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Application of solid backfilling to reduce hard-roof caving and longwall coal face burst potential
International Journal of Rock Mechanics & Mining Sciences 88 (2016) 197–205
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International Journal of Rock Mechanics & Mining Sciences journal homepage: www.elsevier.com/locate/ijrmms
Application of solid backfilling to reduce hard-roof caving and longwall coal face burst potential Jixiong Zhang, Baiyi Li, Nan Zhou n, Qiang Zhang Key Laboratory of Deep Coal Resource Mining, School of Mines, Ministry of Education of China, China University of Mining and Technology, Xuzhou 221116, China
art ic l e i nf o Article history: Received 1 August 2015 Received in revised form 25 April 2016 Accepted 31 July 2016 Available online 5 August 2016 Keywords: Solid backfilling mining Hard roof Face burst Energy evolution Filling ratio
1. Introduction A hard roof refers to the thick strata above a coal seam or above a thin immediate roof.1 Such a roof generally has high strength with few joints. After each cutting, instead of caving occurring immediately, the roof will overhang for a significant length in the gob, resulting in high abutment stress in front of the face and slow down gas release from the seam. As a result, the longwall face spalling and roof hang up occurs frequently, and the unwanted gas accumulates in the gob. With continuing advance of the longwall face and the increase of overhang area, the stress in the hard roof eventually exceeds its ultimate strength and the roof caves suddenly, which causes a rapid release of energy stored in the roof and coal seam. This often results in the coal mine bursts and/or other catastrophic dynamic events such as wind blasting, causing serious damage to mining equipment, significant mining delays and sometimes casualties. Thus, the hard roof is one of the main factors causing coal mine bursts at the longwall face.2,3 There is wide variation in the geology of hard roofs in coal mines in China. For example, the thickness varies from tens to hundreds of meters. However, what is of great importance is that coal reserves under hard roofs account for about one-third of the total reserve in China; moreover, nearly 40% of fully mechanized coal mining panels have hard roofs and more than 50% of mining n
Corresponding author. E-mail address: [email protected] (N. Zhou).
http://dx.doi.org/10.1016/j.ijrmms.2016.07.025 1365-1609/& 2016 Elsevier Ltd. All rights reserved.
areas suffer from problems associated with hard roofs.4 With regard to these problems, researchers both in China and other countries have proposed many solutions, such as using sacrificial coal pillars, weakening roof strength via water injection or fracturing, and forced caving, to prevent longwall face bursts caused by hard roofs. In some cases, these approaches have been effective. However, owing to the wide variability in geological conditions of the hard roofs among coal mines, the above solutions have limited applicability and new approaches to address this challenge are urgently needed.5,6 In recent years, backfill coal mining technique has become a popular and widely applied mining method to safely extract coal resources trapped under buildings, railways, and water bodies (hereinafter referred to as “three-unders”). Backfill mining has both a firm theoretical basis and a well-developed technology in terms of operating methods and equipment.7–10 Numerous case studies have demonstrated that backfill coal mining can effectively control movement of both the roof and the overlying strata. In comparison with the traditional longwall caving method, the backfill coal mining can greatly reduce both the abutment stress around the excavation surface.11–13 For these reasons, in this paper, a solid backfill method for controlling hard-roof-induced face bursts is proposed. The types and mechanisms of hard-roof-induced face bursts, as well as means to mitigate them, are described from the perspective of energy evolution. The interaction between the solid backfill body and the roof under different roofcontrolled backfilling ratios is analyzed. Through the use of a
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Table 1 Classification and control of mining-induced dynamic hazards under a hard roof. Type
Time
Coal-body-compression Before roof fracture Coal-body-rebound During roof fracture Roof-facture During roof fracture
Cause
Control
Stress concentration in advance of the face, strain energy accumulated in coal–rock body exceeds its limit Rapid release of stain energy in coal–rock body ahead of faceline
Decentralize stress at the face, reduce accumulated strain energy Decentralize stress in advance, and reduce strain-energy release during roof fracture Reduce the release of gravitational potential energy of the roof
Gravitational potential energy of the roof is partially transformed into radiation energy and transferred to coal–rock body at critical state of energy accumulation
mechanical model, the deformation characteristics of the hard roof and the energy evolution of the panel for different roof-controlled backfilling ratios are determined, revealing the effectiveness of solid backfilling for minimizing the energy stored in the seam. A field trial is carried out at Panel 6304–1 of the Jisan Coal Mine and the trial outcome will be discussed in detail using the monitoring data.
in the form of a vibration wave. If the accumulated strain energy in the coal/rock mass ahead of the faceline has reached and remains a critical value, the additional contribution from the radiation energy may cause the strain energy to exceed this critical value, causing instability and resulting in a rooffacture face burst. 2.2. Keys to control of hard-roof-induced face bursts
2. Types of hard-roof-induced face bursts and their control 2.1. Types and causes of hard-roof-induced face bursts A large number of studies show that coal mine bursts result from a process of energy accumulation, transformation, and release proceeding from exposure of the roof to its fracture. The precondition for the coal mine bursts to occur in a panel with a hard roof is the presence of the concentrated stress at the coal– rock body,14,15 together with accumulation of strain energy and rapid energy release when the hard roof breaks. Depending on the type of energy that triggers the bursts, the coal mine bursts can be divided into three types, namely coal-body-compression, coalbody-bounce, and roof-fracture bursts, which are explained as follows: (1) Coal-body-compression face burst. As the face continues to advance, the roof loses the support from the underneath coal seam and begins to deform and bend. Strain energy is stored inside the bending hard roof, and, owing to the compression, a large amount is also stored inside coal body in front of the longwall face. The stored energy slowly releases through transformation into surface energy and energy released. When the strain energy accumulated inside the coal body becomes too large, its rate of release increases and part of the strain energy transforms into kinetic energy of the coal (rock) body, thereby causing a coal-body-compression face burst. (2) Coal-body-rebound burst. As bending and deformation reach their limitation, the roof breaks. The gravitational potential energy of the broken strata together with strain energy accumulated inside the roof and coal seam are rapidly released and transformed into kinetic energy, surface energy, and radiation energy, causing a large area of weighting on the roof. At this point, the coal body ahead of the face rebounds owing to stress changes, and the strain energy accumulated owing to compression is instantaneously released, with some being transformed into surface energy and radiation energy, and some into kinetic energy of the coal body. If the kinetic energy is small, spalling of the face will occur, but if it is sufficiently large, there will be a coal-body-rebound burst. (3) Roof-facture burst. The roof vibrates when it fractures. Moreover, when the roof and floor make contact, this results in vibration and rebound. A portion of the gravitational potential energy and strain energy is transformed into energy released and transferred to the roof ahead of the faceline and coal body
From the causes of the three types of hard-roof-induced bursts described above, it can be seen that a burst is actually a kind of dynamic events caused by ejection of the coal/rock mass from its free face as a result of the release of kinetic energy instantaneously transformed from strain energy stored in the coal–rock body and energy released by deformation or fracture of the roof. The basic cause is accumulation or release of a large amount of strain energy inside the coal–rock body. The method of control of a burst may vary depending on the specific type (Table 1). As shown in Table 1, the key to control of hard-roof-induced bursts is to reduce the strain energy accumulated inside the coal/ rock body and the energy released during roof caving.16–18
3. Effects of solid backfilling in preventing hard-roof-induced face bursts 3.1. Principle of solid backfilling As an integrated technology, the backfilling system was developed to handle solid backfilling materials based on the original fully mechanized coal mining system. The gangue, dune sand, and other solids to be used as backfilling materials are fixed first on the surface and then transported typically via belt conveyor and vertical wells to the underground. At underground, there is another set of belt conveyor to transport the solids to the backfill face. Sometime water is added to reduce the dust and void volume of the mixture, but most time no water is needed. The additives can also be used to improve backfill material strength if needed. Compared with traditional hydraulic supports, backfill hydraulic supports involve the addition of critical parts, such as a rear roof canopy, a backfilling scraped conveyor, and a tamping arm, as shown in Fig. 1. The backfill hydraulic supports are key to the success of solid backfilling: its front canopy supports the roof, providing a safe space for operating mining machines, and its back canopy offers the space needed for transporting the backfill material, and for dumping and compacting it in the gob. The backfilling scraper conveyor used for transporting the solid backfilling materials is hung below the rear roof canopy. Another important piece of equipment in the rear part of the backfill hydraulic support is the tamping arm, which can provide a pressure of 2 MPa to push the backfilling materials into the gob and compact them to a sufficient density to support the roof effectively.19
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Fig. 2. Mechanical model of nonuniform segmented beam on an elastic foundation.
Fig. 1. Backfilling hydraulic support at the working face.
3.2. Interaction between solid backfill body and hard roof The placement of backfill material has an immediate impact on the deformation of the main roof at the backfilling face. Therefore, the roof-controlled backfilling ratio, which is defined as the ratio of the height of the backfilled body to the mining height when the gravitational weight of the main roof and its overlying soft strata constitutes the applied load on the backfilled mass, is adopted to describe the backfilling effect. It directly represents the controlling effect of the backfilled body on the roof. Depending on the interaction between the roof and the backfilled body at different roofcontrolled backfilling ratios, the backfilling status can be classified as follows: 3.2.1. No contact between the backfilled body and the roof before roof caving When the roof-controlled backfilling ratio is small, a large gap will exist between the backfilled body and the roof, and the roof may reach its ultimate deflection limit without coming into contact with the backfilled body. In this case, the backfilled body will play no role in supporting the roof before the roof caves. Just like the traditional longwall mining with natural caving, the roof will exhibit obvious first and periodic weightings. After the roof has caved, the backfilled material will restrict its further subsidence. 3.2.2. Contact between the backfilled body and roof before roof caving For this case, as the roof-controlled backfilling ratio increases, the roof will make contact with the backfilled body before it caves. However, as the supporting effect of the backfilled body is limited, the roof will still cave as the face continues to advance. In this case, the solid backfilled body will support the roof before it caves, thereby changing the roof caving interval. 3.2.3. No roof caving With further increase of the roof-controlled backfilling ratio, the backfill body will begin to bear the roof load, restricting the deformation of the roof to such an extent that it bends continuously without caving and will only be subject to local cracking instead of complete caving. In this case, the roof at the face no longer shows obvious first and periodic weighting events, and the deformation characteristics are quite different from those of traditional longwall mining. 3.3. Controlling effect of solid backfill body on disaster-causing energy
disaster-causing energy when the solid backfilling technique is employed. 3.3.1. Development and solution of a mechanical model for roof deformation For all three backfilling conditions for roof control, the coal and backfill bodies are both represented simply as an elastic foundation,20 while the hard roof is represented as a beam. A mechanical model of the hard roof can then be constructed for each condition before and after failure of the beam. In this paper, the condition with contact between the backfilled body and the roof before roof failure is taken as an example to construct a mechanical model of the hard roof before the first weighting (Fig. 2); other models will not be discussed here. Taking the midpoint of the roof right above the longwall cutting face as the origin as illustrated in Fig. 2, a coordinate system is adopted with the face advancing direction as the x-axis and the vertical downside direction as the ω axis. The remaining notation in Fig. 2 is as follows: q0 is the in situ stress, kd is the stress concentration factor of the front abutment pressure, qc is dead load of the hard roof and the soft strata lying on top of the hard roof, L is the supporting area of the solid backfilled body, L0 is the zone of influence of the front abutment pressure, kc and kg are the coefficient of elastic foundation of the coal and the backfilled body, respectively, and ω1(x ) and ω2(x ) are the deflections of the backfilled body and the roof, respectively. Based on Winkler’s elastic foundation beam theory,21 the differential equations describing roof stress equilibrium can be written as:
⎧ d 4w1(x) ⎪ EtIt + kgw1(x) = qc , −L/2 ≤ x < 0 ⎪ dx 4 ⎪ 4 ⎪ ⎨ EtIt d w2(x) + k cw2(x) 0 ≤ x ≤ L0 ⎪ dx 4 ⎪ (kd − 1)q0 ⎪ = kdq0 − x, ⎪ L ⎩
(1)
0
where EtIt is the flexural rigidity of the hard roof, with Et being Young's modulus of the roof and It the second moment of area of the beam representing the roof. The deflection equation of the roof can be obtained as
⎧ qc αx ⎪ w1(x) = e [C1 cos(αx) + C2 sin(αx)] + kg ⎪ ⎪ +e−αx[C3 cos(αx) + C4 sin(αx)], −L/2 ≤ x < 0 ⎪ ⎪ ⎨ kdq0 ⎪ w2(x) = e−βx[C5 cos(βx) + C6 sin(βx)] + kc ⎪ ⎪ k 1 q ( − ) d 0 ⎪ x, 0 ≤ x ≤ L 0 − ⎪ kL ⎩
(2)
c 0
The key to control of hard-roof-induced face bursts is to reduce the accumulated disaster-causing energy and its release. A mechanical model has been developed to show the changes in
where C1, …, C6 are unknown parameters, and the characteristic coefficients are given by α =
4
kg 4EtIt
and β =
4
kc 4EtIt
.
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The relationship between the bending moment M (x ), the deflection angle θ(x ), the shear force Q (x ) and the deflection ω(x ) at any place in the roof are:
⎧ dω(x) θ (x) = ⎪ dx ⎪ ⎪ d2ω(x) ⎨ M (x) = − EtIt dx2 ⎪ ⎪ 3 d ω(x) ⎪ Q (x) = − EtIt ⎪ ⎩ dx 3
(3)
Q 1( − L/2) = 0
(4)
The continuity condition at x ¼0 is
ω1(0) = ω2(0),
M1(0) = M2(0)
θ1( 0) = θ2( 0),
Q 1(0) = Q 2(0)
(5)
where M1 and M2 are the bending moments of the roof above the backfilling body and coal, respectively, θ1 and θ2 are the corner angles of the roof above the backfilling body and coal, respectively, and Q1 and Q 2 are the shear forces on the roof above the backfilling body and coal, respectively. Eqs. (2)–(5) can be solved for the unknown parameters C1, …, C6 . Because the resulting expressions are very long, they are not given here explicitly. Instead, in subsequent calculations, the notation C1, …, C6 will continue to be used. On substituting the expressions for these parameters into Eqs. (2) and (3), the bending moment, corner angle, shear force, and deflection at any place in the roof can be calculated. Given the tensile strength of the hard roof, its fracture distance can be calculated. 3.3.2. Strain-energy solutions for different roof-controlled backfilling ratios The deformation of the roof is mainly bending deformation. A micro-element of the bending beam is subjected to a bending moment and a shear force, with the strain energy produced by the shear force being negligible compared with the bending strain energy.22 The strain energy in a micro-element of the cross section with a neutral-axis offset y is
Δυ =
M2( x) 2EI 2
y2
(6)
The strain energy density per unit width of the hard roof is ht
υt =
∫− h2
t
2
M2( x) 2EtIt 2
1 k cω2(x) 2
(9)
The compressional deformation strain energy in the coal is given by
Vc =
The boundary condition at x = − L/2 is
θ1( − L/2) = 0,
Δυc =
∫0
x
vc dx
(10)
On substituting the roof deflection ω2(x ) above the coal into Eq. (10), the coal compressional deformation energy in the range from 0 to x can be obtained. The relationship between the elastic modulus kg of the backfilled body and the roof-controlled backfilling ratio φk is as follows:
(
kg = qc / hg − hc φk
)
(11)
where qc is the dead load of the hard roof and the overlying soft strata, hg is the initial backfilling height, and hc is the mining height. For different roof-controlled backfilling ratios, the above calculations give the caving length of the roof and the bending strain energy of the roof before failure, as well as equations for computing the compressive strain energy of the coal seam and the strain-energy release and gravitational potential energy during roof failure. These equations will not be given in this paper. 3.4. Case study 3.4.1. Roof caving length Using Panel No. 6304-1 of the Jisan Coal Mine as an example, by performing laboratory tests on coal and rock properties and site test on the abutment pressure, the following parameters are obtained: hard-roof thickness ht ¼ 41.6 m and Young’s modulus Et ¼17 GPa; mining height hc ¼ 3.5 m and Young’s modulus Ec ¼ 5.25 GPa; tensile strength st of coal ¼ 13.5 MPa; the coefficient of elastic foundation coal seam kc ¼ 1.5 GPa; zone of influence of front abutment pressure L0 ¼ 100 m; in situ stress q0 ¼ 16.5 MPa; dead load of hard roof and overlying soft strata qc ¼ 2.27 MPa; kd ¼ 2.5. By substituting these values into the equations, the relationship between the hard roof caving length and the roofcontrolled backfilling ratio is obtained as shown in Fig. 3. The following observations can be seen from Fig. 3. When the roof-controlled backfilling ratio φk < 82.5% , the roof will not come into contact with the backfilled body before caving occurs, and thus the backfilling does not affect roof caving. The caving length is 151.5 m. When 82.5% ≤ φk < 93% , the roof comes into contact with the backfilled body before caving occurs, and the backfilling
y2 dy (7)
where ht is the thickness of the roof and M(x) is the bending moment on the roof. The bending strain energy is
Vt =
∫0
x
M2( x) dx 2EtIt
(8)
On substituting the expression for the bending moment obtained above into Eq. (8), the total bending strain energy in the range from 0 to x will be obtained. The strain energy in coal is mainly compressive deformation energy. A coal seam can be represented simply as an elastic body, and the compressive deformation of the coal can then be regarded as uniaxial compression of this elastic body. Because the compressional deformation of the coal is equal to the roof deflection ω(x ) above the coal, the strain-energy density per unit width is given by
Fig. 3. Relationship between the caving interval of the hard roof and the roofcontrolled backfilling ratio.
J. Zhang et al. / International Journal of Rock Mechanics & Mining Sciences 88 (2016) 197–205
Fig. 4. Relationship between strain energy and roof-controlled backfilling ratio ahead of the face before caving of the hard roof.
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Fig. 5. Release of strain energy in the roof ahead of the face for different roofcontrolled backfilling ratios.
can control roof deformation to a certain degree before caving. As φk increases, the roof caving interval increases; when φk reaches 93%, the roof caving interval reaches 213.4 m, which is 40.9% more than its value of 151.5 m at φk ¼ 82.5%. When φk ≥ 93% , the backfilled body can effectively support the roof, which no longer caves but mainly bends. 3.4.2. Strain-energy accumulation of coal–rock body ahead of the face before roof caving Fig. 4 shows the strain-energy stored in the coal/rock body ahead of the longwall face for different roof-controlled backfilling ratios before determination of the roof caving interval. When the roof-controlled backfilling ratio is zero, i.e., there is no backfilling in the gob, the roof is managed by the natural caving method. According to Fig. 4, when the solid backfilling method is adopted, with increasing roof-controlled backfilling rate, the change in energy of the coal–rock system ahead of the faceline falls into three stages, which correspond to the three types of roof deformation that occur at the face when the solid backfilling method is employed. When φk < 82.5% , backfilling has no impact on the energy ahead of the faceline. The magnitudes of the strain energy ahead of the faceline are the same as those when the natural caving method is employed. In this case, the strain energy accumulated in a coal–rock system of unit width is 37.6 MJ. When 82.5% ≤ φk < 93% , the strain energy ahead of the faceline decreases significantly. When φk reaches 93%, the strain energy accumulated in a coal–rock system of unit width falls to 19.5 MJ, a decrease of 48.1%. When 82.5% ≤ φk < 93% , the strain energy ahead of the faceline decreases slightly with increasing φk ; moreover, the strain energy falls from 20.57 to 19.5 MJ, resulting in a decrease of 5.2%. As φk increases above 93%, the decrease in strain energy ahead of the faceline accelerates. When φk reaches 99%, the maximum strain energy is 14.9 MJ, which is 23.6% less than that at φk ¼ 93%. 3.4.3. Energy release at the face when the roof caves When the roof caves, the releases of strain energy inside the coal–rock body ahead of the face and the gravitational potential energy of the roof in the gob are as shown in Figs. 5 and 6, respectively. It can be seen Figs. 5 and 6 that as φk increases from 0% to 82.5%, the backfilling has a slight impact on the strain-energy release ahead of the face. The energy release is 11.0 MJ, which is similar to that when the natural longwall caving method is employed. When 82.5% ≤ φk < 93% , the strain-energy release ahead of the faceline decreases. When φk reaches 93%, the strain energy release is 5.71 MJ, a reduction of 48.0%. However, as φk increases from 82.5% to 93%, the strain-energy release shows no obvious changes. When φk ≥ 93% , the roof will not cave. So no energy is
Fig. 6. Release of gravitational potential energy of the roof in the backfilled area for different roof-controlled backfilling ratios.
released ahead of the faceline. When the roof caves for the first time as φk increases, the gravitational potential energy release in the roof inside the backfilling area decreases linearly. When φk reaches 93%, the gravitational potential energy release is 64.8 MJ, which is 89.7% less than the value of 630.0 MJ when the natural caving method is employed. When φk ≥ 93% , the roof will not cave. Hence, no gravitational potential energy is released. 3.5. Effects of solid backfilling in preventing different kinds of face bursts According to the above analysis, the effect of solid backfilling in preventing hard-roof-induced face bursts changes in a stage-bystage manner with increasing roof-controlled backfilling ratio φk when the solid backfilling method is employed for mining of panels with hard roofs. When φk < 82.5% , the roof does not come into contact with the backfilled body before caving occurs. In this case, the backfilling can only restrict roof subsidence after caving, with no impact on the caving interval or on changes in energy ahead of the face before and after caving. It can reduce the risk of roof-caving face bursts, but cannot prevent coal-body-compression and coal-bodyrebound face bursts. When 82.5% ≤ φk < 93% , the backfilled body supports the roof before caving. In this case, as φk increases, both the accumulated strain energy ahead of the face before caving and the strain energy and gravitational potential energy releases during caving decrease, greatly reducing the risk of all three types of face bursts. When φk ≥ 93% , the backfilled body can prevent roof caving. In
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this case, the accumulated strain energy at the face is effectively controlled, with no release of strain energy and gravitational potential energy occurring during roof caving. The panel is at no risk of hard-roof-induced face bursts. The above considerations indicate that when the solid backfilling mining method is adopted for panels with a hard roof, the backfilling will have a better effect in reducing disaster-causing energy and preventing face bursts when the roof-controlled backfilling ratio is high. When this ratio reaches a certain value, the hard roof no longer caves, and the panel is not at risk of face bursts. Therefore, when the solid backfilling method is adopted to prevent face bursts, the roof-controlled backfilling ratio should be as high as possible. For the conditions encountered in Panel 6304-1 of the Jisan Coal Mine, face burst risks can be dramatically reduced when the roof-controlled backfilling ratio is greater than 93%.
4. Field application 4.1. Mining and geological conditions of the test site The original Panel No. 6304 that adopted fully mechanized top coal caving technology was located to the north of the main and auxiliary shafts. It was 250 m wide by 2200 m long. Because of the risk of face bursts due to the hard roof and the need to protect the Nanyang Lake Dam, mining of the panel was halted on May 25, 2008, with the remaining length of 548 m unmined and about 652,000 tons of coal left, resulting in a huge waste of coal resources. Three panels were planned in the pilot area. In order to maximize the recovery ratio, the method of backfill mining without pillar support was adopted. The first panel, No. 6304-1, was designed to be 80 m wide by 548 m long, with a recoverable reserve of 230,000 tons; the average panel depth was about 660 m (Fig. 7). By performing burst tendency tests on the coal and rocks in conjunction with an overall consideration of the geological and mining technology factors influencing the face bursts, it was determined that Panel No. 6304-1 had a medium risk of face bursts. The hardness coefficient of No. 3 coal is f ¼1–2, and its density is 1.36 t/m3. The height of the coal seam is about 3.5 m. It basically has no immediate roof. The main roof is fine sandstone of 32.5– 49.75 m thick (with an average of 41.6 m) having a hardness of 8– 10, tensile strength of 13.5 MPa, and elastic modulus of 17.0 GPa, can be regarded as a hard roof. The geological features of the roof and floor are shown in Table 2. 4.2. Measurement and analysis of the panel's backfilling quality 4.2.1. Measurements of the ratio of backfilled mass to mined mass In order to monitor the backfilling quality of the panel, the mass of backfilled waste rock and mined coal is monitored continuously. The backfilled-to-mined coal mass ratio is used to evaluate the backfilling effect. The mass ratio of backfilled to mined coal ( τ ) is related to the roof-controlled backfilling ratio φk by
τ = φkCg /Cc
(12)
Fig. 7. Mine layout around Panel No. 6304-1.
was mostly greater than the design value of 0.93, with an average of 1.19. Through back-calculation, the average roof-controlled backfilling ratio φk was actually 96.4%. 4.2.2. Measurement of roof subsidence During mining, a roof deformation-monitoring device was installed at the rear end of the shield support. It was connected via cables to a data acquisition instrument near the take-off location of the panel, so that changes in roof subsidence were monitored in real time. Three monitoring stations were deployed at the locations of 10, 30, and 60 m from the cutting hole. Each station had five monitoring devices uniformly distributed in the backfilling area. Taking the measurements at the 60 m location for instance, the largest roof subsidence was obtained, as shown in Fig. 9. When the maximum subsidence of the hard roof reached 197.8 mm, the roof-controlled backfilling ratio at location 60 m was around 94.3%. 4.3. Face burst preventing effect
3
where Cg ¼ 16.46 kN/m is the unit weight of the backfilled body obtained under a pressure of 2.27 MPa and Cc ¼ 13.33 kN/m3 is the unit weight of the coal. Based on monitoring data of the amount of backfilled waste rock and coal mined during the period of March 15, 2011 to September 15, 2011, the mass ratio of backfilled to mined coal of Panel No. 6304-1 was shown in Fig. 8, which shows that the mass ratio
4.3.1. Measurements of stresses in the surrounding rocks Three monitoring stations were deployed in each of maingate and tailgate of Panel No. 6304-1. At each monitoring point, four different drillhole depths (i.e. 3 m, 5 m, 10 m, and 15 m) were designed with a total of twenty-four borehole stress meters in total. Taking the measurements of the station 120 m from the
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Table 2 No. 3 coal seam roof and floor. Mine area
Rock category
Thickness/m
Rock features
Main roof
Fine- and medium- grained sandstone
Immediate roof Immediate floor Main floor
Mudstone Aluminous mudstone Fine-grained sandstone
32.5–49.75 41.63 0.0–1.02 0.0–3.20 2.7–8.43 5.85
Gray–white, largely quartz, followed by feldspar and a small amount of dark and green minerals, f ¼ 8–10 Brown–gray, containing many plant root fossils, f¼ 2–3 Light gray, smooth, containing plant fossil fragments, f¼ 2–3 Light gray–dark gray, dense and hard, f¼ 6–8
setup room as an example, the front abutment pressure in the coal block at different depths is shown in Fig. 10, which shows that the backfilled body restricted the deformation of the hard roof, mitigating the effects of mining. The stress concentration factor of the front abutment pressure was 1.44 and the influence distance was about 28 m, both of which were far less than those of the Jisan Coal Mine when the natural caving method was adopted under similar conditions. 4.3.2. Measurement of coal dust in drillholes The amount of drill dust for a certain drilling depth has been used as an important index to evaluate coal outburst potential. In this work, this index was used to determine the effectiveness of backfill mining method in reducing coal outburst risks. Before mining of Panel No. 6304, coal dust was first measured using a drill hole of 10 m in depth with a diameter of 42 mm, and then another four coal dust measurements were conducted during mining of Panel No. 6304-1. Table 3 show the specific locations and results of the five measurements. Table 3 shows that before mining of Panel No. 6304, a hole were drilled at 15 m in front of the face. When the drilling depth reached 5.0 m, the coal dust concentration was 48.96 kg/m; when the depth reached 7.0 m, it was 32.64 kg/m, which is 5.5 kg/m more than the critical threshold value for that time. When the drilling depth reached 4.5 m, the drill was jammed, accompanied by a coal blast, and the dust particles grew in size. When the drilling depth reached 7.5 m, the drill became stuck and the face showed an obvious risk of face bursts. However, during mining of Panel No. 6304-1, the concentrations of coal dust measured at the two stations did not exceed the critical value. The drilling operation was very smooth, with no abnormal phenomena, so coal outburst risks were reduced dramatically.
Fig. 8. Variation of the mass ratio of backfilled waste to mined coal.
Fig. 9. Measured roof subsidence in backfilling area.
4.3.3. Measurement of microseismic energy During mining of Panel No. 6304-1, a microseismic monitoring system was employed to measure microseismic energy. At the same time, Panel No. 16,305, located about 1700 m to the north of Panel No. 6304-1, was mining using the fully mechanized top coal caving mining method. The mining depth and seam dip angle for Panel No. 16,305 were similar to those of Panel No. 6304-1. The coal seam was 3.2–5.8 m thick, with an average of 4.8 m; the main roof was the medium sandstone with a thickness of 10.95–21.3 m. The measured microseismic energy distribution of the two panels
Fig. 10. Measured abutment pressures. Table 3 Measurement of coal dust at drillholes.
Drilling depth/m Coal dust/kg/m
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
In front of 15 m
1.80
2.12
2.45
2.94
48.96
38.08
32.64
–
–
–
Group 1
In front of 18 m In front of 24 m
2.33 3.93
2.57 2.57
3.06 2.82
3.18 3.06
3.67 3.22
3.92 3.67
4.16 3.92
4.53 5.26
5.26 5.26
6.24 5.75
Group 2
In front of 14 m In front of 18 m
1.80 1.63
2.28 2.12
2.77 2.61
3.75 3.43
4.24 4.08
4.57 4.57
5.06 4.73
5.55 5.39
5.88 5.71
6.36 6.20
Before mining of panel 6304 During mining of panel No. 6304–1
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Fig. 11. Energy distribution of microseismicity during mining.
Table 4 Characteristics of microseismicity during mining of Panels 6304-1 and 16,305. Energy range
Number of microseismic events, 6304-1
Number of microseismic events, 16,305
Less than 1000 J 1000–3000 J 3000–5000 J 5000–10,000 J Greater than 10,000 J Total number Max. microseismic energy/J Average microseismic energy/J
202 51 10 2 0 265 8760
1143 949 165 72 15 2344 39,000
733.2
1428.6
is shown in Fig. 11. Table 4 shows the characteristics of the microseismicity. Fig. 11 and Table 4 show that in Panel No. 6304-1, with solid backfilling mining, there were only two microseismic events with energy greater than 5000 J, the biggest one being of 8760 J. Moreover, the two events were located near the gob of the original Panel No. 6304. In comparison, Panel No. 16,305, with fully mechanized top coal caving mining, had 87 microseismic events with energy greater than 5000 J, including 15 events with energy exceeding 10,000 J, the largest being as high as 39,000 J. Furthermore, the total number of microseismic events and the average energy of Panel No. 6304-1 were far lower than those of Panel No. 16,305. Based on the results described above, by employing the solid backfilling method, deformation of the hard roof can be well controlled and the panel energy release kept small, indicating that solid backfilling can minimize hard-roof-induced face burst risks.
5. Conclusions Concentrated stress in the coal–rock body, strain-energy accumulation, and rapid energy release during hard roof caving are the prerequisites for coal mine bursts. Based on the causes of accidents, hard-roof-induced bursts were divided into three types: coal-body-compression, coal-body-rebound, and roof-caving bursts. The key to burst control is to reduce the accumulation and release of disaster-causing energy. As the roof-controlled backfilling ratio increases, the interaction between the backfilled body and the roof falls into the following three conditions: (i) the roof does not come into contact with the backfilled body before caving; (ii) the roof comes into contact with the backfilled body before caving; (iii) the roof does not cave. The criteria for each condition was developed. By adopting a simplified representation of the coal seam and backfilled body as an elastic foundation and the roof as a beam,
mechanical models of the hard roof under different conditions were developed to determine the roof caving interval, the accumulated strain energy before caving, and the strain-energy and gravitational potential energy releases for different roof-controlled backfilling ratios. The effect of solid backfilling in preventing hard-roof-induced face bursts changes in a stage-by-stage manner. Based on the geological conditions of the trial site, when the roof-controlled backfilling ratio φk < 82.5% , backfilling only reduces the risk of roof-caving face bursts. It will not prevent the bursts due to coalbody compression or coal-body rebound. When 82.5% ≤ φk < 93% , the backfilled body will greatly reduce the risk of all three types of face bursts. When φk ≥ 93% , the risk of hard-roof-induced face bursts is minimal. After adopting the solid backfilling mining method in Panel No. 6304-1 of the Jisan Coal Mine, when the average value of φk reached 96.4%, the stress concentration factor decreased to 1.44, the concentration of coal dust at the panel no longer exceeded the critical value, and little microseismic energy was released. This shows the effectiveness of the solid backfilling method in preventing hard-roof-induced face bursts.
Acknowledgments This research was supported by the National Natural Science Foundation of China (51504238), the Fundamental Research Funds for the Central Universities (China University of Mining and Technology, 2014ZDPY02) and Qing Lan Project and National Key Basic Research Program of China (2013CB227905).
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Contents lists available at ScienceDirect
International Journal of Rock Mechanics & Mining Sciences journal homepage: www.elsevier.com/locate/ijrmms
Application of solid backfilling to reduce hard-roof caving and longwall coal face burst potential Jixiong Zhang, Baiyi Li, Nan Zhou n, Qiang Zhang Key Laboratory of Deep Coal Resource Mining, School of Mines, Ministry of Education of China, China University of Mining and Technology, Xuzhou 221116, China
art ic l e i nf o Article history: Received 1 August 2015 Received in revised form 25 April 2016 Accepted 31 July 2016 Available online 5 August 2016 Keywords: Solid backfilling mining Hard roof Face burst Energy evolution Filling ratio
1. Introduction A hard roof refers to the thick strata above a coal seam or above a thin immediate roof.1 Such a roof generally has high strength with few joints. After each cutting, instead of caving occurring immediately, the roof will overhang for a significant length in the gob, resulting in high abutment stress in front of the face and slow down gas release from the seam. As a result, the longwall face spalling and roof hang up occurs frequently, and the unwanted gas accumulates in the gob. With continuing advance of the longwall face and the increase of overhang area, the stress in the hard roof eventually exceeds its ultimate strength and the roof caves suddenly, which causes a rapid release of energy stored in the roof and coal seam. This often results in the coal mine bursts and/or other catastrophic dynamic events such as wind blasting, causing serious damage to mining equipment, significant mining delays and sometimes casualties. Thus, the hard roof is one of the main factors causing coal mine bursts at the longwall face.2,3 There is wide variation in the geology of hard roofs in coal mines in China. For example, the thickness varies from tens to hundreds of meters. However, what is of great importance is that coal reserves under hard roofs account for about one-third of the total reserve in China; moreover, nearly 40% of fully mechanized coal mining panels have hard roofs and more than 50% of mining n
Corresponding author. E-mail address: [email protected] (N. Zhou).
http://dx.doi.org/10.1016/j.ijrmms.2016.07.025 1365-1609/& 2016 Elsevier Ltd. All rights reserved.
areas suffer from problems associated with hard roofs.4 With regard to these problems, researchers both in China and other countries have proposed many solutions, such as using sacrificial coal pillars, weakening roof strength via water injection or fracturing, and forced caving, to prevent longwall face bursts caused by hard roofs. In some cases, these approaches have been effective. However, owing to the wide variability in geological conditions of the hard roofs among coal mines, the above solutions have limited applicability and new approaches to address this challenge are urgently needed.5,6 In recent years, backfill coal mining technique has become a popular and widely applied mining method to safely extract coal resources trapped under buildings, railways, and water bodies (hereinafter referred to as “three-unders”). Backfill mining has both a firm theoretical basis and a well-developed technology in terms of operating methods and equipment.7–10 Numerous case studies have demonstrated that backfill coal mining can effectively control movement of both the roof and the overlying strata. In comparison with the traditional longwall caving method, the backfill coal mining can greatly reduce both the abutment stress around the excavation surface.11–13 For these reasons, in this paper, a solid backfill method for controlling hard-roof-induced face bursts is proposed. The types and mechanisms of hard-roof-induced face bursts, as well as means to mitigate them, are described from the perspective of energy evolution. The interaction between the solid backfill body and the roof under different roofcontrolled backfilling ratios is analyzed. Through the use of a
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Table 1 Classification and control of mining-induced dynamic hazards under a hard roof. Type
Time
Coal-body-compression Before roof fracture Coal-body-rebound During roof fracture Roof-facture During roof fracture
Cause
Control
Stress concentration in advance of the face, strain energy accumulated in coal–rock body exceeds its limit Rapid release of stain energy in coal–rock body ahead of faceline
Decentralize stress at the face, reduce accumulated strain energy Decentralize stress in advance, and reduce strain-energy release during roof fracture Reduce the release of gravitational potential energy of the roof
Gravitational potential energy of the roof is partially transformed into radiation energy and transferred to coal–rock body at critical state of energy accumulation
mechanical model, the deformation characteristics of the hard roof and the energy evolution of the panel for different roof-controlled backfilling ratios are determined, revealing the effectiveness of solid backfilling for minimizing the energy stored in the seam. A field trial is carried out at Panel 6304–1 of the Jisan Coal Mine and the trial outcome will be discussed in detail using the monitoring data.
in the form of a vibration wave. If the accumulated strain energy in the coal/rock mass ahead of the faceline has reached and remains a critical value, the additional contribution from the radiation energy may cause the strain energy to exceed this critical value, causing instability and resulting in a rooffacture face burst. 2.2. Keys to control of hard-roof-induced face bursts
2. Types of hard-roof-induced face bursts and their control 2.1. Types and causes of hard-roof-induced face bursts A large number of studies show that coal mine bursts result from a process of energy accumulation, transformation, and release proceeding from exposure of the roof to its fracture. The precondition for the coal mine bursts to occur in a panel with a hard roof is the presence of the concentrated stress at the coal– rock body,14,15 together with accumulation of strain energy and rapid energy release when the hard roof breaks. Depending on the type of energy that triggers the bursts, the coal mine bursts can be divided into three types, namely coal-body-compression, coalbody-bounce, and roof-fracture bursts, which are explained as follows: (1) Coal-body-compression face burst. As the face continues to advance, the roof loses the support from the underneath coal seam and begins to deform and bend. Strain energy is stored inside the bending hard roof, and, owing to the compression, a large amount is also stored inside coal body in front of the longwall face. The stored energy slowly releases through transformation into surface energy and energy released. When the strain energy accumulated inside the coal body becomes too large, its rate of release increases and part of the strain energy transforms into kinetic energy of the coal (rock) body, thereby causing a coal-body-compression face burst. (2) Coal-body-rebound burst. As bending and deformation reach their limitation, the roof breaks. The gravitational potential energy of the broken strata together with strain energy accumulated inside the roof and coal seam are rapidly released and transformed into kinetic energy, surface energy, and radiation energy, causing a large area of weighting on the roof. At this point, the coal body ahead of the face rebounds owing to stress changes, and the strain energy accumulated owing to compression is instantaneously released, with some being transformed into surface energy and radiation energy, and some into kinetic energy of the coal body. If the kinetic energy is small, spalling of the face will occur, but if it is sufficiently large, there will be a coal-body-rebound burst. (3) Roof-facture burst. The roof vibrates when it fractures. Moreover, when the roof and floor make contact, this results in vibration and rebound. A portion of the gravitational potential energy and strain energy is transformed into energy released and transferred to the roof ahead of the faceline and coal body
From the causes of the three types of hard-roof-induced bursts described above, it can be seen that a burst is actually a kind of dynamic events caused by ejection of the coal/rock mass from its free face as a result of the release of kinetic energy instantaneously transformed from strain energy stored in the coal–rock body and energy released by deformation or fracture of the roof. The basic cause is accumulation or release of a large amount of strain energy inside the coal–rock body. The method of control of a burst may vary depending on the specific type (Table 1). As shown in Table 1, the key to control of hard-roof-induced bursts is to reduce the strain energy accumulated inside the coal/ rock body and the energy released during roof caving.16–18
3. Effects of solid backfilling in preventing hard-roof-induced face bursts 3.1. Principle of solid backfilling As an integrated technology, the backfilling system was developed to handle solid backfilling materials based on the original fully mechanized coal mining system. The gangue, dune sand, and other solids to be used as backfilling materials are fixed first on the surface and then transported typically via belt conveyor and vertical wells to the underground. At underground, there is another set of belt conveyor to transport the solids to the backfill face. Sometime water is added to reduce the dust and void volume of the mixture, but most time no water is needed. The additives can also be used to improve backfill material strength if needed. Compared with traditional hydraulic supports, backfill hydraulic supports involve the addition of critical parts, such as a rear roof canopy, a backfilling scraped conveyor, and a tamping arm, as shown in Fig. 1. The backfill hydraulic supports are key to the success of solid backfilling: its front canopy supports the roof, providing a safe space for operating mining machines, and its back canopy offers the space needed for transporting the backfill material, and for dumping and compacting it in the gob. The backfilling scraper conveyor used for transporting the solid backfilling materials is hung below the rear roof canopy. Another important piece of equipment in the rear part of the backfill hydraulic support is the tamping arm, which can provide a pressure of 2 MPa to push the backfilling materials into the gob and compact them to a sufficient density to support the roof effectively.19
J. Zhang et al. / International Journal of Rock Mechanics & Mining Sciences 88 (2016) 197–205
199
Fig. 2. Mechanical model of nonuniform segmented beam on an elastic foundation.
Fig. 1. Backfilling hydraulic support at the working face.
3.2. Interaction between solid backfill body and hard roof The placement of backfill material has an immediate impact on the deformation of the main roof at the backfilling face. Therefore, the roof-controlled backfilling ratio, which is defined as the ratio of the height of the backfilled body to the mining height when the gravitational weight of the main roof and its overlying soft strata constitutes the applied load on the backfilled mass, is adopted to describe the backfilling effect. It directly represents the controlling effect of the backfilled body on the roof. Depending on the interaction between the roof and the backfilled body at different roofcontrolled backfilling ratios, the backfilling status can be classified as follows: 3.2.1. No contact between the backfilled body and the roof before roof caving When the roof-controlled backfilling ratio is small, a large gap will exist between the backfilled body and the roof, and the roof may reach its ultimate deflection limit without coming into contact with the backfilled body. In this case, the backfilled body will play no role in supporting the roof before the roof caves. Just like the traditional longwall mining with natural caving, the roof will exhibit obvious first and periodic weightings. After the roof has caved, the backfilled material will restrict its further subsidence. 3.2.2. Contact between the backfilled body and roof before roof caving For this case, as the roof-controlled backfilling ratio increases, the roof will make contact with the backfilled body before it caves. However, as the supporting effect of the backfilled body is limited, the roof will still cave as the face continues to advance. In this case, the solid backfilled body will support the roof before it caves, thereby changing the roof caving interval. 3.2.3. No roof caving With further increase of the roof-controlled backfilling ratio, the backfill body will begin to bear the roof load, restricting the deformation of the roof to such an extent that it bends continuously without caving and will only be subject to local cracking instead of complete caving. In this case, the roof at the face no longer shows obvious first and periodic weighting events, and the deformation characteristics are quite different from those of traditional longwall mining. 3.3. Controlling effect of solid backfill body on disaster-causing energy
disaster-causing energy when the solid backfilling technique is employed. 3.3.1. Development and solution of a mechanical model for roof deformation For all three backfilling conditions for roof control, the coal and backfill bodies are both represented simply as an elastic foundation,20 while the hard roof is represented as a beam. A mechanical model of the hard roof can then be constructed for each condition before and after failure of the beam. In this paper, the condition with contact between the backfilled body and the roof before roof failure is taken as an example to construct a mechanical model of the hard roof before the first weighting (Fig. 2); other models will not be discussed here. Taking the midpoint of the roof right above the longwall cutting face as the origin as illustrated in Fig. 2, a coordinate system is adopted with the face advancing direction as the x-axis and the vertical downside direction as the ω axis. The remaining notation in Fig. 2 is as follows: q0 is the in situ stress, kd is the stress concentration factor of the front abutment pressure, qc is dead load of the hard roof and the soft strata lying on top of the hard roof, L is the supporting area of the solid backfilled body, L0 is the zone of influence of the front abutment pressure, kc and kg are the coefficient of elastic foundation of the coal and the backfilled body, respectively, and ω1(x ) and ω2(x ) are the deflections of the backfilled body and the roof, respectively. Based on Winkler’s elastic foundation beam theory,21 the differential equations describing roof stress equilibrium can be written as:
⎧ d 4w1(x) ⎪ EtIt + kgw1(x) = qc , −L/2 ≤ x < 0 ⎪ dx 4 ⎪ 4 ⎪ ⎨ EtIt d w2(x) + k cw2(x) 0 ≤ x ≤ L0 ⎪ dx 4 ⎪ (kd − 1)q0 ⎪ = kdq0 − x, ⎪ L ⎩
(1)
0
where EtIt is the flexural rigidity of the hard roof, with Et being Young's modulus of the roof and It the second moment of area of the beam representing the roof. The deflection equation of the roof can be obtained as
⎧ qc αx ⎪ w1(x) = e [C1 cos(αx) + C2 sin(αx)] + kg ⎪ ⎪ +e−αx[C3 cos(αx) + C4 sin(αx)], −L/2 ≤ x < 0 ⎪ ⎪ ⎨ kdq0 ⎪ w2(x) = e−βx[C5 cos(βx) + C6 sin(βx)] + kc ⎪ ⎪ k 1 q ( − ) d 0 ⎪ x, 0 ≤ x ≤ L 0 − ⎪ kL ⎩
(2)
c 0
The key to control of hard-roof-induced face bursts is to reduce the accumulated disaster-causing energy and its release. A mechanical model has been developed to show the changes in
where C1, …, C6 are unknown parameters, and the characteristic coefficients are given by α =
4
kg 4EtIt
and β =
4
kc 4EtIt
.
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The relationship between the bending moment M (x ), the deflection angle θ(x ), the shear force Q (x ) and the deflection ω(x ) at any place in the roof are:
⎧ dω(x) θ (x) = ⎪ dx ⎪ ⎪ d2ω(x) ⎨ M (x) = − EtIt dx2 ⎪ ⎪ 3 d ω(x) ⎪ Q (x) = − EtIt ⎪ ⎩ dx 3
(3)
Q 1( − L/2) = 0
(4)
The continuity condition at x ¼0 is
ω1(0) = ω2(0),
M1(0) = M2(0)
θ1( 0) = θ2( 0),
Q 1(0) = Q 2(0)
(5)
where M1 and M2 are the bending moments of the roof above the backfilling body and coal, respectively, θ1 and θ2 are the corner angles of the roof above the backfilling body and coal, respectively, and Q1 and Q 2 are the shear forces on the roof above the backfilling body and coal, respectively. Eqs. (2)–(5) can be solved for the unknown parameters C1, …, C6 . Because the resulting expressions are very long, they are not given here explicitly. Instead, in subsequent calculations, the notation C1, …, C6 will continue to be used. On substituting the expressions for these parameters into Eqs. (2) and (3), the bending moment, corner angle, shear force, and deflection at any place in the roof can be calculated. Given the tensile strength of the hard roof, its fracture distance can be calculated. 3.3.2. Strain-energy solutions for different roof-controlled backfilling ratios The deformation of the roof is mainly bending deformation. A micro-element of the bending beam is subjected to a bending moment and a shear force, with the strain energy produced by the shear force being negligible compared with the bending strain energy.22 The strain energy in a micro-element of the cross section with a neutral-axis offset y is
Δυ =
M2( x) 2EI 2
y2
(6)
The strain energy density per unit width of the hard roof is ht
υt =
∫− h2
t
2
M2( x) 2EtIt 2
1 k cω2(x) 2
(9)
The compressional deformation strain energy in the coal is given by
Vc =
The boundary condition at x = − L/2 is
θ1( − L/2) = 0,
Δυc =
∫0
x
vc dx
(10)
On substituting the roof deflection ω2(x ) above the coal into Eq. (10), the coal compressional deformation energy in the range from 0 to x can be obtained. The relationship between the elastic modulus kg of the backfilled body and the roof-controlled backfilling ratio φk is as follows:
(
kg = qc / hg − hc φk
)
(11)
where qc is the dead load of the hard roof and the overlying soft strata, hg is the initial backfilling height, and hc is the mining height. For different roof-controlled backfilling ratios, the above calculations give the caving length of the roof and the bending strain energy of the roof before failure, as well as equations for computing the compressive strain energy of the coal seam and the strain-energy release and gravitational potential energy during roof failure. These equations will not be given in this paper. 3.4. Case study 3.4.1. Roof caving length Using Panel No. 6304-1 of the Jisan Coal Mine as an example, by performing laboratory tests on coal and rock properties and site test on the abutment pressure, the following parameters are obtained: hard-roof thickness ht ¼ 41.6 m and Young’s modulus Et ¼17 GPa; mining height hc ¼ 3.5 m and Young’s modulus Ec ¼ 5.25 GPa; tensile strength st of coal ¼ 13.5 MPa; the coefficient of elastic foundation coal seam kc ¼ 1.5 GPa; zone of influence of front abutment pressure L0 ¼ 100 m; in situ stress q0 ¼ 16.5 MPa; dead load of hard roof and overlying soft strata qc ¼ 2.27 MPa; kd ¼ 2.5. By substituting these values into the equations, the relationship between the hard roof caving length and the roofcontrolled backfilling ratio is obtained as shown in Fig. 3. The following observations can be seen from Fig. 3. When the roof-controlled backfilling ratio φk < 82.5% , the roof will not come into contact with the backfilled body before caving occurs, and thus the backfilling does not affect roof caving. The caving length is 151.5 m. When 82.5% ≤ φk < 93% , the roof comes into contact with the backfilled body before caving occurs, and the backfilling
y2 dy (7)
where ht is the thickness of the roof and M(x) is the bending moment on the roof. The bending strain energy is
Vt =
∫0
x
M2( x) dx 2EtIt
(8)
On substituting the expression for the bending moment obtained above into Eq. (8), the total bending strain energy in the range from 0 to x will be obtained. The strain energy in coal is mainly compressive deformation energy. A coal seam can be represented simply as an elastic body, and the compressive deformation of the coal can then be regarded as uniaxial compression of this elastic body. Because the compressional deformation of the coal is equal to the roof deflection ω(x ) above the coal, the strain-energy density per unit width is given by
Fig. 3. Relationship between the caving interval of the hard roof and the roofcontrolled backfilling ratio.
J. Zhang et al. / International Journal of Rock Mechanics & Mining Sciences 88 (2016) 197–205
Fig. 4. Relationship between strain energy and roof-controlled backfilling ratio ahead of the face before caving of the hard roof.
201
Fig. 5. Release of strain energy in the roof ahead of the face for different roofcontrolled backfilling ratios.
can control roof deformation to a certain degree before caving. As φk increases, the roof caving interval increases; when φk reaches 93%, the roof caving interval reaches 213.4 m, which is 40.9% more than its value of 151.5 m at φk ¼ 82.5%. When φk ≥ 93% , the backfilled body can effectively support the roof, which no longer caves but mainly bends. 3.4.2. Strain-energy accumulation of coal–rock body ahead of the face before roof caving Fig. 4 shows the strain-energy stored in the coal/rock body ahead of the longwall face for different roof-controlled backfilling ratios before determination of the roof caving interval. When the roof-controlled backfilling ratio is zero, i.e., there is no backfilling in the gob, the roof is managed by the natural caving method. According to Fig. 4, when the solid backfilling method is adopted, with increasing roof-controlled backfilling rate, the change in energy of the coal–rock system ahead of the faceline falls into three stages, which correspond to the three types of roof deformation that occur at the face when the solid backfilling method is employed. When φk < 82.5% , backfilling has no impact on the energy ahead of the faceline. The magnitudes of the strain energy ahead of the faceline are the same as those when the natural caving method is employed. In this case, the strain energy accumulated in a coal–rock system of unit width is 37.6 MJ. When 82.5% ≤ φk < 93% , the strain energy ahead of the faceline decreases significantly. When φk reaches 93%, the strain energy accumulated in a coal–rock system of unit width falls to 19.5 MJ, a decrease of 48.1%. When 82.5% ≤ φk < 93% , the strain energy ahead of the faceline decreases slightly with increasing φk ; moreover, the strain energy falls from 20.57 to 19.5 MJ, resulting in a decrease of 5.2%. As φk increases above 93%, the decrease in strain energy ahead of the faceline accelerates. When φk reaches 99%, the maximum strain energy is 14.9 MJ, which is 23.6% less than that at φk ¼ 93%. 3.4.3. Energy release at the face when the roof caves When the roof caves, the releases of strain energy inside the coal–rock body ahead of the face and the gravitational potential energy of the roof in the gob are as shown in Figs. 5 and 6, respectively. It can be seen Figs. 5 and 6 that as φk increases from 0% to 82.5%, the backfilling has a slight impact on the strain-energy release ahead of the face. The energy release is 11.0 MJ, which is similar to that when the natural longwall caving method is employed. When 82.5% ≤ φk < 93% , the strain-energy release ahead of the faceline decreases. When φk reaches 93%, the strain energy release is 5.71 MJ, a reduction of 48.0%. However, as φk increases from 82.5% to 93%, the strain-energy release shows no obvious changes. When φk ≥ 93% , the roof will not cave. So no energy is
Fig. 6. Release of gravitational potential energy of the roof in the backfilled area for different roof-controlled backfilling ratios.
released ahead of the faceline. When the roof caves for the first time as φk increases, the gravitational potential energy release in the roof inside the backfilling area decreases linearly. When φk reaches 93%, the gravitational potential energy release is 64.8 MJ, which is 89.7% less than the value of 630.0 MJ when the natural caving method is employed. When φk ≥ 93% , the roof will not cave. Hence, no gravitational potential energy is released. 3.5. Effects of solid backfilling in preventing different kinds of face bursts According to the above analysis, the effect of solid backfilling in preventing hard-roof-induced face bursts changes in a stage-bystage manner with increasing roof-controlled backfilling ratio φk when the solid backfilling method is employed for mining of panels with hard roofs. When φk < 82.5% , the roof does not come into contact with the backfilled body before caving occurs. In this case, the backfilling can only restrict roof subsidence after caving, with no impact on the caving interval or on changes in energy ahead of the face before and after caving. It can reduce the risk of roof-caving face bursts, but cannot prevent coal-body-compression and coal-bodyrebound face bursts. When 82.5% ≤ φk < 93% , the backfilled body supports the roof before caving. In this case, as φk increases, both the accumulated strain energy ahead of the face before caving and the strain energy and gravitational potential energy releases during caving decrease, greatly reducing the risk of all three types of face bursts. When φk ≥ 93% , the backfilled body can prevent roof caving. In
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this case, the accumulated strain energy at the face is effectively controlled, with no release of strain energy and gravitational potential energy occurring during roof caving. The panel is at no risk of hard-roof-induced face bursts. The above considerations indicate that when the solid backfilling mining method is adopted for panels with a hard roof, the backfilling will have a better effect in reducing disaster-causing energy and preventing face bursts when the roof-controlled backfilling ratio is high. When this ratio reaches a certain value, the hard roof no longer caves, and the panel is not at risk of face bursts. Therefore, when the solid backfilling method is adopted to prevent face bursts, the roof-controlled backfilling ratio should be as high as possible. For the conditions encountered in Panel 6304-1 of the Jisan Coal Mine, face burst risks can be dramatically reduced when the roof-controlled backfilling ratio is greater than 93%.
4. Field application 4.1. Mining and geological conditions of the test site The original Panel No. 6304 that adopted fully mechanized top coal caving technology was located to the north of the main and auxiliary shafts. It was 250 m wide by 2200 m long. Because of the risk of face bursts due to the hard roof and the need to protect the Nanyang Lake Dam, mining of the panel was halted on May 25, 2008, with the remaining length of 548 m unmined and about 652,000 tons of coal left, resulting in a huge waste of coal resources. Three panels were planned in the pilot area. In order to maximize the recovery ratio, the method of backfill mining without pillar support was adopted. The first panel, No. 6304-1, was designed to be 80 m wide by 548 m long, with a recoverable reserve of 230,000 tons; the average panel depth was about 660 m (Fig. 7). By performing burst tendency tests on the coal and rocks in conjunction with an overall consideration of the geological and mining technology factors influencing the face bursts, it was determined that Panel No. 6304-1 had a medium risk of face bursts. The hardness coefficient of No. 3 coal is f ¼1–2, and its density is 1.36 t/m3. The height of the coal seam is about 3.5 m. It basically has no immediate roof. The main roof is fine sandstone of 32.5– 49.75 m thick (with an average of 41.6 m) having a hardness of 8– 10, tensile strength of 13.5 MPa, and elastic modulus of 17.0 GPa, can be regarded as a hard roof. The geological features of the roof and floor are shown in Table 2. 4.2. Measurement and analysis of the panel's backfilling quality 4.2.1. Measurements of the ratio of backfilled mass to mined mass In order to monitor the backfilling quality of the panel, the mass of backfilled waste rock and mined coal is monitored continuously. The backfilled-to-mined coal mass ratio is used to evaluate the backfilling effect. The mass ratio of backfilled to mined coal ( τ ) is related to the roof-controlled backfilling ratio φk by
τ = φkCg /Cc
(12)
Fig. 7. Mine layout around Panel No. 6304-1.
was mostly greater than the design value of 0.93, with an average of 1.19. Through back-calculation, the average roof-controlled backfilling ratio φk was actually 96.4%. 4.2.2. Measurement of roof subsidence During mining, a roof deformation-monitoring device was installed at the rear end of the shield support. It was connected via cables to a data acquisition instrument near the take-off location of the panel, so that changes in roof subsidence were monitored in real time. Three monitoring stations were deployed at the locations of 10, 30, and 60 m from the cutting hole. Each station had five monitoring devices uniformly distributed in the backfilling area. Taking the measurements at the 60 m location for instance, the largest roof subsidence was obtained, as shown in Fig. 9. When the maximum subsidence of the hard roof reached 197.8 mm, the roof-controlled backfilling ratio at location 60 m was around 94.3%. 4.3. Face burst preventing effect
3
where Cg ¼ 16.46 kN/m is the unit weight of the backfilled body obtained under a pressure of 2.27 MPa and Cc ¼ 13.33 kN/m3 is the unit weight of the coal. Based on monitoring data of the amount of backfilled waste rock and coal mined during the period of March 15, 2011 to September 15, 2011, the mass ratio of backfilled to mined coal of Panel No. 6304-1 was shown in Fig. 8, which shows that the mass ratio
4.3.1. Measurements of stresses in the surrounding rocks Three monitoring stations were deployed in each of maingate and tailgate of Panel No. 6304-1. At each monitoring point, four different drillhole depths (i.e. 3 m, 5 m, 10 m, and 15 m) were designed with a total of twenty-four borehole stress meters in total. Taking the measurements of the station 120 m from the
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Table 2 No. 3 coal seam roof and floor. Mine area
Rock category
Thickness/m
Rock features
Main roof
Fine- and medium- grained sandstone
Immediate roof Immediate floor Main floor
Mudstone Aluminous mudstone Fine-grained sandstone
32.5–49.75 41.63 0.0–1.02 0.0–3.20 2.7–8.43 5.85
Gray–white, largely quartz, followed by feldspar and a small amount of dark and green minerals, f ¼ 8–10 Brown–gray, containing many plant root fossils, f¼ 2–3 Light gray, smooth, containing plant fossil fragments, f¼ 2–3 Light gray–dark gray, dense and hard, f¼ 6–8
setup room as an example, the front abutment pressure in the coal block at different depths is shown in Fig. 10, which shows that the backfilled body restricted the deformation of the hard roof, mitigating the effects of mining. The stress concentration factor of the front abutment pressure was 1.44 and the influence distance was about 28 m, both of which were far less than those of the Jisan Coal Mine when the natural caving method was adopted under similar conditions. 4.3.2. Measurement of coal dust in drillholes The amount of drill dust for a certain drilling depth has been used as an important index to evaluate coal outburst potential. In this work, this index was used to determine the effectiveness of backfill mining method in reducing coal outburst risks. Before mining of Panel No. 6304, coal dust was first measured using a drill hole of 10 m in depth with a diameter of 42 mm, and then another four coal dust measurements were conducted during mining of Panel No. 6304-1. Table 3 show the specific locations and results of the five measurements. Table 3 shows that before mining of Panel No. 6304, a hole were drilled at 15 m in front of the face. When the drilling depth reached 5.0 m, the coal dust concentration was 48.96 kg/m; when the depth reached 7.0 m, it was 32.64 kg/m, which is 5.5 kg/m more than the critical threshold value for that time. When the drilling depth reached 4.5 m, the drill was jammed, accompanied by a coal blast, and the dust particles grew in size. When the drilling depth reached 7.5 m, the drill became stuck and the face showed an obvious risk of face bursts. However, during mining of Panel No. 6304-1, the concentrations of coal dust measured at the two stations did not exceed the critical value. The drilling operation was very smooth, with no abnormal phenomena, so coal outburst risks were reduced dramatically.
Fig. 8. Variation of the mass ratio of backfilled waste to mined coal.
Fig. 9. Measured roof subsidence in backfilling area.
4.3.3. Measurement of microseismic energy During mining of Panel No. 6304-1, a microseismic monitoring system was employed to measure microseismic energy. At the same time, Panel No. 16,305, located about 1700 m to the north of Panel No. 6304-1, was mining using the fully mechanized top coal caving mining method. The mining depth and seam dip angle for Panel No. 16,305 were similar to those of Panel No. 6304-1. The coal seam was 3.2–5.8 m thick, with an average of 4.8 m; the main roof was the medium sandstone with a thickness of 10.95–21.3 m. The measured microseismic energy distribution of the two panels
Fig. 10. Measured abutment pressures. Table 3 Measurement of coal dust at drillholes.
Drilling depth/m Coal dust/kg/m
1.0
2.0
3.0
4.0
5.0
6.0
7.0
8.0
9.0
10.0
In front of 15 m
1.80
2.12
2.45
2.94
48.96
38.08
32.64
–
–
–
Group 1
In front of 18 m In front of 24 m
2.33 3.93
2.57 2.57
3.06 2.82
3.18 3.06
3.67 3.22
3.92 3.67
4.16 3.92
4.53 5.26
5.26 5.26
6.24 5.75
Group 2
In front of 14 m In front of 18 m
1.80 1.63
2.28 2.12
2.77 2.61
3.75 3.43
4.24 4.08
4.57 4.57
5.06 4.73
5.55 5.39
5.88 5.71
6.36 6.20
Before mining of panel 6304 During mining of panel No. 6304–1
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Fig. 11. Energy distribution of microseismicity during mining.
Table 4 Characteristics of microseismicity during mining of Panels 6304-1 and 16,305. Energy range
Number of microseismic events, 6304-1
Number of microseismic events, 16,305
Less than 1000 J 1000–3000 J 3000–5000 J 5000–10,000 J Greater than 10,000 J Total number Max. microseismic energy/J Average microseismic energy/J
202 51 10 2 0 265 8760
1143 949 165 72 15 2344 39,000
733.2
1428.6
is shown in Fig. 11. Table 4 shows the characteristics of the microseismicity. Fig. 11 and Table 4 show that in Panel No. 6304-1, with solid backfilling mining, there were only two microseismic events with energy greater than 5000 J, the biggest one being of 8760 J. Moreover, the two events were located near the gob of the original Panel No. 6304. In comparison, Panel No. 16,305, with fully mechanized top coal caving mining, had 87 microseismic events with energy greater than 5000 J, including 15 events with energy exceeding 10,000 J, the largest being as high as 39,000 J. Furthermore, the total number of microseismic events and the average energy of Panel No. 6304-1 were far lower than those of Panel No. 16,305. Based on the results described above, by employing the solid backfilling method, deformation of the hard roof can be well controlled and the panel energy release kept small, indicating that solid backfilling can minimize hard-roof-induced face burst risks.
5. Conclusions Concentrated stress in the coal–rock body, strain-energy accumulation, and rapid energy release during hard roof caving are the prerequisites for coal mine bursts. Based on the causes of accidents, hard-roof-induced bursts were divided into three types: coal-body-compression, coal-body-rebound, and roof-caving bursts. The key to burst control is to reduce the accumulation and release of disaster-causing energy. As the roof-controlled backfilling ratio increases, the interaction between the backfilled body and the roof falls into the following three conditions: (i) the roof does not come into contact with the backfilled body before caving; (ii) the roof comes into contact with the backfilled body before caving; (iii) the roof does not cave. The criteria for each condition was developed. By adopting a simplified representation of the coal seam and backfilled body as an elastic foundation and the roof as a beam,
mechanical models of the hard roof under different conditions were developed to determine the roof caving interval, the accumulated strain energy before caving, and the strain-energy and gravitational potential energy releases for different roof-controlled backfilling ratios. The effect of solid backfilling in preventing hard-roof-induced face bursts changes in a stage-by-stage manner. Based on the geological conditions of the trial site, when the roof-controlled backfilling ratio φk < 82.5% , backfilling only reduces the risk of roof-caving face bursts. It will not prevent the bursts due to coalbody compression or coal-body rebound. When 82.5% ≤ φk < 93% , the backfilled body will greatly reduce the risk of all three types of face bursts. When φk ≥ 93% , the risk of hard-roof-induced face bursts is minimal. After adopting the solid backfilling mining method in Panel No. 6304-1 of the Jisan Coal Mine, when the average value of φk reached 96.4%, the stress concentration factor decreased to 1.44, the concentration of coal dust at the panel no longer exceeded the critical value, and little microseismic energy was released. This shows the effectiveness of the solid backfilling method in preventing hard-roof-induced face bursts.
Acknowledgments This research was supported by the National Natural Science Foundation of China (51504238), the Fundamental Research Funds for the Central Universities (China University of Mining and Technology, 2014ZDPY02) and Qing Lan Project and National Key Basic Research Program of China (2013CB227905).
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