
p
δij Þ;j 5 0

(11.6)

The equilibrium equation is then expressed according to the effective stress principle. According to the continuity conditions, for a perfectly elastic isotropic continuum, the geometrical equation can be expressed as Eq. (11.7) εij 5

1 ðui;j 1 uj;i Þ 2

(11.7)

where, εij is strain tensor, ði; j 5 1; 2; 3Þ; εv is the volumetric strain; εv 5 ε11 1 ε22 1 ε33 ; and u is the displacement of element. The constitutive equation of deformation fields can be expressed as Eq. (11.8) for elastic isotropic materials. σ0ij 5 Kδij εv 1 2Gεij

(11.8)

where, G is shear modulus and K is Lame’s constant. Based on the equilibrium (Eq. 11.6), continuity (Eq. 11.7), and constitutive (Eq. 11.8) equations, the governing equations can be represented as Eq. (11.9) for mathematical model of coal/rock deformation considering the gas pressure in coal/ rock: ðK 1 GÞ
uj; ji 1 Gui; jj 1 fi 1 ðα
pÞ;i 5 0

(11.9)

11.2.3 Damage induced permeability evolution In the RFPA method, only the linear elastic constitutive equations have been introduced for all elements, which are assigned to have different strength and elastic constant parameters depending on the heterogeneity of rock materials. When the stress of the element satisfies the strength criterion (such as the CoulombMohr criterion), the element begins to undergo damage. In elastic damage mechanics, the elastic modulus of the element may degrade gradually as damage progresses. If the element and its damage are assumed to be isotropic, the elastic modulus of the damaged element is defined as follows (Lemaitre and Desmorat, 2005) E 5 E0 ð1 2 DÞ

(11.10)

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where, D represents the damage variable, and E and E0 are the elastic moduli of the damaged and undamaged elements, respectively. The parameters E, E0, and D are all scalar. The constitutive model used in RFPA is shown in Fig. 11.1. Under compression states, the MohrCoulomb criterion is chosen as the strength criterion for the elements: σ1 2 σ3

1 1 sinφ $ fc 1 2 sinφ

(11.11)

where, σ1 and σ3 are the maximum principal stress and minimum principal stress, respectively, φ is the internal friction angle, and fc is the threshold of the compressive strength of the element. Correspondingly, the damage variable D in compression can be expressed as (Tang et al., 2002): 8 <

0

D 5 1 2 fcr : E0 ε

ε , ε c0 (11.12)

ε c0 # ε

where, fcr is the residual compressive strength; ε and εc0 are the compressive strain and the compressive threshold strain, respectively. Similarly, the maximum tensile stress criterion is chosen as the strength criterion for the elements in tension σ 3 # 2 ft

(11.13)

where, ft is the threshold of the tensile strength of the element. σ

σc

σcr εt σrt σt

Figure 11.1 Constitutive model used in RFPA.

εc

ε

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Correspondingly, the damage variable D in tension can be expressed as 8 0 ε t0 # ε > > < ftr D 5 1 2 E ε εtu # ε , εt0 0 > > : 1 ε # ε tu

(11.14)

where ftr is the residual tensile strength of the elements; εt0 and εtu are the tensile threshold strain of the damage elements and the ultimate tensile strain of the failed elements, respectively. Generally, the stress decrease is the main factor leading to the increase of the gas permeability. In the numerical model, gas flow is coupled to stress describing the permeability change induced by the decrease of the stress field. The coupling function can be described as Eq. (11.15) (Louis, 1974) λ 0 5 e2βσ λ0

(11.15)

where, λ is current gas permeability; λ0 is original gas permeability; β is coupling coefficient (stress sensitive factor to be measured by experiment); and σ0 is effective stress. For damage induced permeability change, most of the theories are only valid in pre-failure regions. During elastic deformations, rock permeability may either decrease when the rock compacts or increase when the rock expands. However, a dramatic and remarkable increase in rock permeability can be expected as a result of the generation of numerous micro fractures near and at the peak load. Once passing the peak load, the permeability may gradually drop again if the failed rock is further compacted, or the permeability may increase continuously if the failed rock is further expanded. The gas permeability coefficients in uniaxial compression and tension can be described as Eqs. (11.16) and (11.17), respectively. For the elements in compression, the gas permeability can be described as Eq. (11.16)  λ5

λ0 e2βðσ1 2αpÞ ξλ0 e2βðσ1 2αpÞ

D50 D.0

(11.16)

where, λ0 is the initial gas permeability for unloaded coal and rock; β is the coupling factor of stress to pore pressure; α is the coefficient of pore pressure; and ξ is the coefficient of sudden jump of gas permeability for loaded elements in compression. For the elements in tension, the gas permeability-stress equation is expressed as Eq. (11.17) 8 < λ0 e2βðσ3 2αpÞ D50 (11.17) λ 5 ξλ0 e2βðσ3 2αpÞ 0 , D , 1 : 0 2βðσ3 2pÞ D 5 1 ξ λ0 e

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where, ξ0 is the coefficient of sudden jump of gas permeability for failed elements in tension. The other parameters are the same as those in the equations mentioned above. AEs (acoustic emissions) are transient elastic waves generated by the rapid release of energy within a material, such as the strain energy released during microcrack propagation. Monitoring AE during deformation has become an increasingly important diagnostic tool in material science and has provided a wealth of information regarding the failure process in brittle materials. AE monitoring has shed light on the onset of microcracking during deformation or C0 (Wong et al., 1997), the evolution of the spatial and temporal progression of microcracks (Benson et al., 2007; Brantut et al., 2013; Lockner, 1993a; Ohnaka, 1983), amongst many more. For instance, Lockner (1993b) analyzed catalogs of AE events recorded during compressive loading tests on rock. The events were analyzed in terms of the information they offer about the accumulated state of damage in a material. This measured damage state can be combined with a model for the weakening behavior of cracked solids, showing that reasonable predictions of the mechanical behavior are possible. Based on this prior knowledge it is reasonable to assume that the number of AE events is proportional to the number of damaged elements and that the strain energy released (the strain energy before and after damage) corresponds to the energy of that particular AE event (Tang, 1997; Tang et al., 2007). In our model, we can use the output of AE to indirectly assess the damage evolution. However, it must be mentioned that aseismic damage during rock creep tests could possibly occur. The causes of aseismic damage are numerous, for example: the low surface energy of calcite, radiated energy being absorbed by neighboring dislocation, and/ or intermittent dislocation flow (Schubnel et al., 2006), amongst many more. Although this approximation is obviously a simplification of what occurs in reality, it has been shown that this micromechanical representation of microcracking can yield realistic patterns and can reproduce the macromechanical behavior of heterogeneous rock. The cumulative damage, ψ, in a given volume of rock, due to local failures can be defined as the ratio of the volume of failed rock, Vf, to the total volume, V: ψ5

P s Vf ve
s1 ni 1X 5 5 ni N 1 V N
ve

(11.18)

where ve is the volume of single element, s is the number of calculation steps, ni is the number of failed elements in the ith step and N is the total number of elements in the model. For a perfectly elastic brittle material, the energy, ef released by the failure of each element can be calculated from the element peak strength: ef 5

σ20 ve 2E

(11.19)

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where σ20 is the peak strength of the element and E is the Young’s modulus of the element. The cumulative seismic energy can then be obtained by: X

ef 5

X σ2 0

2E

ve 5

ve X σ20 2 E

(11.20)

Thus, by recording the number of failed elements, the AE associated with the progressive failure of the material can be simulated in our model.

11.2.4 Characterization of heterogeneity In the absence of heterogeneity, the behavior of the model is entirely homogenous, no local damage occurs, and the local behavior is replicated at the macroscopic scale. Thus, it is necessary to introduce heterogeneity to obtain a collective macroscopic behavior different from those of the elements. In order to reflect the material heterogeneity at a mesoscale, the mechanical parameters (e.g., strength and Young’s modulus) of the mesoscopic material elements, which are assumed to be homogeneous and isotropic, are assigned randomly from the Weibull statistic distribution (Weibull, 1951) as defined in the following statistics probability density function: σðuÞ 5

    m  m u m21 u exp 2 u0 u0 u0

(11.21)

where u is the scale parameter of an individual element such as the strength or Young’s modulus and the scale parameter u0 is related to the average element

Figure 11.2 Weibull distribution with various homogeneity index.

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parameter. The shape parameter m reflects the degree of material homogeneity and is defined as a homogeneity index. According to the Weibull distribution and the definition of homogeneity index, a larger m implies that more elements will have the mechanical properties similar to the mean value, resulting in a more homogeneous material, as shown in Fig. 11.2.

11.3

Case Studies

11.3.1 Overburden movement 11.3.1.1 Model setup In this section, RFPA method is used to investigate the characteristics of deformation and fracturing of overburden strata during the extraction of coal seam. The main mineable coal seam of the coal mine is 6 m thick on average and is nearly horizontal. The panel is mined using the longwall retreat mining method with natural roof caving. The immediate roof of the panel is composed of sandy mudstone and is approximately 2 m thick. The average depth of coal seam below surface is approximately 84 m. Coal has been produced by means of longwall retreat method with 5 m mining height. The model domain for the study area is 100 m high and 300 m long. The numerical model is discretized into 30,000 elements (100 3 300) and contains a total of 12 strata based on the site-specific geological conditions as shown in Fig. 11.3. Considering the effect of bedding and weak planes on the failure of rock mass, it is necessary to embed some bedding planes between two contiguous strata in the model. The mining length of coal seam in the model is 115 m, with 23 steps in total, i.e., 5 m each step of mining. According to the site-specific mining conditions, it is assumed that the time interval of each step is a half day, i.e., the extraction of coal seam lasts 11 days and a half day in the model simulation. In the simulation, the elements in the numerical model are characterized by the Young’s modulus (E), uniaxial compressive strength, tensile strength, and Poisson’s ratio. It is crucial to properly assess the properties of the surrounding rocks to obtain acceptable results for numerical modeling. Therefore, the physical and mechanical properties of each geological unit must be determined. In general, the

Figure 11.3 Model for simulation of overburden movement.

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properties of the surrounding rocks are determined by laboratory testing. Samples of the surrounding rock described above were obtained from exploration drilling cores and rock blocks taken directly from the coal mine. According to the available literature (Brady and Brown, 2004), the uniaxial compressive strength, Young’s modulus, Poisson’s ratio, cohesion, and friction angle of the rock are needed to analyze the fracturing and caving process of overburden strata and the distribution characteristics of abutment pressures over coal face and immediate roof in this study. Uniaxial compression tests were used to determine the uniaxial compressive strength, Young’s modulus, and Poisson’s ratio. The cohesion and friction angle of the surrounding rocks were obtained by triaxial compression tests. Based on the results of these tests, the panel stratigraphy and other important geotechnical parameters of the coal seam, roof, and floor strata adopted in the simulation for the numerical model are listed in Table 11.1. The boundary conditions of the numerical model are that the both sides of the model are restricted by displacement in the horizontal direction, the upper boundary of the model is free and the bottom of the

Physical and mechanical properties of coal and surrounding rocks

Table

11.1

No

Formation

Young’s modulus (GPa)

UCS (MPa)

Density (3103 kg m3)

Thickness (m)

1

Weathered sandstone

6

50

2.48

7

2

Medium sandstone

8

60

2.65

10

3

Siltstone

10

60

2.56

8

4

Mudstone

2

36

2.56

8

5

Siltstone

8

50

2.56

8

6

Shale

4

45

2.6

10

7

Medium sandstone

8

30

2.5

12

8

Fine sandstone

6

20

2.5

10

4

20

2.6

9

1.5

5

2.5

2

9 10

Mudstone Sandy mudstone

11

Coal seam

1

33

1.4

6

12

Siltstone

10

100

2.5

10

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model is fixed by displacement. In the simulations, coal mining in 23 steps by 5 m/ step was consecutively carried out to simulate the progressive coal extraction process.

11.3.1.2 Overburden movement process The typical sequences of overburden strata movement and caving as the face advances are presented in Fig. 11.4. As the face advances some distance of 20 m from the set up room, layer 10 of sandy mudstone bends under the gravity, cracks initiate at mid span of lower side and at both ends of the upper side of layer 10 as shown in Fig. 11.4A1. When the span is beyond the limit characteristics of layer 10, the cracks will propagate and coalesce near the end of coal face, and at last lead to the caving of layer 10 with a caving thickness of 2 m, as shown in Fig. 11.4A2. Later on, layer 10 continuously caves as the face advances 35 m as shown in Fig. 11.4B. Layer 9 bends under the gravity with the continuous advance of the face. When the face has advanced to 45 m, fractures initiates and propagates along the bedding planes between layer 8 and layer 9 over layer 10, and tensile fractures also initiate at mid span of lower side and at both ends of the upper side of layer 9, as shown in Fig. 11.4C. As the face advances to 50 m, the caving of layer 9 occurs and cuts and caves along the coal face, as shown in Fig. 11.4D. As the face advances to 60 m, bed separation and cracking in layer 8 can be observed as shown in Fig. 11.4E. As the face continuously advances to 70 m, the adjacent upper strata close to the face begins to fracture, rotate and cave as shown in Fig. 11.4F. As the face advances to 75 m, the caving of the layer 8 occurs in Fig. 11.4G. As the face continues to advance to 85 m, layer 10 over coal face shortly caves, bed separation along the bedding planes (such as between layers 6 and 7, between layers 3 and 4), and cracking above the strata of layer 7 occurs. The adjacent thick upper strong main roof (layer 7) tends to cantilever over the goaf. It gradually bends and finally ruptures at the ends as shown in Fig. 11.4H1 and H2. As the face advances to 105 m, cracking occurs in the layer 6 and bed separation can be observed along the bedding planes in the upper strata as shown in Fig. 11.4I. Thereafter, the overburden rock strata caves in periodically as the face advances as shown in Fig. 11.4J. It follows that the typical sequences of overburden strata movement and caving with the advancing of the face can be summarized as follows: A set-up room is made in the coal seam, the intact stress is disturbed and the pressure redistributes itself. As the face advances away from set-up room, the roof beds above the excavation bend downwards from over the coal across the coal face. As the span across the coal face reaches to a certain limit, the roof bed above the excavation ruptures and caves. The roof beds above goaf will again bend and cause the beds to sag away from each other. The bending of roof beds causes the formation of bed separation in the overburden, and even the rupturing and caving of the roof bed adjacent to the goaf. Thereafter, the overburden roof beds caves in periodically as the coal face gradually advances.

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Figure 11.4 Sequence of overburden strata movement and caving as the face advances. (A1 and A2) Coal face advances 20 m, (B) coal face advances 35 m, (C) coal face advances 45 m, (D) coal face advances 50 m, (E) coal face advances 60 m, (F) coal face advances 70 m, (G) coal face advances 75 m, (H1 and H2) coal face advances 85 m, (I) coal face advances 105 m, and (J) coal face advances 115 m.

Rock failure process analysis method (RFPA) for modeling coal strata movement

Figure 11.4 (Continued).

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Figure 11.4 (Continued).

Figure 11.5 Stress distribution in overburden strata as the face advances. (A1 and A2) Coal face advances 20 m, (B) coal face advances 35 m, (C) coal face advances 45 m, (D) coal face advances 50 m, (E) coal face advances 60 m, (F) coal face advances 75 m, (G) coal face advances 85 m, (H) coal face advances 105 m, and (I) coal face advances 115 m.

11.3.1.3 Associated stress in overburden and weighting characteristics Fig. 11.5 presents the associated stress field in the overburden in panel mining. As the face advances, bed separation first occurs in layer 10, and high tensile stress concentrates at mid span of lower side and at both ends of the upper side of layer 10. Cracks initiate at mid span of lower side and at both ends of the upper side of layer 10 when the mining-induced tensile stress is beyond the tensile strength of layer 10 as shown in Fig. 11.5A1. It follows that the fracturing of layer 10 is

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Figure 11.5 (Continued).

induced by high tensile stress. Later on, layer 10 shortly caves with the advancing of the face, as shown in Fig. 11.5A2 and B. As the face advances to 45 m, bed separation occurs in the middle of layer 9. Meanwhile, the high tensile stress induces cracks at mid span of lower side and at both ends of the upper side of layer 9 as shown in Fig. 11.5C. As the face goes on, layer 9 near the goaf always undergoes the effect of the mining-induced tensile stress and thus is severely damaged, while the maximum stress zone in layer 9 near the face progressively moves forward and correspondingly the mining-induced damage zone also moves forward. Finally layer 9 unsymmetrically ruptures and caves as shown in Fig. 11.5D. Thereafter, as the

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Figure 11.5 (Continued).

face advances, the overburden strata caves in periodically under the mining-induced high tensile stress, as shown in Fig. 11.5EI. As the face advances to 20 m, the layer 10 caves with a caving height of 2 m as shown in Fig. 11.5A1. Then layer 10 shortly caves with the face advancing and the caved zone gradually expands upwards. As the face advances to 50 m, the caving of layer 9 occurs and the height of caved zone goes up to about 11 m, as shown in Fig. 11.5D. As the face advances to 75 m, the first periodic weight of main roof with an interval of 25 m occurs and the height of caved zone reaches to 21 m as shown in Fig. 11.5F. As the face advances to 105 m, the second periodic weight of main roof with an interval of 30 m occurs and the height of caved zone reaches to 43 m as shown in Fig. 11.5H. Thereafter, overburdened caves in periodically and periodic weighting of main roof occur as the face advances. It can be seen from numerical simulations that RFPA model well reproduces the overburden movement process, reveal the rupturing mechanism of overburden strata and captures periodic weighting of the roof even though there is not any pre-existing cracks or fractures in the model.

11.3.1.4 Abutment pressures with the face advancing When an opening is created in the coal seam, the stress that was present before the opening was created is redistributed to the adjacent coal pillars that are left.

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The areas within the remaining coal where the vertical stress is greater than the average are called abutments and hence the stresses in those areas are called abutment pressures. The abutment pressures over coal face gradually increase as the face advances, and Fig. 11.6 shows the distribution and evolution of abutment pressures over coal face with the face advancing. As the face advances to 20 m, stress

Figure 11.6 Abutment pressure distribution as the face advances. (A) Coal face advances 20 m, (B) coal face advances 25 m, (C) coal face advances 35 m, (D) coal face advances 45 m, (E) coal face advances 50 m, (F) coal face advances 60 m, (G) coal face advances 70 m, (H) coal face advances 75 m, (I) coal face advances 85 m, (J) coal face advances 105 m, and (K) coal face advances 115 m.

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Figure 11.6 (Continued).

concentration forms around coal wall and the distribution of abutment pressures exhibits a monotonic curve with a maximum pressure around the coal face as shown in Fig. 11.6A. As the face advances to 25 m, the abutment pressures over the face decrease drastically due to the first immediate roof weighting, as shown in Fig. 11.6B. Later on, the abutment pressures over coal face gradually increase with the face advancing, as shown in Fig. 11.6C and D. As the face advances to 50 m, the abutment pressures over the face again decrease drastically due to the first main roof weighting, as shown in Fig. 11.6E. The main roof strata caves in the goaf and

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takes the dead weight of the caved strata as waste rock, as shown in Fig. 11.6EK, in which the stress in the goaf is that of the waste rock withstands the caved strata. As the face goes on advancing, the abutment pressures over coal face gradually increase again, as shown in Fig. 11.6F and G. As the face advances to 75 m, the first periodic weighting of main roof occurs and abutment pressures over the face again decrease drastically, as shown in Fig. 11.6H. Thereafter, abutment pressures over the face change periodically as the face advances as shown in Fig. 11.6HK. It is noted that the front abutment pressures over coal face are higher than back abutment pressures over coal face due to the continuous effect of mining excavation from coal face.

11.3.2 Gas flow In recent years, a new mining technique called pressure relief gas drainage technology has been applied in mining areas with multiple high-gassy coal seams of low gas permeability (Chen et al., 2004; Yu et al., 2004). The main idea of the technique is to first mine the coal seam with low gas contents and low risks of coal and gas outbursts, which serves to relieve the stress in the protected coal seam and improve the gas permeability of the protected coal seam. Furthermore, the technique promotes considerably the desorption of coal methane in the protected coal seam and the formation of a high-efficiency extraction conditions, which facilitates the subsequent gas drainage in the protected target coal seam. Correspondingly, the technique not only avoids the pollution of air by releasing gas from coal seams and gobs, but also greatly reduces the gas content in the coal seams to eliminate the danger of gas explosion and coal and gas outbursts. Thus, the technique secures a fast and high-efficiency exploitation of gas and coal in the pressure relieved coal seams. Based on the mechanical analysis and engineering practice of pressure relief long-distance gas drainage at the Panyi coal mine of Huainan coal mining Co. Ltd, RFPA is used to model the mining of coal seam at great depth with this technique and to study the deformation and fracture characteristics of overburden strata, the evolution of gas permeability, and the gas migration in target coal seams. It is expected to gain an insight into the gas flow and migration mechanism and to offer some theoretical basis and scientific evidence for the application of the pressure relief long-distance gas drainage technique.

11.3.2.1 Description of the area under study The Panyi coal mine of Huainan Mining Group Co. Ltd, is located in the Huainan coal mining area, one of the largest coal fields in China. It has been mined since 1983. It is a typical highly gassy mine. At present, the coal seams C13, B11, and B8 are being mined. The mining level is between 2530 and 2650 m. The fullymechanized longwall mining is adopted. The initial protective coal seam was the coal seam B11. The panel lengths along the strike and dip directions were 1640 m and 190 m, respectively. The thickness of the coal seam B11 was 1.52.4 m, with an average of 2 m. The slope angle of the coal seam was 613 degrees, with an

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average of 9 degrees. The original gas content of coal seam B11 was 47.5 m3/t, with an average gas content gradient of 75.7 m/(m3/t). The relative gas emission of 5.237.32 m3/t, and the absolute gas emission of 3.073.38 m3/min were far below those of coal seam C13. Thus, the coal seam B11 was not liable to coal and gas outbursts. The thickness of the coal seam B11 was uniform and the geological structure was simple. In contrary, the thickness of the coal seam C13 was 5.576.25 m with an average of 6 m. The slope angle of the coal seam varied from 6 to 13 degrees with 9 degrees on average. The measured gas pressure of the coal seam C13 was about 4.4 MPa on average, with a fluctuation of 5.0 MPa at the level of 2580 m and 5.6 MPa at the level of 2620 m. The gas content was 14.2 m3/t and the average gas pressure gradient was 2.42 3 10 MPa/m. The initial gas permeability coefficient was 2.84 3 1024 in milli-darcy (md) and gas content coefficient was 9 m3/m3 (MPa0.5). The gas permeability coefficients of coal and rock are expressed in the international unit of milli-darcy here though they are commonly expressed in m2/ (MPa2
d) in the coal mines in China. During coal mining, the relative gas emission was 14.838.6 m3/t, with an average gas emission up to 25.0 m3/t. The absolute gas emission was 22.733.1 m3/min and was 27.0 m3/min on average. Fig. 11.7 shows the relationship between gas pressure and gas content of coal seam C13. The geological structure of the coal seam C13 was simple. The panel lengths along the strike and dip directions were 1680 m and 160 m, respectively. Coal and gas outbursts and gas explosion accidents have occurred frequently in the coal seam C13 in recent years. The interburden, viz. the distance from the roof of the coal seam B11 to coal seam C13 was nearly 67 m. Due to the low gas content of coal seam B11, and high gas content and low permeability of coal seam C13, coal seam B11 was first mined to reduce the stress field and increase the gas permeability of the coal seam C13. Meanwhile, the pressure relief gas drainage by long-distance boreholes was performed to capture the released gas and avoid possible coal and gas outbursts. In this study, the panel No 2151 of the coal seam B11 was first extracted and

Figure 11.7 Relationship between gas pressure and gas content.

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the coal seam C13 was considered as the pressure relief coal seam. The panel No 2121 of the coal seam C13 was then mined after the gas permeability of the coal seam C13 was enhanced by the pressure relief induced by the extraction of the coal seam B11. After the coal seam B11 was mined, the height of the fractured zone measured was 30.136.1 m, which was formed by horizontal bedding plane separations and vertical mining induced fractures. The height of the caving zone, characterized as fragmented rock mass, was 8.511.0 m. The coal seam C13 just lied in the bending zone of the overlying strata induced by coal mining. The extraction of the coal seam B11 only caused the coal seam C13 to subside as a whole, in which parallel layered fissures and a few vertical cross fissures formed. Monitoring points were installed by drilling two boreholes at the roof and floor of gas drainage testing roadway below the coal seam C13 to determine the deformation of coal seam by measuring the relative displacement of two points. Fig. 11.8 shows the relative displacement between roof and floor in the coal seam C13 along the face advancing direction. The compressive deformation was up to 27 mm, while the tensile deformation at the center reached 210.44 mm. The largest relative compressive deformation was 3.37m and the largest relative tensile deformation was 26.3m, indicating that the in situ stress in the coal seam C13 decreased considerably, while the fractures in the coal seam C13 increased significantly due to the extraction of the coal seam B11. In addition, monitoring points were installed, too, along the floor of the gas drainage testing roadway in the coal seam C13 before mining the coal seam B11. The final relative displacement of the floor in the coal seam C13 was measured during mining the coal seam B11. Fig. 11.9 shows the relationship between the measured floor subsidence of the coal seam C13 and the mining distance away from the face. The measured floor subsidence reached the maximum average value of 1.56 m at about 40 m from the panel set up room. On the whole,

Figure 11.8 Measured roof and floor deformations (compression is negative).

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Figure 11.9 Measured floor subsidence of Coal seam C13 as the face advanced.

Figure 11.10 Vertical cross-section showing three gas drainage testing boreholes drilled in the gas drainage roadway.

the support of the roadway was still stable and the roadways in the bending zone were suitable for gas drainage. Three gas drainage testing boreholes, that lied within the boundary of unloading stress induced by mining coal seam B11, were installed in the coal seam C13 (Fig. 11.10) to determine the changes of gas pressure, gas flow, and gaspermeability when mining in the coal seam B11. As the face advanced, the gas flow in the borehole No.3 reached 4.48 L/min (6.45 m3/d), specific gas flow 1.82 m3/ (d
m2) and the coefficient of gas permeability 0.284 md. It is noted that the coefficient of gas permeability increased 2800 times within a period of four and one half days, from 11:40 am, July 31, 2000 to 10:00 am, August 4, 2000. Fig. 11.11 shows the variation of the gas pressure measured around the borehole in the coal seam C13 as the face advanced. It can be seen that the gas pressure sharply decrease from 4.4 to 0.4 MPa as the face advanced from about 40 to 80 m away from the borehole in the coal seam C13 along the strike direction. Subsequently, the gas pressure

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Figure 11.11 Measured gas pressures as the face advanced.

Figure 11.12 The gas flow model.

stabilized at approximate 0.4 MPa as the face advanced from about 80 m and beyond away from the borehole. According to the in situ stresses, gas pressures, gas permeability, and other mechanic parameters measured onsite, a plain strain model is setup for the case mentioned above using the RFPA2D-GasFlow code, as shown in Fig. 11.12. The grayness level in the Fig. 11.12 represents the magnitude of elastic modulus. The lighter the gray is, the higher the elastic modulus is. The domain of the numerical model is 110 m in height and 300 m in length. The thickness of the upper coal seam C13 is 6 m, which is located 500 m below the surface. The pressure of 10 MPa is loaded on the model surface boundary to represent the dead weight of 500 m strata in thickness. The thickness of the coal seam B11, 67 m from the upper coal seam C13, is 2 m. The numerical model contains a total of 10 layers based on the site-specific geological conditions. With the effect of bedding and weak planes on the failure of rock mass in mind, it is necessary to embed some bedding planes

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in the model. The mining length of the coal seam B11 is 100 m, with a total of 20 steps, i.e., 5 m/step. Based on the site-specific mining conditions, it is assumed that the time interval of each step is one half day. For example, the extraction of the coal seam B11 lasts 10 days in the simulation. The initial gas pressure in the coal seam C13 is 5 MPa and the initial gas pressure in the strata above and below the coal seam C13 are assumed to be approximately zero, as shown in Fig. 11.12. The gas drainage hole is located in the middle of the coal seam C13 and its initial gas pressure is 20 kPa. The time interval of drainage at the hole is 0.5 day per step. As the face advances, the seepage of gas from the coal seam gradually continues. In this model, gas permeability λ at zero stress condition is 0.284 md, and stress sensitive factor β is 2.0. The physical, mechanical, and seepage parameters in the model are listed in Table 11.2. The numerical simulation was carried out to investigate the strata deformation and failure, the change of gas permeability and the characteristics of gas flow in the coal seam C13 during the extraction of the coal seam B11.

11.3.2.2 Deformation and fracture characteristics of overburden strata Fig. 11.13 depicts the dynamic evolution of fracturing, caving processes and corresponding stress field in the overburden strata during mining. The grayness levels represent the magnitude of stresses. The lighter the gray is, the higher the stress is, and vice versa. It can be seen that, with the face advancing, the strata above the mined coal seam gradually fractures and caves. Correspondingly, the three zones (i.e., caving, fractured, and bending zones) form in the overburden strata above the mined coal seam. Moreover, the numerical results show that the first caving interval is 30 m and subsequent periodic caving interval is about 15 m, which agrees well with the field observations. When the face advances close to 100 m, the height of the caving zone in the overburden strata above the coal seam B11 is between 15 and 20 m, and the height of the fractured zone is between 40 and 45 m. In addition, the results indicate that the extraction of the coal seam B11 does not cause the fracture of the coal seam C13, but does cause the subsidence of the coal seam C13 since it lies in the mining induced bending zone in the overburden strata.

11.3.2.3 Effect of in situ stress on gas permeability of coal seam The extraction of the coal seam B11 leads to the redistribution of the stress fields in the overburden strata, which in turn causes the change of gas permeability in the coal seam. Figs. 11.14 and 11.15 depict the evolution of stress and gas permeability, respectively, along the horizontal direction of the coal seam C13 with the mining of the coal seam B11. It can be seen that there is considerably wide changes in the stress and gas permeability. When the face advances half a day (or 2.5 m in distance), the ground stress in the protected coal seam C13 decreases slightly. A stress-relief area is produced and the gas permeability of the coal seam increases. However, the range of the stress-relief area is small and the degree of pressure relief is low, indicating that the relief effect on the protected coal seam

Table 11.2

Physical, mechanical and seepage parameters used in the numerical model

No.

Litho logy

Elastic modulus E (Gap)

Uniaxial compressive strength Rc (MPa)

Uniaxial tensile strength Rt (MPa)

Poisson ratio λ

Friction angle (degrees)

Cohesion (MPa)

Layer thickness (m)

1

Sandy mudstone

27.0

63

6

0.4

29

5.0

20.0

2

C13 coal seam

2

20

2

0.4

30

8

6.0

3

Siltstnoe

31.1

60

6

0.27

30.5

9.4

6.0

4

Siltstnoe

27.21

68

5

0.12

25.6

11.46

12.0

5

Mediumgrain sandstone

28.05

100.2

8

0.12

30.5

25.7

12.0

6

Sandy mudstone

36.44

60

6

0.38

26.1

7.48

15.0

7

Mudstone

27.58

86.5

4

0.21

20.9

10.68

4

8

Fine sandstone

19.18

48.72

5

0.40

24.4

9.0

12.0

9

Sandy mudstone

57.4.

30.61

3

0.26

37

4.8

6

10

B11b coal seam

2

20

2

0.4

30

8

2

11

Sandy mudstone

12.0

70

7

0.26

26

4.0

14

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Figure 11.13 Simulated caving process of the strata and correspondingly induced shear stress fields.

Figure 11.14 Maximum principal stress changes as the face advances for various periods.

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Figure 11.15 Gas permeability changes as the face advances for various periods, (A) from the beginning of mining to 7 days, (B) from 7 to 10 days.

is small. At the same time, a stress concentration region appears in the protected coal seam. When the working face continuously advances for about 7 days (or 35 m in distance), the overlying coal or rock mass subsides. The stress-relief area of the protected coal seam C13 increases above the gob, the in situ stress decreases and the seam permeability further increases. The magnitude and range of the stress concentration also increase slightly. When the coal face advances for 10 days, the stress-relief area in the protected coal seam further increases and the overburden strata above the face caves. It is found that the magnitude of gas permeability in the coal seam C13 increases by 2100 times as compared to the initial gas permeability. Moreover, the horizontal range of the coal seam with an obvious

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Figure 11.16 The gas permeability changes in the coal seam C13 along the A-A0 section in Fig. 11.12 with different sensitivity parameter β.

increase of gas permeability is up to 70 m when the stress sensitivity factor β is 2.0. The effect of the pressure-relief and the increase of gas permeability are obvious in the protected coal seam C13. In contrary, when the stress sensitivity factor β is 1.6 but other stress conditions are the same, the magnitude of the gas permeability change in the coal seam C13 is 1000 times, as compared to the initial gas permeability and the horizontal range of the coal seam with an obvious increase of the gas permeability is about 50 m. It is concluded that the stress sensitivity factor β is an important factor for modeling the coupled gas flow in coal seams during coal mining. Compared with the field observations, it may be suitable to use the stress sensitivity factor β 5 2.0 for modeling the coupled gas flow in coal seams for the case study. Fig. 11.16 shows the gas permeability changes in the coal seam C13 along the A-A0 section in Fig. 11.12 with different sensitivity parameter β 5 2.0 and β 5 1.6. Meanwhile, these results show that the in-situ stress in the coal seam C13 is greatly decreased by mining the coal seam B11. Pressure relief and tensile deformation are induced by mining, and a large number of bed separation cracks are created in the coal seam C13. The field measurement results are in agreement with the numerical simulation.

11.3.2.4 Subsidence of coal seam and associated gas flow The simulated maximum subsidence of the coal seam are 1400, 1500, and 1600 mm as the face advances 45, 70, and 100 m, respectively, and the range of the stress relief zone in the overburden strata is about 50 m. The numerical results agree well with the field observations depicted in Fig. 11.9. Fig. 11.17 shows the change of the gas pressure in the coal seam C13. The gas pressure around the gas drainage boreholes sharply decreases from pre-drainage to post drainage. For example, it decreases from 5 MPa at pre-drainage to 2.2 MPa at post-drainage after ten days’ gas drainage and 100 m of face advances. Moreover, the decreasing range of gas pressure gradually expands as the days of the gas

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Figure 11.17 The change of gas pressure in the coal seam C13 as the face advances.

drainage extends and continuation of face advance. The reason is that the coal seam C13 above the gob falls into the mining induced stress release zone, where the gas permeability of the coal seam increases significantly, promoting the gas flow in coal seam and leading to an obvious decrease of the gas pressure around the gas drainage boreholes. This is called the effect of reducing the stress to increase the permeability. The simulated results and the field observations in Figs. 11.11 and 11.17 show that the modeled dynamic process of the pressure relief gas drainage during coal mining on the whole agrees with that in field observations, and pressure relief gas drainage during coal mining is an effective technique applied in mining areas with multiple high-gassy coal seams of low gas permeability.

11.3.2.5 Analyses and discussions This section revisited the engineering practices and conducted the numerical simulations for the long-distance pressure relief gas drainage in the coal seam B11 at the Panyi coal mine, where there is about 67 m spacing between the coal seams B11 and C13. Through the fields observations, it was found that the gas pressure around the borehole in the coal seam C13 decreased from 4.4 to 0.4 MPa; the gas content in the coal seam C13 decreased from 13 to 5 m3/t; the gas permeability of the coal seam C13 increased from 2.84 3 1024 to 0.817 md by an increase of 2800 times; dilatational deformation of the coal seam B11 reached up to 26.33m. Comparing with some other panels without the application of the technique, the roadway driving speed in the fully-mechanized longwall top coal caving mining zone increased from the original 4060 m/month to more than 200 m/month. Moreover, the average panel production improved from 1700 to 5100 tons/day (i.e. with an increase of two times) and the relative gas emission of the heading face decreased from 25 to 5.0 m3/min (i.e., with a decrease of five times). Under equal air-conditions in the roadways, the average gas concentration in return airways decreased from 1.15% to

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0.5%. In addition, the average monthly panel production can reach up to 7000 tons/ day based on the current capacity of the gas drainage and the ventilation in the top caving panel. The numerical simulations of the case study reproduced the mining-induced deformation, subsidence and fracture characteristics of overburden strata, and the associated gas flow in the coal seam C13. Moreover, the numerical modeling results revealed the effect of the in-situ stress on gas permeability of the coal seam C13. According to the numerical simulations, the height of the caving zone is between 15 and 20 m and the height of fractured zone is between 40 and 45 m in the overburden strata above the B11 coal seam as the face advances to about 100 m. Due to the extraction of the coal seam B11, the coal seam C13 subsides on a whole in the mining induced sagging zone of the overburden strata. The modeled deformation of the coal seam is basically in accordance with that in field observations. When the face has advanced for 10 days, an obvious increase of the gas permeability is observed in the horizontal range of up to 70 m in the coal seam C13. The magnitude of gas permeability increase is up to about 2100 times higher than the initial gas permeability in the coal seam C13. The predicted increase is close to that observed in the field, which is about 2800 times, considering the simplification here and complex of the problem. The extremely large increase of gas permeability indicated that the pressure relief and the tensile deformation are induced by coal mining and a large number of bed separations are created in the coal seam C13. In addition, the numerical modeling results indicated that the gas pressure around the gas drainage boreholes sharply decreased from 5 to 2.2 MPa after 10 days’ gas drainage and 100 m of face advance. The decreasing range of the gas pressure gradually expanded as the gas drainage extends and the face advances. Generally, the modeled dynamic process of the pressure relief gas drainage during coal mining on the whole agrees well with those in field observations. Moreover, it was found that the gas pressure in the coal seam B11 decreased from 4.4 to 0.5 MPa; the gas content decreased from 13 to 5 m3/t; the gas permeability increased from 2.84 3 1024 to 0.817 md by an increase of 2880 times; the dilatational deformation reached up to 26.33m. Both the field observations and the numerical simulations revealed that the protective coal seam mining method should be adopted in combination with the pressure relief gas drainage technique in high gassy and outburst coal mine. As mentioned above, the method obviously eliminates the outburst hazards and also changes the coal seam C13 from a high gas and outburst coal seam into a low nonoutburst coal seam in the case study. Moreover, the method may lead to high production and high efficiency for high gassy and outburst coal mine.

11.4

Conclusions

RFPA method was used to simulate the deformation and fracture characteristics of overburden strata, the evolution of gas permeability, and the gas flow in the protected coal seam, in which stress, damage, and seepage coupling are taken into

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account. In this chapter, two case studies are presented. The case study on fracturing and caving of overburden strata induced by longwall mining was performed to study the overburden movement and associated abutment pressures distribution over coal face using RFPA code, which considers the heterogeneity and nonlinear characteristics of overburden rock strata. Numerical results show that as the face advances, bed separation and the bending occur in the adjacent upper strata, cracks initiates at mid span of lower side and at both ends of the upper side of the upper roof. The upper strata ruptures and caves when the span reaches to a certain limit. The overburden movement gradually propagates from the goaf to the upper strata. Generally the immediate roof shortly caves with the advancing of the face. The main roof ruptures and caves when the face advances to a certain distance and the first weighting occurs. Thereafter, the overburden rock strata caves in periodically as the face advances. Meanwhile, numerical simulations shows that the equilibrium of stress in overburden is disturbed due to coal mining and that the abutment pressure over coal face is redistributed. The abutment pressure over coal face is characterized with low stress zone, high stress zone, and in situ stress zone along the mining direction. The front abutment pressures over coal face are higher than back abutment pressures over coal face due to the continuous effect of mining excavation from coal face. Numerical simulations show that RFPA model well reproduces the overburden movement process, reveal the rupturing mechanism of overburden strata and captures periodic weighting of the roof even though there is not any preexisting cracks or fractures in the model. Another case study on the pressure relief gas drainage by mining the protective coal seam at Panyi coal mine, Huainan coal mining Co. Ltd., were numerically performed to evaluate the effectiveness of the long-distance pressure relief gas drainage during extraction of the protective coal seams at great depths. The numerical results visualized the three zones (i.e., caving, fractured, and bending zones) in the overburden strata and three stress zones (i.e., stress concentration, stress relaxed and intact stress zones) in the protective coal seam induced by coal mining. The numerical simulations generally agree well with the field observations, especially, the increasing ratio of the gas permeability and the range of the pressure relief induced by the protective coal seam mining. Moreover, the numerical simulations indicate that the extraction of the coal seams causes the deformation, bed separation, and caving of overburden strata at a large scale, and the large enhancement of the gas permeability of the protected coal seam, which in turn leads to obvious decrease of the gas pressure around the boreholes in the surrounding rock mass. It is demonstrated that the technique of the pressure relief gas drainage by longdistance boreholes greatly alleviates the risk of coal and gas outbursts in the highgassy coal seam of the low gas permeability, such as the coal seam C13 in the case study. In addition, the technique also significantly reduces the gas content in the coal seam C13 and the gas emission at the face. Correspondingly, a safe and highefficient mining is secured in the main protected coal seam C13. The numerical results offered some important theoretical indications of the mechanism of the gas flow in the coal seam. Furthermore, the results also provided some practical instructions of using the technique of the pressure relief gas drainage in the highly gassy

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coal seam of low gas permeability to achieve the safe and high efficient mining in fully mechanized top coal caving and prevent the occurrences of coal and gas outbursts in underground coal mining. Therefore, it is concluded that RFPA method is of significant help in better understanding of coal strata movement characteristics and the associated gas flow mechanism in both theory and practice. It is pointed out that the long-distance pressure relief gas drainage technology can effectively improve the safety and productivity in underground coal mines, especially in the mining areas with the multiple highly gassy coal seams of low permeability.

Acknowledgments We thanks for Prof. Syd S. Peng for his careful reading, revision and improvement of the manuscript. The joint support provided by NSFC (Grant No. 51474051, 41672301, 41172265, 51404067) and the National Basic Research Program of China (2013CB227900, 2014CB047100) and Fundamental Research Funds for the Central Universities (N150102002) is highly acknowledged.

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