Research Journal of Applied Sciences, Engineering and Technology 6(2): 249-253, 2013
ISSN: 2040-7459; e-ISSN: 2040-7467
© Maxwell Scientific Organization, 2013
Submitted: December 17, 2012
Accepted: February 01, 2013
Published: June 10, 2013
Analysis of Pre-Blasting Cracks in Horizontal Section Top-coal Mechanized
Caving of Steep Thick Seams
1
Shu-Ren Wang, 2, 3Xing-Ping Lai and 2, 3Peng-Fei Shan
School of Civil Engineering and Mechanics, Yanshan University, Qinhuangdao, China
2
School of Energy Engineering, Xi’an University of Science and Technology, Xi’an, China
3
Key Laboratory of Western Mine and Hazard Prevention, Ministry of Education, Xi’an, China
1
Abstract: 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·m1/2, 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.
Keywords: Fracture toughness, pre-blasting, steep thick seam, top-coal mechanized caving
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 B3-6 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 B3-6 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. Table 1 lists
characteristics of roof and floor geological conditions.
Correspondong Author: Shu-Ren Wang, School of Civil Engineering and Mechanics, Yanshan University, Qinhuangdao,
China, Tel.: 13273373298
249
Res. J. Appl.
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Sci. Engg. Technol., 6(22): 249-253, 20013
Table 1: Physiccal and mechanical parameters
p
of rock mass.
m
Density
L
Lithological
3
Lithology namee
Lithologicss
(kg/m )
trraits
Main roof
Shale
2380
H
Hard,
dark grey, layyer
structure
Immediate rooff
Siltstones
1440
G
Grey,
layer structurre,
jooint well
False roof
2450
S
Soft,
joint obvious
Shale
Immediate flooor
Shale
2600
Joint obvious, fragile
Fig. 1: Geollogical section condition and thickness of B33-6
coal seams
Plan of
o monitoring
g points’ position: As shoown in
Fig. 2, No. 1-3 monnitoring borehooles were locaated at
3.5 m high
h
from the scraper bottom
m to top-coal seam,
with 2.8 m away from
m No. 53 grouup borehole (H
Hole 3,
hole 6 and
a hole 9) and 1.2 m from No.
N 53 group minedm
borehole (Hole 1, holle 4, hole 7 andd hole 10) and 1.2 m
from No.
N 53 group middle
m
boreholee (Hole 2, holee 5 and
hole 8). Both No. 1 and
a No. 3 moniitoring boreholles are
8 m deep and that of No. 2 monitorring borehole is
i 9 m.
Furtherrmore, No. 1 monitoring borehole
b
lies in B4
seam, 1.2
1 m away frrom No. 53 grroup and is inttersect
with hoole 4 at 5.0-6.00 m. No. 2 monnitoring borehoole lies
in B5 seam,
s
1.2 m away
a
from No. 53 group. No. 3
monitoring borehole lies in B6 seam
m, 1.2 m awayy from
No. 53 group and is inntersect with hole
h 10 at 3.0-55.0 m.
EXPERIM
MENTAL ME
ETHODS
(a)
(b)
b
and working
w
face; (a)
Fig. 2: Pre-blasting plan boreholes
holes; (b) Physsics
Proffile of pre-blastting plan boreh
mod
del of working faace during advan
ncing
As shoown in Fig. 1, the
t seam thickkness ranges froom
B6 to B3 reespectively, fou
ur partings lie in B4-B5 and the
t
thickness of
o parting is between
b
0.15 and
a 0.20 m, thuus,
average thiickness is 0.16
6 m.
Basedd on existing teechnologies annd on-site hazaard
identificatiion, a pre-blastting program was
w developed as
follows:
In the top-coal, theree were groups of blasting holes
v
seamss, diameter: 100
(fan-shaped, one-way vertical
mm, blast--hole spacing: 4.0 m, 10 booreholes in eaach
group) werre arranged fro
om B3 to B6 Lane.
L
Pre-blastiing
was conduucted to weakeen the top-coall by latex matrix
explosives. The protectio
on or safe thickkness of top-cooal
should be 3.0 m, with 5.0 m width of coal pillar, usiing
Charge Maachine and yelllow mud sealinng.
In-situ
u real deformation mo
onitoring: Inn situ
borehole-wall opticall systems weree applied to obbserve
crack degrees
d
and thhe distributionn of crack nettwork.
This innformation is gathered byy scopes of testing
t
devicess to reveal thhe top-coal ruppture characteeristics
and rulles. The high-resolution probbes and color display
d
device were adopted to conduct in-ssitu real deform
mation
monitoring, which caan distinguish cracks correcct to 1
mm. The
T
computerr can be connnected directtly to
facilitatte real-time im
mage display and
a preservatiion, as
shown in Fig. 3.
Calcullation method of fracturee toughness: After
the pree-blasting boreehole in the toop-coal, blast stress
wave in
i the wave-front surface generates a radial
compreessive stress and
a
circumfereential tensile stress,
which causes a dynnamic stress concentration
c
at the
borehole-wall along the boreholee center line. As a
p
formed by
b the
result, the initial fraccture can be prior
t borehole center
c
line direection.
tensile stress along the
Under the quasi-stattic stress fieldd of the detoonation
t pre-splittinng borehole wiill produce the stress
gases, the
concenntration area of the initial crack
c
tip, leading to
further expansion of the
t initial crackks (Zong, 19988).
Ass shown in Fig.. 4. The processs of crack expaansion
can be analyzed by a computational model of frracture
mechannics and the crrack can be sim
mplified in thee plane
strain state.
s
In Fig. 1a, due to the velocity that static gasses get
into thee crack is lesss than that of the crack expaansion
and thee static gases efffects in the crracks can be ignored.
So Eq. 1 can be usedd to describe sttress intensity factor
at the crack
c
tip:
1
K   f  Pqs (  as ) 2
where,
KI = Rock
R
stress inteensity factor
Pqs = Quasi-static
Q
preessure of the boorehole
250 (1)
Res. J. Appl.
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Sci. Engg. Technol., 6(22): 249-253, 20013
K

 d  P qs (   a 1 )
1
2
(3)
where,
a1 = The
T final lengthh of crack
D
Diameter
of thee borehole
=
d
Acccording to fracture mechanics, the longer length
of the initial crack, the easier it is to be deveeloped.
Hence, the crack will
w quickly expand to form
m the
rupturee surface along the borehole center
c
line.
Affter emulsion matrix’s detoonation, the avverage
explosiive pressure caan be describedd as:
(a)
Pm 
(b)
Fig. 3: In siitu monitoring pictures:
p
(a): Monitoring workeers;
(b): Monitoring instrruments
(4)
1
0D 2
8
where,
Pm = Average
A
explossive pressure
ρ0 = Density
D
of emuulsion matrix
V
of emuulsion matrix
D = Velocity
Acccording to thee calculation, thhe outcome is 20253164 MPa.
M
In addittion, the equaation of quasii-static
pressurre of the hole can
c be describeed as:
 P
P qs   m
 Pk
(a)
(



k
n
 d 
 c 
 d 
2k
 Pk
(5)
where,
Pk = Critical
C
pressurre of explosivve gas in expaansion
prrocess, is 100 MPa
M
dc = Diameter
D
of dynnamite
n = A constant 3
k = The
T adiabatic cooefficient 1.3-11.4
Thhe calculated quasi-static
q
pressure is 10.02--10.36
Mpa.
(b)
RESULT
TS AND DISCU
USSION
Fig. 4: The computational model: (a): Beefore fracture; (b):
(
Afteer fracture
xtension
as = Lenngth of radial ex
f = Corrrection factor related with as and diameterr d
of thhe borehole
nsion, it must satisfy
s
Eq. 2:
For the inittial crack exten
K  Kc
(22)
where, kIc is rock fracturee toughness.
o the boreholee is
When the crack exteension length of
r
the callculation can be
much longger than the radius,
simplified by Fig. 2b. Then Eq. 3 can be used to
describe thhe stress intensity factor of thhe crack tip streess
field.
Underr such conditio
on, Eq. 3 could describe streess
intensity faactors at the craack tip:
Boreho
ole deforma
ation analysiss: From all above
mentionned, it is posssible for us to observe various
v
cracks in the No. 1--3 monitoring boreholes. Figgure 5
exhibits partial oness collected frrom field boreholes
informaation, which presents cracck degrees annd the
distribuution of fracturre network.
Staatistics analysiis of these craccks in each boorehole
were shhown in Fig. 6. Above all, annalysis of in-sittu real
deform
mation monitoriing can be achiieved:

251 Moonitoring borehhole was locateed at B4 coal (ssofter)
andd was influennced largely by
b hole 4 blaasting.
Cracks developedd apparently frrom 0 to 1.0 m,
m with
verrtical crack dennsely between 1.0 to 2.5 m, with a
maaximum of 4.5 m. Due to thee blasting effeccts, the
bore-hole with broken free surface, it haad the
cruushed loose cooal. Some craacks developedd at a
deppth of aroundd 2.5 m, which may causee roof
cavving.
Res. J. Appl.
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(a)
Fig. 6: S
Statistical regulaarity of cracks with
w advent of boorehole
d
depth
Table 2: Parameters of em
mulsion matrix
Density
Velocity
Loose zone
(g/cm3)
(m/s)
(m)
1.25
3600-4500
6
(b)
(c)
(d)
Fig. 5: All kinds
k
of boreho
ole cracks (a): Oblique
O
crack; (b):
(
Lateeral crack; (c): Circular crack; (d): Lateral craack
and vertical crack


Monittoring boreholee was located at
a B5 coal (Whiich
was haarder than B4) and hole 7 inttersects with No.
N
monitooring 2 boreh
hole at 7.0-9.00 m. Cracks are
a
mainlyy focused on ab
bout 1.0 m andd there was som
me
brokenn coal from 5..5 to 6.5 m. Overall,
O
the innner
wall of
o No. 2 mo
onitoring boreehole maintainned
integriity.
Monittoring boreholee was locatedd at B6 coal and
a
was afffected by hole 10 blasting. This monitoriing
borehoole had dual ciircular cracks from 0 to 1.0 m,
with broken
b
section scattered betw
ween 0 to 3.0 m.
Table 3: Parameters
P
of cracks in
i boreholes
Angeel
Width
Position Borehole
(°)
(mm)
B4
1#
90
1-3
B5
2#
90
2-4
B6
3#
90
3-5
Diameter
D
(mm)
2
25
Lenngth
(mm
m)
100-230
130-270
180-420
Efficciency
(%)
96-100
Fracture touughness
(MPa·m1/2)
0.5616-0.88806
0.6403-0.95541
0.7535-1.19900
Calcullation of frracture toug
ghness: Afterr preblastingg being comppleted, a largge amount off free
surfacees brought aboout laminationn crack phenom
menon
by blaasting stress waves in No.
N
1-3 moniitoring
boreholes, which cauused cracks too grow and deevelop
well. Details of em
mulsion matrixx can be seenn from
Table 2.
2
Parameters of cracks in No.
N
1-3 moniitoring
boreholes were listeed in Table 3.
3 Because avverage
length of
o crack was distinctly
d
more than radius off holes,
Eq. 3 was
w adopted forr computing thhe fracture tougghness
and thee results can bee seen from Tabble 3.
F 7,
Pre-bllasting stress monitoring: As shown in Fig.
the insttruments were applied for thhe stress monittoring.
On onee side, it cann be (a) From
m the workingg face,
abutmeent pressure is not obvious frrom 0 to 10.0 m. (b)
From 10.0
1
to 15.0 m, the pressure was
w 0.2-1.0 MP
Pa and
kept steeady. Because of pre-blastingg, abutment prressure
rose suuddenly to 2.55 MPa and thhen lessened a little
towarded a new balannce. On the othher side, roof load
l
at
v
rarely in the explosive process:
stope varied



252 Froom 0 to 8.0 m,
m the roof loadd went up graadually
andd its peak addeed up to 2.5 MP
Pa.
Based on resultss of the stresss monitoring, 4.0 m
rooof canopy is used
u
to controll caving and sliding
s
aboove
Froom 8.0 to 13.0 m, roof load increased
i
sharpply by
flucctuating stress and its peak added
a
up to 3.55 MPa
andd then also lessened a little towards
t
anotheer new
ballance.
Res. J. Appl.
A
Sci. Engg. Technol., 6(22): 249-253, 20013
Nationaal High Technnology Researcch and Develoopment
Program
m (2008AA0662104), all thhese are grattefully
acknow
wleged.
R
REFERENCES
S
(a)
(b)
Fig. 7: In-siitu monitoring conditions;
c
(a): Layout of geneeral
in-siitu monitoring; (b): Stress moonitoring principle
and field data acquissition
Based on
o results of the stress monitooring, 4.0 m rooof
canopy is used to conttrol caving annd sliding aboove
A
sup
pporting from 0 to 45.0 m is
supports. Advanced
identical too the field situaation.
CON
NCLUSION
After pre-blasting in steep and thick seam, as
explosive effects
e
cause dynamic
d
stress focusing on poore
wall alongg the central lin
ne, so initial crracks form aloong
the centrall line, then und
der the quasi- static
s
stress froom
explosive gases,
g
initial crracks grows further.
The avverage length of
o crack is muuch more than the
t
radius of holes,
h
the fractu
ure toughness of
o B3 and B6 cooal
seams wass 0.5616 and 1.1900 MPa·m
m1/2, respectiveely
which hass been used to
o optimize parrameters of preblasting.
In-situu real monitorring results shhow that miniing
pressure iss not fierce and
d its peak decrreases obvioussly,
abutment pressure mov
ves toward woorking face and
a
stress conccentration zonee lessens appareently.
ACKNOW
WLEDGMEN
NT
This study was financially
f
suppported by the
t
National Natural Scieence Foundattion of Chiina
(510741400, 11002021), the
t China Schholarship Counncil
(CSC) & the Hebei Prrovincial Officce of Educatiion
(2010813124), the Docto
oral Subject Foundation of the
t
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