Analysis of Pre-Blasting Cracks in Horizontal Section Top-coal Mechanized Caving of Steep Thick Seams

In order to achieving safe and efficient exploitation in horizontal section top-coal mechanized caving of steep and thick seams, pre-blasting of top-coal is one of the prerequisites and analysis of crack evolution law is a key method to achieving good pre-splitting effects. Based on investigations of coal seams and mining conditions, theories of fracture mechanics were applied to explain the process of caving cracks and fracture toughness of coal seams in preblasting caving were calculated. The distribution of caving cracks was determined with in-situ borehole-wall real deformation optical monitoring systems. The results showed that the pre-splitting crack could rapidly develop in the direction of borehole center line and form the failure surface along the same direction in the last; the fracture toughness of B3 and B6 coal seams was 0.5616 and 1.1900 MPa·m 1/2 , respectively. The distribution of caving stress from real monitoring instruments provided a theoretical proof for optimizing the parameters of pre-blasting in top-coal safe mining.


INTRODUCTION
Mechanized Top-Coal Caving Technology (MTCCT) has a long history.From the early 1950s, MTCCT has been applied in countries such as the former Soviet Union, France, Poland, Yugoslavia and India.However, less than ideal results, foreign MTCCT began to shrink in the late 1980s (Oosthuizen and Esterhuizen, 1997;Zhang and Qian, 2003;Vakili and Hebblewhite, 2010).
In China, a series of tests on MTCCT had been conducted in Shenyang, Pingdingshan, Lu'an and Yangquan Mining Bureau since 1982 and after a number of technical setbacks, the thick coal seam caving efficiency and security issues were satisfactorily solved in the late 1980s, which promoted the technology to develop rapidly and now it has become a main method of thick seam mining in China (Xie et al., 1999;Ren, 2005;Wang et al., 2006;Liu et al., 2009).
Pre-blasting in top-coal caving is one of the prerequisites to achieving efficient and safe mining.Specifically, to transform top coal from the original state to caving state, three phases of complex processes are required, namely deformation, broken inflation and falling.Crack evaluation law analysis is a crucial method to achieve pre-blasting and weaken coal in steep and thick seams, which includes top-coal sturdiness coefficient testing, scope of loose top-coal, advanced stress influence, blasting parameters, equipment supporting, caving distance and coal spontaneous combustion in mined-out areas, determination of support method after pre-blasting, considering hole parameters, packaging quality, minimal resistance line, the interaction between caving in front of hydraulic mechanized support with initiation sequence, drilling angle and length in pre-blasting (T.H. Kang et al., 2004;Yasitli and Unver, 2005;Unver and Yasitli, 2006;Encina et al., 2010).
According to an investigation of the geological and mining conditions, based on +564 m B 3-6 in Jiangou Coal Mine, China, in order to obtain good pre-splitting effects and to ensure safe mining, optimization parameters of the crack expanding processes in preblasting were achieved by fracture mechanics and insitu stress monitoring

ENGINEERING SITUATION
Engineering geological environment and mining conditions: Pre-blasting lanes, +564 m B 3-6 coal seams, Jiangou Coal Mine, are layed out along strike and the center distance between lanes is 45.0 m.The total coal thickness is 50.0 m and its average angle is 86.5°.The methane levels are low, but there is still a hazard of coal dust explosion.The coal has a propensity for spontaneous combustion.As shown in Fig. 1, the seam thickness ranges from B 6 to B 3 respectively, four partings lie in B thickness of parting is between 0.15 and 0.20 m, thus, average thickness is 0.16 m.
Based on existing technologies and on identification, a pre-blasting program was developed as follows: In the top-coal, there were groups of blasting holes (fan-shaped, one-way vertical seams, diameter: 100 mm, blast-hole spacing: 4.0 m, 10 boreholes in each group) were arranged from B 3 to B 6 Lane.Pre was conducted to weaken the top-coal by latex matrix explosives.The protection or safe thickness of top should be 3.0 m, with 5.0 m width of coal pillar, using Charge Machine and yellow mud sealing.

In-situ real deformation monitoring:
borehole-wall optical systems were applied to observe crack degrees and the distribution of crack network.This information is gathered by scopes of testing devices to reveal the top-coal rupture characteristics and rules.The high-resolution probes and color display device were adopted to conduct in-situ real deformation monitoring, which can distinguish cracks correct to 1 mm.The computer can be connected d facilitate real-time image display and preservation, as shown in Fig. 3.

Calculation method of fracture toughness:
the pre-blasting borehole in the top wave in the wave-front surface generates a radial compressive stress and circumferential tensile stress, which causes a dynamic stress concentration at the borehole-wall along the borehole center line.As a result, the initial fracture can be prior formed by the tensile stress along the borehole center line direction Under the quasi-static stress field of the detonation gases, the pre-splitting borehole will produce the stress concentration area of the initial crack tip, leading to further expansion of the initial cracks As shown in Fig. 4. The process can be analyzed by a computational model of fracture mechanics and the crack can be simplified in the plane strain state.
In Fig. 1a, due to the velocity that static gases get into the crack is less than that of the crack expansion and the static gases effects in the cracks can be ignored.So Eq. 1 can be used to describe stress intensity factor at the crack tip: static stress field of the detonation splitting borehole will produce the stress concentration area of the initial crack tip, leading to further expansion of the initial cracks (Zong, 1998).
. The process of crack expansion can be analyzed by a computational model of fracture mechanics and the crack can be simplified in the plane In Fig. 1a, due to the velocity that static gases get into the crack is less than that of the crack expansion the static gases effects in the cracks can be ignored.So Eq. 1 can be used to describe stress intensity factor (1) static pressure of the borehole For the initial crack extension, it must satisfy Eq. 2: where, k Ic is rock fracture toughness.
When the crack extension length of the borehole is much longer than the radius, the calculation can be simplified by Fig. 2b.Then Eq. 3 can be used to describe the stress intensity factor of the crack tip stress field.
Under such condition, Eq. 3 could describe stress intensity factors at the crack tip: For the initial crack extension, it must satisfy Eq. 2: (2) When the crack extension length of the borehole is much longer than the radius, the calculation can be simplified by Fig. 2b.Then Eq. 3 can be used to describe the stress intensity factor of the crack tip stress Under such condition, Eq. 3 could describe stress According to fracture mechanics, the longer length of the initial crack, the easier it is to be developed.Hence, the crack will quickly expand to form the rupture surface along the borehole center line.
After emulsion matrix's detonation explosive pressure can be described as: The calculated quasi-static pressure is 10.02 Mpa.

RESULTS AND DISCUSSION
Borehole deformation analysis: mentioned, it is possible for us to observe various cracks in the No. 1-3 monitoring boreholes.Figure exhibits partial ones collected from field boreholes information, which presents crack degrees and the distribution of fracture network.
Statistics analysis of these cracks in each borehole were shown in Fig. 6.Above all, analysis of in deformation monitoring can be achieved: • Monitoring borehole was located at B and was influenced largely by hole 4 blasting.
Cracks developed apparently from 0 to 1.0 m, with vertical crack densely between 1.0 to 2.5 m, with a maximum of 4.5 m.Due to the blasting effects, the bore-hole with broken free surface, it ha crushed loose coal.Some cracks developed at a depth of around 2.5 m, which may cause roof caving. (3) According to fracture mechanics, the longer length of the initial crack, the easier it is to be developed.Hence, the crack will quickly expand to form the rupture surface along the borehole center line.emulsion matrix's detonation, the average explosive pressure can be described as: (4) According to the calculation, the outcome is 2025-3164 MPa.In addition, the equation of quasi-static pressure of the hole can be described as: (5) ritical pressure of explosive gas in expansion 1.4 static pressure is 10.02-10.36

RESULTS AND DISCUSSION
analysis: From all above mentioned, it is possible for us to observe various 3 monitoring boreholes.Figure 5 exhibits partial ones collected from field boreholes information, which presents crack degrees and the

Statistics analysis of these cracks in each borehole
. Above all, analysis of in-situ real deformation monitoring can be achieved: Monitoring borehole was located at B 4 coal (softer) and was influenced largely by hole 4 blasting.Cracks developed apparently from 0 to 1.0 m, with vertical crack densely between 1.0 to 2.5 m, with a maximum of 4.5 m.Due to the blasting effects, the hole with broken free surface, it had the crushed loose coal.Some cracks developed at a depth of around 2.5 m, which may cause roof   Calculation of fracture toughness: blasting being completed, a large amount of free surfaces brought about lamination crack phenomenon by blasting stress waves in No. 1 boreholes, which caused cracks to grow and develop well.Details of emulsion matrix can be seen from Table 2. Parameters of cracks in No. 1 boreholes were listed in Table 3.Because average length of crack was distinctly more than radius of holes, Eq. 3 was adopted for computing the fracture toughness and the results can be seen from Table 3.
Pre-blasting stress monitoring: the instruments were applied for the stress monitoring.On one side, it can be (a) From the working face, abutment pressure is not obvious from 0 to 10.0 m.(b) From 10.0 to 15.0 m, the pressure was 0.2 kept steady.Because of pre-blasting, abutment pressure rose suddenly to 2.5 MPa and then lessened a little towarded a new balance.On the other side, roof load at stope varied rarely in the explosive process: • From 0 to 8.0 m, the roof load went up gradually and its peak added up to 2.  3.Because average length of crack was distinctly more than radius of holes, Eq. 3 was adopted for computing the fracture toughness n be seen from Table 3.
As shown in Fig. 7, the instruments were applied for the stress monitoring.On one side, it can be (a) From the working face, abutment pressure is not obvious from 0 to 10.0 m.(b) m, the pressure was 0.2-1.0MPa and blasting, abutment pressure rose suddenly to 2.5 MPa and then lessened a little towarded a new balance.On the other side, roof load at stope varied rarely in the explosive process: From 0 to 8.0 m, the roof load went up gradually and its peak added up to 2.5 MPa.Based on results of the stress monitoring, 4.0 m roof canopy is used to control caving and sliding From 8.0 to 13.0 m, roof load increased sharply by ss and its peak added up to 3.5 MPa and then also lessened a little towards another new

CONCLUSION
After pre-blasting in steep and thick seam, as explosive effects cause dynamic stress focusing on pore wall along the central line, so initial cracks form along the central line, then under the quasi-static stress from explosive gases, initial cracks grows further.
The average length of crack is much more t radius of holes, the fracture toughness of B seams was 0.5616 and 1.1900 MPa•m which has been used to optimize parameters of pre blasting.
In-situ real monitoring results show that mining pressure is not fierce and its peak decreases obviously, abutment pressure moves toward working face and stress concentration zone lessens apparently.The average length of crack is much more than the radius of holes, the fracture toughness of B 3 and B 6 coal seams was 0.5616 and 1.1900 MPa•m 1/2 , respectively which has been used to optimize parameters of presitu real monitoring results show that mining peak decreases obviously, abutment pressure moves toward working face and stress concentration zone lessens apparently.

Fig. 1 :Fig. 2 :
Fig. 1: Geological section condition and thickness of B3 coal seams and thickness of B3-6 and working face; (a) blasting plan boreholes; (b) Physics model of working face during advancing , the seam thickness ranges from respectively, four partings lie in B 4 -B 5 and the thickness of parting is between 0.15 and 0.20 m, thus, ologies and on-site hazard blasting program was developed as coal, there were groups of blasting holes way vertical seams, diameter: 100 hole spacing: 4.0 m, 10 boreholes in each Lane.Pre-blasting coal by latex matrix explosives.The protection or safe thickness of top-coal should be 3.0 m, with 5.0 m width of coal pillar, using Charge Machine and yellow mud sealing.Plan of monitoring points' position: Fig. 2, No. 1-3 monitoring boreholes were located at 3.5 m high from the scraper bottom to top with 2.8 m away from No. 53 group borehole (Hole 3, hole 6 and hole 9) and 1.2 m from No. 53 group borehole (Hole 1, hole 4, hole 7 and hole 10) and 1.2 m from No. 53 group middle borehole (Hole 2, hole 5 and hole 8).Both No. 1 and No. 3 monitoring boreholes are 8 m deep and that of No. 2 monitoring borehole is 9 m.Furthermore, No. 1 monitoring borehole lies in B seam, 1.2 m away from No. 53 group and is intersect with hole 4 at 5.0-6.0 m.No. 2 monitoring borehole lies in B 5 seam, 1.2 m away from No. 53 group.No. 3 monitoring borehole lies in B 6 seam, 1.2 m away from No. 53 group and is intersect with hole 10 at 3.0 EXPERIMENTAL METHODS K I = Rock stress intensity factor P qs = Quasi-static pressure of the borehole osition: As shown in 3 monitoring boreholes were located at 3.5 m high from the scraper bottom to top-coal seam, with 2.8 m away from No. 53 group borehole (Hole 3, hole 6 and hole 9) and 1.2 m from No. 53 group minedborehole (Hole 1, hole 4, hole 7 and hole 10) and 1.2 m from No. 53 group middle borehole (Hole 2, hole 5 and hole 8).Both No. 1 and No. 3 monitoring boreholes are 8 m deep and that of No. 2 monitoring borehole is 9 m.borehole lies in B 4 seam, 1.2 m away from No. 53 group and is intersect 6.0 m.No. 2 monitoring borehole lies seam, 1.2 m away from No. 53 group.No. 3 seam, 1.2 m away from ect with hole 10 at 3.0-5.0m.EXPERIMENTAL METHODSonitoring: In situ wall optical systems were applied to observe crack degrees and the distribution of crack network.This information is gathered by scopes of testing coal rupture characteristics resolution probes and color display situ real deformation monitoring, which can distinguish cracks correct to 1 mm.The computer can be connected directly to time image display and preservation, asCalculation method of fracture toughness: After blasting borehole in the top-coal, blast stress front surface generates a radial stress and circumferential tensile stress, which causes a dynamic stress concentration at the wall along the borehole center line.As a result, the initial fracture can be prior formed by the tensile stress along the borehole center line direction.

Fig. 4 :
Fig. 3: In situ monitoring pictures: (a): Monitoring workers; (b): Monitoring instruments The final length of crack d = Diameter of the borehole m = Average explosive pressure ρ 0 = Density of emulsion matrix D = Velocity of emulsion matrix According to the calculation, the outcome is 2025 3164 MPa.In addition, the equation of quasi pressure of the hole can be described as: Critical pressure of explosive gas in expansion process, is 100 MPa d c = Diameter of dynamite n = A constant 3 k = The adiabatic coefficient 1.3-1.4

Fig. 5 :
Fig. 5: All kinds of borehole cracks (a): Oblique crack; Lateral crack; (c): Circular crack; and vertical crack • Monitoring borehole was located at B was harder than B 4 ) and hole 7 intersects with No. monitoring 2 borehole at 7.0-9.0m.Cracks are mainly focused on about 1.0 m and there was some broken coal from 5.5 to 6.5 m.Overall, the inner wall of No. 2 monitoring borehole maintained integrity.• Monitoring borehole was located at B was affected by hole 10 blasting.This monitoring borehole had dual circular cracks from 0 to 1.0 m, with broken section scattered between 0 to 3.0 m.

Fig. 6 :
Fig. 6: Statistical regularity of cracks with advent of borehole depthTable 2: Parameters of emulsion matrix Density (g/cm 3 ) Velocity (m/s) 5 MPa.• Based on results of the stress monitoring, 4.0 m roof canopy is used to control caving and sliding above • From 8.0 to 13.0 m, roof load increased sharply by fluctuating stress and its peak added up to 3.5 MPa and then also lessened a little towards another new balanceAfter preblasting being completed, a large amount of free about lamination crack phenomenon by blasting stress waves in No. 1-3 monitoring boreholes, which caused cracks to grow and develop atrix can be seen from Parameters of cracks in No. 1-3 monitoring boreholes were listed in Table Fig. 7: In-situ monitoring conditions; (a): in-situ monitoring; (b): Stress monitoring principle and field data acquisition Based on results of the stress monitoring, 4.0 m roof canopy is used to control caving and sliding above supports.Advanced supporting from 0 to 45.0 m is identical to the field situation.
This study was financially supported by the National Natural Science Foundation of China (51074140, 11002021), the China Scholarship Council (CSC) & the Hebei Provincial Office of Education (2010813124), the Doctoral Subject Foundation of the Ministry of Education of China (20070008012) and the ; (a): Layout of general Stress monitoring principle Based on results of the stress monitoring, 4.0 m roof canopy is used to control caving and sliding above supports.Advanced from 0 to 45.0 m is blasting in steep and thick seam, as ve effects cause dynamic stress focusing on pore wall along the central line, so initial cracks form along static stress from explosive gases, initial cracks grows further.
This study was financially supported by the National Natural Science Foundation of China na Scholarship Council (CSC) & the Hebei Provincial Office of Education (2010813124), the Doctoral Subject Foundation of the Ministry of Education of China (20070008012) and the National High Technology Research and Development Program (2008AA062104), all these are acknowleged.
Table 1 lists characteristics of roof and floor geological conditions.

Table 1 :
Physical and mechanical parameters of rock mass.

Table 3 :
Parameters of cracks in boreholes